EX-99.1

EX-99.1 2 d218871dex991.htm EX-99.1 EX-99.1

Exhibit 99.1

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Technical Report on the Kibali Gold

Mine, Democratic Republic of the Congo

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18 March 2022

Effective Date: 31 December 2021

Mr. Rodney B. Quick, MSc, Pr. Sci.Nat

Mr. Simon Bottoms, CGeol, MGeol, FGS, FAusIMM

Mr. Christopher Hobbs, CGeol, MSc, MCSM, FAusIMM

Mr. Graham E. Trusler, MSc, Pr Eng, MIChE, MSAIChE

Dr. Thamsanqa Mahlangu, Pr. Eng, PhD

Mr. Shaun Gillespie, Reg Eng Tech, FAusIMM

Mr. Ismail Traore, MSc, FAusIMM, M.B. Law, DES

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Kibali Gold Mine Technical Report

CAUTIONARY STATEMENT ON FORWARD-LOOKING INFORMATION

This report contains forward-looking statements. All statements, other than statements of historical fact regarding Kibali Goldmines SA, Barrick Gold Corporation, AngloGold Ashanti Limited, or the Kibali Gold Mine, are forward-looking statements. The words “believe”, “expect”, “anticipate”, “contemplate”, “target”, “plan”, “intend”, “project”, “continue”, “budget”, “estimate”, “potential”, “may”, “will”, “can”, “could” and similar expressions identify forward-looking statements. In particular, this report contains forward looking statements with respect to cash flow forecasts, projected capital, operating and exploration expenditure, targeted cost reductions, mine life and production rates, potential mineralization and metal or mineral recoveries and information pertaining to potential improvements to financial and operating performance and mine life at the Kibali Gold Mine. All forward-looking statements in this report are necessarily based on opinions and estimates made as of the date such statements are made and are subject to important risk factors and uncertainties, many of which cannot be controlled or predicted. Material assumptions regarding forward-looking statements are discussed in this report, where applicable. In addition to such assumptions, the forward-looking statements are inherently subject to significant business, economic and competitive uncertainties, and contingencies. Known and unknown factors could cause actual results to differ materially from those projected in the forward-looking statements. Such factors include, but are not limited to: fluctuations in the spot and forward price of commodities (including gold, diesel fuel, natural gas and electricity); the speculative nature of mineral exploration and development; changes in mineral production performance, exploitation and exploration successes; diminishing quantities or grades of reserves; increased costs, delays, suspensions, and technical challenges associated with the construction of capital projects; operating or technical difficulties in connection with mining or development activities, including disruptions in the maintenance or provision of required infrastructure and information technology systems; damage to Kibali Goldmines SA’s, Barrick Gold Corporation’s, AngloGold Ashanti Limited’s, or the Kibali Gold Mine’s reputation due to the actual or perceived occurrence of any number of events, including negative publicity with respect to the handling of environmental matters or dealings with community groups, whether true or not; risk of loss due to acts of war, terrorism, sabotage and civil disturbances; fluctuations in the currency markets; changes in interest rates; changes in national and local government legislation, taxation, controls or regulations and/or changes in the administration of laws, policies and practices including the rules applicable to the repatriation of Kibali Gold Mine’s cash held in the Democratic Republic of the Congo, expropriation or nationalization of property and political or economic developments in the Democratic Republic of the Congo; uncertainty whether the Kibali Gold Mine will meet Barrick Gold Corporation’s capital allocation objectives; the impact of inflation; failure to comply with environmental and health and safety laws and regulations; timing of receipt of, or failure to comply with, necessary permits and approvals; non-renewal of key licences by governmental authorities; litigation; contests over title to properties or over access to water, power and other required infrastructure; risks associated with artisanal and small-scale mining; increased costs and physical risks including extreme weather events and resource shortages, related to climate change; availability and increased costs associated with mining inputs and labour, and risks associated with diseases, epidemics and pandemics, including the effects and potential effects of the global Covid-19 pandemic. In addition, there are risks and hazards associated with the business of mineral exploration, development, and mining, including environmental hazards, industrial accidents, unusual or unexpected formations, pressures, cave-ins, flooding, and gold ore losses (and the risk of inadequate insurance, or inability to obtain insurance, to cover these risks).

Many of these uncertainties and contingencies can affect Kibali Goldmines SA’s actual results and could cause actual results to differ materially from those expressed or implied in any forward-looking statements made by, or on behalf of, Kibali Goldmines SA, Barrick Gold Corporation, or AngloGold Ashanti Limited. All of the forward-looking statements made in this report are qualified by these cautionary statements. Kibali Goldmines SA, Barrick Gold Corporation, AngloGold Ashanti Limited, and the Qualified Persons who authored this report undertake no obligation to update publicly or otherwise revise any forward-looking statements whether as a result of new information or future events or otherwise, except as may be required by law.

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Kibali Gold Mine Technical Report

Table of Contents


1	Executive Summary	1

	1.1 Location	1

	1.2 Ownership	2

	1.3 History	3

	1.4 Geology and Mineralisation	4

	1.5 Mineral Resource Estimate	8

	1.6 Mineral Reserve Estimate	10

	1.7 Mining Method	12

	1.8 Mineral Processing	14

	1.9 Project Infrastructure	20

	1.10 Market Studies	21

	1.11 Environmental, Permitting and Social Considerations	21

	1.12 Capital and Operating Costs	24

	1.13 Economic Analysis	26

	1.14 Interpretation and Conclusions	26

	1.15 Recommendations	32

2	Introduction	34

	2.1 Effective Date	34

	2.2 Qualified Persons	34

	2.3 Site Visit of Qualified Persons	36

	2.4 List of Abbreviations	37

3	Reliance on Other Experts	38

4	Property Description and Location	39

	4.1 Project Location	39

	4.2 Mineral Rights and Land Ownership	39

	4.3 Surface Rights	44

	4.4 Ownership, Royalties and Lease Obligations	44

5	Accessibility, Climate, Local Resources, Infrastructure and Physiography	45

	5.1 Accessibility	45

	5.2 Climate and Physiography	45

	5.3 Infrastructure	47

	5.4 Local Resources	50

6	History	52

	6.1 Ownership	52

	6.2 Previous Exploration	53

	6.3 Previous Resource and Reserve Estimates	53

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	6.4 Past Production	55

7	Geological Setting and Mineralisation	56

	7.1 Regional Geology	56

	7.2 Structural Geology	58

	7.3 Project Geology	60

	7.4 Mineralisation	61

	7.5 Project Deposits	64

8	Deposit Types	73

9	Exploration	74

	9.1 Exploration Concept	74

	9.2 Geology and Geochronology	75

	9.3 Geophysics and Remote Sensing	75

	9.4 Geochemical Sampling	80

	9.5 Proposed 2022 Greenfields Exploration	83

	9.6 Proposed 2022 Brownfields Exploration	83

	9.7 Discussion	85

10	Drilling	86

	10.1 Drilling Definitions	86

	10.2 Drill Planning and Site Preparation	90

	10.3 Downhole Surveying	90

	10.4 Collar Surveys	90

	10.5 Diamond Drilling	90

	10.6 Reverse Circulation Drilling	92

	10.7 Twin Drilling Studies	93

	10.8 Drill Spacing Optimisation	95

	10.9 Independent Audits	95

	10.10 Discussion	96

11	Sample Preparation, Analysis and Security	97

	11.1 Sample Preparation	97

	11.2 Sample Analysis	101

	11.3 Quality Assurance and Quality Control	102

	11.4 Sample Security	128

	11.5 Independent Audit	129

	11.6 Discussion	129

12	Data Verification	130

	12.1 Historical Drill Hole Data Verification	130

	12.2 Current Drill Hole Data Verification	130

	12.3 Independent Audit	131

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	12.4 Discussion	131

13	Mineral Processing and Metallurgical Testing	132

	13.1 Summary	132

	13.2 Test Work Strategy & Sample Selection	135

	13.3 Metallurgical Recoveries	153

	13.4 Deleterious Elements	155

	13.5 Discussion	156

14	Mineral Resource Estimates	157

	14.1 Summary	157

	14.2 Resource Database	159

	14.3 Geological Modelling	175

	14.4 Bulk Density	189

	14.5 Compositing	196

	14.6 Treatment of High-Grade Outliers (Top Capping)	212

	14.7 Variography	220

	14.8 Block Model Estimation	236

	14.9 Block Models	239

	14.10 Resource Classification	246

	14.11 Block Model Depletion	247

	14.12 Block Model Validation	249

	14.13 Resource Cut-Off Grades	253

	14.14 Mineral Resource Statement	267

	14.15 2021 Versus 2020 EOY Mineral Resource Comparison	269

	14.16 Discussion	278

15	Mineral Reserve Estimate	279

	15.1 Summary	279

	15.2 Mineral Reserve Estimation Process	280

	15.3 Economic Parameters	283

	15.4 Pit Optimisation	290

	15.5 Underground Stope Shapes	304

	15.6 Reconciliation	305

	15.7 Mineral Reserves Statement	308

	15.8 Discussion	312

16	Mining Methods	314

	16.1 Summary	314

	16.2 Open Pit Mining	314

	16.3 Underground Mining	345

	16.4 Underground Mining Operations	359

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	16.5 Underground Mining Services and Infrastructure	366

	16.6 Life of Mine Production Schedule	382

17	Recovery Methods	389

	17.1 Processing Plant	389

	17.2 Processing Recovery	400

	17.3 Production History	402

	17.4 Processing Costs	402

18	Project Infrastructure	405

	18.1 Mine Roads	405

	18.2 Supply Chain	405

	18.3 Surface Water Management	406

	18.4 Water Supply	408

	18.5 Tailings Facilities	408

	18.6 Power Supply	409

	18.7 Site Infrastructure	411

	18.8 Communication and Information Technology	413

	18.9 Security	413

19	Market Studies and Contracts	414

	19.1 Markets	414

	19.2 Contracts	414

20	Environmental Studies, Permitting, and Social or Community Impact	416

	20.1 Environmental and Social Management	416

	20.2 Environmental Considerations	420

	20.3 Environmental Studies, Permitting, and Social or Community Impact	425

21	Capital and Operating Costs	431

	21.1 Capital Costs	431

	21.2 Operating Costs	432

22	Economic Analysis	435

23	Adjacent Properties	436

24	Other Relevant Data and Information	437

25	Interpretation and Conclusions	438

	25.1 Geology and Mineral Resources	438

	25.2 Mining and Mineral Reserves	439

	25.3 Processing	439

	25.4 Infrastructure	440

	25.5 Environment and Social Aspects	440

	25.6 Risks	441

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26	Recommendations	444

	26.1 Geology and Mineral Resources	444

	26.2 Mining and Mineral Reserves	444

	26.3 Processing	445

	26.4 Infrastructure	445

	26.5 Environment and Social Aspects	445

27	References	446

28	Date and Signature Page	451

29	Certificate of Qualified Persons	453

	29.1 Rodney B. Quick	453

	29.2 Simon P. Bottoms	454

	29.3 Christopher B. Hobbs	455

	29.4 Graham E. Trusler	456

	29.5 Thamsanqa Mahlangu	457

	29.6 Shaun Gillespie	458

	29.7 Ismail Traore	459

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List of Tables


Table 1-1 Kibali Exploitation Permit Details		2

Table 1-2 Kibali Mineral Resources as of 31December 2021		9

Table 1-3 Kibali Mineral Reserves as of 31December 2021		11

Table 1-4 Summary of Average Recovery for All the Samples		17

Table 1-5 Actual Process and Plant Engineering Operating Costs for 2019, 2020 and 2021		19

Table 1-6 LOM Capital Expenditure Based on Mineral Reserves		25

Table 1-7 LOM Unit Operating Costs Based on Mineral Reserves		25

Table 1-8 Kibali Risk Analysis		31

Table 2-1 QP Responsibilities		35

Table 4-1 Kibali Exploitation Permit Details		39

Table 4-2 Kibali Exploitation Permit Coordinates		41

Table 4-3 Coordinates of the Exclusion Zone in the Kibali Exploitation Permit		44

Table 6-1 Summary of Historical Kibali Trenches, Auger and Pits Summary		53

Table 6-2 Moto Goldmines Ltd. Mineral Resource Estimate as of August 2008		54

Table 6-3 Moto Goldmines Ltd. Ore Reserve Estimate as of August 2008		54

Table 6-4 Past Production Records for the Kibali Mine		55

Table 9-1 Kibali Soil and Stream Sediment Sample Summary		81

Table 9-2 Kibali Trenches, Auger and Pits Summary		81

Table 10-1 Kibali Drilling Summary		87

Table 10-2 Summary of Drill Hole Twinning in 2021		94

Table 11-1 Submitted Samples		101

Table 11-2 List of CRMs Assayed at SGS Doko		105

Table 11-3 CRM Summary for Review Period at SGS Doko		105

Table 11-4 Statistics for Coarse Blank Samples		107

Table 11-5 Statistics for Pulp Blanks (OREAS22f)		108

Table 11-6 Statistics for RC Duplicates at SGS Doko		110

Table 11-7 Statistics of Half-Core Duplicates at SGS Doko		113

Table 11-8 Statistics of Coarse Reject Duplicates at SGS Doko		116

Table 11-9 Statistics of Pulp Reject Duplicates at SGS Doko		119

Table 11-10 Statistics of Pulp Re-Submissions at SGS Doko		122

Table 11-11 Summary of Pulp Duplicates Analysed at ALS		125

Table 13-1 Summary of Test Work		133

Table 13-2 Physical and Extraction Sample Selection and Test Work Logic		137

Table 13-3 Extraction Comparison – Underground Variability		139

Table 13-4 Isolated Samples for Further Analysis		140

Table 13-5 Direct Cyanidation Results		144

Table 13-6 Metallurgical Recoveries per Deposit		151

Table 13-7 KCD Fresh Open Pit Fresh Samples – Lode 5000		152

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Table 13-8 Summary of Average Recovery for All the Samples			154

Table 14-1 Kibali Mineral Resources as of 31December 2021			157

Table 14-2 Summary of Deposits and Model Date			159

Table 14-3 KCD Drill Summary Used in the 2021 Mineral Resource Estimate			160

Table 14-4 KCD 3000 Lodes Composite Data – 2021 Mineral Resource Estimate			161

Table 14-5 KCD 5000 Lodes Composite Data – 2021 Mineral Resource Estimate			162

Table 14-6 KCD 9000 Lodes Composite Data – 2021 Mineral Resource Estimate			162

Table 14-7 KCD 11000 Lodes Composite Data – 2021 Mineral Resource Estimate			163

Table 14-8 Sessenge Drill Summary of Holes Used in the 2021 Mineral Resource Estimate			163

Table 14-9 Sessenge Composite Data – 2021 Mineral Resource Estimate			164

Table 14-10 Drill Summary of Gorumbwa Holes Used in the 2021 Mineral Resource Estimate			164

Table 14-11 Gorumbwa Composite Data – 2021 Mineral Resource Estimate			165

Table 14-12 Drill Summary of Pakaka Holes Used in the 2021 Mineral Resource Estimate			166

Table 14-13 Pakaka Composite Data – 2021 Mineral Resource Estimate			167

Table 14-14 Drill Summary of Kombokolo Holes Used in the 2021 Mineral Resource Estimate			167

Table 14-15 Kombokolo Composite Data – 2021 Mineral Resource Estimate			168

Table 14-16 Drill Summary of Pamao Holes Used in the 2021 Mineral Resource Estimate			168

Table 14-17 Pamao Composite Data – 2021 Mineral Resource Estimate			169

Table 14-18 Pamao South Composite Data – 2021 Mineral Resource Estimate			169

Table 14-19 Drill Summary of Mengu Hill Holes Used in the 2021 Mineral Resource Estimate			170

Table 14-20 Mengu Hill Composite Data – 2021 Mineral Resource Estimate			170

Table 14-21 Drill Summary of Mengu Village Holes Used in the 2021 Mineral Resource Estimate			171

Table 14-22 Mengu Village Composite Data – 2021 Mineral Resource Estimate			171

Table 14-23 Drill Summary of Megi-Marakeke-Sayi Holes Used in the 2021 Mineral Resource Estimate			171

Table 14-24 Megi-Marakeke-Sayi Composite Data – 2021 Mineral Resource Estimate			172

Table 14-25 Drill Summary of Kalimva-Ikamva Holes Used in the 2021 Mineral Resource Estimate			172

Table 14-26 Kalimva Composite Data – 2021 Mineral Resource Estimate			173

Table 14-27 Ikamva Composite Data – 2021 Mineral Resource Estimate			173

Table 14-28 Drill Summary of Aerodrome Holes Used in the 2021 Mineral Resource Estimate			173

Table 14-29 Aerodrome Composite Data – 2021 Mineral Resource Estimate			174

Table 14-30 Drill Summary of Oere Holes Used in the 2021 Mineral Resource Estimate			174

Table 14-31 Oere Composite Data – 2021 Mineral Resource Estimate			174

Table 14-32 Mineralisation Domain Dimensions for all Mineral Resource Deposits			176

Table 14-33 KCD 3000 Lodes Assigned Density Summary			190

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Table 14-34 KCD 5000 Lodes Assigned Density Summary			192

Table 14-35 KCD 9000 Lodes Assigned Density Summary			194

Table 14-36 KCD Waste (99999 & 800) Assigned Density Summary			195

Table 14-37 KCD 11000 Lodes Assigned Density Summary			196

Table 14-38 KCD 3000 Lodes Top Capping Analysis			213

Table 14-39 KCD 5000 Lodes Top Capping Analysis			214

Table 14-40 KCD 9000 Lodes Top Capping Analysis			214

Table 14-41 KCD 11000 Lodes Top Capping Analysis			214

Table 14-42 Sessenge Top Capping Analysis			215

Table 14-43 Gorumbwa Top Capping Analysis			215

Table 14-44 Pakaka Top Capping Analysis			216

Table 14-45 Kombokolo Top Capping Analysis			216

Table 14-46 Pamao Top Capping Analysis			217

Table 14-47 Pamao South Top Capping Analysis			217

Table 14-48 Mengu Village Top Capping Analysis			217

Table 14-49 Megi-Marakeke-Sayi Village Top Capping Analysis			218

Table 14-50 Kalimva Village Top Capping Analysis			218

Table 14-51 Ikamva Village Top Capping Analysis			218

Table 14-52 Mengu Hill Top Capping Analysis			219

Table 14-53 Aerodrome Top Capping Analysis			219

Table 14-54 Oere Top Capping Analysis			219

Table 14-55 QKNA Parameters for KCD 5101 Domain			238

Table 14-56 KCD Global Block Model Extent (With Rotation)			240

Table 14-57 Sessenge Global Block Model Extent (With Rotation)			241

Table 14-58 Gorumbwa Global Block Model Extent (With Rotation)			241

Table 14-59 Pakaka Global Block Model Extent			242

Table 14-60 Pakaka Search Ellipsoid Orientation			242

Table 14-61 Kombokolo Global Block Model Extent (No Rotation)			242

Table 14-62 Kombokolo Search Ellipsoid Orientation			243

Table 14-63 Pamao and Pamao South Global Block Model Extent (With Rotation)			243

Table 14-64 Mengu Village Global Block Model Extent (With Rotation)			244

Table 14-65 Megi-Marakeke-Sayi Global Block Model Extent (With Rotation)			244

Table 14-66 Kalimva-Ikamva Block Model Extent (No Rotation)			244

Table 14-67 Mengu Hill Global Block Model Extent (No Rotation)			245

Table 14-68 Mengu Hill Search Ellipsoid Orientation			245

Table 14-69 Aerodrome Global Block Model Extent			245

Table 14-70 Oere Global Block Model Extent (With Rotation)			246

Table 14-71 Kibali Mineral Resource Classification Parameters			247

Table 14-72 2021 Block Model Volume Comparison			250

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Table 14-73 KCD 2021 Optimisation Parameters			253

Table 14-74 KCD Underground 2021 Optimisation Parameters			254

Table 14-75 Sessenge 2021 Optimisation Parameters			258

Table 14-76 Gorumbwa 2021 Optimisation Parameters			258

Table 14-77 Pakaka 2021 Optimisation Parameters			259

Table 14-78 Pakaka Geometallurgical Domained Recoveries			261

Table 14-79 Kombokolo 2021 Optimisation Parameters			262

Table 14-80 Pamao 2021 Optimisation Parameters			262

Table 14-81 Pamao South 2021 Optimisation Parameters			263

Table 14-82 Mengu Village 2021 Optimisation Parameters			264

Table 14-83 Megi-Marakeke-Sayi 2021 Optimisation Parameters			264

Table 14-84 Kalimva-Ikamva 2021 Optimisation Parameters			265

Table 14-85 Mengu Hill 2021 Optimisation Parameters			265

Table 14-86 Aerodrome 2021 Optimisation Parameters			266

Table 14-87 Oere 2021 Optimisation Parameters			266

Table 14-88 Kibali Mineral Resources as of 31December 2021			268

Table 14-89 KCD Open Pit 2021 vs 2020 Comparison Above the Underground Box			271

Table 14-90 KCD Underground 2021 vs 2020 Comparison within Bounding Box and Within MSO			271

Table 14-91 Sessenge 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell			272

Table 14-92 Sessenge SW 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell			272

Table 14-93 Gorumbwa 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell			273

Table 14-94 Pakaka 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell			273

Table 14-95 Pamao – Pamao South 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell			274

Table 14-96 Mengu Village 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell			275

Table 14-97 Megi-Marakeke-Sayi 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell			276

Table 14-98 Kalimva-Ikamva 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell			276

Table 14-99 Mengu Hill 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell			276

Table 14-100 Aerodrome 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell			277

Table 14-101 Oere 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell			277

Table 15-1 Kibali Mineral Reserves as of 31 December 2021			280

Table 15-2 KCD, Megi-Marakeke-Sayi, Pakaka Open Pits – Marginal and Full Grade Ore (FGO) Cut-Off Grade for Different Material Types			285

Table 15-3 Pamao, Kalimva-Ikamva, Gorumbwa Open Pits – Marginal and Full Grade Ore Cut-Off Grade for Different Material Types			286

Table 15-4 Sessenge, Aerodrome, and Oere Open Pits - Marginal and Full Grade Ore Cut-Off Grade for Different Material Types			287

Table 15-5 Kibali Underground Mine – Cut-Off Grade Calculation			289

Table 15-6 Comparison of Whittle Results for the Gorumbwa Pit at $1,000/oz Au and $1,200/oz Au			291

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Table 15-7 Comparison of Whittle Results for the Pamao Pit with 2020 Results at $1,200/oz Au			292

Table 15-8 Sessenge Gold Price Sensitivities			293

Table 15-9 Pamao Gold Price Sensitivities			294

Table 15-10 Kalimva-Ikamva Gold Price Sensitivities			295

Table 15-11 Pakaka Gold Price Sensitivities			296

Table 15-12 Megi-Marakeke-Sayi Gold Price Sensitivities			297

Table 15-13 Aerodrome Gold Price Sensitivities			298

Table 15-14 Oere Gold Price Sensitivities			299

Table 15-15 MSO Parameters			304

Table 15-16 Kibali 2021 EOY MCF Reconciliation			306

Table 15-17 Kibali EOY 2021 Resource Call Factor Reconciliation			308

Table 15-18 Kibali Open Pit Mineral Reserves as of 31December 2021			309

Table 15-19 Open Pit Mineral Reserve Comparison to Previous Estimate			310

Table 15-20 Kibali Underground Mineral Reserves as of 31 December 2021			310

Table 15-21 Kibali Mineral Reserves by Mining Zone as of 31 December 2021			311

Table 15-22 Underground Mineral Reserve Comparison to Previous Estimate			311

Table 15-23 Kibali Surface Stockpile Mineral Reserve as of 31 December 2021			312

Table 16-1 Kibali Open Pits Historical Production			315

Table 16-2 Kibali Open Pits, Reserves Basis			316

Table 16-3 KCD Geotechnical Geometry			318

Table 16-4 Rock Mass Properties for the Sessenge Pit			319

Table 16-5 Sessenge Slope Design			320

Table 16-6 Gorumbwa Recommended Pit Slope Configuration			321

Table 16-7 Pakaka Recommended Pit Slope Configuration			323

Table 16-8 Pamao Recommended Pit Slope Configuration			325

Table 16-9 Pamao South Recommended Pit Slope Configuration			325

Table 16-10 Oere Slope Design			327

Table 16-11 Gorumbwa South Pit (5,760 mRL) 2021 Dewatering Operations Summary			329

Table 16-12 Gorumbwa North Pit (5,745 mRL) 2021 Dewatering Operations Summary			330

Table 16-13 KCD Pushback 3 Pit 2021 Dewatering Operations Summary			333

Table 16-14 Summary of Pit Design Parameters			335

Table 16-15 Current Primary Open Pit Mine Equipment Fleet			344

Table 16-16 Waste Dump Capacities			344

Table 16-17 Summary of Historical Paste Dilution per Stoping Type			347

Table 16-18 Summary of 70/30 and 90/10 Historical Paste Test Work UCS Strength			348

Table 16-19 Summary of Historical Stope Performance and Dilution Parameters			350

Table 16-20 Summary of Mineral Reserve Estimate Recovery Parameters			351

Table 16-21 Stope Recovery History			351

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Table 16-22 Stope Recovery and Unplanned Dilution History			352

Table 16-23 Summary of Proportion of the Mineral Reserve per Stoping Type			358

Table 16-24 Historical Proportion of the Stopes Mined per Lode and per Stope Type			359

Table 16-25 Summary of Yearly Historical Proportion of the Stopes Mined per Lode			359

Table 16-26 KCD UG Support Categories and Classifications for short life openings (<5 years)			363

Table 16-27 Cable Bolt Requirements for Intersections in General / Short-Term Excavations			364

Table 16-28 KCD UG Support Categories and Classifications for long life openings (>5 years)			364

Table 16-29 Cable Bolt Requirements for Intersections in Long-Term Excavations (Pump Stations, Crushers, etc)			365

Table 16-30 Kibali Underground Mining Equipment			365

Table 16-31 Summary of Paste UCS			368

Table 16-32 Ventilation Requirement per Equipment			378

Table 16-33 Open Pit Mining Sequence Over the Mineral Reserves LOM			385

Table 16-34 Kibali KCD Underground LOM Physicals Based on Mineral Reserves			387

Table 17-1 Plant Availability and Utilisation			391

Table 17-2 Kibali Processing Plant Overall Gold Recovery in 2021 by Month			401

Table 17-3 Kibali Processing Plant Production History			402

Table 17-4 Actual Process and Plant Engineering Operating Costs for 2019, 2020 and 2021			403

Table 21-1 LOM Capital Expenditure Based on Mineral Reserves			432

Table 21-2 LOM Operating Unit Costs Based on Mineral Reserves			433

Table 21-3 LOM Operating Total Costs Based on Mineral Reserves			433

Table 25-1 Kibali Risk Analysis			443

List of Figures

Figure 1-1 Simplified Flowsheet of the Kibali Processing Plant Depicting Two Discrete Streams			15

Figure 1-2 Initial Hole Composite Dissolutions			15

Figure 1-3 Kibali Throughput Performance versus 7.2 Mtpa Design Throughput			18

Figure 1-4 Kibali Processing Plant Process Recovery for Year 2021			18

Figure 4-1 Kibali Mine Location			40

Figure 4-2 Kibali Tenement and Permits			43

Figure 5-1 Kibali Average Monthly Rainfall Statistics			46

Figure 5-2 Kibali Mine infrastructure			49

Figure 5-3 Kibali Deposits and Surrounding Communities			51

Figure 7-1 Regional Geology			57

Figure 7-2 Summary Geologic Map of the Moto (Kibali Greenstone) Belt, Showing Major Geologic Domains, Crosscutting Granitoid Plutons, and the General Structural Architecture			59

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Figure 7-3 Photograph Showing Examples of Altered and Mineralised Rocks from the KCD Deposit			63

Figure 7-4 Summary Geologic Map of the Central Part of the KZ Trend Including the Area Around the KCD Deposit			65

Figure 7-5 Cross Section Through the Central Part of KCD, Extending as Far West as Gorumbwa			66

Figure 7-6 KCD and Sessenge 2021 Block Models with Underground Mine Design			67

Figure 9-1 Kibali Project Area with Airborne Magnetic Response			77

Figure 9-2 Kibali Project Area with Airborne EM Response			78

Figure 9-3 Kibali Project Area with Topographic Surveys			79

Figure 9-4- Kibali Project Area with Stream Sediment Sampling Gold Results for Each Catchment			82

Figure 9-5 Geological Map of the KZ Trend, Showing Targets and Known Deposits			84

Figure 10-1 KDC Deposit Drill Plan			88

Figure 10-2 Representative Cross Section (XS065) through the KCD Deposit			89

Figure 11-1 Diamond Drill Core Sample Flowchart			98

Figure 11-2 Reverse Circulation Sample Flowchart			99

Figure 11-3 Channel Sample Flowchart			100

Figure 11-4 Kibali QA/QC Protocol Flowchart			103

Figure 11-5 Tram Line Graph for CRMs Analysed at SGS Doko			106

Figure 11-6 Performance Graph of Coarse Blanks			107

Figure 11-7 Performance Graph of Pulp Blanks			108

Figure 11-8 Performance Graph for Spiked Blanks			109

Figure 11-9 Normal QQ Plot of RC Field Duplicates ≤ 100 g/t Au Tail Cut			111

Figure 11-10 Normal Scatter Plot of RC Field Duplicates ≥ 0.5 g/t Au and ≤ 100 g/t Au Tail Cuts			111

Figure 11-11 Ranked HARD Plot of RC Field Duplicates ≤ 100 g/t Au Tail Cut			112

Figure 11-12 Precision Plot of RC Field Duplicates vs Original Sample at SGS Doko			113

Figure 11-13 Normal QQ Plot of HC Duplicates ≤ 100 g/t Au Tail Cut			114

Figure 11-14 Normal Scatter Plot of HC Duplicates ≥ 0.5 g/t Au and ≤ 100 g/t Au Tail Cuts			114

Figure 11-15 Ranked HARD Plot of HC Duplicates ≤ 100 g/t Au Tail Cut			115

Figure 11-16 Precision Plot of HC Field Duplicates vs Original Sample at SGS Doko			116

Figure 11-17 Normal QQ Plot of Coarse Reject Duplicates ≤ 100 g/t Au Tail Cut			117

Figure 11-18 Normal Scatter Plot of Coarse Reject Duplicates ≥ 0.5 g/t Au and ≤ 100 g/t Au Tail Cuts			117

Figure 11-19 Ranked HARD Plot of Coarse Reject Duplicates ≤ 100 g/t Au Tail Cut			118

Figure 11-20 Precision Plot of Coarse Reject Duplicates vs Original Sample at SGS Doko			119

Figure 11-21 Normal QQ Plot of Pulp Reject Duplicates ≤ 100 g/t Au Tail Cut			120

Figure 11-22 Normal Scatter Plot of Pulp Reject Duplicates ≥ 0.5 g/t Au and ≤ 100 g/t Au Tail Cuts			120

Figure 11-23 Ranked Plot of Pulp Reject Duplicates ≤ 100 g/t Au Tail Cut			121

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Figure 11-24 Precision Plot of Pulp Reject Duplicates vs Original Sample at SGS Doko			122

Figure 11-25 Normal QQ Plot of Pulp Re-Submissions ≤ 100 g/t Au Tail Cut			123

Figure 11-26 Normal Scatter Plot of Pulp Re-Submissions ≥ 0.5 g/t Au and ≤ 100 g/t Au Tail Cuts			123

Figure 11-27 Ranked HARD Plot of Pulp Re-Submissions ≤ 100 g/t Au Tail Cut			124

Figure 11-28 Precision Plot of Pulp Re-Submissions vs Original Sample at SGS Doko			125

Figure 11-29 Normal QQ Plot of Umpires SGS Doko vs ALS at 100 g/t Au Tails Cut			126

Figure 11-30 Relative Difference Plot of Umpires SGS Doko vs ALS at 100 g/t Au Tails Cut			127

Figure 11-31 Precision Plot of Umpires by SGS Doko vs ALS			128

Figure 13-1 Initial Hole Composite Dissolutions			136

Figure 13-2 Primary Extraction Excluding the Leaching of Flotation Tails			138

Figure 13-3 Primary Extraction Variability Including the Leaching of the Flotation Tails			140

Figure 13-4 Extraction as a Function of Diamond Drill Holes			141

Figure 13-5 Analysis of the Drill Hole Samples Exhibiting Large Variances			141

Figure 13-6 Gold Extractions Obtained for Various Extraction Variability Tests and Master Composite Samples			142

Figure 13-7 Plots of Extraction Using the Primary Process for the Oxide Materials – KCD			143

Figure 13-8 Direct Cyanidation of Grade Control Samples			143

Figure 13-9 DD Composite Samples BBWi (KWh/t)			145

Figure 13-10 Kibali Processing Plant Average P₈₀and Specific Energy Consumption (2021)			145

Figure 13-11 Met_OT Composite Samples: Leach Recovery vs P₈₀			146

Figure 13-12 Kibali Processing Plant Composite Samples: Flotation Recovery by Particle Size Range			146

Figure 13-13 Kibali Ore – General Mineralogy			147

Figure 13-14 Kibali Metallurgical Composite Samples – Oxygen and Recovery Profile			149

Figure 13-15 2020 Aerodrome BRT and Arsenic Distribution			150

Figure 13-16 Sampling Strategy and Classification of Samples at KCD			152

Figure 14-1 Boundary Analysis between KCD High-Grade (5101) and Low-Grade (5005) Domains			176

Figure 14-2 3D View of KCD Sessenge Mineralisation			178

Figure 14-3 3D View of Gorumbwa Mineralisation			180

Figure 14-4 3D View of Pakaka Mineralisation			181

Figure 14-5 3D View of the Kombokolo Mineralisation			182

Figure 14-6 3D View of the Pamao and Pamao South Mineralisation			183

Figure 14-7 3D View of the Mengu Village Mineralisation			184

Figure 14-8 3D View of the Megi-Marakeke-Sayi Mineralisation			185

Figure 14-9 3D View of the Kalimva Ikamva Mineralisation			186

Figure 14-10 3D View of the Mengu Hill Mineralisation			187

Figure 14-11 3D View of the Aerodrome Mineralisation			188

Figure 14-12 3D View of Oere Mineralisation			189

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Figure 14-13 KCD 3000 Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			197

Figure 14-14 KCD 5000 Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			198

Figure 14-15 KCD 9000 Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			199

Figure 14-16 KCD 11000 Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			200

Figure 14-17 Sessenge Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			201

Figure 14-18 Gorumbwa Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			202

Figure 14-19 Pakaka Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			203

Figure 14-20 Kombokolo Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			204

Figure 14-21 Pamao Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			205

Figure 14-22 Pamao South Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			206

Figure 14-23 Mengu Village Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			207

Figure 14-24 Megi-Marakeke-Sayi Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			208

Figure 14-25 Kalimva-Ikamva Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			209

Figure 14-26 Mengu Hill Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			210

Figure 14-27 Aerodrome Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			211

Figure 14-28 Oere Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length Distribution of 2 m Uncapped Composites			212

Figure 14-29 KCD Lode 3101 Normal Score Variogram Models and Nested Back Transformed Variogram Model			221

Figure 14-30 KCD Lode 5102 and 5202 Normal Score Variogram Models and Nested Back Transformed Variogram Model			222

Figure 14-31 KCD Lode 9105 Normal Score Variogram Models and Nested Back Transformed Variogram Model			223

Figure 14-32 KCD Lode 11000 Normal Score Variogram Models and Nested Back Transformed Variogram Model			224

Figure 14-33 Sessenge Lode 1004 Normal Score Variogram Models and Nested Back Transformed Variogram Model			225

Figure 14-34 Gorumbwa Lode 1004 Normal Score Variogram Models and Nested Back Transformed Variogram Model			226

Figure 14-35 Pakaka Lode 1001 Normal Score Variogram Models and Nested Back Transformed Variogram Model			227

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Figure 14-36 Kombokolo Lode 1101 and 1102 Normal Score Variogram Models and Nested Back Transformed Variogram Model			228

Figure 14-37 Pamao Lode 2001 Normal Score Variogram Models and Nested Back Transformed Variogram Model			229

Figure 14-38 Pamao South Lode 1001 and 1002 Normal Score Variogram Models and Nested Back Transformed Variogram Model			230

Figure 14-39 Megi-Marakeke-Sayi Lode 1001 and 1101 Normal Score Variogram Models and Nested Back Transformed Variogram Model			231

Figure 14-40 Kalimva-Ikamva Lode 1001 Normal Score Variogram Models and Nested Back Transformed Variogram Model			232

Figure 14-41 Mengu Hill Lode 1001 Normal Score Variogram Models and Nested Back Transformed Variogram Model			233

Figure 14-42 Aerodrome Lode 2001 Normal Score Variogram Models and Nested Back Transformed Variogram Model			234

Figure 14-43 Oere Lode 1001 and 1101 Normal Score Variogram Models and Nested Back Transformed Variogram Model			235

Figure 14-44 QKNA for KCD Domain 5101 and 5201 Underground GC Zone			236

Figure 14-45 3D View of KCD Domain 5002 and Guiding Dynamic Anisotropy Surface			240

Figure 14-46 3D OP and Development Stope Voids at Gorumbwa			248

Figure 14-47 KCD SWATH Plot of Domains 3106 Along Strike (45°)			250

Figure 14-48 KCD SWATH Plot of Domains 3106 Across Strike (135°)			251

Figure 14-49 KCD SWATH Plot of Domains 3106 Along Z Axis (RL)			251

Figure 14-50 An Example of the KCD Visual Checks on Section for Lode 5000 (Domain 5102)			252

Figure 14-51 COS Plot for Lode 3000 (Domain 3106)			252

Figure 14-52 Decluster Plot for Lode 3000 (Domain 3106)			253

Figure 14-53 3D Oblique Image of KCD Underground Development with the Annual Resource Exclusion Solids			255

Figure 14-54 KCD 3D View of MSO Shapes and the Exclusion Solids			256

Figure 14-55 KCD 3D View of MSO Shapes Against Grade Blocks			257

Figure 14-56 Plan View Map of the Pakaka Geometallurgical Domains and Their Spatial Correlation with the Mineralisation Resource Domains			260

Figure 14-57 2021 Kibali Open Pit Mineral Resource Reconciliation			269

Figure 14-58 2021 Kibali Underground Mineral Resource Reconciliation			270

Figure 14-59 2021 Kibali Total Mineral Resource Reconciliation			270

Figure 15-1 KCD Underground Mining Zones			281

Figure 15-2 Kibali Underground Mineral Reserves – Stopes Above the BCOG			290

Figure 15-3 Aerodrome Pit Size versus Cash Flow Curve at Different Gold Prices			301

Figure 15-4 Sessenge Pit Size versus Cash Flow Curve at Different Gold Prices			302

Figure 15-5 Oere Pit Size versus Cash Flow Curve at Different Gold Prices			303

Figure 15-6 Kibali Underground Mineral Reserve Classification (Looking NW)			305

Figure 15-7 2021 Kibali Production with Weekly Feed Source Ratio versus Pulp Call versus Gold After Smelting			306

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Figure 15-8 2021 Weekly Grades Comparison (Mine Call Grade vs Plant Check Out Grade vs Carbon Loading)			307

Figure 15-9 2021 Weekly Tonnage Comparison (Mine Call Tonnes vs Plant Check Out Tonnes)			307

Figure 16-1 Long Section of the Pakaka Pushbacks			316

Figure 16-2 Long Section of the Gorumbwa Pushbacks			317

Figure 16-3 Gorumbwa Pit (Looking East) Showing Historical Underground Void Mined Out in Pushback 1			317

Figure 16-4 KCD Pushback 3 Geotechnical Domains			318

Figure 16-5 Gorumbwa Pit Geotechnical Domains			321

Figure 16-6 Pakaka Pit Geotechnical Domains			322

Figure 16-7 Cross Section Showing Pakaka PB2 Pit Project MRMM Reviewed from the Old Pit at $1,000/oz toward the $1,200/oz Au Pit Shells			323

Figure 16-8 Pamao Complex Geotechnical Domains			324

Figure 16-9 Oere Pit Geotechnical Domains			326

Figure 16-10 Gorumbwa South Pit Dewatering Design Criteria			331

Figure 16-11 Gorumbwa North Pit Dewatering Design Criteria			331

Figure 16-12 KCD Pushback 3 Pit Dewatering Design Criteria			332

Figure 16-13 Pamao Pit Dewatering Design Criteria			334

Figure 16-14 Pakaka Dam Fresh Water Reservoir Design Criteria			334

Figure 16-15 Gorumbwa Sessenge Upper Pit Designs			336

Figure 16-16 KCD Pit Design			337

Figure 16-17 Kalimva Pit Design			338

Figure 16-18 Ikamva Pit Design			339

Figure 16-19 Oere Pit Design			340

Figure 16-20 Pakaka Pit Design			341

Figure 16-21 Aerodrome Pit Design			342

Figure 16-22 Pamao and Pamao South Pit Design			343

Figure 16-23 Kibali Underground Infrastructure, LOM Development, and As Built EOY 2021			346

Figure 16-24 Stope Mining Sequence Using WebGen Wireless Blasting			352

Figure 16-25 Illustration of the Longitudinal Ring Design Approach in Secondary Stopes			353

Figure 16-26 Kibali Underground Mineral Reserve by Mining Method (Looking NW)			354

Figure 16-27 Database of Cablebolt-Supported Stopes			355

Figure 16-28 Database of Unsupported Stopes			355

Figure 16-29 Transverse Stope Sequencing			356

Figure 16-30 Transverse Advancing Face Sequencing			357

Figure 16-31 Longitudinal Mining Sequencing			358

Figure 16-32 KCD Underground Tonnes (Ore and Waste) Production History			360

Figure 16-33 LOM Paste Reticulation			367

Figure 16-34 Conceptual Understanding of Groundwater Flow Pathways in Northern KCD			372

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Figure 16-35 Kibali Underground Water Flows 2016 to 2021			374

Figure 16-36 Kibali Underground Pumping System Infrastructure (Looking NW)			376

Figure 16-37 Kibali Underground Water Services Infrastructure Flow Sheet			377

Figure 16-38 3D Oblique Image of Kibali Underground LOM Ventilation Network (Looking NW)			379

Figure 16-39 Kibali Underground Infrastructure LOM Electrical Reticulation (Looking NW)			381

Figure 16-40 Historic Rainfall Pattern and Lost Production Hours from Rain			382

Figure 16-41 Historical and Planned Kibali Gold Production			383

Figure 16-42 Kibali Open Pit Mining Rate			384

Figure 16-43 Kibali Open Pit and Underground LOM Feed Schedule Based on Mineral Reserves			388

Figure 17-1 Simplified Flowsheet of the Kibali Processing Plant Depicting Two Discrete Streams			390

Figure 17-2 Kibali Processing Plant Specific Energy Consumption 2015 to 2021			391

Figure 17-3 Kibali Processing Plant Water Demand 2013 to 2021			392

Figure 17-4 Kibali Processing Plant Specific Water Consumption 2013 to 2021			392

Figure 17-5 Kibali Processing Plant Performance Tonnes Treated 2013 to 2021			393

Figure 17-6 Kibali Processing Plant Overall Gold Recovery in 2021			400

Figure 17-7 Kibali Processing Plant Recovery			401

Figure 17-8 Kibali Processing Plant Pumpcell Residue and Throughput			401

Figure 18-1 Kibali Water Management Plan			407

Figure 18-2 Kibali Electrical Supply Mix			410

Figure 20-1 Time Series Satellite Imagery of the Kibali Region Indicating Land Transformation from 2003 to 2018			430

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1	Executive Summary

This Technical Report on the Kibali Gold Mine (Kibali, the Mine, or the Project), located in the Democratic Republic of the Congo (DRC) has been prepared by Barrick Gold Corporation (Barrick). The purpose of this Technical Report is to support public disclosure of Mineral Resource and Mineral Reserve estimates at the Mine as of 31 December 2021.

Kibali Goldmines SA (Kibali Goldmines) is an exploration and mining company, which is currently owned 45% by Barrick and 45% by AngloGold Ashanti Limited (AngloGold). The remaining 10% interest in Kibali Goldmines is held by Congolese parastatal Société Minière de Kilo-Moto SA (SOKIMO) with the shareholding held by the Minister of Portfolio (MoP) of the DRC.

The Project consists of multiple mineral deposits including; an underground mine at Karagba-Chauffeur-Durba (KCD); active open pits at KCD, Sessenge, Aerodrome and Gorumbwa; a partially depleted open pit with planned push backs at Pakaka; depleted open pits with further potential at Kombokolo and Mengu Hill; planned new open pits at Pamao, Megi-Marakeke-Sayi, Kalimva, Ikamva, and Oere, plus deposits under evaluation at Mengu Village and Sessenge SW; a processing plant (7.2 million tonnes per annum (Mtpa) design capacity), three hydropower stations, together with other associated mine operations and regional exploration infrastructure. The Kibali plant produces gold doré bars.

Total mine production from both Kibali underground and open pits in 2021 was 7.8 million tonnes (Mt) at a head grade of 3.62 g/t Au for a total of 812 (koz) Au (89.8% recovery).

The effective date of this report is 31 December 2021.

Unless otherwise stated, all data in this report is reported on a 100% basis.

1.1

Location

Kibali is located in the NE of the DRC in the Haut Uélé Province, approximately 1,800 km NE of the capital city of Kinshasa, approximately 560 km NE of the capital of the Orientale Province, Kisangani, 1,800 km from the Kenyan port of Mombasa, 1,950 km from the Tanzanian port of Dar es Salaam, and 150 km W of the Ugandan border town of Arua, near the international borders with Uganda and Sudan.

Personnel access to the Project is commonly through charter flight directly to site from Entebbe, Uganda which is served daily by commercial flights from European cities.

Road access is available from Kampala, Uganda and is approximately 650 km, which provides the primary route for operational supply chain.

The Project, which covers an area of approximately 1,836 km², is centred at approximately 3.13° latitude and 29.58° longitude, in the Haut Uélé Province.

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1.2

Ownership

Kibali Goldmines is owned 90% by a joint venture between Barrick (45%) and AngloGold (45%), and 10% by SOKIMO. SOKIMO is wholly owned by the DRC with the shareholding held by the MoP. The DRC Governmental Entity Offices des Mines d’Or de Kilo-Moto (OKIMO) was transformed into SOKIMO in December 2010.

Barrick is the operator at Kibali for both exploration and mining.

Kibali Goldmines has been granted ten Exploitation (Mining) Permits under the DRC Mining Code (2002) in respect of the Project, eight of which are valid until 2029 and two of which are valid until 2030.

The principal mineral deposit, KCD, all other open pit mining operations, and the associated infrastructure (processing plant, accommodation, and airport) are within Exploitation Permits 11447 and 11467. All Kibali Exploitation Permits are presented in Table 1-1.

Table 1-1 Kibali Exploitation Permit Details


Arête No.	Permit No.	Surface Area (km²)	Expiry Year
0852/CAB.MIN/MINES/01/2009	11447	226.8	2029
0855/CAB.MIN/MINES/01/2009	11467	248.9	2029
0854/CAB.MIN/MINES/01/2009	11468	45.9	2030
0853/CAB.MIN/MINES/01/2009	11469	91.8	2029
0329/CAB.MIN/MINES/01/2009	11470	30.6	2029
0852/CAB.MIN/MINES/01/2009	11471	113.0	2029
0331/CAB.MIN/MINES/01/2009	11472	85.0	2029
0856/CAB.MIN/MINES/01/2009	5052	302.4	2029
0858/CAB.MIN/MINES/01/2009	5073	399.3	2029
0103/CAB.MIN/MINES/01/2011	5088	292.2	2030

In the Qualified Person’s (QP’s) opinion, all appropriate Exploitation Permits have been acquired and obtained to conduct the work proposed for the property.

The next renewal dates for the Exploitation Permits are 05 November 2029 and 06 March 2030 and the current life of mine (LOM) plan for the Kibali Mineral Reserves extends beyond these dates. The DRC Mining Code (2002) includes a provision for the renewal of all Exploitation Permits for a successive period of 15 years, provided the holder has not breached the obligations in respect of permit fee and annual surface rights fee payments, and upholds all environmental standards set out in the Exploitation Permit. Furthermore, the Exploitation Permit holder must provide the appropriate government departments with a monthly mining activity report and quarterly exploration reports.

All the Exploitation Permit fees and taxes relating to Kibali’s exploitation rights have been paid to date and the concession is in good standing.

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The DRC Mining Code (2002) and associated regulations have been amended with an updated Mining Code which came into force on 09 March 2018 (the “DRC Mining Code (2018)”) and the related amended mining regulations which came into force on 08 June 2018.

The following changes made to the DRC Mining Code (2002) in 2018 introduced a series of changes at Kibali: (i) royalty charges were increased from 3.5% to 4.7%, which is not anticipated to materially impact the LOM profitability; (ii) various increases in import and other duties from 4% to 7% depending on consumable type, which is not anticipated to materially alter the LOM profitability, (iii) a super-tax profit has been promulgated based on the feasibility study prepared at the time the approval was given for the construction of the Project and accordingly, such a tax would be applicable only if the average annual gold price was in excess of $2,000/oz.

The QP is not aware of any risks that could result in the loss of ownership of the deposits or loss of the permits, in part or in whole.

The QP is not aware of any other significant factors and risks that may affect access, title, or the right of ability to perform work on the property.

1.3

History

The first documented discovery of gold in the NE of the DRC is attributed to Hannan and O’Brien in 1903, who were sent by Kind Leopold and found nationals washing alluvial gravels for gold.

Historical gold production from the Kilo and Moto areas between 1906 and 2009 is estimated to be approximately 11 Moz Au, half of which came from alluvial deposits. Mining operations were conducted by the Belgian Government via the SOKIMO, which was established in 1926. Most of the mining activity within the Project area was undertaken during the 1950s but accurate production records have been lost over the years of civil unrest in the region. Gorumbwa, Agbarabo and Durba deposits are believed to have produced more than 60% of the over 3 Moz of recorded gold production from the Moto area.

After independence in 1960, gold production dropped sharply as mining was mainly undertaken by artisanal workers and small-scale alluvial operations. SOKIMO changed its name to OKIMO in 1966 and was the main operator in the Project area. Sporadic underground mining was conducted in the Project area after 1960, however this is believed to be of a remnant nature and as such negligible amounts of gold were produced. Accurate production records are not available due to the civil unrest in the region during the 1980s and 1990s. The DRC Governmental Entity OKIMO was later transformed back into SOKIMO in December 2010.

The KCD deposit was originally discovered by a joint venture (JV) between Barrick and AngloGold Ashanti in 1998. The Barrick and AngloGold Ashanti JV completed several drilling programmes, mainly concentrated at KCD and Pakaka. AngloGold Ashanti and Barrick withdrew from the Project in 1998 due to local unrest and civil war.

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Moto Goldmines Limited (Moto) acquired the available 70% stake in the Project in 2004. Moto completed a pre-feasibility study in 2006, a Feasibility Study in December 2007, and an Optimised Feasibility Study in March 2009.

In July 2009, Randgold Resources Limited (Randgold) and AngloGold entered a 50/50 JV, which acquired Moto and their 70% ownership of the Project. In December 2009, the JV acquired an additional 20% shareholding of Kibali Goldmines from SOKIMO. The DRC Government remained a partner in the Project through OKIMO retaining a 10% interest.

Kibali Goldmines undertook an update of the previous feasibility study in 2009, which doubled the declared Mineral Reserve to over 10 Moz Au. Subsequently, construction was approved in 2012 and the Mine has been developed in two key phases.

Phase 1 encompassed the construction and mining of the KCD open pit operation. Mining commenced in July 2012, with construction of the processing plant and commissioning of the oxide processing circuit which began in the third quarter of 2013. Kibali poured its first gold in September 2013, ahead of plan, and started commercial production in the fourth quarter of 2013. A 36-unit high speed thermal power station was constructed to support the power generation requirements of the mine, together with the first of three hydropower stations, Nzoro 2. This first phase of development was completed in December 2014.

Phase 2 compromised of construction for underground mine development, including the vertical shaft and twin declines, in addition to the associated Project infrastructure to support mining of other satellite open pit operations including Mengu-Hill, Pakaka, Kombokolo and Rhino. During Phase 2, Ambarau, was commissioned at the start of 2017 and, Azambi was completed in 2018. A battery energy storage system was incorporated in 2020 to improve power stability. Nzoro 2, Ambarau and Azambi hydropower plants significantly reduce dependence on diesel fired power plants and thus also reduce operational greenhouse gas emissions (GHG). In addition, the existing Nzoro 1 hydrostation was refurbished and is exclusively used to provide power to the local community. Commissioning of the sulphide circuit began early in 2014 and production has steadily ramped-up since then with the mine now consistently exceeding its processing design capacity.

1.4

Geology and Mineralisation

The Kibali deposits are hosted within the Kibali Greenstone Belt (otherwise referred to as Moto granite-greenstone terrane), bounded to the north by the West Nile Gneiss and to the south by plutonic rocks of the Watsa district. The Kibali Greenstone Belt is an elongate WNW-ESE trending terrane containing Archean aged volcano-sedimentary conglomerate, carbonaceous shales, siltstone, banded iron formations, sub aerial basalts, mafic intermediate intrusions (dykes and sills) and multiple intrusive phases that range from granodiorite, to gabbroic in composition. Based on textures and types of lithologies present in the stratigraphy, the rocks within the Project area are interpreted as having been laid down in an aqueous environment.

The majority of the primary lithologies are clastic (sedimentary) in origin, possibly being developed in a regional extensional environment such as a rift graben or half graben. At Kibali,

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the gold deposits are largely hosted in siliciclastic rocks, banded iron formations, and cherts that were metamorphosed under greenschist facies conditions, situated along a curvilinear zone 20 km long and up to one km in width, known as the ’KZ Trend’. Gold mineralisation is concentrated in gently NE to NNE-plunging fold axes whose orientations are generally parallel with a prominent lineation in the mineralised rocks.

The Kibali deposits differ from many orogenic gold deposits as they are hosted within a thrust stack sequence with ductile to brittle-ductile deformational structures and a complex folding history. There are two principal structure sets: NW-SE striking, NE dipping thrust faults and a series of sub-vertical NE-SW shear structures both of which, in association with the folding, are considered important mineralising controls. Unlike many other orogenic gold deposits, mineralisation within the Kibali district typically lacks significant phases of quartz-rich veins.

The mineralised deposits of the Kibali district are associated with halos of quartz, ankerite, and sericite (ACSA-A) alteration that extend into the adjacent rocks. Areas of economic mineralisation are defined where the project scale ACSA-A alteration is locally overprinted by the ankerite-siderite, pyrite alteration assemblage (ACSA-B) that hosts the gold mineralisation. The gold bearing sulphides consist of disseminated pyrite, minor pyrrhotite, and arsenopyrite. The auriferous pyrite occurs as both ‘salt and pepper’ disseminated fine grains and bleb-like clusters of disseminated grains.

The KCD deposit is the principal mineralised occurrence along the Sessenge-KCD Trend. It consists of five semi-vertically stacked lodes; 3000, 5000, 9000, 11000 and 12000, hosted within the volcano-sedimentary units. The location of the individual lodes within the KCD deposit are intimately controlled by the position, shape, and orientation of a series of gently NE-plunging tight to isoclinal folds. The lodes may be linked genetically by large-scale recumbent folding developed between two bounding NE trending structures.

Higher grade zones of strong to intense alteration overprint and texturally-destructive foliation and lithological textures. These are broadly categorised as the 3000 lodes, 5000 lodes, and the 9000 lodes, all of which plunge towards the NE at low to moderate angles (approximately 30°) with drilling intercepts indicating a down plunge continuation of approximately 2,000 m (remaining open down plunge).

The 3000 lode crops out in the present open pit (Karagba) and is the western-most lode. It is approximately 300 m in width, 30 m thick, and has a broad gentle and open semi-synclinal form to its plunge. The 5000 lode outcrops slightly east and south of the 3000 lode (Chauffeur and Durba) and forms the majority of the topographically elevated area known as Durba Hill. The lodes are more sub-vertical in attitude than the 3000 and 9000 lodes and are of a consistently higher grade. The 9000 lode outcrops out to the south of Durba Hill, forming the Mineral Resources in the Sessenge open pit. The 9000 lode is comprised of two main lodes: 9101 and 9105. The 9105 lode is of a similar shape and attitude as the 5000 lode, with the flat underlying 9101 lode extending up-plunge to Sessenge open pit mineralisation. The 11000 and 12000 lodes were discovered during deep drilling, and were subsequently followed up plunge, where the 11000 merges with the 5000 and 9000 lodes and the 12000 lode crops out at Sessenge SW.

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Both the Gorumbwa and Kombokolo deposits occur along a NE trending mineralised corridor located 800 m to the west of the main Sessenge-KCD structural zone. Both are considered to be formed from the same mineralising event, with similar alteration and structural characteristics to the KCD deposit but significantly smaller in size. The Gorumbwa deposit was mined by SOKIMO in 1955 from both underground and via a small open pit operation, with total production estimated to have been approximately 2.8 Mt at 7 g/t Au for 630 koz. The underground and open pit workings are presently collapsed and flooded. The mineralisation consists of a series of stacked ‘lenses’ that variably extend down plunge for a length of 1,000 m at an average width of 200 m and which have been mined to a depth of 400 m below topographic surface.

The Mengu Hill deposit lies on the KZ North structure, to the NW of Pakaka and to the south of Mofu-Oere. The stratigraphy in the vicinity of the deposit is dominated by a meta conglomerate unit that is interbedded with fine-grained sediments, siliceous sericite schist and minor mafic volcanic rocks. These lithologies overlay a massive magnetite and specular hematite ironstone-chert unit that has weathered to create the topographic high, Mengu Hill, the ironstone protecting the northern face from weathering and erosion. Mineralisation is associated with silica-ankerite-pyrite alteration that is focused within the ironstone unit and along its contact with the overlying conglomerate unit. The mineralised lens is cigar like in shape and plunges shallowly to the NNE within a zone of ACSA-B altered rock in the upper part of a thick banded iron formation (BIF) unit. Further barren ACSA-B altered rock occurs widely within the surrounding BIF unit beyond the mineralised lode, unlike ACSA-B altered rocks at the KCD deposit that are invariably mineralised to some degree (Allibone et al, 2020). The Mengu Hill mineralisation averages a width of 150 m and continues 700 m down plunge to a depth of 250 m below the topographical surface. Mineralisation remains open down plunge.

The Aerodrome-Pakaka-Pamao deposits are located along the KZ North trend, in the gently NNE- to E-dipping shear zone (Allibone et al., 2020). Gold mineralisation at Aerodrome-Pakaka-Pamao is hosted in altered (iron carbonate and chlorite with lesser amounts of silica, sericite, pyrrhotite, pyrite, and auriferous disseminated arsenopyrite) and sheared meta volcaniclastic rocks (Allibone and Vargas, 2017) interbedded with minor tuffaceous units. The presence of significant arsenopyrite at Pakaka distinguishes it from other deposits and prospects along the northern half of the KZ Trend (Allibone et al., 2020). The mineralised zones are associated with pervasive silicification with local preservation of breccia textures that have been overprinted by the dominant S1 fabric. Higher gold grades appear to correlate well with the presence and abundance of pyrite-arsenopyrite, and empirically the high-grades appear to be spatially associated with the intersection of the NW trending D1 thrust surface, and a NE trending strain corridor. The structures combine to produce a broad NE plunging open anticlinal structure, with Pamao on the west limb, and Pakaka on the east. The Pakaka mineralisation continues down plunge beyond the limits of the drilling and represents a further exploration potential. The Pakaka mineralisation extends over a strike length of 1,000 m, averages a thickness of 30 m and has been identified to a depth of 350 m below surface. The weathering profile at Pakaka is relatively deep up to 70 m.

The Mengu Village deposit is located near the NW end of the Pakaka-Mengu Trend, the mineralisation is tabular in form, trending NW and dipping shallowly to the NE. The mineralisation is approximately 150 m in strike length with an average thickness of 15 m and has been identified

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to a depth of 150 m below the surface. The mineralisation is hosted by conglomerates with thin ironstone and carbonaceous shale intercalations.

The Megi-Marakeke-Sayi deposit comprises three individual deposits, Megi, Marakeke, and Sayi, separated by lower grade mineralisation but are mined in a single open pit. Megi-Marakeke-Sayi is located midway along the Pakaka-Mengu Trend with mineralisation developed in a variably carbonate-sericite-silica altered basalt and ironstone-chert, that dips to the NE at approximately 30° and strikes to the WNW. The Megi-Marakeke-Sayi deposit occurs as multiple tabular lenses typically between 10 m to 30 m thick that trends NW and dips gently to the NE. The mineralised zone has a strike length of approximately 1,000 m and extends 200 m down dip.

The Ikamva-Kalimva and Oere deposits are all located along the major lineament of the KZ North Trend, north of Mengu Hill where the KZ structure rotates to the NNE. These deposits are broadly similar in geology, consisting of hanging wall volcanic to volcaniclastic succession associated with BIF, intrusions, and carbonaceous shale. These deposits are characterised by an intense shear deformation associated with widespread carbonate-chlorite-quartz altered rocks. However, Kalimva is mostly dominated by a variable-intensity chlorite-quartz-carbonate-pyrrhotite±pyrite-ilmenite assemblage, while Ikamva and Oere are dominated by ACSA alteration (quartz-carbonate-sericite±subordinate chlorite-pyrite) with a distinctive buff-coloured variant of ACSA-A and a texturally destructive ACSA-B assemblage (FeCO₃-quartz±chloritoid±magnetite-pyrite) often spatially associated with mineralisation (Stenhouse, 2020). The mineralisation lodes in Kalimva, show a shallowly NNE-plunging ore-shoot along a moderate to steeply E-dipping structure locally called the Kalimva Deformation Zone and interpreted as an equivalent of lower layer-parallel shear at the Ikamva deposit, characterised by a narrower chlorite-altered high strain BIF (Stenhouse, 2020).

Status of Exploration

Greenfields exploration will include testing a number of grassroots targets identified by the 2018 stream sediment survey. Initially during 2022, follow up works will include geological mapping, local soil sampling grids, and rock chip channel sampling at Makoro, Abimva and Marabi. If successful, targets will be further tested with scout drilling. Additional anomalous catchments will also be tested during following three to five years to sustain a level of exploration target turnover that ultimately supports the mines depletion replenishment pipeline for several years.

Brownfields exploration includes the current underground drilling at KCD, aimed at defining additional extensions to mineralisation to increase the underground Mineral Resources and Mineral Reserves over the next five years. Drilling is completed from dedicated exploration drill drives particularly in the down and up plunge of the 3000 lodes and down plunge of the 5000, 9000, and new 11000 lodes. Analysis of deeper underground opportunities below the base of the existing shaft is planned to be conducted, including down plunge extensions, testing in both the hanging wall and foot wall of the KCD system, refining of the 12000 lode conceptual model, and identification of any new potential lodes that can be connected to the existing KCD underground infrastructure. Execution of 2D seismic lines in the KCD area is also planned to support exploration deeper mineralisation.

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Additionally, brownfields exploration will continue across a number of satellite pit, including but not limited to; Gorumbwa, Pakaka, Kombokolo, Mengu Hill and Ikamva. These pits will be drill tested for down plunge extensions to mineralisation and evaluate their economic viability for further smaller satellite underground operations to support the mine life extension outside of the existing LOM.

Combined exploration efforts are planned to target the delineation of satellite deposits within the gaps between and along the structural corridors of existing Mineral Resources and Mineral Reserves. This is planned with the goal of identifying and evaluating additional targets to add to the open pit Mineral Resources and Mineral Reserve, maintaining a robust depletion replenishment pipeline for several years. During 2022 drill programmes are planned at Oere (North and South extensions), Mengu Village and Ikamva East. Ongoing drilling is also planned in the Gorumbwa-Sessenge-KCD gap to test the concept of combining the three pits especially considering that the Gorumbwa and Sessenge pits now merge.

1.5

Mineral Resource Estimate

The Mineral Resource estimates have been prepared according to the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) 2014 Definition Standards for Mineral Resources and Mineral Reserves dated 10 May 2014 (CIM (2014) Standards) as incorporated with National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101). Mineral Resource estimates were also prepared using the guidance outlined in CIM Estimation of Mineral Resources and Mineral Reserves (MRMR) Best Practice Guidelines 2019 (CIM (2019) MRMR Best Practice Guidelines).

Quality assurance and quality control (QA/QC) has been undertaken across the life of exploration to minimise errors. A standard operating procedure (SOP) outlines Kibali Goldmines approach to QA/QC which meets industry best practice.

The Kibali Mineral Resources consists of the KCD, Sessenge, Pakaka, Mengu Hill, Gorumbwa, Megi-Marakeke-Sayi, Pamao (inclusive of Pamao South, also known as Tete Bakangwe), Kombokolo, Kalimva-Ikamva, Aerodrome, Oere and Mengu Village deposits. Only KCD (underground and open pit), Sessenge, Gorumbwa, Pamao (including Pamao South), Aerodrome, and Oere were updated in 2021, following additional data from drilling, and/or updated geological mapping. Pamao South, Mengu Village, and Oere are new additions to the Kibali Mineral Resources for 2021.

Total Mineral Resources for Kibali are estimated to be 140 Mt at an average grade of 3.41 g/t Au for 15 Moz Au in the Measured and Indicated categories and 23 Mt at an average grade of 2.7 g/t Au for 2 Moz Au in the Inferred category.

A summary of the Kibali Mineral Resources is presented in Table 1-2. These Mineral Resources have been depleted to 31December 2021 using the mined-out surfaces and voids.

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Table 1-2 Kibali Mineral Resources as of 31December 2021



Type	Category	Tonnes (Mt)	Grade (g/t Au)	Contained Gold (Moz Au)	Attributable Gold¹ (Moz Au)

Stockpiles	Measured	0.32	3.17	0.032	0.015
Open Pits	Measured	15	2.24	1.1	0.50
	Indicated	45	2.25	3.3	1.5
	Inferred	8.2	2.1	0.55	0.25
Underground	Measured	32	4.63	4.7	2.1
	Indicated	48	4.06	6.3	2.8
	Inferred	15	3.0	1.4	0.64
Total Mineral Resources	Measured	48	3.84	5.9	2.6
	Indicated	93	3.18	9.5	4.3
	Measured and Indicated	140	3.41	15	6.9
	Inferred	23	2.7	2.0	0.89

Notes:

Attributable Gold (Moz) refers to the quantity attributable to Barrick based on Barrick’s 45% interest in Kibali Goldmines. Mineral Resources are reported on a 100% and attributable basis.

The Mineral Resource estimate has been prepared according to CIM (2014) Standards and using CIM (2019) MRMR Best Practice Guidelines.

All Mineral Resources tabulations are reported inclusive of that material which is then modified to form Mineral Reserves.

Open pit Mineral Resources are reported within the $1,500/oz Au pit shell at a tonnage weighted average cut-off grade of 0.77 g/t Au.

Underground Mineral Resources in the KCD deposit are Mineral Resources which meet a cut-off grade of 1.62 g/t Au and are reported insitu within a minimum mineable stope shape, at a gold price of $1,500/oz Au.

Mineral Resources were estimated by Christopher Hobbs CGeol, MSc, MCSM, FAusIMM, an officer of the company and QP, and reviewed by Simon Bottoms CGeol, MGeol, FAusIMM, an officer of the company and QP.

Numbers may not add due to rounding. Tonnes and contained gold are rounded to 2 significant figures. All Measured and Indicated grades are reported to 2 decimal places whilst Inferred Mineral Resource grades are reported to 1 decimal place.

The QP is not aware of any environmental, permitting, legal, title, taxation socioeconomic, marketing, political, metallurgical, fiscal, or other relevant factors, that could materially affect the Mineral Resource estimate.

The cut-off grade selected for reporting each of the open pit Mineral Resources corresponds to the in-situ marginal cut-off grade at either fresh, transitional or saprolite oxidation states, using a gold price of $1,500/oz Au. The pit shell selected for limiting each of the Mineral Resources also corresponds to a gold price of $1,500/oz Au. Reasonable prospects for eventual economic extraction are demonstrated as a result of this pit optimisation process.

Underground Mineral Resources were reported using Mineable Stope Optimiser (MSO), effectively within a minimum mineable stope shape, applying reasonable mineability constraints, including a minimum mining width, a reasonable distance from current or planned development, and a measure of assumed profitability at the related resource cut-off grade, thus deemed as having reasonable prospects for eventual economic extraction.

Stockpiles are comprised of mineralised material stored at the surface run of mine (ROM) pad, originating from both open pit and underground production. Each stockpile is filled with similar material types, with an established grade range and oxidation state, tracked as part of normal mining operations and metal accounting. The stockpiles are measured by weekly drone survey. Grade and tonnage of open pit stocks are estimated according to source dig blocks and number of truck counts, using a weighbridge to adjust for fluctuations in both density and truck fill factor. Grade and tonnage of underground floor stocks are estimated according to shaft skip weights

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and ore pass truck counts and their source blasts from stopes, adjusting for the presence of paste dilution.

All models have been depleted using the December 2021 mined out stopes and surfaces as appropriate.

1.6

Mineral Reserve Estimate

As of 31December 2021, the total Proven and Probable Mineral Reserves in open pits, underground, and stockpiles (100% basis) is estimated to be 83 Mt at an average grade of 3.60 g/t Au, containing approximately 9.6 Moz Au.

The Mineral Reserve estimates have been prepared according to the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) 2014 Definition Standards for Mineral Resources and Mineral Reserves dated 10 May 2014 (CIM (2014) Standards) as incorporated with National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101). Mineral Resource estimates were also prepared using the guidance outlined in CIM Estimation of Mineral Resource and Mineral Reserve Best Practice Guidelines 2019 (CIM (2019) MRMR Best Practice Guidelines).

The Mineral Reserves have been estimated from the Measured and Indicated Mineral Resources and do not include any Inferred Mineral Resources. The estimate uses updated economic factors, the latest Mineral Resource and geological models, geotechnical and hydrological inputs, and metallurgical processing and recovery updates. The QPs responsible for estimating the Mineral Reserves have performed an independent verification of the block model tonnes and grade, and in their opinion the process has been carried out to industry standards.

For the open pit mines, economic pit shells were generated using the Lerch-Grossman algorithm within Whittle software and then used in the open pit mine design process and Mineral Reserve estimation.

For the KCD underground mine, the Datamine MSO was used to evaluate the geological block model to create overall mining shapes. Preliminary stope wireframes were created and planned dilution was added to the mineable stope shape. Datamine’s Enhanced Production Scheduler (EPS) software was used to estimate the diluted mined tonnes, grade, and contained metal of the Mineral Reserves. Stopes with a diluted grade below the cut-off grade (2.02 g/t Au) were excluded from Mineral Reserves.

The planning process incorporated appropriate modifying factors and the use of cut-off grades and other technical-economic investigations. Mineral Reserves are stated:

●

As of 31December 2021

●

At a gold price of $1,200/oz Au

●

As ROM grades and tonnage as delivered to the plant

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A financial model was constructed to demonstrate that the Mineral Reserves are economically viable.

The total Kibali open pit and underground Mineral Reserves as of 31December 2021 are summarised in Table 1-3.

Table 1-3 Kibali Mineral Reserves as of 31December 2021



Type	Category	Tonnes (Mt)	Grade (g/t Au)	Contained Gold (Moz Au)	Attributable Gold¹ (Moz Au)
Stockpiles	Proven	0.32	3.17	0.032	0.015
Open Pits	Proven	11	2.28	0.79	0.35
Open Pits	Probable	26	2.51	2.1	0.95
Underground	Proven	21	4.54	3.0	1.4
Underground	Probable	25	4.54	3.7	1.6
Total Mineral Reserves	Proven	32	3.76	3.9	1.7
	Probable	51	3.50	5.8	2.6
	Proven and Probable	83	3.60	9.6	4.3

Notes:

Attributable Gold (Moz) refers to the quantity attributable to Barrick based on Barrick’s 45% interest in the Kibali Goldmines. Mineral Reserves are reported on a 100% and attributable basis.

The Mineral Reserve estimate has been prepared according to CIM (2014) Standards and using CIM (2019) MRMR Best Practice Guidelines.

Open pit Mineral Reserves are reported at a gold price of $1,200/oz Au, and an overall weighted average cut-off grade of 0.96 g/t Au, including dilution and ore loss factors

Underground Mineral Reserves are reported at a gold price of $1,200/oz Au and a cut-off grade of 2.02 g/t Au.

Open pit Mineral Reserves were estimated by Shaun Gillespie, Reg Eng Tech, FAusIMM, an officer of the company and QP, and reviewed by Simon Bottoms CGeol, MGeol, FAusIMM, an officer of the company and QP.

Underground Mineral Reserves were estimated by Ismail Traore, MSc, FAusIMM, an officer of the company and QP, and reviewed by Simon Bottoms CGeol, MGeol, FAusIMM, an officer of the company and QP.

Numbers may not add due to rounding. Tonnes and contained gold are rounded to 2 significant figures. All Proven and Probable grades are reported to 2 decimal places.

The year-end 2021 Mineral Reserve estimate shows a net increase of 0.19 Moz Au when compared to the estimate for year-end 2020. This is mainly due to positive model changes resulting from infill grade control drilling, new deposits, pit size changes and various adjustments to the economic parameters, partially offset by mining depletion.

The QPs have performed an independent verification of the block model tonnes and grade, and in their opinion, the process has been carried out to industry standards.

The QPs are not aware of any environmental, legal, title, socioeconomic, marketing, mining, metallurgical, infrastructure, permitting, fiscal, or other relevant factors that could materially affect the Mineral Reserve estimate.

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1.7

Mining Method

Over the Kibali Mineral Reserve LOM, a total of 83 Mt of ore at 3.60 g/t Au is expected to be produced over 13 years up to 2034. Mineral Reserves supplied to the plant during this period, including stockpile changes, will be 83 Mt at an average grade of 3.60 g/t Au resulting in 8.6 Moz Au recovered at an average processing recovery of 89 %.

The Kibali open pit operation will continue until 2033 and the underground until 2034 based on current Mineral Reserves.

A total of 46 Mt of ore will be mined from the underground operations with a further 37 Mt mined from the open pits based on Mineral Reserves.

In the QP’s opinion, the parameters used in the Mineral Resource to Mineral Reserve conversion process are reasonable.

Open Pit

Open pit mining is carried out using conventional drill, blast, load, and haul surface mining methods.

From 2022 onwards, open pit production will come from the Sessenge, Aerodrome, Pamao, Gorumbwa, Megi-Marakeke-Sayi, Kalimva-Ikamva, Oere, Pakaka, and KCD deposits. The Mengu Hill, Mofu, Kombokolo and Rhino pits were depleted in 2017.

Open pit mining is conducted by contractor Kibali Mining Services (KMS), a local subsidiary of DTP Terrassement, using either free-dig or conventional drill, blast, load, and haul methods.

The mining equipment is jointly owned by a subsidiary of Barrick and the contractor’s parent, who also operates at Barrick’s Loulo-Gounkoto Mine in Mali and Tongon Mine in Côte d’Ivoire.

All the mineral deposits are characterised by the presence of a near-surface groundwater table with the potential for high groundwater inflows into the pits. The possible impacts of ingress of groundwater are investigated prior to mining and during the mining activities. Dewatering well systems are installed for all pits to lower the groundwater level prior to commencement of mining. A system of dewatering trenches are procedurally established prior to commencement of mining in each of the pits, preventing the inflow of any surface water to the active mining areas.

The upper levels of the open pits are usually in weathered material, which typically is free digging material. Once fresh (unweathered) rock is encountered, drilling and blasting is required. Emulsion explosives are supplied as a down-the-hole service by the contracted explosive supplier Orica.

Free digging in the upper levels uses 5 m high benches, with 10 m benches used for drilling and blasting operations. The 10 m benches containing ore are excavated in three flitches of equal height.

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Opportunities exist to upgrade and convert the Inferred Mineral Resources within the current pits to Mineral Reserves with further conversion drilling, but any Inferred Resources within pit designs are not reported as Mineral Reserves.

Under current Mineral Reserves, the open pit end of life is estimated at year 2033 based on current Mineral Reserves. The addition of future open pit Reserves from additional exploration sites have the potential to extend open pit mining post-2033.

Underground

The Kibali KCD underground mine is designed to extract the KCD deposit directly beneath the KCD open pit. A 50 m crown pillar separates the pit bottom from the top of the underground mine. The Kibali underground mine is a long hole stoping operation producing at a rate of 3.8 million ore tonnes per year. Development of the underground mine commenced in 2013. Stoping commenced in 2015 and ore production has ramped up to 1.8 Mt in 2017 and 3.6 Mt in 2021. Initial production was truck hauled by a twin decline to surface. In 2017, the haulage shaft (740 m deep) and materials handling system was commissioned. From 2018 onwards, underground ore has predominantly been hoisted up the shaft. The decline to surface will be used to haul some of the shallower zones and to supplement shaft haulage.

A major pump station has been installed near the shaft bottom with redundant capacity in the pumps and pipelines to the surface.

A significant portion of the capital and access development for the mine is in place. To date 43,609 m of capital and access development has been completed. The current LOM plan contains a further 9,928 m of capital lateral development based on Mineral Reserves.

There are four main mineralised zones, 5101, 5102, 9101 and 9105 that contribute the bulk of Mineral Reserves. Five other mineralised zones, 3101, 3102, 5104, 5105 and 5110 contribute approximately 18 % of the remaining Mineral Reserves.

Ore from stopes is loaded (both by teleremote and conventional manual loaders) from the stopes into the eight ore passes via finger raises on the respective levels. This ore is then transferred by Autonomous load haul dumpers (LHDs) into two coarse ore bins and then into two primary crushers, followed by two fine ore bins and independent skip loadout conveyors near the shaft bottom.

The proposed mining methods are variants of long hole open stoping with cemented paste:

●

Primary / Secondary long hole open stoping (primary 20% of Mineral Reserve tonnes, secondary 33% of Mineral Reserve tonnes) is used in the wider zones, with 35 m interval heights where stopes are mined either as single lift or multiple (up to four) lifts, depending on stope geometry and the geotechnical stable span.

●

Advancing face long hole open stoping (29% of Mineral Reserve tonnes) is used where the mineralisation has a shallower plunge (approximately 20° to the NE), where stopes are mined with variable interval heights between 25 m and 35 m to optimise extraction.

●

Longitudinal open stoping (18% of Mineral Reserve tonnes) is used in narrow zones (< 15 m width) with variable interlevel heights between 20 m and 30 m.

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No significant failures of the openings in the underground workings have occurred. The rock assessed for the rock mass model is ranked as good.

The underground mining operations have been owner operated by Kibali Goldmines since 2018.

The paste backfill plant treats the Kibali tailings from the flotation circuit by de-watering processes (filtration) to produce a paste containing binder, which is delivered to underground stopes under gravity or pump via a distribution piping system. The paste plant has been designed to treat a feed rate of 292 tph of dry tailings solids and produce nominally 190 m³/hr of paste fill. The paste plant is fully automated with its own fully equipped laboratory. The paste is transported to the stopes underground via a single borehole (duty and stand-by). Paste is subsequently transported horizontally along the levels to the upper stopes. Internal boreholes take paste fill to the lower levels.

Under current Mineral Reserves, the underground end of life is estimated at year 2034. The addition of future underground Mineral Reserves from additional lodes such as the 11000 lode has the potential to extend surface mining post 2034.

1.8

Mineral Processing

The Kibali gold processing plant comprises two largely independent processing circuits, the first one designed for oxide and transition ores and the second for sulphide refractory ore. However, both circuits are designed to process sulphide ore when the oxide and transition ore sources are no longer available. The flow sheet, depicted in Figure 1-1 comprises crushing, ball milling, classification, gravity recovery, a conventional Carbon-in-leach (CIL) circuit, flash flotation, also conventional flotation, together producing a concentrate which goes to ultra-fine-grinding and a dedicated intensive cyanide leach. This process consists of well tested technology in the gold industry and is appropriate for the style of mineralisation present at Kibali.

Extensive metallurgical test work campaigns have been completed across all mineral deposits in Kibali that form part of the declared Mineral Reserves. These have consistently demonstrated two distinct behavioural patterns, the first of which exhibits free-milling characteristics suitable for gold extraction by a conventional CIL metallurgical process. The second of which exhibits a degree of refractoriness, where straight cyanidation returns gold dissolutions considered to be too low for optimal plant operation due to the presence of occluded gold particles within sulphide minerals. It has been demonstrated that a finer grind will expose a portion of this additional gold for leaching so that the recovery is enhanced to economically acceptable levels.

Figure 1-1 provides a simplified flowsheet of the Kibali Processing Plant depicting the two discrete streams.

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Source: Kibali Goldmines, 2021

Figure 1-1 Simplified Flowsheet of the Kibali Processing Plant Depicting Two Discrete Streams

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Source: Kibali Goldmines, 2021

Notes:

Blue Markers = Fresh, Green Markers = Transition, Red Markers = Oxide

Figure 1-2 Initial Hole Composite Dissolutions

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The current metallurgical recoveries expected at Kibali derived from both test work and actual plant operational knowledge and used in the financial model can be found in Table 1-4.

Within the existing Kibali process plant, ore is blended using both KCD underground ore plus ore sourced from satellite open pits at Kibali. The process plant has demonstrated improvements in throughput capability, performing beyond design capacity of 7.2 Mtpa as depicted in Figure 1-3, at reasonably consistent recovery performance (Figure 1-4). The fluctuations seen in the monthly recovery performance is mainly driven by the different feed blends that were fed during the course of the year. The October low recovery of 87.8% was mainly due to the high residues emanating from the circuit changes between full sulphide and sulphide/oxide campaign treatment. The changeover often results in flashing out of high residues that build up in the CIL circuit coupled with reduced residence time during the changeover period.

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Table 1-4 Summary of Average Recovery for All the Samples



Ore Source	Weathering	Average Recovery Primary Only (%)	Average Recovery Primary + Tail Leach (%)	Average Recovery Oxide Process (%)	Feasibility or Financial Model Recovery (%)

KCD	Fresh_OP	86.4	89.2		86.1
	Fresh_UG	89.0	93.4		90.0
	Transition	66.6	91.3		90.1
	Oxide			90.1	90.1

Sessenge	Fresh	72.7	81.2		81.0
	Transition		80.3		75.9
	Oxide			90.4	90.3

Pakaka	Fresh	78.1	82.3		80.2
	Transition				81.3
	Oxide			96.9	88.7

Mengu Hill	Fresh	69.2	72.2		70.1
	Transition	84.4	89.9		89.3
	Oxide			92.6	89.3

Kombokolo	Fresh	70.3	75.2		85.0
	Transition	78.9	95.3		95.9
	Oxide			85.0	85.0

Pamao	Fresh	74.5	85.5		85.0
	Transition				85.0
	Oxide			95.8	90.9

Kalimva-Ikamva	Fresh	89.38	93.64		89.0
	Transition			89.88	89.0
	Oxide			91.05	90.0.

3000 lode Down Plunge KCD UG	Fresh	88.36	89.56		89.4
	Transition
	Oxide

5000 lode Down Plunge KCD UG	Fresh	78.58	88.03		89.5
	Transition
	Oxide

Aerodrome	Fresh	79.05	85.83		85.83
	Transition			88.96	88.0
	Oxide			90	90.0

Pamao South	Fresh	81.2	88.0		86.5
	Transition			90.07
	Oxide			90.64

Megi-Marakeke-Sayi	Fresh	87.37	90.33		89.5
	Transition			92.58	90.0
	Oxide			94.29	90.0

Oere	Fresh	82.49	88.2		87.0
	Transition			87.43	86.5
	Oxide			88.57	88.0

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Figure 1-3 Kibali Throughput Performance versus 7.2 Mtpa Design Throughput

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Figure 1-4 Kibali Processing Plant Process Recovery for Year 2021

Actual process and plant engineering operating costs for 2019, 2020, and 2021 are denoted in Table 1-5.

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Table 1-5 Actual Process and Plant Engineering Operating Costs for 2019, 2020 and 2021



Cost	Units	2019	2020	2021

Fixed Cost

Consultants	$ 000	537	307	256

Contractors – Assays	$ 000	1,676	1,639	1,680

Contractors – Oxygen	$ 000	109	193	-52¹

Equipment Hire	$ 000	2,597	1,898	1,815

General Costs	$ 000	10,396	11,038	13,789

Gold Refining	$ 000	3,443	5,817	5,870

Labour	$ 000	7,796	8,710	8,695

Stores – Other	$ 000	1,730	1,544	1,824

Total Fixed	$ 000	28,284	31,148	33,878

Tonnes Processed	kt	7,513	7,632	7,783

Total Fixed	$/t	3.76	4.08	4.35

Variable Costs

Power	$/t	3.27	1.93	2.12

Reagents – Cyanide	$/t	2.71	2.60	2.69

Reagents – Lime	$/t	1.03	0.65	0.65

Good Issues – Caustic Soda	$/t	0.44	0.41	0.43

Good Issues – Activated Carbon	$/t	0.12	0.11	0.11

Reagents – Other	$/t	1.36	1.39	1.33

Stores – Grinding Media	$/t	0.81	0.90	0.97

Stores – Liners	$/t	0.54	0.55	0.48

Stores – Screens and Panels	$/t	0.05	0.01	0.03

Total Variable	$/t	10.34	8.56	8.81



Total	$/t	14.11	12.64	13.16



Plant Engineering	$/t	3.11	3.22	3.31



Combined Process and Plant Engineering	$/t	17.21	15.86	16.47

Notes:

Included in this amount is a cumulative catchup amount related to an FRS 16 (Financial Reporting Standard 16) adjustment processed to account for the extension of a lease agreement, with the corresponding right of use asset being unwound in the income statement. The impact of the adjustment is not considered material.

LOM processing costs are modelled to be $17.49/t milled (which includes Plant Engineering) based on Mineral Reserves. The actual costs for 2021 were $16.47/t milled, with the key improvements over the LOM as a result of:

Power cost due to the installation of a grid stabiliser and diesel heaters in the elution circuit. The first item has improved the hydro blend, while the second item has reduced plant power consumption from the elution circuit.

Lime cost due to the change from hydrated lime to quick lime, which has lowered price and consumption over the LOM.

Optimisation of cyanide and caustic consumptions.

However, the improvement in 2021 has been partially offset by an increase in gold refining fees due to the lack of direct commercial flights from Nairobi to South Africa following Covid-19 travel restrictions.

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1.9

Project Infrastructure

Kibali is located in the NE of the DRC, approximately 560 km NE of the city of Kisangani and near the international borders with Uganda and South Sudan.

The Project is situated in a rural setting that lacks local infrastructure. Infrastructure in the DRC is generally poor as a result of limited investment in the maintenance of road networks established during colonial times. Historically, the lack of investment was the result of civil unrest and diminished government revenue collection.

The main access points for equipment and supplies for the operation include the major ports of Mombasa, Kenya (1,800 km) and Dar es Salaam, Tanzania (1,950 km). The routes are paved up to the DRC border. Road access is from Kampala, Uganda and is approximately 650 km. The arterial road between Arua and site is unpaved but has been upgraded and serves as the main access route for materials to site. Local roads are generally in very poor states of repair. Supplies typically require two weeks to arrive from Mombasa.

A local certified airstrip with passport control, serves as the primary access point to site for personnel on charter flights from Entebbe, Uganda, which is approximately 470 km SE of the Mine. International air carriers service Entebbe – Doko – Entebbe daily with exception of Saturday and Sunday.

The primary source of raw water supply is rain and spring water catchments with top-up from a borehole system and a final backup from the Kibali River. Raw water is collected and stored in the raw water dam, which has a storage capacity of 9,500 m³. The processing plant requires approximately 46,000 m³ of water per day, which is sourced by reclaiming water from the Flotation Tailings Storage Facility (FTSF) and CTSF1 and CTSF2.

There are two TSFs at Kibali; one for the cyanide containing (CIL) tails and the second one for the sulphide flotation tails. The CIL tailings contain residual cyanide and are contained in an HDPE lined dam. The flotation tails contain are benign and therefore the dam is not lined. The cyanide containing TSFs comprising of concentrate tailings storage facility one, (CTSF1) and concentrate tailings storage facility two (CTSF2) for the CIL tails and the FTSF is dedicated to flotation tails.

Approximately half of the sulphide tailings generated will be used to produce paste backfill for the stoping operations. A paste fill plant filters the sulphide tailings, which are mixed with cement to form a paste fill that is delivered to the underground via a distribution pipe network from the surface.

Since there is no national grid power supply to the site, Kibali is fully dependent on its own generation facilities. The power supply currently comes from a mix of on-site, high-speed diesel generator sets and off-site hydropower stations; Nzoro II is currently producing approximately 22 MW, Ambarau produces 10.6MW and Azambi produces a further 10.2MW, with total peak hydropower capacity of 42.8 MW, which is sufficient to meet the mine power demand. A battery energy storage system was incorporated in 2020 to improve power stability. Nzoro 2, Ambarau

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and Azambi hydropower plants. The site is connected to the hydrostations via a 66 kV overhead line network.

The hydropower system has a combined potential capacity of 42.8 MW of hydropower (at peak) and has backup installed capacity for 43 MW of thermal generation. The load demand of the mine is not constant, power demand at full production is currently between 39 MW and 43 MW, averaging approximately 41 MW.

1.10

Market Studies

Financial evaluation of all Mineral Reserves uses a gold price of $1,200/oz Au. This is in line with Barrick’s corporate guidelines. Gold price sensitivities were run for all the pits.

Royalties payable to the DRC government has increased following the introduction of the DRC Mining Code (2018). A total royalty payable to the DRC government of 4.7% of gold revenue inclusive of 1% shipment fees was used for the Open Pit Mineral Reserve estimate.

Kibali currently pays income tax at a rate of 30% to the DRC government. Due to accelerated depreciation charged on capital expenditure, tax payable is currently 30% of 40% of the total tax rate, whilst 60% of the tax is offset against the accumulated loss. The full rate of 30% tax is expected to commence in 2026, if the assessed loss is depleted.

Gold doré produced at the Mine is shipped from site under secured conditions and sold under agreement to Rand Refinery in South Africa. Under the agreement, Kibali Goldmines receives the ruling gold price on the day after dispatch, less refining and freight costs, for the gold content of the doré gold. Kibali Goldmines has an agreement to sell all gold production to only one customer. The “customer” is chosen periodically on a tender basis from a selected pool of accredited refineries and international banks to ensure competitive refining and freight costs. Gold mines do not compete to sell their product given that the price is not controlled by the producers.

1.11

Environmental, Permitting and Social Considerations

An independent Environmental and Social Impact Assessment (ESIA) for the Kibali mine was completed as part of the Kibali Goldmines Feasibility Study completed in December 2012. Subsequent ESIAs for various Project extensions and new elements have been completed, the most recent of which was in 2020. An Environmental Adjustment Plan (EAP) has been approved by the Direction de Protection de l’Environnement Minier (DPEM) with the purpose of describing any measures that have been or will be taken for the purpose of the protection of the environment. An environmental management plan is in place, and the Kibali operations are ISO 14001:2015 certified and independently audited to continuously improve environmental management. Audits are also carried out to gauge conformance with the International Cyanide Management Code (ICMC); ICMC certification and construction of a cyanide detox plant for the tailings stream is planned to commence in 2022.

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Waste rock is generated and disposed of on Waste Rock Dumps (WRDs) that are located adjacent to the open pits and underground shaft. The waste rock characterisation assessment returned a negative acid generating status. Waste rock is used to build various infrastructural platforms on site, while the remainder is stockpiled on surface or deposited in stopes as backfill. The waste rock has been demonstrated to have moderate to high acid neutralising capacity for the majority of lithologies tested.

Tailings are generated from the plant and disposed of in two separate tailings storage facilities, the FTSF and CTSF, which consists of the CTSF1 and CTSF2. The CTSF is lined and contains materials which are acid producing and which also contain cyanide residues and materials with a higher arsenic content. The concentrate tails are acid producing and contain cyanide residues and arsenic containing materials. A portion of the flotation tailings are used for paste backfill in the KCD underground mine.

Routine environmental monitoring takes place across the site, including dust deposition, noise, arsenic, and weak-acid dissociated (WAD) cyanide sampling, TSF seepage water and tails streams as well as sample collection of drinking water, ground water, surface water and the TSF borehole water.

Environmental incidents are noted in a register which forms part of the Environmental Management System (EMS); the causes and responses are identified, and once completed, the incident is closed out.

A comprehensive water balance model has been compiled for the site, which models flows, inputs and losses throughout the operations (i.e., the open pits, underground workings, process plant, TSFs, water management structures, offices, camp, and sewage treatment facilities). The model includes inputs regarding river water use (e.g., discharges, gains, and losses, volumes of potential savings/recycling opportunities). Opportunities to reuse water within operations has significantly reduced the volume of freshwater abstracted from the Kibali River.

The original vegetation of the Project area has been largely transformed through human activity. Site clearance for the establishment of infrastructure, together with anthropogenic activities has occurred across all vegetation habitat types. Alien invasive plant species occur throughout all habitat types. Despite human pressure, most protected plants species remain within gallery forests (Digby Wells, 2015) that are associated with drainage lines and water courses.

Biodiversity monitoring is ongoing, such as the use of camera traps to detect fauna within the concession. The Biodiversity Management Plan is being updated to reflect additional information on the biodiversity which has been collected. The mine site lies around 65 km south of the Garamba National Park, which lies on the border with South Sudan. A partnership with the Park has been established to support the Garamba National Park’s goals. This partnership provides a wider strategic support for game protection from poachers from the north, and connections with local enforcement networks.

Mine closure costs are updated each year, with increases or decreases in disturbed areas noted and costed; the current cost for rehabilitation and closure of the mine according to the calculation model is estimated to be $23.76 million as of 31^st December 2021 (Digby Wells, 2021).

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The mine is a significant employer to members of the local communities. The mining operations contribute to extended LOM, employment of local Congolese and the growth of the DRC economy. Kibali Goldmines’ policy is to promote nationals to manage the Project. The policy of promoting local employment also extends to its contractors. The mine prioritises local employment and in October 2021, the employee bases were made up of 88% Congolese nationals; more than 70% from the local area. More than 70% of management positions were held by Congolese Nationals. Local procurement is also promoted and is a contractual requirement for contractors. Kibali Goldmines procured in excess of $110 million of goods and services from DRC suppliers in 2021. This includes produce from agribusinesses (e.g., producers of eggs, pork, maize), which is purchased for use in the mine canteens.

Kibali Goldmines follows a resettlement and compensation process that will leave project-affected people (PAPs) in the same or better off position than before the Project intervention, which is in conformance International Finance Corporation (IFC) Performance Standards (PS).

Due to the construction of the Project, since 2012 to date, it was necessary to resettle approximately 36,700 people, from 7,504 households. The Project also displaced around 134 items of community infrastructure, including 13 communal agricultural projects, five communal business/commercial facilities, 12 education facilities, 19 health facilities, nine recreational/community facilities, 39 religious facilities, and 41 water sources. A Resettlement Working Group (RWG) was established as the primary consultation forum to develop and implement a Resettlement Action Plan (RAP). The RAP process was carried out by independent consultants. All primary stakeholders are represented on the RWG.

The Moratorium Zone was expanded in 2020 to incorporate new deposits at Pamao, as well as Kalimva-Ikamva (Moratorium Zone C). These areas have been rezoned and allocated to Kibali for the mine and associated infrastructure. The land was used for residential sites, agricultural, and artisanal and small-scale mining (ASM) before mining.

The Pamao RAP initiated in 2020 includes Pamao North and Pamao South as expansion areas to the Moratorium Zone A to allow mining activities of the Pamao pit. It involves resettling 628 households from two villages whereby 222 households will be physically displaced and 406 will be economically displaced, who were engaged in farming activities within the affected zone but did not reside there. An additional 250 households were affected by the Pamao Diversion Road and Gatanga-Surur Diversion Road, which are both deviating the RN26 National Road and whose section is affected by the Pamao North Zone. The physically affected households will be resettled at the Avokala host site, along with the Kalimva-Ikamva PAPs.

The Kalimva-Ikamva RAP was initiated in 2019 and is still under development. It involves 1,888 households from six villages, whereby 1,141 households are physically displaced and 747 will be economically displaced. An additional 232 households are affected by the host site work at Avokala, and two diversion roads created heading to the host site. Through the RWG consultation, Kibali Goldmines has made funds available in the event that PAPs decide to build infrastructure themselves. In such cases, payments are made in three instalments and full payment will only be made upon completion of construction.

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The Kalimva-Ikamva-Pamao RAP includes construction of water sources, schools, solar power energy, road infrastructure, sports infrastructures, health facilities, cemetery, places of prayer and adequate sanitation at the host site. Guidance was provided by Congolese town planners, as well as the RWG, for a town plan outlining the development of the host site that improves the provision of basic services and social infrastructure.

Stakeholder engagement activities, community development projects and local economic development initiatives contribute to the maintenance and strengthening of Kibali Goldmines’ Social License to Operate (SLTO). A grievance mechanism is in place, and all registered grievances in 2021 were successfully resolved.

ASM remains a concern in the Kibali Exploitation Permit area and the mine is working with provincial authorities to prevent and relocate ASM within the Exploitation Permits.

The QPs consider the extent of all environmental liabilities to which the property is subject to have been appropriately met.

1.12

Capital and Operating Costs

Capital Costs

Kibali is a sustaining capital combined open pit and underground mining operation with the necessary facilities, equipment, and manpower in place to produce gold.

The open pit and underground LOM and capital and operating cost estimates have been completed in sufficient detail to be satisfied that economic extraction of the Proven and Probable Mineral Reserves is justified.

The total capital expenditure from 2018 to 2021 amounted to $484 million. This included $201 million spent on underground mining capital, which represented 42% of total capital expenditure. A total of $61 million, representing 13% of total capital expenditure, was spent on deferred stripping to remove mine waste materials (overburden) to gain access to mineral ore deposits in new pits. A further $43.5 million, representing 9% of total capital expenditure, was spent on capitalised drilling which resulted in LOM extensions and conversion of Mineral Resources to Mineral Reserves. $18 million was spent on permit wide exploration for resource replacement, representing 4% of total capital expenditure. Completion of the hydropower stations accounted for $26 million, or 5% of total capital expenditure, and $33 million for the refurbishment of open pit equipment, or 7% of total capital expenditure.

Capital expenditure over the remaining LOM is estimated to be $715 million (from 2022) based on Mineral Reserves, made up from the allocation of costs as summarised in Table 1-6.

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Table 1-6 LOM Capital Expenditure Based on Mineral Reserves



Description		Value ($M)

Grade control drilling		41

Capitalised deferred stripping		35

Underground capital development and drilling		185

RAP growth capital		18

Drilling capitalised		6

Other sustaining capital		430

Total LOM Capital Expenditure		715

Operating Costs

Kibali Goldmines maintains detailed operating cost records that provide a sound basis for estimating future operating costs.

Costs used for the open pit optimisations were derived from KMS open pit mining contractor’s pricing of the open pit LOM schedule. Underground operations were costed starting in mid-year 2018 as owner costs, when underground mining changed to owner operated.

Labour costs for national employees were based on actual costs. Local labour laws regarding hours of work, employment conditions were also considered, and overtime costs included.

During 2021, costs for processing and general and administration (G&A) were updated based on actuals adjusted with the latest forward estimates, production profiles and personnel levels.

Customs duties, taxes, charges, and logistical costs have been included.

Unit costs used to estimate LOM operating costs based on Mineral Reserves (from 2022) are summarised in Table 1-7. The annual fluctuation in production levels is relatively low, such that the effect of fixed versus variable expenses is minimised.

Table 1-7 LOM Unit Operating Costs Based on Mineral Reserves



Activity	Units	Value

Open Pit Mining	$/t mined	3.44

Open Pit Mining	$/t ore mined	33.00

Underground Mining	$/t mined	36.16

Underground Mining	$/t ore mined	37.95

Processing	$/t milled	17.49

G&A	$/t milled	9.35

Mining Total¹	$/t milled	35.60

Total LOM Net OPEX¹	$/t milled	62.44

Notes:

Total LOM Net of Opex in this table, represents the total amount, before capitalised cost and royalty costs of 4.7% based on the total revenue

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Cost inputs have been priced in real Q4 2021 dollars, without any allowance for inflation or consideration to changes in foreign exchange rates.

The QPs are satisfied that the open pit LOM and cost estimates have been completed in sufficient detail to justify the economic extraction of the open pit Proven and Probable Mineral Reserves.

The QPs are satisfied that the underground LOM and cost estimates have been completed in sufficient detail to justify the economic extraction of the underground Proven and Probable Mineral Reserves.

1.13

Economic Analysis

This section is not required as Barrick, the operator of Kibali for both exploration and mining, is a producing issuer, the property is currently in production, and there is no material expansion of the current annual production planned.

The QP has verified the economic viability of the Mineral Reserves via cash flow modelling, using the inputs discussed in this report.

1.14

Interpretation and Conclusions

Geology and Mineral Resources

QA/QC

Kibali Goldmines has documented standard procedures for the drilling, logging, and sampling processes, which meet industry standards. The geological and mineralisation modelling at Kibali is based on visibly identifiable geological contacts, which ensure a geologically robust interpretation can be developed.

Kibali has a QA/QC programme in place to ensure the accuracy and precision of the assay results from the analytical laboratory. Checks conducted on the quality control database indicated that the results are of acceptable precision and accuracy for use in Mineral Resource estimation.

Mineral Resources

Geological models and subsequent Mineral Resource estimations have evolved and improved with each successive model update from added data within both the open pit and underground. Significant grade control drill programmes, and mapping of exposures in mine developments have been completed to increase the confidence in the resulting Mineral Resources and Mineral Reserves.

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deemed to reflect best practice by external project audits. The QP considers the Mineral Resources at Kibali as appropriately estimated and classified.

The strategic focus of Kibali exploration is to prioritise higher grade underground resource definition targets, particularly with down plunge extension drilling at depth, thereby increasing years of production with complimentary underground and open pit sources and filling a gap at the end of the LOM.

Mining and Mineral Reserves

The open pit mining operations at Kibali consists of multiple open pits. The open pits are being operated by KMS mining contractor and a down-the-hole blasting service is provided by an appropriate blasting contractor. Opportunities exist within the current pits with the Inferred Mineral Resource for upgrade and conversion to Mineral Reserves. The end of the current open pit mine life is estimated at year 2033 based on current Mineral Reserves.

The KCD underground mine is designed to extract the KCD deposit directly beneath the KCD pit. A 50 m crown pillar separates the pit bottom from the top of the underground mine. The underground mine is a long hole stoping operation planned to produce ore at a rate of 3.6 Mtpa to 3.8 Mtpa for 10 years, tapering off to 3.3 Mtpa in year 11 and 2.5 Mtpa for the last two years. Most of the underground mine infrastructure is already in place. A vertical production shaft was fully commissioned during 2018. Most ore is currently hoisted up the shaft, however, throughout the underground LOM the decline to surface is being used to haul ore from some of the shallower zones and to supplement the shaft haulage. The schedule will be progressively optimised as the underground and open pit continue to convert and update down plunge extensions and new deposits.

Barrick, as the owner operator of the Project, has significant experience in other mining operations within Africa and these production rates, modifying factors, and costs are benchmarked against other African operations to ensure they are suitable.

The current Mineral Reserves for Kibali support a total mine life of 13 years, twelve years of open pit operations, and thirteen years of underground mining. LOM gold production averages approximately 730 koz Au per year for 10 years based only on Mineral Reserves.

The QP considers the modelled recoveries for all ore sources and combined process and plant engineering unit costs, used within the Mineral Resource and Mineral Reserve process to be acceptable.

The QP is not aware of any environmental, legal, title, socioeconomic, marketing, mining, metallurgical, infrastructure, permitting, fiscal, or other relevant factors that could materially affect the Mineral Reserve estimate.

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Processing

Extensive metallurgical test work campaigns have been completed across all mineral deposits in Kibali that form the Mineral Reserve. These have consistently demonstrated two distinct behavioural patterns, the first of which exhibits free-milling characteristics suitable for gold extraction by a conventional CIL metallurgical process, and the second of which exhibits a degree of refractoriness, where straight cyanidation returns gold dissolutions considered to be too low for optimal plant operation due to the presence of occluded gold particles within sulphide minerals. It has been demonstrated that a finer grind will expose a portion of this additional gold for leaching so that the recovery is enhanced to economically acceptable levels.

The Kibali process plant operational risks are materially reduced as a function of the two separate process streams and independent milling circuits. The process plant has demonstrated excellent improvements in throughput capability, even performing beyond design capacity at 7.2 Mtpa at consistent recovery performance.

The ore feed plan is blended using both KCD underground ore plus ore sourced from satellite open pits at Kibali in order to provide a stable feed grade blend. The Kibali feed plan utilises geometallurgical models that estimate the arsenic content within arsenic bearing deposits, such that any ore with high arsenic content is stockpiled separately and blended into the CIL process route to ensure minimal impact on recovery and reagents consumption.

The QP considers the modelled recoveries for all ore sources and the process and plant engineering unit costs applied to the Mineral Resource and Mineral Reserve process to be acceptable.

Infrastructure

Kibali is a mature operation that has all necessary support infrastructure already in place.

For purposes of reducing Kibali’s reliance on thermal generation and reducing the mine operating costs, three hydropower stations with a combined potential capacity of 42.8 MW of hydropower (at peak) and has backup installed capacity for 43 MW of thermal generation. The load demand of the mine is not constant, power demand at full production is currently between 39 MW and 43 MW, averaging approximately 41 MW.

Environment and Social Aspects

Three ESIAs, and two ESIA updates have been completed at the project since 2010. The ESIAs and associated Environmental and Social Management Plans (ESMP) have been consolidated and incorporated into the ESIA updates which occur every five years in accordance with the DRC Mining Regulations (2018). The most recent ESIA update was completed in 2020 in compliance with both DRC national legislation and IFC PS. Kibali’s EMS is ISO14001:2015 certified. The ESIA, ESMP and EMS considers all current and proposed activities, as well as rehabilitation and closure planning requirements.

All permits are in place and an Environmental Adjustment Plan has been approved by the DPEM.

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The mine prioritises local employment and in 2021, the employees were made up of 88% Congolese nationals; more than 70% from the local area. More than 70% of management positions were held by Congolese Nationals.

Stakeholder engagement is ongoing, and all senior management are involved in regular meetings with the community.

Three significant resettlement campaigns have taken place, one in 2012/2013, one in 2016/2017 (Gorumbwa), and the Pamao-Kalimva-Ikamva RAP is ongoing. Ongoing monitoring of affected households to ensure that their livelihoods, often previously based on artisanal mining, are not adversely affected by the resettlement, is undertaken. Economic displacement has also been significant across the area.

ASM remains a concern in the Kibali Exploitation Permit area and the mine is working with provincial authorities to prevent and relocate artisanal miners within the permit area.

Kibali Goldmines continues to invest in community development initiatives, focussing on potable water supplies, primary school education, health care education, investment in medical clinics and local economic development projects.

The QP considers the extent of all environmental liabilities, to which the property is subject, to have been appropriately met.

Risks

Kibali Goldmines has undertaken an analysis of the Project risks. Table 1-8 summarises the Project risks and the QP’s assessment of the risk degrees and consequences, as well as ongoing/required mitigation measures. The QPs, however, note that the degree of risk refers to our subjective assessment as to how the identified risk could affect the achievement of the Project objectives. Kibali has been in production since 2013 and is a mature operation

In the QP’s opinion, there are no significant risks and uncertainties that could reasonably be expected to affect the reliability or confidence in the exploration information, Mineral Resource or Mineral Reserve estimates.

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Risk Analysis Definitions

The following definitions have been employed by the QPs in assigning risk factors to the various aspects and components of the Project:

●

Low – Risks that are considered to be average or typical for a deposit of this nature and could have a relatively insignificant impact on the economics. These generally can be mitigated by normal management processes combined with minor cost adjustments or schedule allowances.

●

Minor – Risks that have a measurable impact on the quality of the estimate but not sufficient to have a significant impact on the economics. These generally can be mitigated by normal management processes combined with minor cost adjustments or schedule allowances.

●

Moderate – Risks that are considered to be average or typical for a deposit of this nature but could have a more significant impact on the economics. These risks are generally recognisable and, through good planning and technical practices, can be minimised so that the impact on the deposit or its economics is manageable.

●

Major – Risks that have a definite, significant, and measurable impact on the economics. This may include basic errors or substandard quality in the basis of estimate studies or project definition. These risks can be mitigated through further study and expenditure that may be significant. Included in this category may be environmental/social non-compliance, particularly regarding Equator Principles and IFC PS.

●

High – Risks that are largely uncontrollable, unpredictable, unusual, or are considered not to be typical for a deposit of a particular type. Good technical practices and quality planning are no guarantee of successful exploitation. These risks can have a major impact on the economics of the deposit including significant disruption of schedule, significant cost increases, and degradation of physical performance. These risks cannot likely be mitigated through further study or expenditure.

In addition to assigning risk factors, the QPs provided opinion on the probability of the risk occurring during the LOM. The following definitions have been employed by the QPs in assigning probability of the risk occurring:

●

Rare – The risk is very unlikely to occur during the Project life.

●

Unlikely – The risk is more likely not to occur than occur during the Project life.

●

Possible – There is an increased probability that the risk will occur during the Project life.

●

Likely – The risk is likely to occur during the Project life.

●

Almost Certain – The risk is expected to occur during the Project life.

Risk Analysis Table

Table 1-8 details the Kibali Risk Analysis as determined by the QPs.

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Table 1-8 Kibali Risk Analysis



Issue	Likelihood	Consequence Rating	Risk Rating	Mitigation

Geology and Mineral Resources – Confidence in Mineral Resource Models	Unlikely	Minor	Low	Additional scheduled infill drilling. Resource model updated on a regular basis using production reconciliation results.

Mining and Mineral Reserves – Open Pit Slope Stability	Unlikely	Moderate	Minor	Continued in-pit monitoring, geotechnical drilling, instrumentation, and continued updating of geotechnical and hydrology models.

Mining and Mineral Reserves – Underground Recovery and Dilution	Possible	Moderate	Low	Change in drilling and blasting practices and paste filling binder to reduce dilution and increase recovery.

Processing - Salts build up in the process water – leading to carbon fouling in the CIL and elution circuits	Possible	Moderate	Low	A full salt and water balance has been completed and tracked in the plant to ensure that correct water dilution into the critical streams of elution is managed with minimum impact on carbon fouling and gold recovery.

Environmental -Groundwater contamination (arsenic) -Tailings failure and Waste Rock	Possible	Major	Low	Manage arsenic levels through feed profile and capture run-off. All high arsenic feed reports to lined tailings facility. Encapsulate and rehabilitate waste dumps. Continuing monitoring and external or third-party audits.

Social – SLTO	Possible	Moderate	Moderate	Dedicated community engagement by company social and sustainability department. Accessible Grievance Mechanism

Country & Political – Security – Governmental	Possible	Major	Moderate	Dedicated government liaison team in Kinshasa. Government participation/ownership.

Capital and Operating Costs	Unlikely	Moderate	Low	Continue to track actual costs and LOM forecast costs, including considerations for inflation and foreign exchange.

Fiscal Stability	Possible	Moderate	Moderate	Dedicated government liaison team in Kinshasa Government participation/ownership

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1.15

Recommendations

The QPs make the following recommendations:

Geology and Mineral Resources

●

Investigate the potential to transfer from explicit strings to implicit lithological modelling.

●

Improve on modelling of barren centimetric quartz felsic porphyry (QSF) units to enable separate volume assignment and estimation, especially underground in the up plunge 3000 lodes and across the 9000 lodes. Use long sections to improve continuity of QSF, particularly in the 9000 lodes.

●

Pamao South domaining generates poor grade distributions and is challenging to link up mineralised wireframes from one section to the next. Extra geological observation and adjustments to short term grade control (GC) models will be required in 2022 with added GC drilling.

●

Identify and refine ‘higher grade risk stopes’ by use of bespoke drilling and follow up mapping to ensure vertical and lateral high-grade lode terminations or other areas with edge uncertainty are modelled robustly to reduce short term variability, particularly in the 3000 and 9000 lodes.

●

Re-establish the regular use of blast movement monitoring in the open pits (OrePro3D) to adjust dig polygons and reduce dilution.

●

As identified in the 2021 Mineral Resource and Mineral Reserve audit by RSC consultants, an update of SOPs is required. Updating SOPs will be assigned to all senior geologists as part of team KPIs. Consider using the database for document control.

●

Address all outstanding recommendations from the RSC independent audit, as outlined in 14.16 (External Resource Audits).

●

Address each low-risk recommendation and value add comment from the RSC independent audit and collate findings in a presentation and review results at year end 2022.

Mining and Mineral Reserves

●

Improve the drill and blast practices by introducing a wireless blasting initiation system. Implementing a wireless blasting technology will help in optimizing the firing sequences, improving the ore recovery, and reducing the dilution.

●

Monitor the 90/10 slag cement binder performance over time and ensure that witness QA/QC samples are kept beyond the normal testing period and break at the time of drilling the adjacent stope.

●

Integrate paste filling within the existing stope closure note process to ensure that a thorough analysis of the filling performance during and upon completion of the filling is undertaken for each stope.

●

Due to the high variability in the 9101 lode stope, ensure that a ratio of no more than 30% to 35% of the 9101 Lode is mined and fed at any given time.

●

Implement bigger chamfer for mining the 9101 and 9101 upper transverse advancing face stope in order to optimize the bogging and improve the stope recovery.

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Processing

●

Implementation of the cyanide recovery process must be completed to realise process efficiencies in cyanide consumption.

●

Continuous process improvement and geometallurgical work on new satellite orebodies must remain in place to ensure that the plant performance remains optimal for both sulphide and free milling ores.

Infrastructure

●

Further decrease the mine’s reliance on thermal power, increase grid stability, and potentially reduce operating costs in dry season, by increasing current battery storage capacity integration with the current power model and commence a feasibility study on Solar Power.

Environment and Social Aspects

●

An ASM cessation strategy should be agreed with the Haut Uélé governor so that the local community and local chiefs are sensitised to the importance of limiting ASM activities within the government identified ‘corridors’.

●

Continued stakeholder engagement and re-enforcement of the accessibility of the grievance mechanism.

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2	Introduction

This Technical Report on the Kibali Gold Mine, located in the DRC, has been prepared by Barrick on behalf of Kibali Goldmines. The purpose of this Technical Report is to support public disclosure of Mineral Resource and Mineral Reserve estimates at the Mine as of 31 December 2021.

Kibali Goldmines is an exploration and mining company, which is currently owned 45% by Barrick and 45% by AngloGold. The remaining 10% interest in Kibali Goldmines is held by Congolese parastatal SOKIMO with the shareholding held by the MoP of the DRC.

The Project consists of multiple mineral deposits including; an underground mine at KCD; active open pits at KCD, Sessenge, Aerodrome and Gorumbwa; a partially depleted pit with planned push backs at Pakaka; depleted pits with further potential at Kombokolo and Mengu Hill; planned new pits at Pamao, Megi-Marakeke-Sayi, Kalimva, Ikamva, and Oere, plus deposits under evaluation at Mengu Village and Sessenge SW; a processing plant (7.2 Mtpa design capacity), three hydropower stations, together with other associated mine operation and regional exploration infrastructure. The Kibali plant produces gold doré bars.

Total mine production from both Kibali underground and open pits in 2021 was 7.8 million tonnes (Mt) at a head grade of 3.62 g/t Au for a total of 812 thousand ounces (koz) Au (89.8% recovery).

Total production since mining commenced in 2013 to end of year (EOY) 2021 is 59 Mt milled at a head grade of 3.48 g/t Au for 5.7 Moz Au (85.7% recovery).

The Mineral Resource and Mineral Reserve estimates have been prepared according to the Canadian Institute of Mining, Metallurgy and Petroleum CIM (2014) Standards as incorporated by reference in NI 43-101. Mineral Resource and Mineral Reserve estimates were also prepared using the guidance outlined in CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines 2019 (CIM (2019) MRMR Best Practice Guidelines).

2.1

Effective Date

The effective date of this Technical Report is 31December 2021.

2.2

Qualified Persons

This Technical Report was prepared by Barrick on behalf of Kibali Goldmines and incorporates the work of Digby Wells and Associates Pty Ltd. (Digby Wells).

The Qualified Persons (QPs) and their responsibilities for this Technical Report are listed in Section 29 Certificates of Qualified Persons and summarised in Table 2-1.

The documentation reviewed, and other sources of information, are listed at the end of this report in Section 27 References.

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Table 2-1 QP Responsibilities



Qualified Person	Company	Title/Position	Sections

Rodney B. Quick, MSc, Pr. Sci.Nat	Barrick Gold Corporation	Mineral Resource Manager and Evaluation Executive	1.1, 1.2, 1.3, 1.10, 2, 4 to 6, 19, and 23

Simon P. Bottoms, CGeol, MGeol, FGS, FAusIMM	Barrick Gold Corporation	Senior Vice President, Africa and Middle East, Mineral Resource Manager	1.4, 1.12, 1.13, 1.15¹, 3, 7 to 9, 21, 22, 24, and 26.1¹

Christopher B. Hobbs, CGeol, MSc, MCSM, FAusIMM	Barrick Gold Corporation	Group Resource Geologist	1.5, 1.14², 1.15², 10 to 12, 14, 25.1, and 26.1²

Graham E. Trusler, MSc, Pr Eng, MIChE, MSAIChE	Digby Wells and Associates Pty Ltd.	CEO	1.11, 1.14⁷, 1.15⁷, 20, 25.5, and 26.5

Thamsanqa Mahlangu, Pr. Eng, PhD	Barrick Gold Corporation	Head of Metallurgy, Africa and Middle East	1.8, 1.9, 1.14⁵, 1.14⁶, 1.15⁵, 1.15⁶, 13, 17, 18, 25.3, 25.4, 26.3, and 26.4

Shaun Gillespie, Reg Eng Tech, FAusIMM	Barrick Gold Corporation	Group Planning Manager, Africa and Middle East	1.6³, 1.7³, 1.14³, 1.15³, 15.1³ to 15.3³, 15.4, 15.6³ to 15.8³, 16.1³, 16.2, 16.6³, 25.2³, and 26.2³

Ismail Traore, MSc, FAusIMM, M.B. Law, DES	Barrick Gold Corporation	Group Underground Planning Manager, Africa and Middle East	1.6⁴, 1.7⁴, 1.14⁴, 1.15⁴, 15.1⁴ to 15.3⁴, 15.5, 15.6⁴ to 15.8⁴, 16.1⁴, 16.3 to 16.5, 16.6⁴, 25.2⁴, and 26.2⁴

All			1.14 (Risks), 25.6, and 27

Notes:

Geology

Mineral Resources

Mining and Mineral Reserves - Open Pit and Stockpiles

Mining and Mineral Reserves – Underground

Processing

Infrastructure

Environment and Social Aspects

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2.3

Site Visit of Qualified Persons

Below are the most recent site visit dates for the QPs:

●

Mr. Rodney B. Quick – 06 to 09 July 2021. Mr Quick has made two separate visits in 2021 upon which he reviewed the exploration programme results, Mineral Resource and grade control model updates, mine plans, mining performance results, mine strategy, results of external audits, and board meeting reviews.

●

Mr. Simon Bottoms – 07 to 09 October 2021. Mr Bottoms has made four separate visits in 2021 upon which he reviewed the exploration programme results, Mineral Resource and grade control model updates, mine plans, mining performance results, mine strategy, results of external audits, and board meeting reviews.

●

Mr. Christopher Hobbs – 07 to 09 October 2021. Mr Hobbs has made four separate visits in 2021 upon which he reviewed the exploration programme results, Mineral Resource and grade control model updates, mine strategy, results of external audits, and board meeting reviews.

●

Mr. Graham E. Trusler – 19 to 23 July 2021. Mr Trusler made a visit to Kibali from 19 to 23 July 2021 when he visited all major establishments within the mining area including the mining pits, tailings dams, water dams, some community projects, and the resettlement sites near to the mine. Reviews were held with management teams from the social, safety, and environmental departments.

●

Dr Thamsanqa Mahlangu – 07 to 09 October 2021. Dr Mahlangu has made four separate visits in 2021 upon which he reviewed the processing plant operations performance, and geometallurgical test work on new and current deposits. Also covered were reviews on the process improvement projects and board meeting reviews.

●

Mr. Shaun Gillespie – 07 to 09 October 2021. Mr Gillespie has made four separate visits in 2021 upon which he reviewed mining performance results, Mineral Reserve and grade control model updates, mine strategy, results of external audits, and board meeting reviews.

●

Mr. Ismail Traore – 15 to 20 July 2021. Mr Traore has made two separate visits in 2021 upon which he reviewed mining performance results, Mineral Reserve and grade control model updates, mine strategy, results of external audits, and board meeting reviews.

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2.4

List of Abbreviations

Units of measurement used in this Technical Report conform to the metric system. All currency in this Technical Report is in US dollars (US$ or $) unless otherwise noted.


Mm	micron	kW	kilowatt

°C	degree Celsius	kWh	kilowatt-hour

°F	degree Fahrenheit	L	litre

mg	microgram	L/s	litres per second

A	ampere	m	metre

a	annum	M	mega (million)

bbl	barrels	m²	square metre

Btu	British thermal units	m³	cubic metre

C$	Canadian dollars	Ma	million years ago

cal	calorie	min	minute

cfm	cubic feet per minute	MASL	metres above sea level

cm	centimetre	mm	millimetre

cm²	square centimetre	mph	miles per hour

d	day	MVA	megavolt-amperes

dia.	Diameter	MW	megawatt

dmt	dry metric tonne	MWh	megawatt-hour

dwt	dead-weight ton	m³/h	cubic metres per hour

ft	foot	opt, oz/st	ounces per short ton

ft/s	feet per second	oz	Troy ounce (31.10348 g)

ft²	square foot	koz	thousand ounces

ft³	cubic foot	Moz	million ounces

g	gram	ppm	parts per million

G	giga (billion)	psia	pounds per square inch absolute

Gal	Imperial gallon	psig	pounds per square inch gauge

g/L	grams per litre	RL	relative elevation

g/t	grams per tonne	s	second

gpm	Imperial gallons per minute	st	short ton

gr/ft³	grains per cubic foot	stpa	short tons per annum

gr/m³	grains per cubic metre	stpd	short tons per day

hr	hour	t	metric tonne

ha	hectare	kt	thousand metric tonnes

hp	horsepower	Mt	million metric tonnes

in	inch	tpa	metric tonnes per annum

in²	square inch	tpd	metric tonnes per day

J	joule	Mtpa	million metric tonnes per annum

k	kilo (thousand)	US$	United States dollar

kcal	kilocalorie	Usg	United States gallon

kg	kilogram	Usgpm	US gallon per minute

km	kilometre	V	volt

km/h	kilometres per hour	W	watt

km²	square kilometre	wmt	wet metric tonne

kPa	kilopascal	yd³	cubic yard

kVA	kilovolt-amperes	yr	year

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3	Reliance on Other Experts

This report has been prepared by Barrick. The information, conclusions, opinions, and estimates contained herein are based on:

●

Information available at the time of preparation of this Technical Report,

●

Assumptions, conditions, and qualifications as set forth in this Technical Report.

For the purpose of this report, the QPs have relied upon information provided by Barrick’s legal counsel regarding the validity of the Exploitation Permits and the changes to the fiscal regime outlined in the DRC Mining Code (2018) as part of ongoing annual reviews. This opinion has been relied upon in Section 4 (Property Description and Location) and in the summary of this report.

Except for the purposes legislated under provincial securities laws, any use of this Technical Report by any third party is at that party’s sole risk.

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4	Property Description and Location

4.1

Project Location

Kibali is located in the NE of the DRC in the Haut Uélé Province, approximately 1,800 km NE of the capital city Kinshasa, approximately 560 km NE of the capital of the Orientale Province, Kisangani, 1,800 km from the Kenyan port of Mombasa, 1,950 km from the Tanzanian port of Dar es Salaam, and 150 km west of the Ugandan border town of Arua, near the international borders with Uganda and Sudan.

The location of the Project is shown in Figure 4-1.

The Project, which covers an area of approximately 1,836 km², is centred at approximately 3.13º latitude and 29.58° longitude, in the Haut Uélé Province.

The Project consists of multiple mineral deposits including an underground mine at KCD; active pits at KCD, Sessenge, Aerodrome and Gorumbwa; a partially depleted pit with planned push backs at Pakaka; depleted pits with further potential at Kombokolo and Mengu Hill; planned new pits at Pamao, Megi-Marakeke-Sayi, Kalimva, Ikamva, and Oere, plus deposits under evaluation at Mengu Village and Sessenge SW.

4.2

Mineral Rights and Land Ownership

Table 4-1 provides Exploitation Permit details, Table 4-2 provides the Exploitation Permit perimeter coordinates and Figure 4-2 shows the Exploitation Permit locations. All coordinates use UTM Zone 35N datum WGS84 grid.

Table 4-1 Kibali Exploitation Permit Details



Arête No.	Permit No.	Surface Area (km²)	Expiry Year

0852/CAB.MIN/MINES/01/2009	11447	226.8	2029

0855/CAB.MIN/MINES/01/2009	11467	248.9	2029

0854/CAB.MIN/MINES/01/2009	11468	45.9	2030

0853/CAB.MIN/MINES/01/2009	11469	91.8	2029

0329/CAB.MIN/MINES/01/2009	11470	30.6	2029

0852/CAB.MIN/MINES/01/2009	11471	113.0	2029

0331/CAB.MIN/MINES/01/2009	11472	85.0	2029

0856/CAB.MIN/MINES/01/2009	5052	302.4	2029

0858/CAB.MIN/MINES/01/2009	5073	399.3	2029

0103/CAB.MIN/MINES/01/2011	5088	292.2	2030

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Table 4-2 Kibali Exploitation Permit Coordinates



Permit	Lat	Long	Easting	Northing	Permit	Lat	Long	Easting	Northing

5088	3°00’30”	29°51’00”	816830	332928	11472	2°57’00”	29°57’00”	827974	326501
	3°01’00”	29°51’00”	816828	333850		2°57’00”	29°56’30”	827047	326499
	3°01’00”	29°51’30”	817755	333853		2°58’00”	29°56’30”	827042	328344
	3°04’00”	29°51’30”	817741	339386		2°58’00”	29°56’00”	826115	328341
	3°04’00”	29°51’00”	816813	339384		2°58’30”	29°56’00”	826112	329263
	3°10’00”	29°51’00”	816783	350451		2°58’30”	29°54’00”	822403	329254
	3°10’00”	29°57’30”	828835	350485		2°59’00”	29°54’00”	822401	330176
	3°09’30”	29°57’30”	828838	349562		2°59’00”	29°53’30”	821474	330173
	3°09’30”	29°59’30”	832547	349573		3°00’00”	29°53’30”	821469	332018
	3°10’00”	29°59’30”	832544	350495		3°00’00”	30°01’00”	168332	332045
	3°10’00”	30°00’00”	166529	350498		2°55’00”	30°01’00”	168307	322822
	3°00’00”	30°00’00”	166477	332051	11471	3°10’00”	29°31’30”	780634	350357
	3°00’00”	29°53’30”	821469	332018		3°10’00”	29°28’00”	774147	350342
	3°00’30”	29°53’30”	821466	332940		3°19’30”	29°28’00”	774104	367859
5052	3°10’00”	29°31’30”	780634	350357		3°19’30”	29°31’30”	780590	367875
	3°19’30”	29°31’30”	780590	367875	11470	3°00’00”	30°00’00”	166477	332051
	3°19’30”	29°32’00”	781517	367878		3°09’00”	30°00’00”	166524	348653
	3°18’30”	29°32’00”	781522	366034		3°09’00”	30°01’00”	168378	348648
	3°18’30”	29°32’30”	782448	366036		3°00’00”	30°01’00”	168332	332045
	3°17’30”	29°32’30”	782453	364192	11469	2°55’30”	29°37’00”	790894	323642
	3°17’30”	29°33’00”	783380	364194		2°55’30”	29°31’00”	779770	323617
	3°16’00”	29°33’00”	783387	361428		3°00’00”	29°31’00”	779751	331915
	3°16’00”	29°36’00”	788947	361443		3°00’00”	29°37’00”	790874	331941
	3°16’30”	29°36’00”	788945	362365	11468	2°55’30”	29°31’00”	779770	323617
	3°16’30”	29°36’30”	789872	362367		2°55’30”	29°28’00”	774208	323605
	3°17’00”	29°36’30”	789869	363289		3°00’00”	29°28’00”	774189	331902
	3°17’00”	29°37’30”	791723	363294		3°00’00”	29°31’00”	779751	331915
	3°16’30”	29°37’30”	791725	362372	11467	3°10’00”	29°35’00”	787122	350373
	3°16’30”	29°38’30”	793579	362377		3°06’30”	29°35’00”	787138	343919
	3°17’00”	29°38’30”	793576	363299		3°06’30”	29°35’30”	788065	343921
	3°17’00”	29°39’00”	794503	363301		3°00’00”	29°35’30”	788093	331934
	3°18’00”	29°39’00”	794498	365145		3°00’00”	29°28’00”	774189	331902
	3°18’00”	29°40’00”	796352	365150		3°10’00”	29°28’00”	774147	350342
	3°19’30”	29°40’00”	796344	367917	11447	3°00’00”	29°37’00”	790874	331941
	3°19’30”	29°42’30”	800978	367929		3°00’00”	29°35’30”	788093	331934
	3°18’30”	29°42’30”	800983	366085		3°06’30”	29°35’30”	788065	343921
	3°18’30”	29°44’30”	804690	366095		3°06’30”	29°35’00”	787138	343919
	3°19’00”	29°44’30”	804688	367017		3°10’00”	29°35’00”	787122	350373
	3°19’00”	29°45’30”	806541	367023		3°10’00”	29°40’00”	796390	350397
	3°19’30”	29°45’30”	806539	367945		3°15’30”	29°40’00”	796364	360540
	3°19’30”	29°47’00”	809319	367953		3°15’30”	29°44’00”	803778	360560
	3°19’00”	29°47’00”	809322	367030		3°15’00”	29°44’00”	803781	359638
	3°19’00”	29°48’30”	812102	367038		3°15’00”	29°46’00”	807488	359648
	3°18’00”	29°48’30”	812108	365194		3°14’30”	29°46’00”	807491	358726
	3°18’00”	29°48’00”	811181	365191		3°14’30”	29°45’30”	806564	358723
	3°17’00”	29°48’00”	811186	363347		3°14’00”	29°45’30”	806567	357801
	3°17’00”	29°47’30”	810259	363344		3°14’00”	29°44’00”	803786	357793

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Permit	Lat	Long	Easting	Northing	Permit	Lat	Long	Easting	Northing

	3°16’00”	29°47’30”	810264	361500		3°10’30”	29°44’00”	803803	351338
	3°16’00”	29°47’00”	809337	361497		3°10’30”	29°44’30”	804730	351341
	3°15’00”	29°47’00”	809342	359653		3°10’00”	29°44’30”	804733	350418
	3°15’00”	29°44’30”	804708	359640		3°10’00”	29°45’00”	805660	350421
	3°15’30”	29°44’30”	804705	360562		3°09’30”	29°45’00”	805662	349499
	3°15’30”	29°40’00”	796364	360540		3°09’30”	29°46’00”	807516	349504
	3°10’00”	29°40’00”	796390	350397		3°08’00”	29°46’00”	807523	346737
11472	2°55’00”	29°58’30”	830766	322819		3°08’00”	29°47’00”	809377	346742
	2°55’30”	29°58’30”	830764	323742		3°04’00”	29°47’00”	809397	339364
	2°55’30”	29°57’30”	828909	323737		3°04’00”	29°46’00”	807543	339360
	2°56’00”	29°57’30”	828906	324659		3°07’30”	29°46’00”	807526	345815
	2°56’00”	29°57’00”	827979	324657		3°07’30”	29°37’00”	790841	345772

All Mineral Resources and Mineral Reserves summarised in this report are contained within these Exploitation Permits. The Exploitation Permits occur within two territories, namely Watsa and Faradje, which fall under the administrative district of Haut Uélé.

The principal mineral deposit, KCD, forms both an open pit and underground mine. This operation and the associated infrastructure (processing plant, accommodation, and airport) are within Exploitation Permits 11447 and 11467 (Table 4-1 and Figure 4-2).

In the QP’s opinion, all appropriate Exploitation Permits have been acquired and obtained to conduct the work proposed for the property.

The next renewal dates for the Exploitation Permits are 05 November 2029 and 06 March 2030 and the current LOM plan for the Kibali Mineral Reserves extends beyond these dates. The DRC Mining Code (2002) includes provision for renewal of all Exploitation Permits for a successive period of 15 years, providing the holder has not breached the obligations of permit fee and annual surface rights fee payments, and upholds all environmental standards set out in the Exploitation Permit. Furthermore, the Exploitation Permit holder should provide the appropriate government departments with a monthly mining activity report and quarterly exploration reports.

All the Exploitation Permit fees and taxes relating to Kibali Goldmines’ exploitation rights have been paid to date and the concession is in good standing.

The QPs are not aware of any risks that could result in the loss of ownership of the deposits or loss of the Exploitation Permits, in part or in whole.

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4.3

Surface Rights

Surface rights in the area of the Kibali Permits belong to the DRC Government. Utilisation of the surface rights is granted by the Kibali Exploitation Permits under the condition that the current users are properly compensated. All the surface rights fees relating to Kibali’s exploitation rights have been paid to date and the concession is in good standing.

One exclusion zone with an area of 10.26 km² exists within the Permit surrounding the Kibali South deposit, which was transferred to SOKIMO from Kibali Goldmines in December 2012. The coordinates of the limits of this exclusion zone are presented in Table 4-3.

Table 4-3 Coordinates of the Exclusion Zone in the Kibali Exploitation Permit



ID	Lat	Long	Easting	Northing
A	03°05’00”	29°35’30”	788071	341155
B	03°05’00”	29°34’00”	785291	341148
C	03°06’00”	29°34’00”	785286	342992
D	03°06’00”	29°32’30”	782506	342986
E	03°05’30”	29°32’30”	782508	342064
F	03°05’30”	29°33’30”	784361	342068
G	03°04’00”	29°33’30”	784368	339302
H	03°04’00”	29°35’30”	788076	339311

The QPs are not aware of any other significant factors and risks that may affect access, title, or the right or ability to perform work on the property.

4.4

Ownership, Royalties and Lease Obligations

Barrick is the operator at Kibali for both exploration and mining.

The DRC Mining Code (2002) and associated regulations have been amended with the DRC Mining Code (2018), which came into force on 09 March 2018, and the related amended mining regulations, which came into force on 08 June 2018.

The following changes made to the DRC Mining Code (2002) in 2018 introduced a series of changes at Kibali: (i) royalty charges were increased from 3.5% to 4.7%, which is not anticipated to materially impact the LOM profitability; (ii) various increases in import and other duties from 4% to 7% depending on consumable type, which is not anticipated to materially alter the LOM profitability, and (iii) a super-tax profit has been promulgated based on the feasibility study prepared at the time the approval was given for the construction of the Project and accordingly, such a tax is applicable only if the average annual gold price was in excess of $2,000/oz Au. No other parties own a royalty interest other than the DRC government.

The QPs are not aware of any risks that could result in the loss of ownership of the deposits or loss of the Permits, in part or in whole.

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5	Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1

Accessibility

Kibali is located in the NE of the DRC, approximately 560 km NE of the city of Kisangani and near the international borders with Uganda and South Sudan.

5.2

Climate and Physiography

The DRC has a total area of 2.3 million km². The country straddles the equator and is characterised by dense tropical rain forest in the central Congo River basin and highlands in the east.

The climate is tropical – hot and humid in the equatorial river basin; cooler and drier in the southern highlands; cooler and wetter in the eastern highlands where Kibali is located.

The Watsa territory wet season occurs between March to November, with the dry season occurring between November and March (Figure 5-1). Watsa experiences extreme seasonal variation in monthly rainfall with most rain occurring in heavy tropical thunderstorms. Precipitation is highest in October, and January and December are the driest months. Humidity levels are highest in the wet season.

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Source: Kibali Goldmines, 2021

Note:

Data collected from 2012 to 2021.

Figure 5-1 Kibali Average Monthly Rainfall Statistics

The Watsa territory dry season lasts from January to March, with average daily high temperatures above 30°C and average daily low temperatures of approximately 19°C. The cool season occurs between May and November, with average daily high temperatures below 29°C and average daily low temperatures of approximately 18°C.

The average wind speed experiences mild seasonal variation over the course of the year, generally averaging 8.0 km/h in the wet season and 6.5 km/h in the dry season.

Climatic conditions do not materially affect either exploration, development, or mining operations allowing these activities to be conducted year-round.

The topography of the area is gently hilly, ranging in elevation between 700 m to 1,500 m above sea level (MASL). The immediate Project area is characterised as generally hilly, which includes several discrete hills up to 170 m high. The plant site is located on a flat plain area which lies at approximately 860 MASL. The Project lies in a low seismic rated area.

Vegetation is dominated by elephant grass with forested areas along drainages. It is likely that the entire area comprised rainforest prior to modification by human activity.

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5.3

Infrastructure

The local Project area lacks any substantial infrastructure to support the mining operation, other than that which has been constructed by Kibali Goldmines. All existing infrastructure supports the local subsistence and small-scale agriculture.

Remnants of historical mining activities can be found on the property (residential buildings, processing plant, underground mine shafts, and surface workings) in various states of repair. Although remnants of the historical mining activities remain, the mine is essentially a greenfield development, with new facilities having been built to support the current mining and processing activities, because the current mine is of a much larger scale than any of the historical mining infrastructure.

The key on-site surface and underground infrastructure at Kibali include the following:

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Mine access and internal road network.

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A 7.2 Mtpa process plant.

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TSFs comprising of concentrate tailings storage facility one, (CTSF1) and concentrate tailings storage facility two (CTSF2) for the CIL tails and the FTSF is dedicated to flotation tails.

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Accommodation village for married and single staff and employees.

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Administrative buildings, stores warehouses, laboratory, workshops for surface and underground equipment, security buildings, medical and emergency response facilities.

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Fuel Storage.

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Raw and process water containment and storage dams and water distribution network.

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Communications and data transmission networks.

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Airstrip.

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Twin declines and vertical production shaft and a series of ramp-connected levels.

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Diesel generator station installed with CAT 3516B-HD (1.5 MW) generators.

As there is no national grid power supply to the site. As a result, Kibali is fully dependent on its own generation facilities.

The power supply currently comes from a mix of on-site, high-speed diesel generator sets and off-site hydropower stations; Nzoro II is currently producing approximately 22 MW, Ambarau produces 10.6 MW and Azambi produces a further 10.2 MW, with total peak hydropower capacity of 42.8 MW, which is sufficient to meet the mine power demand. A battery energy storage system was incorporated in 2020 to improve power stability. Nzoro 2, Ambarau and Azambi hydropower plants. The site is connected to the hydrostations via a 66 kV overhead line network.

The hydropower system has combined potential capacity of 42.8 MW of hydropower (at peak) and has backup installed capacity for 43 MW of thermal generation. The load demand of the mine is not constant, power demand at full production is currently between 39 MW and 43 MW, averaging approximately 41 MW.

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Figure 5-2 shows the location of the underground and open pit mines, open pit designs and associated infrastructure with respect to the Mineral Reserves within Kibali Exploitation Permits.

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5.4

Local Resources

The mine office of Kibali Goldmines is located in the village of Doko, which is centrally located within the Project area and approximately 180 km by road from Arua on the Ugandan border. The district capital of Watsa lies approximately 9 km to the south of the Project, which is situated just north of the Kibali River on the road to Faradje and the Sudan. The town of Bunia, which is the United Nations controlled entry point to NE DRC, lies approximately 200 km to the south of the Mine.

The population in the Project area is approximately 65,000. The Watsa territory population is approximately 300,000. Figure 5-3 presents a plan view of the Kibali deposits and surrounding communities.

There are generally limited services in the Project area that are suitable to directly support the Mine, as such, Kibali Goldmines has completed extensive infrastructure upgrades during construction.

As per Barrick’s strategy, the Mine continues to focus on host country employment and skills transfer, steadily increasing the Congolese component to approximately 92% of the full time Kibali manpower. Congolese contractors are also utilised for construction projects. The Mine has assisted, and continues to assist, local start-up businesses.

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6	History

6.1

Ownership

The first documented discovery of gold in the NE of the DRC is attributed to Hannan and O’Brien in 1903, who were sent by Kind Leopold and found nationals washing alluvial gravels for gold. Historical gold production from the Kilo and Moto areas between 1906 and 2009 is estimated to be approximately 11 Moz Au, half of which came from alluvial deposits. Mining operations were conducted by the Belgian Government via SOKIMO, which was established in 1926. Most of the mining activity within the Project area was undertaken during the 1950s but accurate production records have been lost over the years due to civil unrest in the region. Gorumbwa, Agbarabo and Durba deposits are believed to have collectively produced more than 60% of the over 3 Moz of recorded gold production from the Moto area. The SOKIMO processing plant was located near the old Durba mine. The plant comprised crushing and ball milling circuits, followed by gravity, cyanide leach and mercury amalgamation circuits.

After independence in 1960, gold production dropped sharply as mining was mainly undertaken by artisanal workers and small-scale alluvial operations. SOKIMO changed its name to OKIMO in 1966 and was the main operator in the Project area. Sporadic underground mining was conducted in the Project area after 1960, however this is believed to be of a remnant nature and as such negligible amounts of gold were produced. Accurate production records are not available due to the civil unrest in the region during the 1980s and 1990s as described above. The DRC Governmental Entity OKIMO was later transformed back into SOKIMO in December 2010.

Davy McKee undertook a detailed assessment of the area on behalf of the Government of Zaire in 1991, with funding from the African Development Bank. This assessment included a significant amount of drilling to verify historical data.

Barrick acquired exploration rights over most of the Kilo-Moto belts in 1996 in a 70/30 joint venture with the government entity OKIMO and drilled several targets as well as completing regional and detailed soil sampling programmes. Subsequently Barrick formed a JV with AngloGold Ashanti to split equally their 70% holding of the Project.

KCD was discovered by the Barrick and AngloGold Ashanti JV in 1998 and AngloGold Ashanti became the operator of the Project. The Barrick and AngloGold Ashanti JV completed several drilling programmes, mainly concentrated at KCD and Pakaka. The Barrick and AngloGold Ashanti JV also carried out soil sampling over most of the concession area, and a regional aeromagnetic survey was completed by World Geoscience Limited (WGC). The survey was undertaken at 200 m line spacings and the data was interpreted by WGC. AngloGold Ashanti and Barrick withdrew from the Project in 1998 due to local unrest and civil war. Very little information prior to 1998 is available regarding drilling by OKIMO or the Barrick and AngloGold JV.

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Moto Goldmines Limited (Moto) acquired the available 70% stake in the Project in 2004. Moto completed a Feasibility Study in 2008 (Moto Goldmines Ltd, 2008) and an Optimised Feasibility Study in 2009 (Moto Goldmines Ltd, 2009).

In July 2009, Randgold and AngloGold entered into a 50/50 JV, which acquired Moto and its 70% ownership of the Project. In December 2009, the JV acquired an additional 20% shareholding in the Project from SOKIMO. The DRC State remained a partner in the Project through OKIMO retaining a 10% interest.

On 01 January 2019, Barrick acquired 100% of the issued and outstanding shares of Randgold (the “Merger”) and from there on, the 45% ownership of Kibali Goldmines JV was transferred to the new Barrick company created by the merger in continued partnership with both AngloGold Ashanti (retaining a 45% interest) and SOKIMO (retaining a 10% interest).

6.2

Previous Exploration

Table 6-1 presents a summary of the known historical trenches, auger, and pit exploration results at Kibali prior to Randgold’s (now Barrick’s) involvement. Historical drilling is discussed in Section 10 of this Technical Report.

Table 6-1 Summary of Historical Kibali Trenches, Auger and Pits Summary



Year	Company	Trenches		Auger		Pits		Total
		Meters	No.	Meters	No.	Meters	No.	Meters	No.

1950 to 1960	OKIMO	167	9	-	-	1,144	79	1,311	88

1980	MOTO	No Information Available

1996	Barrick – AngloGold Ashanti	No Information Available

2006 to 2007	MOTO	-	-	-	-	12	2	12	2

2008 to 2009	MOTO	-	-	260	135	-	-	260	135

Total		167	9	260	135	1,156	81	1,583	225

6.3

Previous Resource and Reserve Estimates

The following estimates are historical in nature and should not be relied upon. A QP has not completed sufficient work to classify the historical estimates as a current Mineral Resource or Mineral Reserve and Barrick is not treating the historical estimates as current Mineral Resources or Mineral Reserves. They have been superseded by the Mineral Resource and Mineral Reserve estimates presented in this Technical Report.

Table 6-2 and Table 6-3, respectively, list the significant historical Mineral Resource and Mineral Reserve estimates completed prior to Randgold’s (now Barrick’s) involvement in the Kibali Project. These Mineral Resource and Mineral Reserve estimates were completed by Moto as part of the 2008 Feasibility Study (Moto Goldmines, 2008). These historical Mineral Resources and Ore Reserves were classified and reported in accordance with the 2004 Australasian Code for Reporting of Mineral Resources and Ore Reserves (2004 JORC Code).

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Table 6-2 presents a tabulation of the Mineral Resources within the Moto Gold Project, estimated above a nominal 1.0 g/t Au cut-off grade within the interpreted mineralised domains.

Table 6-2 Moto Goldmines Ltd. Mineral Resource Estimate as of August 2008



Deposit	Indicated Mineral Resources			Inferred Mineral Resources
Deposit	Tonnes (Mt)	Grade (g/t Au)	Contained (Moz Au)	Tonnes (Mt)	Grade (g/t Au)	Contained (Moz Au)

Pakaka	16.9	2.5	1.4	-	-	-

Gorumbwa	-	-	-	8.3	5.2	1.4

Kibali	-	-	-	17	2.2	1.2

Mengu Hill	8.8	3.0	0.8	-	-	-

Mengu Village	1.2	1.9	0.07	0.08	1.4	0.004

KCD	67	3.6	7.7	74	3.4	8.1

Megi	-	-	-	4.1	2.1	0.3

Marakeke	-	-	-	2.4	1.7	0.1

Kombokolo	2.3	2.4	0.2	-	-	-

Sessenge	8.6	2.3	0.6	-	-	-

Ndala	-	-	-	0.3	4.0	0.03

Pamao	7.9	1.9	0.5	1.2	1.9	0.07

Total	112	3.1	11	107	3.3	11

Note:

Mineral Resources are reported at a nominal cut-off grade of 1.0 g/t Au

All Mineral Resource tabulations are reported inclusive of that material which is then modified to form Mineral Reserves.

Mineral Resources and Ore Reserves were classified and reported in accordance with the 2004 Australasian Code for Reporting of Mineral Resources and Ore Reserves (2004 JORC Code)

Table 6-3 presents a tabulation of the Probable Ore Reserves estimated within the Moto Gold Project, based upon the Feasibility Study pit design (Moto Goldmines, 2008). These Ore Reserves were based on a $600/oz Au price and were contained within the Indicated Mineral Resources of the Moto Gold Project.

Table 6-3 Moto Goldmines Ltd. Ore Reserve Estimate as of August 2008



Pit	Tonnes (Mt)	Grade (g/t Au)	Contained Gold (Moz Au)

KCD	14	3.6	1.6

Kombokolo	0.5	3.0	0.05

Mengu Hill	5.4	3.4	0.6

Pakaka	6.1	2.7	0.5

Pamao	1.5	2.1	0.1

Sessenge	2.9	2.5	0.2

Total	31	3.2	3.2

Note:

Mineral Resources and Ore Reserves were classified and reported in accordance with the 2004 Australasian Code for Reporting of Mineral Resources and Ore Reserves (2004 JORC Code)

Ore Reserves were reported using a gold price of $600/oz Au.

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6.4

Past Production

Since commencing mining operations in 2013 to the end of 2021, 59 Mt of ore have been mined from the various deposits at Kibali. Table 6-4 summarises the past mill production for the Project.

The historical gold production by previous operators and artisanal miners is unknown.

Table 6-4 Past Production Records for the Kibali Mine



Year	Tonnes Milled (kt)	Grade (g/t Au)	Contained Gold (oz Au)	Recovery (%)

2013	808	3.87	88,199	91.5

2014	5,546	3.81	526,627	79.0

2015	6,833	3.55	642,720	83.8

2016	7,299	3.10	586,530	79.8

2017	7,621	2.87	596,226	83.6

2018	8,218	3.45	807,251	88.6

2019	7,513	3.80	814,027	88.7

2020	7,632	3.68	808,134	89.4

2021	7,783	3.62	812,152	89.8

Total	59,254	3.48	5,681,866	85.7

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7	Geological Setting and Mineralisation

7.1

Regional Geology

The Kibali gold deposits are hosted within the Moto Greenstone Belt, a Kibalian Neoarchean terrane that lies in the NE Congo Craton. The northeastern part of the Congo craton is formed of Archean rocks, which extend eastward from the northern part of the DRC across the Cenozoic East African rift into Uganda, southern Kenya, and northern Tanzania (Allibone et al, 2020). Plutonic rocks underlie 80 to 90% of the area, volcano-sedimentary rocks are largely metamorphosed under greenschist facies conditions and form isolated belts for the remaining 10% to 20% of the craton (Figure 7-1).

The Moto Greenstone Belt is elongated, WNW-ESE trending, and is comprised primarily of two distinct litho-stratigraphically blocks. To the north, the belt is bounded by the West Nile Gneiss complex, a Meso- or Paleoarchean granite gneiss that extends northward into the Sahara Desert (U-Pb ages > 2670 Ma; Turnbull et al., 2017). To the south, the belt is bounded by the Upper Zaire Granitic Massif, an Archean granite-gneiss terrane that dominates the NE Congo Craton. The Massif is locally represented by the Watsa Igneous Complex.

The Moto Greenstone Belt contains Archean aged volcano-sedimentary conglomerate, carbonaceous shales, siltstone, BIFs, sub aerial basalts, mafic intermediate intrusions (dykes and sills) and multiple intrusive phases that range from granodiorite, tonalite and gabbroic in composition. The Kibali deposits are predominantly hosted within sedimentary lithologies that have undergone complex structural deformation and metamorphism. Metamorphic grade varies from lower greenschist facies in the west, progressively increasing to amphibolite facies in the east. Granitoid plutons as old as 2640 Ma intrude these rocks, constraining the lithologies minimum age.

Intrusive units from both the West Nile Gneiss and Moto Belt Greenstones are bimodal in geochemistry, with trace element distribution indicating formation in an island arc environment (Allibone et al, 2020). Extrusive units from both terranes show trace element signatures that are more typical of Mid Oceanic Ridge Basalts (MORB) (Allibone et al, 2020). This is supported by the presence of ‘pillow’ textured basalts in the Project area.

Regional geological interpretations suggest that the belt is a thrust stack that developed during the collision of an island arc along the northern margin of the Upper Zaire Granitic Massif with the West Nile Gneiss thrust southward over the Moto Greenstone Belt. Ductile and brittle deformation events are observed in the lithological units, with polyphase isoclinal and recumbent folding mapped in some of the deposits. The belt is cut by two principal structure sets: NW-SE striking, NE dipping thrust faults, and a series of sub-vertical NE-SW shear structures, both of which in association with the folding are considered important mineralising controls.

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7.2

Structural Geology

The Kibali gold deposits are scattered along a curvilinear zone approximately 20 km long and up to 1 km wide known as the KZ North Structure (KZS), part of the 60 km KZ Trend (Figure 7-2). Gold is concentrated in gently NE to NNE-plunging shoots whose orientations are generally parallel with a prominent lineation in the mineralised rocks. It has been concluded that the structure of the Kibali district is the product of at least seven phases of deformation. Key features of each event are listed below:

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D_1/14: ductile faults generally parallel to lithologic layering but which locally cut across lithologic layering.

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D_2/14: isoclinal recumbent folds whose axial planes dip approximately 25° to 30° NNE, axes that plunge approximately 25° NE, and an associated generally layer-parallel foliation.

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D_3/14: upright folds whose axial planes dip steeply towards either the NW or SE and axes plunge approximately 25° NE.

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D_4/14: sericite-rich spaced foliation largely restricted to altered rocks at the KCD deposit.

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D_5/14: NE-striking steeply dipping brittle faults close to parallel with axial planes of the F3/14 folds.

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D_6/14: SSW-dipping folds with near horizontal axes that trend WNW or ESE, an associated axial plane-parallel crenulation cleavage, and related contractional faults.

●

D_7/14: Minor SSW-dipping normal faults, fractures, and associated barren en-echelon quartz veins.

D_1/14 through D_4/14 are all ductile in character and each involved the formation of ductile faults, folds, penetrative foliations, and/or penetrative linear fabrics. D_2/14 and D_3/14 clearly occurred in a contractional setting, but evidence of the tectonic settings of D_1/14 and D_4/14 are more ambiguous. D_5/14 is a phase of essentially brittle faulting that was followed by a return to a more ductile style of contractional deformation during D_6/14. The D_7/14 event likely represents some type of minor tectonic relaxation following cessation of D_6/14 shortening.

Most aspects of the district-scale structural architecture formed during D_1/14 to D_3/14, although the effects of D_5/14 faulting are locally apparent in district-scale geological maps. Mineralised lodes formed at some time between the S_4/14 sericite foliation, which they overprint, and movement on the D_5/14 faults.

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7.3

Project Geology

The mineralised KZ Trend, which hosts the Kibali deposits, is located in the central part of the Moto Greenstone Belt. The Project hosts most of the gold endowment in the Moto Greenstone Belt. The KZ Trend marks an important boundary between older and younger parts of the belt with different provenances (Allibone et al., 2020). The Project includes a wide complex of variably deformed and metamorphosed basalts, dacitic volcanic and volcaniclastic rocks, siliciclastic sedimentary rocks, BIF, and chert (Figure 7-2).

The lithology to the east of the KZ Trend consists of variably deformed and metamorphosed basalt, dacitic volcaniclastic, psammo pelitic schists, amphibolite, BIF, carbonaceous argillite, chert, and granitoid intrusions between the western part of the KZ Trend and Belengo (Bird, 2016; Allibone et al., 2020). Low- to mid- greenschist facies mineral assemblages, which include sericite, chlorite, actinolite-tremolite, carbonate, epidote, titanite, pyrrhotite, and rutile, have partly to largely replaced all primary minerals in rocks between the KZ Trend and Belengo (Allibone et al., 2020). The rocks occurring in the eastern part of Belengo are metamorphosed under upper greenschist to low-mid-amphibolite facies conditions, which comprises garnet-bearing amphibolite facies psammo-pelitic schists, rare unfoliated epidote-hornblende-carbonate-magnetite-bearing calc-silicate rocks, metabasic amphibolites, and relatively coarse-grained recrystallised BIF (Allibone et al., 2020).

The lithology to the west of the KZ Trend consists of a thick package of immature sandstone, gritstone, pebble conglomerate, carbonaceous argillite, banded iron formation, chert, granitoid intrusion, mafic intermediate intrusions (dykes and sills) and probably acid tuffs. These rocks host the KCD deposit (Figure 7-2). Radiometric dating of detrital zircons does not differ from the emplacement ages of the larger tonalitic plutons in the eastern part of the Moto Greenstone Belt and in the region to the south. This suggests that these siliclastic rocks were deposited during a belt-wide basin extension event between 2629 to 2626 Ma, with much of the detritus derived from adjacent older parts of the belt (Allibone and Vargas, 2017; Allibone et al., 2020).

Geochronological and provenance data imply the proto-KZ Trend initiated as a network of extensional faults 2629 Ma (Allibone et al., 2020). An extensive basin subsequently developed west of the extensional proto-KZ Trend, and filled with volcanic, volcaniclastic, and sedimentary rocks between 2629 Ma and 2626 Ma (Allibone et al., 2020). Intra-basin transfer and extensional faults controlled the development of the internal basin structure and likely persisted during later contractional deformation, coinciding with the location of many gold occurrences in the western part of the Moto Greenstone Belt.

Emplacement of minor porphyry intrusions (e.g., Sessenge Tonalite) 2626 Ma was followed by the onset of contractional deformation throughout the Moto Greenstone Belt (Allibone et al., 2020). A complex history of reverse faulting, folding, and ultimately mineralisation occurred over the following 10 to 15 Ma. Older rocks of the eastern Moto Greenstone Belt were thrust across younger rocks of the western Moto Greenstone Belt in the vicinity of the proto-KZ Trend, establishing the altered shear zones which mark the current position of the KZ Trend. At KCD, and likely elsewhere, thrust faults and klippes which formed early in this contractional event were subsequently folded and cut by younger reverse faults. Plutonism during this period was confined to the extended crust of the western Moto Greenstone Belt, west of the KZ Trend. This implies

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that a higher thermal gradient and potential for gold mineralisation persisted in the western part of the Moto Greenstone Belt during this series of contractional deformation events.

7.4

Mineralisation

At Kibali, the gold deposits are largely hosted in siliciclastic rocks, BIF, and chert that were metamorphosed under greenschist facies conditions. Ore-forming H₂O-CO₂-rich fluids migrated along a linked network of gently NE dipping shears and NE to NNE-plunging fold axes of the KZ Trend. On-going deformation during hydrothermal activity resulted in development of lodes in a variety of related structural settings within the KZ Trend. The source(s) of metal and fluids which formed the deposits remain unknown, but metamorphic devolatilisation reactions within the supracrustal rocks of the Moto Greenstone Belt and/or deeper fluid and metal sources may have contributed.

In general, the KZ Trend shows mainly three styles of mineralisation including disseminated, vein and replacement style, all dominated by an Fe-sulphide phase, mostly pyrite, with variable chalcopyrite, arsenopyrite and pyrrhotite (Bird, 2016). Gold generally occurs as inclusions in both disseminated and vein pyrite, and along the margins of pyrite grains (Lawrence, 2011; Bird, 2016). A second phase of gold mineralisation is identified, occurring as fracture hosted gold grains and in the case of the KCD as isolated gold grains within the groundmass (Bird, 2016).

●

Disseminated mineralisation is characterised by sulphide minerals over-printing and replacing chlorite and Fe-carbonate mineral phases in the phyllosilicate-rich inter-clast zones in the deformed volcano-sedimentary conglomerates, constituting the low-grade mineralisation in most deposits.

●

Vein style mineralisation is characterised by the formation of quartz-siderite (±aluminoceladonite) sulphide veins in lithologies that have undergone extensive Fe-carbonate alteration (Bird, 2016; Allibone et al., 2020).

●

Replacement mineralisation is characterised by ankerite-siderite, pyrite alteration (ACSA-B) that is typically texturally destructive.

The gold deposits are associated with halos of quartz, ankerite, sericite, ± albite (ACSA-A) alteration that extend into the adjacent rocks. This widespread ACSA-A alteration assemblage is superimposed on older greenschist facies metamorphic assemblages. Gold is directly associated with ACSA-B alteration. ACSA-B alteration is ACSA-A alteration overprinted by ankerite-siderite, pyrite alteration. In smaller peripheral deposits a late chlorite, carbonate, pyrite assemblage is associated with the mineralisation rather than the ACSA-B assemblage, implying a district-wide zonation of mineral assemblages along and across the KZ Trend. Zones of mineralised ACSA-B alteration are commonly developed along the margins of BIFs, or contacts between chert, carbonaceous phyllite, and BIFs (Figure 7-3). Local remobilisation and upgrading of ACSA-B related mineralisation occurred adjacent to the margins of some post-ore crosscutting chlorite, carbonate, ± pyrite, ± magnetite-altered diorite dykes.

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Mineralised rocks at Kibali typically lack significant infill quartz-rich veins, unlike many other orogenic gold deposits. Gold is instead associated with pyrite in zones of alteration that replaced the earlier mineralogy of the host rocks.

Most gold mineralisation is texturally associated with fine disseminated pyrite, with minor pyrrhotite and arsenopyrite. The auriferous pyrite occurs as both ‘salt and pepper’ disseminated fine grains and clusters of disseminated grains forming blebs and pseudo-vein mosaics. A petrographic study has identified several sulphide phases with arsenopyrite, chalcopyrite, pyrrhotite and pyrite dominating the assemblage with multiple generations of each identified. Gold is hosted within the dominant second pyrite phase and as late fracture fillings associated with chalcopyrite and galena. The gold bearing pyrite is hosted by the sequence of coarser clastic sedimentary unit’s conglomerate and chert-ironstone assemblage, often with an envelope of ACSA-A (Figure 7-3).

Higher grades are associated with ACSA-B with disseminated sulphides. This is interpreted as being a result of silicified and altered host units becoming brecciated as deformation progressed, producing competency contrasts, and increasing permeability. A similar setting is interpreted for the host lithologies, where coarser clastics and chert/ironstone units seem to have behaved in a more brittle fashion (finer grained sediments behaved more ductile).

The preliminary mineralisation model for the area suggests ore-forming fluids were produced in a convergent tectonic environment as part of a thickening thrust stack. Progressive metamorphism and devolitisation of the lower stack generated fluids which ascended upwards along faults, scavenging sulphur and metals. The fluids migrated upward and southward along NE dipping thrust faults and NE trending shears. The shears contributed to development of sheath folding, which in turn contributed to the ACSA alteration in proximal host rocks. The brittle ACSA alteration shattered with progressive deformation, allowing further infiltration of fluids and deposition of gold and sulphides (pyrite). The alteration varies in intensity from weak to texturally destructive ACSA-B.

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Source: After Allibone et al., 2020

A. Carbonate, quartz, sericite (ACSA-A) altered sandstone and siltstone in which sericite is largely confined to spaced folia that cut relict bedding at an oblique angle.

B. Strong carbonate, quartz, sericite (ACSA-A) alteration which has largely destroyed all the primary textures within the protolith. Early-formed carbonate-quartz veinlets have been dismembered along the sericite folia.

C. Siderite-pyrite (ACSA-B) alteration front overprinting ACSA-A alteration and destroying the sericite folia associated with this earlier assemblage.

D. Typical ore from the KCD deposit, comprising numerous irregular-shaped mineralised pyrite veinlets surrounded by siderite, ± quartz, ± magnetite (ACSA-B) alteration. Relicts of the BIF protolith remain within the altered and mineralised rocks.

Figure 7-3 Photograph Showing Examples of Altered and Mineralised Rocks from the KCD Deposit

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7.5

Project Deposits

Most of the mineralisation currently delineated at the Project occurs along the KZ Trend (Figure 7-2). The KCD deposit and satellite deposits (Kombokolo and Gorumbwa) are located in the central part of the KZ Trend (Figure 7-4). Most deposits are located along the north branch of the KZ Trend (Aerodrome to Ikamva), with other targets along the south branch of the KZ Trend (KZ South).

Karagba-Chauffeur-Durba (KCD) Deposit

The KCD deposit is the principal mineralised occurrence along the Sessenge-KCD Trend (Figure 7-4). It consists of five semi-vertically stacked lodes hosted within volcano-sedimentary units. The mineralisation shows a strong correlation with texture destructive ACSA-B alteration (quartz-ankerite-siderite-chlorite) (Figure 7-5).

The lodes are broadly categorised as the upper 3000 lodes, 5000 lodes, and the deeper 9000 lodes, 11000 lodes, and 12000 lodes (Figure 7-6). All generally plunge from surface to the NE at low to moderate angles (approximately 25°) with mineralised wireframes based on drilling intercepts indicating a down plunge continuation of over approximately 2,000 m (remaining open down plunge).

The 3000 lode crops out in the present open pit (Karagba) and is the western-most lode (Figure 7-6). It is approximately 300 m in width, 30 m thick, and has a broad gentle and open semi-synclinal form to its plunge.

The 5000 lode outcrops slightly east and south of the 3000 lode (Chauffeur and Durba) and forms most of the topographically elevated area known as the Durba Hill, on which the historic Durba Plant is situated (Figure 7-6). The lodes are more sub-vertical in attitude than the 3000 and 9000 lodes and are consistently of higher grade.

The 9000 lode does not outcrop in the KCD open pit but crop out to the south of the Durba Hill at Sessenge (Figure 7-6). The 9000 lode is comprised of two main lodes 9101 and 9105. The 9105 is of a similar shape and attitude as the 5000 lode and is connected in part. The 9101 lode joins Sessenge and is a shallow dipping lens with a similar plunge to the 5000 lode.

The 11000 and 12000 lodes were discovered during deep drilling, and were subsequently followed up plunge, where the 11000 merges with the 5000 and 9000 lodes and the 12000 lode crops out at Sessenge SW (Figure 7-6).

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The location of the individual lodes within the KCD deposit are intimately controlled by the position, shape, and orientation of a series of gently NE-plunging tight to isoclinal folds. The lodes may be linked genetically by large-scale recumbent folding developed between two bounding NE trending structures locally termed the Eastern Transfer Fault and the Western Transfer Fault. Almost all the anomalous and economic mineralisation at the KCD deposit occurs in areas located between the Eastern Transfer Fault and Western Transfer Fault, in conjunction with increases in alteration and structural deformation and localised refolded fold (previously called sheath fold) development. Mineralisation is hosted with conglomerate units, and ironstone and chert assemblages, enveloped within a halo of weak to moderate siderite-ankerite-silica-sericite-sulphide alteration. Higher grade (generally with increased sulphide content) developed in zones of strong to intense alteration that overprinted and texturally destroyed previous breccia, foliation, and lithological textures.

The ACSA-A alteration developed during the formation of these folds, and the sericite foliation which is an integral part of the ACSA-A assemblage formed parallel to their axial planes. Zones of later auriferous ACSA-B alteration developed along the axis, limbs, and more rarely the axial planes of these folds, locally wrapping around the hinges of the folds to form elongate NE plunging concave-shaped rods. ACSA-B alteration is also commonly focused along the margins of more extensive banded iron formations, indicating a stratigraphic as well as structural control on the distribution of ore, both within KCD, and other parts of the wider KZ Trend. Shear zones that were active during folding are a third key structural control on the location of ore within KCD and the wider KZ Trend. At KCD a folded carbonaceous shear in the core of the deposit juxtaposes stratigraphically distinct blocks, separating the 3000 lodes from the 5000, 9000 and 11000 lodes. The 3000 lodes above this shear are hosted by locally ferruginous cherts, carbonaceous argillites, and minor greywacke, whereas the 5000, 9000 and 11000 lodes below are hosted by siliciclastic rocks and banded iron formation. Fold shapes and wavelength differ between the two blocks reflecting their different rheology during folding, and this is reflected in the scale, shape, and continuity of lodes in each block.

Sessenge (Sessenge-KCD Trend)

The Sessenge deposit is located approximately 1 km to the SW of the KCD deposit (Figure 7-6). Interpretations of drill data suggest the Sessenge deposit mineralisation represents the up-plunge continuation of the KCD 9000 lode (Figure 7-6). Mineralisation at Sessenge forms one main mineralisation lode comprised of multiple small high-grade shoots that plunge to the NE towards KCD. Gold mineralisation is associated with ACSA alteration, and occurs with an upper BIF unit, and a lower intercalated ironstone and clastic sedimentary unit. Strong chlorite alteration occurs in places, which has been erroneously logged as basic intrusive in historic RC drilling. The mean grade of the main low-grade halo is 1.5 g/t Au with several high-grade shoots between 4.0 g/t Au and 5.0 g/t Au.

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Gorumbwa and Kombokolo Deposits (KCD area)

Both deposits occur to the W and NW of the KCD deposit (Figure 7-4) They are considered to be part of the same mineralising event, however mineralised zones are of significantly smaller dimensions as compared to the KCD deposit. High-grade shoots occur in the order of 30 m to 50 m as observed at KCD, but the surrounding low-grade mineralised halo occurs in the order of tens of metres. The Gorumbwa lode plunges at a low to moderate angle to the NE. The mineralisation consists of a series of stacked lenses that variably extend down plunge for a length of 1,000 m at an average width of 200 m and has been identified at a depth of 400 m below topographic surface. In 1995, SOKIMO commenced mining from underground and small open pit operations. Historical underground workings extend to 380 m below surface.

The lithological sequence based on mapping and core logging include a coarse meta sandstone sequence and meta conglomerate (coarse quartz clast) packages with intercalated medium to fine grained sediment horizons (meta arenite and meta siltstone units). The contact between the meta sandstone sequence and upper coarse clastic sequences is marked by a thin matrix supported polymictic red chert or jasper clast bearing conglomerate horizon, which is useful as a marker bed horizon. A dolerite unit sub parallel to lithological layering mapped in the SW is the main “Banc Vert”, another marker horizon that sits predominantly within the coarse meta sandstone unit immediately above in the hanging wall of the main mineralised 1004 lens. From drill intersection interpretations, the unit is interpreted to be sill-like. The Banc Vert is generally a medium grained mafic unit (dolerite) which becomes fine-grained at the contact with the host rock. The main constituents are chlorite, feldspar, and ankerite-calcite, with the ankerite appearing as porphyroblastic euhedral crystals up to 2 cm in diameter. Some parts of the mafic intrusion contain magnetite. The Banc Vert is not mineralised.

Mineralisation at Gorumbwa is hosted almost exclusively within the meta sandstone unit, with minor sporadic mineralisation noted in a conglomerate unit that occurs beneath the meta sandstone. Mineralisation is divided into twelve lensoidal lodes (1001 to 1012) which broadly trend west to WSW, dip to the NW, and plunge to the ENE at approximately 30°. The lenses are echelon-like in vertical stacking, with only the main 1004 lens being the most consistent in continuity. The stratigraphically higher upper lenses include 1001 to 1003, with 1005 to 1008 for the deeper footwall lenses. Higher grades within the lodes occur in the central area of the shoots where a higher strain environment increased space and hydrothermal fluid inflows. Historic mining focussed on the extraction of the main (1004) lode. The styles of mineralisation vary from KCD, with the dominant style being moderate to strong silicification and sericitisation with minimal pyrite. There is low correlation between sulphide and gold content. The second style is the typical ACSA B style noted at KCD where the gold is proportional to pyrite percentages, though the iron carbonate is predominantly ankerite, unlike KCD which is dominated by siderite. This style is mainly observed in the main 1004 lode. The third style is visible gold within late, moderate to strong silicification. Mineralisation is structurally controlled within a NE trending corridor where the foliation strikes EW within the central area rotating approximately 30° to 240° on the western edge of the corridor. The corridor is bounded by NE crosscutting structures on the eastern and western margins. These NE structures are indicated by the discontinuity of the red pebble conglomerate horizon in the west, micro folding within the lithological units near to the structures and rotation of S1 foliation laterally across the ± 200 m wide mineralised area. The structures

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may be ductile. Refolded folds are not observed but may have been seen in the high-grade shoot, which was mined and makes up the void area.

The Kombokolo deposit lies approximately 1500 m to the NE of the Gorumbwa deposit, on the east side of an elongated topographic high locally named Kombokolo Hill (Figure 7-4). The hill is capped by an ironstone unit that strikes NE and dips gently-to-moderately to the NW. Mineralisation is located in a clastic conglomerate unit in the footwall of the ironstone, with moderate pervasively carbonate-sericite-silica-pyrite alteration. The mineralisation plunges at low to moderate angles (30°) to the NE. The lode has a down plunge continuation of over 300 m to a depth of 170 m below surface, an average width of 100 m, with an average thickness of 20 m. The lode remains open down plunge.

Aerodrome, Pakaka and Pamao Deposits (KZ North Trend)

The Aerodrome-Pakaka-Pamao deposits are located along the KZ North trend, in the gently NNE- to E-dipping shear zone (Allibone et al., 2020). The Pakaka deposit is the largest gold deposit discovered to date in the Project area other than KCD. The stratigraphic section at Aerodrome-Pakaka-Pamao is comprised of three main lithological packages. The hanging wall rock package consists of older rocks deposited before 2640 Ma (Allibone and Vargas, 2017; Allibone et al., 2020) and by an upper tholeiitic basalt flow sequence with interbedded argillite and graphitic carbonaceous shale horizons called the Pakaka-Pamao Hanging Wall Formation; a middle sequence of meta conglomerate interbedded with abundant felsic crystal tuff, undifferentiated tuff, as well as lesser horizons of siltstone and at Pamao, localised magnetite alteration. A lower footwall rock package, deposited between 2630 and 2625 Ma, comprises an immature sandstones and gritstones inter-layered with minor beds of pebble conglomerate and BIF.

Gold mineralisation at Aerodrome-Pakaka-Pamao is hosted in altered (iron carbonate and chlorite with lesser amounts of silica, sericite, pyrrhotite, pyrite, and auriferous disseminated arsenopyrite) and sheared meta volcaniclastic rocks (Allibone and Vargas, 2017) interbedded with minor tuffaceous units. The presence of significant arsenopyrite at Pakaka distinguishes it from other deposits and prospects along the northern half of the KZ Trend (Allibone et al., 2020). The mineralised zones are associated with pervasive silicification with local preservation of breccia textures that have been overprinted by the dominant S1 fabric. Higher gold grades appear to correlate well with the presence and abundance of pyrite-arsenopyrite, and empirically the high-grades appear to be spatially associated with the intersection of the NW trending thrust surface, and a NE trending strain corridor. The structures combine to produce a broad NE plunging open anticlinal structure, with Pamao on the west limb, and Pakaka on the east. The Pakaka mineralisation continues down plunge beyond the limits of the drilling and represents a further exploration potential. The Pakaka mineralisation extends over a strike length of 1,000 m, averages a thickness of 30 m and has been identified to a depth of 350 m below surface. The weathering profile at Pakaka is relatively deep up to 70 m.

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Mengu Hill Deposit (KZ North Trend)

The Mengu Hill deposit lies on the KZ North structure, to the NW of Pakaka and to the south of Mofu-Oere (Figure 7-4). The stratigraphy in the vicinity of the deposit is dominated by a meta conglomerate unit that is interbedded with fine-grained sediments, siliceous sericite schist and minor mafic volcanic rocks. These lithologies overlay a massive magnetite and specular hematite ironstone-chert unit that has weathered to create the topographic high, Mengu Hill, the ironstone protecting the northern face from weathering and erosion. Mineralisation is associated with silica-ankerite-pyrite alteration that is focused within the ironstone unit and along its contact with the overlying conglomerate unit. The mineralised lens is cigar like in shape and plunges shallowly to the NNE within a zone of ACSA-B altered rock in the upper part of a thick BIF unit. Further barren ACSA-B altered rock occurs widely within the surrounding BIF unit beyond the mineralised lode, unlike ACSA-B altered rocks at the KCD deposit that are invariably mineralised to some degree (Allibone et al, 2020). Narrow zones of intensely carbonate altered rock, locally containing conspicuous prisms of chloritoid, overprint a layer of siliciclastic rock sandwiched between layers of BIF adjacent to the lode (Allibone et al, 2020). Zones of strong chlorite-carbonate alteration are developed along other BIF siliciclastic rock contacts above and below the BIF unit. The Mengu lode coincides with an asymmetric bend in the upper surface of the host BIF unit. The deposit does not coincide with either a major shear zone or a larger tight to isoclinal fold like those that host the deposits. Drilling in the footwall of Mengu and geologic mapping across the district suggest it is located 100 to 200 m above the shear zone that hosts the Pakaka deposit. The Mengu Hill mineralisation averages a width of 150 m and continues 700 m down plunge to a depth of 250 m below the topographical surface. Mineralisation remains open down plunge.

Mengu Village and Megi-Marakeke-Sayi Deposits (KZ North Trend)

At Mengu Village, located near the NW end of the Pakaka-Mengu Trend, the mineralisation is tabular in form, trending NW and dipping shallowly to the NE. The mineralisation is approximately 150 m in strike length with an average thickness of 15 m and has been identified to a depth of 150 m below the surface. The mineralisation is hosted by conglomerates with thin ironstone and carbonaceous shale intercalations.

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Ikamva, Kalimva, Oere and Mofu Deposits (KZ North Trend)

Ikamva-Kalimva & Oere deposits are all located along the major lineament of on KZ North Trend, north of Mengu Hill where the KZ structure rotates to the NNE (Figure 7-2).

These deposits are broadly similar in geology, consisting of hanging wall volcanic to volcaniclastic succession associated with BIF, intrusions, and carbonaceous shale. These are dated as older rocks deposited before 2640 Ma (Stenhouse, 2020; Allibone et al., 2020). The footwall comprises siliciclastic sedimentary rocks and is dated as a younger basin deposited between 2630 and 2625 Ma (Allibone et al., 2020).

These deposits are characterised by an intense shear deformation associated with widespread carbonate-chlorite-quartz altered rocks. However, Kalimva is mostly dominated by a variable-intensity chlorite-quartz-carbonate-pyrrhotite±pyrite-ilmenite assemblage, while Ikamva and Oere are dominated by ACSA alteration (quartz-carbonate-sericite±subordinate chlorite-pyrite) with a distinctive buff-coloured variant of ACSA-A and a texturally destructive ACSA-B assemblage (FeCO3-quartz±chloritoid±magnetite-pyrite) often spatially associated with mineralisation (Stenhouse, 2020). The mineralisation lodes in Kalimva, show a shallowly NNE-plunging ore-shoot along a moderate to steeply E-dipping structure locally called Kalimva Deformation Zone and interpreted as an equivalent of lower layer-parallel shear at Ikamva deposit, characterised by a narrower chlorite-altered high strain BIF (Stenhouse, 2020).

The deformation event interpreted from structural features recorded over the area (field mapping and drill core logging) suggest a complex deformation history with pervasive foliation as a dominant fabric, commonly folds and transposes small-scale layering but is oblique to project-scale lithological units; sub-parallel to the lithology, mainly observed in Ikamva (Stenhouse, 2020). The mineralisation formed during the emplacement of older eastern succession rocks over younger western succession rocks.

The Kalimva deposit was the first of these deposits discovered in 1950 by historical miners who focussed workings along on a N to NNE-striking fault on the western margin of the prospect. The coincident distribution of these workings and the late fault has led to a belief that the fault, the Kalimva Deformation Zone, is a key control on the distribution of mineralised rocks at Kalimva (Allibone and Vargas, 2017).

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8	Deposit Types

Gold deposits of the Kibali district are part of the globally significant group of Neoarchean orogenic gold deposits, examples of which are found in most Neoarchean cratons around the world. However, the gold deposits of the Kibali district are hosted within the sulphide vein and disseminated in the altered country rocks instead of in mineralised quartz veins, found generally in most Neoarchean cratons around the world (Allibone et al., 2020). Vein hosted gold is identified in the Project area, but to date only represents small scale mineral occurrences.

Gold mineralisation within the Neo-Archean Moto Greenstone Belt is associated with epigenetic mesothermal style mineralisation, consistent with the majority of Archaean and Proterozoic greenstone terranes worldwide. The type of deposit has been termed orogenic gold and is generally associated with regionally metamorphosed terranes that have experienced a long history of thermal and deformational events and intrusion by igneous complexes. As such, the gold deposits are invariably structurally controlled. The most common style of mineralisation in this setting is fracture, vein-type and disseminated gold bearing sulphide mineralisation in zones of brittle fracture to ductile folding and dislocation.

The Kibali deposits differ from many orogenic gold deposits in terms of structural setting. Rather than being linked to a major large scale steeply dipping strike slip fault with brittle-ductile deformational evolution, many of the deposits are hosted within a thrust stack sequence with ductile to brittle-ductile deformational structures and complex folding history. Some Kibali deposits, like Kalimva and Oere, are more typical orogenic gold deposits, with planar mineralised lodes associated with mineralised brittle ductile fault systems, and with high-grade shoots associated with geological intersections and/or flexures of the host fault zone.

The richly mineralised KZ Trend appears to have initiated as an extensional fault system along the boundary between the relatively young basin in the western part of the belt and older rocks to the east. Mineralisation occurred during the later stages of subsequent regional contractional deformation which resulted in inversion of the basin, and development of reverse faults and folds.

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9	Exploration

The Project has been explored since the early 1990s by geochemistry sampling, mapping, trenching, geophysical surveys, and drilling. Kibali Goldmines has been exploring the Project area since 2010. Exploration on the Kibali property prior to Kibali Goldmines ownership is described in Section 6 of this Technical Report.

During 2021, Kibali Goldmines spent approximately $7.3 million on greenfield and brownfield exploration.

9.1

Exploration Concept

The Kibali district is extremely prospective for gold mineralisation, with a relatively low exploration maturity when compared to mature districts such as in Canada, USA, or Australia. The full potential of the district remains undefined with mineralised rocks intersected >1,000 m below the surface in the deepest holes drilled to date (only seven holes to date drilled >1,000 m, representing only 0.02% of total drill holes).

A fundamental exploration approach in the Kibali district involves the mapping deep crustal, long lived gold bearing structures (using geophysical, geochemical, isotope data and regional geological mapping) that have the potential to supply volumes of fertile hydrothermal fluids sufficient to host world-class gold deposits. Second order structures are delineated to target prospective gold depositional sites within prospective host lithologies (such as chemically reactive or rheologically contrasting units, such as BIF, cherts, or carbonaceous shale) or structural dilation zones, which have the potential to concentrate gold in sufficient concentrations to form an economic deposit. Existing and identified targets are ranked using Barrick’s Area Selection Criteria, based on each target’s geological potential and confidence scores, the results form a framework for target prioritisation and budget allocation.

Exploration at Kibali is structured to simultaneously advance brownfields targets to rapidly feed into the mine plan, and to develop greenfield targets to replenish the target pipeline and sustain the long term growth of the mine. Brownfields exploration efforts at Kibali test for extensions of open pit and underground deposits, testing lode extensions using aggressive step out exploration, and for gap opportunities within the mine area. Once a geological model is defined and tested by exploration and the target demonstrates potential, the target is shared with the Mineral Resource Management department for follow-up drill testing and resource evaluation.

Satellite deposits and gaps between existing Mineral Resources are periodically re-evaluated to define Mineral Resource extensions based on conceptual targets. During 2022, key exploration programmes will target extensions and gaps to KCD-Gorumbwa-Kombokolo-Agbarabo, Kalimva, Oere, Sessenge SW, Gorumbwa SW, and Mengu Hill with the aim of identifying and defining new Inferred Mineral Resources.

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9.2

Geology and Geochronology

Since 2011, in-depth geologic and geochronologic investigations have been undertaken on a variety of scales within the KCD deposit, along the KZ Trend, and throughout the MotoGreenstone Belt to define the internal structure, hydrothermal character, and geologic context of gold deposits more clearly in the Kibali district (Lawrence, 2011; Bird, 2016; Jongens et al., 2016; Allibone and Vargas, 2017, and Allibone et al, 2021).

9.3

Geophysics and Remote Sensing

Detailed interpretation of multi-source remote sensing datasets with ground checking of geologic and geophysical features forms the basis of the Kibali exploration programmes. Remote airborne data sets include high-resolution magnetics, radiometrics, and electromagnetic (EM) and detailed topographic surveys (LiDAR). The distribution and form of the ironstone units, carbonaceous shale horizons, and intrusives in the Project area can be mapped out by the airborne data sets with confidence. Targets with coincident magnetic highs (BIF), EM conductive highs (carbonaceous shales), structural complexity with folding and dislocations, evidence of alteration and/or geochemical anomalism are of particular interest.

Spectrem Air Limited completed an airborne EM, magnetic and radiometric survey in 2010 over the Project area (Figure 9-1 and Figure 9-2). A total of 10,559 line-km was surveyed at a nominal line spacing of 200 m, the KCD area was in filled to 100 m line spacing.

To improve the detail of mapping prospective host lithological units and structures, in January 2020, Xcalibur Airborne Geophysics completed a high resolution aeromagnetic and radiometric survey along the KZ Trend at nominal line spacing of 50 m, for a total of 7,221 line-km (Figure 9-1).

The airborne EM and magnetic data have both indirectly contributed to target generation by enhancing lithological and structural interpretations, and directly through detecting and outlining several NE plunging highly conductive linear shapes. Although the EM anomalies do not directly map gold mineralisation, it is thought that the conductive linear shapes highlight structurally prospective areas and have been interpreted as representing graphitic carbonaceous shale, which has been deformed into a rod like shape by NE trending structures. The magnetic anomalies delineate trends of BIF units and highlight some of the intrusive bodies.

Geophysical datasets have been combined with a longer-term study to develop a tectonostratigraphy for Kibali, and to improve the understanding of the controls to gold mineralisation and regional geologic architecture. This Project-wide geologic framework is driving a re-assessment of exploration work to date as part of greenfields target generation.

In 2020, a high-resolution topographical survey was undertaken by Southern Mapping to produce a digital terrain model (DTM) and rectified colour images of the KZ South area, thus completing high-resolution DTM coverage over the entire KZ Trend (Figure 9-3). The topographical survey was carried out using an aircraft mounted LiDAR system to create a high-resolution DTM of the ground surface and objects above the ground (>6 cm vertical accuracy). Digital colour images

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were also captured from the aircraft and used to produce colour orthophotos of the project area, with a 7 cm pixel resolution. The remainder of the Project area utilised topography data from Shuttle Radar Topography Mission (SRTM).

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9.4

Geochemical Sampling

Soil samples are the first pass geochemical exploration technique used in the western portion of the Project, where ease of access and suitable terrain aids field activities. Despite historical ASM workings and potential surface contamination, the thin horizons of transported cover, shallow depths of paleo weathering surfaces (marked by quartz gravel layers in the district), and weak laterite development produce robust geochemical anomalies which are in general proximal to sources of mineralisation. Geochemical anomalies correlate well with the KZ Trend (as anomalies at the Pakaka-Mengu trend and Kalimva) and NE trending structure corridors (as at KCD and Gorumbwa).

Prior to conducting a soil sampling programme, a regolith map is produced, by interpreting remote datasets (including DTM, satellite imagery and radiometrics) and field validation. Test pits are excavated to further understand the regolith profile, thickness, validate regolith mapping and ultimately to identify any regolith characteristics that may impact soil results. Once a grid is designed, each sampling station is cleared of surface vegetation prior to sampling. A hole is excavated to approximately 30 cm depth to sample the B horizon and a 1 kg sample is collected. If quartz fragments are abundant, the sample is sieved to <5 mm. Samples are collected at 50 m centres along lines spaced 200 m and 400 m apart. Anomalous lines are in filled with samples at 50 m centres along lines spaced 100 m and 200 m apart. Soil samples are analysed by aqua regia-atomic absorption spectroscopy for gold and X-Ray Fluorescence (XRF) for multi-elements. Table 9-1 summarises the soil samples collected annually.

In the eastern portion of the exploitation permit, thicker horizons of transported cover (>2 m) and higher-grade metamorphism demonstrated that further refinement of the interpretation of historic geochemical results was required. Therefore, a stream sediment sampling programme was executed in 2018 to cover the whole Kibali permit. The purpose was to generate potential new greenfield targets with a greater confidence than the historic soil sampling alone. Samples were analysed for low detection gold and for 53 elements to define pathfinders. Gold shows moderate to good correlation with As-Sb-W. Anomalous catchments have been ranked and selected for follow up soil sampling and mapping. A review of multi-element and gold results in conjunction with one another highlights trends that can aid discrimination between real and transported anomalies. Table 9-1 summarises the stream sediment samples collected annually.

Stream sediment sampling identified grassroots targets Makoro, Abimva, Kialo, Lanza and Marabi in the east of the licence (Figure 9-4). The Lanza target was field tested in 2020 and the anomalous catchment was found not to have the potential to host a deposit meeting Barrick’s area selection criteria.

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Table 9-1 Kibali Soil and Stream Sediment Sample Summary



Year	Company	Number of Soil Samples	Stream	Total Number of Samples

2008	Moto	28,864	-	28,864

2009	Kibali Goldmines	5,030	-	5,030

2010	Kibali Goldmines	617	-	617

2013	Kibali Goldmines	205	-	205

2014	Kibali Goldmines	1,673	-	1,673

2015	Kibali Goldmines	2,295	-	2,295

2016	Kibali Goldmines	-	-	0

2017	Kibali Goldmines	4,073	-	4,073

2018	Kibali Goldmines	-	313	313

2019	Kibali Goldmines	2,420	-	2,420

2020	Kibali Goldmines	1,528	-	1,528

2021	Kibali Goldmines	447	-	447

Total		47,152	313	47,465

Geophysical and geochemical targets are investigated with geologic mapping, pitting, and trenching prior to drill testing. Table 9-2 presents the Kibali trenches, auger, and pit lithosamples collected annually.

Table 9-2 Kibali Trenches, Auger and Pits Summary



Year	Company	Trenches		Auger		Pits		Total
		Meters	No.	Meters	No.	Meters	No.	Meters	No.

2010	Kibali Goldmines	481	5	-	-	273	48	754	53

2011	Kibali Goldmines	398	2	350	185	538	147	1,286	334

2012	Kibali Goldmines	1,050	43	1,083	181	691	131	2,823	355

2013	Kibali Goldmines	3,216	61	11	2	498	165	3,725	228

2014	Kibali Goldmines	8,570	83	83	23	1,115	383	9,768	489

2015	Kibali Goldmines	12,240	110	800	360	3,727	1,128	16,767	1,598

2016	Kibali Goldmines	8,066	101	1,799	843	1,830	648	11,694	1,592

2017	Kibali Goldmines	8,712	58	-	-	1,596	605	10,308	663

2018	Kibali Goldmines	7,751	53	5791.75	1128	1,137	334	14,680	1515

2019	Kibali Goldmines	4,073	30	1178.57	265	314	87	5,565	382

2020	Kibali Goldmines	3,336	21	-	-	123	50	3,459	71

2021	Kibali Goldmines	361	5	-	-	43	24	527	33

Total		58,255	572	11,096	2,987	11,885	3,750	81,357	7,313

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9.5

Proposed 2022 Greenfields Exploration

Targets are identified through the review of geophysical, geochemical, and remote sensing data in conjunction with mapping and historic drilling datasets. During 2022, exploration programmes are planned for the following targets (Figure 9-5):

●

Kolapi – scout and framework drilling to test the southern extension of Oere, along the KZ North structure.

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MMR (Madungu-Memekazi-Renzi) located SE of KCD – scout and framework drilling following up prospective historic drilling results, rock chips, and trench within a sericite altered high strain corridor with prospective host rocks (BIF, chert, carbonaceous shales, and conglomerate).

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Aindi Watsa, SW of KCD on the KCD trend – geological model review and framework drilling.

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Korogo – geochemical anomaly along the Gorumbwa-Kombokolo trend, SW of Gorumbwa. Geological model review and scout drilling.

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KZ South, Zakitoko and Zambula – scout and framework drilling to test targets along the 15 km structure with anomalous mineralisation.

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KZ North, Mengu Hill to Ikamva – scout and framework drilling to test targets along the KZ North structure with anomalous mineralisation.

In addition, the exploration programme will include testing a number of grassroots targets identified by the 2018 stream sediment survey. Initially during 2022, follow up works will include geological mapping, local soil sampling grids, and rock chip channel sampling at Makoro, Abimva and Marabi (Figure 9-4). If successful, targets will be further tested with scout drilling. Additional anomalous catchments will also be tested during following three to five years to sustain a level of exploration target turnover that ultimately supports the mines depletion replenishment pipeline for several years.

9.6

Proposed 2022 Brownfields Exploration

The current underground drilling at KCD is aimed at defining additional extensions to mineralisation to increase the underground Mineral Resources and Mineral Reserves over the next five years. Drilling is completed from dedicated exploration drill drives particularly in the down and up plunge of the 3000 lodes and down plunge of the 5000, 9000, and new 11000 lodes. Analysis of deeper UG opportunities below the base of the existing shaft is planned to be conducted, including down plunge extensions, testing in both the hanging wall and footwall of the KCD system, refining of the 12000 lode conceptual model, and identification of any new potential lodes that can be connected to the existing KCD underground infrastructure. Execution of 2D seismic lines in the KCD area is also planned to support exploration deeper mineralisation.

A number of satellite pits, including Gorumbwa, Pakaka, Kombokolo, Mengu Hill and Ikamva will also be drill tested for down plunge extensions to mineralisation and evaluate their economic viability for further smaller satellite underground operations to support the mine life extension outside of the existing LOM.

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Satellite deposits and gaps between existing Mineral Resources and Mineral Reserves are planned to be evaluated to develop additional targets to add to the open pit Mineral Resources and Mineral Reserves, maintaining a robust depletion replenishment pipeline for several years. During 2022 drill programmes are planned at Oere (North and South extensions), Mengu Village and Ikamva East. Ongoing drilling is also planned in the Gorumbwa-Sessenge-KCD gap to test the concept of combining the three pits especially considering that the Gorumbwa & Sessenge pits now merge.

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Source: Kibali Goldmines, 2021

Figure 9-5 Geological Map of the KZ Trend, Showing Targets and Known Deposits

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9.7

Discussion

Kibali has a detailed SOP Manual for Exploration and Drilling Practices that provides standardisation and consistency to ensure the collection of quality data by all field technical personnel. In the opinion of the QP, all samples collected are representative and unbiased.

In the opinion of the QP the exploration programmes completed to date are appropriate to the style of the deposits and prospects within the Kibali Project. Kibali retains significant brownfields exploration potential, and additional work is planned.

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10	Drilling

Reverse circulation (RC) and diamond drilling (DD) are used to support Mineral Resource estimation. Rotary air blast (RAB) drilling has previously been used in regional first pass exploration and for sterilisation purposes. Sample data from RAB drilling trenches (TR), open-pit rip-lines, and underground channels are not used for Mineral Resource estimation.

Table 10-1 presents the known drilling by year, company, and type at the Project. The cut-off date for drilling is 30 September 2021 except for the 11000 lode which is 15November 2021.

Since 2009, 2,257,938 m of drilling from 29,324 DD and RD holes have been drilled. This data has been used for estimation of Mineral Resources.

Prior to 2009, a total of 442,423 m of historical drilling was conducted by previous operators in different drilling campaigns described in Section 6. This historical drill hole data now constitutes a minority (11%) of the total database used in the geological framework and for the estimation of Kibali Mineral Resources and Mineral Reserves (see Section 14). This data is used for exploration targeting but has been effectively superseded by current drill holes within the declared Mineral Resources.

10.1

Drilling Definitions

Kibali is an advanced project with operating open pits and an underground mine. As such drilling is completed regularly as part of ongoing operations. All drilling falls into three categories each with specific objectives and outcomes as follows:

●

Exploration Drilling (EXP) - Wide spaced exploration drilling intended to grow the Mineral Resource base.

●

Advanced Grade Control (AGC) Drilling – Consists of first pass wide spaced grade control drilling to increase confidence in open pit and underground Mineral Resources to a sufficient level of confidence to support Probable Mineral Reserves.

●

Infill Grade Control (GC) Drilling – Consists of close spaced grade control drilling for final production definition to inform Measured Mineral Resources/Proven Mineral Reserves. Generally, Kibali Goldmines’ inventory of infill GC drilling is approximately six to 12 months of production coverage for open pits and between 18 and 24 months for underground.

DD is used for exploration, resource evaluation work, hydrogeological work, geotechnical work, collecting metallurgical samples, and for checking/twinning previous RC intercepts.

RC holes are used for exploration, AGC, and GC drilling. If penetration rates of the RC drilling decrease materially or if groundwater inflows prevented the collection of a dry sample, then the drill hole is continued with a DD tail. In some cases, in the hanging wall units where mineralisation is not intersected, the RC pre-collars are continued through zones of significant groundwater and associated wet samples to achieve the planned pre-collar depth prior to commencing the DD tail.

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Table 10-1 Kibali Drilling Summary



Year	Company	Diamond Drilling		Reverse Circulation		RC Collar + DD Tail		Total
		Metres (m)	No. of Holes	Metres (m)	No. of Holes	Metres (m)	No. of Holes	Metres (m)	No. of Holes

1950	OKIMO	35,153	242	2,856	102	-	-	38,009	344

1951	OKIMO	1,259	15	-	-	-	-	1,259	15

1952	OKIMO	294	5	-	-	-	-	294	5

1960	OKIMO	16,162	175	-	-	-	-	16,162	175

1980	Moto	1,484	10	-	-	-	-	1,484	10

1996	Barrick	8,988	70	-	-	-	-	8,988	70

2004	Moto	9,840	50	42,133	655	-	-	51,973	705

2005	Moto	42,672	201	51,685	739	-	-	94,357	940

2006	Moto	50,396	227	34,658	558	178	1	85,232	786

2007	Moto	51,540	125	19,574	402	-	-	71,114	527

2008	Moto	50,516	98	-	-	-	-	50,516	98

2009	Moto	23,035	67	-	-	-	-	23,035	67

	Sub-Total	291,339	1,285	150,906	2,456	178	1	442,423	3,742

2009	Kibali Goldmines	2,938	9	-	-	-	-	2,938	9

2010	Kibali Goldmines	28,403	64	24,166	483	-	-	52,569	547

2011	Kibali Goldmines	10,507	28	59,192	1,811	-	-	69,699	1,839

2012	Kibali Goldmines	23,166	79	94,764	1,834	-	-	117,930	1,913

2013	Kibali Goldmines	18,794	77	80,036	1,487	-	-	98,830	1,564

2014	Kibali Goldmines	34,079	176	140,283	2,941	417	3	174,779	3,120

2015	Kibali Goldmines	52,375	311	112,260	2,372	2,715	17	167,350	2,700

2016	Kibali Goldmines	71,834	559	210,908	2,950	8,691	48	291,433	3,557

2017	Kibali Goldmines	122,074	700	202,680	2,854	-	-	324,754	3,554

2018	Kibali Goldmines	112,571	616	114,867	1,701	772	3	228,209	2,320

2019	Kibali Goldmines	79,584	409	102,002	1,514	-	-	181,586	1,923

2020	Kibali Goldmines	116,729	551	133,902	1,900	-	-	250,631	2,451

2021	Kibali Goldmines	113,698	672	182,739	3,152	793	3	297,230	3,827

	Sub-Total	786,752	4,251	1,457,799	24,999	13,388	74	2,257,938	29,324

Total		1,074,269	1,078,091	5,536	1,608,705	27,455	13,566	75	2,700,361

Notes:

OKIMO = Office des Mines du Kilo-Moto

Moto = Moto Goldmines Ltd

Figure 10-1 and Figure 10-2 show a drill plan and representative cross section through the largest deposit, KCD.

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10.2

Drill Planning and Site Preparation

Drill holes are planned in Vulcan, LeapFrog and Micromine software. Consideration is given to the orientation of the drilling in relation to the geological structures, to provide for unbiased sampling. Drilling directions are optimised on an individual deposit basis to ensure that the preferred drilling direction is on a cross plunge basis, cutting the geological trend perpendicularly, or at high angle close to true thickness. Efforts to avoid low angle intercepts that may introduce bias are ongoing during drill programme design and budgeting.

The senior geologist, drill contractor, mine planner, mine surveyor, and mineral resource manager all sign off on the drill hole plan prior to initiating drilling.

Open pit drill collars, as well as back sights and foresights are surveyed using a Differential Global Positioning System (DGPS) and are then staked by Kibali Goldmines mine surveyors or geologists. Underground drill collars, as well as back sights and foresights, are surveyed using total station underground survey instruments, and marked on the drift walls, by Kibali Goldmines mine surveyors.

10.3

Downhole Surveying

Reflex EZ-Trac tools were used prior to mid-2016 but were replaced by Reflex EZ-Gyro. When both EZ-Trac and conventional Gyro surveys were being completed, the results of the Gyro survey took higher priority than those of Reflex EZ-TRAC surveys.

Orientation surveys are completed on all holes using either a Reflex EZ-Gyro or a Reflex Sprint-Gyro (new gyro tool introduced in 2020). Reflex EZ-Gyro surveys were undertaken in both up hole and down hole directions every 5 m and Reflex Sprint-Gyro surveys are undertaken in an up-hole direction every 3 m.

Downhole survey equipment is calibrated yearly and checked every quarter by Reflex technicians during site visits.

10.4

Collar Surveys

All drill collar locations are surveyed using differential GPS to 10 mm accuracy.

The Mine uses the UTM Zone 35N datum WGS84 grid for drill hole coordinates.

10.5

Diamond Drilling

DD is primarily used to establish a robust geological understanding of the controls on mineralisation, for Mineral Resource extension work, for geotechnical, hydrogeological, or metallurgical investigation, and to confirm deep (>200 m) very high-grade intersections in RC holes, via twinning.

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From surface, 85.0 mm diameter core (PQ) is generally drilled for the first 100 m down hole with 63.3 mm diameter core (HQ), or 47.6 mm diameter core (NQ) used from 100 m to 200 m depending on the drilling depth requirement. All underground grade control diamond drilling is completed in NQ.

Recent diamond drilling has generally been completed by Boart Longyear and Ore Zone, though some has been carried out by COMISEMI, a local contractor.

Core recoveries are in general excellent, with an average of 98.8% recovery in the unweathered rock, 94.3% recovery in the transitional zone and 73.6% in saprolite zone. Average mineralised interval recovery was 98.7% with a range of between 70% and 100%.

Drilling Procedure

A project geologist must be on site prior to drilling commencing and ensuring that the drill rig is lined up as per the drill plan and supervising drilling, core orientation and down hole surveying. Once each drilling run is complete, the drill core is removed from the drill rod and placed in an angled iron rack to mark up an orientation line with red chinagraph pencil or crayon, as indicated by a Reflex ACT II Core Orientation Tool. The apex of the structure is also marked on the core in a chinagraph pencil or crayon by the core technician. If the orientation and apex lines are overlapping, then the apex line is offset by 5 mm.

DD core is transferred into metal core trays and a plastic down hole depth marker is placed at the end of each core run with the depth marked on it. All areas of core loss are identified, and the run markers are updated with the core recovery. Each drill core box is marked with the hole ID, top and bottom depth of the core and the box number. The core is then transferred to the core yard facility for logging and sampling.

Core Logging

DD core is geologically logged into standardised paper log sheets that include weathering, grain size, mineralisation, alteration style, lithology, structural measurements with sketches and redox data. This is manually transcribed to Excel before being stored in a central database, after the responsible geologist has validated their inputs.

Geologists create a sampling plan using the same paper sheets and label the boxes and core with sample codes. The core (both wet and dry) is then digitally photographed using a purpose built imaging station, high resolution camera and Imago software. These photos are stored on the cloud for ease of sharing and future viewing in 3D modelling software.

All DD core is oriented and where orientation is not possible the core is assembled with previous runs, where possible, to extend the orientation line.

A dedicated geotechnical logging team digitally captures detailed geotechnical logging using tablets for all OP and UG drill core, not just for holes drilled specifically for geotechnical assessment. Since 2018 logging data is synchronised with the main database at the end of the shift.

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All Excel drillhole log sheets are imported directly into the database. Direct digital capture of geological logging was tested, but loss of sketched structures and issues with ease of re-logging and collaboration mean that paper logs are preferred.

Sampling

DD core samples selected are usually between 0.8 m and 1.2 m long. The drill core is split in half along a cutting line (CL), 10° clockwise from the orientation line (OL), using diamond saws utilising fresh water. When looking down hole, the right-hand side half core is submitted for primary assay. Quarter core is submitted when taking a field duplicate to ensure that there is some core preserved in the box. However, as of the end of 2021, half core will be used for field duplicates.

All remaining core is stored for future reference.

10.6

Reverse Circulation Drilling

RC is only used at surface, primarily to infill gaps and improve grade confidence (Advanced Grade Control) and ultimately provide infill grade control ahead of open pit mining.

RC chip samples are logged with the same lithological, mineralogical and alteration information as DD core but are logged on regular 2 m RC sample intervals split through a riffle splitter.

Recent RC drilling has generally been completed by Boart Longyear and Ore Zone, with smaller amounts completed by local contractors AMAZONE (now TTS) and BMS. RC holes typically use 131 mm diameter rods with a 5.5-inch face-sampling bit.

RC recovery measured by weighing the total weight of the sample collected over a meter drilled and comparing it to the theoretical expected weight for each material type (lithological unit) and weathering type. RC recovery is good with an average of 94.6% recovery in the unweathered rock, 91.5% recovery in the transitional zone and 81.6% in saprolite zone. Average mineralised interval recovery was 89.2% with a range of between 76.4% and 100%.

Drilling Procedure

A project geologist and sampling technicians must be on site prior to drilling. They will ensure that the drill rig is lined up as per the drill plan, supervise the drilling contractor and carry out manual sampling exterior to the cyclone, plus quality check all down hole surveying.

RC samples are collected in pre-numbered rice bags, arranged in numerical order away from the cyclone area. Samples are weighed to allow estimation of recovery, which is recorded in a sample book. After homogenisation and splitting, chips are sieved from the reject material and collected in chip-trays labelled with the hole ID, depth interval and the sample number. The samples and chip trays are then transferred to the core yard facility by Kibali Goldmines staff for logging and sampling.

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Reverse Circulation Logging

RC chips are logged in the field by the site geologist. Geological logging is completed digitally using Maxwell LogChief installed on tablets that captures weathering, grain size, mineralisation, alteration style, lithology, and redox data, for each 1 m run interval.

Sampling

RC samples are collected from the rig in fixed 2 m intervals using an external Gilson splitter. The total mass is collected from the cyclone in 1 m run intervals, split by 50% to reduce manual handling. Two consecutive runs are combined to be mixed and further homogenised twice through a splitter. This mass is split three further times to a final 6.25% mass that gives a 3 kg to 4 kg sample. Auxiliary booster units are used to ensure that most of the samples collected are already dry. On the rare occasion a wet sample is obtained, it is dried before being manually split.

10.7

Twin Drilling Studies

Twin drilling studies are regularly undertaken as part of the Mineral Resource Management programme at Kibali. During 2021, twinning took place at Pamao, Pamao South, and Oere for grade confirmation and geotechnical assessment, as well as for metallurgical test work (Table 10-2). Comparisons of twin holes have shown that although there can be local variations in grade, as expected from the nugget effect at each deposit (which ranges from 15% to 35% of total variance), the broad intercepts and relative grade of the intersections are comparable across the twin holes.

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Table 10-2 Summary of Drill Hole Twinning in 2021


DD Hole ID	DD Interval		DD Interval Length (m)	RC Twin ID	RC Interval		RC Interval Length (m)	Zone	DD Fire Assay (g/t Au)	RCH Fire Assay (g/t Au)	Lode	Year
DD Hole ID	From	To	DD Interval Length (m)	RC Twin ID	From	To	RC Interval Length (m)	Zone	DD Fire Assay (g/t Au)	RCH Fire Assay (g/t Au)	Lode	Year

PMDD064	63.00	72.00	9.00	PMGC0297	68.00	72.00	4.00	Pamao	1.01	0.71	2001	2021
PMDD064	79.00	92.00	13.00	PMGC0297	82.00	96.00	14.00	Pamao	1.58	1.31	2002	2021

PMDD065	52.75	64.50	11.75	PMGC3888	44.00	70.00	26.00	Pamao	1.08	1.10	2002	2021

PMDD067	49.60	71.50	21.90	PMGC0352	52.00	70.00	18.00	Pamao	1.10	1.11	2002	2021

TDD011	-	13.40	13.40	TBGC0196	-	12.00	12.00	Pamao South	2.35	8.08	1002	2021
TDD011	23.00	31.00	8.00	TBGC0196	20.00	30.00	10.00	Pamao South	1.50	3.48	1003	2021

TDD012	1.50	5.40	3.90	TBGC0227	-	4.00	4.00	Pamao South	1.24	0.69	1001	2021
					12.00	22.00	10.00	Pamao South	-	1.22	1001	2021
	35.25	46.00	10.75		36.00	48.00	12.00	Pamao South	0.67	1.10	1002	2021
	61.00	77.40	16.23					Pamao South	1.23		1002	2021
					72.00	76.00	4.00	Pamao South		0.73	1003	2021

TDD013	-	35.90	35.90	TBGC0267	-	36.00	36.00	Pamao South	2.35	3.15	1002	2021

TDD014	30.00	38.00	8.00	TBGC0280	28.00	38.00	10.00	Pamao South	0.62	0.64	1002	2021
	46.00	54.00	8.00		60.00	66.00	6.00	Pamao South	6.49	1.57	1002	2021
	60.00	67.20	7.20					Pamao South	0.99		1003	2021

TDD015	64.70	69.00	4.30	TBGC0294	64.00	72.00	8.00	Pamao South	1.66	7.55	1002	2021
TDD015	86.00	88.00	2.00	TBGC0294	86.00	96.00	10.00	Pamao South	1.01	3.75	1002	2021

TDD016	84.00	88.65	4.65	TBGC0304	68.00	86.00	18.00	Pamao South	11.33	3.95	1002	2021

TDD017	93.00	107.00	14.00	TBGC0325	74.00	92.00	18.00	Pamao South	0.85	1.22	1002	2021
TDD017				TBGC0325	98.00	106.00	8.00	Pamao South		1.01	1002	2021

ORDD0016	72.00	78.00	6.00	ORRC0020	56.00	78.00	22.00	Oere	5.27	1.37	1001	2021

ORDD0018	63.00	67.00	4.00	ORRC0024	50.00	68.00	18.00	Oere	1.32	1.27	1001	2021
ORDD0018	84.00	94.50	10.50	ORRC0024	82.00	88.00	6.00	Oere	2.19	3.99	1001	2021

ORDD0019	10.50	31.50	21.00	ORGC0097	-	20.00	20.00	Oere	1.54	1.27	1001	2021

ORDD0020	83.10	93.00	9.90	ORGC0102	96.00	110.00	14.00	Oere	1.26	1.31	1001	2021
ORDD0020	101.63	116.90	15.27	ORGC0102	116.00	124.00	8.00	Oere	0.80	1.32	1001	2021

ORDD0021	3.40	18.20	14.80	ORGC0131	6.00	24.00	18.00	Oere	1.63	1.15	1001	2021
ORDD0021	23.00	34.55	11.55	ORGC0131	26.00	28.00	2.00	Oere	1.70	1.03	1001	2021

ORDD0022	79.00	96.00	17.00	ORGC0151	74.00	86.00	12.00	Oere	1.31	0.66	1001	2021
ORDD0022	107.40	115.00	7.60	ORGC0151	100.00	114.00	14.00	Oere	2.29	2.23	1001	2021

ORDD0023	109.00	114.50	5.50	ORGC0075	106.00	112.00	6.00	Oere	0.90	1.04	1001	2021
ORDD0023	116.50	146.00	28.40	ORGC0075	122.00	160.00	38.00	Oere	2.50	4.09	1001	2021

ORDD0024	51.50	63.00	11.50	ORGC0238	42.00	54.00	12.00	Oere	1.67	2.08	1001	2021

RDD0025	39.50	51.00	11.50	ORRC0007	44.00	58.00	14.00	Oere	1.44	0.73	1001	2021

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10.8

Drill Spacing Optimisation

Drilling directions for all open pit Mineral Resources are optimised on an individual deposit basis to ensure that the preferred drilling direction for Advanced GC and Infill GC drilling is on a cross plunge basis.

The data distribution or ‘drill campaign’ is one of several classification specifications for the resource estimate.

Measured classification infill GC drill spacing has been independently optimised using closely spaced variance drilling grids, supported by change of support analysis. In general, the infill drill spacings range between 10 m to 20 m along the principal direction and 5 m to 10 m across strike within the ore zones and are sampled at 2 m downhole intervals.

Indicated classification advanced GC drill spacing has been optimised using change of support analysis. In general, AGC is spaced approximately 40 m by 40 m with geological continuity of 100 m or more along strike. All open pit Mineral Resources that also form Mineral Reserves, namely KCD, Pakaka, Pamao, Gorumbwa, Sessenge, Megi-Marakeke-Sayi, Kalimva-Ikamva, Aerodrome, and Oere have been drilled to an advanced GC spacing.

Inferred classification Mineral Resource drill holes are generally on an 80 m to 100 m by 80 m or less drill spacing.

All drilled holes are composited to 2 m down hole during resource estimation; this is supported by a sample interval optimisation study completed that showed 2 m to be optimal for sampling within the Kibali Exploitation Permits.

10.9

Independent Audits

Independent audits are completed on a regular basis.

An external audit on the drilling procedures was previously completed in August 2017 by Optiro, which identified good performance for data collection and a high performance for data quality (Optiro, 2018a).

In September 2021, RSC Ltd (RSC) completed an independent audit of the Mineral Resource and Mineral Reserve processes used at Kibali (RSC Ltd, 2021). This included the drilling procedures used to collect the data informing Mineral Resource estimates. The audit demonstrated that Mineral Resource and Mineral Reserve processes conform to good practices. However, RSC made a number of recommendations to Kibali Goldmines for continual improvement including a review of RC drilling and sampling practices, particularly concerning testing of alternative splitters to provide better sample quality.

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10.10

Discussion

In the QP’s opinion, the drilling and sampling procedures at Kibali are robust, suitable for the style of mineralisation and are at, or above, industry standard practices. There are no drilling, sampling, or recovery factors that could materially impact the accuracy and reliability of the results.

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11	Sample Preparation, Analysis and Security

11.1

Sample Preparation

All samples submitted for assay are prepared and analysed at the SGS Doko laboratory, which is independently managed by SGS but is located at the Kibali site for exclusive use by Kibali Goldmines. The laboratory is operated using ISO/IEC 17025:2005 for testing and calibration, and ISO 9001:2015 for quality management.

Grade control and exploration drill samples are prepared in the same manner. Once the samples are received by SGS Doko, the sample is weighted and entered into a Laboratory Information Management System (LIMS). Samples are dried in an oven at 105°C. Channel and trench samples are disaggregated to remove dry lumps. All dried samples are crushed to ensure that 75% of the sample is below 2 mm.

The crushed sample is then passed through a Rocklab BOYD crusher with auto rotary splitter and the 75% reject material is retained. The 25% split sample is then pulverised in an LM2 pulveriser until 85% passes through a 75 µm (200 mesh) screen and after mat rolling, approximately 350 g is spooned into a packet. The LM2 pulveriser is cleaned with an air hose every sample, and with blank material every sixth sample. SGS Doko undertakes regular screen sieve tests on the crushing and pulverising. The coarse (2 mm) reject and the pulp (75 µm) reject material are returned to Kibali for storage at the mine site for future re-analysis, if required. Details on security measures taken to ensure validity/integrity of samples and the relevant chain of custody are documented within 11.4.

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Figure 11-1 outlines the preparation and analysis flow chart for DD core samples.

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Figure 11-2 outlines the preparation and analysis flow chart for RC samples.

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Figure 11-3 outlines the preparation and analysis flow chart for channel samples.

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11.2

Sample Analysis

All samples are analysed at the independently operated SGS Doko laboratory or SGS Mwanza, Tanzania. SGS Mwanza is used for sample overflow and analyses that could not be completed at SGS Doko including multi element, arsenic for selected samples and soils analysis. SGS Mwanza is an independent laboratory, accredited ISO / IEC 17025. ALS Johannesburg is used as an independent umpire laboratory. ALS Johannesburg is ISO 17025 accredited by the South African National Accreditation System (SANAS).

All samples are analysed using lead collection 50 g fire assay with atomic absorption finish with a gravimetric finish for any samples reporting above 100 g/t Au.

Results discussed include samples from exploration, resource evaluation, and both open pit and underground grade control. A total of 261,940 samples were submitted in 2021. Approximately 27% of the total samples received are check samples inserted into the sample streams (Table 11-1). Check samples consist of field duplicates for RC, pulp duplicates for diamond cores, certified referenced materials (CRM) and coarse blanks.

Table 11-1 Submitted Samples



Sample Type	Number of Samples	Percentage

DD	110,774	42%

RC	80,115	31%

Others	290	0%

Subtotal	191,179	73%



Standards	8,304	3%

Coarse Blanks	8,545	3%

Pulp Blanks	5,821	2%

Spiked Blanks	1,157	0%

Field Duplicates	8,575	3%

Coarse Reject Duplicates	8,471	3%

Pulp Reject Duplicates	6,757	3%

Pulp Resubmitted	10,605	4%

Umpires	12,526	5%

Subtotal	70,761	27%



Total	261,940	100%

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11.3

Quality Assurance and Quality Control

Kibali has an extensive QA/QC programme in place. This section covers the QA/QC from 01 October 2020 to 30 September 2021 (the review period) for exploration, resource evaluation and grade control assay data. Previous QA/QC reporting periods have not been observed to contain any significant sources of error or bias which would have a material effect on the Mineral Resource.

Quality Assurance (QA) is the protocols and procedures used and deemed to be appropriate and optimal for the Kibali district, based on orientation study at the beginning of the project, but regularly reviewed and updated.

Quality Control (QC) is using control samples and other measurements in real-time monitoring, using statistical analysis to ensure that the assay results are reliable, and that the sampling system is ‘in control.’

Quality control checks are inserted into the sample stream prior to dispatch to the laboratory, except for coarse and pulp duplicates, which are taken as a split by Kibali staff in the laboratory using a rotary splitter after crushing, or from the pulp reject after mat rolling. Overall, the QA/QC sampling includes 10% duplicates, 6% blanks, and 3% CRM. Independent umpire laboratories are also used on a quarterly basis to verify the primary laboratory, as well as to check the consistency in sampling protocols.

During 2021, 34,408 duplicates (RC field duplicates, core duplicates, coarse and pulp rejects, re-submitted pulp duplicates and umpire repeats), 15,523 blanks, and 8,304 CRMs were included within the routine sample submission.

All laboratories undertake their own internal QA/QC which includes blanks, duplicates, and CRMs, which are reported alongside Kibali Goldmines’ results. The results of the laboratory internal QA/QC are reviewed monthly with the Kibali team but are not included below.

The Kibali QA/QC protocol flowchart is illustrated in Figure 11-4.

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Certified Reference Materials (CRMs)

CRMs or ‘standards’ are inserted into batches at a frequency of 1 in 20 (5%) samples to check for bias over time and to test for laboratory handling errors. These monitor the accuracy of results received from the laboratory by comparing against the certified reference value.

All CRMs used in the review period are sourced from Ore Research and Exploration Pty Ltd, Australia, and are oxide or sulphide type with a matrix of feldspar minerals, basalt, and iron pyrite. CRMs are purchased in pre-packaged 50 g samples that require no preparation before being submitted to the laboratory. A sub-set of the total CRMs available are used and rotated on a quarterly basis to reduce laboratory identification.

CRM results are monitored and classified as a failure if one sample point falls outside of three standard deviations from the certified mean, or three consecutive samples fall outside of two standard deviations (on the same side) of the mean.

CRM results that have a failure outside of three standard deviations are checked for possible CRM swaps. This is investigated by comparing the returned assay grade to the list of known CRM grades values. The CRM samples are supplied by OREAS with the CRM ID printed on the bag. This printed ID is photographed during CRM insertion and then removed prior to submission of the CRM to the laboratory. This CRM photograph is used to help identify CRM swaps. A normal sample swap is also investigated to check if a normal drill sample has been incorrectly labelled as CRM.

In addition to the CRM photographs, swaps can be investigated using the technician’s sampling plan document, verifying used sample numbers, reviewing the sample booklet, and comparing against the other CRMs in the batch. When all the above investigations are complete, and it has been established that a genuine failure has occurred, the following actions are initiated:

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When two or more CRMs fail in batch and the failure is a result of sample swap, the entire batch is called for re-assay.

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When one or more CRMs fail in batch and the failure is not a result of sample swap, the entire batch is called for re-assay.

Based on the above controls when a batch is re-assayed and fails again, the samples are flagged but committed into the database whilst new samples are prepared for re-analysis. If a CRM is observed as repeatedly failing over a period, then it is removed from storage and is no longer used.

A total of 8,304 commercial standards have been submitted to SGS Doko during the reporting period. Table 11-2 lists all CRMs analysed at SGS Doko with their respective minimum and maximum control limits, as based on 3 standard deviations (3SD).

Overall performance across all standards indicates approximately 91% of the CRMs from SGS-Doko are within 2SD tolerance and approximately 98% are within 3SD tolerance. However, the vast majority of follow-up re-assay sampling did not return poor or questionable results.

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Table 11-2, Table 11-3, and Figure 11-5 summarise the performance of the CRMs relative to their upper and lower limits.

Table 11-2 List of CRMs Assayed at SGS Doko



Standard ID	Expected (g/t Au)				Assayed (g/t Au)				No. of Samples	First Used	Last Used
	Value	Std Dev	Min	Max	Min	Max	Mean	Std Dev

OREAS210	5.49	0.15	5.04	5.94	5.31	5.7	5.50	0.15	25	25/12/2020	30/08/2021

OREAS220	0.87	0.02	0.81	0.93	0.82	0.93	0.88	0.02	30	13/10/2020	27/08/2021

OREAS222	1.22	0.03	1.12	1.32	1.12	1.33	1.23	0.03	884	01/10/2020	30/08/2021

OREAS228	8.73	0.28	7.89	9.57	8.23	9.03	8.64	0.28	39	14/10/2020	27/08/2021

OREAS228b	8.57	0.20	7.97	9.17	7.52	9.24	8.64	0.20	2169	01/10/2020	30/09/2021

OREAS229b	11.95	0.29	11.09	12.81	12.4	12.9	12.72	0.29	5	26/11/2020	22/12/2020

OREAS232	0.90	0.02	0.83	0.97	0.81	1.03	0.90	0.02	3337	01/10/2020	27/09/2021

OREAS250	0.31	0.01	0.27	0.35	0.27	0.35	0.31	0.01	233	01/10/2020	08/09/2021

OREAS254	2.55	0.08	2.32	2.78	2.37	2.81	2.55	0.08	1582	03/10/2020	30/09/2021

Table 11-3 CRM Summary for Review Period at SGS Doko



Standard ID	Minimum Assay (g/t Au)	Maximum Assay (g/t Au)	No of Samples	%Pass of -/+1STD	%Pass of -/+2STD	%Pass of -/+3STD

All CRM	0.27	12.9	8,304	62.5%	91.1%	98.3%

Tramline analysis shown in Figure 11-5 is used to view any longer term trends and identify possible CRM sample swaps with regular samples inserted into the sample stream. This has been grouped into low-grade, medium-grade and high-grade ranges that cover the typical grade profile at Kibali.

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Blanks

Blank samples are assayed to help ensure no false-positives are obtained from laboratory analysis, checking for contamination during sample preparation, or to detect analytical contamination. These samples should return gold assay values below the analytical detection limit (i.e., <0.01 g/t Au). The coarse blank samples used originate from barren granite material sourced from Matiko and Kalimva, approximately 20 km NW of the project area.

During the collection of samples, blank sample materials are inserted at a rate of approximately 1 in 20 (5%) of the total submitted samples. These samples undergo the same sample preparation as the drill samples to detect inter-contamination due to poor cleaning of sample preparation equipment throughout the various sub-sampling processes.

Coarse Blanks

A total of 8,545 coarse blank samples have been submitted to SGS Doko. The results are evaluated against twice the standard deviation as an acceptable limit. The overall performance shows more than 98.7% of the blank samples assayed fell within the 2SD (Table 11-4 and Figure 11-6). In the opinion of the QP the overall performance of coarse blanks shows sample preparation is good.

Table 11-4 Statistics for Coarse Blank Samples



Assayed Values (g/t Au)				No Samples	Expected Value (g/t Au)	StDev (g/t Au)	% Pass			Bias
Min	Max	Mean	SD				-/+1SD	-/+2SD	-/+3SD

0.005	0.18	0.02	0.02	8,545	0.03	0.02	50.1	98.7	99.9	-44.84

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Figure 11-6 Performance Graph of Coarse Blanks

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Pulp Blanks

A total of 5,821 certified pulp blanks OREAS22f have been assayed and evaluated against two times the standard deviation as the acceptable limit. More than 99% of the samples analysed are within 2SD (Table 11-5 and Figure 11-7). In the opinion of the QP the overall performance of pulp blanks shows analytical contamination is not an issue.

Table 11-5 Statistics for Pulp Blanks (OREAS22f)



Assayed Values (g/t Au)				No Samples	Expected Value (g/t Au)	StDev (g/t Au)	% Pass			Bias
Min	Max	Mean	SD				-/+1SD	-/+2SD	-/+3SD

0.005	0.19	0.01	0.01	5,821	0.03	0.02	41.30	99.20	99.90	-58.81

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Figure 11-7 Performance Graph of Pulp Blanks

Spiked Blanks

A spiked blank is inserted in an occasional batch to test whether a laboratory is actively spotting them and making adjustments. Blanks are deliberately contaminated by mixing one pulp blank with a low-grade CRM in equal proportions. These contaminated samples are inserted and submitted to the laboratory blind. The laboratory has reported all samples correctly (Figure 11-8).

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Figure 11-8 Performance Graph for Spiked Blanks

Duplicates

Duplicate samples are primarily used to assess precision (repeatability) of the assay data and can also be used to assess for the presence of bias in the sample preparation chain, from each sample reduction stage. A duplicate sample is a second split from the original, prepared and analysed separately with a unique sample number, inserted after every 25^th sample and only in mineralised zones.

Duplicate samples are obtained from five sources, with the error being cumulative:

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Field Duplicate: a duplicate sample taken from the RC rig splitter or the second half of DD core, which quantifies the combined errors from field splitting through to analysis.

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Coarse (Reject) Duplicate: a duplicate sample off the crusher which quantifies a coarse crush splitting error and pulverising error through to analysis.

Typically, crusher and field duplicates are viewed using an absolute relative error of 20% (equates to ±10% precision level).

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Pulp (Reject) Duplicate: a duplicate sample after pulverising, which quantifies pulp sub-sampling and analytical error. Typically pulp duplicates range between 5 % to 10 % precision of the primary sample (preferably within ±5% of the primary sample).

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Pulp Repeat: a duplicate sample from the same pulp packet, submitted later and blind to the same laboratory, which quantifies the analytical error, but crucially can help identify bias trends over time (accuracy determination).

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Umpire: a duplicate sample from the same pulp packet, submitted later to an alternative laboratory, to independently confirm the accuracy of the primary laboratory.

RC and DD samples are reviewed separately to quantify and address the source of bias in duplicates more accurately.

The scatter plots in Figure 11-10, Figure 11-14, Figure 11-18, Figure 11-22, Figure 11-26, Figure 11-28 and other charts generally show all duplicate samples >0.5 g/t Au analysed at SGS Doko, to ensure they are relevant to the mineralisation.

RC Field Duplicates

A total of 1,747 RC field duplicate samples were analysed at SGS Doko within the period under review.

Statistical analysis of the results in Table 11-6 and the QQ plot in Figure 11-9 suggest that there is no significant bias between the original samples and field duplicates. Duplicate grades are noted to be slightly higher over 2.0 g/t Au.

Table 11-6 Statistics for RC Duplicates at SGS Doko



Statistics	Discrete Statistics			Percentile Statistics
Statistics	Original	Duplicate	Units	Distribution	Original	Duplicate	Units

Population	1,747			25.0%	0.90	0.85	g/t Au

Minimum	0.50	0.02	g/t Au	50.0%	1.00	0.97	g/t Au

Maximum	115.00	151.00	g/t Au	75.0%	1.26	1.26	g/t Au

Mean	3.24	3.42	g/t Au	80.0%	1.61	1.58	g/t Au

Std Dev	5.29	6.27	g/t Au	90.0%	2.19	2.20	g/t Au

CV	1.63	1.83	g/t Au	97.5%	2.98	3.08	g/t Au

Correlation	0.937			99.9%	3.49	3.62	g/t Au

The scatterplot in Figure 11-10 shows a strong correlation between the original and duplicate samples. There is moderate scatter around the 45° line and there is more variability at lower grades, but most results lay within the ±10% acceptable range. There are some obvious errors, highlighting where continuous monitoring of the sub-sampling processes and on-going review is required.

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Figure 11-9 Normal QQ Plot of RC Field Duplicates ≤ 100 g/t Au Tail Cut

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Figure 11-10 Normal Scatter Plot of RC Field Duplicates ≥ 0.5 g/t Au and ≤ 100 g/t Au Tail Cuts

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Figure 11-11 shows 90% of the samples have a precision that is less than 24% Half Absolute Relative Difference (HARD). This indicates that the precision of field duplicates could be improved. Follow-up re-assays also returned similar results.

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Figure 11-11 Ranked HARD Plot of RC Field Duplicates ≤ 100 g/t Au Tail Cut

Precision pairs plot, also known as a relative difference plot (Figure 11-12), shows that the 87% of RC field duplicates analysed at SGS Doko had sample pairs within a 20% difference of each other and 71% of these samples had samples pairs within 10% of each other, which is considered acceptable given the natural variability in the deposits.

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Figure 11-12 Precision Plot of RC Field Duplicates vs Original Sample at SGS Doko

Half-Core Field Duplicates

A total of 2,649 core duplicates were analysed at SGS Doko during the review period.

Table 11-7 summarises the statistical analysis of the core duplicates submitted to SGS Doko. The data suggests that there is no significant bias between the original samples and duplicates. Overall pair data is acceptable but can still be improved to reduce variability.

Table 11-7 Statistics of Half-Core Duplicates at SGS Doko



Statistics	Discrete Statistics			Percentiles statistics
Statistics	Original	Duplicate	Units	Distribution	Original	Duplicate	Units

Population	2,649			25.00%	1.40	1.39	g/t Au

Minimum	0.5	0.01	g/t Au	50.00%	1.58	1.59	g/t Au

Maximum	311	344	g/t Au	75.00%	2.23	2.21	g/t Au

Mean	6.58	6.66	g/t Au	80.00%	3.07	3.05	g/t Au

Std Dev	11.58	11.98	g/t Au	90.00%	4.38	4.63	g/t Au

CV	1.76	1.80		97.50%	6.44	6.38	g/t Au

Correlation	0.904			99.90%	7.62	7.95	g/t Au

The QQ plot presented in Figure 11-13 shows a strong correlation with an evenly distributed grade in both original and duplicates. There is no bias seen between original and duplicate samples and most of the samples are well within acceptable limits.

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Figure 11-13 Normal QQ Plot of HC Duplicates ≤ 100 g/t Au Tail Cut

Scatter plot analysis (Figure 11-14) shows a good correlation most samples falling within a 10% acceptable range. Some duplicate results report lower than originals, resulting in a trend line that is very slightly below the 1:1 line. Grade variability is consistent across the grade range.

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Figure 11-14 Normal Scatter Plot of HC Duplicates ≥ 0.5 g/t Au and ≤ 100 g/t Au Tail Cuts

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The ranked HARD plot (Figure 11-15) shows that 90% of the samples have a precision that is less than 20% HARD, indicating reasonable precision. However, this is at the top end of expectations, so efforts should be made to review the process with the technicians involved.

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Figure 11-15 Ranked HARD Plot of HC Duplicates ≤ 100 g/t Au Tail Cut

Precision plots (Figure 11-16) show that 90% of the coarse duplicates analysed at SGS Doko returned paired results within a 20% difference of each other and 73% of the coarse duplicates within a 10% difference. The overall performance is in line with the expected threshold and deemed acceptable.

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Figure 11-16 Precision Plot of HC Field Duplicates vs Original Sample at SGS Doko

Coarse Reject Duplicates

A summary of coarse reject duplicates analysed at SGS Doko is presented in Figure 11-17 to Figure 11-20. Table 11-8 presents a summary of the SGS Doko coarse reject duplicate statistics.

A total of 5,477 coarse rejects duplicates were analysed at SGS Doko during the review period.

Table 11-8 Statistics of Coarse Reject Duplicates at SGS Doko



Statistics	Discrete Statistics			Percentiles statistics
Statistics	Original	Duplicate	Units	Distribution	Original	Duplicate	Units

Population	5,477			25.00%	1.22	1.23	g/t Au

Minimum	0.5	0.02	g/t Au	50.00%	1.40	1.41	g/t Au

Maximum	503	452	g/t Au	75.00%	1.93	1.94	g/t Au

Mean	5.91	5.98	g/t Au	80.00%	2.58	2.61	g/t Au

Std Dev	12.21	11.90	g/t Au	90.00%	3.54	3.62	g/t Au

CV	2.07	1.99		97.50%	5.48	5.44	g/t Au

Correlation	0.984			99.90%	6.71	6.80	g/t Au

The QQ plot presented in Figure 11-18 shows a strong correlation with an evenly distributed grade in both original and duplicates.

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The scatter plot in Figure 11-18 shows only a few failures with most of the data following the trend line with minimal scatter. There is a slight increase in variability at lower grade.

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Figure 11-17 Normal QQ Plot of Coarse Reject Duplicates ≤ 100 g/t Au Tail Cut

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Figure 11-18 Normal Scatter Plot of Coarse Reject Duplicates ≥ 0.5 g/t Au and ≤ 100 g/t Au Tail Cuts

Figure 11-19 shows 90% of the samples have a precision that is less than 14% HARD, indicating that the precision of coarse reject duplicates is reasonable.

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Figure 11-19 Ranked HARD Plot of Coarse Reject Duplicates ≤ 100 g/t Au Tail Cut

Precision plots (Figure 11-20) show that 95% of the coarse duplicates analysed at SGS Doko returned paired results within a 20% difference of each other and 84% of the coarse duplicates within a 10% difference. The overall performance is in line with the expected threshold and deemed acceptable.

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Figure 11-20 Precision Plot of Coarse Reject Duplicates vs Original Sample at SGS Doko

Pulp Reject Duplicates

A total of 5,146 pulp reject duplicates were analysed at SGS Doko during the review period.

Table 11-9 summarises the statistical analysis of the pulp reject duplicates submitted to SGS Doko.

Table 11-9 Statistics of Pulp Reject Duplicates at SGS Doko



Statistics	Discrete Statistics			Percentiles Statistics
Statistics	Original	Duplicate	Units	Distribution	Original	Duplicate	Units

Population	5,146			25.00%	1.26	1.25	g/t Au

Minimum	0.50	0.02	g/t Au	50.00%	1.48	1.47	g/t Au

Maximum	503.00	382.00	g/t Au	75.00%	2.02	1.99	g/t Au

Mean	6.21	6.15	g/t Au	80.00%	2.74	2.76	g/t Au

Std Dev	12.58	11.32	g/t Au	90.00%	3.82	3.90	g/t Au

CV	2.03	1.84		97.50%	5.88	5.76	g/t Au

Correlation	0.976			99.90%	7.18	7.14	g/t Au

The QQ plot presented in Figure 11-21 shows a strong correlation with an evenly distributed grade in both original and duplicates. There is no visible bias seen between original and duplicate samples.

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Figure 11-21 Normal QQ Plot of Pulp Reject Duplicates ≤ 100 g/t Au Tail Cut

The scatter plot in Figure 11-22 shows strong correlation with most samples falling within a 10% acceptable range. There are outliers where the duplicate value is lower than the original resulting in a trend line slightly below the 1:1 line.

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Figure 11-22 Normal Scatter Plot of Pulp Reject Duplicates ≥ 0.5 g/t Au and ≤ 100 g/t Au Tail Cuts

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Figure 11-23 shows ranked HARD plot of pulp reject duplicates. 90% of the samples have a precision that is less than 11% HARD, indicating that the precision of coarse reject duplicates is reasonably good.

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Figure 11-23 Ranked Plot of Pulp Reject Duplicates ≤ 100 g/t Au Tail Cut

Precision pairs plot (Figure 11-24) shows that 98% of the pulp reject duplicates analysed at SGS Doko returned paired results within a 20% difference of each other and 88% of the coarse duplicates within a 10% difference. The overall performance is in line with the expected threshold and deemed acceptable.

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Figure 11-24 Precision Plot of Pulp Reject Duplicates vs Original Sample at SGS Doko

Pulp Repeats – SGS Doko

Pulp repeat samples were re-submitted to SGS Doko for a second test for possible bias in assays received at time of original analysis.

3,783 pulp duplicate samples were submitted during the review period from different grade ranges, all showing good correlation between the original and duplicates analyses.

Table 11-10 summarises the statistical analysis of the pulp duplicates submitted to SGS Doko.

Table 11-10 Statistics of Pulp Re-Submissions at SGS Doko



Statistics	Discrete Statistics			Percentile Statistics
Statistics	Original	Duplicate	Units	Distribution	Original	Duplicate	Units

Population	3,783			25.0%	1.13	1.13	g/t Au

Minimum	0.50	0.01	g/t Au	50.0%	1.32	1.30	g/t Au

Maximum	251.00	264.00	g/t Au	75.0%	1.82	1.82	g/t Au

Mean	5.72	5.77	g/t Au	80.0%	2.51	2.50	g/t Au

Std Dev	10.80	10.90	g/t Au	90.0%	3.47	3.43	g/t Au

CV	1.89	1.89		97.5%	5.11	5.10	g/t Au

Correlation	0.984			99.9%	6.04	6.14	g/t Au

The QQ plot presented in Figure 11-25 shows very good correlation between all grade ranges.

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Figure 11-25 Normal QQ Plot of Pulp Re-Submissions ≤ 100 g/t Au Tail Cut

The scatter plot in Figure 11-26 shows good correlation between the original and duplicate sample. There is an insignificant number of outliers.

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Figure 11-26 Normal Scatter Plot of Pulp Re-Submissions ≥ 0.5 g/t Au and ≤ 100 g/t Au Tail Cuts

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Figure 11-27 shows ranked HARD plot of pulp repeat duplicates. 90% of the samples have a precision that is less than 15% HARD, indicating that the precision of pulp repeats is good, but could improve.

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Figure 11-27 Ranked HARD Plot of Pulp Re-Submissions ≤ 100 g/t Au Tail Cut

Precision pairs plot (Figure 11-28) shows that 95% of the pulp duplicates analysed at SGS Doko returned paired results within a 20% difference of each other and 80% of the pulp duplicates within a 10% difference. The overall performance is in line with the expected threshold and deemed acceptable. However, an increase in variability relative to the original pulp duplicates, indicates more care is required during preparation and submission of pulp repeats.

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Figure 11-28 Precision Plot of Pulp Re-Submissions vs Original Sample at SGS Doko

Umpire Analysis

ALS Johannesburg, South Africa is used as an independent umpire laboratory. Samples are submitted quarterly, along with CRMs, to check for and quantify bias between laboratories.

A total of 4,408 pulp duplicate samples were submitted across the grade range, showing no significant bias between SGS Doko and ALS (Table 11-11).

Table 11-11 Summary of Pulp Duplicates Analysed at ALS



Statistic	Discrete Statistics			Percentiles Statistics
Statistic	Original	Duplicate	Units	Distribution	Origin	Duplicate	Units

Population	4,408			25.0%	1.24	1.25	ppm

Minimum	0.50	0.10	ppm	50.0%	1.47	1.47	ppm

Maximum	290.00	295.00	ppm	75.0%	2.01	1.98	ppm

Mean	5.76	5.93	ppm	80.0%	2.68	2.68	ppm

Std Dev	11.10	11.42	ppm	90.0%	3.65	3.70	ppm

CV	1.91	1.93		97.5%	5.20	5.12	ppm

Correlation	0.969			99.9%	6.26	6.15	ppm

Figure 11-29 illustrates that there is no systematic bias above or below the 1:1 (45°) line. There are some very high-grades >10 g/t Au reported higher by ALS, but this may reflect the paucity of sample pairs at this grade and is not considered meaningful.

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Figure 11-29 Normal QQ Plot of Umpires SGS Doko vs ALS at 100 g/t Au Tails Cut

There is an even but wide spread of points either side of zero difference, though the majority are within +/- 20%. ALS generally reports higher grades above 10 g/t Au compared to SGS-Doko, but the majority are within 10% tolerance (Figure 11-30).

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Figure 11-30 Relative Difference Plot of Umpires SGS Doko vs ALS at 100 g/t Au Tails Cut

Precision plots (Figure 11-31) show that 90% of the umpire duplicates analysed at SGS Doko returned paired results within a 20% difference of each other and 75% of the umpire duplicates within a 10% difference. The overall performance is in line with the expected threshold and deemed acceptable. However, an increase in variability relative to the original pulp duplicates, indicates more care is required during preparation and submission of umpires.

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Figure 11-31 Precision Plot of Umpires by SGS Doko vs ALS

11.4

Sample Security

Samples are under secure observation by geologists from collection at rig, to processing at the site core yard, to delivery at the laboratory.

RC samples on the rig are bagged, tied with custom tags, weighed, and documented. The samples are stored in a secure warehouse facility. DD samples are stored in core boxes with the appropriate numbering and markings, at the core shed area.

Sample submission forms are completed and sent to the laboratory with the samples as part of the chain of custody. These are checked at the laboratory to ensure that all samples are received. Sample security relies on samples always being attended or locked in appropriate sample storage areas, prior to dispatch to sample preparation facilities.

Coarse reject samples from infill grade control are discarded immediately but are stored for two months for exploration and advanced grade control. Pulp rejects are discarded immediately if the deposit is actively mined but for deposits under exploration or resource evaluation pulps are stored until the area is mined.

They are stored in the core yard in a dedicated storage area, under clean and dry conditions to avoid contamination. The pulp sample boxes are catalogued with details such as dispatch

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number, laboratory job number, and sample from and to information on each box. Samples sent to SGS Mwanza are also kept in a secured samples yard.

Samples are analysed at the independently operated SGS Doko laboratory, located at Kibali, except on infrequent occasions when the laboratory has had short-term reduced operating capabilities. In these instances, samples are prepared onsite in SGS Doko and then the pulps are shipped for analysis at SGS Mwanza, Tanzania. The samples are securely and directly shipped by logistics partner TCFF to Entebbe and then onward by road to the laboratory. Umpire samples are shipped from Entebbe to the laboratory in South Africa via air by DHL.

11.5

Independent Audit

Independent audits on Mineral Resources and all supporting data including QA/QC programmes are completed on a regular basis with previous audits completed by QG Australia Ltd. (QG) in 2012 (Quantitative Group, 2013) and Optiro in 2017 (Optiro, 2018a).

In September 2021, RSC completed an independent audit of the Mineral Resource and Mineral Reserve processes used at Kibali (RSC Ltd, 2021). This included the sampling procedures used to collect the data informing Mineral Resource estimates. The audit demonstrated that Mineral Resource and Mineral Reserve processes conform to good practices. However, RSC made a number of recommendations to Kibali Goldmines for continual improvement including a review of RC drilling and sampling practices, particularly concerning testing of alternative splitters to provide better sample quality, which will be implemented by Kibali in 2022.

11.6

Discussion

The QP is of the opinion that the sample collection, preparation, analysis, and security used at Kibali are performed in accordance with best practice and industry standards, and are appropriate for the style of deposit.

The QA/QC procedures and management are consistent with industry standards and the assay results within the database are suitable for use in Mineral Resource estimation. The QP has not identified any issues that could materially affect the accuracy, reliability, or representativeness of the results.

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12	Data Verification

12.1

Historical Drill Hole Data Verification

Historical data constitutes 11% of the drill hole database. This data is used for exploration targeting but has been effectively superseded in by current drill holes within the declared Mineral Resources.

In general, twin holes completed to date have shown that assayed intercepts in historic holes are mostly repeatable. However, some twin holes have identified that the mineralised intercept is at a different depth down hole relative to the historic data, thereby indicating that either the down-hole survey or collar survey data of the historic data is not reliable. These limitations are widely recognised, so grade data from historic drilling is generally not included and these holes have typically been re-drilled during Mineral Resource evaluation.

The QPs do not consider historical data to have a material impact on the current Mineral Resource estimate.

12.2

Current Drill Hole Data Verification

All forms of Project data are securely stored in an industry standard Maxwell Geoservices (Maxwell) DataShed SQL database. Data must pass validation through constraints, library tables, triggers, and stored procedures prior to importing. Failed data is either rejected or stored in buffer tables awaiting correction. A full-time database administrator employed at site manages the database.

Daily, weekly, monthly, and quarterly backups are made and stored on a hard disk onsite and automatically stored on the cloud which is in the UK but can be accessed globally.

A custom MS Access front end application has been designed for data entry, reporting, and viewing via Open Database Connectivity (ODBC), which utilises the data validation procedures from the SQL database. All other geological and mining software databases on site use ODBC link to retrieve information from the DataShed SQL database.

Assay data is imported directly from assay certificates from the laboratory and validated. Only fully trained and authorised network users can upload laboratory data. Assay data is stored in a normalised format and multiple assays are stored for each sample. Ranking of different assay formats is performed automatically so that one assay result is displayed in the final table. Any change to the rankings in the assay table must be approved by the onsite Database Manager.

Downhole survey data is directly uploaded from an associated handheld unit to Reflex Hub, a cloud based database server where each hole is reviewed by the respective geologist. Once approved, survey data is directly integrated with the Kibali database under an initial temporary table using a customised integration key. After further validation, it is written to the final survey table.

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The data is reported to and reviewed by the QP on a weekly and monthly basis. The QP completes an additional data review prior to Mineral Resource estimation.

12.3

Independent Audit

An independent external database audit was completed by Maxwell in 2020 (Maxwell, 2020). Maxwell identified that the Mineral Resource data within the SQL database was in good order and only minor data issues were identified.

Continued training and mentoring are ongoing for the database administrators as recommended by Maxwell.

12.4

Discussion

In the QP’s opinion the data verification program, as well as the sample collection, preparation, analysis, and security procedures comply with industry standards and are adequate for the purposes of Mineral Resource estimation.

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13	Mineral Processing and Metallurgical Testing

13.1

Summary

Metallurgical test work has been conducted on representative samples from Kibali ore bodies starting with project initiation in 2006 and continuing to date as additional deposits are developed.

Metallurgical and mineralogical characterisation has informed the initial plant design criteria and ongoing process optimisation initiatives to maximise cost effective gold recovery from a reasonably complex and variable ore mix delivered to the plant. Test work has led to the following features being incorporated in the gold recovery process:

●

Centrifugal gravity concentrators in conjunction with flash flotation to recover gravity recoverable gold (GRG) early in the milling circuit.

●

In-line leach reactor to dissolve concentrated gravity gold facilitating a short pipeline to bullion dispatch of GRG (± 23% of total gold produced).

●

Processing fresh ore through conventional flotation to recover refractory gold bearing sulphide/arsenopyrite concentrate for fine grinding and high shear partial oxidation resulting in improved leach recovery and reduced cyanide consumption.

●

Processing free milling oxide/transition ore through conventional CIL minimising the occasional preg-robbing effect from natural carbon in ore.

There have been several test work programmes completed at Kibali. Test work programmes for some satellite deposits were completed after initial plant commissioning and others targeted characterisation. More recently, studies have been completed for Pamao, Kalimva – Ikamva, the 3000 and 5000 lodes, Sessenge-KCD gap, Aerodrome, and Megi-Marakeke-Sayi as part of the definition or validation of modifying factors for Mineral Reserves. A summary of the test work to date can be found in Table 13-1.

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Table 13-1 Summary of Test Work



Name of Programme	Laboratory	Report ID or Number	Publication Date

Metallurgical Test Work Including Risk Reduction and Variability Tests	AMTEC/Orway Mineral Consultants (OMC)	A12949 A TO D	2011

Bankable Feasibility Study¹	AMTEC/Orway Minerals Consultants/Senet Engineering (SA)	Senet KGM Feasibility Report	2010

Feasibility Study²	AMTEC (Now ALS)/Lycopodium Engineering	1329/16.15/1329-STY-002/S5-B	2007

Prefeasibility Study³	AMTEC (Now ALS)/Lycopodium Engineering	1329/16.15/1329-STY-001/S5-B	2006

Satellite Pits and Additional Work

Mengu Hill

Deportment of Gold in Mengu Hill feed and flotation products	AMTEL	Amtel Report 12/55	2013

Mengu Hill Test Work Summary (Appendices available with all details of sample selection and compositing strategies)	AMTEC/OMC	Report No. 8888 Rev 1	2012

Pakaka

Metallurgical Performance of the Pakaka Feed Blends in the CIL – Review Relative to Feasibility and Geomet Arsenic domains	Kibali Goldmines Internal Review and Geomet Report	Internal Report	2017

Laboratory flotation test work on Pakaka gold samples (also includes in APP reports work on mineralogy)	Outotec Research Finland	15142-ORC-T	2016

Gold deportment analysis of Pakaka major ore types	AMTEL	Amtel Report 14/14	2014

Gorumbwa

Metallurgical Test Work conducted upon samples from the Gorumbwa Project for Kibali	ALS Metallurgy (Formerly AMTEC)	Report No. A16184	2016

Gorumbwa Feasibility Study – Metallurgical Test Work Report	Kibali Goldmines Internal Review and Summary of all tests conducted – T. Mahlangu	Internal Report	2014

Gold Deportment in Gorumbwa ores by CN leach	AMTEL	Amtel Report 14/42	2014

Sessenge

Processing of three samples from the Kibali – Sessenge Pit according to the current Kibali flowsheet	Maelgwyn Mineral Services Africa	REP 18-008	2018

Kibali Met Laboratory Sessenge Geomet Work_2018	Kibali Geomet Internal Test Work and Review Report	Internal Report	2018

Deportment of Gold in Kibali Sessenge ores	AMTEL	Amatel Report 16/38	2016

Pamao

Pamao Gravity Test Work	Peacocke & Simpson	PS394A to F	2017

Pamao BRT and Arsenic Distribution	Kibali Geomet Internal Review Report	Internal Report	2017

Metallurgical Test Work – Pamao_2017	Kibali Internal Pamao Metallurgical Review – T. Mahlangu	Internal Report	2017

Metallurgical Test Work – Pamao extension & low recovery zone_2021	Kibali Internal Pamao Metallurgical Review – T. Kapotwe	Internal Report	2021

Ore characterisation – Pamao extension & low recovery zone_2021	AMTEL	Amtel report 21-51	2021

Kalimva – Ikamva

Metallurgical Test Work – Kalimva-Ikamva_2019	Maelgwyn Mineral Services Africa	Report N0. 19-059	2019

Metallurgical Test Work – Kalimva-Ikamva_2019	Kibali Internal Ikamva-Ikamva Metallurgical Review – T. Kapotwe	Internal Report	2019

Deportment of Gold in Kalimva & Ikamva ore _2019	AMTEL	Amtel report 19-39	2019

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Name of Programme	Laboratory	Report ID or Number	Publication Date

3000 lode & 5000 lode DP

Metallurgical Test Work – 3000 lode & 5000 lode DP_2020	Kibali Internal Metallurgical test work – T. Kapotwe	Internal Report	2020

Deportment of Gold in 3000 lode & 5000 lode DP ore _2020	AMTEL	Amtel report 20-41	2020

Sessenge-KCD GAP

Metallurgical Test Work – Sessenge-KCD Gap_2020	Kibali Internal Metallurgical test works – T. Kapotwe	Internal Report	2020

Aerodrome

Metallurgical Test Work – Aerodrome_2020	Internal Metallurgical test works – T. Kapotwe	Internal Report	2020

Megi-Marakeke-Sayi

Metallurgical Test Work_Megi-Marakeke-Sayi_2020	Maelgwyn Mineral Services Africa	Report No. 20-197	2020

Metallurgical Test Work_Megi-Marakeke-Sayi_2020	Kibali Internal Metallurgical Review – T. Kapotwe	Internal Report	2020

Deportment of Gold in Megi-Marakeke-Sayi ore _2019	AMTEL	Amtel report 20-50/20-51	2019

Notes

Randgold, 2010

Moto Goldmines Ltd, 2008

The extensive metallurgical test work campaigns demonstrate two distinct behavioural patterns, particularly in the oxides but sometimes in the sulphides. Some ore sources exhibit free-milling characteristics suitable for gold extraction by a conventional CIL metallurgical process. Other ore sources exhibit a degree of refractoriness, albeit never extreme, where straight cyanidation returns gold dissolutions in the region of 70%, which is too low for optimal plant operation. This refractoriness is invariably due to the presence of occluded gold particles within sulphide minerals. It has been determined that a finer grind will expose a portion of this additional gold for leaching, thus enhancing the recovery such that it exceeds 80%. In addition, many of the Kibali ore sources, exhibit a preg-robbing tendency, which points to the need for rapid carbon adsorption. Thus, the Kibali plant was designed to cater to these observations through two distinct processing circuits:

●

Free-milling ore sources – conventional CIL circuit.

●

Refractory ore sources – flotation circuit with ultra-fine-grinding (UFG) and dedicated intensive leaching of the concentrate generated. Float tails leaching is optional and dependent on profitability.

More detailed descriptions of the discrete metallurgical test work campaigns follow.

Open pit extraction variability had a lower average extraction of 83.7% for KCD.

The LOM average gold extractions are 89% excluding the leach tails with minimum and maximum recoveries of 78.4% and 96.4%, respectively.

The resultant strategy is to:

●

Maximise gold recovery into the flotation concentrate – less through increased mass pull, due to the capacity limitations imposed by the downstream concentrate treatment processes, in particular UFG, and rather by reagent suite optimisation including optimal and steady flotation operation.

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Maximise gold dissolution from the concentrate – mineralogical effects might have an effect, but regular diagnostic leach tests will help keep track and identify where the problems come from.

●

Additional residence time for concentrate can be provided by the CIL – pumpcell product is provided for the benefit of further gold dissolution in the larger tanks.

Gold Recovery

A detailed explanation of the test work programme, results and interpretation of these results are detailed in the Amtec Reports (A12949 Parts (A – D) Full Report, Table 13-1) covering the:

●

Metallurgical sample characterisation in terms of grade, mineralogy, and physical characteristics of crushing and grinding parameters.

●

Experimental procedures, collation, and analysis of leach test results.

●

Extraction variability tests as well as comminution variability tests.

13.2

Test Work Strategy & Sample Selection

Extraction

The physical and extraction sample selection and test work logic was developed by Lycopodium and used in the feasibility and optimised feasibility study for Moto Gold ores (Table 13-2). Extraction results presented in the figures below include the OFS results and extraction variability tests (Moto Goldmines Ltd, 2009).

A total of 136 drill hole composite samples, composited at 10 m to 12 m intervals, were subject to direct cyanidation. The test procedure involved milling the samples to 80% passing 75 µm, bottle roll leaching in the presence of oxygen at 40% solids, pH 10.5 and 0.2%w /v nominal strength of cyanide for 24 hours. Note that the Master Composite and extraction variability were selected for detailed metallurgical investigation based on the geological description of the oxidation state and not the metallurgical behaviour of the hole composite samples.

The results, as depicted in Figure 13-1 indicated significant spatial changes in the cyanidation response of the deposit. The scattered nature of the results indicated that certain samples logged as primary sulphide material responded very positively to direct cyanidation.

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Notes:

Blue Markers = Fresh, Green Markers = Transition, Red Markers = Oxide

Figure 13-1 Initial Hole Composite Dissolutions

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Table 13-2 Physical and Extraction Sample Selection and Test Work Logic



Select Drill Hole Metre Interval Intercepts from Site

Select Comminution Test Work Samples and Comminution Variability Samples

Select samples for mineralogical thin section investigation

Conduct JK Drop Weight tests, Apparent SG, Abrasion index, Bwi, Rwi, SMC tests and Levin open circuit grindability test work on selected samples.

Crush remainder (P100 2 mm), Mix, Split, Assay and Leach a sub-sample of each 10 m Interval for each hole to determine direct cyanidation characteristics of individual hole composites

Select Master Composite samples for Primary and Oxide material individually

Select Variability samples, both spatially and by rock type

Oxide Master Composite		Primary Master Composite

Head Assays		Head Assays

Mineralogical Investigation		Mineralogical Investigation

Grind optimisation and leach tests		Grind optimisation and flotation tests

Gravity gold recovery, including intensive cyanidation of gravity concentrates at “as received” and ultra-fine-grind		Gravity gold recovery, including intensive cyanidation of gravity concentrates at “as received” and ultra-fine-grind

Leach optimisation, including reagents, oxygen vs. air sparging, diagnostic analysis and retention time		Direct cyanidation tests, including reagents, oxygen vs. air sparging, diagnostic analysis and retention time

Flotation test work		Flotation reagent optimisation test work, including flotation tests in site water

Oxygen Uptake Rate determination		Bulk gravity separation and pilot flotation

Viscosity measurements at varying pulp densities		Flotation Tail

Flocculation and thickener test work		Head assays

Sequential Triple Contact CIP (Carbon-in-pulp) test work and Equilibrium Carbon Loading test work		Leach tests

Geochem analysis on leach tail		Viscosity measurements at varying pulp density

Cyanide Detoxification test work		Thickener and flocculation test work

		Geochem analysis

		Flotation Concentrate

		Head assays, true SG determination, mineralogical examination

		Ultra-fine-grind test work and leach optimisation, including reagents, oxygen vs. air sparging, and retention time

		Indicative oxidation test work: - Pressure Oxidation, - Roast Calcination, - Bio-oxidation, - Albion Process

		Oxygen Uptake Rate determination

		Viscosity measurements at varying pulp densities

		Flocculation and thickener test work

		Sequential Triple Contact CIP test work and Equilibrium Carbon Loading test work

		Geochem analysis on leach tail

		Cyanide Detoxification test work

Upon completion of the extraction test work, the Process Route is defined for Oxide and Primary Material

Subject Primary variability samples to optimal recovery conditions as determined for the Primary Master Composite material

Subject Oxide variability samples to optimal recovery conditions as determined for the Oxide Master Composite material

Subject Transition variability samples to optimal recovery conditions as determined for the Primary Master Composite material

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Besides the extraction variability samples, metallurgical test work was conducted on the risk reduction samples, for both oxide and sulphide/fresh samples. The data represented in Figure 13-2 gives the extraction variability for the primary process of gravity float – float concentrate leach with the exclusion of flotation tails. Samples OFS_UG 1 to 7 represent the underground samples for the OFS (Moto Goldmines, 2009).

The average extraction of all fresh samples, that is, open pit and underground excluding the leaching of tails, is 88.1%. Also included in the plots are the underground feasibility recovery (89.8%) and open pit feasibility recovery (86.1%). Except for the underground OFS samples, the extraction data is plotted as a function of the DD holes.

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Figure 13-2 Primary Extraction Excluding the Leaching of Flotation Tails

The underground OFS variability samples were composited according to Table 13-3; the results of which are not characteristic of each drill hole, but of the composite.

Extraction variability data including the leaching of flotation tails is shown in Figure 13-3. The resultant impact of leaching flotation tails can be realised in terms of the variance in recovery between the two process routes as well as other factors that include, gold recovery into the flotation concentrate and concentrate residue values. There is significant variance for the data, as presented (Figure 13-3).

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Table 13-3 Extraction Comparison – Underground Variability



Sample Number	Hole Number	Intercept		Oxidation State	Source	Extraction (Oxide Process Route) % of Total Au	Extraction (Primary Process Route) % of Total Au	Extraction (Primary Process Route plus Flotation Tail Leach) % of Total Au
		From (m)	To (m)
1	DDD228	507	528	Fresh	Lode 910	75.6	86.3	89.8
	DDD080	537	588
	DDD227	546	554
	DDD213	482	496
	DDD128	471	475
2	DDD207	474	488	Fresh	Lode 910	76.5	90.8	92.6
	DDD212	468	518
	DDD293	472	480
	DDD297	476	488
	DDD306	530	542
3	DDD297	474	494	Fresh	Lode 910	72.3	84.2	86.1
	DDD073	446	452
	DDD219	488	512
	DDD130	482	516
4	DDD225	544	546	Fresh	Lode 910	76.5	86.4	91.1
	DDD031	646	650
	DDD175	476	488
	DDD211	480	512
5	DDD271	472	492	Fresh	Lode 910	71.2	87.7	90.4
	DDD227	592	612
	DDD269	663	675
	DDD207	488	502
	DDD103	308	330
6	DDD129	542	570	Fresh	Lode 910	78.9	92.1	94.5
	DDD069	470	488
	DDD240	484	498
	DDD206	481	519
7	DDD070	480	488	Fresh	Lode 910	60.9	91.1	93.3
	DDD085	438	456
	DDD205	504	526
	DDD214	508	542
Master Composite				Fresh	Lode 910 & 920	76 .9	91.3	93

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Figure 13-3 Primary Extraction Variability Including the Leaching of the Flotation Tails

Although these samples are considered fresh, it is apparent that their leaching response differs. What is not clear is why some respond positively to the leaching of flotation tails, meaning that there are non-floating minerals such as silicate mineral that contain gold or alternatively there is depression of passivated minerals during flotation. Samples in Table 13-4 and Figure 13-4 were isolated for further analysis to define the reasons causing the variances. The variances were mainly identified to be as a result of the presence of slow floating materials and poor grind in the concentrate leach stages. To deal with slow floating materials, higher mass pulls were adopted while at the same time increasing the downstream ultrafine grinding circuit capacity. This also enabled the circuit to cope with higher concentrate throughput without compromising the grind and thus ensure stable concentrate leach.

Table 13-4 Isolated Samples for Further Analysis



No.	Drill Hole	From	To	Lithology	Lode	Variance

1	DDD011	249	263	acs(vag)	j-3000	12.77

2	DDD224	101	114	vag/cs	i-3000	7.72

3	DDD005	450	464	vag/tuff	ug-9000	9.45

4	DDD290	294	308	vag/cs	ug-9000	8.44

5	DDD211	546	560	vag/xtaltuff	ug-9000	7.45

6	DDD127	496	510	is/vag	uga-9000	6.91

7	DDD084	184	198	is/sbx	ia-5000	6.71

8	DDD195	150	164	vag	n-5000	4.88

9	DDD162	40	54	vag	m-5000	4.72

10	OFS_UG 4				Lode 920	4.7

11	2010 MC					4.61

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Figure 13-4 Extraction as a Function of Diamond Drill Holes

Figure 13-5 represents the variation of gold recovery into the flotation concentrate (top left), initial sample head grade (top right), concentrate grade (bottom left) and float tail and concentrate leach residue (bottom right) for the drill holes with relatively high variances between the two processing routes.

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Figure 13-5 Analysis of the Drill Hole Samples Exhibiting Large Variances

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Gold losses, hence, variances, between the two process routes can be defined in terms of (i) losses to the flotation tails, that is, poor flotation recovery or (ii) losses to the concentrate leach residue.

In the first case, the presence of non-floating or slow floating mineralised material in the ore leads to the observed gold losses and is the reason why leaching the flotation of tails provides an improvement in gold recovery. This requires an optimised flotation circuit. From the available data, low-grade ore samples such as DDD224 and DDD005, present poor flotation recoveries which transcends into poor overall recovery on the exclusion of the leaching of the flotation tails. The split into the float concentrate and tails is around 7:93 (7% mass pull), leaching of the low-grade flotation tails will compensate for any losses into the concentrate residue. The fundamental point lies with the maximisation of flotation recovery and subsequent dissolution of gold from the flotation concentrate.

The gold losses into the concentrate leach residue are either a function of inadequate reagents or residence time on treating high-grade flotation concentrates (DDD162) or mineralogical effects related to gold occlusion in the form of finely disseminated particles such that even ultrafine grinding does not liberate it.

The results in Figure 13-6 were extracted from the OMC report and include the oxide and transition material leach (red plots) results. These results are included in this analysis for comparison purposes.

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Figure 13-6 Gold Extractions Obtained for Various Extraction Variability Tests and Master Composite

Samples

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Extraction Results – Oxides, Transition, and Other Ores

The average recovery on this data is 89.9% with a minimum value of 78% and maximum of 99.12% (Figure 13-7). The direct leach tests results from the grade control samples are shown in Figure 13-8. The results have an average recovery of 90% with a range of 82% to 96%.

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Figure 13-7 Plots of Extraction Using the Primary Process for the Oxide Materials – KCD

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Figure 13-8 Direct Cyanidation of Grade Control Samples

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It is clear from the two sets of data that the strategy adopted for the treatment of the oxide/transition material of the KCD is sufficient to minimise any gold losses. The benefit of treating the transition material through the oxide route with the flash flotation component ensures that both sulphides and non-floating materials are treated in the UFG – leach and CIL, respectively.

The leach results for the gravity – direct cyanidation tests on the grade control samples are detailed in the Table 13-5.

Table 13-5 Direct Cyanidation Results



Sample No.	Assay Head (g/t Au)	Calc Head (g/t Au)	Solids Tail Value (g/t Au)	Gravity Recovery (%)	Dissolution (%)	Total Extraction (%)	Lime Cons (kg/t)	NaCN Cons (kg/t)

DCRC0049 4.0-1 4.0 m	10.3/10.7	10.2	1.37	27.12	59.49	86.61	0.56	0.51

DCRC00 37 22.0-32.0 m	18.6/18.2	20.3	0.91	17.22	78.3	95.52	0.68	0.55

DCRC0047 14.0-24.0 m	9.08/9.28	9.75	0.96	17.69	72.46	90.15	0.59	0.70

DCRC00 50 4.0-14.0 m	12/11.2	12	1.19	33.80	56.26	90.06	0.55	0.65

DCRC0008 26.0-36.0 m	5.27/4.81	5.36	0.23	25.21	70.5	95.71	0.46	1.46

DCRC0007 46.0-58.0 m	2.79/2.26	2.76	0.22	13.72	78.31	92.03	1.79	0.82

DCRC0047 4.0-14.0 m	3.56/3.7	3.6	0.51	15.42	70.42	85.84	0.83	0.76

DCRC0040 16.0-26.0 m	1.63/1.83	1.75	0.21	8.46	79.56	88.02	0.53	0.87

DCRC0046 4.0-14.0 m	1.8/1.76	1.75	0.31	10.11	72.14	82.25	1.23	0.67

DCRC000S 8.0-1 8.0 m	0.8/0.72	0.84	0.03	16.07	80.36	96.43	0.75	0.38

DCRC0013 68.0-78.0 m	0.56/0.44	0.58	0.05	3.45	87.93	91.38	0.87	1.26

Comminution Characterisation Tests – Oxides, Transition and Sulphide Ores

Kibali ore source bond work index (BBWi) numbers for fresh unweathered sulphide material are presented on Figure 13-9. These tests were conducted at a limiting screen of 106 µm since the targeted grind size is 75 µm. The Pakaka sulphide material BBWi was very high, which resulted in extra cost in terms of energy, steel balls, and liners. The BBWi of other material lies within the design BBWi. The average plant operating work index recorded in 2020 was 10.19 KWh/t and 10.78 KWh/t for Stream #1 and Stream #2, respectively, with mills specific energy consumptions of 10.75 KWh/t and 11.63 KWh/t. Mill products average P₈₀ were at 78 µm on Mill #1 and 80 microns on Mill #2 (Figure 13-10).

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Figure 13-9 DD Composite Samples BBWi (KWh/t)

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Figure 13-10 Kibali Processing Plant Average P₈₀and Specific Energy Consumption (2021)

Particle size reduction is critical to achieve targeted direct leach and flotation recoveries as shown on Figure 13-11 and Figure 13-12. These results were generated from fresh rock KCD, Gorumbwa, and Sessenge orebody samples process in the plant post-2018 reporting. A size-by-size flotation recovery analysis conducted on plant composite samples confirmed that higher performance was achieved between 75 µm and 53 µm.

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Figure 13-11 Met_OT Composite Samples: Leach Recovery vs P₈₀

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Figure 13-12 Kibali Processing Plant Composite Samples: Flotation Recovery by Particle Size Range

Gravity Amenability Tests: All Material Types

The original plant design for the gravity circuit had an estimated gravity recovery below 18%. However, actual plant performance has shown that recovery is consistently above 21%. One key change to the circuit has been to reconfigure the gravity circuit with the installation of a falcon centrifugal concentrator which primarily targets fine gold recovery. The feed to this unit is the flash flotation concentrate, which had been reporting to the final concentrate thickener creating unintended ultrafine grinding and leach inefficiencies.

As part of the evaluation and optimisation of new satellite pits and new underground stopes, onsite and external laboratory gravity recoverable gold test work has become an integral part of routine work at the Project.

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Mineralogical Assessment

Extensive mineralogical examination data exists for Kibali ores, primarily from the original feasibility work and more recently generated as part of pre-production geometallurgical studies for either new pits and/or new mining and feeding domains of the existing ore bodies. The primary objective of this work is to identify all forms and carriers of gold, assess mineralogical factors affecting gold recovery, and determine target recoveries along with opportunities to optimise. The data in Figure 13-13 includes such examples of pre-production work and the new pits such as Kalimva-Ikamva and Megi Marakeke-Sayi. The later pits are yet to be processed but have already been developed to understand the bulk mineralogy and identify possible recovery impact issues.

Kibali Goldmines submits composite samples to an independent external laboratory for a full gold deportment, especially in cases where lower than initially predicted recoveries are encountered. Examples of exposed residual gold grains accounting for more than 3% have been identified. These have surface build-up of silver+arsenate/Fe which interfere with gold dissolution. However, up to 75% of gold losses in the tailings is accounted for by the natural refractoriness of the ore in form of sub-microscopic gold in pyrite and arsenopyrite. The latter has been consistent and elevated in the satellite orebodies that carry significant content of arsenopyrite minerals and generally retain sub-80% recoveries, exemplified by high arsenic domains at Pakaka.

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Figure 13-13 Kibali Ore – General Mineralogy

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Ultra Fine Grinding

The deportment of gold determined for Kibali Goldmines orebodies confirmed that over 15% of gold is enclosed in sulphide mineral matrix, mainly in grains with sizes range between 25 µm and 13 µm. The original plant design for the UFG circuit defined an 18 µm size fraction but at a lower dissolution of less than 83%. Additional optimisation work in the plant post commissioning and better understanding of the different orebodies have further generated benefits in the gold dissolution profiles. The average P80 of re-grinded floatation concentrate achieved in 2020 was around 25.5 µm, the grinding energy required was 38.4 KWh/t for an average gold dissolution of 88.1%.

Aachen Assisted Preoxidation and Leaching

Sulphide minerals specifically display elevated oxygen uptake rate (OUR) patterns leading to increased demand in oxygen supply. The result is elevated oxygen demand which must be met where sulphide minerals are specifically targeted for oxidation rather than passivation, normally attributed to the pre-oxidation phase of the leach. If not efficiently carried out, the ongoing oxygen demand will suppress the leach kinetics downstream, often resulting in recovery losses. The Project operates a two stage Aachen assisted peroxidation followed by an extended leach system, with dissolved oxygen levels in excess of 15 ppm in the leach circuit (Figure 13-14). In the aggressively sulphide concentrate leach circuit, this provides both:

●

Additional liberation from partially oxidised pyrite, pyrrhotite and arsenopyrite, especially if the gold particles are found near interface areas of different minerals and

●

More predictable, stable leach kinetics with a degree of safety not leading to dips in recovery based on inconclusive dissolution within the given residence time.

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Figure 13-14 Kibali Metallurgical Composite Samples – Oxygen and Recovery Profile

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Operations

Processing of the new Aerodrome pit commenced in Q4 2021, and a geo-metallurgical work programme was conducted in 2020 as part of the Mineral Reserve conversion studies. Two significant geo-metallurgical zones have been identified within the main envelope (Figure 13-15). Zone 1 is located up plunge in the chemical sediments where recovery exceeds 90%. Zone 2 is at the bottom of the pit where gold recovery tends to drop to <70% coupled with an increase in arsenic content (>4000 ppm).

Based on the geo-metallurgical test work, a simplified model (Arsenic-Recovery) has been defined to establish a robust feeding strategy to avoid plant underperformance.

The oxide and transition ore recoveries have returned a positive response during the CIL simulation with over 90% recovery achieved. Fresh material picked up plunge also related to Zone 1 whose arsenic content was below 2,000 ppm, did not reflect any dissolution issue. Flotation and leaching performance confirmed their suitability to the Kibali Goldmines sulphide circuit, the overall recovery recorded was 89%. In Zone 2, where the arsenic content exceeds 4,000 ppm, returned a low recovery of 75.18%.

The blend simulation performed on high arsenic material with both KCD and Gorumbwa to mitigate the negative impact of arsenic in the system, established a maximum limit of 20% in the feed blend to maintain plant stability in terms of leaching efficiency and main reagent consumption.

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Figure 13-15 2020 Aerodrome BRT and Arsenic Distribution

Additional geo-metallurgical test work on KCD Open Pit Pushback 3, confirmed oxide recoveries and re-defined the reagent consumptions and blending strategies with the current suite of free milling ores from Pakaka and Kombokolo (Table 13-6).

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Overall, the current geo-metallurgical work is focused on closing gaps and continuous improvement to enhance efficiencies in terms of recoveries and reagent consumption. These have been built into the budget models to ensure that operating costs and efficiencies are managed live to reflect the current ore feed blends (Table 13-6).

Table 13-6 Metallurgical Recoveries per Deposit



Ore Source	Recovery
Ore Source	Oxide (%)	Transitional (%)	Fresh (%)

KCD	90.1	90.1	86.1

KCD UG	-	-	90.0

Sessenge	90.3	75.9	81.0

Pamao	90.9	85.0	85.0

Kombokolo	85.0	85.0	85.0

Pakaka	89.0	89.0	80.2

Mengu Hill	81.0	77.0	70.0

Gorumbwa	90.0	90.0	90.0

Kalimva-Ikamva	90.0	89.0	89.0

Aerodrome	90.0	88.0	85.9

Megi-Marakeke-Sayi	90.0	90.0	89.5

Pamao South	89.0	88.0	86.5

Oere	88.0	86.5	87.0

Variability of Results and Confidence Levels

The data available for the original feasibility metallurgical sampling and extraction test work was from KCD, Kombokolo, Mengu Hill, Pakaka, Pamao, and Sessenge. While all the samples have been tested, the selection of the process routes and subsequent plant design has been based on the results from KCD, which consists of 70% of the feasibility study ore feed to the plant. The most significant increase in tonnage is likely to come from the KCD deposit. The sampling strategy and classification of samples in the KCD area has followed, in principle, the process depicted in Figure 13-16.

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Figure 13-16 Sampling Strategy and Classification of Samples at KCD

The KCD Metallurgical Analysis and Extraction Data can be found in Table 13-7.

Table 13-7 KCD Fresh Open Pit Fresh Samples – Lode 5000



Hole ID	Sample ID	From (m)	To (m)	Weathering	UG/OP	Lode	Overall Extraction (excl float tail leach) %	Overall Extraction (incl float tail leach) %

DDD072	EV2009	120	130	Fresh	OP	5000	89.8	92.8

DDD257	EV2009	140	170	Fresh	OP	5000	81.5	83.6

DDD165	EV2009	86	96	Fresh	OP	5000	84.1	85.5

DDD165	EV2009	116	126	Fresh	OP	5000	81.8	82.9

DDD160	EV2009	90	100	Fresh	OP	5000	78.4	79.4

DDD195	EV2010	150	164	Fresh	OP	5000	90. 7	95.5

DDD162	EV2010	40	54	Fresh	OP	5000	92.0	96.7

DDD166	EV2010	96	110	Fresh	OP	5000	92.4	94.0

DDD164	Master Comp	92	110	Fresh	OP	5000	-	-

DDD455	Master Comp	113	148	Fresh	OP	5000	-	-

DDD424	Master Comp	54	89	Trans	OP	5000	-	-

Master Comp Final							87.3	92.4

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The selection was similarly performed for all ore bodies, either open pit or underground and based on redox state as oxides, transition, or fresh.

The process routes selection strategy ensures that fast floating material and gravity recoverable gold are recovered upfront with minimal size reduction requirements. Provision has been made for this in both process streams in the form of flash flotation and gravity concentration units. The benefits and possible shortcomings of this arrangement have been briefly discussed Section 17.1 of this report covering the strategy around handling oxide/transition material classification.

13.3

Metallurgical Recoveries

The sample selection for the ore bodies and metallurgical recoveries expected at Kibali and used in the financial model can be found in Table 13-8.

The samples have been selected by site geologists and metallurgists and, in the opinion of the QPs, are representative of the ore bodies across the Permits.

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Table 13-8 Summary of Average Recovery for All the Samples



Ore Source	Weathering	Average Recovery Primary Only (%)	Average Recovery Primary + Tail Leach (%)	Average Recovery Oxide Process (%)	Feasibility or Financial Model Recovery (%)

KCD	Fresh_OP	86.4	89.2		86.0
	Fresh_UG	89	93.4		90.0
	Transition	66.6	91.3		90.1
	Oxide			89.1	85.8

Sessenge	Fresh	72.7	81.2		81
	Transition		80.3		75.9
	Oxide			90.4	90.3

Pakaka	Fresh	78.1	82.3		80.2
	Transition				81.3
	Oxide			96.9	88.7

Mengu Hill	Fresh	69.2	72.2		70.1
	Transition	84.4	89.9		89.3
	Oxide			92.6	89.3

Kombokolo	Fresh	70.3	75.2		85.0
	Transition	78.9	95.3		95.9
	Oxide			96.4	95.6

Pamao	Fresh	74.5	85.5		85.0
	Transition				85.0
	Oxide			95.8	90.9

Kalimva-Ikamva	Fresh	89.38	93.64		89.0
	Transition			89.88	89.0
	Oxide			91.05	90.0

3000 lode DP_KCD UG	Fresh	88.36	89.56		89.4
	Transition
	Oxide

5000 lode DP_KCD UG	Fresh	78.58	88.03		89.5
	Transition
	Oxide

Aerodrome	Fresh	79.05	85.83		85.83
	Transition			88.96	88
	Oxide			89.96	89.5

Megi-Marakeke- Sayi	Fresh	87.37	90.33		89.5
	Transition			92.58	90.0
	Oxide			94.29	90.0

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13.4

Deleterious Elements

Kibali Goldmines needs to consider the remediation of cyanide species as well as arsenic.

The QP confirms that there are no processing factors that could have significant effect on potential economic extraction.

Cyanide

Kibali Goldmines abides by the guidelines of the ICMC to which both Barrick and AngloGold are formal signatories. The two cyanide tailings storage facilities (CTSFs) are both lined with a high density polyethylene (HDPE) liner. Protocols call for limited threshold discharges to the CTSF and cyanide discharge concentrations are controlled through use of an on-line cyanide analyser and controller. The presence of two CTSFs allows management of the cyanide containing liquor streams and moreover, most of the water is recycled to the plant area where there also exists an additional cyanide detoxification pond facility.

Aiming to mitigate the risk of long term International Cyanide Management Institute non-compliance and possible detrimental environmental impact of discharging high weak acid dissociable (WAD) cyanide levels in the CTSF tailings dam, Kibali evaluated the possibility of conducting peroxide detox versus the cyanide recovery process (CRP) developed by AZMET consultants, who designed the process flow sheet.

The in-plant peroxide detox trial with copper sulphate addition was performed on the CIL tails slurry with close monitoring using an online cyanoprobe WAD analyser. WAD Cyanide below 50ppm was achieved at a total detox cost estimated at $1.50/t during the oxide campaign and $0.31/t during the sulphide campaign. The full-scale plant detox has been budgeted for 2022 and half of 2023.

AZMET CRP pilot plant Phase 1 and Phase 2 trials were performed on the Pumpcell tails (PCT) and CIL tails slurry during both campaigns (full sulphide and oxide-sulphide), in which very high WAD cyanide levels were observed in the CIL tails slurry. Results showed that WAD CN levels below 20ppm were achieved on the CIL tails with an expected benefit due to additional gold and cyanide recovery. Analysis and review is ongoing for a potential $26.75 million capital expenditure project to build a 4.8 Mtpa plant, which will allow treatment of the full oxide-sulphide CIL tails stream with an OPEX cost estimated at $0.43/t. The AZMET CRP plant is slated for commissioning in Q2 2023.

Based on the Kibali LOM, the onsite test work, trade-off studies and financial analysis showed that the AZ-CRP is more effective and economic when compared to other detox methods such as the Peroxide Detox & INCO process. Another benefit of the AZMET CRP plant is the additional gold and cyanide recovery leading to an after-tax payback period of less than 4 years.

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Arsenic

The main deleterious element in the Kibali ore is arsenic. Certain isolated ore types exhibit higher levels of arsenic (for example Pakaka, Sessenge and Aerodrome), which can result in dissolution during the recovery process. During the recovery process, the arsenic dissolves into solution and is captured by the leach of flotation concentrate in the intensive oxygenation/cyanidation circuit.

Mitigation can occur for either of the cyanide containing streams or non-cyanide containing streams, that is, flotation tails, which report to a dedicated but unlined flotation storage facility (FTSF). Arsenic remediation can occur through oxidation of ferrous sulphate and arsenic species to the valency state (V). Alternatively, ferric chloride may be used directly, though is associated with corrosion issues. Both methods result in the formation of a stable ferric arsenate precipitate. The primary mitigation method utilised at Kibali is the application of a blending strategy where high arsenic content ores are intentionally blended with ores with low content, thereby restricting the arsenic solution tenors within the circuit.

Arsenic content more than 2,000 ppm has a negative effect on gold dissolution. Dissolution values as low as 70% are attained when arsenic content increases as high as 9,000 ppm.

Subsequently detailed geo-metallurgical analysis has been completed on Pakaka and Sessenge where the arsenic content has been modelled as part of the Mineral Resource block model.

Metrics have been developed for stockpiling and blending, to dilute and minimise the impact of high arsenic in the overall plant feed. Additional work was carried out to identify the poor recovery related to the refractory component of the ore, while the pre-oxidation processes of the concentrate post ultrafine grinding was controlled or restricted to minimise arsenic mobilisation to solution.

13.5

Discussion

Mineral processing and metallurgical testing fundamentals are well established at Kibali. The ore characterisation insights gained have contributed to achievement of ongoing relatively high consistent predictable gold recoveries.

In the opinion of the QP, the rigorous representative sampling and testing of new deposits provides a sound geometallurgical understanding of process requirements as mining activities advance.

Test work and gold recovery variability characterisation has in the QP’s opinion resulted in provision of considerable flexibility and rigor within the plant processes enable the operation to target and customise parameters appropriate for different ore types.

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14	Mineral Resource Estimates

14.1

Summary

Geological interpretation and Mineral Resource estimation were completed by Kibali Goldmines with an effective date of 31December 2021. Table 14-1 presents a summary of the Kibali Mineral Resource estimate, as of 31December 2021.

Table 14-1 Kibali Mineral Resources as of 31December 2021



Type	Category	Tonnes (Mt)	Grade (g/t Au)	Contained Gold (Moz Au)	Attributable Gold¹ (Moz Au)

Stockpiles	Measured	0.32	3.17	0.032	0.015

Open Pits	Measured	15	2.24	1.1	0.50
	Indicated	45	2.25	3.3	1.5
	Inferred	8.2	2.1	0.55	0.25

Underground	Measured	32	4.63	4.7	2.1
	Indicated	48	4.06	6.3	2.8
	Inferred	15	3.0	1.4	0.64

Total Mineral Resources	Measured	48	3.84	5.9	2.6
	Indicated	93	3.18	9.5	4.3
	Measured and Indicated	140	3.41	15	6.9
	Inferred	23	2.7	2.0	0.89

Notes:

Attributable Gold (Moz) refers to the quantity attributable to Barrick based on Barrick’s 45% interest in Kibali Goldmines. Mineral Resources are reported on a 100% and attributable basis.

The Mineral Resource estimate has been prepared according to CIM (2014) Standards and using CIM (2019) MRMR Best Practice Guidelines.

All Mineral Resources tabulations are reported inclusive of that material which is then modified to form Mineral Reserves.

Open pit Mineral Resources are reported within the $1,500/oz Au pit shell at a tonnage weighted average cut-off grade of 0.77 g/t Au.

Underground Mineral Resources in the KCD deposit are Mineral Resources, which meet a cut-off grade of 1.62 g/t Au and are reported in-situ within a minimum mineable stope shape, at a gold price of $1,500/oz Au.

The Mineral Resource estimates have been prepared according to the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) 2014 Definition Standards for Mineral Resources and Mineral Reserves dated 10 May 2014 (CIM (2014) Standards) as incorporated with National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101). Mineral Resource estimates were also prepared using the guidance outlined in CIM Estimation of Mineral Resource and Mineral Reserve Best Practice Guidelines 2019 (CIM (2019) MRMR Best Practice Guidelines).

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Definitions for resource categories used in this report are consistent with those defined by CIM (2014) and adopted by NI 43-101. In the CIM classification, a Mineral Resource is defined as ‘a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction’. Mineral Resources are classified into Measured, Indicated, and Inferred categories.

Underground Mineral Resources were reported using MSO, effectively within a minimum mineable stope shape, applying reasonable mineability constraints, including a minimum mining width, a reasonable distance from current or planned development, and a measure of assumed profitability at the related resource cut-off grade, thus deemed as having a reasonable prospects for eventual economic extraction.

The Kibali Mineral Resources consist of the KCD, Sessenge, Pakaka, Mengu Hill, Gorumbwa, Megi-Marakeke-Sayi, Pamao (inclusive of Pamao South), Kombokolo, Kalimva-Ikamva, Aerodrome, Oere, and Mengu Village deposits.

KCD (underground and open pit), Sessenge, Gorumbwa, Pamao, Aerodrome, and Mengu Village were updated following additional data and/or updated geological interpretations. The updates for each of these deposits are summarised as follows:

●

The KCD underground model update incorporates data from GC, ACG, and EXP drilling up until July 2021 for the 3000, 5000, 9000, and 11000 lodes. Mineral Resources are reported for the first time for the 11000 lodes.

●

Sessenge was updated in August 2021 with GC drilling. Mineral Resources for Sessenge include a small adjacent satellite deposit known as Sessenge SW.

●

The Gorumbwa deposit has been updated using GC and AGC, an updated void shape, and the additional data in the gap between this deposit and Sessenge.

●

Pamao was updated following GC drilling inside the $1,500/oz Au pit shell.

●

Aerodrome was updated in June 2021 following GC drilling within the $1,500/oz Au pit shell.

●

No new data were added to the Mengu Village deposit. The deposit was previously estimated in 2006 using the uniform conditioning (UC) method. New geological

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interpretations based on additional drilling data, collected during 2020, prompted a review and update of the geological model, which was re-estimated using the ordinary kriging method in 2021.

Models for actively producing deposits are updated on a quarterly basis to incorporate all additional grade control drilling results throughout 2021 with a budget model produced once a year for Mineral Resource reporting (Table 14-2).

Pamao South and Oere are new additions to the Kibali Mineral Resources based on AGC drilling up to 2021 (Table 14-2).

Megi-Marakeke-Sayi and Kalimva-Ikamva are unmined deposits where no significant drilling has been completed since 2020 and 2019 respectively (Table 14-2).

Mengu Hill, Kombokolo, and Pakaka are depleted deposits where no significant drilling has been completed since 2018 (Mengu Hill and Kombokolo) or 2019 (Pakaka).

Where appropriate, all models have been depleted using the December 2021 mined out stopes and surfaces.

Table 14-2 Summary of Deposits and Model Date



Deposit	Producing Status	Model Date

KCD Underground	Active	07/07/2021

KCD Open Pit	Active	07/07/2021

Sessenge	Active	11/08/2021

Sessenge SW	Unmined	11/08/2021

Gorumbwa	Active	22/07/2021

Aerodrome	Active	05/05/2021

Pamao and Pamao South	Unmined	30/11/2021

Mengu Village	Unmined	30/06/2021

Oere	Unmined	26/08/2021

Megi-Marakeke-Sayi	Unmined	15/08/2020

Pakaka	Depleted, $1,200/oz Au pushback in LOM	06/06/2019

Kombokolo	Depleted, awaiting 2022 drilling for UG	10/08/2018

Mengu Hill	Depleted	03/04/2018

Kalimva-Ikamva	Unmined	25/07/2019

14.2

Resource Database

KCD

During 2021, a total of 794 DD and RC drill holes were completed for a total of 105,253 m. The drilling at KCD consists of a combination of grade control at 10 m by 5 m spacing within high-grade shoot zones, 20 m by 5 m spacing within low-grade zones, 20 m by 20 m spacing in waste zones for open pit, and approximately 20 m by 15 m spaced diamond drilling for underground grade control.

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A summary of KCD data used for the current Mineral Resource estimate is presented in Table 14-3.

Table 14-3 KCD Drill Summary Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min Depth (m)	Max Depth (m)	Total Drilled (m)

Kibali Goldmines	2021	DD	458	15	1,729	79,676
Kibali Goldmines	2021	RC	336	7	198	25,577

Kibali Goldmines	2020	DD	502	27	709	87,343
Kibali Goldmines	2020	RC	409	6	120	17,927

Kibali Goldmines	2019	DD	470	30	930	92,446
Kibali Goldmines	2019	RC	589	6	132	25,935

Kibali Goldmines	2018	DD	575	25	682	102,388
Kibali Goldmines	2018	RC	460	10	156	29,714

Kibali Goldmines	2017	DD	643	16	1,491	111,688
Kibali Goldmines		GT	8	100	280	1,073
Kibali Goldmines		RC	615	12	150	32,833

Kibali Goldmines	2016	DD	501	12	565	64,661
Kibali Goldmines	2016	GT	14	66	427	3,227

Kibali Goldmines	2015	DD	250	45	464	41,793
Kibali Goldmines		GT	21	24	230	2,973
Kibali Goldmines		RC	872	6	110	35,751

Kibali Goldmines	2014	DD	109	26	800	19,329
Kibali Goldmines		GT	8	135	321	1,894
Kibali Goldmines		RC	1,769	6	800	87,877

Kibali Goldmines	2013	DD	32	16	801	13,010
Kibali Goldmines		GT	3	195	723	1,364
Kibali Goldmines		RC	1,336	12	220	70,700

Kibali Goldmines	2012	DD	24	80	1,092	11,027
Kibali Goldmines		GT	19	20	801	2,989
Kibali Goldmines		RC	1,778	0	150	93,526

Kibali Goldmines	2011	DD	15	9	1,347	4,991
Kibali Goldmines		GT	67	11	860	10,058
Kibali Goldmines		RC	1,616	4	150	51,263

Kibali Goldmines	2010	DD	58	25	942	27,166
Kibali Goldmines		GT	11	101	728	5,261
Kibali Goldmines		RC	74	6	150	4,355

Kibali Goldmines	2009	DD	9	72	798	2,938

Moto	2009	DD	67	71	790	23,035

Moto	2008	DD	97	53	861	50,250
Moto	2008	GT	12	403	650	6,445

Moto	2007	DD	73	82	953	41,657
Moto		GT	6	170	420	1,871
Moto		RC	1	78	78	78

Moto	2006	DD	127	23	699	36,256
Moto		GT	2	230	263	493
Moto		RC	15	154	200	2,853

Moto	2005	DD	51	116	666	14,890
Moto	2005	RC	24	60	120	2,482

Moto	2004	DD	9	150	421	1,904
Moto	2004	RC	34	31	100	2,596

Moto	1960	DD	11	83	127	1,192

Moto	1952	DD	5	18	107	294

Moto	1951	DD	15	55	129	1,259

Total		14,200				1,350,306

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KCD currently comprises four main mineralised lode systems; 3000, 5000, 9000, and 11000, each with different structural and/or lithological controls. These systems are further sub-divided into individual lodes according to grade, position, and host lithology. There are a total of 24 lodes in the 3000 system, 16 in the 5000 system, eight in the 9000 system, and four in the 11000 system.

Table 14-4 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for KCD underground 3000 domains.

Table 14-4 KCD 3000 Lodes Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

3001	14,573	0.00	69.68	0.91	1.96	26.72	0.90	1.72	11

3002	15,451	0.01	132.88	1.56	2.45	43.1	1.52	2.00	20

3003	20,425	0.00	499.70	1.84	3.35	51.5	1.75	2.39	50

3004	4,774	0.01	33.00	2.03	1.72	27	2.03	1.71	9

3006	3	0.07	0.22	0.16	0.50	-	0.16	0.50	-

3101	4,302	0.01	123.38	5.97	1.26	64.39	5.94	1.22	10

3102	3,769	0.01	180.00	6.02	1.52	60.1	5.87	1.31	23

3103	173	0.05	37.65	4.03	1.06	9.67	3.62	0.67	8

3105	449	0.01	75.70	3.77	2.04	9.32	2.69	0.90	27

3106	2,651	0.01	514.37	6.53	2.78	80	6.00	1.70	12

3107	2,325	0.03	354.78	9.97	1.94	100	16.00	1.56	16

3108	691	0.02	25.49	4.23	0.90	22	4.21	0.88	4

3109	108	0.17	20.60	5.12	0.79	20	5.12	0.78	1

3110	100	0.06	58.15	9.56	1.27	22.3	7.74	0.91	10

3111	201	0.02	123.20	4.70	2.06	10.86	3.74	0.77	14

3112	65	0.01	58.81	6.42	1.64	12.46	4.41	0.89	7

3114-3117	122	0.03	34.90	4.23	1.27	10.85	3.60	0.77	6

3119	52	0.01	20.55	2.64	1.40	-	2.64	1.40	-

3120	45	0.04	27.58	4.51	1.34	-	4.51	1.34	-

3121	319	0.01	193.00	11.59	1.89	112.75	11.18	1.72	4

3122	59	0.60	173.95	17.90	1.65	74.41	15.55	1.30	5

Total	70,657			2.65	2.4		2.49		237

Table 14-5 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for KCD underground 5000 domains.

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Table 14-5 KCD 5000 Lodes Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

5002	8,429	0.00	63.81	0.99	1.96	25.8	0.98	1.70	5

5003	13,140	0.00	240.00	2.65	2.45	77.1	2.60	2.20	23

5004	1,352	0.00	120.83	0.48	8.15	19.56	0.38	4.78	6

5005	34,291	0.00	433.59	1.30	3.36	61.50	1.27	2.34	13

5006	847	0.01	240.00	1.52	5.89	12.9	1.09	1.71	8

5007	52,14	0.01	178.00	2.31	2.32	50.3	2.24	1.88	12

5101	16,299	0.00	3,008.00	7.36	3.48	119	7.11	1.18	18

5102	8,505	0.01	184.17	6.17	1.11	57.78	6.10	0.94	12

5104	360	0.03	240.00	10.63	2.08	55.93	9.12	1.31	7

5105	6,713	0.01	727.02	5.63	2.77	86.12	5.32	1.38	10

5106	76	0.01	14.68	2.83	0.95	7.7	2.65	0.81	4

5110	1,684	0.02	540.00	7.36	2.39	61	6.79	1.37	11

5201	1,657	0.08	194.44	16.87	0.73	100	16.75	0.66	4

5202	890	0.02	340.00	18.92	1.12	100	18.21	0.80	7

5101&5201	17,948	0.01	3,008.00	8.24	3.02	122.45	8.17	1.13	19

5102&5202	9,305	0.01	340.00	7.41	1.35	119.36	7.34	1.17	9

Total	126,710	-	-	4.66	-	-	4.55	-	168

Table 14-6 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for KCD underground 9000 domains.

Table 14-6 KCD 9000 Lodes Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

9004	40,537	-	103.65	0.94	2.48	35.41	0.93	2.16	24

9101	7,179	0.01	242.66	7.10	1.31	72.9	7.05	1.22	9

9102	1,306	0.01	87.37	5.11	1.08	28.31	5.03	0.97	9

9103	101	0.17	42.65	7.12	1.02	19.70	6.58	0.81	5

9104	818	0.03	205.87	7.46	1.41	37.8	7.04	0.91	9

9105	7,426	0.01	579.53	6.80	1.56	77.95	6.71	1.21	7

9106	45	1.60	19.36	5.73	0.67	12	5.47	0.58	3

9107	725	0.01	200.68	5.20	2.10	20.43	4.31	0.97	17

Total	58,137	-	-	2.70	-	-	2.66	-	83

Table 14-7 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for KCD underground 11000 domains.

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Table 14-7 KCD 11000 Lodes Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

11000	1385	0.01	26.8	1.33	1.33	12.08	1.29	1.08	5

11101	348	0.08	125.15	6.84	1.34	34.6	6.59	0.98	3

11102	214	0.04	119.41	7.63	1.42	52.67	7.31	1.12	1

11103	188	0.12	58.55	7.46	1.02	34.7	7.33	0.94	1

Total	2135	-	-	3.40	-	-	3.29	-	10

Sessenge and Sessenge SW

During 2021, a total of 115 drill holes were completed for a sum of 13,620 m. A summary of Sessenge data used for the current Mineral Resource estimate is presented in Table 14-8.

Table 14-8 Sessenge Drill Summary of Holes Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min Depth (m)	Max Depth (m)	Total Drilled (m)

Kibali Goldmines	2021	RC	115	73	170	13,620

Kibali Goldmines	2020	RC	61	15	80	2,819

Kibali Goldmines	2019	DD	4	60	474	1,320
Kibali Goldmines	2019	RC	311	14	127	14,941

Kibali Goldmines	2018	DD	3	89.7	90.6	270.9
Kibali Goldmines	2018	RC	275	6	150	16,348

Kibali Goldmines	2017	DD	11	86	245.2	1,807
Kibali Goldmines		GT	2	101.2	101.47	202.67
Kibali Goldmines		RC	791	10	180	50,227

Kibali Goldmines	2016	DD	7	74.3	193.7	831.45
Kibali Goldmines	2016	RC	206	19	160	11,604

Kibali Goldmines	2015	DD	1	152	152	152
Kibali Goldmines	2015	RC	108	40	130	6,609

Kibali Goldmines	2011	GT	5	30.23	30.5	151.63

Kibali Goldmines	2010	DD	6	119.88	300.28	1236.5
Kibali Goldmines	2010	RC	160	15	150	8,344

Moto	2017	DD	1	227.05	227.05	227.05

Moto	2008	DD	1	266	266	266
Moto	2008	GT	3	150.65	156.85	458.15

Moto	2006	DD	15	39.15	352.95	2,911.45
Moto	2006	RC	23	82	100	2282

Moto	2005	DD	13	158	248.95	2,520.9
Moto	2005	RC	87	40	160	8,288

Moto	2004	RC	41	50	60	2,210

Total			2,250	-	-	149,648

Sessenge was remodelled in 2021 and domains 9104 and 9105, with mean grades above 4 g/t Au, were added.

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Table 14-9 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for Sessenge.

Table 14-9 Sessenge Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

9004	10,878	0.01	60.10	1.47	1.47	27	1.46	1.37	10

9008	622	0.01	31.90	1.91	1.91	18	1.86	1.51	5

9009	321	0.02	5.06	1.06	1.06	-	1.06	1.06	0

9102	1,912	0.03	68.30	3.91	3.91	19.00	3.81	0.78	10

9103	1,431	0.04	41.40	4.82	4.82	22	4.79	0.72	5

9104	297	0.22	36.10	5.57	5.57	19	5.42	0.71	6

9105	90	0.06	21.90	5.12	5.12	13	4.85	0.69	4

Total	15,551	-	-	2.19	-	-	2.14	-	40

Gorumbwa

During 2021, a total of 945 drill holes were completed for a sum of 30,122 m.

A summary of the data used for the current Mineral Resource estimates are presented in Table 14-10.

Table 14-10 Drill Summary of Gorumbwa Holes Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Kibali Goldmines	2021	RC	943	4	120	29,740
Kibali Goldmines	2021	RC_DD	2	156	226	382

Kibali Goldmines	2020	DD	18	147	765	8,823
Kibali Goldmines	2020	RC	160	10	90	7,162

Kibali Goldmines	2019	DD	15	68	807	6,620
Kibali Goldmines	2019	RC	80	15	221	4,766

Kibali Goldmines	2018	DD	1	183	183	183
Kibali Goldmines	2018	RC	359	10	258	15,217

Kibali Goldmines	2017	RC	200	18	177	9,255

Kibali Goldmines	2016	RC	183	23	245	21,132
Kibali Goldmines	2016	RC_DD	48	98	225	8,691

Kibali Goldmines	2015	DD	31	53	407	4,835
Kibali Goldmines	2015	RC	150	20	150	8,796

Kibali Goldmines	2014	DD	67	44	434	14,750
Kibali Goldmines		GT	6	60	279	1,074
Kibali Goldmines		RC	55	12	204	4,586
Kibali Goldmines		RC_DD	4	125	178	595

Kibali Goldmines	2012	DD	12	63	615	3,414

Kibali Goldmines	2011	DD	1	900	900	900
Kibali Goldmines	2011	GT	1	150	150	150

Moto	2006	GT	1	150	150	150

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Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Moto		RC	24	61	450	4,373

Moto	2005	DD	2	41	300	341
Moto	2005	RC	53	34	180	5,545

Moto	2004	DD	24	60	400	4,865
Moto	2004	RC	7	50	64	374

Moto	1996	DD	3	207	395	907

Moto	1960	DD	63	5	59	1,314

Moto	1950	DD	157	42	570	25,203

Total			2,670	-	-	194,143

Table 14-11 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for Gorumbwa.

Table 14-11 Gorumbwa Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

1001	2,824	0.01	43.3	2.22	1.48	27.1	2.21	1.42	9

1003	1,294	0.01	66.4	2.13	1.85	20.4	2.04	1.49	10

1004	8,005	0.01	175	3.47	2.48	62.4	3.32	2.09	32

1006	1,221	0.03	80.5	2.14	2.12	23.5	2.01	1.65	9

1008	2,892	0.01	55.2	2.40	1.30	24.2	2.38	1.22	4

1013	281	0.04	13	1.12	1.40	7.68	1.08	1.12	4

1015	184	0.04	16	1.42	1.36	-	1.42	1.36	0

9001	1,181	0.01	18	1.94	0.98	10.3	1.91	0.91	6

Total	17,882	-	-	2.75	-	-	2.65	-	74

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Pakaka

A summary of the historical and recent drilling undertaken at Pakaka is presented in Table 14-12.

Table 14-12 Drill Summary of Pakaka Holes Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Kibali Goldmines	2021	DD	6	120	419	1,322

Kibali Goldmines	2020	DD	4	507	752	2,569

Kibali Goldmines	2019	DD	4	143	725	2,010
Kibali Goldmines	2019	RC	6	84	114	624

Kibali Goldmines	2018	RC	32	9	26	481

Kibali Goldmines	2017	RC	488	9	213	52,216

Kibali Goldmines	2016	RC	1,146	15	182	84,745

Kibali Goldmines	2015	DD	4	103	130	474
Kibali Goldmines	2015	RC	690	6	131	38,180

Kibali Goldmines	2014	RC	3	100	150	350

Kibali Goldmines	2013	DD	6	97	295	1,075

Kibali Goldmines	2012	DD	9	26	700	1,921
Kibali Goldmines	2012	GT	5	171	323	1,161

Kibali Goldmines	2011	DD	1	700	700	700
Kibali Goldmines		GT	1	169	169	169
Kibali Goldmines		RC	6	100	150	650

Kibali Goldmines	2010	RC	3	100	100	300

Moto	2007	DD	10	341	449	3,910
Moto	2007	RC	7	48	151	806

Moto	2006	DD	41	45	182	4,475
Moto		GT	5	80	191	692
Moto		RC	27	45	80	1,575

Moto	2005	DD	73	100	347	15,434
Moto	2005	RC	34	40	140	3,279

Moto	2004	DD	16	120	230	2,892
Moto	2004	RC	160	30	130	10,799

Moto	1996	DD	9	84	188	1,015

Moto	1980	DD	10	69	237	1,484

Moto	1960	DD	101	65	351	13,656

Total			2,907	-	-	248,964

The Pakaka Mineral Resource estimate was updated in 2019 including revised wireframes, updated cut off grades, grade interpolation parameters. The same 2019 model will be used for 2021 Mineral Resource reporting. Drilling completed at Pakaka since 2019 is for geotechnical and hydrogeological purposes and has no material impact on the Mineral Resource estimate.

Table 14-13 presents statistics of the composite samples used in the current Mineral Resource estimate.

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Table 14-13 Pakaka Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

1001	16,597	0.01	65.42	1.36	1.48	25	1.35	1.32	18

1007	164	0.03	6.99	1.21	1.00	-	1.21	1.00	-

1101	2,777	0.01	89.43	6.84	0.97	34	6.76	0.90	12

1102	102	0.37	520.00	10.16	5.06	45	5.50	1.41	1

1103	652	0.10	36.70	3.22	0.95	-	3.22	0.95	-

1105	1,768	0.01	60.00	3.97	1.21	50	3.96	1.18	4

1106	328	0.01	32.49	3.17	0.93	-	3.17	0.93	-

Total	22,388	-	-	2.37	-	-	2.33	-	35

Kombokolo

Table 14-14 summarises the drilling that has been undertaken at Kombokolo since 2005. The last model update occurred in 2018. A total of six holes were drilled in 2020 to target potential underground opportunity, the results of which have not identified intercepts capable of supporting future underground, but further drilling is planned to re-test this from different drill directions. The 2020 added data will not have no impact on the current Mineral Resource. No drilling activity occurred during 2021.

Table 14-14 Drill Summary of Kombokolo Holes Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Kibali Goldmines	2020	DD	5	365	509	2,222
Kibali Goldmines	2020	RC	1	150	150	150

Kibali Goldmines	2018	DD	7	52	360	1,860
Kibali Goldmines		RC	229	8	300	19,442
Kibali Goldmines		RC_DD	2	276	322	599

Kibali Goldmines	2017	DD	10	179	301	2,251
Kibali Goldmines		GT	5	80	120	480
Kibali Goldmines		RC	217	0	180	9,411

Kibali Goldmines	2016	DD	27	44	263	3,928
Kibali Goldmines	2016	RC	747	20	189	60,932

Kibali Goldmines	2015	GT	1	161	161	161

Kibali Goldmines	2014	GT	1	86	86	86

Moto	2006	DD	14	100	210	2,062
Moto	2006	GT	1	90	90	90

Moto	2005	RC	30	59	170	3,382

Total			1,297	-	-	107,056

Drilling completed at Kombokolo since 2018 is for down plunge exploration purposes of evaluating underground potential and has no material impact on the Open Pit Mineral Resource estimate.

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The Kombokolo deposit is comprised of nine lodes, seven of which are low-grade domains (1001 to 1007) and two are high-grade domains (1101 and 1102). The high-grade domains 1101 and 1102 were identified in the 2018 model update through specific geological criteria observed in DD. These domains are hosted within the two main lodes 1001 and 1002 respectively.

Table 14-15 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for Kombokolo.

Table 14-15 Kombokolo Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

1001	4,852	0.01	72.20	1.33	1.97	18.20	1.28	1.49	15

1002	2,459	0.01	152.00	1.41	2.61	14.00	1.33	1.17	15

1003	37	0.01	4.10	1.10	0.84	-	1.10	0.84	0

1004	11	0.11	3.13	0.90	0.96	-	0.90	0.96	0

1005	33	0.01	4.00	1.85	0.86	-	1.85	0.86	0

1006	11	0.50	5.41	2.33	0.69	-	2.33	0.69	0

1007	3	0.02	2.63	1.28	1.03		1.28	1.03	0

1101	2,466	0.01	239.67	6.67	1.35	39.60	6.42	0.97	17

1102	673	0.06	124.40	5.56	1.32	16.00	4.99	0.78	25

Total	10,487	-	-	2.83	-	-	2.72	-	72

Pamao and Pamao South

The Pamao Mineral Resource estimate was updated in 2021, with the addition of 3,001 holes (546 drilled in the Pamao South portion of the deposit) for a total of 142,994 m. The mineralisation model was fully revised with significant changes in interpretation and the modelling of intercepts of a QSF internal waste unit within a meta-sandstone host

Pamao South has been included in 2021 Mineral Resources as a new addition within the overall Pamao Mineral Resource.

A summary of the historical and recent drilling undertaken at Pamao is presented in Table 14-16.

Table 14-16 Drill Summary of Pamao Holes Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Kibali Goldmines	2021	DD	42	18	332	6,000
Kibali Goldmines	2021	RC	2,959	6	240	136,994

Kibali Goldmines	2020	DD	14	3	224	2,018
Kibali Goldmines	2020	RC	268	10	195	21,858

Kibali Goldmines	2019	RC	39	36	240	5,368

Kibali Goldmines	2018	DD	9	56	155	858
Kibali Goldmines	2018	RC	170	11	179	15,032

Kibali Goldmines	2017	RC	330	35	176	27,187

Kibali Goldmines	2016	DD	7	89	189	911

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Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Kibali Goldmines		RC	62	18	96	2,642

Kibali Goldmines	2012	DD	6	120	210	1,022
Kibali Goldmines	2012	RC	1	5	5	5

Moto	2007	GT	4	80	105	385

Moto	2005	DD	24	81	200	3,495
Moto	2005	RC	26	60	130	2,640

Moto	2004	RC	118	40	160	7,920

Total			4,079			4,079

Table 14-17 and Table 14-18 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for Pamao and Pamao South respectively.

Table 14-17 Pamao Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

2001	1,400	0.01	10.5	0.92	0.83	5.90	0.92	0.79	3

2002	13,537	0.01	29.6	1.09	1.14	22.99	1.09	1.12	18

2003	2,424	0.01	53.1	1.21	1.44	10.9	1.18	1.13	4

2102	2,106	0.02	33.6	3.81	0.82	23.39	3.79	0.79	5

Total	19,467	-	-	1.39	-	-	1.38	-	30

Table 14-18 Pamao South Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

1001	1,009	0.01	122.0	1.93	2.40	20.0	1.80	1.34	4

1002	2,137	0.01	86.69	1.92	1.93	33.1	1.87	1.62	5

1003	436	0.04	39.4	1.89	1.72	13.2	1.75	1.26	4

Total	3,582			1.92			1.84		13

Mengu Hill

The last drillhole for Mengu Hill was completed in 2019 and the Mineral Resource estimate was updated in same the year.

A summary of the historical and recent drilling undertaken at Mengu Hill is presented in Table 14-19.

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Table 14-19 Drill Summary of Mengu Hill Holes Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Kibali Goldmines	2019	RC	1	50	50	50

Kibali Goldmines	2018	DDH	2	322	329	651

Kibali Goldmines	2016	RC	317	6	112	14,991

Kibali Goldmines	2015	DDH	2	180	806	986
Kibali Goldmines		GT	2	150	150	300
Kibali Goldmines		RC	443	12	84	16,369
Kibali Goldmines		RC_DDH	17	78	332	2,715

Kibali Goldmines	2014	RC	1,011	8	175	42,160

Kibali Goldmines	2013	DDH	22	40	330	2,621
Kibali Goldmines	2013	RC	1	200	200	200

Kibali Goldmines	2012	DDH	6	153	395	1,555
Kibali Goldmines	2012	GT	11	10	266	1,105

Moto	2006	DDH	1	85	85	85

Moto	2005	DDH	37	53	260	5,764
Moto	2005	RC	16	60	240	3,887

Moto	2004	RC	78	60	60	4,680

Total			1,967	-	-	98,119

Drilling completed at Mengu Hill since 2018 was for down plunge exploration at depth for the purposes of evaluating underground potential, the results of which were inconclusive, and so further closer spaced drilling is planned to evaluate this. The drilling since 2018 is not included in the current Mineral Resource but would have no material impact on the Mineral Resource estimate.

Table 14-20 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for Mengu Hill.

Table 14-20 Mengu Hill Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

1001	9,812	0.01	138.38	1.29	1.59	34.40	1.28	1.53	3

1002	135	0.03	4.89	1.15	0.93	2.50	1.03	0.79	14

1101	7,601	0.02	117	5.71	1.16	61.90	5.69	1.11	6

1102	26	1	31.09	7.94	1	8.01	5.69	1.11	6

1103	151	0.05	11.3	2.95	0.7	6.69	2.81	0.59	7

Total	17,725	-	-	3.21	-	-	3.19	-	10

Mengu Village

The most recent drillholes for Mengu Village was completed in 2020 adding 40 drillholes for 4,610 m. The Mengu Village deposit was re-modelled in 2021 based on the new data and new geological interpretations, whilst also using ordinary kriging for the block model estimates (rather than uniform conditioning used in 2006).

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A summary of the historical and recent drilling undertaken at Mengu Village is presented in Table 14-21.

Table 14-21 Drill Summary of Mengu Village Holes Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Kibali Goldmines	2020	RC	40	56	194	4,610

Kibali Goldmines	2019	RC	1	90	90	90

Kibali Goldmines	2013	RC	4	50	95	265

Kibali Goldmines	2012	DDH	3	170	395	793

Moto	2005	RC	26	60	120	2,111

Moto	2004	RC	11	60	60	660

Total			85	-	-	8,529

Table 14-22 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for Mengu Village.

Table 14-22 Mengu Village Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

1004	256	0.08	18	1.62	1.06	5.88	1.53	0.78	6

Megi-Marakeke-Sayi

No additional drilling has taken place at Megi-Marakeke-Sayi in 2021. The latest drilling was carried out in 2020 where 33 DD for 3,325 m and 732 RC holes for 60,396 m were added.

A summary of the historical and recent drilling undertaken at Megi-Marakeke-Sayi is presented in Table 14-23.

Table 14-23 Drill Summary of Megi-Marakeke-Sayi Holes Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Kibali Goldmines	2020	DD	33	66	196	3,325
Kibali Goldmines	2020	RC	692	20	220	55,786

Kibali Goldmines	2019	RC	117	29	150	10,609

Kibali Goldmines	2016	DD	4	100	152	502

Kibali Goldmines	2015	DD	16	22	347	2,327
Kibali Goldmines		GT	7	113	252	1,175
Kibali Goldmines		RC	93	18	102	5,308

Kibali Goldmines	2014	RC	15	42	120	1,106

Kibali Goldmines	2013	RC	15	24	95	663

Kibali Goldmines	2012	DD	2	251	252	502

Moto	2005	RC	56	40	130	4,732

Moto	2004	RC	84	50	120	6,178

Moto	1950	RC	102	1	170	2,856

Total			1,236	-	-	95,069

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Table 14-24 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for Megi-Marakeke-Sayi.

Table 14-24 Megi-Marakeke-Sayi Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

1001	529	0.02	32	1.50	1.67	7.26	1.33	1.10	16

1002	934	0.01	20.4	1.31	1.19	9.27	1.28	1.17	6

1003	418	0.01	67.2	2.29	1.97	14.1	2.08	1.27	7

1004	1,928	0.01	19.7	1.46	1.07	11.1	1.44	0.98	5

1005	2,015	0.01	40.9	1.31	1.27	12.9	1.29	1.06	5

1101	127	0.04	193	6.71	2.57	14.31	5.07	0.76	8

1102	141	0.06	42.6	5.11	1.11	20.1	4.90	0.84	4

1105	147	0.09	10.1	3.56	0.57	9.46	3.56	0.52	3

Total	6,239	-	-	1.69	-	-	1.60	-	54

Kalimva-Ikamva

During 2020 and 2021, a total of 59 holes were completed for a sum of 7,139 m, targeting a down plunge underground opportunity and an exploration target to the east of the main Ikamva deposit, outside of the open pit optimised pit shells. The drilling results have not yet been incorporated in a block model update. This will be done once the underground target investigations and exploration work are completed in 2022. This drilling has no material impact on the 2021 Mineral Resource estimate.

A summary of the historical and recent drilling undertaken at Kalimva-Ikamva is presented in Table 14-25.

Table 14-25 Drill Summary of Kalimva-Ikamva Holes Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Kibali Goldmines	2021	DD	7	214	397	2,097
Kibali Goldmines	2021	RC_DD	1	411	411	411

Kibali Goldmines	2020	DD	1	120	120	120
Kibali Goldmines	2020	RC	50	35	247	4,511

Kibali Goldmines	2019	DD	38	22	428	5,851
Kibali Goldmines	2019	RC	656	19	298	58,150

Kibali Goldmines	2018	DD	11	141	255	2,072
Kibali Goldmines		RC	173	40	258	19,555
Kibali Goldmines		RC_DD	1	173	173	173

Kibali Goldmines	2017	RC	76	40	230	9,599

Kibali Goldmines	2016	RC	79	37	88	4,624

Kibali Goldmines	2015	DD	4	56	473	753

Kibali Goldmines	2012	DD	14	102	252	2,686

Total			1,111	-	-	110,602

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Table 14-26 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for Kalimva.

Table 14-26 Kalimva Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

1001	4,440	0.01	25.85	0.95	1.11	6.66	0.94	0.95	19

1101	1,926	0.01	69.39	3.67	1.01	20.6	3.60	0.83	9

Total	6,366	-	-	1.78	-	-	1.74	-	28

Table 14-27 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for Ikamva.

Table 14-27 Ikamva Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

2001	1,960	0.01	7.35	0.83	0.95	5.60	0.83	0.95	3

2101	642	0.01	23.90	4.37	0.74	15.90	4.33	0.71	8

Total	2,602	-	-	1.71	-	-	1.70	-	11

Aerodrome

During 2021, a total of 74 DD and RC drill holes were completed for a sum of 2,918 m taking the total meterage to 41.7 km of drilling. A summary of Aerodrome data used for the current Mineral Resource estimate is presented in Table 14-28.

Table 14-28 Drill Summary of Aerodrome Holes Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Kibali Goldmines	2021	DD	2	70	81	151
Kibali Goldmines	2021	RC	72	4	98	2,767

Kibali Goldmines	2020	DD	7	126	176	1,034
Kibali Goldmines	2020	RC	104	18	138	8,227

Kibali Goldmines	2018	RC	11	42	150	783

Kibali Goldmines	2017	DD	16	90	210	2,417
Kibali Goldmines	2017	RC	140	8	123	7,736

Kibali Goldmines	2013	RC	79	36	118	4,921

Kibali Goldmines	2011	RC	2	20	73	93

Moto	2007	DD	34	73	375	5,687
Moto	2007	RC	19	40	140	1,870

Moto	2006	RC	35	50	160	3,435

Moto	2004	RC	52	47	60	2,617

Total			573	-	-	41,737

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Table 14-29 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for Aerodrome.

Table 14-29 Aerodrome Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

2001	1,087	0.25	13.8	1.45	0.95	6.19	1.43	0.90	10

2002	542	0.01	122	1.69	2.40	12.8	1.50	1.29	6

Total	1,629	-	-	1.53	-	-	1.45	-	16

Oere

During 2021, a total of 163 DD and RC drill holes were completed for a sum of 19,782 m bringing the total meterage in that deposit to 36.6 km of drilling. A summary of Oere data used for the current Mineral Resource estimate is presented in Table 14-30.

Table 14-30 Drill Summary of Oere Holes Used in the 2021 Mineral Resource Estimate



Company	Year Completed	Drill Type	No. of Holes	Min. Depth (m)	Max. Depth (m)	Total Drilled (m)

Kibali Goldmines	2021	DD	22	50	190	2,824
Kibali Goldmines	2021	RC	141	55	198	16,958

Kibali Goldmines	2020	RC	23	65	198	3,009

Kibali Goldmines	2019	DD	8	134	434	2,172
Kibali Goldmines	2019	RC	48	41	216	5,660

Kibali Goldmines	2018	RC	19	54	150	1,939

Kibali Goldmines	2017	RC	1	84	84	84

Moto	2006	RC	56	60	60	3,360

Moto	2005	RC	10	60	60	600

Total			328	-	-	36,606

Table 14-31 presents statistics of the composite samples used in the 2021 Mineral Resource estimate for Oere.

Table 14-31 Oere Composite Data – 2021 Mineral Resource Estimate



Domain	Raw					Capped
Domain	No. of Samples	Min (g/t Au)	Max (g/t Au)	Mean (g/t Au)	CV	Grade (g/t Au)	Mean (g/t Au)	CV	No. of Samples

1001	1,661	0.01	17.78	1.32	1.32	17.8	1.31	1.13	2

1101	257	0.01	30.90	4.48	0.83	30.9	4.48	0.83	0

Total	1,918	-	-	1.74	-	-	1.73	-	2

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14.3

Geological Modelling

Geological interpretation and modelling are based on the following standard procedures:

●

Hard copy geological cross sections and long sections are generated and updated during drill campaigns. These are then scanned and georeferenced to be used as a basis for 3D modelling.

●

Geological interpretations are digitised as polylines on cross sections spaced 10 m apart. Lithological, weathering, oxidation, low and high-grade polylines are snapped on each section to the corresponding sample interval. In areas of complex folding, additional polylines are wireframed between sections to build a valid 3D solid. Most of the open-pit sections were based on flitch-plans and used for updating sub-surface geology, with special attention paid to short range barren internal waste lithologies.

●

Mineralisation domains are sub domained into low-grade (>0.5 g/t Au), high-grade (>2.0 g/t Au), and very high-grade (>7.5 g/t Au) domains, utilising contact analysis and domain stationarity tests.

●

For active mining areas, the geological and mineralisation models are updated quarterly when additional grade control data is available.

●

Interpretations are regularly cross checked with DD core and RC chips to ensure the model is representative.

●

Chip samples are used within the underground development area to provide an additional source of information regarding the mineralisation associated with the alteration, particularly when mapping low-grade halo contacts. This data is recorded on the underground geological maps, which are then scanned and georeferenced for wireframe model updating. However, this data is used only for modelling of geological contacts and is not directly used for Mineral Resource estimation.

●

Rip-line samples are used within the open cast exposed benches to provide an additional source of information regarding lithologies and mineralisation, particularly when mapping contacts and updating the exact dimensions of modelled internal dilution, artisanal depletion, and carbonaceous shale units. This data is used for refining geological models and is not used for Mineral Resource estimation.

Statistical analysis of the data shows that a suitable geological related threshold grade is approximately 0.5 g/t Au for the KCD and Sessenge deposits. This same 0.5 g/t Au modelling threshold has been applied at all other deposits. The resulting low-grade mineralised envelopes incorporate minor amounts of internal sub-grade material to preserve continuity. During interpretation, efforts were made to minimise the amount of sub-grade material included within each of the lode wireframes.

Mineralisation domains were built with a combination of grade, lithology, alteration, structural data, and the presence of pyrite. In areas where further contiguous high-grade shoots are evident and supported by the geological logging, high-grade continuity wireframes were also considered. The intention of the geological domaining is to generate a single stationary geostatistical population for each of the domains.

The dimensions and orientations of the modelled mineralised domains for all deposits are summarised in Table 14-32.

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Table 14-32 Mineralisation Domain Dimensions for all Mineral Resource Deposits



Deposit	Down Plunge (m)	Down Dip (m)	Thickness (m)	Plunge Direction

KCD 3000	1,900	450	200	NE

KCD 5000	2,200	250	80	NE

KCD 9000	2,280	470	80	NE

KCD 11000	670	400	100	NE

Mengu Hill	850	100	90	NNE

Sessenge	500	400	40	NE

Gorumbwa	1,240	500	150	NE

Kombokolo	730	300	30	ENE

Pakaka	1700	500	30	NE

Pamao	1250	690	45	NW

Megi-Marakeke-Sayi	1900	450	100	NW

Kalimva	1470	270	30	NNE

Ikamva	1580	120	50	NE

Aerodrome	350	200	40	NNW

Sessenge SW	520	150	25	NE

Oere	2580	600	30	NNE

Pamao South	830	150	35	NE

Mengu Village	1050	560	30	NW

Boundary analysis (Figure 14-1) is completed to check the nature of the grade transition across domain contacts, most profiles being sharp (hard) and rarely gradual (soft). This helps delineate the rod-like high-grade mineralisation shoots noted in the KCD, Sessenge, Kombokolo, and Pakaka deposits.

LOGO

Figure 14-1 Boundary Analysis between KCD High-Grade (5101) and Low-Grade (5005) Domains

Composites are coded by domain. These codes are used for statistical analysis and domain control during the estimation process. The coding of the composites and block model is prioritised

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to ensure that the high-grade domain codes are preserved when they are situated within surrounding low-grade mineralisation envelopes.

To ensure consistency of the domaining controls used, the database and geological block model are both flagged with the same codes defining the mineralised envelopes that a particular composite falls within. The high-grade mineralised envelopes are predominantly situated within low-grade mineralisation wireframes, which are built independently of each other. Since Boolean operations are not utilised to remove these overlaps between internal high-grade shoot models and surrounding low-grade mineralisation envelope wireframes, care is taken to avoid the double-counting of samples and blocks.

All Mineral Resource pits were optimised on a $1,500/oz Au pit shell selected, following an analysis of the pit size against value and gold price.

All Mineral Reserve pits were designed on a $1,200/oz Au pit shell selected, following an analysis of the pit size against value and gold price. The exceptions are the Sessenge and Oere pits, where the reserve pit designs were both based on a $1,300/oz Au optimised pit shell and Aerodrome, where the reserve pit designs were based on a $1,500/oz Au optimised pit shell. All Mineral Reserves, including Aerodrome, Sessenge, and Oere are profitable at $1,200/oz Au, and thus the Mineral Reserve and supporting cash flow statements are reported at $1,200/oz Au. This is in line with Barrick corporate guidelines, which considers long-term gold price forecasts.

KCD, Sessenge and Sessenge SW

These areas have been modelled using a combination of grade continuity, alteration, mineralisation, and structural readings, where available.

KCD Mineralisation Modelling

The mineralisation wireframes generally followed a threshold grade of 0.5 g/t Au but did include some internal dilution of the lower grade material, where applicable, to create reasonably continuous envelopes. Some of these broad envelopes often contain multiple rod-like high-grade zones within them. These high-grade zones are associated with strong ACSA alteration and fine-grained disseminated pyrite and have been modelled separately at a nominal threshold grade of 2.0 g/t Au. The strong ACSA alteration and pyrite zones were also modelled.

Mineralisation at KCD and Sessenge has been grouped into four lodes (3000, 5000, 9000 and 11000). The 9000 lode extends up plunge to surface at Sessenge, enabling the two deposit areas to be joined together into a single unit (Figure 14-2). Since 2018, there has been significant growth in the 3000 and 9000 lodes from underground Mineral Reserve conversion drilling, plus the discovery of the new 11000 lode Mineral Resource.

Thin continuous barren intrusive/volcanic late dolerite units and QSF early felsic intrusives are interspersed within the metasediment units and have been used as marker units during geological interpretation. Where possible, these larger barren units have been modelled independently and flagged in both the composites and the block model.

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LOGO

Source: Kibali Goldmines, 2021

Notes:

3000 Lodes (Yellow), 5000 (Red), 9000 (Purple), 11000 Lodes (Brown), and $1,200/oz Au Optimised Pit Shell (Grey)

Looking Northwest

Figure 14-2 3D View of KCD Sessenge Mineralisation

Structure, Alteration, and Mineralisation

Since ACSA alteration and mineralisation was spatially linked with the distribution and fold shape of BIF, it has been possible to trace the mineralisation, and modelling the continuity and morphology of the lodes from one section to the next has become more continuous. Confidence within the geological models has therefore been greatly improved. Folding, with alteration and mineralisation following the main fold plunge direction, can be traced with confidence in the same structural setting from one cross-section to the next.

Like the host rocks in the 5000 to 9000 lodes, the most striking feature of the 3000 lode host rocks is the pervasive sericitic foliation. This foliation is subparallel to the fold axial surfaces of the mapped folds. The axial surfaces of the folds in the 3000 lodes are subparallel to the 5000 to 9000 axial surfaces, although these generally become steeper and locally more curved (due to subsequent deformation) down section in the 5000 to 9000 lodes. The similar orientation, form, and relative paragenetic timing of the sericitic foliation in both areas suggests that the associated folds in both areas are the same generation and not fundamentally different in terms of timing or original orientation.

ACSA-A alteration is controlled by the fold axial surface foliation in both areas and is most intense in subparallel shears. The overprinting ACSA-B alteration and mineralisation is significantly different in the 3000 lodes, but this is interpreted as the result of local host rock variation rather than a significant change in the mineralising hydrothermal fluids. Mineralisation in the 3000 lode is hosted primarily in brecciated cherts and BIF in the hinges of folds, whereas mineralisation in the 5000 to 9000 lodes is hosted predominantly in the hinge zones and limbs of folded BIF units.

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Due to the predominance of chert in the 3000 lodes, the pyrite-pyrrhotite assemblage is more common than the pyrite dominated 5000 to 9000 lodes that are hosted in BIF.

Mineralised zones in the 3000 lodes are typically hosted in tightly folded hinge zones where relatively brittle host rocks (cherts and BIF) have resisted folding and are brecciated and sheared. The 3000 lode contains abundant cm to m scale early felsic dykes and sills that crosscut these units and, although not particularly mineralised, are signs of localised strong shearing and ACSA-A alteration. This shearing has strongly affected the host rocks, and cherts and BIF are typically brecciated and mineralised along the contacts with these altered, rheologically weak, units.

Shearing has also focussed on the carbonaceous phyllite units and carbonaceous shears; typically containing lenticular clasts of other sheared lithologies including siltstone, chert, and igneous clasts. Clasts are commonly replaced by pyrite ± pyrrhotite that commonly carry elevated gold values, but mineralised sections are volumetrically minor in the 3000 lode. The best mineralised areas are hosted in sheared and brecciated cherts and BIFs on the margins of such zones.

Each of the mineralised wireframes were modelled separately and snapped to drill holes where applicable. Lithological, weathering, and redox wireframes (also modelled from drill hole data) were flagged into both the database and the block model with their respective priorities.

Gorumbwa

Mineralisation at Gorumbwa is hosted almost exclusively within the meta-sandstone unit, with minor sporadic mineralisation noted in a conglomerate unit that occurs beneath the meta-sandstone. Mineralisation is divided into eight lodes shown together in Figure 14-3, which coarsely trend west to WSW, dip to the NW, and plunge to the ENE at approximately 30°. The lenses are echelon like in vertical stacking, with only the main 1004 lode being the most consistent in continuity. Higher grades within the lodes tend to the central area of the shoots where a higher strain environment increased accommodation space and hydrothermal fluid inflows. Historic UG mining focussed on the extraction of the main (1004) lode. Historical extraction is shown by the underground depletion voids (Figure 14-3).

The styles of mineralisation vary from KCD, with the dominant style being moderate to strong silicification and sericitisation with minimal pyrite, with a low correlation between sulphide and gold content. The second style is the typical ‘ACSA’ style noted at KCD where the gold is proportional to pyrite percentages, though the iron carbonate is predominantly ankerite, unlike KCD which is dominated by siderite. This style is mainly observed in the main 1004 lode. The third style is visible gold within late, moderate to strong silicification.

Mineralisation is structurally controlled within a NE trending corridor where the S1 foliation strikes E-W. The corridor is bounded by NE crosscutting structures on the eastern and western margins. These NE structures are indicated by the discontinuity of the red pebble conglomerate horizon in the west, micro folding within the lithological units near to the structures, and rotation of S1 foliation laterally across the 200 m wide mineralised area. The structures may be ductile; whilst refolded folds are not observed, they are postulated to have been present in the high-grade shoot, which was mined out and makes up the void area.

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Each of the mineralisation wireframes are modelled separately and snapped to drill holes where applicable. Lithological, weathering, and redox wireframes also modelled from drill hole data, which were flagged into the database and the block model with their respective codes.

Historical mining in these areas has resulted in the highest-grade mineralisation being mined out and thus all the current lodes have been classified as low-grade envelopes.

LOGO

Source: Kibali Goldmines, 2021

Notes:

Gorumbwa Lodes (Yellow), Underground Depletion (Magenta), and $1,200/oz Au Optimised Pit Shell (Grey)

Looking Northwest

Figure 14-3 3D View of Gorumbwa Mineralisation

Pakaka

Gold mineralisation at Pakaka is hosted by volcano-sedimentary conglomerates interbedded with minor tuffaceous units. Ironstone units are rare which is unusual for Kibali. The mineralised zones are characterised by silica-ankerite/siderite-pyrite alteration, mainly in well foliated siliceous rocks. The mineralised zones are associated with pervasive silicification with local preservation of breccia textures that have been overprinted by the dominant S1 fabric. Higher gold grades appear to correlate well with the presence and abundance of pyrite, which appears to be spatially associated with the intersection of the NW trending D1 thrust surface, and a NE trending strain corridor.

The Pakaka mineralisation is open down plunge (18°) towards the NE, representing further exploration potential. The Pakaka mineralisation extends over a strike length of 700 m, has variable thickness and has been identified to a depth of 350 m below surface.

The mineralisation is interpreted as an open fold with a thick (30 m) western limb dipping 7° toward SE and a very thin (average 12 m) eastern limb dipping 18° towards the SE (Figure 14-4). The axial plan is trending NE and plunging 18° towards the NE.

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The Pakaka mineralisation is a single mineralised lode with internal high-grade zones.

LOGO

Source: Kibali Goldmines, 2021

Notes:

Low-Grade Mineralisation (Light Orange), High-grade Mineralisation (Orange) and Optimised Pit Shells ($1,200/oz Au (Grey) and $1,500/oz Au (Blue))

Looking West-Northwest

Figure 14-4 3D View of Pakaka Mineralisation

Kombokolo

The Kombokolo mineralisation is located between a clastic meta-conglomerate unit and the footwall of ironstone, following the brittle deformation within those units. Carbonate, sericite, silica, and chlorite alteration is associated with the mineralisation at varying intensities, plus disseminated sulphides, predominantly pyrites. High grades generally sit in the very strongly deformed meta-conglomerate with strong and dark chlorite alteration and high percentage of pyrite content. Grain size becomes very fine, and foliations are rarely identifiable.

In general, the mineralisation, in general, plunges at low to moderate angles (30°) to the NE/ENE. The dip direction is mainly to the NW and dip angle is around 23° but flattens out up-plunge.

Nine low-grade lodes have been modelled, with the principal lodes being 1001 and 1002. These have a down plunge continuation of approximately 490 m within the $1,500/oz Au pit shell to a depth around 200 m below surface, an average width of 150 m, with an average thickness of 15 m which can reach to approximately 25 m to 35 m locally. The lode remains open down plunge.

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LOGO

Source: Kibali Goldmines, 2021

Notes:

Low-Grade Mineralisation (Yellow), High-grade Mineralisation (Orange) and Optimised Pit Shells ($1,200/oz Au (Grey) and $1,500/oz Au (Blue))

Looking Southwest

Figure 14-5 3D View of the Kombokolo Mineralisation

Pamao and Pamao South

Pamao is located in a part of the KZ Trend where NE-SW thrusting controls the planar mineralised structures (Figure 14-6). The Pamao mineralisation is predominantly hosted in meta-conglomerate (+ minor BIF), dipping at low to moderate angles (25°) to the NE, with a generally flat plunge (5°). Pamao South is stratigraphically lower and slightly away from the thrusting, in a structural setting characterised by multiple folded BIFs within a package of meta-conglomerate. The Pamao South mineralisation is predominantly hosted along BIF-conglomerate lower sheared contacts (+ brecciated conglomerate), with dip of the lodes increasing up to 50° SE in the up-plunge area, with overall low to moderate plunge (25°) to the NE.

The lithological sequence starts by a hanging wall basaltic unit with the lower contact marked in some places by meta-siltstone rock with carbonaceous shale (graphitic shear) interpreted as a thrust. This is known as the upper bounding package.

Beneath the upper bounding package is a strongly foliated meta-sandstone package (middle formation), containing a sheared meta-conglomerate (with chloritic matrix) package. A minor up-dip BIF unit (sometimes appearing as a magnetite alteration layer – potentially relic BIF), sits within this meta-conglomerate near surface and is itself sometimes intercalated with a thin unit of sandstone. This unit is the main host rock to the mineralised system at Pamao.

A sericitic schist marks the contact between the meta-sandstone below and the meta-conglomerate above and constitutes the lower limit of the Pamao deposit.

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Dolerite dykes intrude this meta-conglomerate, verging towards the contact with the meta-sandstone at surface. The mineralised unit is also intruded by a centimetric felsic porphyry (QSF) unit that is strongly sheared.

At Pamao, silica alteration is dominant, but ACSA appears locally within the high-grade envelope. Pyrite is the dominate sulphide with chlorite widely present. At Pamao South, weak ACSA alteration is dominant, with fine disseminated pyrite.

LOGO

Source: Kibali Goldmines, 2021

Notes:

Low-Grade Mineralisation (Orange), High-grade Mineralisation (Red) and Optimised Pit Shells ($1,200/oz Au (Grey) and $1,500/oz Au (Blue))

Looking West

Figure 14-6 3D View of the Pamao and Pamao South Mineralisation

Mengu Village

The mineralisation at Mengu Village occurs within a geological sequence consisting of meta-siltstone with intercalated graphitic layers, meta-basalts, and banded iron formation. The whole sequence is intruded by late dolerite dykes. The mineralisation is associated with the meta-sediment and characterised by weak ACSA-A and strong silica alteration.

Mineralisation at Mengu Village is modelled as a single domain that is approximately 1,000 m in strike length with an average thickness of 4 m and has been identified to a depth of 200 m below the surface (Figure 14-7). The mineralisation trends NW and dips 25° to the NE. The Mengu Village mineralisation shows good continuity as a NW extension of lode 1004 from the Sayi deposit, albeit with zones of barren material.

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LOGO

Source: Kibali Goldmines, 2021

Notes:

Low-Grade Mineralisation (Orange), and Optimised Pit Shells ($1,200/oz Au (Grey) and $1,500/oz Au (Blue))

Looking West-Northwest

Figure 14-7 3D View of the Mengu Village Mineralisation

Megi-Marakeke-Sayi

The Megi-Marakeke-Sayi complex is located on the KZ Trend along the Pakaka-Mengu trend, approximately 1 km NW of Pamao and 6 km north of the KCD pit.

The lithological sequence consists mostly of meta-siltstones/sandstones with intercalations of carbonaceous shale, metabasalt, BIF, and a dolerite unit.

In Megi-Marakeke, the mineralisation is hosted within the BIF and meta-sandstone units, associated with strong silica and fine pyrite. While at Sayi, the mineralisation is related to the fine pyrite/strong silica alteration hosted solely in meta-sandstone.

Re-interpretations of the mineralisation at Megi-Marakeke-Sayi were carried out on NE-SW 20 m spaced sections depending on drilling spacing. At a 0.5 g/t Au threshold and 20 m section spacing, the resulting interpretation demonstrates consistent geometry and excellent continuity of the mineralised zone from Megi to the east through Sayi to the west.

The 3D wireframes generated for the interpreted mineralised volume (Figure 14-8), demonstrate that the Megi-Marakeke-Sayi deposits occurs as layered tabular bodies typically between 10 m to 30 m thick, that trend N310° and dip 25°NE. The modelled plunges of the high-grade shoots are limited, with continuity best developed down-dip. The mineralised zone tested by drilling has a strike length of approximately 1,000 m and extends 200 m down dip.

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Source: Kibali Goldmines, 2021

Notes:

Low-Grade Mineralisation (Yellow), High-grade Mineralisation (Orange) and Optimised Pit Shells ($1,200/oz Au (Grey) and $1,500/oz Au (Blue))

Looking West

Figure 14-8 3D View of the Megi-Marakeke-Sayi Mineralisation

Kalimva-Ikamva

Kalimva-Ikamva are located towards the north end of the KZ Trend. The deposits are located approximately 1 km apart, with Ikamva to the NW of Kalimva. The lithological sequence consists of felsic intrusive, BIF, and basal meta-conglomerate intercalated at the footwall by sheared lenses of carbonaceous shale (CS).

The mineralisation is hosted within metasediment and at the contact with siliciclastic BIF, dipping SE along the regional NNE shear with high-grade shoots dispersed and plunging NE. Mineralisation is driven by strong chlorite, silica, and fine disseminated pyrite within the mineralised envelope. The high-grade envelope is characterised by pyrite, pyrrhotite, and silica.

Manual interpretations of the mineralisation at Kalimva-Ikamva were carried out on NW-SE 20 m spaced sections at a threshold of 0.5 g/t Au. The spacing of the sections were based on a drilling grid of 20 m by 20 m.

The 3D wireframes generated from the interpreted sections are shown in Figure 14-9, which demonstrate that Kalimva occurs as a steep layered deposit, while Ikamva was interpreted as a recumbent fold, where the mineralisation is hosted on the hinge and the limb of the fold. The thickness of the deposits is between 5 m to 10 m thick for the high-grade lodes and 20 m to 30 m for the low-grade lodes.

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The mineralisation at Kalimva trends NNE and dips approximately 65°ESE. The modelled plunge of the high-grade shoot is approximately 15° towards NNE and slightly flatter at 5° for the low-grade domain. The mineralised zone tested by drilling has a strike length of approximately 1,400 m and extends 240 m down dip.

The mineralisation at Ikamva trends NE and dips approximately 25° to the SE. The modelled plunge of the low-grade domain is approximately 12° towards NE. The mineralised zone tested by drilling has a strike length of approximately 1,400 m and extends 200 m down dip. The wireframes at Ikamva have been extrapolated beyond the drilling to form a conceptual underground exploration target.

LOGO

Source: Kibali Goldmines, 2021

Notes:

Low-Grade Mineralisation (Yellow), High-grade Mineralisation (Red) and Optimised Pit Shells ($1,200/oz Au (Grey))

Looking Northwest

Figure 14-9 3D View of the Kalimva Ikamva Mineralisation

Mengu Hill

Mineralisation at Mengu Hill consists of one large and two small low-grade domains with high-grade rod-like shoots embedded within. All wireframes were modelled separately and coded to represent the domain that they belong to. Mineralisation is predominantly hosted in the BIF, which forms the cap of the hill. Lithological, weathering, redox and mineralisation wireframes were all built based on drillhole data and surface mapping.

The mineralisation wireframes (Figure 14-10) were snapped to drill holes as much as possible to create a more accurate interpretation. This resulted in minimising the influence of unwanted zones in the final composite files.

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Source: Kibali Goldmines, 2021

Notes:

Low-Grade Mineralisation (Yellow), High-grade Mineralisation (Orange) and Optimised Pit Shells ($1,200/oz Au (Grey) and $1,500/oz Au (Blue))

Looking West-Northwest

Figure 14-10 3D View of the Mengu Hill Mineralisation

Aerodrome

The Aerodrome lithology sequences start with fine-medium grained sediments overlaying the basalts, followed by a thick package of meta-conglomerate with BIF lenses. Quartz-feldspar units cap the mineralised system.

The mineralisation is mainly associated with silica, chlorite alteration, fine disseminated pyrite and pyrrhotite. The interpreted mineralised volume shows that the Aerodrome deposit (Figure 14-11) is a layered tabular deposit between 4 m to 25 m thick, trending NNW and dipping 30° to the ENE. The mineralised zone tested by drilling has a strike length of approximately 350 m and extends approximately 300 m down dip.

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Source: Kibali Goldmines, 2021

Notes:

Mineralisation (Yellow), and Optimised Pit Shells ($1,500/oz Au (Blue))

Looking Northwest

Figure 14-11 3D View of the Aerodrome Mineralisation

Oere

The mineralisation at Oere occurs within a geological sequence consisting of a package of meta-volcanoclastic material with intercalations of banded iron formation and graphitic layers. The whole sequence is intruded by late dolerite and early felsic intrusions, sub parallel to the shear corridor.

The mineralisation at Oere is controlled by a N-NE trending brittle shear corridor characterised by a moderate to steep dip (40° to 50° on average) and related to very fine-grained disseminated sulphides, largely pyrite, within a chlorite and silica alteration halo. Figure 14-12 shows the mineralisation lodes at Oere.

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LOGO

Source: Kibali Goldmines, 2021

Notes:

Low-Grade Mineralisation (Yellow), High-grade Mineralisation (Orange) and Optimised Pit Shell ($1,300/oz Au (Grey))

Looking Northwest

Figure 14-12 3D View of Oere Mineralisation

Topography

The Topography has been defined using a 2 m contoured LiDAR digital terrain model (DTM). This DTM covers the entire project area as required for mine design purposes. The surface was checked against known drill hole collar elevations, and an acceptable match was found.

Original data was captured in UTM WGS84 Zone 35N with elevation. For the purposes of converting the elevation from UTM to Mine Grid, +5,000 m was applied to the elevation. Once the conversion was completed, all data (i.e., drill holes, DTM, 3D wireframes, and block models) were checked to ensure that they all use the same mine grid system.

14.4 Bulk Density

Density values were measured from diamond drill core samples by applying the Archimedes Principle:

density = weight (in air) ÷ (weight (in air) – weight (in water)

Bulk density measurements were carried out on the fresh, transition, and saprolite material for both mineralised and waste rock using this water immersion method.

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A single density value is hard-coded to the block model for each estimation domain based on lithology and weathering status. The data is reviewed to remove any outliers that may exist and coded for the different mineralisation lodes. These outliers are noted to be mostly at the contacts of different weathering zones. All diamond drilling undertaken is sampled for density on routine basis. The sample selection is divided by the logged lithology, alteration and weathering type.

The depth of the weathering interfaces has been interpreted from drill hole geological logs and are divided into three categories:

●

Fresh rock is the unweathered underlying lithology.

●

Saprock and Transition materials are the first appearance of recognisable chips of the underlying lithology. These chips are clearly still oxidised but constitute material that requires drill and blast. Depending on the topographic morphology, this transitional zone may be quite thin.

●

Saprolite and Oxide material is a zone of red/orange coloured silt/clay fragments with no recognisable lithology, generally rich in clay.

The lithology used comes directly from the logged lithology fields in the database.

Where there are no density measurements, or the volume of density data is not sufficient to make an unbiased estimate for the subgroup, a substitute density is applied. This substitute density is obtained from the assigned value in the previous model or has been calculated using the density obtained from other lodes with similar rock and mineralisation characteristics.

Kibali Goldmines performs quarterly truck factor tests and broken density calibration using weighbridge and cavity monitoring scanners (CMS). This helps to verify assigned bulk density values against actual production data.

Table 14-33 to Table 14-37 are example summaries of the densities assigned for KCD, the largest deposit at Kibali.

Table 14-33 KCD 3000 Lodes Assigned Density Summary



Lodes		Weathering	Lithology	Mean Density (g/cm³)	No. of Samples	Assigned Density (g/cm³)

3000	3001	1	100	2.90	3,795	2.97
	3001	1	250	2.92	13	2.97
	3001	1	300	2.86	420	2.86
	3001	1	500	3.25	444	3.25
	3001	1	550	3.05	360	3.05
	3001	1	600	2.76	57	2.97
	3001	1	800	2.80	6	2.97
	3001	1	1000	2.72	2	2.72
	3002	1	100	2.84	316	2.99
	3002	1	300	2.82	4	2.97
	3002	1	500	3.34	98	3.34
	3002	1	600	2.73	29	2.73
	3002	1	800	2.87	8	2.84
	3002	1	1000	2.69	3	2.69
	3003	1	100	2.82	13	2.99
	3003	1	250	2.76	14	2.76

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Lodes		Weathering	Lithology	Mean Density (g/cm³)	No. of Samples	Assigned Density (g/cm³)

	3003	1	300	2.84	68	2.84
	3003	1	500	3.34	13	3.16
	3003	1	550	3.14	97	3.14
	3003	1	600	2.75	10	2.97
	3003	1	800	2.80	3	2.97
	3004	1	100	2.86	10	2.97
	3004	1	250	2.62	1	2.62
	3004	1	600	2.73	16	2.97
	3005	1	100	2.80	1	2.97
	3101	1	100	2.95	868	3.10
	3101	1	300	2.91	64	3.03
	3101	1	500	3.06	130	3.25
	3101	1	550	3.06	426	3.15
	3101	1	600	2.93	5	2.93
	3101	1	1000	2.73	1	2.73
	3102	1	100	2.94	83	3.24
	3102	1	250	2.87	1	2.87
	3102	1	300	2.81	1	3.08
	3102	1	500	3.38	36	3.27
	3102	1	550	3.32	3	3.32
	3102	1	600	2.73	5	2.73
	3102	1	800	2.91	2	3.26
	3103	1	100	2.97	16	3.10
	3103	1	300	2.90	6	3.08
	3103	1	500	3.27	19	3.25
	3103	1	550	3.11	28	3.15
	3105	1	100	2.75	1	3.10
	3105	1	300	3.00	8	3.05
	3105	1	550	3.23	39	3.23
	3106	1	100	2.93	247	2.93
	3106	1	250	2.93	6	2.93
	3106	1	300	2.91	43	2.91
	3106	1	500	3.34	57	3.34
	3106	1	550	3.16	74	3.16
	3106	1	600	2.89	4	3.08
	3107	1	100	2.98	3	3.10
	3107	1	250	2.88	3	2.88
	3107	1	300	2.86	4	2.86
	3107	1	500	2.89	6	2.89
	3107	1	550	2.93	8	2.93
	3107	1	600	2.78	2	2.78
	3107	1	800	2.82	1	2.82
	3110	1	100	2.85	50	3.08
	3110	1	300	2.92	1	2.92
	3110	1	500	3.27	1	3.19
	3111	1	100	2.90	11	3.17
	3111	1	500	3.29	18	3.25
	3111	1	600	2.93	1	2.93
	3112	1	500	3.33	1	3.25
	3112	1	550	3.19	8	3.15
	3112	1	600	2.71	1	2.71
	3114	1	100	2.75	2	3.10
	3114	1	600	2.76	4	2.76
	3115	1	100	2.73	2	3.10
	3116	1	100	2.66	2	3.10
	3117	1	100	2.99	26	3.10
	3119	1	100	2.83	25	3.10
	3121	1	100	2.97	156	2.97

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Lodes		Weathering	Lithology	Mean Density (g/cm³)	No. of Samples	Assigned Density (g/cm³)

	3121	1	500	3.13	9	3.13
	3121	1	800	2.89	1	2.89
	3122	1	100	2.88	20	2.88
	3122	1	500	2.93	13	2.93
	3122	1	800	2.85	2	2.85
	3001	2	300	2.20	1	2.20
	3002	2	100	2.43	10	2.38
	3002	2	300	2.07	1	2.07
	3002	2	500	2.46	2	2.46
	3002	2	600	1.99	2	1.99
	3003	2	100	2.21	5	2.38
	3003	2	250	2.10	9	2.10
	3003	2	300	2.23	16	2.30
	3003	2	500	1.87	1	2.67
	3003	2	550	2.58	15	2.33
	3003	2	600	2.54	3	2.54
	3004	2	100	2.24	6	2.38
	3004	2	250	2.80	1	2.80
	3004	2	300	2.23	5	2.23
	3004	2	600	2.54	7	2.54
	3005	2	100	2.35	3	2.38
	3107	2	100	2.16	3	2.38
	3107	2	300	2.33	3	2.33
	3002	3	100	1.77	45	1.77
	3002	3	300	1.91	3	1.91
	3002	3	550	1.99	1	1.99
	3002	3	600	1.70	2	1.70
	3003	3	100	1.74	6	1.72
	3003	3	250	1.84	2	1.84
	3003	3	300	1.81	8	1.67
	3003	3	550	2.10	4	1.72
	3003	3	600	1.51	1	1.51
	3003	3	800	1.15	1	1.15
	3004	3	100	1.66	35	1.67
	3004	3	600	1.82	12	1.82
	3005	3	100	1.53	5	1.72
	3005	3	250	1.94	1	1.94
	3102	3	500	1.78	5	1.78
	3108	3	100	1.46	1	1.72

Table 14-34 KCD 5000 Lodes Assigned Density Summary



Lodes		Weathering	Lithology	Mean Density (g/cm³)	No. of Samples	Assigned Density (g/cm³)

5000	5002	1	100	2.85	1,925	3.01
	5002	1	250	2.88	3	3.01
	5002	1	300	2.81	211	2.82
	5002	1	500	3.29	953	3.29
	5002	1	550	3.06	6	3.06
	5002	1	600	2.74	9	3.01
	5002	1	800	2.85	28	3.01
	5003	1	100	2.90	12	3.12
	5003	1	250	2.90	5	2.96
	5003	1	300	2.85	52	2.92
	5003	1	500	3.25	33	3.25
	5003	1	550	3.12	73	3.12

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Lodes		Weathering	Lithology	Mean Density (g/cm³)	No. of Samples	Assigned Density (g/cm³)

	5003	1	600	2.98	6	2.98
	5004	1	100	2.82	296	2.95
	5004	1	250	2.77	1	2.77
	5004	1	300	2.80	23	2.91
	5004	1	500	3.11	15	3.15
	5004	1	550	3.26	9	3.26
	5004	1	600	2.78	5	2.78
	5005	1	100	2.89	3,188	2.97
	5005	1	250	2.93	84	2.93
	5005	1	300	2.89	339	2.89
	5005	1	500	3.29	1,569	3.29
	5005	1	550	3.11	416	3.11
	5005	1	600	2.83	200	2.96
	5005	1	800	2.85	16	2.96
	5005	1	1000	2.75	1	2.75
	5006	1	100	2.85	70	2.96
	5006	1	250	2.83	9	2.96
	5006	1	300	2.82	67	2.85
	5006	1	550	2.95	18	3.05
	5007	1	100	2.96	5	3.12
	5007	1	300	2.80	2	2.90
	5101	1	100	2.92	1,046	3.24
	5101	1	250	2.76	1	3.20
	5101	1	300	3.03	9	2.98
	5101	1	500	3.39	1,585	3.39
	5101	1	550	3.14	66	3.15
	5101	1	600	2.91	32	3.18
	5101	1	800	2.93	12	2.93
	5101	1	1000	2.79	1	2.79
	5102	1	100	3.00	1,103	3.20
	5102	1	250	3.02	1	3.02
	5102	1	300	2.82	31	3.00
	5102	1	500	3.33	1,198	3.33
	5102	1	550	3.03	28	3.12
	5102	1	600	3.05	2	3.16
	5102	1	800	2.99	18	3.16
	5104	1	100	2.91	19	3.13
	5104	1	300	2.82	28	2.99
	5105	1	100	2.97	173	3.13
	5105	1	250	3.07	2	3.07
	5105	1	300	2.90	2	2.90
	5105	1	500	3.26	371	3.26
	5105	1	550	3.18	13	3.12
	5105	1	600	2.92	28	3.13
	5106	1	500	3.24	2	3.31
	5106	1	550	3.22	6	3.14
	5110	1	100	2.95	109	3.11
	5110	1	250	2.89	10	3.11
	5110	1	300	2.91	61	3.10
	5110	1	500	3.30	36	3.30
	5110	1	550	3.13	496	3.13
	5110	1	600	2.84	14	2.84
	5201	1	100	3.03	74	3.22
	5201	1	300	3.22	2	3.00
	5201	1	500	3.31	139	3.34
	5201	1	550	3.34	6	3.12
	5202	1	100	2.96	197	3.20
	5202	1	250	3.15	1	3.15

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Lodes		Weathering	Lithology	Mean Density (g/cm³)	No. of Samples	Assigned Density (g/cm³)

	5202	1	500	3.22	60	3.27
	5202	1	550	2.92	15	3.12
	5003	2	100	2.85	1	2.37
	5003	2	250	2.77	1	2.77
	5003	2	300	2.07	2	2.24
	5003	2	550	2.48	2	2.67
	5005	2	100	2.25	5	2.42
	5005	2	300	2.13	4	2.24
	5005	2	500	2.02	3	2.67
	5005	2	550	2.43	1	2.43
	5007	2	100	2.06	1	2.42
	5007	2	550	2.23	2	2.53
	5105	2	100	2.08	1	2.61
	5005	3	100	1.94	22	1.83
	5005	3	250	1.96	1	1.96
	5005	3	300	2.50	1	2.50
	5005	3	500	2.04	2	2.11
	5007	3	100	1.81	4	1.83
	5007	3	500	2.47	1	2.11
	5007	3	550	2.33	1	1.85
	5101	3	100	1.76	2	1.83
	5105	3	100	2.17	2	1.83

Table 14-35 KCD 9000 Lodes Assigned Density Summary



Lodes		Weathering	Lithology	Mean Density (g/cm³)	No. of Samples	Assigned Density (g/cm³)

9000	9004	1	100	2.86	9,471	2.96
	9004	1	250	2.94	20	2.95
	9004	1	300	2.82	273	2.83
	9004	1	500	3.34	5,164	3.35
	9004	1	550	2.84	24	2.84
	9004	1	600	2.80	203	2.95
	9004	1	800	2.89	20	2.95
	9004	1	1000	2.70	22	2.70
	9101	1	100	2.91	1,531	3.09
	9101	1	250	3.08	2	3.08
	9101	1	300	2.86	15	3.04
	9101	1	500	3.34	1,883	3.34
	9101	1	550	3.36	2	3.36
	9101	1	600	2.84	8	3.09
	9101	1	800	2.76	1	2.76
	9101	1	1000	2.83	10	2.83
	9102	1	100	2.92	48	3.09
	9102	1	500	3.40	538	3.40
	9102	1	600	3.37	4	3.37
	9103	1	100	2.91	10	3.09
	9103	1	500	3.33	45	3.35
	9103	1	600	0		3.09
	9104	1	100	2.92	29	3.38
	9104	1	500	3.49	504	3.49
	9104	1	600	0		3.09
	9105	1	100	2.89	1,065	3.14
	9105	1	300	2.84	14	3.04
	9105	1	500	3.40	861	3.40
	9105	1	550	2.68	2	2.68

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Lodes		Weathering	Lithology	Mean Density (g/cm³)	No. of Samples	Assigned Density (g/cm³)

	9105	1	600	3.06	6	3.09

	9105	1	800	2.78	2	2.78

	9105	1	1000	2.70	2	2.70

	9106	1	100	3.22	1	3.21

	9107	1	100	2.96	119	3.14

	9107	1	300	2.77	3	2.99

	9107	1	500	3.25	21	3.22

	9107	1	600	2.76	7	3.14

	9004	2	500	2.46	3	2.76

	9004	3	500	1.71	1	2.05

Table 14-36 KCD Waste (99999 & 800) Assigned Density Summary



Lodes		Weathering	Lithology	Mean Density (g/cm³)	No. of Samples	Assigned Density (g/cm³)

99999	99999	1	100	2.83	48,255	2.83

	99999	1	250	2.85	1,063	2.85

	99999	1	300	2.82	3,909	2.82

	99999	1	500	3.25	8,690	3.25

	99999	1	550	3.04	933	3.04

	99999	1	600	2.76	1,499	2.84

	99999	1	800	2.83	481	2.92

	99999	1	1000	2.73	116	2.73

	99999	2	100	2.43	230	2.43

	99999	2	250	2.44	45	2.44

	99999	2	300	2.29	141	2.29

	99999	2	500	2.56	32	2.56

	99999	2	550	2.43	36	2.53

	99999	2	600	2.34	37	2.34

	99999	2	800	2.37	5	2.43

	99999	3	100	1.60	420	1.60

	99999	3	250	1.62	41	1.62

	99999	3	300	1.86	91	1.86

	99999	3	500	2.01	48	2.01

	99999	3	550	1.87	57	1.87

	99999	3	600	1.67	39	1.67

	99999	3	800	1.49	7	1.63

800	800	1	100	2.81	466	2.81

	800	1	250	2.86	8	2.86

	800	1	300	2.89	23	2.89

	800	1	500	3.18	94	3.18

	800	1	550	2.91	8	2.91

	800	1	600	2.72	8	2.72

	800	1	800	2.90	2,897	2.92

	800	2	100	2.53	1	2.53

	800	2	800	2.54	24	2.54

	800	3	250	1.86	3	1.86

	800	3	550	1.76	1	1.76

	800	3	800	1.77	12	1.77

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Table 14-37 KCD 11000 Lodes Assigned Density Summary



Lodes	Weathering	Lithology	Mean Density (g/cm³)	No. of Samples	Assigned Density g/cm³)

11000	1	100	2.84	638	2.84

11000	1	200	2.84	66	2.84

11000	1	300			3

11000	1	500	3.54	14	3.39

11000	1	1000	2.75	10	2.81

11101	1	100	2.93	133	2.93

11101	1	500	3.46	11	3.56

11101	1	1000	2.85	1	2.81

11102	1	100	2.84	37	2.84

11102	1	500	3.57	46	3.57

11103	1	100	2.93	70	2.93

11103	1	500	3.58	10	3.58

11103	1	1000	2.86	1	2.81

14.5

Compositing

All samples were composited to 2 m lengths honouring domain boundaries.

Prior to selecting the composite length, the data was analysed using a histogram of sample length to identify the mode of length. The coefficient of variation, standard deviation, and mean plots were produced with several composite lengths to ensure that they remain stable and do not vary with compositing.

Compositing is completed in Maptek Vulcan software using the merge option for small composites, which adds the last composite (if smaller than the tolerance), to the previous interval. For Kibali, a tolerance length of 0.5 m is used. In deposits where the merge option was not selected, residual composites were filtered out and disregarded during estimation.

KCD

Compositing was undertaken at 2 m on the drillhole data. The 3000, 5000, 9000, and 11000 lodes show cumulative length distributions with approximately 90% of the composited data showing lengths of 2 m or more.

Figure 14-13 to Figure 14-16 illustrates a log histogram, log probability plot of the gold grades and the length distributions after compositing for the 2 m uncapped composites within the 3000, 5000, 9000 and 11000 lodes at KCD.

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Figure 14-13 KCD 3000 Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

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Figure 14-14 KCD 5000 Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

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Figure 14-15 KCD 9000 Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

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Figure 14-16 KCD 11000 Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

Sessenge and Sessenge SW

Compositing was undertaken at 2 m on the drillhole data. The samples were composited within each of the Sessenge lode domains, honouring the wireframe boundaries. The lodes show cumulative length distributions with approximately 98% of the composited data with lengths of 2 m or more. Residual composites less than 0.5 m were excluded from the estimation.

Figure 14-17 illustrates a log histogram, log probability plot of the gold grades and the length distributions after compositing for the 2 m uncapped composites within the mineralised domains at Sessenge.

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Figure 14-17 Sessenge Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

Gorumbwa

Compositing was undertaken at 2 m on the drillhole data. The samples were composited within each of the Gorumbwa lodes, honouring wireframe boundaries. The lodes show cumulative length distributions with approximately 95% of the composited data with lengths of 2 m or more.

Figure 14-18 illustrates a log histogram, log probability plot of the gold grades and the length distributions after compositing for the 2 m uncapped composites for Gorumbwa.

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Figure 14-18 Gorumbwa Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

Pakaka

Compositing was undertaken at 2 m on the drillhole data. The samples were composited within each of the Pakaka lodes, honouring wireframe boundaries. The lodes show cumulative length distributions with approximately 92% of the composited data with lengths of 2 m or more.

Figure 14-19 illustrates a log histogram, log probability plot of the gold grades and the length distributions after compositing for the 2 m uncapped composites at Pakaka.

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Figure 14-19 Pakaka Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

Kombokolo

Compositing was undertaken at 2 m on the drillhole data. The samples were composited within each of the Kombokolo lodes, honouring wireframe boundaries. The lodes show cumulative length distributions with approximately 89% of the composited data with lengths of 2 m or more.

Figure 14-20 illustrates a log histogram, log probability plot of the gold grades and the length distributions after compositing for the 2 m uncapped composites at Kombokolo.

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Figure 14-20 Kombokolo Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

Pamao and Pamao South

Compositing was undertaken at 2 m on the drillhole data. The samples were composited within each of the Pamao and Pamao South lodes, honouring wireframe boundaries. The lodes show cumulative length distributions with approximately 96% of the composited data with lengths of 2 m or more. At Pamao, residual composites were created with a length of less than 0.5 m, which were excluded from the estimation process.

Figure 14-21 and Figure 14-22 illustrate log histograms, log probability plots of the gold grades and the length distributions post compositing for the 2 m uncapped composites at Pamao and Pamao South respectively.

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Figure 14-21 Pamao Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

A comparison of the drilling data used in the 2021 estimate at Pamao versus 2018 report shows an increase of 345% in composite samples (from 2,182 in 2018 to 9,737 in 2021) resulting in a 7% increase in the capped mean grade (from 1.30 g/t in 2018 versus 1.39 g/t in 2021).

At Pamao South, a new deposit added to the Kibali Mineral Resource, a total of 1,498 composites were added with an overall capped mean grade of 1.84 g/t.

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Figure 14-22 Pamao South Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

Mengu Village

Compositing was undertaken at 2 m on the drillhole data. The samples were composited within the lode, honouring wireframe boundaries. The lode shows a cumulative length distribution with approximately 96% of the composited data with lengths of 2 m or more. Two residual composites smaller than 0.5 m were created, which were filtered out and excluded from the estimation.

Figure 14-23 illustrates a log histogram, log probability plot of the gold grades and the length distributions post compositing for the 2 m uncapped composites at Mengu Village.

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Figure 14-23 Mengu Village Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

Megi-Marakeke-Sayi

Compositing was undertaken at 2 m on the drillhole data. The samples were composited within the lodes at Megi-Marakeke-Sayi, honouring the wireframe boundaries. The lodes show a cumulative length distribution with approximately 95% of the composited data with lengths of 2 m or more.

Figure 14-24 illustrates a log histogram, log probability plot of the gold grades and the length distributions after compositing for the 2 m uncapped composites at Megi-Marakeke-Sayi.

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Figure 14-24 Megi-Marakeke-Sayi Log Histogram, Log Probability Plot, Length Histogram, and Cumulative

Length Distribution of 2 m Uncapped Composites

Kalimva-Ikamva

Compositing was undertaken at 2 m on the drillhole data. The samples were composited within the Kalimva-Ikamva lodes, honouring the wireframe boundaries. The lodes show a cumulative length distribution with approximately 94% of the composited data with lengths of 2 m or more.

Figure 14-25 illustrates a log histogram, log probability plot of the gold grades and the length distributions after compositing for the 2 m uncapped composites at Kalimva-Ikamva.

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Figure 14-25 Kalimva-Ikamva Log Histogram, Log Probability Plot, Length Histogram, and Cumulative

Length Distribution of 2 m Uncapped Composites

The Kalimva data show a total of 6,239 composites at an overall capped mean grade of 1.74 g/t.

The Ikamva data show a total of 2,602 composites at an overall capped mean grade of 1.70 g/t.

Mengu Hill

The lodes show a cumulative length distribution with approximately 86% of the composited data with lengths of 2 m or more. Figure 14-26 illustrates a log histogram and log probability plot of the gold grades for the 2 m uncapped composites within all mineralised domains at Mengu Hill.

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Figure 14-26 Mengu Hill Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

Aerodrome

The lodes show a cumulative length distribution with approximately 96% of the composited data with lengths of 2 m or more. Two residual composites smaller than 0.5 m were created, which were filtered out and excluded from the estimation.

Figure 14-27 illustrates a log histogram and log probability plot of the gold grades for the 2 m uncapped composites within all mineralised domains at Aerodrome. A total of 1,629 composites were generated with a capped mean grade of 1.45 g/t.

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Figure 14-27 Aerodrome Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

Oere

Compositing was undertaken at 2 m on the drillhole data. The samples were composited within the lodes, honouring wireframe boundaries. The lode shows a cumulative length distribution with approximately 96% of the composited data with lengths of 2 m or more. Six residual composites smaller than 0.5 m were created, which were filtered out and excluded from the estimation.

Figure 14-28 illustrates a log histogram and log probability plot of the gold grades for the 2 m uncapped composites within all mineralised domains at Oere. A total of 1,918 composites at a capped mean grade of 1.73 g/t were generated.

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Figure 14-28 Oere Log Histogram, Log Probability Plot, Length Histogram, and Cumulative Length

Distribution of 2 m Uncapped Composites

14.6

Treatment of High-Grade Outliers (Top Capping)

Top capping was applied to reduce the effect of high-grade outliers during resource estimation. Generally, the top capping occurred within the top percentile ranges, between the 95^th to 99.9^th percentiles within the individual mineralised lodes. A multi-variate analysis method was used to select the top cap, analysing a combination of histograms, probability plot, and disintegration.

In addition, high-grade yields were occasionally used to further restrict the distance of influence of significant gold grades, with thresholds typically aligned with values observed in the histogram. Above a value threshold and beyond a specified local distance (typically drill spacing), composites are not included in the Mineral Resource estimate to limit smearing.

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KCD

At the KCD 3000 lodes, a total of 70,657 samples were included in the database for top capping analysis. In total, 237 samples were top capped, ranging between 9.32 g/t Au and 112.65 g/t Au for the high-grade lodes. Top capping reduced the average mean grade from 2.65 g/t Au to 2.49 g/t Au and resulted in a reduction of the coefficient of variation from 2.40 to 1.89. In total, the metal reduction was -6% overall.

For the 5000 lodes, a total of 126,710 samples were included in the database for top capping analysis. In total, 168 samples were top capped, ranging between 7.70 g/t Au and 122.45 g/t Au. Top capping reduced the average mean grade from 4.66 g/t Au to 4.55 g/t Au and resulted in a reduction of the coefficient of variation from 2.77 to 1.68. In total, the metal reduction was -2% overall.

For the 9000 lodes, a total of 58,137 samples were included in the database for top capping analysis. In total, 83 samples were top capped, ranging between 12.00 g/t Au and 77.95 g/t Au. Top capping reduced the average mean grade from 2.70 g/t Au to 2.66 g/t Au and resulted in a reduction of the coefficient of variation from 2.16 to 1.86. In total, the metal reduction was -1% overall.

For the 11000 lodes, about 2,135 samples were included in the database for top capping analysis. In total, 10 samples were top capped, ranging between 12.08 g/t Au and 52.67 g/t Au. Top capping reduced the average mean grade from 3.40 g/t Au to 3.29 g/t Au and resulted in a reduction of the coefficient of variation from 1.87 to 1.06. In total, the metal reduction was -3% overall.

A detailed breakdown of the statistical analysis for top capping at KCD is presented in Table 14-38 to Table 14-41.

Table 14-38 KCD 3000 Lodes Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

3001	14,573	0	69.68	0.91	1.96	26.72	0.9	1.72	11	-1%

3002	15,451	0.01	132.88	1.56	2.45	43.1	1.52	2	20	-3%

3003	20,425	0	499.7	1.84	3.35	51.5	1.75	2.39	50	-5%

3004	4,774	0.01	33	2.03	1.72	27	2.03	1.71	9	0%

3006	3	0.07	0.22	0.16	0.5	-	0.16	0.5	0	0%

3101	4,302	0.01	123.38	5.97	1.26	64.39	5.94	1.22	10	-1%

3102	3,769	0.01	180	6.02	1.52	60.1	5.87	1.31	23	-2%

3103	173	0.05	37.65	4.03	1.06	9.67	3.62	0.67	8	-10%

3105	449	0.01	75.7	3.77	2.04	9.32	2.69	0.9	27	-29%

3106	2,651	0.01	514.37	6.53	2.78	80	6	1.7	12	-8%

3107	2,325	0.03	354.78	9.97	1.94	100	16	9.51	16	4.6%

3108	691	0.02	25.49	4.23	0.9	22	4.21	0.88	4	0%

3109	108	0.17	20.6	5.12	0.79	20	5.12	0.78	1	0%

3110	100	0.06	58.15	9.56	1.27	22.3	7.74	0.91	10	-19%

3111	201	0.02	123.2	4.7	2.06	10.86	3.74	0.77	14	-20%

3112	65	0.01	58.81	6.42	1.64	12.46	4.41	0.89	7	-31%

3114-3117	122	0.03	34.9	4.23	1.27	10.85	3.6	0.77	6	-15%

3119	52	0.01	20.55	2.64	1.4	-	2.64	1.4	0	0%

3120	45	0.04	27.58	4.51	1.34	-	4.51	1.34	0	0%

3121	319	0.01	193	11.59	1.89	112.75	11.18	1.72	4	-4%

3122	59	0.6	173.95	17.9	1.65	74.41	15.55	1.3	5	-13%

Total	70,657	-	-	2.65	-	-	2.491	-	237	-6%

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Table 14-39 KCD 5000 Lodes Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

5002	8,429	0	63.81	0.99	1.96	25.8	0.98	1.7	5	-1%

5003	13,140	0	240	2.65	2.45	77.1	2.6	2.2	23	-2%

5004	1,352	0	120.83	0.48	8.15	19.56	0.38	4.78	6	-21%

5005	34,291	0	433.59	1.3	3.36	61.5	1.27	2.34	13	-2%

5006	847	0.01	240	1.52	5.89	12.9	1.09	1.71	8	-28%

5007	5,214	0.01	178	2.31	2.32	50.3	2.24	1.88	12	-3%

5101	16,299	0	3,008.00	7.36	3.48	119	7.11	1.18	18	-3%

5102	8,505	0.01	184.17	6.17	1.11	57.78	6.1	0.94	12	-1%

5104	360	0.03	240	10.63	2.08	55.93	9.12	1.31	7	-14%

5105	6,713	0.01	727.02	5.63	2.77	86.12	5.32	1.38	10	-6%

5106	76	0.01	14.68	2.83	0.95	7.7	2.65	0.81	4	-6%

5110	1,684	0.02	540	7.36	2.39	61	6.79	1.37	11	-8%

5201	1,657	0.08	194.44	16.87	0.73	100	16.75	0.66	4	-1%

5202	890	0.02	340	18.92	1.12	100	18.21	0.8	7	-4%

5101& 5201	17,948	0.01	3,008.00	8.24	3.02	122.45	8.17	1.13	19	-1%

5102& 5202	9,305	0.01	340	7.41	1.35	119.36	7.34	1.17	9	-1%

Total	126,710	-	-	4.66	-	-	4.550	-	168	-2%

Table 14-40 KCD 9000 Lodes Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

9004	40,537	-	103.65	0.94	2.48	35.41	0.93	2.16	24	-1%

9101	7,179	0.01	242.66	7.1	1.31	72.9	7.05	1.22	9	-1%

9102	1,306	0.01	87.37	5.11	1.08	28.31	5.03	0.97	9	-2%

9103	101	0.17	42.65	7.12	1.02	19.7	6.58	0.81	5	-8%

9104	818	0.03	205.87	7.46	1.41	37.8	7.04	0.91	9	-6%

9105	7,426	0.01	579.53	6.8	1.56	77.95	6.71	1.21	7	-1%

9106	45	1.6	19.36	5.73	0.67	12	5.47	0.58	3	-5%

9107	725	0.01	200.68	5.2	2.1	20.43	4.31	0.97	17	-17%

Total	58,137	-	-	2.7	-	-	2.660	-	-	-1%

Table 14-41 KCD 11000 Lodes Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

11000	1,385	0.01	26.80	1.33	1.33	12.08	1.29	1.08	5	-3%

11101	348	0.08	125.15	6.84	1.34	34.60	6.59	0.98	3	-5%

11102	214	0.04	119.41	7.63	1.42	52.67	7.31	1.12	1	-4%

11103	188	0.12	58.55	7.46	1.02	34.7	7.33	0.94	1	-2%

Total	2,135	-	-	3.40	-	-	3.29	-	10	-3%

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Sessenge and Sessenge SW

At Sessenge, a total of 15,551 samples were included in the database for top capping analysis. In total, 40 samples were capped, ranging between 13.00 g/t Au and 27.00 g/t Au. Top capping reduced the average mean grade from 2.19 g/t Au to 2.16 g/t Au and resulted in a reduction of the coefficient of variation from 2.19 to 1.22. The total metal reduction was -1%. A detailed breakdown of the statistical analysis for top capping at Sessenge is presented in Table 14-42.

Table 14-42 Sessenge Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

9004	10,878	0.01	60.1	1.47	1.47	27	1.46	1.37	10	-1%

9008	622	0.01	31.9	1.91	1.91	18	1.86	1.51	5	-3%

9009	321	0.02	5.06	1.06	1.06	-	1.06	1.06	-	-

9102	1,912	0.03	68.3	3.91	3.91	19	3.81	0.78	10	-3%

9103	1,431	0.04	41.4	4.82	4.82	22	4.79	0.72	5	-1%

9104	297	0.22	36.1	5.57	5.57	19	5.42	0.71	6	-3%

9105	90	0.06	21.9	5.12	5.12	13	4.85	0.69	4	-5%

Total	15,551	-	-	2.19	-	-	2.14	-	40	-1%

Gorumbwa

At Gorumbwa, a total of 17,882 samples were included in the database for top capping analysis. In total, 2,650 samples were capped, ranging between 7.68 g/t Au and 62.40 g/t Au. Top capping reduced the average mean grade from 2.75 g/t Au to 2.65 g/t Au and resulted in a reduction of the coefficient of variation from 1.93 to 1.66. The total metal reduction was -3%. A detailed breakdown of the statistical analysis for top capping at Gorumbwa can be found in Table 14-43.

Table 14-43 Gorumbwa Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

1001	2,824	0.01	43.3	2.22	1.48	27.1	2.21	1.42	9	0%

1003	1,294	0.01	66.4	2.13	1.85	20.4	2.04	1.49	10	-4%

1004	8,005	0.01	175	3.47	2.48	62.4	3.32	2.09	32	-4%

1006	1,221	0.03	80.5	2.14	2.12	23.5	2.01	1.65	9	-6%

1008	2,892	0.01	55.2	2.4	1.3	24.2	2.38	1.22	4	-1%

1013	281	0.04	13	1.12	1.4	7.68	1.08	1.12	4	-4%

1015	184	0.04	16	1.42	1.36	-	1.42	1.36	-	-

9001	1,181	0.01	18	1.94	0.98	10.3	1.91	0.91	6	-2%

Total	17,882	-	-	2.75	.93	-	2.66	-	74	-3%

Pakaka

At Pakaka, a total of 22,388 samples were included in the database for top capping analysis. In total, 35 samples were capped, ranging between 25.00 g/t Au and 50.00 g/t Au. Top capping reduced the average mean grade from 2.37 g/t Au to 2.33 g/t Au and resulted in a reduction of the coefficient of variation from 1.38 to 1.24. The total metal reduction was -2%. A detailed breakdown of the statistical analysis for Top Capping at Pakaka can be found in Table 14-44.

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Table 14-44 Pakaka Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

1001	16,597	0.01	65.42	1.36	1.48	25	1.35	1.32	18	-1%

1007	164	0.03	6.99	1.21	1.00	-	1.21	1.00	-	0%

1101	2,777	0.01	89.42	6.84	0.97	34	6.76	0.90	12	-1%

1102	102	0.37	520.00	10.16	5.06	45	5.5	1.41	1	-46%

1103	652	0.10	36.69	3.22	0.95	-	3.22	0.95	-	0%

1105	1,768	0.01	60.00	3.97	1.21	50	3.96	1.18	4	0%

1106	328	0.01	32.48	3.17	0.93	-	3.17	0.93	-	0%

Total	22,388	-	-	2.37	1.38	-	2.330	-	35	-2%

Kombokolo

At Kombokolo, a total of 10,547 samples were included in the database for top capping analysis. In total, 72 samples were top capped between 14.00 g/t Au and 39.60 g/t Au. Top capping reduced the average mean grade from 2.87 g/t Au to 2.73 g/t Au and resulted in a reduction of the coefficient of variation from 1.92 to 1.24. The total metal reduction was -5%. A detailed breakdown of the statistical analysis for top capping at Kombokolo can be found in Table 14-45.

Table 14-45 Kombokolo Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

1001	4,852	0.01	72.20	1.33	1.97	18.20	1.28	1.49	15	-4%

1002	2,459	0.01	152.00	1.41	2.61	14.00	1.33	1.17	15	-6%

1003	38	0.01	4.10	1.07	0.86	-	1.07	0.86	0	0%

1004	12	0.11	3.13	0.95	0.89	-	0.95	0.89	0	0%

1005	33	0.01	4.00	1.85	0.86	-	1.85	0.86	0	0%

1006	11	0.50	5.41	2.33	0.69	-	2.33	0.69	0	0%

1007	3	0.02	2.63	1.28	1.03		1.28	1.03	0	0%

1101	2,466	0.01	239.67	6.67	1.35	39.60	6.42	0.97	17	-4%

1102	673	0.06	124.40	5.56	1.32	16.00	4.99	0.78	25	-10%

Total	10,547	-	-	2.87	-	-	2.73	-	72	-5%

Pamao and Pamao South

At Pamao, a total of 9,737 samples were included in the database for top capping analysis. In total, 32 samples were top capped between 4.50 g/t Au and 25.00 g/t Au. Top capping reduced the average mean grade from 1.41 g/t Au to 1.39 g/t Au and resulted in a reduction of the coefficient of variation from 1.30 to 0.99. The total metal reduction was -1%. A detailed breakdown of the statistical analysis for top capping at Pamao can be found in Table 14-46.

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Table 14-46 Pamao Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

2001	1,400	0.01	10.5	0.92	0.83	5.9	0.92	0.79	3	0%

2002	13,537	0.01	29.6	1.09	1.14	22.99	1.09	1.12	18	0%

2003	2,424	0.01	53.1	1.21	1.44	10.9	1.18	1.13	4	-2%

2102	2,106	0.02	33.6	3.81	0.82	23.39	3.79	0.79	5	-1%

Total	19,467	-	-	1.39	-	-	1.38	-	30	-1%

At Pamao South, a total of 1,498 samples were included in the database for top capping analysis. In total, 18 samples were treated and ranged between 4.50 g/t Au and 15.73 g/t Au. The top capping reduced the average mean grade from 2.03 g/t Au to 1.84 g/t Au and resulted in a reduction of the coefficient of variation from 2.13 to 1.30. In total, the metal reduction was -9% overall, mainly due to some extreme maximum values within a predominantly low-grade deposit. A detailed breakdown of the statistical analysis for top capping at Pamao can be found in Table 14-47.

Table 14-47 Pamao South Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

1001	1,009	0.01	122	1.93	2.4	20	1.8	1.34	4	-7%

1002	2,137	0.01	86.69	1.92	1.93	33.1	1.87	1.62	5	-3%

1003	436	0.04	39.4	1.89	1.72	13.2	1.75	1.26	4	-7%

Total	3,582			1.92			1.84		13	-4%

Mengu Village

At Mengu Village, a total of 256 samples were included in the database for top capping analysis. In total, 6 samples were top capped at 5.88 g/t Au. Top capping reduced the average mean grade from 1.62 g/t Au to 1.53 g/t Au and resulted in a reduction of the coefficient of variation from 1.06 to 0.78. The total metal reduction was -6%. A detailed breakdown of the statistical analysis for top capping at Mengu Village is presented in Table 14-48.

Table 14-48 Mengu Village Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

1004	256	0.08	18	1.62	1.06	5.88	1.53	0.78	6	-6%

Total	256	-	-	1.62	-	-	1.53	-	6	-6%

Megi-Marakeke-Sayi

At Megi-Marakeke-Sayi, a total of 6,239 samples were included in the database for top capping analysis. In total, 54 samples were top cut between 7.26 g/t Au and 20.10 g/t Au. Top capping reduced the average mean grade from 1.69 g/t Au to 1.60 g/t Au and resulted in a reduction of the coefficient of variation from 1.28 to 1.05. The total metal reduction was -5%. A detailed breakdown of the statistical analysis for top capping at Megi-Marakeke-Sayi is presented in Table 14-49.

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Table 14-49 Megi-Marakeke-Sayi Village Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

1001	529	0.02	32	1.5	1.67	7.26	1.33	1.1	16	-11%

1002	934	0.01	20.4	1.31	1.19	9.27	1.28	1.17	6	-2%

1003	418	0.01	67.2	2.29	1.97	14.1	2.08	1.27	7	-9%

1004	1,928	0.01	19.7	1.46	1.07	11.1	1.44	0.98	5	-1%

1005	2,015	0.01	40.9	1.31	1.27	12.9	1.29	1.06	5	-2%

1101	127	0.04	193	6.71	2.57	14.31	5.07	0.76	8	-24%

1102	141	0.06	42.6	5.11	1.11	20.1	4.9	0.84	4	-4%

1105	147	0.09	10.1	3.56	0.57	9.46	3.56	0.52	3	0%

Total	6,239	-	-	1.69	-	-	1.60	-	54	-5%

Kalimva-Ikamva

At Kalimva, a total of 6,366 samples were included in the database for top capping analysis. In total, 28 samples were top capped between 6.66 g/t Au and 20.60 g/t Au. Top capping reduced the average mean grade from 1.78 g/t Au to 1.74 g/t Au and resulted in a reduction of the coefficient of variation from 1.08 to 0.91. The total metal reduction was -2%.

At Ikamva, a total of 2,602 samples were included in the database for top capping analysis. In total, 11 samples were top capped between 5.60 g/t Au and 15.90 g/t Au. Top capping reduced the average mean grade from 1.71 g/t Au to 1.70 g/t Au and resulted in a reduction of the coefficient of variation from 0.9 to 0.89. The total metal reduction was -1%.

A detailed breakdown of the statistical analysis for top capping at Kalimva and Ikamva is presented in Table 14-50 and Table 14-51 respectively.

Table 14-50 Kalimva Village Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

1001	4,440	0.000	25.85	0.95	1.10	6.66	0.94	0.95	19	-2%

1101	1,926	0.01	69.39	3.67	1.01	20.6	3.60	0.83	9	-2%

Total	6,366	-	-	1.78	-	-	1.74	-	28	-2%

Table 14-51 Ikamva Village Top Capping Analysis



Domain	No of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No of Samples Capped	% Metal Reduction

2001	1,960	0.01	7.35	0.83	0.95	5.6	0.83	0.95	3	0%

2101	642	0.01	23.9	4.37	0.74	15.9	4.33	0.71	8	-1%

Total	2,602	-	-	1.71	-	-	1.70	-	11	-1%

Mengu Hill

At Mengu, a total of 17,725 samples were included in the database for top capping analysis. In total, 36 samples were top capped between 2.50 g/t Au and 61.90 g/t Au. Top capping reduced the average mean grade from 3.21 g/t Au to 3.19 g/t Au and resulted in a reduction of the coefficient of variation from 1.39 to 1.34. A detailed breakdown of the statistical analysis for top capping at Mengu is presented in Table 14-52.

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Table 14-52 Mengu Hill Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

1001	9,812	0.01	138.38	1.29	1.59	34.40	1.28	1.53	3	-1%

1002	135	0.03	4.89	1.15	0.93	2.50	1.03	0.79	14	-10%

1101	7,601	0.02	117	5.71	1.16	61.90	5.69	1.11	6	0%

1102	26	1	31.09	7.94	1	8.01	5.69	1.11	6	-28%

1103	151	0.05	11.3	2.95	0.7	6.69	2.81	0.59	7	-5%

Total	17,725	-	-	3.21	-	-	3.19	-	36	-1%

Aerodrome

At Aerodrome, a total of 1,629 samples were included in the database for top capping analysis. In total, 16 samples were top cut between 1.53 g/t Au and 1.45 g/t Au. Top capping reduced the average mean grade from 1.53 g/t Au to 1.45 g/t Au and resulted in a reduction of the coefficient of variation from 1.43 to 1.03. The total metal reduction was -5%. A detailed breakdown of the statistical analysis for top capping at Aerodrome is presented in Table 14-53.

Table 14-53 Aerodrome Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

2001	1,087	0.25	13.8	1.45	0.95	6.19	1.43	0.9	10	-1%

2002	542	0.01	122	1.69	2.4	12.8	1.50	1.29	6	-11%

Total	1,629	-	-	1.53	-	-	1.45	-	16	-5%

Oere

At Oere, a total of 1,918 samples were included in the database for top capping analysis. In total, 2 samples were top cut between 17.80 g/t Au and 30.90 g/t Au. Top capping reduced the average mean grade from 1.74 g/t Au to 1.73 g/t Au and resulted in a reduction of the coefficient of variation from 1.25 to 1.09. The total metal reduction was negligible. A detailed breakdown of the statistical analysis for top capping at Oere is presented in Table 14-54.

Table 14-54 Oere Top Capping Analysis



Domain	No. of Samples	Min Raw (g/t Au)	Max Raw (g/t Au)	Mean Raw (g/t Au)	CV Raw	Capped (g/t Au)	Mean Capped (g/t Au)	CV Capped	No. of Samples Capped	% Metal Reduction

1001	1,661	0.01	40.70	1.32	1.32	17.8	1.31	1.13	2	-1%

1101	257	0.01	30.90	4.48	0.83	30.9	4.48	0.83	0	0

Total	1,918	-	-	1.74	-	-	1.730	-	2	-1%

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14.7 Variography

Exploratory data analysis (EDA) was conducted using Snowden Supervisor statistical software, with all modelling and estimation has been completed in Maptek Vulcan. Values less than the detection limit (<0.01 g/t Au) were replaced by half the limit (0.005 g/t Au).

Variography has been used to analyse the spatial continuity and relation within the individual mineralised lodes and to determine the appropriate search strategy and estimation parameters. The Variogram modelling process involved the following steps:

●

A normal score transform was applied to all data prior to undertaking variography on the top capped, declustered composite dataset. The data was transformed into a normal score space using Snowden Supervisor.

●

Calculate and model the omni-directional or down hole variogram to characterise the nugget effect.

●

Systematically calculate orientated variograms in three dimensions to identify the plane of greatest continuity.

●

Calculate a variogram fan within the plane of greatest continuity to identify the direction of maximum continuity within this plane.

●

Model experimental variogram in the direction of maximum continuity and the orthogonal directions.

●

Apply a back transform to all variogram models to obtain the appropriate variogram models for interpolation of the capped composite data.

Within the domains, the relative nugget effect ranged between 18% and 58%, with most of the deposits showing nuggets of 25% to 35%, indicating a low to moderate grade variability, which is typical for these type of gold deposits. Variogram ranges interpreted, were typically significantly greater than the average drill hole spacing.

In some areas which contain infill grade control drilling, such as KCD, variograms were required for nested structures, thus multiple ranges were used.

Where an individual domain has insufficient samples to undertake variography, the variography parameters from a comparative domain with a similar trend was used and the orientation adjusted to match the domain with insufficient data.

Variogram Validation

Prior to interpolation runs, each semi-variogram model is cross validated to ensure that any bias in estimated grades compared to the actual sample grades is minimal. This was checked by estimating a grade value at each composite sample point, which ignored said sample point. The resulting grade is compared to the actual sample grade in the same location and is plotted on a scatter plot to establish a possible trend or bias and relative standard error. In most cases, there is an expected level of smoothing in an estimated grade compared to the actual sample grade, but overall, estimated grades and sample grades match well and conditional bias is minimised.

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KCD

At KCD, the three lodes return significantly different variography results. The lodes are broadly categorised as the upper 3000 lodes, 5000 lodes, and at the deeper 9000 lodes.

3000 Lodes

At the KCD 3000 lodes, the relative nugget effect ranged from 10% to 34% for the high-grade domains, and from 10% to 21% for the low-grade domains. Figure 14-29 illustrates an example of the KCD 3101 normal score and nested back transformed variogram models.

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Figure 14-29 KCD Lode 3101 Normal Score Variogram Models and Nested Back Transformed Variogram

Model

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5000 Lodes

At the KCD 5000 lodes, the relative nugget effect ranged from 15% to 21% for the high-grade domains, and from 10% to 17% for the low-grade domains. Figure 14-30 illustrates an example of the KCD 5102 and 5202 normal score and nested back transformed variogram models.

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Figure 14-30 KCD Lode 5102 and 5202 Normal Score Variogram Models and Nested Back Transformed

Variogram Model

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9000 Lodes

At the KCD 9000 lodes, the relative nugget effect ranged from 15% to 30% for the high-grade domains and was 12% for the low-grade domains. Figure 14-31 illustrates an example of the KCD 9105 normal score and nested back transformed variogram models.

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Figure 14-31 KCD Lode 9105 Normal Score Variogram Models and Nested Back Transformed Variogram

Model

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11000 Lodes

The 11000 lode is a new addition to the Mineral Resource at Kibali. In this regard, data density is lower and so lodes with the same orientation were combined for variography. The relative nugget effect was 21%, with a resultant back transformed nugget of 30%. Lode 1101 differed in orientation and is linked to 5000 lode, particularly lode 5101. Therefore, the variogram model for lode 5101 was used for the estimation of lode 11101. Figure 14-32 illustrates the KCD 11000 lodes normal score and nested back transformed variogram models.

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Figure 14-32 KCD Lode 11000 Normal Score Variogram Models and Nested Back Transformed Variogram

Model

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Sessenge and Sessenge SW

At Sessenge, the relative nugget effect ranged from 20% to 26% for both high-grade and low-grade domains, with resultant back transformed nuggets ranging between 29% and 32%. Figure 14-33 illustrates an example of the Sessenge 9004 normal score and nested back transformed variogram models.

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Figure 14-33 Sessenge Lode 1004 Normal Score Variogram Models and Nested Back Transformed Variogram

Model

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Gorumbwa

At Gorumbwa, the relative nugget effect ranged from 20% to 26%, with resultant back transformed nuggets ranging between 25% and 35%. Figure 14-34 illustrates an example of the Gorumbwa 1004 normal score and nested back transformed variograms models.

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Figure 14-34 Gorumbwa Lode 1004 Normal Score Variogram Models and Nested Back Transformed

Variogram Model

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Pakaka

At Pakaka, the relative nugget effect ranged from 10% to 22%. Figure 14-35 illustrates an example of the Pakaka 1001 normal score and nested back transformed variograms models.

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Figure 14-35 Pakaka Lode 1001 Normal Score Variogram Models and Nested Back Transformed Variogram

Model

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Kombokolo

At Kombokolo, the relative nugget effect ranged from 25% to 46%. Kombokolo is observed to have a higher nugget effect than the other domains at Kibali. Figure 14-36 illustrates an example of the Kombokolo 1101 and 1002 normal score and nested back transformed variogram models.

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Figure 14-36 Kombokolo Lode 1101 and 1102 Normal Score Variogram Models and Nested Back

Transformed Variogram Model

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Pamao and Pamao South

At Pamao, the relative nugget effect ranged from 20% to 25%, with resultant back transformed nuggets ranging between 29% and 33%. Figure 14-37 illustrates an example (domain 2001) of the Pamao normal score and nested back transformed variograms models.

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Figure 14-37 Pamao Lode 2001 Normal Score Variogram Models and Nested Back Transformed Variogram

Model

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At Pamao South, the relative nugget effect was modelled at 25%, with resultant back transformed nuggets ranging between 34% and 36%. Figure 14-38 illustrates an example of the Pamao South normal score and nested back transformed variograms models for domain 1001 and 1002, which were combined for variography.

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Figure 14-38 Pamao South Lode 1001 and 1002 Normal Score Variogram Models and Nested Back

Transformed Variogram Model

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Megi-Marakeke-Sayi

At Megi-Marakeke-Sayi, the relative nugget effect ranged from 21% to 25% for both high-grade and low-grade domains. Figure 14-39 illustrates an example of the Megi-Marakeke-Sayi normal score and nested back transformed variograms models for domain 1001 and 1101 combined.

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Figure 14-39 Megi-Marakeke-Sayi Lode 1001 and 1101 Normal Score Variogram Models and Nested Back

Transformed Variogram Model

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Kalimva-Ikamva

At Kalimva-Ikamva, the relative nugget effect ranged from 33% to 36% for both high-grade and low-grade domains. Figure 14-40 illustrates an example of the Kalimva-Ikamva normal score and nested back transformed variograms models for domain 1001.

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Figure 14-40 Kalimva-Ikamva Lode 1001 Normal Score Variogram Models and Nested Back Transformed

Variogram Model

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Mengu Hill

At Mengu Hill, the relative nugget effect ranged from 14% to 15% for both high-grade and low-grade domains. Figure 14-41 illustrates an example of the Mengu Hill normal score and nested back transformed variograms models for domain 1001.

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Figure 14-41 Mengu Hill Lode 1001 Normal Score Variogram Models and Nested Back Transformed

Variogram Model

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Aerodrome

At Aerodrome, the relative nugget effect was modelled at 29%, with a resultant back transformed nugget of 33%. Aerodrome consists of two mineralisation domains, 2001 and 2002. Not enough data points exist in lode 2002 to define a decent variogram structure. The variogram model from 2001 was used in the estimation of domain 2002. Figure 14-42 shows the Aerodrome 2001 normal score and nested back transformed variograms models.

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Figure 14-42 Aerodrome Lode 2001 Normal Score Variogram Models and Nested Back Transformed

Variogram Model

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Oere

At Oere, the relative nugget effect was modelled at 22%, with a resultant back transformed nugget of 28%. Figure 14-43 shows the combined normal score and nested back transformed variogram models for lode 1001 and lode 1101.

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Figure 14-43 Oere Lode 1001 and 1101 Normal Score Variogram Models and Nested Back Transformed

Variogram Model

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Mengu Village

The Mengu Village Mineral Resource is still in the Inferred category, with limited data, making it difficult to generate robust variograms. The one lode (1004) modelled at Mengu Village extends from the Megi-Marakeke-Sayi deposit. The variograms from Megi-Marakeke-Sayi for lode 1004 were used in the estimation of Mengu Village.

14.8 Block Model Estimation

Ordinary Kriging (OK) was used to estimate all Mineral Resources. Quantitative Kriging Neighbourhood Analysis (QKNA) was applied to help to determine the minimum number of samples, search radius, and block discretisation for each domain. Almost all domains use hard boundaries to ensure that separate grade populations do not influence the grades (exception for between high and very high-grade domains at KCD which employ a semi-hard boundary).

In certain cases, the input estimation parameters were adjusted following block model validation checks employed by Kibali, which involved visual checks, swath plots, decluster plots, change of support checks and global mean block model versus data comparisons. Figure 14-44 illustrates the results of the QKNA for domain 5101/5201 at KCD.

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Figure 14-44 QKNA for KCD Domain 5101 and 5201 Underground GC Zone

At KCD, Sessenge, Sessenge SW, Gorumbwa, Pamao, Pamao South, Aerodrome, and Oere deposits, the search ellipse in run 1 was setup in line with the second ranges of the variogram models, which typically corresponded to 80% of the total sill. The typical variograms modelled in

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KCD show extended ranges associated with the last C3 structure (typically, representative of the final 20% of the sill). The search ellipse in run 2 was setup in line with the full ranges of the variogram models. The third pass was one and a half times the full variograms model ranges and the fourth pass was double the variogram model range. In rare cases, a fifth pass was employed to ensure that a limited amount of edge blocks received grade estimates.

In some cases, for example Pakaka, Kombokolo, Megi-Marakeke-Sayi, Kalimva, Ikamva, Mengu Hill or where drill spacing is too wide to inform the variogram at short spacing (Mengu Village), a similar multiple pass interpolation was employed, but slightly different search distances in relation to the variograms were used.

Due to the large number of domains within each of the Kibali deposits, a small subset of the KCD QKNA parameters is shown in Table 14-55.

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Table 14-55 QKNA Parameters for KCD 5101 Domain



Domain	OP/ UG	Block Size (m)			Run	Search Radius (m)			No. of Samples		Max Samples Per Drill Hole	Discretisation			High- Grade Yield (g/t Au)	High-Grade Yield Restriction
Domain	OP/ UG	X	Y	Z	Run	Y	X	Z	Min	Max	Max Samples Per Drill Hole	X	Y	Z	High- Grade Yield (g/t Au)	X	Y	Z

5101 Infill GC	OP	5	5	2.5	1	35	15	10	9	15	3	5	5	5	62.63	10	10	5
					2	70	30	20	9	12	3	5	5	5	62.63	10	10	5
					3	105	45	30	6	12	-	5	5	5	62.63	10	10	5
					4	140	60	40	4	12	-	5	5	5	62.63	10	10	5
					5	525	225	150	4	12	-	5	5	5	62.63	10	10	5

5101 Infill GC	UG/OP	5	10	5	1	35	15	10	12	18	4	5	5	5	62.63	10	10	5
					2	70	30	20	10	16	4	5	5	5	62.63	10	10	5
					3	105	45	30	6	12	-	5	5	5	62.63	10	10	5
					4	140	60	40	4	12	-	5	5	5	62.63	10	10	5
					5	525	225	150	4	12	-	5	5	5	62.63	10	10	5

5003 Advance GC	UG	5	10	5	1	35	15	10	12	18	4	5	5	5	62.63	10	10	5
					2	70	30	20	10	16	4	5	5	5	62.63	10	10	5
					3	105	45	30	6	12	-	5	5	5	62.63	10	10	5
					4	140	60	40	4	12	-	5	5	5	62.63	10	10	5
					5	525	225	150	4	12	-	5	5	5	62.63	10	10	5

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14.9 Block Models

Setup

Consideration is given for selectivity during mine design and planning when selecting an appropriate block size. Selective Mining Units (SMUs) reflect the geological knowledge of the deposit and balancing equipment efficiency and anticipated ore loss and dilution. Block sizes used on each deposit and domain are based on the data density, directly linked to the drill campaign (GC, AGC or exploration). Block sizes are typically one half to one third the drilling spacing.

Wireframes are built to define the three drill campaign areas at Kibali, which are grade control, advance grade control and exploration/resource drilling, listed in order of decreasing drillhole density. The drill campaign wireframes control the maximum size of the blocks that are built in a specified block model area, allowing the estimation to be carried out on a parent block size appropriate to each drill campaign, within a single block model. Sub-blocking was used to define the geological and domain contacts to an acceptable level of accuracy within the block model, allowing a higher resolution when the model is interpolated.

The search strategy used was based on the variogram results obtained through considering the data distribution for each of the domains. The search ellipsoids were orientated optimally for each domain, considering the plunge and dip of the wireframe.

Each pass is completed using a varying degree of restrictions before any given block can be estimated. In total, four passes were used on every block model, each with increasing search radius representing the decreasing confidence in the blocks for each subsequent run. In rare situations, a fifth pass was considered to fill a small number of edge blocks with grades, typically in conceptual/exploration target zones.

Dolerite dykes were wireframed and coded into the block with the relevant grade field set to zero as default.

All block models use a standardised attribute field setup to ensure consistency of nomenclature and data capture across all deposits within Kibali.

For all deposits, gold grades are estimated using OK. For Sessenge, Pakaka, and Aerodrome arsenic grades are also estimated using OK.

Dynamic Anisotropy

Many of the models since 2017 were estimated using the dynamic anisotropy (DA) functionality within Vulcan software. At Kibali, DA surfaces are modelled for each domain. These are usually simple surfaces that trend through the middle of the 3D mineralisation wireframes, orientation data from which is written to the block model and used to orientate the search neighbourhood.

Figure 14-45 shows and example of a DA surface from KCD domain 5002.

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Notes:

Dynamic Anisotropy surface (Red), and 5002 Mineralised domain (Brown)

Looking Northwest

Figure 14-45 3D View of KCD Domain 5002 and Guiding Dynamic Anisotropy Surface

KCD

The KCD block model has a parent block size of 10 m by 20 m by 10 m with a minimum sub cell size of 2.5 m by 2.5 m by 2.5 m. This considers that most of the higher grade open pit drill holes were on a 10 m by 5 m grid spacing. Underground drilling was drilled on an approximate 15 m by 20 m spacing. The block model was flagged by each mineralisation domain separately by priority. Table 14-56 summarises the KCD block model extents.

Table 14-56 KCD Global Block Model Extent (With Rotation)



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	784,250	343,750	4,500

Minimum Offset	0	0	0

Maximum Offset	2,120	4,400	1,550

Parent Block Size (m)	10	20	10

Sub Cell Size (m)	2.5	2.5	2.5

Rotation (°)	135	0	0

In total, KCD contains 103 estimation domains. The KCD model was constructed using DA during estimation to capture any slight change in the orientation of the mineralisation.

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Sessenge and Sessenge SW

The Sessenge and Sessenge SW block model has a parent block size of 10 m by 20 m by 5 m with a minimum sub cell size of 1.25 m by 1.25 m by 1.25 m. The block model was flagged by each mineralisation domain separately by priority. Table 14-57 summarises the Sessenge block model extents.

Table 14-57 Sessenge Global Block Model Extent (With Rotation)



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	784,700	343,460	5,500

Minimum Offset	0	0	0

Maximum Offset	1,250	1,740	500

Parent Block Size (m)	10	20	5

Sub Cell Size (m)	1.25	1.25	1.25

Rotation (°)	135	0	0

The Sessenge block model was constructed using DA during estimation.

Gorumbwa

The Gorumbwa block model has a parent block size of 10 m by 20 m by 10 m with a minimum sub cell size of 2.5 m by 2.5 m by 2.5 m. The block model was flagged by each mineralisation domain separately by priority. Table 14-58 summarises the Gorumbwa block model extents.

Table 14-58 Gorumbwa Global Block Model Extent (With Rotation)



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	784,919	344,425	4,625

Minimum Offset	0	0	0

Maximum Offset	1,030	2,100	1550

Parent Block Size (m)	10	20	10

Sub Cell Size (m)	2.5	2.5	2.5

Rotation (°)	135	0	0

The Gorumbwa block model was constructed using DA during estimation.

Pakaka

The block Pakaka model has a parent block size of 10 m by 20 m by 10 m with a minimum sub cell size of 1.25 m by 1.25 m by 1.25 m. The block model was flagged by each mineralisation domain separately by priority. Table 14-59 summarises the Pakaka block model extents.

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Table 14-59 Pakaka Global Block Model Extent



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	787,880	347,800	5,400

Minimum Offset	0	0	0

Maximum Offset	1,400	1,500	700

Parent Block Size (m)	10	20	10

Sub Cell Size (m)	1.25	1.25	1.25

Rotation (°)	90	0	0

The Pakaka block model estimation did not make use of DA. The search ellipsoid was orientated individually for each domain (Table 14-60).

Table 14-60 Pakaka Search Ellipsoid Orientation



Domain	Domain Orientation (°)
Domain	Bearing	Plunge	Dip

1001	30	-15	-13

1007	54	-13	38

1101	37	-18	-16

1102	125	-5	-1

1103	44	-8	-18

1105	44	-14	-6

1106	41	-14	5

Kombokolo

The Kombokolo block model has a parent block size of 10 m by 15 m by 5 m with a minimum sub cell size of 1.25 m by 1.25 m by 0.625 m. The block model was flagged by each mineralisation domain separately by priority. Table 14-61 summarises the Kombokolo block model extents.

Table 14-61 Kombokolo Global Block Model Extent (No Rotation)



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	786,000	345,000	5,500

Minimum Offset	0	0	0

Maximum Offset	1,100	1,125	900

Parent Block Size (m)	10	15	5

Sub Cell Size (m)	1.25	1.25	0.625

Rotation (°)	90	0	0

The Kombokolo block model estimation did not use DA. The search ellipsoid was orientated individually for each domain (Table 14-62).

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Table 14-62 Kombokolo Search Ellipsoid Orientation



Domain	Domain Orientation (°)
Domain	Bearing	Plunge	Dip

1101	55	-25	25

1102	55	-25	25

1001	45	-30	25

1002	55	-30	25

1003	45	-30	25

1004	45	-30	25

1005	45	-20	25

1006	60	-25	20

1007	60	-25	20

Pamao and Pamao South

The Pamao and Pamao South block model has a parent block size of 20 m by 20 m by 5 m with a minimum sub cell size of 2.5 m by 2.5 m by 2.5 m. The block model was flagged by each mineralisation domain separately by priority listed. Table 14-63 summarises the Pamao block model extents.

Table 14-63 Pamao and Pamao South Global Block Model Extent (With Rotation)



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	786,150	349,100	5,400

Minimum Offset	0	0	0

Maximum Offset	2,200	2,000	600

Parent Block Size (m)	20	20	5

Sub Cell Size (m)	2.5	2.5	2.5

Rotation (°)	145	0	0

The Pamao and Pamao South block model was constructed using DA during estimation.

Mengu Village

The Mengu Village block model has a parent block size of 20 m by 10 m by 5 m with a minimum sub cell size of 2.5 m by 2.5 m by 2.5 m. The block model was flagged by each mineralisation domain separately by priority.

Table 14-64 summarises the Mengu Village block model extents.

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Table 14-64 Mengu Village Global Block Model Extent (With Rotation)



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	783,112	351,059	5,500

Minimum Offset	0	0	0

Maximum Offset	1,360	1,000	500

Parent Block Size (m)	20	10	5

Sub Cell Size (m)	2.5	2.5	2.5

Rotation (°)	130	0	0

The Mengu Village block model was constructed using DA during estimation.

Megi-Marakeke-Sayi

The Megi-Marakeke-Sayi block model has a parent block size of 20 m by 10 m by 5 m with a minimum sub cell size of 1.25 m by 1.25 m by 1.25 m. The block model was flagged by each mineralisation domain separately by priority. Table 14-65 summarises the Mengu block model extents.

Table 14-65 Megi-Marakeke-Sayi Global Block Model Extent (With Rotation)



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	784,085	350,100	5,500

Minimum Offset	0	0	0

Maximum Offset	2,300	1,200	500

Parent Block Size (m)	20	10	5

Sub Cell Size (m)	1.25	1.25	1.25

Rotation (°)	130	0	0

The Megi-Marakeke-Sayi block model was constructed using DA during estimation.

Kalimva-Ikamva

The Kalimva-Ikamva block model has a parent block size of 10 m by 20 m by 5 m with a minimum sub cell size of 2.50 m by 2.50 m by 2.50 m. The block model was flagged by each mineralisation domain separately by priority. Table 14-66 summarises the Kalimva-Ikamva block model extents.

Table 14-66 Kalimva-Ikamva Block Model Extent (No Rotation)



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	782,189	358,048	5,500

Minimum Offset	0	0	0

Maximum Offset	2,000	3240	600

Parent Block Size (m)	10	20	5

Sub Cell Size (m)	2.50	2.50	2.50

Rotation (°)	90	0	0

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The Kalimva-Ikamva block model was constructed using DA during estimation.

Mengu Hill

The Mengu Hill block model has a parent block size of 10 m by 15 m by 5 m with a minimum sub cell size of 2.5 m by 2.5 m by 1.25 m. The block model was flagged by each mineralisation domain separately by priority. Table 14-67 summarises the Mengu Hill block model extents.

Table 14-67 Mengu Hill Global Block Model Extent (No Rotation)



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	782,510	350,610	5,300

Minimum Offset	0	0	0

Maximum Offset	1,100	1,290	700

Parent Block Size (m)	10	15	5

Sub Cell Size (m)	2.5	2.5	1.25

Rotation (°)	90	0	0

The Mengu Hill estimation did not make use of DA, and the search ellipsoid was orientated individually for each estimation domain as shown in (Table 14-68).

Table 14-68 Mengu Hill Search Ellipsoid Orientation



Domain	Domain Orientation (°)
Domain	Bearing	Plunge	Dip

1001	35	-20	-25

1002	35	-20	-25

1101	35	-20	-25

1102	35	-20	-25

1103	35	-20	-25

Aerodrome

The Aerodrome block model has a parent block size of 10 m by 20 m by 10 m with a minimum sub cell size of 2.5 m by 2.5 m by 2.5 m. The block model was flagged by each mineralisation domain separately by priority. Table 14-69 summarises the Aerodrome model extents.

Table 14-69 Aerodrome Global Block Model Extent



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	788,000	347,800	5,600

Minimum Offset	0	0	0

Maximum Offset	800	600	400

Parent Block Size (m)	10	20	10

Sub Cell Size (m)	2.5	2.5	2.5

Rotation (°)	160	0	0

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The Aerodrome block model was constructed using DA during estimation.

Oere

The Oere block model has a parent block size of 10 m by 20 m by 5 m with a minimum sub cell size of 2.5 m by 2.5 m by 2.5 m. The block model was flagged by each mineralisation domain separately by priority. Table 14-70 summarises the Oere block model extents.

Table 14-70 Oere Global Block Model Extent (With Rotation)



Block Extents	Easting (X)	Northing (Y)	Elevation (Z)

Origin	781,390	354,360	5,275

Minimum Offset	0	0	0

Maximum Offset	1,300	3,600	650

Parent Block Size (m)	10	20	5

Sub Cell Size (m)	2.5	2.5	2.5

Rotation (°)	110	0	0

The Oere block model was constructed using DA during estimation.

14.10

Resource Classification

Current Resources

Under the CIM definitions (CIM Standards on Mineral Resources and Reserves Definitions and Guidelines, 2014), a “Measured Mineral Resource is that part of a Mineral Resource for which quantity, grade, density, shape, and physical characteristics need to be established with sufficient confidence sufficient to allow the application of modifying factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit”.

An Indicated Mineral Resource is “that part of a Mineral Resource for which “quantity, grade, density, shape, and physical characteristics need to be established with sufficient confidence sufficient to allow the appropriate application of modifying factors in sufficient detail to support mine planning and evaluation of the economic viability of the deposit”. An Indicated Mineral Resource has a lower level of confidence than a Measured Mineral Resource.

An Inferred Mineral Resource is “that part of a Mineral Resource for which quantity and grade or quality are estimated on the basis of limited geological evidence and sampling. Geological evidence is sufficient to imply but not verify geological and grade continuity or quality continuity’. An Inferred Mineral Resource has a lower level of confidence than an Indicated and Inferred Mineral Resource and must not be converted to a Mineral Reserve.

Resource Classification was based on geological continuity and drill data density, variogram range continuity and stability, as well as estimation quality in form of slope of regression (SR) and kriging efficiency (KE). This was carried out by displaying the estimated blocks (SR and KE), together with the supporting data as a guide.

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The general Mineral Resource classification parameters are presented in Table 14-71.

Table 14-71 Kibali Mineral Resource Classification Parameters



Statistic	Deposit	Measured	Indicated	Inferred

Minimum Samples		8	6	4

Minimum Consecutive Sections		4	Good Geological Continuity	-

Max Drilling Density	KCD OP	10 m by 5 m or 20 m by 5 m	40 m by 30 m	80 m by 80 m
	KCD UG	25 m by 10 m	40 m by 40 m	80 m by 80 m
	Gorumbwa	10 m by 5 m or 15 m by 10 m	20 by 10 or 30 by 30	80 m by 80 m
	Pakaka	20 m by 10 m or 20 m by 5 m	40 m by 40 m	80 m by 60 m
	Sessenge	10 m by 10 m	40 m by 40 m	80 m by 80 m
	Pamao	10 m by 10 m	20 by 20	80 m by 80 m
	Pamao South	NA	20 by 20	40 m by 40 m
	Kombokolo	10 m by 5 m or 10 by 10 m	30 m by 30 m	80 m by 80 m
	Mengu Village	-	-	40 m by 40 m to 80 m by 40 m
	Megi-Marakeke-Sayi	-	30 m by 30 m	80 m by 80 m
	Kalimva-Ikamva	10 m by 5 m	20 m by 20 m	40 m by 20 m
	Aerodrome	10 m by 10 m	20 m by 20 m	40 m by 40 m
	Mengu Hill	10 m by 5 m	30 m by 20 m	80 m by 60 m
	Oere	NA	20 m by 20 m	40 m by 40 m

For Indicated Mineral Resources, there are some allowances for areas where drilling density is lower but successive drilling campaigns have shown there is grade and geological continuity.

14.11

Block Model Depletion

Active mining areas are scanned using cavity monitoring laser scanners on a monthly basis and detailed drone photometry surface scans are completed on a weekly basis.

Every block model was flagged with the regional 2 m DTM with any blocks falling above the surface being flagged as air.

The following deposits have not been mined and no depletion was applied:

●

Megi-Marakeke-Sayi

●

Kalimva-Ikamva

●

Mengu Village

●

Oere

●

Sessenge SW

KCD

As an active mine, KCD requires depletion to represent the blocks mined until the end of the reporting period. Depletion pit surveys at KCD were updated in December 2021 and used to flag the block model. The KCD underground resource block model was likewise depleted using the

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final EOY 2021 CMS stope scans. Accuracy was maintained despite differing block sizes as a partial block percentage is stored.

Gorumbwa Underground

The high-grade central lode (1004) of the Gorumbwa deposit was mined at surface and underground between the 1950s and 1990s. The original plans were digitised from paper drawings and can be seen in Figure 14-46. However, actual results can differ from plan, as demonstrated by probe drilling used to identify the voids and build an interpretation. This difference is local and globally not material. During 2018, 3DMSI carried out a 3D laser scan of the void to increase confidence. The open pit surface and overall development and stope void wireframes were used to deplete the model and will continuously be updated (Figure 14-46).

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Figure 14-46 3D OP and Development Stope Voids at Gorumbwa

Gorumbwa, Aerodrome, and Sessenge

Gorumbwa, Aerodrome, and Sessenge are all currently in operation and were depleted with pit surfaces as of 31 December 2021.

Pakaka, Kombokolo, and Mengu Hill

Pakaka, Kombokolo, and Mengu Hill are not in operation and were depleted with pit surfaces as of 28 February 2018, and 31 December 2018 and 30 June 2017, respectively.

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14.12

Block Model Validation

Before, during and after the block models were classified, validation checks were undertaken on the block model volumes and estimated grades to ensure that no major errors occurred during the mode build or estimation process, as well as testing the precision, accuracy, and assess any bias in the estimated grades.

The block models were validated using the following steps:

Volume Reconciliation between the block model estimation domains and related wireframes. Table 14-72 summarises the variances between the wireframe and block model volumes across all deposits.

A check of the number of the blocks estimated with negative grades due to excessive negative kriging weights have been reset to the anisotropic nearest block grade of the closest sample.

A comparison between the data minimum, maximum, mean, declustered mean and the estimation mean for each of the domains (within the open pit or underground drill campaigns is created). This is completed to check for possible over or under estimation.

Swath plots are created for each geological domain to validate the estimated grade variability compared to the composite along strike, across strike and Z axis. This is to check that the model estimate follows the trends seen in the data and that there is no general bias with over or under estimation. Areas with less data support are also highlighted for further drilling and geological work. The swath plots for Kibali show the confidence for the deposit is within acceptable limits and that conditional bias is kept to a minimum. An example for KCD lode 3000, domain 3106 is shown in Figure 14-47 to Figure 14-49.

Visual check comparing the composite data to the block estimates to check for an acceptable correlation. An example of the visual checks is shown in Figure 14-50.

Change of support (COS) histogram plots which, compare the distribution of the block estimate with the distribution of the change of support local block estimate. These COS graphs demonstrate how the variance is reduced from the composited data to the change of support value of each composite (Figure 14-51). In addition, decluster plots are generated to compare the ordinary kriged block estimate against the local change of support block estimate (Figure 14-52).

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Table 14-72 2021 Block Model Volume Comparison



Deposit	Wireframe Volume (m³)	Block Model Volume (m³)	Variance (%)

KCD UG	101,607,249	101,527,016	0%

KCD OP	23,599,755	23,593,672	0%

Sessenge	5,992,984	5,921,703	0%

Gorumbwa	11,025,860	11,031,844	0%

Pakaka	12,365,957	12,366,826	0%

Kombokolo	2,984,241	2,984,110	0%

Pamao	13,697,374	13,688,344	0%

Pamao South	1,729,009	1,728,547	0%

Mengu Village	4,123,637	4,119,742	0%

Megi-Marakeke-Sayi	10,265,978	10,266,791	0%

Kalimva Ikamva	14,551,715	14,553,578	0%

Mengu Hill	4,123,637	4,119,742	0%

Aerodrome	1,059,927.82	1,059,593.75	0%

Oere	29,364,412.50	29,350,812.50	0%

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Figure 14-47 KCD SWATH Plot of Domains 3106 Along Strike (45°)

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Figure 14-48 KCD SWATH Plot of Domains 3106 Across Strike (135°)

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Figure 14-49 KCD SWATH Plot of Domains 3106 Along Z Axis (RL)

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Figure 14-50 An Example of the KCD Visual Checks on Section for Lode 5000 (Domain 5102)

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Figure 14-51 COS Plot for Lode 3000 (Domain 3106)

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Figure 14-52 Decluster Plot for Lode 3000 (Domain 3106)

14.13

Resource Cut-Off Grades

The assumptions used to generate cut-off grades for Mineral Resource estimation are based on operational data. A gold price of $1,500 is used in line with Barrick corporate guidelines, which considers long-term gold price forecasts.

KCD

KCD Open Pit Resources

The cut-off grade calculations for the KCD open pit Mineral Resources are summarised in Table 14-73.

Table 14-73 KCD 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	2.92	2.97	3.09

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	1.27	1.27	1.27

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	0.00	0.00	0.00

Process Cost	$/t milled	15.04	15.04	17.85

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Material Type	Unit	Oxide	Trans	Fresh

Processing Recovery	%	90.1	90.1	86.1

General and Administration (G&A)	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost	$/t milled	15.04	15.04	17.85

Total Mining Cost	$/t ore mined	24.69	25.08	26.04

Marginal In-situ Cut-off Grade	g/t Au	0.60	0.60	0.70

Strip Ratio	7.0

The average tonnage weighted cut-off grade for the KCD Open Pit is 0.69 g/t Au.

KCD Underground Resources

The cut-off grade calculations for KCD underground Mineral Resources are summarised in Table 14-74.

Table 14-74 KCD Underground 2021 Optimisation Parameters



Material Type	Unit	Fresh

Mine Production	$/t mined	36.17

Capital	$/t mined	3.97

Backfill	$/t mined	0.00



Process Cost	$/t milled	17.85

Processing Recovery	%	90%



G&A	$/t milled	8.47



Gold Royalties (4.7%)	$/oz Au	70.50

Gold Price (Resource)	$/oz Au	1,500



Total Unit Cash Cost	$/t milled	66.46

Mining Cut-Off Grade	g/t Au	1.62

KCD Underground Optimised Minable Stope Shapes

For the current KCD Mineral Resource MSO shapes were used to differentiate blocks that demonstrate reasonable prospects of eventual economic extraction. This reporting method of using stopes, not blocks, excludes high-grade blocks that are geometrically isolated and can in fact include blocks at lower grades, but that are geometrically contiguous.

For KCD, 3D exclusion solid shapes were manually constructed post MSO computation, to ensure no accumulation of unrecoverable mineralised blocks in the current Mineral Resource (Figure 14-53).

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Figure 14-53 3D Oblique Image of KCD Underground Development with the Annual Resource Exclusion

Solids

A marginal mined cut-off grade of 1.62 g/t Au at $1,500/oz Au defines the KCD UG optimised mineable stope shapes, within an underground reporting limit wireframe solid, with varying upper elevation (RL). This varying RL now limits the 5000 lode to 5680 mRL, and 3000 lode to 5682.5 mRL for the current Mineral Resource estimate. This varying RL was put in place to ensure that all material that forms part of the UG Mineral Resource is excluded from the OP Mineral Resource.

The MSO is executed with parameters that are less restrictive than those used for Mineral Reserve calculation. Stope orientation changes and stope sizes are more flexible, as well as a proportion of waste included. All stope orientations are set to follow wireframe surfaces modelled on deposit structure.

Visual checks were undertaken on blocks that were not included in the MSO shapes primarily due to geology and the shapes of mineralised lodes. These blocks would have been included in the Mineral Resource estimation if a cut-off grade only approach had been used.

Figure 14-54 shows the MSO generated stope for KCD UG and Figure 14-55 shows the above cut-off grade blocks not included in an MSO generated stope due to low reasonable prospects for eventual economic extraction and not reaching the required cut-off grade if a mineable shape is considered.

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Sessenge and Sessenge SW

The cut-off grade calculations for the Sessenge open pit Mineral Resource are summarised in Table 14-75.

Table 14-75 Sessenge 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	2.62	2.68	2.80

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	1.24	1.24	1.24

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	0.00	0.00	0.00

Process Cost	$/t milled	15.04	15.04	17.85

Processing Recovery	%	90.3	75.9	81.0

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	15.04	15.04	17.85

Total Mining Cost (per ore tonne mined)	$/t ore	9.79	9.98	10.37

Marginal In-situ Cut-off Grade	g/t Au	0.60	0.71	0.74

Strip Ratio	2.3

The average tonnage weighted cut-off grade for Sessenge and Sessenge SW is 0.76 g/t Au

Gorumbwa

The cut-off grade calculations for Gorumbwa open pit Mineral Resource are summarised in Table 14-76.

Table 14-76 Gorumbwa 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	3.29	3.14	3.24

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	1.28	1.28	1.28

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	0.00	0.00	0.00

Process Cost	$/t milled	15.04	15.04	17.85

Processing Recovery	%	90.0	90.0	90.0

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	15.04	15.04	17.85

Total Mining Cost (per ore tonne mined)	$/t ore	37.43	35.77	36.88

Marginal In-situ Cut-off Grade	g/t Au	0.60	0.60	0.67

Strip Ratio	10.0

The average tonnage weighted cut-off grade for Gorumbwa is 0.67 g/t Au.

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Pakaka

The cut-off grade calculations for the Pakaka open pit Mineral Resource summarised in Table 14-77.

Table 14-77 Pakaka 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	2.72	2.80	2.88

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	1.38	1.38	1.38

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	1.05	1.05	1.05

Process Cost	$/t milled	15.04	15.04	17.85

Processing Recovery	%	89.0	89.0	80.2

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	16.09	16.09	18.90

Total Mining Cost (per ore tonne mined)	$/t ore	32.44	33.37	34.33

Marginal In-situ Cut-off Grade	g/t Au	0.63	0.63	0.78

Strip Ratio	10.4

The average tonnage weighted cut-off grade for Pakaka is 0.78 g/t Au.

Pakaka haulage costs are incorporated into the mining costs as both the haulage and mining are operated by the same contractor.

Geometallurgical work initiated in early 2016 has primarily focused on Pakaka, where feasibility test work had identified mainly two domains of high and low-grade arsenic domains. With limited metallurgical test work data available, it was demonstrated that there was:

●

Direct correlation between gold grade and arsenic content.

●

Inverse correlation between recovery and arsenic grade.

Consequently, the recoveries shown in Table 14-77 are an average for each weathering classification. The detailed Pakaka domain recoveries are detailed in Table 14-78 and the geometallurgical domains are shown in Figure 14-56.

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Source: Kibali Goldmines, 2021

Figure 14-56 Plan View Map of the Pakaka Geometallurgical Domains and Their Spatial Correlation with the

Mineralisation Resource Domains

In addition to applying these recoveries to the LOM optimisation, the delineation of the six geometallurgical domains is used to optimise the blending strategy during feeding in the plant (Table 14-78). Apart from understanding the recoveries associated with the individual domains, arsenic concentration in the plant feed blend is used to maintain the thresholds (<2,000 ppm) that ensures, not only stable recovery, but reagents consumption.

A small amount of silver is also within the ore and doré bars for which a penalty is applied by the smelters. However, this is not material and silver grades are not required to be estimated in the model.

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Table 14-78 Pakaka Geometallurgical Domained Recoveries



Domain	Description	Weathering	BRTs Average Dissolution (%)	Arsenic Assay (ppm)	Feasibility Direct Leach (%)	Comments

1	Low Au grade / Low As / Low Recovery	Saprolite	84.1	<1,000	-	-
		Oxidised Transition	86.8	<1,000	-	-
		Reduced Transition	81.6	<1,000	-	-

2	High Au grade / High As / High Recovery	Saprolite	90.8	>2,000	-	-
		Oxidised Transition	90.4	>2,000	-	-
		Reduced Transition	86	>2,000	-	-

3	High Au grade / High As / Low Recovery	Fresh	75.2	>2,000	59.6	Feasibility dissolution exclude gravity, so use BRT value to cater for gravity

4	High Au grade / High As / High Recovery	Saprolite	85.5	<1,000	-	-
		Reduced Transition	92.6	<1,000	-	-
		Fresh	93.4	<1,000	87.3	Use feasibility number and the BRT number for plant performance tracking

5	Medium Au grade / Medium As / Medium Recovery	Saprolite	87.4	1,000 – 2,000	-	-
5	Medium Au grade / Medium As / Medium Recovery	Fresh	88.3	1,000 – 2,000	87.3	Feasibility split only caters for above and below 0.2% arsenic content . Sample represents below 0.2%

6	High Au grade / High As / High Recovery	Saprolite	89	>2,000	-	-
		Oxidised Transition	89.6	>2,000	-	-
		Fresh	88.8	>2,000	-	-

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Kombokolo

The cut-off grade calculations for the Kombokolo open pit Mineral Resource are summarised in Table 14-79.

Table 14-79 Kombokolo 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	2.65	2.72	2.84

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	1.19	1.19	1.19

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	0.00	0.00	0.00

Process Cost	$/t milled	15.04	15.04	15.04

Processing Recovery	%	85.0	85.0	85.0

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	14.34	14.34	14.34

Total Mining Cost (per ore tonne mined)	$/t ore	28.75	29.53	30.68

Marginal In-situ Cut-off Grade	g/t Au	0.63	0.63	0.70

Strip Ratio	9.40

The average tonnage weighted cut-off grade for Kombokolo is 0.78 g/t Au.

Pamao

The cut-off grade calculations for the Pamao open pit Mineral Resource are summarised in Table 14-80.

Table 14-80 Pamao 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	2.85	2.88	2.95

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	1.31	1.31	1.31

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	1.05	1.05	1.05

Process Cost	$/t milled	15.04	15.04	17.85

Processing Recovery	%	90.9	85.0	85.0

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	16.09	16.09	18.90

Total Mining Cost (per ore tonne mined)	$/t ore	18.59	18.80	19.19

Marginal In-situ Cut-off Grade	g/t Au	0.62	0.66	0.73

Strip Ratio	5.1

The average tonnage weighted cut-off grade for Pamao is 0.71 g/t Au.

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Pamao South

The cut-off grade calculations for the Pamao South open pit Mineral Resource are summarised in Table 14-81.

Table 14-81 Pamao South 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	2.85	2.88	2.95

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	1.31	1.31	1.31

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	1.05	1.05	1.05

Process Cost	$/t milled	15.04	15.04	17.85

Processing Recovery	%	89.0	88.0	86.5

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	16.09	16.09	18.90

Total Mining Cost (per ore tonne mined)	$/t ore	14.97	15.14	15.45

Marginal In-situ Cut-off Grade	g/t Au	0.63	0.64	0.72

Strip Ratio	3.8

The average tonnage weighted cut-off grade for Pamao South is 0.71 g/t Au.

Mengu Village

The cut-off grade calculations for the Mengu Village open pit Mineral Resource are summarised in Table 14-82.

Specific studies for haulage and processing recovery have not yet been undertaken for Mengu Village. As such values for these parameters are taken from the nearest deposit, Meg-Marakeke-Sayi, which has similar mineralisation, host lithology, and distance to ROM pad.

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Table 14-82 Mengu Village 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	2.46	2.81	2.95

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	4.25	4.25	4.25

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost (after Megi-Marakeke-Sayi)	$/t mined	2.24	2.24	2.24

Process Cost	$/t milled	15.04	15.04	17.85

Processing Recovery (after Megi-Marakeke-Sayi)	%	90.0	90.0	90.0

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	15.04	15.04	17.85

Total Mining Cost (per ore tonne mined)	$/t ore	19.97	22.26	23.16

Marginal In-situ Cut-off Grade	g/t Au	0.73	0.73	0.80

Strip Ratio	5.40

The average tonnage weighted cut-off grade for Mengu Village is 0.74 g/t Au.

Megi-Marakeke-Sayi

The cut-off grade calculations for the Megi-Marakeke-Sayi open pit Mineral Resource are summarised in Table 14-83.

Table 14-83 Megi-Marakeke-Sayi 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	3.18	3.23	3.28

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	1.25	1.25	1.25

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	2.24	2.24	2.24

Process Cost	$/t milled	15.04	15.04	17.85

Processing Recovery	%	90.0	90.0	89.5

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	17.28	17.28	20.09

Total Mining Cost (per ore tonne mined)	$/t ore	15.24	15.44	15.70

Marginal In-situ Cut-off Grade	g/t Au	0.65	0.65	0.72

Strip Ratio	3.4

The average tonnage weighted cut-off grade for Megi-Marakeke-Sayi is 0.74 g/t Au.

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Kalimva-Ikamva

The cut-off grade calculations for the Kalimva-Ikamva open pit Mineral Resource are summarised in Table 14-84.

Table 14-84 Kalimva-Ikamva 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	2.35	2.59	2.92

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	6.25	6.25	6.25

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	5.00	5.00	5.00

Process Cost	$/t milled	15.04	15.04	17.85

Processing Recovery	%	90.0	89.0	89.0

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	20.04	20.04	22.85

Total Mining Cost (per ore tonne mined)	$/t ore	31.72	34.36	37.92

Marginal In-situ Cut-off Grade	g/t Au	0.84	0.85	0.92

Strip Ratio	9.9

The average tonnage weighted cut-off grade for Kalimva-Ikamva is 0.94 g/t Au.

Mengu Hill

The cut-off grade calculations for the Mengu Hill open pit Mineral Resource are summarised in Table 14-85.

Table 14-85 Mengu Hill 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	2.98	3.16	3.22

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	4.25	4.25	4.25

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	2.50	2.50	2.50

Process Cost	$/t milled	15.04	15.04	17.85

Processing Recovery	%	81.0	77.0	70.0

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	17.54	17.54	20.35

Total Mining Cost (per ore tonne mined)	$/t ore	16.92	19.75	20.01

Marginal In-situ Cut-off Grade	g/t Au	0.81	0.86	1.03

Strip Ratio	3.9

The average tonnage weighted cut-off grade for Mengu Hill is 0.99 g/t Au.

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Aerodrome

The cut-off grade calculations for the Aerodrome open pit Mineral Resource are summarised in Table 14-86.

Table 14-86 Aerodrome 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	1.70	2.01	2.33

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	1.25	1.25	1.25

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	1.05	1.05	1.05

Process Cost	$/t milled	15.04	15.04	17.85

Processing Recovery	%	90.0	88.0	85.9

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	16.09	16.09	18.90

Total Mining Cost (per ore tonne mined)	$/t ore	10.75	12.45	14.24

Marginal In-situ Cut-off Grade	g/t Au	0.62	0.64	0.72

Strip Ratio	4.6

The average tonnage weighted cut-off grade for the Aerodrome is 0.69 g/t Au.

Oere

The cut-off grade calculations for the Oere open pit Mineral Resource are summarised in Table 14-87.

Table 14-87 Oere 2021 Optimisation Parameters



Material Type	Unit	Oxide	Trans	Fresh

Waste Cost	$/t mined	2.69	2.97	3.02

Extra Ore Cost – GC + Ore – Rehandle + Overhaul	$/t mined	6.25	6.25	6.25

GC Only	$/t mined	0.75	0.75	0.75

Dilution	%	10%	10%	10%

Ore Loss	%	3%	3%	3%

Haulage Cost	$/t mined	4.50	4.50	4.50

Process Cost	$/t milled	15.04	15.04	17.85

Processing Recovery	%	88.0	86.5	87.0

G&A	$/t milled	8.47	8.47	8.47

Gold Price (Resource)	$/oz Au	1,500	1,500	1,500

Gold Royalty (4.7%)	$/oz Au	70.50	70.50	70.50

Total Process Cost (per ore tonne mined)	$/t ore	19.54	19.54	22.35

Total Mining Cost (per ore tonne mined)	$/t ore	37.61	40.87	41.54

Marginal In-situ Cut-off Grade	g/t Au	0.85	0.86	0.93

Strip Ratio	10.7

The average tonnage weighted cut-off grade for the KCD Open Pit is 0.93 g/t Au.

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14.14 Mineral Resource Statement

The Mineral Resource estimates have been prepared according to the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) 2014 Definition Standards for Mineral Resources and Mineral Reserves dated 10 May 2014 (CIM (2014) Standards) as incorporated with National Instrument 43-101 Standards of Disclosure for Mineral Projects (NI 43-101). Mineral Resource estimates were also prepared using the guidance outlined in CIM Estimation of Mineral Resource and Mineral Reserve Best Practice Guidelines 2019 (CIM (2019) MRMR Best Practice Guidelines).

Underground Mineral Resources were reported using MSO, effectively within a minimum mineable stope shape, applying reasonable mineability constraints, including a 4.5 m minimum mining width, a reasonable distance from current or planned development, and a measure of assumed profitability at the related resource cut-off grade, thus deemed as having reasonable prospects for eventual economic extraction.

Stockpiles are comprised of mineralised material stored at the surface ROM pad, originating from both OP and UG production. Each stockpile is filled with similar material types, with an established grade range and oxidation state, tracked as part of normal mining operations and metal accounting. The stockpiles are measured by weekly drone survey. Grade and tonnage of OP stocks are estimated according to source dig blocks and number of truck counts, using a weighbridge to adjust for fluctuations in both density and truck fill factor. Grade and tonnage of UG stocks are estimated according to shaft skip weights and ore pass truck counts and their source blasts from stopes, adjusting for the presence of paste dilution.

The Kibali Measured and Indicated Mineral Resources, as of 31 December 2021 (Table 14-88), are estimated at 140 Mt at 3.41 g/t Au containing 15 Moz of gold, with an additional Inferred Resource of 23 Mt at 2.7 g/t Au containing 2.0 Moz of gold (100% basis). This represents a 4% drop in grade and a 6% increase in tonnes, resulting in an overall 3% increase in contained gold ounces, relative to the 2020 Mineral Resource estimate.

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Table 14-88 Kibali Mineral Resources as of 31December 2021



Deposit	Cut -Off Grade (g/t Au)	Measured			Indicated			Measured + Indicated			Inferred
		Tonnes (Mt)	Grade (g/t Au)	Contained (Moz Au)	Tonnes (Mt)	Grade (g/t Au)	Contained (Moz Au)	Tonnes (Mt)	Grade (g/t Au)	Contained (Moz Au)	Tonnes (Mt)	Grade (g/t Au)	Contained (Moz Au)

Open Pit

Stockpiles	0.54	0.32	3.17	0.032	-	-	-	0.32	3.17	0.032	-	-	-

KCD	0.69	2.5	3.13	0.25	3.9	2.27	0.28	6.4	2.61	0.53	0.79	1.9	0.048

Sessenge	0.76	0.68	2.36	0.051	0.80	1.72	0.044	1.5	2.01	0.096	0.12	1.6	0.0060

Sessenge SW	0.76	-	-	-	-	-	-				0.47	1.7	0.026

Pakaka	0.78	3.2	2.70	0.28	5.5	2.54	0.45	8.7	2.60	0.73	0.76	3.4	0.083

Mengu Hill	0.99	-	-	-	0.80	2.56	0.066	0.80	2.56	0.066	0.48	3.5	0.054

Gorumbwa	0.67	1.3	2.38	0.10	5.0	3.21	0.52	6.3	3.04	0.62	1.5	2.5	0.12

Megi-Marakeke-Sayi	0.74	-	-	-	11	1.73	0.62	11	1.73	0.62	0.84	1.9	0.050

Pamao and Pamao South	0.71	6.6	1.65	0.35	6.6	1.76	0.37	13	1.71	0.73	0.00091	0.8	0.000024

Kombokolo	0.78	0.24	2.01	0.015	0.31	3.00	0.030	0.55	2.57	0.046	0.13	1.8	0.0072

Kalimva Ikamva	0.94	0.52	2.57	0.043	7.8	2.61	0.66	8.3	2.61	0.70	0.016	4.4	0.0023

Aerodrome	0.69	0.38	1.56	0.019	0.020	1.38	0.00072	0.39	1.55	0.020	0.0075	1.6	0.00039

Oere	0.93	-	-	-	3.1	2.15	0.21	3.1	2.15	0.21	2.0	1.7	0.11

Mengu Village	0.74	-	-	-	-	-	-	-	-	-	1.1	1.4	0.048

OP Total		16	2.26	1.1	45	2.25	3.3	60	2.25	4.4	8.2	2.1	0.55

Underground

KCD UG	1.62	32	4.63	4.7	48	4.06	6.3	80	4.29	11	15	3.0	1.4

Open Pit + Underground

Total Resources		48	3.84	5.9	93	3.18	9.5	140	3.41	15	23	2.7	2.0

Notes:

The Mineral Resource estimate has been prepared according to CIM (2014) Standards and using CIM (2019) MRMR Best Practice Guidelines.

All Mineral Resources tabulations are reported inclusive of that material which is then modified to form Mineral Reserves.

Open pit Mineral Resources are Mineral Resources within the $1,500/oz Au pit shell reported at a weighted average cut-off grade of 0.77 g/t Au.

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14.15

2021 Versus 2020 EOY Mineral Resource Comparison

Annual comparisons of Mineral Resources are completed to quantify and verify changes due to model change, depletion, and changes due to the cut-off grade, where a calculated 2021 model value is compared to the actual declared 2020 Mineral Resources. Model changes and depletion at KCD, Sessenge, Pamao, Gorumbwa and Aerodrome were updated in 2021, with the rest remaining the same as 2020.

Oere, Pamao South and Mengu Village were new additions to the Kibali Mineral Resources in 2021 and as a result such a comparison to actual declared 2020 Mineral Resources could not be completed for these Mineral Resources.

A summary of the Year on Year Mineral Resource reconciliation is shown in Figure 14-57 to Figure 14-59.

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Figure 14-57 2021 Kibali Open Pit Mineral Resource Reconciliation

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Figure 14-58 2021 Kibali Underground Mineral Resource Reconciliation

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Figure 14-59 2021 Kibali Total Mineral Resource Reconciliation

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KCD

KCD Open Pit Resources

Open pit Mineral Resources are reported within the 2021 $1,500/oz Au pit shell, depleted with December 2021 mined surfaces, and reported above the underground reporting box wireframe solid (Table 14-89).

Table 14-89 KCD Open Pit 2021 vs 2020 Comparison Above the Underground Box



KCD Open Pit	M&I Mineral Resource			Inferred Mineral Resource
KCD Open Pit	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	6,957,619	2.80	625,880	874,492	1.8	49,270

2021	6,372,213	2.61	534,465	794,841	1.9	48,115

Net Change	-8%	-7%	-15%	-9%	7%	-2%

KCD open pit changes show a -93 koz Au decrease as a result of:

●

Depletion, which accounted for -219 koz Au.

●

Model update based on 380 additional drillholes with higher gold grades in the 3000 and 5000 lodes, which accounted for +115 koz Au.

●

A cut-off grade change (+12 koz Au), oxide/transition cut-off grade (COG) decreasing from 0.64 g/t Au to 0.60 g/t Au, and fresh COG decreasing from 0.78 g/t Au to 0.70 g/t Au, resulting in a tonnage weighted average COG of 0.69 g/t Au.

KCD Underground Resources

The current underground Mineral Resources were reported within optimised MSO shapes at 1.62 g/t Au and within the underground reporting limit wireframe solid. The Mineral Resource is depleted with EOY 2021 CMS stope scans and updated reporting exclusion solids. Results of the reconciliations are presented in Table 14-90.

Table 14-90 KCD Underground 2021 vs 2020 Comparison within Bounding Box and Within MSO



KCD UG	M&I Mineral Resource			Inferred Mineral Resource
KCD UG	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	82,399,207	4.29	11,362,474	11,294,579	3.0	1,103,549

2021	79,954,464	4.29	11,023,518	14,653,080	3.0	1,427,483

Net Change	-3%	0%	-3%	30%	0%	29%

KCD underground changes show a -15 koz Au decrease as a result of:

●

Depletion, which accounted for -572 koz Au (CMS scans and development as at month end December 2021).

●

Model update based on the addition of 615 new drillholes, which resulted in a net gain of 500 koz Au.

●

Gains in 3000 lode down plunge were offset by thinner lenses due to the steepening of the mineralisation in line with the structural measurements from core logging (-50 koz Au).

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Removal of the high yields for estimation within the very high-grade domains in the 5000 lode (+ 50 koz Au).

●

Additional drilling in the 9000 lode, resulted in a model with less continuous mineralisation shoots (-104 koz Au).

●

The addition of Inferred Mineral Resources in the 11000 lode, based on 17 drillholes spaced at approximately 50 m (+ 604 koz Au)

●

A slight COG grade reduction from 1.67 g/t Au to 1.62 g/t Au, resulting in a gain of 84 koz Au,

●

New stope exclusion solids to remove unrecoverable blocks around mined out areas (where both secondary and primary mined), which accounted for -27 koz Au.

Sessenge

Table 14-91 presents the Sessenge 2021 versus 2020 Mineral Resource comparison within the $1,500/oz Au pit shell.

Table 14-91 Sessenge 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell



Sessenge Main	M&I Mineral Resource			Inferred Mineral Resource
Sessenge Main	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	1,368,136	2.11	92,759	19,715	1.2	781

2021	1,482,920	2.01	95,798	117,806	1.6	6,098

Net Change	8%	-5%	3%	498%	31%	681%

Sessenge changes show a net +8koz Au increase as a result of:

●

Depletion, which accounted for -9 koz Au.

●

A model update based on 53 new drillholes, which resulted in a 10% volume increase in the mineralisation wireframes for 0.2 Mt and 14 koz Au.

●

Oxide COG decreasing from 0.64 g/t Au to 0.60 g/t Au, transitional material decreasing from 0.76 g/t Au to 0.71 g/t Au and fresh COG decreasing from 0.85 g/t Au to 0.76 g/t Au, resulting in a tonnage weighted average COG of 0.76 g/t Au and a gain of 3 koz Au for 2021.

Sessenge SW

Table 14-92 presents the Sessenge SW 2021 versus 2020 Mineral Resource comparison within the $1,500/oz Au pit shell.

Table 14-92 Sessenge SW 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell



Sessenge SW	M&I Mineral Resource			Inferred Mineral Resource
Sessenge SW	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	-	-	-	468,110	1.7	25,886

2021	-	-	-	468,110	1.7	25,886

Net Change	-	-	-	0%	0%	0%

There has been no change in model, depletion, or COG at Sessenge SW.

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Gorumbwa

Table 14-93 presents the Gorumbwa 2021 versus 2020 Mineral Resource comparison within the $1,500/oz Au pit shell.

Table 14-93 Gorumbwa 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell



Gorumbwa	M&I Mineral Resource			Inferred Mineral Resource
Gorumbwa	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	8,111,069	2.74	714,284	1,078,715	2.1	72,400

2021	6,324,341	3.04	617,501	1,549,560	2.5	123,403

Net Change	-22%	11%	-14%	44%	19%	70%

Gorumbwa changes show a -46 koz Au decrease as a result of:

●

Depletion, which accounted for -138 koz Au.

●

A model update based on 786 new grade control drillholes, which resulted in a net increase of 0.4 Mt and 91 koz Au.

The primary modelling direction/section lines orientated at 345° (across strike) were changed to 45° (along plunge) which, resulted in smoother more continuous mineralisation wireframes.

A high-grade zone was drilled in the main 1004 lode, below the historical SOKIMO pit, which resulted in an increase in grade.

Domain 1008 has been extended down plunge based on data from the new drilling.

●

Oxide/transition COG decreasing from 0.65 g/t Au to 0.60 g/t Au, and fresh COG decreasing from 0.75 g/t Au to 0.67 g/t Au, resulting in a tonnage weighted average COG of 0.67 g/t Au and a gain of 1 koz Au for 2021.

Pakaka

Table 14-94 outlines results for Pakaka 2021 versus 2020 Mineral Resource comparison within the $1,500 pit shell.

Table 14-94 Pakaka 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell



Pakaka	M&I Mineral Resource			Inferred Mineral Resource
Pakaka	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	8,287,890	2.68	715,066	718,439	3.5	80,377

2021	8,708,120	2.60	727,095	760,175	3.4	82,681

Net Change	5%	-3%	2%	6%	-3%	3%

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The Pakaka model was completed in 2019, with no subsequent drilling or model updates. The net Mineral Resource change at Pakaka reflects an increase of 0.5 Mt for 14 koz Au within the $1,500/oz Au pit shell, due to a cut-off grade change:

●

Oxide COG decreasing from 0.69 g/t Au to 0.64 g/t Au, transitional COG decreasing from 0.75 g/t Au to 0.69 g/t Au and fresh COG decreasing from 0.87 g/t Au to 0.78 g/t Au, resulting in a tonnage weighted average cut-off grade of 0.78 g/t Au and a gain of 14 koz Au for 2021.

Kombokolo

Table 14-95 outlines results for Pakaka 2021 versus 2020 Mineral Resource comparison within the $1,500 pit shell.

There has been no change in model, depletion, or cut-off grade for the Kombokolo Mineral Resource.



Kombokolo	M&I Mineral Resource			Inferred Mineral Resource
Kombokolo	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	551,287	2.57	45,594	127,949	1.8	7,199

2021	551,287	2.57	45,594	127,949	1.8	7,199

Net Change	0%	0%	0%	0%	0%	0%

Pamao and Pamao South

Table 14-95 outlines the Pamao 2021 versus 2020 Mineral Resource comparison within the $1,500/oz Au pit shell.

Table 14-95 Pamao – Pamao South 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell



Pamao & Pamao South	M&I Mineral Resource			Inferred Mineral Resource
Pamao & Pamao South	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	9,883,102	1.69	535,587	818,761	1.6	41,881

2021	13,237,692	1.71	725,860	906	0.8	24

Net Change	34%	1%	36%	-100%	-49%	-100%

Pamao and Pamao South changes show a net increase of 2.5 Mt for 148 koz Au as a result of:

●

The addition of Pamao South to the current Mineral Resource, adding 133 koz Au and a loss of 2 koz Au in Pamao with the model update based on new data.

●

289 grade control drillholes added to Pamao.

●

234 holes drilled to define the Pamao South mineral deposit.

●

A COG change in Pamao, oxide material decreasing from 0.67 g/t Au to 0.62 g/t Au, transitional material decreasing from 0.71 g/t Au to 0.66 g/t Au and fresh material decreasing from 0.83 g/t Au to 0.75 g/t Au, resulting in a tonnage weighted average COG of 0.71 g/t Au resulting in a gain of 17koz Au for 2021.

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Mengu Village

Table 14-96 outlines the Mengu Village 2021 Mineral Resources within the $1,500/oz Au pit shell. Mengu Village is a new addition to the Kibali Mineral Resource as an Inferred Mineral Resource.

Historically, Mengu Village was estimated using UC with allowance for an information effect incorporating important modifying factors such as likely grade control drilling, mining selectivity, and cut-off grade criteria. The application of UC technique is based on the premise that mining would be by open pit extraction. Kibali Goldmines took a decision in 2020 to exclude the Mengu Village Mineral Resource from the overall Kibali Mineral Resource base, due to the estimation technique used, which was a deviation from the standard practices employed at Kibali.

A review of the geology and geological interpretations in 2021, improved understanding on the controls to mineralisation and prompted an update of the model, using OK in the estimation. The drillhole spacing at Mengu Village is approximately 40 m by 40 m and in this regard, Mengu Village has been re-admitted to the Kibali Mineral Resources as an Inferred Mineral Resource in the 2021 declaration.

Table 14-96 Mengu Village 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell



Mengu Village	M&I Mineral Resource			Inferred Mineral Resource
Mengu Village	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	-	-	-	-	-	-

2021	-	-	-	1,058,128	1.4	47,994

Net Change	-	-	-	-	-	-

Megi-Marakeke-Sayi

A COG change, oxide material decreasing from 0.70 g/t Au to 0.65 g/t Au, transitional material decreasing from 0.75 g/t Au to 0.69 g/t Au and fresh material decreasing from 0.85 g/t Au to 0.76 g/t Au, resulting in a tonnage weighted average COG of 0.74 g/t Au and a gain of 57 koz Au for 2021.

Table 14-97 outlines the Megi-Marakeke-Sayi 2021 versus 2020 Mineral Resource comparison within the $1,500/oz Au pit shell.

Changes show a net increase of 1.2 Mt for 57 koz Au as a result of:

●

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Table 14-97 Megi-Marakeke-Sayi 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell



Megi Marakeke Sayi	M&I Mineral Resource			Inferred Mineral Resource
Megi Marakeke Sayi	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	10,110,572	1.77	574,853	670,522	1.9	40,459

2021	11,211,194	1.73	622,650	839,563	1.9	50,090

Net Change	11%	-2%	8%	25%	-1%	24%

Kalimva-Ikamva

Table 14-98 outlines the Kalimva-Ikamva 2021 versus 2020 Mineral Resource comparison within the $1,500/oz Au pit shell.

Table 14-98 Kalimva-Ikamva 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell



Kalimva Ikamva	M&I Mineral Resource			Inferred Mineral Resource
Kalimva Ikamva	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	6,914,012	2.91	645,953	126,560	4.4	17,950

2021	8,334,331	2.61	699,661	16,404	4.4	2,307

Net Change	21%	-10%	8%	-87%	-1%	-87%

Kalimva-Ikamva changes show a net increase of 1 Mt for 38koz Au as a result of:

●

A cut-off grade change, oxide material decreasing from 0.93 g/t Au to 0.86 g/t Au, transitional material decreasing from 0.95 g/t Au to 0.88 g/t Au and fresh material decreasing from 1.07 g/t Au to 0.97 g/t Au, resulting in a tonnage weighted average cut-off grade of 0.95 g/t Au and a gain of 38 koz Au for 2021.

Mengu Hill

Table 14-99 outlines the Mengu Hill 2021 versus 2020 Mineral Resource comparison within the $1,500/oz Au pit shell.

Table 14-99 Mengu Hill 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell



Mengu Hill	M&I Mineral Resource			Inferred Mineral Resource
Mengu Hill	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	754,506	2.66	64,425	431,097	3.7	51,956

2021	803,570	2.56	66,148	482,072	3.5	53,735

Net Change	7%	-4%	3%	12%	-8%	3%

Mengu Hill changes show a net increase of 0.10 Mt for 4 koz Au as a result of:

●

A cut-off grade change, oxide material decreasing from 0.78 g/t Au to 0.72 g/t Au, transitional material slightly increasing from 0.84 g/t Au to 0.86 g/t Au and fresh material decreasing from 1.14 g/t Au to 1.03 g/t Au, resulting in a tonnage weighted average COG of 0.99 g/t Au and a gain of 4 koz Au for 2021.

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Aerodrome

Aerodrome is a new addition to the Kibali Mineral Resources. Table 14-100 outlines Aerodrome 2021 Mineral Resource within the $1,500/oz Au pit shell.

Table 14-100 Aerodrome 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell



Aerodrome	M&I Mineral Resource			Inferred Mineral Resource
Aerodrome	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	564,667	1.46	26,433	17,894	1.4	787

2021	391,703	1.55	19,514	7,480	1.6	388

Net Change	-31%	6%	-26%	-58%	18%	-51%

Aerodrome changes show a net decrease of 0.2 Mt for -7 koz Au as a result of:

●

Depletion, which accounted for -3 koz Au.

●

An updated model based on an additional 71 drillholes completed in 2021, resulting in a wireframe volume loss of 5% due to the new data added.

●

A cut-off grade change, oxide material decreasing from 0.67 g/t Au to 0.58 g/t Au, transitional material decreasing from 0.71 g/t Au to 0.62 g/t Au and fresh material decreasing from 0.82 g/t Au to 0.62 g/t Au, resulting in a tonnage weighted average COG of 0.66 g/t Au and a gain of 0.3 koz Au for 2021.

Oere

Oere is a new addition to the Kibali Mineral Resources. Table 14-101 outlines Oere 2021 Mineral Resource within the $1,500/oz Au pit shell.

Table 14-101 Oere 2021 vs 2020 Comparison Within $1,500/oz Au Pit Shell



Oere	M&I Mineral Resource			Inferred Mineral Resource
Oere	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)	Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

2020	-	-	-	-	-	-

2021	3,111,561	2.15	214,649	1,987,480	1.7	106,273

Net Change	-	-	-	-	-	-

The Oere Mineral Resource is declared for the first time comprising of 3.1 Mt at 2.15 g/t Au for 215 koz Au Measured & Indicated Mineral Resources, in addition to 2 Mt at 1.7 g/t for 106 koz Au Inferred Mineral Resources. These new Mineral Resources are supported by:

●

163 drillholes completed at 40 m by 40 m in 2020.

●

An additional 140 drillholes completed in 2021, decreasing the drillhole spacing to between 20 m by 20 m and 20 m by 40 m in the $1,500/oz Au pit shell and 40 m by 100 m outside the shell.

●

An improved understanding on the geology and mineralisation controls with the addition of 22 interpreted geological sections; model showed good continuity even outside the shell where data are spaced at 100 m on strike.

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14.16 Discussion

External Mineral Resource Audits

An independent audit was undertaken in 2012 on the Mineral Resource estimate by Quantitative Group (QG) (Quantitative Group, 2013). The audit focussed primarily on KCD due to its dominant size. The results of the audit set out some minor recommendations to improve QA/QC compliance, sampling procedures, and modelling methodologies, all of which were since implemented.

Subsequently, an additional full resource audit by Optiro was completed in 2017, after the majority of the underground lodes were covered with an initial pass of underground resource definition drilling. Optiro concluded that the Mineral Resource estimation processes used by Kibali Goldmines were commensurate with industry best practice (Optiro, 2018b).

In September 2021, RSC completed an independent audit of the Mineral Resource and Mineral Reserve processes used at Kibali (RSC Ltd, 2021). The audit demonstrated that Mineral Resource and Mineral Reserve processes conform to good practices. However, RSC made a number of recommendations to Kibali Goldmines from a Mineral Resource perspective including:

●

Sensitivity analysis using different domain thresholds to assess potential impact on grade control – more readily achieved using implicit techniques to create domains, particularly high-grade ones or in areas with complexity or more diffuse grade boundaries, and

●

Create a comprehensive reconciliation code of practice (SOP) that includes a process flow map. Share a clear and consistent method with established checks and balances to improve calibration of resource and reserve estimates

Relative Accuracy / Confidence of the 2021 Mineral Resource Estimate

The QP offers the following conclusions regarding the relative accuracy / confidence of the 2021 Mineral Resource Estimate:

●

The application of optimised resource shapes (MSO) applies reasonable mineability constraints including a minimum mining width, a reasonable distance from current or planned development, and a measure of assumed profitability at the related resource cut-off grade. This change in UG reporting method has removed isolated areas of mineralisation and lowered the grade of the reported underground resource by reporting all material, geologically classified as mineralisation, within each mineable shape, whilst ensuring the overall shape meets the resource cut-off grade. Thereby ensuring that the Mineral Resources are reported in line with industry best practise with specific regard to underground resources only being reported if there is an intention to mine the material.

The QP is not aware of any mining, metallurgical, infrastructure, permitting or other relevant factors that could materially affect the Mineral Resource estimates.

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15	Mineral Reserve Estimate

15.1

Summary

For the open pit mines, economic pit shells were generated using the Lerch-Grossman algorithm within Whittle software and then used in the open pit mine design process and Mineral Reserve estimation.

The planning process incorporated appropriate modifying factors and the use of cut-off grades and other technical-economic investigations. Mineral Reserves are stated:

●

As of 31December 2021

●

At a gold price of $1,200/oz Au

●

As ROM grades and tonnage as delivered to the plant

A financial model was constructed to demonstrate that the Mineral Reserves are economically viable.

The total Kibali open pit and underground Mineral Reserves as of 31December 2021 are summarised in Table 15-1.

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Table 15-1 Kibali Mineral Reserves as of 31 December 2021



Type	Category	Tonnes (Mt)	Grade (g/t Au)	Contained Gold (Moz Au)	Attributable Gold¹ (Moz Au)

Stockpiles	Proven	0.32	3.17	0.032	0.015

Open Pits	Proven	11	2.28	0.79	0.35
Open Pits	Probable	26	2.51	2.1	0.95

Underground	Proven	21	4.54	3.0	1.4
Underground	Probable	25	4.54	3.7	1.6

Total Mineral Reserves	Proven	32	3.76	3.9	1.7
	Probable	51	3.50	5.8	2.6
	Proven and Probable	83	3.60	9.6	4.3

Notes

Attributable refers to the quantity attributable to Barrick based on Barrick’s 45% interest in the Kibali Goldmines. Mineral Reserves are reported on a 100% and attributable basis.

The Mineral Reserve estimate has been prepared according to CIM (2014) Standards and using CIM (2019) MRMR Best Practice Guidelines.

Open pit Mineral Reserves are reported at a gold price of $1,200/oz Au, and an overall weighted average cut-off grade of 0.96 g/t Au, including dilution and ore loss factors.

Underground Mineral Reserves are reported at a gold price of $1,200/oz Au and a cut-off grade of 2.02 g/t Au

Underground Mineral Reserves were estimated by Ismail Traore, MSc, FAusIMM, an officer of the company and QP, and reviewed by Simon Bottoms CGeol, MGeol, FAusIMM, an officer of the company and QP.

Numbers may not add due to rounding. Tonnes and contained gold are rounded to 2 significant figures. All Proven and Probable grades are reported to 2 decimal places.

The QPs have performed an independent verification of the block model tonnes and grade, and in their opinion, the process has been carried out to industry standards.

15.2 Mineral Reserve Estimation Process

Resource Models

The Mineral Reserve estimates use the block models prepared by the QP responsible for Mineral Resource estimation.

KCD, Sessenge, Aerodrome, and Gorumbwa are active open pits and therefore, the block models were depleted with EOY pit surveys.

The KCD block model is used for both the underground and open pit Mineral Reserve estimation. Four main mineralised zones, 5101, 5102, 9101, and 9105, comprise most of the underground

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Mineral Reserves, while five other mineralised zones, 3101, 3102, 5104, 5105, and 5110, contribute the remaining 12% of the Mineral Reserve (Figure 15-1).

LOGO

Source: Kibali Goldmines, 2021

Figure 15-1 KCD Underground Mining Zones

Open Pits

The estimation of Kibali open pit Mineral Reserves is based on the following key inputs:

●

Mineral Resource models for the estimated gold content and material weathering type.

●

Estimated processing and G&A costs.

●

Metallurgical recovery by material type and by deposit.

●

Geotechnical wall angle parameters.

●

KMS (mining contractor) 2021 pricing, which was used for mining costs.

●

Cut-off grade analysis using final estimated costs derived from pit designs and pit schedules, and finalised processing and administration costs.

●

The cut-off grades have been estimated for each material type for all nine reserve pits included in the 2021 Mineral Reserve estimate. These are based on a gold price of $1,200/oz Au for all pits, with the exception of $1,300/oz Au for the Sessenge and Oere pits, and $1,500/oz Au for the Aerodrome pit and include dilution, royalties, processing costs and recoveries, G&A costs, and ore mining costs. All Mineral Reserves, including Aerodrome, Sessenge, and Oere are profitable at $1,200/oz Au sales price and thus the Mineral Reserve and supporting cash flow statements are reported at $1,200/oz Au.

●

All Mineral Reserves, including Aerodrome, Sessenge, and Oere are profitable at $1,200/oz Au sales price and thus the Mineral Reserve and supporting cash flow statements are reported at $1,200/oz Au.

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●

Open pit Mineral Reserves are reported at a gold price of $1,200/oz Au and a tonne-weighted average cut-off grade of 0.96 g/t Au, including dilution and ore loss factors. KCD pit shares an interface with the underground at the 5685 mRL.

The Open Pit Mineral Reserves were estimated as follows:

●

Open pit stockpiles estimated as of 31December 2021.

●

Only Measured and Indicated Mineral Resources were used for conversion.

●

Depletion of the KCD, Sessenge, Aerodrome, and Gorumbwa pit block models with EOY actual survey face positions as of 31December 2021.

●

Use of an integrated mine and feed schedule.

Underground Mine

The estimation was undertaken using Datamine Studio 5D software. The block models used were sub cell block models. The geological zones (including mineralised zones) were defined by three dimensional wireframes solids and surfaces. Both the block models and wireframes were created in Maptek Vulcan by the Barrick geological team. The block models and wireframes were converted to a Datamine format for use in Datamine Studio 5D. The 2021 Mineral Reserves estimation process was estimated by manually updating MSO generated stope shapes, which had been generated using the July 2021 block model.

The process undertaken for estimation of the 2021 Underground Mineral Reserves was as follows:

●

Define mining method by area, based on the geometry, geotechnical considerations, and the mine development requirement to access the orebody.

●

Review the historical and LOM planned costs to determine cut off grades.

●

Use the Datamine MSO to evaluate the geological block model mineralisation and determine the areas to be included and the overall mining shapes. Due to geotechnical, productivity and practical mining constraints, the MSO shapes have not been used for the Mineral Reserve Estimate. The resulting stope shapes were digitised as required. The parameters used for generating the MSO shapes are discussed in Section 15.5.

●

Manually create stope section strings to follow geological block model mineralisation above cut-off grade, using the MSO shapes as a guide. Strings are based on level intervals determined in the previous Mineral Reserve estimate. Planned dilution is included in the stope shape to create a mineable stope shape.

●

Create mineable stope wireframes from the strings.

●

Deplete the stope wireframes by the parts of the mine survey solids that intersected them to remove development drives and parts of stopes.

●

Evaluate stope wireframes against the geological block model (estimate the tonnes, grade and ounces of the stopes).

●

Design the development required to access the mineable stopes.

●

Use Datamine EPS to calculate the diluted mined tonnes, grades and contained metal. This included mining dilution added as a varying percentage depending on hanging wall exposure, stope sequence (primary, secondary, advancing transverse or longitudinal) and number of paste fill exposures. Mining loss was subtracted as a percentage from diluted tonnes and contained metal.

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●

Assess economics of mining areas and mining individual stopes.

●

Exclude sub-economic stopes from the short term and LOM plan.

●

Classify the Mineral Reserve into Proven and Probable Reserves on a proportional basis as described in Section 15.5.

Stockpiles

Stockpiles are comprised of mineralised material stored at the surface ROM pad, originating from both open pit and underground production. Each stockpile is filled with similar material types, with an established grade range and oxidation state, tracked as part of normal mining operations and metal accounting. The stockpiles are measured by weekly drone survey. The grade and tonnage of the open pit stocks are estimated according to source dig blocks and number of truck counts, using a weighbridge to adjust for fluctuations in both density and truck fill factor. Grade and tonnage of underground stocks are estimated according to shaft skip weights and ore pass truck counts and their source blasts from stopes, adjusting for the presence of paste dilution.

Location of the stockpiles is shown in Figure 5-2 in Section 5.3.

15.3 Economic Parameters

Open Pit

Gold Price and Royalties

With the exception of the Oere and Sessenge pits, which use a gold price of $1,300/oz Au, and the Aerodrome Pit which uses $1,500/oz Au to define their optimum pit shell; all the other reserve pit shells are based on a gold price of $1,200/oz Au. All Mineral Reserves, including Aerodrome and Sessenge are profitable at $1,200/oz Au sales price and thus the Mineral Reserve and supporting cash flow statements are reported at $1,200/oz Au. This is in line with Barrick corporate guidelines, which considers long-term gold price forecasts. Gold price sensitivities were prepared for all the pits, the decision on a higher price for the other pits is discussed in detail under Section 15.4.

Royalties payable to the DRC government remained unchanged from the year end 2020 estimate. A total royalty of 4.7% of gold revenue inclusive of 1% shipment fees was used for the year-end 2021 estimate.

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Processing Costs

Processing costs for the year were reviewed as there was a slight change compared to the 2020 LOM projections.

General and Administration Cost

The G&A cost was reviewed based on LOM expectations and actuals for the year end 2021. A downward adjustment of 9% was noted and this was subsequently applied in the 2021 Mineral Reserve estimation.

Mining Costs

The mining costs used for the 2021 pit optimisations were derived from the KMS 2020 budget unit plan (BUP) and Long-Term Review (LTR) pricing for the Kibali open pit operations.

Mining Cost Adjustment Factors (MCAF) were generated from various bench by bench waste mining costs received for all deposits. The waste mining cost is inclusive of fuel cost, drill and blast cost per bench, pre-split cost, explosive cost per tonne, mining departmental cost, pit dewatering, rehabilitation cost, and contractor fixed costs.

The MCAF were then imported into their respective block models and assigned to the corresponding benches in Surpac software for the creation of economic block models.

Cut-Off Grade

The Mineral Reserves are based on a marginal cut-off grade. Mineral Resources contained within the final pit designs were evaluated against these cut-off grades to produce the Open Pit Proven and Probable Mineral Reserves.

Cut-off grade sensitivities were trialled by adjusting the gold price (Table 15-2 to Table 15-4).

The QP responsible for the open pit Mineral Reserve estimate considers that the process used is appropriate for the estimation of Kibali open pit Mineral Reserves.

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Table 15-2 KCD, Megi-Marakeke-Sayi, Pakaka Open Pits – Marginal and Full Grade Ore (FGO) Cut-Off Grade for Different Material Types



	Deposit			KCD			Megi Marakeke Sayi			Pakaka
	Material Type	Unit	Constant	Oxide	Trans	Fresh	Oxide	Trans	Fresh	Oxide	Trans	Fresh

Mining	Waste Cost (per tonne mined)	$/t		2.92	2.97	3.09	3.18	3.23	3.28	2.72	2.80	2.88
	Extra Ore Cost (per ore tonne) – GC + Core - Rehandle + Overhaul	$/t		1.27	1.27	1.27	1.25	1.25	1.25	1.38	1.38	1.38
	GC Only (per tonne mined)	$/t		0.75	0.75	0.75	0.75	0.75	0.75	0.75	0.75	0.75
	Dilution	%	10	10
	Ore Loss	%	3	3

Process	Haulage Cost per ore tonne	$/t	2.50				2.24	2.24	2.24	1.05	1.05	1.05
	Process Cost (per ore tonne milled)	$/t	14.93	15.04	15.04	17.85	15.04	15.04	17.85	15.04	15.04	17.85
	Process Recovery	%		90.1	90.1	86.1	90.0	90.0	89.5	89.0	89.0	80.0
	Plant Throughput	Mtpa	7.2

G&A	G&A (per ore tonne milled)	$/t	8.47	8.47

Revenue	Gold Price (Reserve)	$/oz Au	1,200	1,200
	Gold Price	$/g Au	31.10348	38.58
	Gold Royalties	$/oz Au	4.70%	56.4
	Net Gold Price	$/oz Au		1,143.6
	Net Gold Price	$/g Au		36.8

	Total Process Cost (per ore tonne milled)	$/t		15.04	15.04	17.85	17.28	17.28	20.09	16.09	16.09	18.90
	Total Mining Cost (per ore tonne mined)	$/t		31.92	32.43	33.69	16.60	16.82	17.10	38.09	39.20	40.34
	Marginal Cut-off Grade	g/t Au		0.82	0.82	0.96	0.90	0.90	1.00	0.87	0.87	1.07
	Strip Ratio			9.5			3.8			12.5
	FGO Cut-off Grade	g/t Au		1.84	1.86	2.09	1.41	1.42	1.53	2.11	2.14	2.53

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Table 15-3 Pamao, Kalimva-Ikamva, Gorumbwa Open Pits – Marginal and Full Grade Ore Cut-Off Grade for Different Material Types



	Deposit			Pamao			Kalimva-Ikamva			Gorumbwa
	Material Type	Unit	Constant	Oxide	Trans	Fresh	Oxide	Trans	Fresh	Oxide	Trans	Fresh

Mining	Waste Cost (per tonne mined)	$/t		2.85	2.88	2.95	2.35	2.59	2.92	3.29	3.14	3.24
	Extra Ore Cost (per ore tonne) – GC + Core - Rehandle + Overhaul	$/t		1.31	1.31	1.31	6.25	6.25	6.25	1.28	1.28	1.28
	GC Only (per tonne mined)	$/t		0.75	0.75	0.75	0.75	0.75	0.75	0.75	0.75	0.75
	Dilution	%	10	10
	Ore Loss	%	3	3

Process	Haulage Cost per ore tonne	$/t	2.50	1.05	1.05	1.05	5.00	5.00	5.00
	Process Cost (per ore tonne milled)	$/t	14.93	15.04	15.04	17.85	15.04	15.04	17.85	15.04	15.04	17.85
	Process Recovery	%		90.9	85.0	85.0	90.0	89.0	89.0	90.0	90.0	90.0
	Plant Throughput	Mtpa	7.2

G&A	G&A (per ore tonne milled)	$/t	8.47	8.47

Revenue	Gold Price (Reserve)	$/oz Au	1,200	1,200
	Gold Price	$/g Au	31.10348	38.58
	Gold Royalties	$/oz Au	4.70%	56.4
	Net Gold Price	$/oz Au		1,143.6
	Net Gold Price	$/g Au		36.8

	Total Process Cost (per ore tonne milled)	$/t			16.09	18.9	20.04	20.04	22.85	15.04	15.04	17.85
	Total Mining Cost (per ore tonne mined)	$/t			20.40	20.82	31.72	34.36	37.92	39.93	38.15	39.34
	Marginal Cut-off Grade	g/t Au			0.91	1.01	1.16	1.17	1.26	0.82	0.82	0.92
	Strip Ratio						9.9			10.7
	FGO Cut-off Grade	g/t Au			1.58	1.70	2.00	2.11	2.33	2.11	2.05	2.18

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Table 15-4 Sessenge, Aerodrome, and Oere Open Pits - Marginal and Full Grade Ore Cut-Off Grade for Different Material Types



	Deposit			Sessenge			Aerodrome			Oere
	Material Type	Unit	Constant	Oxide	Trans	Fresh	Oxide	Trans	Fresh	Oxide	Trans	Fresh

Mining	Waste Cost (per tonne mined)	$/t		2.62	2.68	2.80	1.70	2.01	2.33	2.69	2.97	3.02
	Extra Ore Cost (per ore tonne) – GC + Core - Rehandle + Overhaul	$/t		1.24	1.24	1.24	1.25	1.25	1.25	6.25	6.25	6.25
	GC Only (per tonne mined)	$/t		0.75	0.75	0.75	0.75	0.75	0.75	0.75	0.75	0.75
	Dilution	%	10	10
	Ore Loss	%	3	3

Process	Haulage Cost per ore tonne	$/t	2.50				1.05	1.05	1.05	4.5	4.5	4.5
	Process Cost (per ore tonne milled)	$/t	14.93	15.04	15.04	17.85	15.04	15.04	17.85	15.04	15.04	17.85
	Process Recovery	%		90.3	75.9	81.0	90.0	88.0	85.0	88.0	86.5	87.0
	Plant Throughput	Mtpa	7.2

G&A	G&A (per ore tonne milled)	$/t	8.47	8.47

Revenue	Gold Price (Reserve)	$/oz Au	-	1,300			1,500			1,300
	Gold Price	$/g Au	31.10348	41.80
	Gold Royalties	$/oz Au	4.70%	61.10
	Net Gold Price	$/oz Au		1,238.9
	Net Gold Price	$/g Au		39.8

	Resource Gold Price	$/oz Au	1,500	1,500

	Total Process Cost (per ore tonne milled)	$/t		15.04	15.04	17.85	16.09	16.09	18.90	19.54	19.54	22.35
	Total Mining Cost (per ore tonne mined)	$/t		14.35	14.65	15.24	15.12	17.60	20.20	53.44	58.35	59.35
	Marginal Cut-off Grade	g/t Au		0.76	0.90	0.94	0.79	0.81	0.92	1.08	1.09	1.18
	Strip Ratio			4.0			7.1			16.6
	FGO Cut-off Grade	g/t Au		1.16	1.39	1.42	1.22	1.32	1.53	2.56	2.76	2.86

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Underground

Cut-Off Grade

The underground Mineral Reserve cut-off grade is updated once a year using inputs parameters based on recent operating experience, projected costs, and Barrick corporate guidance. The cut-off grade parameters are as follows:

●

Gold price per ounce

●

LOM production costs

●

Processing recovery

●

Processing costs

●

G&A costs

●

Royalty costs

A break-even cut-off grade (BCOG) is used for Mineral Reserve estimation. All stopes and development material that fail to meet the BCOG are classified as waste. Incremental cut-off grade (ICOG) is used on the case-by-case basis.

BCOG is the grade of material that will generate revenue from the sale of the finished product at the metal price after applying the cost of mining, transporting/hauling, processing, royalties, and G&A. It is defined using the following formula:

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PC: Total processing operating costs (include process sustaining capital) ($/t)

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MC: Total mine operating costs (include secondary development and mining sustaining capital, exclude capital development costs) ($/t)

●

G&A: General and administrative costs ($/t)

●

REC: Planned recovery of the metal (%)

●

MP: Selling price of metal ($/oz)

●

RO: Royalty (%)

●

SC: Selling costs (include smelter, refinery and transportation costs as required)

ICOG is applied to the mineralised part of the deposit below the BCOG that can incrementally add value to the operation under certain circumstances. It is used in the following circumstances:

●

When mine development goes through low-grade material in order to expose higher grade production areas or stopes.

●

When there is low-grade material near an already developed part of the mine. However, this low-grade material should never displace available higher grade material above the BCOG. These materials are assessed on a case-by-case basis and may be scheduled for mining toward the end of the LOM if practical.

●

When the mill is operating at capacity and the mine has the ability to provide material for placements in the stockpiles that can be economically processed at a later stage.

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The ICOG carries only the variable portion of the mining costs (drilling, blasting, mucking, hoisting), process operating costs, G&A costs, royalties, and re-handling costs if stockpiling is required. Development costs (capital or operating) are only included if the development is required to mine the incremental ore.

ICOG is calculated using the following formula:

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PC (var): Variable processing operating costs (excludes process sustaining capital)

●

MC (var): Variable mining operating costs (excludes mining sustaining capital)

●

G&A (var): Variable G&A operating costs (excludes G&A sustaining capital)

●

REC: Planned recovery of the metal (%)

●

MP: Selling price of metal ($/oz)

●

RO: Royalty (%)

●

SC: Selling costs (include smelter, refinery and transportation costs as required)

Table 15-5 shows the BCOG and ICOG calculation for the Underground Mineral Reserves. Figure 15-2 shows the stopes above the BCOG.

Table 15-5 Kibali Underground Mine – Cut-Off Grade Calculation



Description	Units	BCOG	ICOG Development	ICOG Stoping

Gold price	$/oz Au	1,200	1,200	1,200

Process plant gold recovery	%	90.0	90.0	90.0

Royalty	%	4.7	4.7	4.7

Mine Production and Backfill	$/t mined	36.17	5.18	25.32

Sustaining Capital	$/t mined	3.97

Processing	$/t milled	17.85	17.85	17.85

Site G & A	$/t milled	8.47	8.47	8.47

Total unit cash costs	$/t milled	66.47	31.51	51.65

Mining Cut Off Grade	g/t Au	2.02	0.96	1.57

The 2021 Underground Mineral Reserve cut-off grade is 2.02 g/t Au, compared to 2.09 g/t Au used in 2020. The decrease in the cut-off grade is mainly driven by higher process recovery and lower processing and G&A costs. The reduction of the G&A and processing cost is mainly driven by the additional tonnes mined and processed on an annual basis within the new LOM.

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Source: Kibali Goldmines, 2021

Figure 15-2 Kibali Underground Mineral Reserves – Stopes Above the BCOG

The QP considers that the process used to calculate the cut-off grade is appropriate for the Kibali underground mine.

15.4 Pit Optimisation

Data checks on block models received from the Mineral Resource Department were conducted. This included checks for missing cells, absent values, density checks, grade errors and correctly assigned weathering profiles. All models received had waste blocks built into them.

Economic models were generated from the Mineral Resource block models with the inclusion of the MCAFs for each of the nine target deposits. Approved geotechnical slope domains and angles based on rock characteristics and behaviour were also assigned to the block models before converting them to block models suitable for optimisation. These were then imported into Geovia Whittle software version 4.7.2 for the pit optimisation exercise.

The initial optimisation run considered the Measured and Indicated Resources, with Inferred Mineral Resources excluded. These were run with a gold price of $1,200/oz Au for Mineral Reserves.

A second set of optimisations was conducted with the inclusion of Inferred Mineral Resources. These optimisations were used to quantify the Inferred portions of the deposits, determine the impact on the mine plan, and to provide direction to the Mine Resource Management (MRM) and Exploration departments for possible targets for drilling and resource conversion.

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The advanced grade control and infill drilling campaign in 2021 resulted in major changes to Pamao and brought in Aerodrome and Oere pits as new deposits within Mineral Reserves. The Pakaka and Gorumbwa pits were updated using a gold price of $1,200/oz Au from $1,000/oz Au previously. All the other pits remained unchanged compared to the 2020 Mineral Reserves. Table 15-6 shows the comparison between the pit shells for Gorumbwa from the 2021 Whittle at a gold price of $1,000/oz Au and the 2021 Whittle at $1,200/oz Au, while Table 15-7 shows a comparison of the 2020 and 2021 Whittle results for Pamao at $1,200/oz Au.

For the KCD deposit, which is also extracted by underground methods, the optimisation was run by constraining the block model to the underground-open pit interface at the 5680 mRL. The block model imported into the Whittle was therefore constrained to this elevation. The optimal pit shell is the $1,200/oz Au shell, which delivers higher ounces and no downside risk at a lower gold price. Pushback One, Two and Three of the deposit have successfully been fully depleted as of end of 2021, with only Pushback three North remaining to be mined in 2027.

No other physical surface infrastructures constraints were applied for any of the orebodies in the Mineral Reserves.

Table 15-6 Comparison of Whittle Results for the Gorumbwa Pit at $1,000/oz Au and $1,200/oz Au



Parameter	2021 Whittle Run ($1,000/oz Au)	2021 Whittle Run ($1,200/oz Au)	Diff	% Diff
Proven

Ore Tonnes (t)	1,202,271	1,620,577	418,306	35%

Ore Grade (g/t Au)	2.43	2.18	1.46	-10%

Ounces (oz Au)	94,018	113,653	19,635	21%
Probable

Ore Tonnes (t)	3,342,432	4,407,870	1,065,438	32%

Ore Grade (g/t Au)	3.33	3.20	2.80	-4%

Ounces (oz Au)	357,599	453,422	95,823	27%
Inferred

Inferred Tonnes (t)	174,257	658,056	483,799	278%

Inferred Grade (g/t Au)	3.00	2.17	1.87	-28%

Inferred Ounces (oz Au)	16,813	45,847	29,034	173%
Total

Total Ore Tonnes (t)	4,718,960	6,686,503	1,967,543	42%

Grade (g/t Au)	3.09	2.85	2.28	-8%

Total Ounces (oz Au)	468,430	612,922	144,492	31%

Total Waste Tonnes (t)	36,431,364	61,095,182	24,663,818	68%

Strip Ratio with Inferred Ore	7.7	9.1	1.42	18%

Strip Ratio without Inferred Ore	8.1	10.2	2.19	27%

% of Inferred Tonnes	4%	10%	-	-

% of Inferred Ounces	4%	7%	-	-

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Table 15-7 Comparison of Whittle Results for the Pamao Pit with 2020 Results at $1,200/oz Au



$1,200/oz Au Shell	2020 Whittle Run	2021 Whittle Run	Diff	% Diff

Proven

Ore Tonnes (t)	322,552	328,604	6,052	2%

Ore Grade (g/t Au)	2.29	2.24	- 0.05	-2%

Ounces (oz Au)	23,739	23,676	-62	0%

Probable

Ore Tonnes (t)	4,775,636	6,795,260	2,019,624	42%

Ore Grade (g/t Au)	2.12	1.92	-0.20	-9%

Ounces (oz Au)	325,435	419,718	94,283	29%

Inferred

Inferred Tonnes (t)	179,719	-	-179,719	-100%

Inferred Grade (g/t Au)	2.74	-	-2.74	-100%

Inferred Ounces (oz Au)	15,821	-	-15,821	-100%

Total

Total Ore Tonnes (t)	5,277,907	7,123,864	1,845,957	35%

Grade (g/t Au)	2.15	1.94	-0.22	-10%

Total Ounces (oz Au)	364,995	443,395	78,400	21%

Total Waste Tonnes (t)	27,828,970	30,104,549	2,275,579	8%

Strip Ratio with Inferred Ore	5.3	4.2	-1.0	-20%

Strip Ratio without Inferred Ore	5.5	4.2	-1.3	-23%

% of Inferred Tonnes	3%	0%

% of Inferred Ounces	4%	0%

Sensitivity Analysis

An initial optimisation was run on the standard $1,200/oz Au Mineral Reserve gold price. Gold price sensitivities were then run for gold prices of $400/oz Au to $2,000/oz Au at an increment of $100/oz Au to produce a set of nested pits shells (Table 15-8 to Table 15-14)

Various sensitivities at different gold prices were conducted for the different deposits to determine the optimal gold price to be used for the 2021 Mineral Reserves on a case-by-case basis. Analysis was completed on cash cost, strip ratios, cash flows generated as well as geological drill coverage for each deposit (Table 15-8 to Table 15-14).

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Table 15-8 Sessenge Gold Price Sensitivities



Pit Size ($/oz Au)	Cash Flow ($M)	Mineralised Material (Mt)	Grade (g/t Au)	Waste (Mt)	Mining Cost ($M)	Process Cost ($M)	Royalty ($M)	Mining Cost ($/t)	Stripping Ratio (t:t)	Ounces Mined (koz Au)	Recovery (%)	Gold Produced (koz Au)	Cash Cost ($/oz Au)

400	2.2	0.0	3.18	0.03	-0.2	-1.0	-0.1	3.18	0.7	4	78%	3	450

500	3.9	0.1	3.20	0.08	-0.5	-1.7	-0.3	3.19	1.2	7	77%	5	463

600	5.0	0.1	2.96	0.12	-0.7	-2.6	-0.3	3.20	1.2	9	77%	7	497

700	6.0	0.1	2.69	0.19	-1.0	-3.6	-0.4	3.20	1.3	12	77%	9	545

800	6.4	0.2	2.60	0.24	-1.3	-4.1	-0.5	3.20	1.5	13	77%	10	569

900	9.5	0.3	2.70	1.52	-5.9	-8.0	-1.0	3.26	5.0	26	78%	21	728

1,000	9.8	0.3	2.66	1.68	-6.5	-8.7	-1.0	3.25	5.1	28	78%	22	744

1,100	10.8	0.5	2.66	3.41	-12.6	-12.7	-1.5	3.25	7.3	40	78%	32	849

1,200	10.9	0.5	2.56	3.69	-13.7	-14.2	-1.6	3.25	7.0	43	78%	34	870

1,300	10.1	1.3	2.22	9.93	-36.0	-34.7	-3.3	3.22	7.9	90	79%	71	1,047

1,400	9.7	1.3	2.17	10.36	-37.6	-36.9	-3.5	3.22	7.7	93	79%	74	1,060

1,500	6.8	1.6	2.11	12.84	-46.3	-43.5	-4.0	3.21	8.1	107	79%	85	1,110

1,600	6.4	1.6	2.10	13.06	-47.1	-44.4	-4.0	3.21	8.1	109	79%	86	1,116

1,700	5.9	1.6	2.09	13.35	-48.1	-45.3	-4.1	3.21	8.1	110	79%	87	1,123

1,800	-117.8	6.6	1.90	87.01	-298.5	-184.6	-15.0	3.19	13.1	404	79%	391	1,560

1,900	-127.8	7.0	1.89	92.23	-316.3	-194.7	-15.7	3.19	13.2	424	79%	335	1,572

2,000	-135.3	7.2	1.88	95.82	-328.4	-200.2	-16.2	3.19	13.3	435	79%	344	1,584

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Table 15-9 Pamao Gold Price Sensitivities



Pit Size ($/oz Au)	Cash Flow ($M)	Mineralised Material (Mt)	Grade (g/t Au)	Waste (Mt)	Mining Cost ($M)	Process Cost ($M)	Royalty ($M)	Mining Cost ($/t)	Stripping Ratio (t:t)	Ounces Mined (koz Au)	Recovery (%)	Gold Produced (koz Au)	Cash Cost ($/oz Au)

400	10.2	0.1	3.26	0.19	-1.0	-3.9	-0.7	3.93	1.3	15	86%	13	425

500	30.1	0.6	2.52	0.64	-3.7	-16.7	-2.5	4.04	1	51	87%	44	519

600	54.7	1.3	2.5	2.57	-11.3	-34.5	-5.0	3.66	2	102	86%	88	577

700	73.3	1.9	2.44	5.16	-20.6	-51.2	-7.2	3.51	2.7	148	86%	127	622

800	93.5	2.8	2.27	8.30	-32.4	-77.3	-10.0	3.47	2.9	208	86%	178	674

900	105.8	3.7	2.15	11.09	-42.9	-99.7	-12.2	3.45	3	254	86%	217	713

1,000	120.9	4.7	2.16	19.92	-71.2	-129.2	-15.9	3.31	4.2	329	85%	281	770

1,100	128.8	6.1	2.03	26.33	-93.6	-166.3	-19.2	3.3	4.3	398	85%	340	821

1,200	130.4	7.1	1.94	30.10	-107.5	-194.6	-21.3	3.31	4.2	443	85%	378	855

1,300	127.4	8.8	1.82	36.86	-131.6	-240.5	-24.6	3.31	4.2	513	85%	437	908

1,400	122.3	9.7	1.78	42.64	-151.1	-268.2	-26.7	3.29	4.4	556	85%	474	942

1,500	114.0	10.6	1.74	48.85	-171.2	-292.4	-28.5	3.27	4.6	594	85%	505	974

1,600	103.1	11.4	1.71	54.13	-188.4	-314.3	-29.9	3.26	4.8	623	85%	530	1,005

1,700	78.3	12.6	1.67	66.14	-226.0	-349.1	-32.2	3.22	5.3	673	85%	571	1,063

1,800	63.2	13.2	1.64	72.75	-246.5	-366.7	-33.4	3.21	5.5	697	85%	591	1,093

1,900	23.3	14.3	1.61	88.83	-295.6	-400.4	-35.5	3.17	6.2	742	85%	629	1,163

2,000	7.1	14.7	1.6	95.95	-316.9	-411.0	-36.2	3.16	6.5	758	85%	643	1,189

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Table 15-10 Kalimva-Ikamva Gold Price Sensitivities



Pit Size ($/oz Au)	Cash Flow ($M)	Mineralised Material (Mt)	Grade (g/t Au)	Waste (Mt)	Mining Cost ($M)	Process Cost ($M)	Royalty ($M)	Mining Cost ($/t)	Stripping Ratio (t:t)	Ounces Mined (koz Au)	Recovery (%)	Gold Produced (koz Au)	Cash Cost ($/oz Au)

400	19.9	0.2	4.85	0.71	-2.5	-6.6	-1.1	4.19	3.1	36	86%	30	335

500	66.2	1.0	4.28	6.29	-19.3	-30.1	-4.2	3.51	6.1	142	85%	121	443

600	109.0	2.0	3.99	13.71	-41.3	-58.0	-7.6	3.42	6.9	255	85%	218	490

700	136.7	2.9	3.79	21.27	-63.5	-83.6	-10.4	3.37	7.4	348	85%	297	530

800	147.1	3.4	3.69	27.00	-80.0	-98.4	-11.9	3.33	8	400	85%	341	558

900	151.6	3.7	3.6	30.68	-90.7	-109.2	-12.9	3.31	8.2	432	85%	368	578

1,000	159.6	4.6	3.45	41.88	-123.2	-136.0	-15.3	3.27	9	514	85%	439	626

1,100	160.2	5.3	3.3	48.59	-143.2	-156.3	-16.8	3.27	9.1	565	85%	481	657

1,200	159.5	5.9	3.09	49.03	-146.0	-173.9	-17.6	3.33	8.3	589	85%	502	672

1,300	156.5	6.7	2.86	49.05	-148.1	-196.8	-18.4	3.41	7.3	616	85%	525	692

1,400	149.2	7.7	2.62	49.80	-152.9	-226.1	-19.4	3.5	6.5	649	85%	553	720

1,500	132.2	9.1	2.42	57.12	-176.6	-267.4	-21.1	3.53	6.3	708	85%	603	771

1,600	116.0	10.5	2.24	60.72	-190.3	-308.9	-22.5	3.59	5.8	756	85%	644	810

1,700	98.3	11.9	2.08	63.43	-201.2	-350.2	-23.8	3.66	5.3	798	85%	680	846

1,800	78.2	13.2	1.97	66.73	-213.6	-389.2	-25.0	3.71	5	837	85%	713	880

1,900	55.8	14.4	1.88	71.21	-228.7	-422.7	-25.9	3.72	5	869	85%	740	915

2,000	42.1	15.4	1.81	71.92	-233.4	-451.8	-26.7	3.77	4.7	894	85%	762	935

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Table 15-11 Pakaka Gold Price Sensitivities



Pit Size ($/oz Au)	Cash Flow ($M)	Mineralised Material (Mt)	Grade (g/t Au)	Waste (Mt)	Mining Cost ($M)	Process Cost ($M)	Royalty ($M)	Mining Cost ($/t)	Stripping Ratio (t:t)	Ounces Mined (koz Au)	Recovery (%)	Gold Produced (koz Au)	Cash Cost ($/oz Au)

400	23.1	0.2	6.59	0.86	-2.9	-5.6	-1.5	2.97	4.4	41	80%	33	302

500	31.5	0.4	5.56	1.96	-6.3	-10.2	-2.3	2.93	5.6	63	80%	50	373

600	42.6	0.6	5.05	4.87	-14.9	-17.6	-3.5	2.88	8.1	98	80%	79	458

700	50.0	0.8	4.87	7.66	-23.2	-23.4	-4.6	2.87	9.5	126	80%	101	506

800	64.1	1.8	3.77	15.59	-47.5	-50.9	-7.7	2.88	8.9	212	80%	170	623

900	65.0	1.9	3.64	16.17	-49.5	-55.0	-8.0	2.88	8.6	221	80%	177	634

1,000	65.8	2.1	3.47	17.89	-54.9	-62.3	-8.6	2.88	8.3	239	80%	192	656

1,100	60.6	3.4	3.38	41.55	-123.4	-98.8	-13.3	2.85	12.2	369	80%	296	795

1,200	53.0	4.3	3.17	53.49	-158.9	-126.0	-15.9	2.85	12.3	441	80%	354	850

1,300	48.3	4.6	3.12	57.99	-172.1	-134.3	-16.7	2.85	12.6	463	80%	371	870

1,400	44.7	4.7	3.11	60.90	-180.5	-137.9	-17.1	2.85	12.8	474	80%	380	882

1,500	-33.0	6.9	3.13	124.25	-362.3	-199.8	-24.9	2.83	18.1	691	80%	554	1,059

1,600	-44.9	7.3	3.04	130.01	-379.5	-212.1	-25.8	2.84	17.8	714	80%	573	1,078

1,700	-60.2	7.7	3	138.11	-403.2	-223.1	-26.7	2.84	18	739	80%	593	1,102

1,800	-68.0	7.9	2.96	141.60	-413.4	-228.8	-27.1	2.84	18	750	80%	601	1,113

1,900	-90.1	8.3	2.91	152.07	-443.8	-242.1	-28.1	2.84	18.3	778	80%	624	1,144

2,000	-99.4	8.5	2.89	156.28	-456.1	-246.8	-28.4	2.84	18.4	788	80%	632	1,157

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Table 15-12 Megi-Marakeke-Sayi Gold Price Sensitivities



Pit Size ($/oz Au)	Cash Flow ($M)	Mineralised Material (Mt)	Grade (g/t Au)	Waste (Mt)	Mining Cost ($M)	Process Cost ($M)	Royalty ($M)	Mining Cost ($/t)	Stripping Ratio (t:t)	Ounces Mined (koz Au)	Recovery (%)	Gold Produced (koz Au)	Cash Cost ($/oz Au)

400	32.4	0.5	3.26	0.28	-2.5	-12.5	-1.9	4.12	0.6	47	88%	41	406

500	48.6	0.8	2.86	0.68	-5.1	-23.4	-3.0	4.02	0.8	77	88%	67	468

600	70.8	1.4	2.67	2.07	-11.8	-40.4	-4.8	3.83	1.4	123	87%	108	530

700	87.0	2.0	2.49	3.27	-17.7	-57.5	-6.4	3.79	1.6	163	87%	142	575

800	101.5	2.7	2.33	4.83	-25.3	-77.7	-8.1	3.76	1.8	205	87%	179	621

900	118.4	3.8	2.21	8.89	-42.2	-108.3	-10.6	3.67	2.3	270	87%	235	685

1,000	129.7	4.9	2.08	12.16	-56.7	-140.2	-12.8	3.65	2.5	328	87%	286	735

1,100	136.0	5.9	1.98	14.78	-68.6	-169.1	-14.7	3.64	2.5	376	87%	327	772

1,200	138.6	7.0	1.91	19.66	-88.3	-201.7	-16.9	3.61	2.8	431	87%	375	819

1,300	135.4	9.0	1.83	30.65	-130.9	-261.6	-20.8	3.56	3.4	532	87%	462	895

1,400	129.9	10.0	1.79	35.93	-151.4	-290.9	-22.5	3.55	3.6	577	87%	500	929

1,500	119.8	11.0	1.77	43.27	-178.5	-319.4	-24.3	3.52	3.9	623	87%	540	967

1,600	110.5	11.6	1.75	48.58	-197.9	-338.6	-25.5	3.51	4.2	652	87%	566	993

1,700	94.8	12.4	1.73	56.27	-225.5	-362.4	-26.9	3.49	4.5	689	87%	597	1,030

1,800	83.1	12.9	1.71	61.06	-242.7	-376.9	-27.7	3.48	4.7	709	87%	615	1,053

1,900	50.3	14.0	1.69	75.31	-292.9	-410.9	-29.7	3.46	5.4	761	87%	659	1,112

2,000	22.3	14.9	1.67	86.48	-331.9	-436.4	-31.1	3.44	5.8	799	87%	691	1,156

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Table 15-13 Aerodrome Gold Price Sensitivities



Pit Size ($/oz Au)	Cash Flow ($M)	Mineralised Material (Mt)	Grade (g/t Au)	Waste (Mt)	Mining Cost ($M)	Process Cost ($M)	Royalties ($M)	Mining Cost ($//t)	Stripping Ratio (t:t)	Ounces Mined (koz Au)	Recovery (%)	Gold Produced (koz Au)	Cash Cost ($/oz Au)

700	0.6	0.0	1.9	0.0	-0.1	-0.4	-0.4	3.6	2.0	0.8	90.0	0.8	713.5

800	1.1	0.0	1.7	0.1	-0.3	-0.7	-0.7	3.5	2.1	1.7	90.0	1.5	771.6

900	1.7	0.1	1.6	0.1	-0.5	-1.4	-1.2	3.6	2.0	2.8	90.0	2.5	841.5

1,000	2.4	0.1	1.5	0.2	-0.9	-2.3	-1.8	3.5	2.4	4.3	89.8	3.9	905.8

1,100	4.5	0.2	1.5	0.8	-2.8	-5.4	-4.2	3.4	3.7	10.0	89.0	8.9	1,016.7

1,200	5.4	0.3	1.5	1.1	-4.0	-7.2	-5.5	3.3	4.0	13.1	88.6	11.6	1,064.8

1,300	6.0	0.4	1.5	1.6	-5.7	-9.5	-7.0	3.3	4.5	16.9	88.1	14.9	1,123.3

1,400	6.3	0.4	1.4	2.0	-7.0	-11.3	-8.1	3.3	4.7	19.6	87.9	17.2	1,163.3

1,500	6.4	0.5	1.4	2.4	-8.5	-13.3	-9.3	3.3	4.9	22.5	87.7	19.7	1,207.8

1,600	5.7	0.8	1.4	5.0	-16.9	-22.2	-14.7	3.2	6.2	36.0	87.1	31.3	1,350.8

1,700	4.5	1.1	1.3	7.0	-23.2	-29.4	-18.8	3.2	6.6	45.9	86.9	39.9	1,421.6

1,800	2.9	1.3	1.3	8.9	-29.3	-35.4	-22.2	3.2	7.0	54.4	86.8	47.3	1,474.2

1,900	2.6	1.3	1.3	9.1	-29.9	-36.1	-22.6	3.2	7.0	55.3	86.8	48.0	1,479.5

2,000	2.2	1.3	1.3	9.4	-30.9	-36.8	-23.0	3.1	7.1	56.3	86.8	48.8	1,489.3

Notes:

Sensitivities not provided below $700/oz pit size because this represents the lower limit of economic pit shells within the Aerodrome deposit

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Table 15-14 Oere Gold Price Sensitivities



Pit Size ($/oz Au)	Cash Flow ($M)	Mineralised Material (Mt)	Grade (g/t Au)	Waste (Mt)	Mining Cost ($M)	Process Cost ($M)	Royalties ($M)	Mining Cost ($//t)	Stripping Ratio (t:t)	Ounces Mined (koz Au)	Recovery (%)	Gold Produced (koz Au)	Cash Cost ($/oz Au)

700	2.9	0.1	2.5	0.3	-1.1	-2.4	-0.3	4.2	3.6	6.5	85.7	5.6	667.3

800	3.6	0.1	2.4	0.4	-1.5	-3.3	-0.3	4.1	3.7	8.6	85.8	7.4	702.1

900	3.9	0.1	2.3	0.5	-1.7	-4.0	-0.4	4.2	3.4	9.7	86.0	8.4	722.5

1,000	8.5	0.6	2.0	2.9	-10.0	-17.0	-1.4	3.8	5.2	36.3	85.7	31.1	915.8

1,100	12.9	1.4	1.9	8.3	-27.7	-43.9	-3.3	3.7	5.7	86.4	85.6	73.9	1,013.8

1,200	14.1	2.0	1.8	11.7	-38.9	-61.6	-4.5	3.7	5.9	117.3	85.4	100.2	1,048.4

1,300	7.8	3.9	2.0	37.1	-116.3	-124.3	-9.8	3.4	9.4	255.0	85.2	217.3	1,152.6

1,400	6.6	4.1	2.0	38.9	-121.7	-129.6	-10.1	3.4	9.5	264.6	85.2	225.5	1,159.4

1,500	2.8	4.4	2.0	42.5	-132.7	-138.0	-10.8	3.4	9.7	280.7	85.2	239.2	1,176.9

1,600	0.4	4.5	2.0	43.9	-137.2	-143.6	-11.1	3.4	9.7	288.6	85.2	245.9	1,186.9

1,700	-3.9	4.7	2.0	46.6	-145.4	-149.4	-11.4	3.4	9.9	298.6	85.2	254.4	1,203.9

1,800	-11.6	5.0	1.9	50.8	-158.1	-157.9	-12.0	3.3	10.2	312.5	85.2	266.2	1,232.1

1,900	-16.5	5.1	1.9	53.0	-164.8	-163.1	-12.3	3.3	10.3	319.6	85.2	272.3	1,249.2

2,000	-61.1	6.5	1.8	72.5	-223.8	-208.2	-14.6	3.3	11.1	380.9	85.2	324.4	1,376.7

Notes:

Sensitivities not provided below $700/oz pit size because this represents the lower limit of economic pit shells within the Oere deposit

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Pit Selection

All Mineral Reserve pits were designed on a $1,200/oz Au pit shell following an analysis of the pit size against value and gold price. The exceptions are the Sessenge and Oere pits, where the reserve pit designs were based on a $1,300/oz Au optimised pit shell, and Aerodrome where the reserve pit designs were based on a $1,500/oz Au optimised pit shell.

All Mineral Reserves, including Aerodrome, Sessenge, and Oere are profitable at a $1,200/oz Au sales price, and thus the Mineral Reserve and supporting cash flow statements are reported at $1,200/oz Au. This is in line with Barrick corporate guidelines, which considers long-term gold price forecasts.

The analysis, prior to the final pit selections, considered the mining of larger pits at higher gold prices and associated risks. This is mainly driven by ounces, changes in strip ratio, life of the pit, and the value of the pits at different metal prices.

Figure 15-3 shows the pit size for the $1,000/oz Au to $2,000/oz Au price pits and the potential net cash flow generated for each price scenario for the Aerodrome deposit. This analysis demonstrates that choosing the $1,500/oz Au pit shell at a sales price of $1,200/oz Au, provides potential for additional production within the current high gold price environment, whilst generating a Mineral Reserve net cash flow of $0.73 million at a $1,200/oz sales price. This demonstrates profitability of the selected $1,500/oz Au pit shell at a $1,200/oz Au sales price, and thus can be declared as a Mineral Reserve at a $1,200/oz Au declaration price. Although Aerodrome is a low-grade deposit, an additional reason that the pit is mined at a higher pit size is because the CTSF phase two lift is planned for early 2022 and geochemical characterisation of the waste rock has determined that it is appropriate for use within the tailings wall buttressing. The life span of the Aerodrome pit is less than one year (planned to be exhausted by July 2022).

Figure 15-4 shows the pit size for the $1,000/oz Au to $2,000/oz Au price pits and the potential net cash flow generated for each price scenario for the Sessenge deposit. This analysis demonstrates that choosing the $1,300/oz Au pit shell at a sales price of $1,200/oz Au, provides potential for additional production within the current high gold price environment, whilst generating a Mineral Reserve net cash flow of $10.78 million at a $1,200/oz sales price. This demonstrates profitability of the selected $1,300/oz Au pit shell at a $1,200/oz Au sales price, and thus can be declared as a Mineral Reserve at $1,200/oz Au declaration price.

Figure 15-5 shows the pit size for the $1,000/oz Au to $2,000/oz Au price pits and the potential net cash flow generated for each price scenario for the Oere deposit. This analysis demonstrates that choosing the $1,300/oz Au pit shell at a sales price of $1,200/oz Au, provides potential for additional production within the current high gold price environment, whilst generating a Mineral Reserve net cash flow of $8.35 million at a $1,200/oz sales price. This demonstrates profitability of the selected $1,300/oz Au pit shell at a $1,200/oz Au sales price, and thus can be declared as a Mineral Reserve at a $1,200/oz Au declaration price.

This same approach was applied to the selection of all other Mineral Reserve pits, all of which have a $1,200/oz Au selected pit shell with positive cash flow at a $1,200/oz sales price.

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Figure 15-3 Aerodrome Pit Size versus Cash Flow Curve at Different Gold Prices

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Figure 15-4 Sessenge Pit Size versus Cash Flow Curve at Different Gold Prices

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Figure 15-5 Oere Pit Size versus Cash Flow Curve at Different Gold Prices

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15.5 Underground Stope Shapes

Economic assessment is performed on a short term and long-term basis. The short-term economic assessment involves the evaluation of individual stope economics considering the direct costs (i.e., rehabilitation, production, paste filling) required to mine specific stopes. Long term economic assessment consists of assessing the economics of a mining area by considering the capital, development, and operating costs of mining that area.

For the conversion of underground resources to underground reserves, dilution and mining loss are applied. The parameters used for creating MSO shapes are summarised in Table 15-15.

Table 15-15 MSO Parameters



MSO Parameters		Value

Slice Interval		0.5

Minimum Mining Width		5 m

Cut-off grade		2.02 g/t Au

Footwall Minimum Dip		45°

Hanging wall Minimum Dip		45°

Maximum Stope Thickness Ratio		4

Near Dilution		0

Far Dilution		0

Section and Level Intervals		Variable based on the mining lode

Sections (U)		Variable based on mining method and mining lode

The confidence categories of the Mineral Reserves are assigned as per CIM (2014) Standards. On a proportional basis, Mineral Resources that are classified as Measured or Indicated are converted to Proven and Probable Mineral Reserves. Inferred Mineral Resources are excluded and not classified as Mineral Reserves.

The following formulae are used for proportionally converting Mineral Resources into Mineral Reserves:

●

Proven Mineral Reserves = (Measured Material + % Measured Material x Waste Material) x Recovery x Dilution

●

Probable Mineral Reserves = (Indicated Material + % Indicated Material x Waste Material) x Recovery x Dilution

The location of the Proven and Probable Mineral Reserves are shown in Figure 15-6.

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Source: Kibali Goldmines, 2021

Figure 15-6 Kibali Underground Mineral Reserve Classification (Looking NW)

15.6 Reconciliation

Kibali Goldmines has a standard weekly, end of month (EOM), and end of quarter production measurement system that reports and provides reconciliation between grade control and monthly mine production.

The measurement system tracks daily, weekly, monthly, quarterly, and year to date production grade control results versus the plant. The system tracks both underground and open pit domain production against the block model. Summary reports are prepared weekly, monthly, and quarterly.

The reconciliation between the mine call and plant check-out was good during 2021.Some local issues were investigated and fixed within the year, but overall, these have had a negligible impact on the overall cumulative mine to mill reconciliation.

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Table 15-16 shows the EOY Mine Call Factor (MCF) reconciliation.

Table 15-16 Kibali 2021 EOY MCF Reconciliation



Dept	Recon Ore Mined, Stockpiles and Plant Check Out	Year End 2021
Dept		Tonnes (t)	Grade (g/t Au)	Ounces (oz Au)

GC	Mine	6,434,586	4.34	896,966

GC	Stockpile Change	-1,137,969	1.03	-37,707

GC	GC Actual Feed	7,852,647	3.68	929,565

Plant	Cone Change	1,779	3.70	212

Plant	Scat Stock Change	2,393	3.61	278

GC	Mine Call	7,848,475	3.68	929,076

Plant	Plant Check Out	7,783,337	3.61	902,613

GC vs Plant	MCF (%) GC Call vs Plant Check Out	99	98	97

Figure 15-7 presents a chart of mine production with weekly feed source ratio versus pulp call versus gold after smelting. As shown in Figure 15-9, there was a loss in grade observed in Q1 2021, which relates to underperformance of the Gorumbwa fresh medium grade stockpile. This was mined with poorly defined polygons, resulting in mineralisation projected through an actual barren zone.

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Figure 15-7 2021 Kibali Production with Weekly Feed Source Ratio versus Pulp Call versus Gold After

Smelting

As shown in Figure 15-8, there was good reconciliation between the Mine call and plant check-out grade, averaging 98% MCF for the entire year.

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Figure 15-8 2021 Weekly Grades Comparison (Mine Call Grade vs Plant Check Out Grade vs Carbon

Loading)

As shown in Figure 15-9, there was a good reconciliation between Mine call and plant check-out tonnes, averaging 99% MCF for the entire year.

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Figure 15-9 2021 Weekly Tonnage Comparison (Mine Call Tonnes vs Plant Check Out Tonnes)

Table 15-17 shows the EOY Resource Call Factor (RCF) reconciliation, comparing 3D volumes mined from the Mineral Resource model against production data from the plant. The 2020 budget models performed well in 2021, with a negligible difference to the MCF reconciliation providing reasonably robust prediction of production.

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Table 15-17 Kibali EOY 2021 Resource Call Factor Reconciliation



Parameter	2020 Resource Model
Parameter	Tonnes (t)	Grade (g/t Au)	Grams (g)	Ounces (oz Au)

Depletion Kibali OP models	2,956,702	2.72	8,054,405	258,955

Diluted (10%)	3,252,373	2.48	8,054,405	258,955

Ore Loss (3%)	97,571	2.48	241,632	7,769

Depletion Kibali OP diluted	3,154,801	2.48	7,812,773	251,186

Depletion Kibali UG model	3,296,296	5.74	18,912,411	608,048



Total Ore Source	6,451,097	4.14	26,725,184	859,235



Stockpile Change	-1,137,969	1.03	-1,172,814	-37,707

Plant cone Change	1,779	3.70	6,585	212

Scats	2,393	3.61	8,642	278

Tails	7,783,337	0.37	2,877,284	92,507

Gold produced	-	-	25,260,737	812,152

Change in Au lockup	-	-	-77,123	-2,480

Unaccounted loss/gain	-	-	461,538	14,839



Total	6,649,540	4.12	27,364,850	879,800



RCF	97%	100%	98%	98%

15.7 Mineral Reserves Statement

Open Pit

The Kibali open pit Mineral Reserve estimate (100% basis) as of 31 December 2021 is presented in Table 15-18.

The following factors are updated in the 31December 2021 open pit Mineral Reserve update compared to the previous estimate:

●

Depletion of Mineral Reserves with the open pit mined shape.

●

Infill grade control drilling resulting in resource model changes.

●

Economic changes to G&A costs.

●

A higher gold price used for KCD.

●

Depletion of stockpiles.

The net change between the 2020 Mineral Reserve estimate and 2021 Mineral Reserve estimate has been an increase of approximately 0.19 Moz Au (+2%)

The Mineral Reserve changes are summarised in Table 15-19.

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Table 15-18 Kibali Open Pit Mineral Reserves as of 31December 2021



Category	Tonnes (Mt)	Grade (g/t Au)	Ounces (Moz Au)

KCD

Proven	0.59	2.53	0.048

Probable	1.2	2.38	0.091

Total	1.8	2.43	0.14

Megi-Marakeke-Sayi

Proven	-	-	-

Probable	7.3	1.83	0.43

Total	7.3	1.83	0.43

Kalimva-Ikamva

Proven	0.36	3.02	0.035

Probable	4.8	3.26	0.51

Total	5.2	3.25	0.54

Gorumbwa

Proven	1.3	2.28	0.097

Probable	4.1	3.23	0.43

Total	5.5	3.00	0.53

Pakaka

Proven	2.3	3.34	0.25

Probable	2.4	2.73	0.21

Total	4.7	3.02	0.46

Oere

Proven	-	-	-

Probable	2.2	2.35	0.17

Total	2.2	2.35	0.17

Pamao

Proven	5.1	1.77	0.29

Probable	3.4	2.15	0.23

Total	8.4	1.92	0.52

Aerodrome

Proven	0.49	1.37	0.022

Probable	0.034	1.15	0.0013

Total	0.53	1.36	0.023

Sessenge

Proven	0.61	2.56	0.050

Probable	0.63	1.79	0.036

Total	1.2	2.17	0.087

Stockpiles

Proven	0.32	3.17	0.032

Probable	-	-	-

Total	0.32	3.17	0.032

Open Pits and Stockpile

Proven	11	2.31	0.82

Probable	26	2.51	2.1

Total	37	2.45	2.9

Notes

Mineral Reserves are reported on a 100% basis.

The Mineral Reserve estimate has been prepared according to CIM (2014) Standards and using CIM (2019) MRMR Best Practice Guidelines.

Open pit Mineral Reserves are reported at a gold price of $1,200/oz Au, and an overall weighted average cut-off grade of 0.96 g/t Au, including dilution and ore loss factors.

Open pit Mineral Reserves were estimated by Mr. Shaun Gillespie, Reg Eng Tech, FAusIMM, an officer of the company and QP, and reviewed by Simon Bottoms CGeol, MGeol, FAusIMM, an officer of the company and QP.

Numbers may not add due to rounding. Tonnes and contained gold are rounded to 2 significant figures. All Proven and Probable grades are reported to 2 decimal places.

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Table 15-19 Open Pit Mineral Reserve Comparison to Previous Estimate



Change		Tonnes (Mt)	Grade (g/t Au)	Ounces (Moz Au)	Comments

	2020 Mineral Reserve Estimate	31.69	2.47	2.52	2020 Declared Reserves

Mining Depletion	Open Pit	-2.64	2.64	-0.22	2021 Depletion

Model Changes/Additions	Open Pit	2.23	2.24	0.16	Model changes from Infill GC

Cut-off Grade Changes	Open Pit	-0.06	0.22	-0.00	Changed G&A and Process cost

Gold Price Changes	Open Pit	4.18	2.38	0.32	Increase in gold Price for Gorumbwa and Pakaka

New Deposit	Open Pit	2.74	2.16	0.19	Conversion of Aerodrome and Oere deposits to Reserves

Design Change	Open Pit	0.07	1.83	0.00	Sessenge slope update

Stockpile Changes	Open Pit	-1.14	1.03	-0.04	Depletion of stockpiles by Dec 2020

	2021 Theoretical	37.16	2.46	2.933

	2021 OP Mineral Reserve Estimate	37.26	2.45	2.932	2021 Declared Reserves

Underground

The Kibali underground Mineral Reserves (100% basis) as of 31 December 2021 are shown in Table 15-20.

Table 15-20 Kibali Underground Mineral Reserves as of 31 December 2021



Kibali (KCD) – Underground	Tonnes (Mt)	Grade (g/t Au)	Ounces (Moz Au)

Proven	21	4.54	3.0

Probable	25	4.54	3.7

Total	46	4.54	6.7

Notes:

Mineral Reserves are reported on a 100% basis.

The Mineral Reserve estimate has been prepared according to CIM (2014) Standards and using CIM (2019) MRMR Best Practice Guidelines.

Underground Mineral Reserves are reported at a gold price of $1,200/oz Au and a cut-off grade of 2.02 g/t Au

Underground Mineral Reserves were estimated by Ismail Traore, MSc, FAusIMM, M.B. Law, DES, an officer of the company and QP, and reviewed by Simon Bottoms CGeol, MGeol, FAusIMM, an officer of the company and QP.

Numbers may not add due to rounding. Tonnes and contained gold are rounded to 2 significant figures. All Proven and Probable grades are reported to 2 decimal places.

The Kibali underground Mineral Reserves (100%) are shown by mining zone in Table 15-21. Table 15-22 shows a comparison of the current Mineral Reserve to the previous Mineral Reserve estimate in 2020.

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Table 15-21 Kibali Mineral Reserves by Mining Zone as of 31 December 2021



Zone	Tonnes (Mt)	Grade (g/t Au)	Ounces (Moz Au)

Proven

3101	0.0043	8.53	0.0012

3102	0.0087	5.75	0.0016

5101	5.2	5.09	0.85

5102	6.5	4.47	0.93

5105	0.32	3.38	0.035

5110	0.11	2.95	0.010

9101	3.3	4.60	0.49

9105	5.4	4.15	0.71

Total Proven	21	4.54	3.0

Probable

3101	7.6	3.98	0.97

3102	0.90	3.02	0.087

5101	1.3	4.65	0.20

5102	2.2	5.10	0.37

5104	0.51	5.21	0.086

5105	0.43	3.77	0.052

5110	0.0037	1.39	0.00017

9101	10	5.12	1.7

9105	1.9	3.75	0.23

Total Probable	25	4.54	3.7

Total UG Proven and Probable Reserves	46	4.54	6.7

Notes:

Mineral Reserves are reported on a 100% basis.

The Mineral Reserve estimate has been prepared according to CIM (2014) Standards and using CIM (2019) MRMR Best Practice Guidelines.

Underground Mineral Reserves are reported at a gold price of $1,200/oz Au and a cut-off grade of 2.02 g/t Au

Underground Mineral Reserves were estimated by Mr. Ismail Traore, MSc, FAusIMM, an officer of the company and QP, and reviewed by Simon Bottoms CGeol, MGeol, FAusIMM, an officer of the company and QP.

Numbers may not add due to rounding. Tonnes and contained gold are rounded to 2 significant figures. All Proven and Probable grades are reported to 2 decimal places.

Table 15-22 Underground Mineral Reserve Comparison to Previous Estimate



Change	Tonnes (Mt)	Grade (g/t Au)	Ounces (Moz Au)

December 2020 Mineral Reserve Estimate	44.76	4.81	6.92



Depletion of 2020 Reserve	-3.59	5.63	-0.65

3000 Lode down plunge – Mineral Reserve Addition	1.74	5.40	0.30

9000 Lode – Model change with additional drilling	0.59	-2.79	-0.05

5000 Lode Mineral Reserve addition & redesign below 245L	1.02	3.16	0.10

Change in cut-off grade and modifying factors	1.31	1.61	0.68





December 2021 Mineral Reserves Estimate	45.84	4.54	6.69

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Stockpiles

Details of surface stockpiles of ore sourced from open pits and underground are presented in Table 15-23 for reference purposes.

Table 15-23 Kibali Surface Stockpile Mineral Reserve as of 31 December 2021



Location	Actual EoY Surface Stockpiles
Location	Tonnes (kt)	Grade (g/t Au)	Ounces (koz Au)

FGO (Full Grade Ore)	0.23	3.96	0.030

MO (Marginal Ore)	0.086	1.06	0.0029

Total Ore SP excl. Scats	0.32	3.17	0.032

Notes:

Stockpiles are reported on a 100% basis.

The Mineral Reserve estimate has been prepared according to CIM (2014) Standards and using CIM (2019) MRMR Best Practice Guidelines.

Stockpile Mineral Reserves were estimated by Mr. Shaun Gillespie, Reg Eng Tech, FAusIMM, an officer of the company and QP, and reviewed by Simon Bottoms CGeol, MGeol, FAusIMM, an officer of the company and QP.

Numbers may not add due to rounding. Tonnes and contained gold are rounded to 2 significant figures. All Proven and Probable grades are reported to 2 decimal places. Tonnes and contained gold are rounded to 2 significant figures. All Proven and Probable grades are reported to 2 decimal places.

15.8

Discussion

External Reserve Audits

An independent audit was undertaken in 2017 on the Mineral Reserve estimate by Optiro, who concluded that the Mineral Reserve estimation processes used by Kibali Goldmines are considered, by Optiro, to be at a level commensurate with industry best practice (Optiro, 2017).

An independent technical review of the annual Mineral Resource and Mineral Reserve estimates for the Mine was carried out by the RSC during 2021, including a site visit by RSC QPs (RSC Ltd, 2021). The audit demonstrated that Mineral Resource and Mineral Reserve processes conform to good practices. However, RSC made a number of recommendations to Kibali Goldmines from a Mineral Reserve perspective including:

●

Investigate the cost/benefit of installing a proper, broken ore, falling stream sampler on the crushed mine feed belts to help analyse particular source material.

●

Investigate the implementation of a broader integrated reconciliation system (i.e., Snowden Reconciler or proprietary), as all current informing data sits in a range of Excel spreadsheets.

●

Degradation of paste fill has been allowed for in the 2021 reserves schedule set-up; however, it remains an item requiring monitoring, along with other allowances for dilution and ore losses.

●

Factors for open pit mining dilution and ore losses are standardised across the site and operations. These have been applied across all facets of mine planning including optimisation, cut-off grade calculations, and generation of physicals for scheduling. RSC considers that, with the established history of mining, individual mining loss and dilution factors can and should be applied for physicals generation for each mining area.

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The QP is not aware of any environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that could materially affect the Mineral Reserve estimate.

The QP is not aware of any mining, metallurgical, infrastructure, permitting, and other relevant factors that could materially affect the Mineral Reserve estimate.

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16	Mining Methods

16.1 Summary

The Mine comprises both open pit and underground mining operations. The general layout of the mine is shown in Figure 5-2 in Section 5.3.

16.2 Open Pit Mining

Open pit mining is carried out using conventional drill, blast, load, and haul surface mining methods. Mining of the main pits is carried out by a mining contractor, KMS.

From 2022 onwards, open pit production will come from the Sessenge, Aerodrome, Pamao, Gorumbwa, Megi-Marakeke-Sayi, Kalimva-Ikamva, Oere, Pakaka, and the KCD deposits. The Mengu Hill, Mofu, Kombokolo and Rhino pits were depleted in 2017.

Free digging in the upper levels uses 5 m high benches, with 10 m benches used for drilling and blasting operations. The 10 m benches containing ore are excavated in three flitches of equal height.

Historical ore production from the Kibali open pits, up to 2021, is detailed in Table 16-1.

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Table 16-1 Kibali Open Pits Historical Production



Source	2013	2014	2015	2016	2017	2018	2019	2020	2021	Total
Source	Tonnes (kt)

KCD	4,335	5,516	4,458	764	366	-	-	-	-	15,439

KCD PB3	-	-	-	-	-	2,082	1,688	1,561	1,154	6,484

Mofu	-	83	84	-	-	-	-	-	-	167

Mengu Hill	-	-	1,191	1,220	725	2	-	-	-	3,138

Kombokolo	-	-	-	278	686	1,572	69	-	-	2,605

Pakaka	-	-	-	2,350	3,386	230	-	-	-	5,965

Rhino	-	-	-	67	95	-	-	-	-	161

Gorumbwa	-	-	-	-	-	-	234	1,163	1,367	2,764

Sessenge	-	-	-	-	-	1,572	1,771	342	235	3,920

Aerodrome	-	-	-	-	-	-	-	-	86	86

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The estimated LOM production of ore and waste for the open pits schedule are as presented in Table 16-2 based on Mineral Reserves.

Table 16-2 Kibali Open Pits, Reserves Basis



Open Pit	Ore		Waste	Total
Open Pit	Tonnes (kt)	Grade (g/t Au)	Tonnes (kt)	Tonnes (kt)	Strip Ratio

KCD PB2 N	1,778	2.43	16,896	18,673	9.5

Sessenge	1,243	2.17	4,977	6,221	4.0

Pakaka	4,735	3.02	59,301	64,036	12.5

Megi-Marakeke-Sayi	7,343	1.83	28,093	35,436	3.8

Aerodrome	529	1.36	3,778	4,307	7.1

Pamao	8,434	1.92	48,821	57,255	5.8

Gorumbwa	5,468	3.00	58,729	64,197	10.7

Kalimva-Ikamva	7,416	2.98	98,578	105,995	13.3

Total	36,946	2.44	319,174	356,120	8.6

Figure 16-1 shows a longitudinal section of the pushbacks in the Pakaka pit. Figure 16-2 illustrates a longitudinal section of the pushbacks in the Gorumbwa pit.

LOGO

Source: Kibali Goldmines, 2021

Figure 16-1 Long Section of the Pakaka Pushbacks

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Source: Kibali Goldmines, 2021

Figure 16-2 Long Section of the Gorumbwa Pushbacks

Geotechnical

Void Management

At Gorumbwa, the historical underground void was safely mined out in Pushback 1 during 2021, from elevation 5830 mRL to 5760 mRL. The void management procedure, which was developed to manage mining around these areas where personnel and equipment are exposed to higher risk associated with instability from sub-surface excavations, was strictly followed. Figure 16-3 shows the location of the historical voids within the Gorumbwa pit.

LOGO

Source: Kibali Goldmines, 2021

Figure 16-3 Gorumbwa Pit (Looking East) Showing Historical Underground Void Mined Out in Pushback 1

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Slope Angles

No extra work was carried out for the KCD Pushback 3 pit. The pit still has three main domains which were generated based on the rock properties and the inter-ramp angles provided to accommodate the ramps in the final pit designs (Figure 16-4 and Table 16-3).

Pushback 3 was mined out in 2021 and only left with Pushback 3 North, which is planned to be mined from 2027.

LOGO

Source: Kibali Goldmines, 2021

Figure 16-4 KCD Pushback 3 Geotechnical Domains

Table 16-3 KCD Geotechnical Geometry



Domain	From	To	Bench Height (m)	Berm Width (m)	Batter Angle (°)	IRA¹ (°)	Design Consideration

CB1	Surface	5880	5	4	40	27	Weathered and Weathered Shale
CB1	5880	5810	10	5	65	48	Transition to Fresh

CB2	Surface	5880	5	4	40	27	Weathered Shale
CB2	5880	5810	5	4	40	27	Weathered Shale

CB3	Surface	5860	5	4	40	27	Weathered Shale
CB3	5860	5810	5	5	50	30	Weathered Material

Notes:

1 Inter ramp slope angle (IRA)

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Dempers & Seymour Pty Ltd (D&S) was commissioned by Kibali Goldmines to undertake the pit slope design for the Sessenge Pit. A 3D Mining Rock Mass Model (MRMM) was constructed based on geotechnical logging of drill core undertaken at Kibali (Table 16-4). Based on this mass model, five distinctive geotechnical domains were identified and defined for the Sessenge Pit.

Rigorous analyses, including rock bridge/structure failure criteria for each rock type per geotechnical domain, were completed and pit slope designs excluding haul ramps were recommended (Table 16-5). Slopes were considered as dry slopes and that the necessary dewatering would take place timely as scheduled.

Pushback one was mined out by March 2020. Pushback two, at a higher gold price, is currently being mined and will be completed by July 2022 following the same design parameters.

Table 16-4 Rock Mass Properties for the Sessenge Pit



Rock Unit	Rock Strength	Joint Condition	Fracture Frequency	RMR	MRMR

Weathered	1 Mpa – 5 Mpa	Smooth and Undulating with Soft Sheared Fine Infill	>40 frac/m	8 – 12	7 – 10
			Spacing <0.0.2 m	Average 11	Average 9

Transition	25 Mpa – 50 Mpa	Smooth and Undulating with No Infill (Clean)	7 frac/m	9 – 56	7 – 45
			Spacing 0.15 m	Average 39	Average 31

MCP	100 Mpa – 130 Mpa	Smooth and Undulating with No Infill (Clean)	1 frac/m	62 – 77	50 – 62
			Spacing 1.0 m	Average 69	Average 56

CHS	100 Mpa – 130 Mpa	Smooth and Undulating with No Infill (Clean)	1 frac/m	60 – 71	48 – 57
			Spacing 1.0 m	Average 65	Average 52

ORE 9001	100 Mpa – 130 Mpa	Rough and Undulating with No Infill (Clean)	1 frac/m	61 – 78	49 – 63
			Spacing 1.0 m	Average 66	Average 53

ORE 9003	100 Mpa – 130 Mpa	Rough and Undulating with No Infill (Clean)	1.5 frac/m	60 – 75	48 – 61
			Spacing 0.67 m	Average 66	Average 53

SCH	100 Mpa – 130 Mpa	Rough and Undulating with No Infill (Clean)	1.2 frac/m	60 – 70	48 – 57
			Spacing 0.8 m	Average 64	Average 52

BIF	100 Mpa – 130 Mpa	Rough and Undulating with No Infill (Clean)	0.4 frac/m	68 – 78	55 – 63
			Spacing 2.5 m	Average 73	Average 59

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Table 16-5 Sessenge Slope Design



Domain	Material	From (mRL)	To (mRL)	Batter Height (m)	Berm Width (m)	Batter Angle (°)	IRA (°)

West	Weathered	Surface	5840	5	4	50	50

West	Transition	5840	5830	10	5	55	55

West	Fresh	5830	5790	10	5	75	57

NW	Weathered	Surface	5830	5	4	50	36

NW	Transition	5830	5820	10	5	55	55

NW	Fresh	5820	5790	10	5	75	59

NE	Weathered	Surface	5850	5	4	50	35

NE	Transition	5850	5830	10	5	55	47

NE	Fresh	5830	5790	10	5	75	57

SW	Weathered	Surface	5820	5	4	50	35

SW	Transition	5820	5800	10	5	55	47

SW	Fresh	5800	5770	10	5	75	59

SE	Weathered	Surface	5860	5	4	50	37

SE	Transition	5860	5840	10	5	55	47

SE	Fresh	5840	5770	10	5	75	55

There were no significant geotechnical works carried for the Gorumbwa pit other than the void management as stated earlier. Although the pit changed from a $1,000/oz Au size to a $1,200/oz Au size, the geotechnical parameters where tested revealed that the same parameters were still relevant.

Four distinctive geotechnical domains were identified and defined for the Gorumbwa Pit (Figure 16-5). The same slope angles were then applied for the various domains (Table 16-6).

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Source: Kibali Goldmines, 2021

Figure 16-5 Gorumbwa Pit Geotechnical Domains

Table 16-6 Gorumbwa Recommended Pit Slope Configuration



Domain	Design Section	From (mRL)	To (mRL)	Batter Height (m)	Berm Width (m)	Batter Angle (°)	IRA (°)

NE	NE1	Surface	5830	5	5	50	32

NE	NE1	5830	5800	10	4	55

NE	NE1	5800	5750	10	4	65

NE	NE1	5750	5650	10	4	70	50

South	S1	Surface	5820	5	5	50	32

South	S1	5820	5790	10	5	55

South	S1	5790	5750	10	4	60

South	S1	5750	5670	10	4	65

South	S1	5670	5650	10	4	70	48

NW	NW1	Surface	5820	5	5	50	32

NW	NW1	5820	5790	10	4	55

NW	NW1	5790	5750	10	4	60

NW	NW1	5750	5670	10	4	65

NW	NW1	5670	5650	10	4	70	48

Footwall	-	Surface	5820	5	5	50	32

Footwall	-	5820	5650	10	6	55	38

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For the Pakaka pit, three dominant geotechnical domains were originally identified (Figure 16-6). Appropriate slope angles were then designed for the various domains, as presented in Table 16-7.

The Pakaka pit changed from a $1,000/oz Au size in 2020 to $1,200/oz Au size in 2021, and thus a review of the MRMM model was completed using a total of 550 m diamond drilling in the NE extension. The top of the fresh extension was confirmed, as well as the continuation of the MRMM (Figure 16-7). Therefore, the three dominant domains were maintained.

LOGO

Source: Kibali Goldmines, 2021

Figure 16-6 Pakaka Pit Geotechnical Domains

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Table 16-7 Pakaka Recommended Pit Slope Configuration



Domain	Material	From (mRL)	To (mRL)	Batter Height (m)	Batter Angle (°)	Berm Width (m)	IRA¹ (°)

North Domain	Weathered	Surface	5820	5	50	4	33

North Domain	Fresh	5820	5810	10	50	6

North Domain	Fresh	5810	5770	20	60	6

North Domain	Fresh	5770	5730	20	65	6

North Domain	Fresh	5730	5710	20	70	6	52

South Domain	Weathered	Surface	5840	5	50	4	33

South Domain	Fresh	5840	5830	10	50	6

South Domain	Fresh	5830	5810	20	55	6

South Domain	Fresh	5810	5770	20	60	6

South Domain	Fresh	5770	5750	20	65	6	50

Footwall Domain	Weathered	Surface	5830	5	50		33

Footwall Domain	Fresh	5830	5750	20	55

Note:

Maximum depth 150 m.

LOGO

Source: Kibali Goldmines, 2021

Figure 16-7 Cross Section Showing Pakaka PB2 Pit Project MRMM Reviewed from the Old Pit at $1,000/oz

toward the $1,200/oz Au Pit Shells

The Pamao pit was evaluated in two stages. The first stage comprised the main Pamao Pit and the second the Pamao South Pit (Figure 16-8). The main pit had four dominant domains identified and generated, while Pamao South had only three. Appropriate slope angles were then designed for the various domains as presented in Table 16-8 and Table 16-9.

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LOGO

Source: Kibali Goldmines, 202,

Figure 16-8 Pamao Complex Geotechnical Domains

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Table 16-8 Pamao Recommended Pit Slope Configuration



Domain	Material	From (mRL)	To (mRL)	Bench Height (m)	Berm Width (m)	Batter Angles (°)	IRA (°)

FW	Saprolite/Oxide	Surface	5840	5	5	45	38
FW	Fresh	5840	5720	10	5	80	55

West	Saprolite/Oxide	Surface	5800	5	5	50	34
West	Fresh	5800	Bottom	10	5	80	55

HW	Saprolite/Oxide	Surface	5840	5	5	55	42
HW	Fresh	5840	Bottom	10	5	80	56

East	Saprolite/Oxide	Surface	5810	5	5	50	36
East	Fresh	5810	Bottom	10	5	80	55

Table 16-9 Pamao South Recommended Pit Slope Configuration



Domain	Material	From (mRL)	To (mRL)	Bench Height (m)	Berm Width (m)	Batter Angles (°)	IRA (°)

FW	Saprolite/Oxide	Surface	5850	5	4	55	34
	Fresh	5850	5815	5	4	55	34
	Fresh	5815	Bottom	10	4	70	53

HW	Saprolite/Oxide	Surface	5845	5	4	55	34
	Fresh	5845	5775	10	4	70	53
	Fresh	5575	Bottom	10	4	75	56

NE	Saprolite/Oxide	Surface	5845	5	4	55	34
	Fresh	5845	5575	10	4	70	53
	Fresh	5575	Bottom	10	4	75	56

For the Oere pit, two dominant domains were defined (Figure 16-9) and appropriate slope angles were then designed (Table 16-10).

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LOGO

Source: Kibali Goldmines, 2021

Note:

Legend = Rock Mass Rating (RMR)

Figure 16-9 Oere Pit Geotechnical Domains

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Table 16-10 Oere Slope Design



Domain	From (mRL)	To (mRL)	Bench Height (m)	Berm Width (m)	Batter Angle (°)	IRA (°)

FW	Surface	5840	5	4	55	34
	5840	5820	10	4	60	46
	5820	5780	10	4	70	53
	5780	Bottom	10	4	75	56

HW	Surface	5820	5	4	55	34
	5820	5780	10	4	70	53
	5780	Bottom	10	4	75	56

Hydrogeology

The geology of Kibali is dominated by highly deformed metasedimentary and metavolcanic rocks. The most notable units are metaconglomerates, carbonaceous shales and BIF, which are intruded by later dolerite dykes. The rocks are highly deformed and many show significant signs of alteration by albite, carbonate, silica and sericite. Alteration changes the fabric of the rock mass and may therefore cause significant changes to its hydraulic conductivity. Silicic alteration in a rock mass has the tendency to sustain open fracturing, while fractures tend to heal due to sertictic alteration, reducing their potential to transmit groundwater.

Apart from alteration, the pores in the fresh bedrocks are too well cemented to allow any significant permeation of water. All groundwater movement occurs within fractures and joints, created by two major deformation events (D1 and D2). The D2 event is interpreted to have formed the main structural corridor for gold mineralisation.

Interpretations from a historic Packer test completed in 2012 at KCD, indicate a geometric mean hydraulic conductivity of 2.38E-08 m/s. Conductivities of up to 1e-5 m/s were interpreted, indicating the presence of high yielding structures, even at greater depths. From a Packer test done in 2018, the conductivities of up to 4.19E-5 m/s were interpreted.

The pit dewatering strategy was reviewed and is now primarily based on using in-pit sump pumping to manage ingress on the mining floor. Perimeter dewatering boreholes were shown to have an insignificant impact on the in-pit groundwater and there was no possibility of implementing long life in-pit boreholes due to excess vibration generated by the blasting method used. Temporary in-pit sumps of 3 m depth across a bench are excavated to support operations during the rainy season.

The maximum daily rainfall record for Kibali over the last fifteen years ranges from 60 mm to 110 mm in 24 hours. The months likely to register the maximum daily rainfall are May and September (four counts), July (six counts), August and October (five counts).

For water management planning, 1:100 year rainfall events derived from on-site data are used across all Mineral Reserve open pit designs.

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Pit Dewatering

The dewatering plans and pumping station capacity for the two active pits Gorumbwa and KCD are summarised below in (Figure 16-10, Figure 16-11, and Figure 16-12). These plans provide full depressurisation programme and pumping of all water such that all rainwater would be pumped out within less than 30 days, even in a major 1:100 year storm event.)

Table 16-11 summarises the 2021 dewatering operations for the Gorumbwa South and Table 16-12 for Gorumbwa North.

Figure 16-10 and Figure 16-11 present the Gorumbwa South and Gorumbwa North dewatering design criteria, respectively.

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Table 16-11 Gorumbwa South Pit (5,760 mRL) 2021 Dewatering Operations Summary



Month	Rain (m)	Run-off Co-efficient	Pit Area (m²)	In-Pit Rain Inflow Vol (m³)	Ground Water Inflow	Dynamic Head (Head+ Loss)	Diesel Pump Capacity (m³/h)	Running Hours	Diesel Consum. (50L/hr)	Diesel Cost ($1.20/L)	Operating Cost ($7.16/hr)	Pump Rental Cost ($9.50/hr)	Total Cost ($)

Jan	0.036	0.68	139,000	3,403	2,678	85	250	24	1,216	1,459	174	231	1,865

Feb	0.045	0.68	139,000	4,253	2,419	85	250	27	1,334	1,601	191	254	2,046

Mar	0.142	0.68	139,000	13,422	2,678	85	250	64	3,220	3,864	461	612	4,937

Apr	0.226	0.68	139,000	21,362	2,592	85	250	96	4,791	5,749	686	910	7,345

May	0.222	0.68	139,000	20,983	2,678	85	250	95	4,732	5,679	678	899	7,256

Jun	0.185	0.68	139,000	17,486	2,592	85	250	80	4,016	4,819	575	763	6,157

Jull	0.219	0.68	139,000	20,700	2,678	85	250	94	4,676	5,611	670	888	7,169

Aug	0.212	0.68	139,000	20,038	2,678	85	250	91	4,543	5,452	651	863	6,966

Sep	0.233	0.68	139,000	22,023	2,592	85	250	98	4,923	5,908	705	935	7,548

Oct	0.283	0.68	139,000	26,749	2,678	85	250	118	5,885	7,063	843	1,118	9,024

Nov	0.147	0.68	139,000	13,894	2,592	85	250	66	3,297	3,957	472	626	5,055

Dec	0.039	0.68	139,000	3,686	2,678	85	250	25	1,273	1,527	182	242	1,952

Total				188,000	31,533			878	43,907	52,688	6,287	8,342	67,318

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Table 16-12 Gorumbwa North Pit (5,745 mRL) 2021 Dewatering Operations Summary



Month	Rain (m)	Run-off Co-efficient	Pit Area (m²)	In-Pit Rain Inflow Vol (m³)	Ground Water Inflow	Dynamic Head (Head+ Loss)	Diesel Pump Capacity (m³/h)	Running Hours	Diesel Consum. (50L/hr)	Diesel Cost ($1.20/L)	Operating Cost ($7.16/hr)	Pump Rental Cost ($9.50/hr)	Total Cost ($)

Jan	0.036	0.68	495,270	12,124	37,498	85	250	198	5,955	7,146	1,421	1,886	10,452

Feb	0.045	0.68	495,270	15,155	33,869	85	250	196	9,805	11,766	1,404	1,863	15,033

Mar	0.142	0.68	495,270	47,823	37,498	85	250	341	17,064	20,477	2,444	3,242	26,163

Apr	0.226	0.68	495,270	76,113	36,288	85	250	450	22,480	26,976	3,219	4,271	34,467

May	0.222	0.68	495,270	74,766	37,498	85	250	449	22,453	26,943	3,215	4,266	34,424

Jun	0.185	0.68	495,270	62,305	36,288	85	250	394	19,719	23,662	2,824	3,747	30,233

Jull	0.219	0.68	495,270	73,756	37,498	85	250	445	22,251	26,701	3,186	4,228	34,115

Aug	0.212	0.68	495,270	71,398	37,498	85	250	436	21,779	26,135	3,119	4,138	33,392

Sep	0.233	0.68	495,270	78,471	36,288	85	250	459	22,952	27,542	3,287	4,361	35,190

Oct	0.283	0.68	495,270	95,310	37,498	85	250	531	26,561	31,874	3,804	5,047	40,724

Nov	0.147	0.68	495,270	49,507	36,288	85	250	343	17,159	20,591	2,457	3,260	26,308

Dec	0.039	0.68	495,270	13,135	37,498	85	250	203	10,126	12,152	1,450	1,924	15,526

Total				669,863	441,504			4,445	218,304	261,964	31,830	42,232	336,026

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Source: Kibali Goldmines, 2021

Figure 16-10 Gorumbwa South Pit Dewatering Design Criteria

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Source: Kibali Goldmines, 2021

Figure 16-11 Gorumbwa North Pit Dewatering Design Criteria

Mining of the Gorumbwa-Sessenge-gap pit is complete. No pumping is required until the second quarter of 2022 when the water storage raises to the level 5,785mRL and spills over into active Pushback 1 for a volume of 155,000m³.

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The dewatering capacity at KCD was designed at 12,000 m³/day to be discharged at the bottom of the old pit, to provide mitigation to the risk of handling a significant volume in a short period of time. The pumping station at the bottom of KCD Pushback two is now running continuously to keep the pit bottom dry and reduce the risk of flooding the underground workings. Figure 16-12 presents the pit dewatering design criteria for KCD Pushback 3. Table 16-13 presents a summary of the KCD Pushback 3 Pit 2021 dewatering operations.

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Source: Kibali Goldmines, 2021

Figure 16-12 KCD Pushback 3 Pit Dewatering Design Criteria

Ahead of mining at Pamao, a total of ten boreholes are needed to produce effective dewatering over a period of six months. Five dewatering boreholes spaced approximately 500 m apart, located in the pit and around the pit perimeters, have been drilled to date. These were equipped with submersible pumps to start depleting the five identified fractured rock aquifers at a dewatering rate of 3,300 m³/day. The remaining five dewatering boreholes will be drilled early in 2022. This is primarily due to the low storativity calculated for the aquifers. The Transmission 11kv Overhead powerline was extended from Pakaka to supply energy to borehole pumps at Pamao. Surface water was also assessed to minimise the seasonal inflow of run-off by designing the surface water diversion along induced topographic gradients through open channels, while pumps will be used to evacuate water from the in-pit sumps and the upstream catchment.

Figure 16-13 presents the Pamao Pit dewatering design criteria.

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Table 16-13 KCD Pushback 3 Pit 2021 Dewatering Operations Summary



Month	Rain (m)	Run-off Co-efficient	Pit Area (m²)	In-Pit Rain Inflow Vol (m³)	Ground Water Inflow	Dynamic Head (Head+ Loss)	Diesel Pump Capacity (m³/h)	Running Hours	Diesel Consum. (50L/hr)	Diesel Cost ($1.20/L)	Operating Cost ($7.16/hr)	Pump Rental Cost ($9.50/hr)	Total Cost ($)

Jul	0.219	0.68	307,104	45,734	2,678	83	300	161	8,069	9,682	1,155	1,533	12,371

Aug	0.212	0.68	307,104	44,272	2,678	83	300	157	7,825	9,390	1,121	1,487	11,997

Sep	0.233	0.68	307,104	48,658	2,592	83	300	171	8,542	10,250	1,223	1,623	13,096

Oct	0.283	0.68	307,104	59,099	2,678	83	300	206	10,296	12,356	1,474	1,956	15,786

Nov	0.147	0.68	307,104	30,698	2,592	83	300	111	5,548	6,658	795	1,054	8,507

Total				228,461	13,219			806	40,280	48,336	5,768	7,653	61,757

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Source: Kibali Goldmines, 2021

Figure 16-13 Pamao Pit Dewatering Design Criteria

To improve the site water balance and reduce abstraction in the Kibali River; all abandoned pits are used as dams to collect and store both seepage and rainfall as fresh water that is now recirculated as service water in process plant operations via a newly installed 250 mm HDPE water line. This will assist in mitigating the negative impact on the Kibali River during the period of dry season. The current volume of fresh water held in the dams (Pakaka, Kombokolo and Sessenge) equal to 3.1 million cubic metres (Mm³). Figure 16-14 presents the Pakaka Dam freshwater reservoir design criteria.

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Figure 16-14 Pakaka Dam Fresh Water Reservoir Design Criteria

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Dilution and Mining Recovery

Global 3% and 10% factors have been used for ore losses and ore dilution in the estimation of open pit Mineral Reserves.

Testing of Ore Pro software is being completed to define approximate measured numbers for the dilution.

The QP considers that the dilution and loss factors are reasonable assumptions for the estimation of the Mineral Reserves.

Mine Design

The selected pit shells were used as guidelines to design the practical ultimate pits with internal phases. Pit design parameters were selected based on the overall pit geometry, geotechnical data and information, and the mine production rate. Pit and internal phases were designed using Surpac software, integrating the recommended standards for road width and minimum mining width based on an efficient operation for the size of mining equipment chosen for the open pit operations.

Comparisons of 2020 Whittle shells to 2021 shells and 2021 reserve pits were completed to assess the changes. The Gorumbwa, Pamao, and Pakaka pits were redesigned.

All designs were based on approved geotechnical slope angles provided by the Geotechnical department and consultants; these have been detailed previously in this Section under Slope Angles, and they are summarised again in Table 16-14 (Refer to sub-section (16.2, Geotechnical) detailing the Slope Angles.).

Table 16-14 Summary of Pit Design Parameters



Material	Bench Height (m)	Berm Width (m)	Batter Angle (°)	IRA (°)

Weathered	5-10	4-5	27-50	27-50

Transition	10	4-6	27-65	38-55

Fresh Rock	10-20	4-6	50-80	48-59

The three-staged Gorumbwa pit design strategy was maintained for 2021 (Figure 16-15). Mining at Gorumbwa for 2021 concentrated on Pushback one which is expected to be mined out by Q1 2022. The second stage and the final stage are planned for 2022.

All designs were checked to ensure that they provided enough width for the mining fleet to avoid constraints and difficulties during excavations. The selected and designed pushbacks ensured adequate waste deferral in the early stages and provided a continuous supply of the appropriate blend of ore to the plant.

The year end 2021 Kibali open pit designs are shown in Figure 16-15 to Figure 16-22.

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Source: Kibali Goldmines, 2021

Figure 16-15 Gorumbwa Sessenge Upper Pit Designs

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Source: Kibali Goldmines, 2021

Figure 16-16 KCD Pit Design

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Source: Kibali Goldmines, 2021

Figure 16-17 Kalimva Pit Design

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Source: Kibali Goldmines, 2021

Figure 16-18 Ikamva Pit Design

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Source: Kibali Goldmines, 2021

Figure 16-19 Oere Pit Design

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Source: Kibali Goldmines, 2021

Figure 16-20 Pakaka Pit Design

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Source: Kibali Goldmines, 2021

Figure 16-21 Aerodrome Pit Design

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Source: Kibali Goldmines, 2021

Figure 16-22 Pamao and Pamao South Pit Design

Open Pit Mining Fleet

During 2021, local mining contractors were used to strip the Aerodrome pit, and this will continue up until the pit is mined out mid-2022.

The mining fleet is presented in Table 16-15. The fleet size for 2022 and beyond is projected to remain consistent with no further material additions of equipment required. The maintenance schedule allows for some annual rebuilds of the equipment each year. In opinion of the QP, the fleet size is adequate to achieve the LOM production targets based on Mineral Reserves.

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Table 16-15 Current Primary Open Pit Mine Equipment Fleet



Fleet	Current Quantity	Planned	Planned
Fleet	2021	2022	2023-2025

LIEBHERR 9350 Excavators	1	1	2

LIEBHERR 984 Excavators	0	0	0

LIEBHERR 9200 Excavators	4	4	4

CAT777G DUMP Trucks	22	22	22

CAT992WHEEL Loaders	2	2	2

CAT D9R Dozer	7	7	7

CAT 16M Graders	3	3	3

CAT 834 Pusher	2	2	2

Blast Drill Drigs	8	8	8

Water Bowsers	2	2	2

The contractor overall workforce is 531 people, of which 242 are working for load and haul, 74 for drilling and blasting, 174 for plant maintenance of equipment, and the remaining for the administration and Environmental Health and Safety (EHS) team.

Waste Dumps

An estimated 300 Mt of waste will be mined over the remaining LOM based on Mineral Reserves.

The capacity of the Kibali open pit waste dumps has been evaluated based on the latest pit designs to confirm that there is adequate dump capacity for the estimated LOM tonnage of waste based on Mineral Reserves. A swell factor of 30% was considered in all waste dump capacity evaluations. Haul roads were also adjusted, where necessary, to ensure they provide easy access where pit ramps are day lighting.

No in-pit dumping was carried out in 2021 and none is planned for 2022, as the Kombokolo, Mengu Hill, KCD, and Sessenge mines continue to explore potential underground opportunities. Future work will, however, consider the use of some of the satellite pits with low potential for future underground operations for waste disposal, based on the mining sequence. Within the current LOM, Pamao has been planned to be backfilled with tailings upon exhaustion of the mineral reserve.

Table 16-16 summarises the related waste dump capacities.

Table 16-16 Waste Dump Capacities



Waste Dump	Design Capacity (Mm³)	Waste Dumped to Date (Mm³)	Planned Future Waste (Mm³)

KCD/Sessenge	41.72	37.18	4.54

Gorumbwa	29.98	12.65	17.33

Pamao	37.82	0.00	37.82

Kalimva-Ikamva	59.46	0.00	59.46

Megi Marakeke Sayi	53.53	0.00	53.53

Pakaka/Aerodrome	49.20	14.73	34.48

Total	271.71	64.56	207.15

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16.3 Underground Mining

The Kibali underground mine is a long hole stoping operation producing at a rate of 3.8 million ore tonnes per year. Development of the underground mine commenced in 2013. Stoping commenced in 2015 and ore production has ramped up to 1.8 Mt in 2017 and 3.6 Mt in 2021. Initial production was truck hauled by a twin decline to surface. In 2017, the haulage shaft (740 m deep) and materials handling system was commissioned. From 2018 onwards, underground ore has predominantly been hoisted up the shaft. The decline to surface will be used to haul some of the shallower zones and to supplement shaft haulage.

A significant portion of the capital and access development for the mine is in place. To date, 43,609 m of capital and access development has been completed. The current LOM plan contains a further 9,928 m of capital lateral development based on Mineral Reserves. The key capital infrastructure remaining to be developed are the 9101 decline, 9101 incline, southern exhaust raises and the 3101 / 3102 access development. Figure 16-23 shows the current (December 2021) mine as-built and the LOM development. Existing infrastructure comprises:

●

A vertical shaft

●

Mobile equipment mining fleet

●

Backfill plant

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Batch plant

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Underground dewatering facility

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Surface compressor house

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Multiple surface workshop facilities

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Electrical power line connection to the grid

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Office building

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Warehouse

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Water clarifying plant

Ore from stopes is loaded (both by teleremote and conventional manual loaders) from the stopes into the eight ore passes via finger raises on the respective levels. This ore is then transferred by Autonomous LHDs into two coarse ore bins and then into two primary crushers, followed by two fine ore bins and independent skip loadout conveyors near the shaft bottom. The proposed mining methods are variants of long hole open stoping with cemented paste:

●

Longitudinal open stoping (18% of Mineral Reserve tonnes) is used in narrow zones (< 15 m width) with variable interlevel heights between 20 m and 30 m.

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Dilution and Mining Recovery

The actual stope performance is routinely reconciled against the planned performance. A dilution and mining loss matrix was developed based on Kibali site-specific experience, with allowances for certain expected problems associated with geotechnical structure, paste fill exposure and stope sequence configuration.

Dilution

Two forms of dilution have been considered in the 2021 Mineral Reserve estimate, including: rock dilution and paste dilution.

Rock dilution and paste dilution have been combined and applied as a combined dilution percentage in Mineral Reserves. The factors are shown in the dilution matrix in Table 16-19.

Rock dilution is added as a percentage of stope tonnes. Rock dilution is dilution outside of the designed mining shapes on the footwall, hanging wall or sidewall of the stopes. Unplanned dilution is added at a grade of 0.00 g/t Au. The unplanned dilution applied is based on the Kibali historical stope performance database.

Paste dilution is the dilution from adjacent paste fill exposures. Paste dilution is added where a stope has paste fill exposure. Paste dilution exposures have been estimated as 2% per paste fill exposure for primary, secondary, transverse advancing face stope, and longitudinal stope. Hence, primary or secondary stopes with two paste exposure walls will have 4% paste dilution and stopes with three paste exposure walls will have 6% paste dilution.

Table 16-17 shows that there has been considerable improvement since last year on the paste dilution experienced in the secondary stope. This is mainly due to a change to a 90/10 slag cement instead of the 70/30 slag cement previously used as the paste fill binder.

Table 16-17 Summary of Historical Paste Dilution per Stoping Type

2016

2017

2018

2019

2020

2021

Paste

Dilution

(%)

1°

2°

Lon

1°

2°

Lon

1°

2°

Lon

1°

2°

Lon

1°

2°

Lon

1°

2°

Lon

1.40

2.50

0.15

0.35

0.4

3.9

0.86

9.36

1.04

3.85

Notes:

1°=Primary, 2°=Secondary, Lon=Longitudinal

The 90/10 slag cement was introduced as the 70/30 slag cement experienced faster Uniaxial Compressive Strength (UCS) deterioration over the 180 day curing period. Better UCS was achieved using the 90/10 slag cement (Table 16-18). The 70/30 slag cement started losing strength after 90 days of curing, whereas the 90/10 slag cement strength was more stable over time. The strength deterioration of the 70/30 is due to the higher calcium ion content within the paste.

Despite the yearly improvement, the paste dilution in the secondary stope has been higher than the overall dilution expected. This is mainly due to higher paste dilution experienced while mining some stopes adjacent to the remaining stopes filled with 70/30 slag cement.

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Table 16-18 Summary of 70/30 and 90/10 Historical Paste Test Work UCS Strength



Curing Time (days)	70- 30 @ 1%	70- 30 @ 1.5%	70- 30 @ 2.0%	70- 30 @ 2.5%	70- 30 @ 3.0%	70- 30 @ 4.0%	70- 30 @ 4.5%	90- 10 @ 1%	90- 10 @ 1.5%	90- 10 @ 2.0%	90- 10 @ 2.5%	90- 10 @ 3.0%	90- 10 @ 4.0%	90- 10 @ 4.5%

3	189	188	197	309	421	699	667	179	222	261	301	560	708	697

14	208	274	365	330	924	920	1002	214	341	354	765	984	804	994

28	180	282	377	783	1028	1166	1800	230	370	388	913	1165	987	1611

56	158	266	438	652	1030	1330	1811	288	388	469	938	1213	1493	1683

90	162	246	431	635	905	1104	1757	246	368	457	948	1457	1681	1757

112	140	233	341	481	771	1003	1452	243	358	452	780	1353	1678	1720

180	121	137	0	481	655	912	1449	231	188	398	684	1225	1610	1696

240	114	129	0	435	583	888	1269	132	165	350	600	937	1497	1491

365	88	0	0	231	0	820	1043	0	147	204	305	752	1248	1371

The 90/10 slag and the introduction of WebGen technology for blasting secondary stope is anticipated to lead to lower backfill dilution in the future. In addition, higher mining losses and dilution have been applied to secondary stopes that will be mined adjacent to poor UCS legacy 70/30 stopes

Dilution Matrix

Based on historical stope reconciliation data and stope closure note data, the equivalent linear over-break/slough (ELOS) dilution matrix was developed. The ELOS Dilution Matrix will evolve and will be updated as new and additional stope performance data becomes available over time.

This is a more reliable basis for forecasting dilution and ore loss rates. This strategy is followed at other large stoping operations and is an appropriate approach for Kibali.

The (ELOS) used for quantifying total dilution are presented in Table 16-19 below.

Based on the historical data, the mining method, and the stope configuration different dilution factor are applied.

For stopes adjacent to poor UCS quality stopes:

●

To mitigate higher dilution that could result in stope sterilisation, mining losses of 25% were applied. The 25% mining losses assumed that a 4 m to 5 m pillar width will be left when mining adjacent to 70/30 poor strength paste. In addition, a dilution factor of 15% was applied as well.

For transverse primary stopes with 20 m average thickness:

●

Transverse primary hanging stope dilution – 2.0%.

●

Transverse primary footwall stope dilution – 4.0%.

●

Transverse primary hanging stope intersected by geotechnical structure dilution – 4.3%.

●

Transverse primary footwall stope intersected by geotechnical structure dilution – 6.4%.

For transverse stopes with 30 m average thickness the following dilutions were applied:

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●

Transverse secondary hanging stope dilution – 7%.

●

Transverse secondary footwall stope dilution – 12.5%.

●

Transverse secondary hanging stope intersected by geotechnical structure dilution –10%.

●

Transverse secondary footwall stope intersected by geotechnical structure dilution – 12.5%.

For transverse advancing stopes with 25 m average thickness and 30 m average height the following dilutions were applied:

●

Transverse advancing face stope dilution – 6.0%.

●

Transverse advancing face stope dilution – 6.0%.

●

Transverse advancing face stope intersected by geotechnical structure dilution –6.0%.

●

Transverse advancing face stope not intersected by geotechnical structure dilution –6.0%.

For longitudinal stopes with a 10 m to 15 m average thickness the following dilutions were applied:

●

Longitudinal stope not intersected by geotechnical structure dilution – 4%.

●

Longitudinal stope intersected by geotechnical structure dilution – 4.5%.

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Table 16-19 Summary of Historical Stope Performance and Dilution Parameters



Stoping Type	Sequence Configurations	Stope Name	Stope Dimension												Calculated Rock Dilution	Applied Dilution Factor	Number of Fill Exposures	Fill Dilution by Exposures	Paste Dilution	Total Unplanned + Paste Dilution
Stoping Type	Sequence Configurations	Stope Name	Average Width (m)	Height (m)	Length (m)	Area (m²)	Perimeter (m)	Hydraulic Radius (m)	Dip of Stope Surface HW/ FW (°)	HW Calculated Overbreak (m³)	FW Calculated Overbreak (m³)	HW ELOS	FW ELOS	%	Calculated Rock Dilution	Applied Dilution Factor	Number of Fill Exposures	Fill Dilution by Exposures	Paste Dilution	Total Unplanned + Paste Dilution

Transverse primary stope	Hanging wall	CB3_420_XC20_315	19.71	119.83	19.71	2109.80	278.54	7.57	82.00	482.00	-	0.23	-	0.01	1.80%	2.00%	0.00	N/A	0.00%	2.00%
		C_455_XC18_350	20.06	115.00	20.06	2134.20	262.53	8.13	53.00	307.00	-	0.14	-	0.72%
		C_455_XC18_525	20.00	79.27	20.00	1521.62	198.15	7.68	52.00	261.00	-	0.17	-	0.86%
		C_595_XC14_560_HW	21.20	44.00	21.20	898.54	131.63	6.83	58.00	63.00	-	0.07	-	0.33%
		C_595_XC12_490	20.40	124.34	20.40	2343.96	285.58	8.21	57.00	1371.00	-	0.58	-	2.87%
		CB3_490_XC#8_385	20.00	99.00	20.00	1908.53	238.40	8.01	75.00	1011.00	-	0.53	-	2.65%
		C_630_XC10_490	20.20	166.00	20.20	3310.56	370.68	8.93	56.00	2528.00	-	0.76	-	3.78%
		C_275_XC32_315	19.75	31.29	19.75	620.59	102.39	6.06	90.00	113.50	-	0.18	-	0.93%
		C_XC8 (665-595)_FW	20.01	88.80	20.01	1720.70	213.97	8.04	57.00	752.00	-	0.44	-	2.18%
		C_595_XC12_UH	28.40	24.90	28.40	707.12	106.62	6.63	64.00	448.30	-	0.63	-	2.23%
		B-630-605-XC6(LS#11)	20.50	30.90	20.50	546.26	111.02	4.92	50.00	246.00	-	0.45	-	2.20%
		C_525_XC12_D_560	24.95	37.36	24.95	875.84	124.36	7.04	49.00	410.52	-	0.47	-	1.88%
		C_350_XC30_420_FW	19.89	23.19	19.89	461.51	86.19	5.35	90.00	149.86	-	0.32	-	1.63%
	Footwall - (Not Hanging wall)	C_455_XC18_420_FW	20.30	36.92	20.30	753.30	115.01	6.55	65.00	-	77.90	-	0.10	0.51%	1.79%	2.00%	1.00	2.00%	2.00%	4.00%
		C_525_XC26_420_FW	20.14	116.56	20.14	2150.87	272.23	7.90	80.00	-	1075.20	-	0.50	2.48%
		C_490_XC28_385_FW	20.00	117.60	20.00	2233.47	278.92	8.01	58.00	-	1066.70	-	0.48	2.39%
	Hanging wall Intersected by geotech structure	C_350_XC26_275	20.08	81.54	20.08	1554.15	214.19	7.26	49.00	981.00	-	0.63	-	3.14%	3.43%	4.30%	0.00	N/A	0.00%	4.30%
		CB3_455_XC#12_350	20.06	107.93	20.06	2164.52	257.91	8.39	57.00	1746.00	-	0.81	-	4.02%
		C_525_XC24_420	20.16	118.68	20.16	2355.54	277.76	8.48	66.00	2608.00	-	1.11	-	5.49%
		CB3_315_420_XC#14	19.99	110.82	19.99	2204.57	267.84	8.23	74.00	3735.60	-	1.69	-	8.48%
		CB3_490_XC#10_350	20.00	145.23	20.00	2788.04	330.60	8.43	69.00	287.00	-	0.10	-	0.51%
		C_525_XC30_350	20.00	210.07	20.00	3684.05	460.29	8.00	61.00	2744.38	-	0.74	-	3.72%
		C_315_XC30_275	20.00	32.34	20.00	643.84	4.43	145.47	78.00	474.00	-	0.74	-	3.68%
		C_350_XC28_385	20.20	42.28	20.20	821.45	132.86	6.18	49.00	273.00	-	0.33	-	1.65%
		CB3_275_385_XC#22	19.86	120.80	19.86	2113.91	282.93	7.47	60.00	2488.00	-	1.18	-	5.93%
		C_245_XC32_315	19.75	54.35	19.75	728.08	113.22	6.43	84.00	142.15	-	0.20	-	0.99%
		CB3_420_XC6_460	19.01	40.02	19.01	759.41	118.01	6.43	59.00	273.78	-	0.36	-	1.90%
		C_385_XC22_455_Hw	19.93	71.73	19.93	1489.77	207.71	7.17	51.00	105.23	-	0.07	-	0.35%
		C_525_XC14_D_560	24.99	33.00	24.99	817.30	124.28	6.58	73.00	859.32	-	1.05	-	4.21%
		CB3_245_XC18_315	20.00	76.16	20.00	1487.71	201.66	7.38	60.00	1983.67	1002.22	1.33	-	6.67%
		CB3_315_XC18_350	20.00	36.65	20.00	733.08	113.31	6.47	72.00	249.68	130.74	0.34	-	1.70%
		C_315_XC24_420	20.01	110.18	20.01	1431.24	197.20	7.26	79.00	677.65	637.46	0.47	-	2.37%
	Footwall (Not Hanging wall Intersected by geotech structure)	CB3_245_350_XC#22_FW	19.94	108.28	19.94	2109.56	256.08	8.24	86.00	-	1265.10	-	0.60	3.01%	3.09%	4.40%	1.00	2.00%	2.00%	6.40%
		C_420_315_XC#20_FW	19.97	109.19	19.97	2163.51	258.10	8.38	88.00	-	2379.00	-	1.10	5.51%
		CB3_245_XC24_315_FW	20.01	73.62	20.01	1389.87	191.38	7.26	73.00	-	213.99	-	0.15	0.77%

Transverse secondary stope	Hanging wall	C_595_XC7_560	25.48	47.88	25.48	1132.69	141.44	8.01	49.00	592.00	-	0.52	-	2.05%	1.49%	2.00%	2.00	2.50%	5.00%	7.00%
		C_490_XC13_525	21.20	37.00	21.20	782.58	116.54	6.72	62.00	303.00	-	0.39	-	1.83%
		B1-605-580-XC5(LS#17)	30.10	33.67	30.10	1013.36	127.73	7.93	44.00	412.00	-	0.41	-	1.35%
		C_455_XC17_385	29.54	86.86	29.54	1935.20	210.41	9.20	55.00	743.00	-	0.38	-	1.30%
		CB3_460_XC9_AB3_490	27.49	38.04	27.49	896.12	120.18	7.46	81.00	211.35	130.01	0.24	-	0.86%
		C_385_XC21_420_UH	29.30	24.54	29.30	679.95	113.97	5.97	80.00	209.89		0.31	-	1.05%
		C_245_XC25_315	28.10	78.50	28.10	1811.25	201.59	8.98	61.00	1003.68		0.55	-	1.97%
	Footwall - (Not Hanging wall)	C_420_XC17_385_FW	29.69	75.50	29.69	2201.27	211.81	10.39	67.00	-	3355.90	-	1.52	5.13%	5.13%	5.00%	3.00	2.50%	7.50%	12.50%
	Hanging wall Intersected by geotech structure	C_490_XC13_455	18.74	42.70	18.74	796.72	122.64	6.50	53.00	845.58	-	1.06	-	5.66%	3.44%	5.00%	2.00	2.50%	5.00%	10.00%
		CB3_350_385_XC#13	30.36	36.20	30.36	1113.85	134.19	8.30	66.00	2481.00	-	2.23	-	7.34%
		CB3_315_XC#15_350	30.15	39.52	30.15	1220.46	141.19	8.64	59.00	886.50	-	0.73	-	2.41%
		C_490_XC15_420	30.10	87.27	30.10	2375.93	288.55	8.23	46.00	1826.00	-	0.77	-	2.55%
		B1-630-605-XC5(LS#18)	30.08	35.33	30.08	1061.26	130.75	8.12	50.00	1397.00	-	1.32	-	4.38%
		C_275_XC27_315	29.00	40.30	29.00	1184.63	138.97	8.52	54.00	386.94	565.67	0.33	-	1.13%
		C_275_XC29_315	26.48	35.35	26.48	921.90	122.42	7.53	72.00	147.67	1032.18	0.16	-	0.60%
	Footwall (Not Hanging wall Intersected by geotech structure)	CB3_350_XC19_385_FW	29.30	34.49	29.30	1083.44	131.78	8.22	74.00	-	124.52	-	0.11	0.39%	1.42%	5.00%	3.00	2.50%	7.50%	12.50%
		C_385_XC27_455_FW	29.80	78.81	29.80	2193.55	221.36	9.91	74.00	-	1601.45	-	0.73	2.45%	1.42%	5.00%	3.00	2.50%	7.50%	12.50%

Transverse advancing face stope	Hanging wall	N/A	-	-												2.00%	2.00	2.00%	4.00%	6.00%
	Not Hanging wall	N/A	-	-												2.00%	2.00	2.00%	4.00%	6.00%
	Hanging wall Intersected by geotech structure	N/A	-													2.00%	2.00	2.00%	4.00%	6.00%
	Not Hanging wall Intersected by geotech structure	N/A	-													2.00%	2.00	2.00%	4.00%	6.00%

Longitudinal stope	Longitudinal (Not intersected with geotech structure)	-	-		-											2.00%	1.00	2.00%	2.00%	4.00%
		-	-		-											2.00%	1.00	2.00%	2.00%	4.00%
	Longitudinal (intersected with geotech structure)	B1_OD_680_655	42.44	33.57	42.44	1370.189	153.585	8.92	54.00	1417.00		1.03	-	2.44%	-	2.50%	1.00	2.00%	2.00%	4.50%
	Longitudinal (intersected with geotech structure)	-	-												-	2.50%	1.00	2.00%	2.00%	4.50%

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Mining Losses

Mining loss occurs where there is under break in the stope or broken ore is left in the stope at the completion of mining. Like the dilution matrix, a mining losses matrix was developed based on historical stope reconciliation data and stope closure note data.

The mining losses experienced in 2021 have been within the range of estimated mining losses in 2020 Table 16-21.

The mining losses matrix used in the 2021 Mineral Reserve is presented in Table 16-20 and the historical performances in Table 16-22.

Table 16-20 Summary of Mineral Reserve Estimate Recovery Parameters



Stoping Type	Sequence Configurations	Recovery

Transverse primary stope	Hanging wall	95.4%
	Not Hanging wall	92.5%
	Hanging wall Intersected by geotechnical structure	90.5%
	Not Hanging wall Intersected by geotechnical structure	91%

Transverse secondary stope	Hanging wall	90.7%
	Not Hanging wall	88.5%
	Hanging wall Intersected by geotechnical structure	85%
	Not Hanging wall Intersected by geotechnical structure	86%

Transverse advancing face stope	Hanging wall	90%
	Not Hanging wall	90%
	Hanging wall Intersected by geotechnical structure	88%
	Not Hanging wall Intersected by geotechnical structure	86%

Longitudinal stope	Hanging wall	92%
Longitudinal stope	Hanging wall Intersected by geotechnical structure	90%

Table 16-21 Stope Recovery History



Stoping type	Sequence configurations	2020 Recovery	2021 Recovery

Transverse primary stope	Hanging wall	95%	96%
	Not Hanging wall	92%	93%
	Hanging wall Intersected by geotechnical structure	90%	91%
	Not Hanging wall Intersected by geotechnical structure	91%	91%

Transverse secondary stope	Hanging wall	90%	91%
	Not Hanging wall	88%	89%
	Hanging wall Intersected by geotechnical structure	85%	87%
	Not Hanging wall Intersected by geotechnical structure	86%	86%

Longitudinal stope	Hanging wall	92%	-
Longitudinal stope	Hanging wall Intersected by geotechnical structure	90%	-

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Table 16-22 Stope Recovery and Unplanned Dilution History



Year	2016	2017	2018	2019	2020	2021

Ore Recovery	90.0%	86.0%	90.0%	92.0%	91.0%	91.6%

Unplanned Dilution	8.0%	5.0%	4.0%	7.0%	7.0%	4.73%

Dilution and Mining Loss Improvement

Kibali has made a significant effort toward improving the drill and blasting practices. The continuous optimisation of the drill and blast designs has been providing step improvement in the dilution and mining losses. Further improvements are planned for 2022 onwards. These include:

●

The implementation of the WebGen technology for blasting secondary stopes

●

The implementation of the Sandvik OptiMine^® measurement while drilling

●

The implementation of longitudinal ring design approach in secondary stope

WebGen Technology

WebGen is an efficient blasting technology for production stopes and pillars developed by Orica that has been successfully implemented at other Barrick underground mines. Preliminary investigations in 2021 indicate WebGen technology will likely result in reducing paste dilution, improving the stope recovery, productivity, and safety.

With the implementation of WebGen wireless blasting, most of the secondary stopes will be blasted and muck cleaned before firing the WebGen pillar that is adjacent to the paste fill stope. As shown in the firing sequence in Table 16-33, the inaccessible pillar adjacent to the paste fill primary stope will be pre-charged with WebGen allowing the firing against the paste to take place wirelessly at a later stage. This reduces overall paste dilution and mining losses.

Figure 16-24 illustrates the stope mining sequence using WebGen wireless blasting.

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Source: Kibali Goldmines, 2021

Figure 16-24 Stope Mining Sequence Using WebGen Wireless Blasting

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Measurement while Drilling

The implementation of up and down hole drilling over the past year has significantly reduced the hole length and the downhole deviation. However, there are circumstances where drilling up and down holes is not practical and thus, Sandvik OptiMine^® measurement while drilling system will be implemented.

Initial investigation shows that the implementation measurement while drilling will likely result in reducing the deviation, hence the underbreak and the overbreak. The measurement while drilling system will provide timely analysis of the hole deviation while drilling. This will allow the production driller to better control drilling parameters such as the feed and percussion pressure, rotation speed, water pressure.

Longitudinal Ring Design Approach in Transverse Secondary Stope

The longitudinal ring design approach was recently implemented in secondary stopes to reduce the impact of the explosive energy on weaker hanging wall material. The longitudinal approach allows the blasting energy to be parallel to the weak hanging wall or structure. As shown in Figure 16-25, this design approach will likely result in reducing the overbreak due to the favourable orientation of the blast.

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Source: Kibali Goldmines, 2021

Figure 16-25 Illustration of the Longitudinal Ring Design Approach in Secondary Stopes

Mine Design

The Kibali underground mining methods are variants of long hole open stoping with cemented paste backfill. Figure 16-26 demonstrates the mining methods in use.

The mine is accessed via a twin decline, a vertical shaft, and a system of internal ramps. Ore from stopes is loaded (both by teleremote and conventional manual loaders) from the stopes into the eight ore passes via finger raises on the respective levels. This ore is then transferred by Autonomous LHDs into two coarse ore bins and then into two primary crushers, followed by two fine ore bins and independent skip loadout conveyors near the shaft bottom.

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No significant failures of the openings in the underground workings have occurred. In general, the rock mass is classified as good with average rock mass rating values between 64 and 73, and Rock Mass Quality (Q’) values between 31 and 47. The mine is currently producing an average of 10,500 t/day.

The mining methods used are supported by operational data and are reviewed periodically as further resource infill and grade control drilling changes the shape of the ore zones.

There are three distinct sequencing patterns for the various mining methods, including transverse primary and secondary stoping, advancing face stoping and longitudinal stoping.

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Source: Kibali Goldmines, 2021

Figure 16-26 Kibali Underground Mineral Reserve by Mining Method (Looking NW)

Stope Design

Original stope dimensions were developed by SRK Consulting (du Plooy, 2011) The stope design approach adopted is primarily based on the allowable stope Hydraulic Radius (HR).

When undertaking a stope stability assessment, the approach adopted is the Modified Stability Graph. There are two stability graphs that are used for assessing the stope dimensions, dependent on whether cable bolt reinforcement is used or not. The ‘Database of Cablebolt-Supported Stopes’ (Figure 16-27) is applied to the stability assessment for the crown (whereas per standard practice, cable bolts are installed in the crown of each stope, regardless of the dimensions). The Modified Stability Graph that encompasses the ‘Database of Unsupported Stopes’ is applied to the stability assessment for the four sidewalls (Figure 16-28).

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As part of the assessment, the analysis also applies the Mining Rock Mass Model (MRMM) and structural model. By incorporating the MRMM and appreciating the structural model, stopes are dimensioned considering the rock mass quality. The MRMM has enabled stope dimension optimising based on the geotechnical domains within each orebody zone rather than a single ‘one size fits all’ approach. Areas within each orebody that are competent allow for potentially larger stopes to be established, while areas of reduced rock mass quality will require smaller sized stopes.

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Figure 16-27 Database of Cablebolt-Supported Stopes

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Figure 16-28 Database of Unsupported Stopes

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Primary / Secondary Transverse Stoping

The transverse primary and secondary sequencing concept is that primary stopes are mined from hanging wall to footwall and multi-level stopes are mined concurrently up to the design vertical height. Secondary stopes follow the primary stopes. A secondary stope cannot start mining until the primary stopes on either side have been mined and filled.

The level interval is 35 m (floor to floor), and stopes are mind as either single lift or multiple lifts (up to four) depending on stope geometry and stable span analysis. Primary stopes are 20 m along strike and secondary stopes are 30 m along strike. The width of primary stopes can be up to 40 m across strike. The controlling span for primary stope size is general the side (north and south) rock walls. Secondary stopes are up to the 30 m across strike. The controlling span for secondary stope size is generally the side wall paste exposure of the adjacent primary stopes. Where the orebody is too wide for a single stope span (> 30 m to 40 m wide), multiple primary and secondary stopes are mined retreating from hanging wall to footwall.

Figure 16-29 shows that the primary stopes are mined, and paste filled prior to mining of the adjacent secondary stopes. The stopes are mined in a bottom-up fashion. Production drill holes are a combination of up and down holes or down holes of 102 mm diameter.

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Source: Kibali Goldmines, 2021

Figure 16-29 Transverse Stope Sequencing

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Advancing Face Transverse Stoping

Advancing face transverse stoping is used in the 9101 zone which has a shallow plunge (20° to 30°) to the NE. The level interval varies from 25 m to 30 m to optimise extraction. The stopes are 25 m down plunge and 25 m across plunge. Figure 16-30 shows that the stoping front (F) advances from NE to SW and toward the shallower part of the 9101 lode.

●

Stopes located in mining front 1 (F1) must be mined and paste filled before the stope located in mining front 2 (F2) can be mined, and mining front 2 (F2) has to be mined and paste filled before mining can take place in mining front 3 (F3).

●

Stopes located in the same mining front that are accessed through different ore drives can be mined simultaneously. As an example, different stopes located in mining front 5 (F5) can be mined simultaneously since they are being access through different ore drive.

●

Stopes are mined as a single lift or multiple lift (up to three lifts) depending on ore zone thickness. Stopes are paste filled prior to the mining of adjacent stopes.

●

A slot raise is developed by production drilling machine or by raise boring. Production drill holes are either a combination of up and down holes or down holes of 102 mm diameter.

This mining sequence is designed to avoid the creation of pillars, which may potentially become highly stressed as mining progress.

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Source: Kibali Goldmines, 2021

Figure 16-30 Transverse Advancing Face Sequencing

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Longitudinal Stoping

Longitudinal stoping is used as the main extraction method to mine the narrower stopes (< 15 m width). In the steeper areas (>60°), the level interval varies from 20 m to 35 m. In the flatter areas (5 to 60°) ore drives are located on the footwall and level interval is controlled by the dip, minimising footwall waste and limiting stope width (up dip) to 20 m. Stopes are paste filled prior to mining of adjacent stopes to maintain hanging wall stability (Figure 16-31).

●

Block 1 must be mined, and paste filled prior to mining of block 2 and block 2 stopes have to be mined and backfilled prior to mining of block 3.

●

Stopes located in the same block are mined as a single lift or multiple lift (up to three lifts).

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Source: Kibali Goldmines, 2021

Figure 16-31 Longitudinal Mining Sequencing

Summary of Stoping Types

The proportion of the reserve per lode and per stoping type is shown in Table 16-23.

Table 16-23 Summary of Proportion of the Mineral Reserve per Stoping Type



Mining Zone	Transverse Primary Stope	Transverse Secondary Stope	Longitudinal Stope	Transverse Advance Face Stope

3101 Zone	43%	44%	13%	-

3102 Zone	5%	4%	91%	-

5101 Zone	25%	75%	-	-

5102 Zone	23%	43%	34%	-

5104 Zone	-	-	100%	-

5105 Zone	15%	36%	50%	-

5110 Zone	23%	-	77%	-

9101 Zone	-	-	-	100%

9105 Zone	34%	48%	18%	-

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The historical proportion mined per lode and per stope type are shown in in Table 16-24. The yearly proportion of stopes mined per lode are shown in Table 16-25. These tables show that from the beginning of the underground project to date, the majority of the underground stopes have been produced from the 5101, 5102, 5105, 9105, 5110, and currently the 3000 lode. However, mining of the 9101 lode will start from 2022 onward, and their modifying factors will be refined over time as more data become available.

Table 16-24 Historical Proportion of the Stopes Mined per Lode and per Stope Type



Mining Zone	Transverse Primary Stope	Transverse Secondary Stope	Longitudinal Stope	Transverse Advance Face Stope

3101 Zone	1%	-	-	-

3102 Zone	-	-	-	-

5101 Zone	21%	18%	-	-

5102 Zone	16%	2%	-	-

5104 Zone	-	-	-	-

5105 Zone	6%	5%	1%	-

5110 Zone	11%	6%	-	-

9101 Zone	-	-	-	-

9105 Zone	6%	7%	-	-

Table 16-25 Summary of Yearly Historical Proportion of the Stopes Mined per Lode



Mining Zone	2015	2016	2017	2018	2019	2020	2021

3101 Zone	-	-	-	-	-	-	6%

3102 Zone	-	-	-	-	-	-	-

5101 Zone	79%	51%	49%	21%	55%	59%	53%

5102 Zone	-	-	-	43%	6%	23%	8%

5104 Zone	-	-	-	-	-	-	-

5105 Zone	21%	49%	44%	14%	2%	-	-

5110 Zone	-	-	7%	22%	1%	-	-

9101 Zone	-	-	-	-	-	-	-

9105 Zone	-	-	-	1%	37%	19%	33%

16.4 Underground Mining Operations

The Kibali underground mine is a long hole stoping operation currently producing at a rate of 3.8 million ore tonnes per year. Development of the underground mine commenced in 2013. Stoping commenced in 2015 and ore production has ramped up to 1.8 Mt in 2017 and 3.6 Mt in 2021. Initial production was truck hauled by a twin decline to surface. In 2017, the haulage shaft (740 m deep) and materials handling system was commissioned. From 2018 onwards, underground ore has been predominantly hoisted up the shaft. The decline to surface will be used to haul some of the shallower zones and to supplement shaft haulage.

There are four main mineralised zones, 5101, 5102, 9101 and 9105 that contribute the bulk of the Mineral Reserve (Figure 16-32). Five other mineralised zones, 3101, 3102, 5104, 5105 and 5110 contribute approximately 18 % of the remaining Mineral Reserve.

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Figure 16-32 KCD Underground Tonnes (Ore and Waste) Production History

Unit Operations

The development cycle consists of the drilling, charging, firing, mucking, and supporting of the development faces. The development activities are carried out by the development crew. Development operation is undertaken using four twin boom jumbos. The mine development cycle is optimised in such a way that a specific Jumbo performs drilling activity, while others perform ground supports. The drilling cycle takes approximately 2 to 3 hours per development face. Once drilling is completed, charging of the face is completed by the charge up crew using a MacLean machine and it takes 1 to 2 hours to load the development round.

The stoping cycle consists of cable bolting, drilling, charging, firing, and mucking of the stope to ore passes or to the truck. The cable bolting usually takes place prior to the drilling activity and consists of supporting the stope crown, hanging wall or stope brow. The cable bolting and grouting is done at a rate of 120 to150 m/day. Once the cable bolting is complete, the production drilling starts. Drilling is carried out using five production drilling rigs at a rate of 250 to 300 m/day.

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At the end of the stope drilling, the production hole charging is completed by the production charge up crew using the charge up machine.

Both the development and the stope cycle use emulsion for charging the holes. Due to the size of the stope and the firing sequence, electronic detonators are used in stope blasting. The blasting of the development heading and production stope take place at the end of the shift.

The loading operation is undertaken using 11 loaders, split between development, stope production, and haulage level loading to the crusher.

Hauling is completed using seven trucks. The hauling operations consist of hauling development and stope material to ore passes or to the surface waste dump or ROM pad.

The shaft is used for hoisting most of the underground ore material, while trucks are mainly used for hauling the underground waste material, which is the stope material located in the shallower part of the mine or for internal trucking to the ore passes.

Geomechanics

Since the completion of the Geomechanical study completed by SRK in 2011 (SRK, 2011), mining rock mass and structural models, stress measurements, and empirical and numerical analyses have been completed. The completion of this work gave a significantly improved understanding of the expected geotechnical environment. In line with this compilation of work, the mine adopted the Western Australian Mines Safety and Inspection Regulations 1995, in particular, compliance with Regulation 10.28 (Western Australian Department of Mines and Petroleum, 1995), thereby ensuring adequate geotechnical consideration when planning, designing and operating the mine. This approach continues today.

A number of underground mining geotechnical assessments for Kibali have been undertaken through several different external consultancy companies, namely Dempers & Seymour (Dempers & Seymour, 2021, 2014, 2015, 2017, 2018) Coffey Mining (Coffey Mining (2012, 2014), Beck Engineering (2014, 2015, 2017, 2018a, 2018b), Applied Geomechanics Consulting (Applied Geomechanics Consulting, 2020), KSCA Geomechanics Pty Limited (KSCA Geomechanics Pty Ltd, 2012, 2016, 2017, 2018) and the Western Australian School of Mines (2011, 2012, 2020). Combined, these consultancy companies have covered and completed different geotechnical aspects of work. These are summarised as follows:

●

Development and construction of mining rock mass and geotechnically significant structural models between 2012 and 2021. These models have incorporated raw geotechnical data comprised of rock mass and structural logs of exploration, geotechnical drill holes, grade control diamond holes and underground mapping.

●

3D mine wide numerical modelling, completed as a means of understanding the resultant mining induced effects from proposed mining sequences in terms of stress effects, damage and seismic potential in relation to the different ore zones and surrounding infrastructure. This approach of numerical modelling is considered to be a best practice method to assist in improving the understanding of the various mining induced interactions, testing new concepts and designs, which aid in improving the overall confidence to continue to successfully operate the underground mine.

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In situ pre-mining stress measurements – Five measurements covering depths below surface from 220 m to 590 m. Further additional stress measurements took place in 2020 at respectively 650, 850 and 1140mbs.

●

Established a Stope Performance Database to allow for the collection of relevant geotechnical stope design parameters, and then enable a comparison to be made between predicted and actual stope behaviour for each stope (considering each of the five stope surfaces).

●

Establishing the Kibali Stope Performance in relation to the Stability Graphs. This work is ongoing with the aim to develop site specific stability graphs with optimised Stable, Transition and Caving Zones (to allow for optimisation and refinement of stope designs). Results to date are presented in terms of the predicted performance for each individual stope surface compared to the actual performance of each stope surface.

●

Damage mapping undertaken by the Geotechnical Engineers is continuously updated. The resultant data from this mapping are being used to calibrate the numerical model, as well as an aid to understanding mining induced stress effects and the response of the rock mass to development and production mining.

●

Installation and operation of the underground seismic system (Phase 1 and 2). Similar to the damage mapping, the resultant seismic data are used to calibrate the numerical model. While the seismic potential for Kibali does not appear to as be high as it is in Western Australia for example, seismic monitoring is still necessary as a normal part of seismic risk management.

With underground stope production starting in December 2014 (and the subsequent voids backfilled with paste), assessing the geotechnical aspects of the underground mine has continued over the recent years. This will continue as experience is gained from understanding how the rock mass responds overtime to the mining induced effects from stope production. This work will be particularly pertinent if there are any future changes to the geology block model that then requires modifications to stope shapes and the subsequent extraction sequence.

As with previous Mineral Reserve updates for Kibali, LOM sequence numerical modelling has been used to confirm that new stope shapes and extraction sequence are ‘fit for purpose’ as part of the Mineral Reserve process. The modelling is seen as an important means of providing a quantification of stability and performance of stope extraction and identify stope extraction optimisation potential. The numerical modelling is also a method from which risk can be gauged, understood and mitigated.

This approach continued for the 2021 Mineral Reserve update, with the work again being undertaken externally by Applied Geomechanics Consulting (Applied Geomechanics Consulting, 2020). This numerical modelling consisted of a base case simulation and an additional simulation that incorporated modifications to the sequence as a means of managing and mitigating mining induced effects.

Ground Support

As per the feasibility study, the ground support systems used at Kibali are typically made up of reinforcing elements (such as end-anchored rock bolts, grouted tendons or friction rock stabilisers) that act directly upon the rockmass to increase its inherent strength. In addition to this, fabric (mesh) or coatings (shotcrete) are utilised to contain any potentially unstable rockmass between the reinforcing units.

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The fundamental purpose of the ground support system, as applied to Kibali, is to maintain the integrity of the rockmass under static load conditions for the expected service life of the underground operation. The suggested ground support regime adopted at Kibali takes due cognisance of several factors including:

●

The expected service life of excavations

●

Geology

●

Hydrogeological conditions

●

Rockmass classification data (Barton’s Q Classification)

●

Structural analysis

●

Seismic considerations

●

Perimeter control

Using the empirical design methodology known as the Norwegian Tunnelling Index method, which is based on Barton’s Q classification, the suggested ground support categories are summarised in Table 16-26.

Table 16-26 KCD UG Support Categories and Classifications for short life openings (<5 years)



Projected life of the opening: < 5 years (short life)

Span range	< 6 m (Standard development)	> 6 m (Wide span)

Q range	0.4 - 4	4 - 1000	0.4 - 4	4 - 1000

Rock Class	Very Poor, Poor, all Carb. Shale	Fair and better	Very Poor, Poor, all Carb. Shale	Fair and better

Ground Support Category	A	B	C	D

Primary Ground Support

Surface support	50 mm Fibrecrete	Mesh	50 mm Fibrecrete	Mesh

Rock bolt	2.4 m galvanised splitsets		3 m galvanised splitsets

Bolting pattern	1.5 m x 1.5 m pattern

Secondary Ground Support

Cable bolt intersections as per Table 16-27 and Table 16-29	Cable bolt backs as per Table 16-27 and Table 16-29

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Table 16-27 Cable Bolt Requirements for Intersections in General / Short-Term Excavations



D¹ (m)	M² (t)	No. of twin c/b FOS = 1.2	No. of twin c/b FOS = 1.5	L³ (m)

6	79	2	2	5

7	126	3	4	5.3

8	188	5	6	5.7

9	267	6	8	6

10	367	9	11	6.3

11	488	12	15	6.7

12	633	15	19	7

13	805	19	24	7.3

14	1006	24	30	7.7

15	1237	30	37	8

Notes:

D = Diameter of the inscribed circle

M = dead weight of potentially loose rock mass above intersection

L = minimum length of cable bolt

Table 16-28 KCD UG Support Categories and Classifications for long life openings (>5 years)



Projected life of the opening: > 5 years (LOM)

Span range	< 6 m (Standard development)		6 m (Wide span)

Q range	0.4 - 4	4 - 1000	0.4 - 4	4 - 1000

Rock Class	Very Poor, Poor	Fair and better	Very Poor, Poor	Fair and better

Ground Support Category	E		F

Primary Ground Support

Surface support	50 mm Fibrecrete	100 mm Fibrecrete

Rock bolt	2.4 m fully encapsulated bolts	3 m fully encapsulated bolts

Bolting pattern	1.5 m x 1.5 m pattern

Secondary Ground Support

Cable bolt intersections Table 16-27 and Table 16-29	Cable bolt backs as per Table 16-27 and Table 16-29

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Table 16-29 Cable Bolt Requirements for Intersections in Long-Term Excavations (Pump Stations, Crushers, etc)



D¹ (m)	M² (t)	No. of twin c/b FOS = 1.2	No. of twin c/b FOS = 1.5	L³ (m)

6	119	3	4	6

7	189	5	6	6.5

8	281	7	8	7

9	401	10	12	7.5

10	550	13	16	8

11	732	18	22	8.5

12	950	23	29	9

13	1208	29	36	9.5

14	1509	36	45	10

15	1856	45	56	10.5

Notes:

D = Diameter of the inscribed circle

M = dead weight of potentially loose rock mass above intersection

L = minimum length of cable bolt

Underground Mining Fleet

The underground equipment consists of mainly development drills, production drills, trucks, loader, loader setup on Sandvik multi lite remote-control system, and loader setup on Sandvik automation control system. The list of underground equipment is presented in the Table 16-30.

Table 16-30 Kibali Underground Mining Equipment



Manufacturer	Model	Type	Number

Sandvik	TH551	Truck	6

Sandvik	LH621	Loader	12

Sandvik	LH410	Loader	1

Sandvik	DL421	Drill	4

Sandvik	DD421	Drill	4

Sandvik	DS421	Drill	2

All light	A9	Light Plant	2

ASOKE	BUS	Light Vehicle	3

TOYOTA	Hilux	Light Vehicle	14

TOYOTA	Land cruiser	Light Vehicle	40

CAT	140K	Grader	1

CAT	TH414	Tele Handler	1

CAT	930K	Integrated tool carrier	3

VOLVO	L120GZ	Integrated tool carrier	1

CAT	420K	Backhoe	1

TCM	FORKLIFT	Forklift	1

MANITOU	MHT10220	Tele Handler	1

NORMET	Spraymec	Shot Crete machine	1

NORMET	Trans mixer	Mixer truck	1

MACLEAN	SL3	Scissor Lift	1

MACLEAN	EC3	Charge Machine	4

MACLEAN	BT3	Flat bed	2

MACLEAN	TM2	Mixer	1

BTI	MRH	Rock Breaker	1

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16.5

Underground Mining Services and Infrastructure

Mine Access

The mine is accessed via a twin decline, a vertical shaft, and a system of internal ramps. In general, mining zone access (to zones A, B, C, D, E and F) consists of six internal ramps. The underground mine internal ramp system connects various mining lodes.

Material Handling

The key components of the materials handling system are:

●

Teleremote (and manually operated) loaders tramming from the stopes and development faces to ore pass finger raises on that level.

●

Eight raise bored ore passes with finger raises on production levels.

●

Haulage level (210 level) – up to three remote operated automated loaders tram ore from the passes to two coarse ore bins.

●

Two crushers.

●

Two fine ore bins.

●

Conveyor transport of ore from crushed ore bin to shaft loading pocket.

●

Shaft haulage (740 m deep).

●

Headframe ore bin.

●

Conveyor haulage from shaft to process plant (including facility to place waste on an interim stockpile).

The materials handling system is also supplemented by truck haulage up the decline to the ROM pad at the process plant.

Backfill

The paste backfill plant treats the Kibali tailings from the flotation circuit by de-watering processes (filtration) to produce a 73% to 76% solids (by weight) paste containing binder, which is delivered to underground stopes under gravity or using a Putzmeister pump via a distribution piping system. The paste plant has been designed to treat a feed rate of 292 tph of dry tailings solids and produce nominally 190 m³/hr of paste fill. The tailings slurry from the tailings thickener reports to the paste plant at nominally 50% solids by weight.

The paste system and backfill plant consist of the following:

●

Tailings transport system

●

Filter Feed Surge Tank

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Belt Filtration

●

Conveying

●

Paste Mixing

●

Cement

●

Paste Tailings

●

Underground Reticulation

As described, the underground mining method is top-down long hole stoping with cemented paste fill. The extraction sequence required both horizontal and vertical exposure of the paste filled stopes.

The paste backfill underground reticulation is shown in Figure 16-33

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Source: Kibali Goldmines, 2021

Figure 16-33 LOM Paste Reticulation

Numerical analyses have been performed by Mining One^® to assess the stability of the paste fill exposures. It was recommended that a factor of safety of 1.3 is applied to the strength estimate to cover variability in material properties, mixing and placement and to allow a sufficient safety margin for the non-entry mining method (Mining One, 2017). For a multi-lift scenario, a factor of safety of 1.3 was applied. Therefore, the strength assumed in the LOM plan is based on the paste UCS matrix, and the age of the paste as represented in Table 16-31. An UCS of 910 kPa is assumed for vertical exposure and 150 kPa for the cap of secondary stope.

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Table 16-31 Summary of Paste UCS



Binder (%)	Age (days)	Slag Cement 90/10 Solid Percentage
Binder (%)	Age (days)	73.5%	74.0%	74.5%	75.0%	75.5%	76.0%	76.5%	77.0%	77.5%	78.0%	78.5%	79.0%	80.0%

1%	3		85	105	138	152	179	211	103	208	122	201	206	153
	7								213	354	151	266	203	194
	14	112	131	158	160	188	214	224	190	431	208	336		216
	28	144	159	180	188	209	230	247	339	387	388
	56	187	201	219	241	265	288	316
	90	128	131	166	205	231	246	311
	112	122	128	161	205	233	243	310
	120
	180	119	122	146	172	191	231	299	-	-	-	-	-	-
	240	66	-	90	97	115	132	175	-	-	-	-	-	-
	365	-	85	88	94	108	-	-	-	-	-	-	-	-

1.5%	3	-	121	143	174	195	222	259	133	207	165	220	224	222
	7	-	-	-	-	-	-	-	237	408	322	323	297	340
	14	150	194	227	266	303	341	385	352	492	418	450	-	368
	28	201	251	275	311	332	370	412	294	442	-	-	-	-
	56	211	239	299	321	342	388	432	-	-	-	-	-	-
	90	199	204	265	301	312	368	402	-	-	-	-	-	-
	112	139	186	256	297	311	358	388	-	-	-	-	-	-
	120	-	-	-	-	-	-	-	-	-	-	-	-	-
	180	101	128	142	156	159	188	200	-	-	-	-	-	-
	240	96	#DIV/0!	132	146	160	165	183	-	-	-	-	-	-
	365	-	99	-	123	131	147	203	-	-	-	-	-	-

2%	3	145	163	199	211	242	261	287	262	268	197	352	343	226
	7	-	-	-	-	-	-	-	391	570	479	478	421	379
	14	158	211	244	281	312	354	387	567	680	522	634	-	532
	28	213	240	268	286	361	388	406	516	555	-	-	-	-
	56	287	339	370	399	429	469	506	-	-	-	-	-	-
	90	281	315	358	371	411	457	502	-	-	-	-	-	-
	112	259	311	350	366	402	452	501	-	-	-	-	-	-
	120	-	-	-	-	-	-	-	-	-	-	-	-	-
	180	198	225	256	336	377	398	414	-	-	-	-	-	-
	240	145	172	192	240		350	387	-	-	-	-	-	-
	365	98	124	135	157	162	204	224	-	-	-	-	-	-

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Binder (%)	Age (days)	Slag Cement 90/10 Solid Percentage
Binder (%)	Age (days)	73.5%	74.0%	74.5%	75.0%	75.5%	76.0%	76.5%	77.0%	77.5%	78.0%	78.5%	79.0%	80.0%

2.5%	3	209	230	244	256	289	301	321	213	213	297	343	387	305
	7	-	-	-	-	-	-	-	453	453	559	530	605	508
	14	315	545	604	705	742	765	782	511	511	599	708	-	703
	28	540	691	705	841	903	913	953	590	590	-	-	-	-
	56	509	599	659	810	907	938	999	513	513	-	-	-	-
	90	507	611	674	833	914	948	1,001	-	-	-	-	-	-
	112	468	587	640	742	779	780	961	-	-	-	-	-	-
	120	-	-	-	-	-	-	-	-	-	-	-	-	-
	180	399	472	557	590	653	684	744	-	-	-	-	-	-
	240	292	371	388	413	486	600	669	-	-	-	-	-	-
	365	215	226	248	267	289	305	336	-	-	-	-	-	-

3%	3	197	200	399	402	470	560	628	343	343	459	399	-	378
	7	-	-	-	-	-	-	-	476	476	713	608	671	666
	14	498	524	602	615	766	984	1,002	519	519	849	840	-	948
	28	640	729	742	840	998	1,165	1,350	755	755	-	-	-	-
	56	920	934	1,040	1,188	1,200	1,213	1,425	704	704	-	-	-	-
	90	1,048	1,131	1,204	1,284	1,320	1,457	1,218	-	-	-	-	-	-
	112	1,051	1,260	1,302	1,303	1,336	1,353	1,399	-	-	-	-	-	-
	120	-	-	-	-	-	-	-	-	-	-	-	-	-
	180	963	1,020	1,072	1,107	1,171	1,225	1,269	-	-	-	-	-	-
	240	589	641	745	784	845	937	1,041	-	-	-	-	-	-
	365	454	552	661	705	705	752	797	-	-	-	-	-	-

4%	3	280	389	444	541	634	708	-	354	354	508	444	-	458
	7	-	-	-	-	-	-	-	555	555	908	892	819	752
	14	410	641	687	711	788	804	900	943	943	1,248	-	-	-
	28	541	691	700	780	901	987	1,011	1,181	1,181	-	-	-	-
	56	1,258	1,282	1,302	1,390	1,396	1,493	1,612	1,339	1,339	-	-	-	-
	90	1,369	1,399	1,444	1,563	1,598	1,681	1,664	-	-	-	-	-	-
	112	1,374	1,400	1,402	1,463	1,578	1,678	1,742	-	-	-	-	-	-
	120	-	-	-	-	-	-	-	-	-	-	-	-	-
	180	1,202	1,329	1,377	1,411	1,508	1,610	1,652	-	-	-	-	-	-
	240	1,117	1,253	1,349	1,417	1,453	1,497	1,583	-	-	-	-	-	-
	365	816	982	1,068	1,197	1,239	1,248	1,300	-	-	-	-	-	-

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Binder (%)	Age (days)	Slag Cement 90/10 Solid Percentage
Binder (%)	Age (days)	73.5%	74.0%	74.5%	75.0%	75.5%	76.0%	76.5%	77.0%	77.5%	78.0%	78.5%	79.0%	80.0%

4.5%	3	497	527	581	619	655	697	-	374	374	425	524	-	549
	7	-	-	-	-	-	-	-	493	493	826	950	-	890
	14	611	745	805	888	921	994	1,040	986	986	1,179	-	-	-
	28	988	1,310	1,469	1,482	1,520	1,611	1,791	1,338	1,338	-	-	-	-
	56	1,005	1,248	1,439	1,468	1,479	1,683	1,879	1,709	1,709	-	-	-	-
	90	1,383	1,432	1,493	1,590	1,683	1,757	1,953	-	-	-	-	-	-
	112	1,111	1,314	1,465	1,508	1,523	1,720	1,897	-	-	-	-	-	-
	120	-	-	-	-	-	-	-	-	-	-	-	-	-
	180	1,109	1,305	1,460	1,586	1,663	1,696	1,736	-	-	-	-	-	-
	240	1,002	1,263	1,359	1,490	1,467	1,491	1,592	-	-	-	-	-	-
	365	825	918	1,010	1,089	1,284	1,371	1,409	-	-	-	-	-	-

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Hydrogeology

Conceptual Permeability Framework

In 2021, SRK concluded the main geological elements contributing to the permeability framework are as follows:

●

Foliation-parallel fractures: foliation is very planar in high strain zones. A review of photos of core from 16 drillholes suggest that fractures with aperture parallel to the foliation do occur but are distributed on the order of metres. Whether or not the sheared contacts of dykes and alteration zones with significant competency contrast, contribute a down-dip fracture permeability fabric is not possible to say. If so, it would not be unreasonable to expect that some joints detach into these surfaces, perhaps allowing groundwater to collect and percolate down- dip.

●

Joints and minor faults: these structures are clearly the most abundant structures in the mine exposures and drillholes. However, overall rock quality is excellent in the drill core, with relatively low frequencies of fractures or faults affecting the rock mass. They are likely to form a tortuous connectivity pathway but may contribute to the groundwater permeability. The main N-S orientation of the joints is likely to influence the overall preferred groundwater circulation pathway.

●

Major structures: in general, it is believed that the major modelled structures (D1, D2 shears, late brittle reverse faults) do not contribute greatly to the deposit-scale groundwater flow. They are more likely, from a geological perspective, to behave as conduits locally over short distances where irregularities in their structures persist (i.e., local bends, branch-lines etc.).

●

Dykes: foliation-parallel dolerite sheets do not appear more preferentially fractured than other lithologies at KCD. Despite the limited observations, it is not inconceivable that certain portions of the intrusives are affected by fracturing due to competency contrasts.

●

Matrix permeability: large volumes of the deposit are strongly altered, including the alteration by silicification, which may have occluded any remnant porosity in the rock mass. Therefore, matrix permeability is likely to be negligible throughout the mine area.

In addition to the geological features that may form a connected network of permeable structures, SRK believes that an important non-geological contributor to the permeability framework are ungrouted exploration drillholes (SRK, 2021).

It follows that the permeability framework comprises of both geological and non-geological features with joints and faults accentuating the N-S preferential connectivity and fractures parallel to foliation and dykes causing the partial compartmentalisation of flow of the groundwater. In addition to the permeability framework, Kibali also exhibits a depth dependent permeability distribution that is typical of tropical basement aquifers.

Conceptual Hydrogeological Model of the Northern KCD Area

A revised conceptual model for the Northern KCD area that considers both current (pre-mitigation) and future (post-mitigation) groundwater pathways. The revised conceptual model is presented as an annotated diagram (Figure 16-34) and is used as the basis for the proposed dewatering and depressurisation plan.

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Dewatering

Hydrogeological modelling and monitoring at KCD underground are managed by a hydrogeological team on site. SRK provides support and review on hydrogeological aspects.

A summary of the dewatering history at KCD is provided below for context:

●

The KCD open pit was dewatered through a combination of boreholes and sump pumping.

●

As underground workings developed, and the pit deepened, the majority of dewatering wells ran dry due to declining water levels, and there are now no active dewatering wells at KCD.

●

Permeable structures associated with dolerite dikes and ironstone formations has also presented an inflow risk, especially at the base of the shaft, prior to commissioning of the clean water pumping station.

●

The objective of the dewatering programme at KCD is full depressurisation of deep and shallow aquifers. This is being achieved by allowing grade control and exploration holes to flow when intersected (wherever possible).

●

Drain hole drilling has been undertaken during the development of the declines and from the base of the shaft to pre-drain permeable structures, avoiding uncontrolled inflows to development, stopes, and the southern haulage roadway. The drain hole drilling has also increased the clean water make and reduced dirty water make, allowing for excess capacity within the dirty water treatment system at the base of shaft pumping station.

●

Inflows from the open pits have caused temporary flooding of B zone and has delayed depressurisation of the upper aquifers above C zone. These issues have been mitigated by lining of KCD South with clay backfill and an HDPE liner and installation of a pumping system. A pumping system has also been installed in KCD North, keeping it almost dry even during the rainy season, but installation of an engineered liner is pending. Fines from the adjacent waste dump have been deposited in the base of the pit after grass growth was noticed, and evidence suggests they have formed a natural low permeability liner.

●

Dewatering flow rates are broadly in line with those predicted during with groundwater inflows being slightly lower and inflows from the open pit being higher than predicted.

●

The primary deviation from the Feasibility Study prediction relates to the shallow aquifers to north of KCD, which are discussed further below.

●

Despite rainfall levels remaining above average over the year, pump outputs from the decline area remained significantly below maximum capacity (Figure 16-35).

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Figure 16-35 Kibali Underground Water Flows 2016 to 2021

Underground Pumping System

Mine water is diverted to the main pump station at C 615 Level via B 580 pump station located in B Zone, and C 390 pump station located in C Zone with the pumping capacity of 60 L/s and 80 L/s, respectively. The maximum pump capacity in the decline section is 120 L/s at C 615 pump station, which is designed to handle inflows from C Zone, B Zone and A Zone, then a pump to surface clarifier. The surface clarifier recycles dewatered water.

In addition, there is also a pump station at the shaft which consists of two (2 x 1143.4 KW) of 240 L/s pumping capacity. This is located at the crusher level and is fed by the clarifier located in the production level. The main shaft pump station pumps water straight up to surface through the vertical shaft.

Currently, C Zone water can also be diverted towards the shaft via sump #13 which, is the lowest sump in the C Declines.

The water in the A and the D Zone is temporarily managed by a 20 kW Flygt pump and Fish tank equipped with a 90 kW Flygt pump, until the completion of the development of their respective pump stations.

The underground pumping infrastructure consists of the following:

●

C 615 Level main pump station: 120 L/s as maximum pump capacity with six 110 kW challenge pumps.

●

B 580 Level pump station: 60 L/s with two 90 kW Flygt pumps.

●

C 390 Level pump station: maximum pump capacity, 80 L/s with four 110 kW challenge pumps.

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A 270 and A 360 set in series: 80 L/s Stalker pump that pumps to the C 390 pump station.

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Off-shaft: temporary pump capacity of 16 L/s with two 110 W Scamont pumps on the production level to manage dirty water, and the main pump station of 240 L/s installed on the crusher level with two 1143 kW pumps for clear water through the clarifiers (one pump is backup).

The current average outflow rates for underground are as follow:

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The total outflow rate is 87 L/s or 315 m³/h from the Declines (615_Pump Station) and 15 L/s or 54 m³/h from the Shaft (Main pump station).

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The outflow rate has increased up to 115 L/s over the year, as result of direct and diffused recharge from rainfall since there is a clear fluctuation of monthly pumping rates and monthly rainfall. Hydrograph analysis indicates that approximately 30% of the total pumping rate is as a result of monthly rainfall.

The underground pumping system is shown in Figure 16-36.

Underground Water Services

The service water rate of use is approximately 45 L/s or 162 m³/h for the decline section and 5 L/s 18 m³/h for the off-shaft section. Both the decline and shaft discharged (post-clarifying process) are collected in the surface dams and provide water service supply to the mine.

The underground water services infrastructure flow sheet is presented in Figure 16-37.

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Figure 16-37 Kibali Underground Water Services Infrastructure Flow Sheet

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Ventilation

The ventilation of the underground workings is provided via four surface primary fans (4291,27 AHP) and four auxiliary fans (590,05 AHP) that provide approximately 3 million cfm (1,420 m³/s) to adequately ventilate the mine equipment and underground workings at a dilution rate of 0.05 m³/s. Table 16-32 presents the ventilation requirements per unit of equipment.

Table 16-32 Ventilation Requirement per Equipment



Equipment	Manufacture	Engine Rate (kW)	Ventilation Requirement (m³/s)

Truck	SANDVIK TH551	515	25.75

Loader	SANDVIK LH621	345	17.25

Truck + Loader		860	43

Charge up rig	MACLEANS AC3	150	7.5

IT	CAT 962H	172	8.6

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The fresh air supply consists of five intakes including, the West and East declines (5.5 m wide x 6.0 m Height, each), Central and Southern fresh airways of 4.1 m diameter each, the main shaft.

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The return airway consists of four Raises of 4.5 m diameter each, including Southern RAR, Central RAR, Northern RAR and Crusher RAR. These raises are equipped with a vertical Zitron fan (ZVN 1-36-800/8, 0°), providing a total airflow of approximately 1250 m³/s.

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Currently, the shaft return system is via the refrigeration raise of 4.1 m diameter and equipped with four 110kW ClemCorp fans mounted in parallel at approximately 170 m³/s. This is temporary until the change of the ventilation system when it will be converted into intake.

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In total, the mine has a capacity of 1,420 m³/s, which represents the airflow ratio of 12 tons of air per tonne of ore mined. The LOM ventilation network is presented in Figure 16-38.

There is a system of internal exhaust ventilation raises for exhausting return air from the different levels up to the main return airways. All the levels are connected with a series of internal raises and drop board regulators are installed in the vent access drive at different levels to allow regulation of the airflows in the levels.

Secondary fans are installed in the decline and in the levels to convey fresh air from the decline to the active headings through flexible ducts, which are connected to the fans.

Fifteen refuge stations are placed in strategic location underground to ensure safe areas for personnel in case of emergency.

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Maintenance and Communications

The underground workshops and facilities consist of a haulage level workshop, decline and surface workshop. An additional workshop is currently being developed in connector 4 for a machine daily service bay and will also include oil dispensing system & pumps, air and water services, parts storage containers, tool storage area, work benches, fuel bay with Sat Stat and a container office. Radio communications are mine-wide. Maintenance infrastructure currently consists of the following:

Haulage Level workshop:

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Two machine service bays.

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Air and water services.

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Bounded oil dispensing system (pneumatic pumps).

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20” Sea container for parts storage and offices.

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2 x fuel tanks with 4200L capacity.

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Work benches and tool locker.

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Small tooling and jacks and stands.

Decline workshop:

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Three machine daily service bays.

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Air and water services.

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Work benches.

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Parts lockers.

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Bounded oil storage area (maximum two drums).

Surface workshop:

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Five machine service bays.

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Five apron and awning for daily servicing.

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Two service pits.

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Two machine wash pad and sump.

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20/5-tonne overhead crane.

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Containerised oil dispensing system with 10,000 litre tanks and dispensing station.

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Oil/Water sump with oil separators.

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Hydraulic hose container including press.

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Separate fabrication shop with the ability to perform gas and manual metal arc welding, line boring, air arc gouging and general fabrication.

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Separate rebuild workshop to complete component rebuilds. Includes Air-conditioned 20” container for drifter repairs.

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Multiple tooling from hand tools to shop tools.

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Variety of jacking and blocking equipment.

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Power Supply

The current load draws between 300 and 360 A on the dual 95 mm² 11kV cables originating from a single feeder on the surface. There is a plan to integrate a second electrical supply from the surface to essentially split the growing power demand across two power supplies. This offers the benefit of reducing the load on the current feeder. The second feed via the shaft will create a secure supply of power to much of the system as the ability to switch between two sources of electricity throughout the mine is introduced.

With the 185 mm² cable link between the surface and the 350 L substation in place and 90% of the installation completed, the commissioning phase of stage 1 for the Kibali underground HV Reticulation Upgrade Project has been carried out in 2021.

The second stage of the upgrade will involve increasing the current size of the power cable from 95 mm² to 185 mm² between 610 L substation and the B1_Decline substation.

The electrical reticulation and infrastructure throughout the production and development mining areas of the Kibali underground mine has been under ongoing assessment following the release of the LOM design in 2021 (Figure 16-39). This has resulted in the necessity for a long-term plan to be implemented encompassing an upgrade to the existing 11 Kv reticulation for the purpose of expansion as the push for the A decline continues down toward the lower levels off the C decline.

The new mining area, namely the A, D, and F zones required additional eight to ten 2MVA substations. The D zone feed will require an upgrade to a larger size of at least 9 mm² up to 185 mm² cable.

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Figure 16-39 Kibali Underground Infrastructure LOM Electrical Reticulation (Looking NW)

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16.6

Life of Mine Production Schedule

Open Pits

Production Scheduling

The nine Mineral Reserve pit designs were scheduled with their respective updated block models with signed off cut-off grades in MineSched software for the 2021 LOM schedule and budget. The mine schedule was based on a marginal grade cut-off. Material classes for various material types were created with different grade categorisation of high, medium and low-grades. These categorisations were based on the grade and tonnage distribution of ore in each deposit.

The mine schedule was generated based on historic rainfall patterns and scheduled calendar days. With nine months of rainfall expected in the year, the monthly budget was aligned to this to cater for lost days as a result of heavy rainfall. New pits are usually brought into production between December and March so that the saprolite can be mined in dry periods. Sheeting (with fresh waste rocks) of haul roads, ramps, and pit floors is practiced where possible to keep haul trucks running in wet conditions.

Figure 16-40 shows analysis of rainfall pattern and lost hours from rain over an eight-year period. Mining operations are carried out seven days per week, three shifts per day, utilising four shift crews.

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Figure 16-40 Historic Rainfall Pattern and Lost Production Hours from Rain

Based only on Mineral Reserves, Kibali gold production is planned to be approximately 730 koz Au per year for 10 years as shown in Figure 16-41.

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Figure 16-41 Historical and Planned Kibali Gold Production

All Inferred Mineral Resources that may fall within any Mineral Reserve reporting areas have been treated as zero value and are thus excluded from the Mineral Reserve estimate.

Pit Sequencing

Changes were made to the sequencing of the Kibali open pits based on the revised optimisation following model changes and gold price sensitivity reviews for each deposit. Pamao pit, which was scheduled to commence in January 2022, was re-sequenced to March 2022 to allow adequate time for the completion of the RAP in nearby communities surrounding the pit. However, the ‘Pamao South’ deposit has also been added to the Pamao Pit and is being mined alongside the main Pamao deposit. With Pamao ending in 2025, the pit will be ready for preparatory works for tailings facilities. This has resulted in the deferral of mining in Megi-Marakeke-Sayi pit to 2025.

The open pit end of life is estimated for 2033 based on current Mineral Reserves, as shown in Figure 16-42 and Table 16-33.

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Figure 16-42 Kibali Open Pit Mining Rate

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Table 16-33 Open Pit Mining Sequence Over the Mineral Reserves LOM



Open Pit	2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	Total

Total OP Waste (Mt)	31.27	26.49	29.73	29.23	22.99	30.05	26.38	28.14	29.10	30.74	29.09	5.93	319

Total OP Ore (Mt)	5.57	2.42	2.79	3.26	2.99	3.30	3.71	2.69	2.59	3.18	3.01	1.44	37

Average S/R (Waste:Ore)	5.6	11.0	10.6	9.0	7.7	9.1	7.1	10.5	11.2	9.7	9.7	4.1	8.6

OP Grade Mined (g/t Au)	1.85	2.15	2.19	2.24	2.48	2.47	2.82	3.68	2.90	2.33	2.29	2.47	2.44

Total OP Ounces (Moz Au)	0.33	0.17	0.20	0.23	0.24	0.26	0.34	0.32	0.24	0.24	0.22	0.11	2.9

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Underground Mine

Production Schedule

The underground LOM plan is based upon Proven and Probable Mineral Reserves and is scheduled using Datamine 5DP / EPS software.

It is a task-based dependency schedule:

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Each task in the schedule is assigned a duration or a rate. Durations are used for time dependant activities and rates are used for productivity dependant activities.

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Resources are assigned to tasks and their capacity is profiled over the LOM to produce a practical schedule.

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EPS Optimisation Tool (EPSOT) is used for the LOM schedule optimisation.

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EPSOT is a scheduler software which optimises the schedule based on constraints provided by the user. Its objective is to maximise Net Present Value (NPV).

As part of the scheduling process, attributes are assigned to each of the tasks, and these are used for interrogating and reporting the schedule.

The underground physicals by year are shown in Table 16-34.

Based only on Mineral Reserves, the KCD underground is planned to sustain a production rate of 3.6 to 3.8 Mtpa for 10 years, tapering off to 3.3 Mtpa in year 11 and 2.5 Mtpa for the last two years.

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Table 16-34 Kibali KCD Underground LOM Physicals Based on Mineral Reserves



Underground	2022	2023	2024	2025	2026	2027	2028	2029	2030	2031	2032	2033	2034	Total

Total UG Ore (Mt)	3.63	3.83	3.86	3.89	3.83	3.90	3.90	3.88	3.76	3.17	2.91	2.87	2.38	46

UG Grade (g/t)	5.30	5.34	5.18	5.02	4.84	4.56	4.18	4.08	3.86	4.17	4.16	4.80	3.14	4.54

Total UG Ounces (Moz Au)	0.62	0.66	0.64	0.63	0.60	0.57	0.52	0.51	0.46	0.45	0.39	0.40	0.24	6.7

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Open Pit and Underground

Figure 16-43 shows the combined open and underground LOM feed.

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Figure 16-43 Kibali Open Pit and Underground LOM Feed Schedule Based on Mineral Reserves

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17	Recovery Methods

17.1

Processing Plant

Kibali ore is blended using both KCD underground ore and ore sourced from satellite open pits. The process plant has been treating KCD underground ore since 2015 and has demonstrated reasonably consistent recovery performance. The flow sheet comprises crushing, ball milling, classification, gravity recovery, a conventional CIL circuit, flash flotation, and conventional flotation, together producing a concentrate which goes to ultra-fine-grinding, and a dedicated intensive cyanide leach circuit. This process consists of industry standard technology and is appropriate for the style of mineralisation present at Kibali.

The Kibali gold processing plant comprises two largely independent processing circuits, the first one designed for oxide, transition, and free milling ore sources and the second for sulphide refractory ore. However, both circuits are designed to be switched to process sulphide ore when the oxide, transition, and free milling ore sources have been depleted. A simplified flowsheet can be seen in Figure 17-1.

The oxide ore is recovered through a standard crushing, milling, and gravity plus CIL operation.

The sulphide ore requires crushing; milling; flotation; UFG; a Pumpcell circuit preceded by a three-tank gravity flow pre-oxidation circuit to passivate cyanide consuming sulphides as well as liberate the gold. The first two tanks are subject to highly intensive oxidation with cyanide being introduced into the third to fifth tanks for pre-leaching, where the resultant product gravitates to a Pumpcell Carbon-in-Pulp (CIP) circuit with high concentrations of activated carbon. The Pumpcell residue stream may still contain some residual gold, which is then pumped to the main CIL circuit for final leaching to scavenge the remaining leachable gold. The flexibility of the plant design allows for an extended pre-oxidation and pre-leach step within the CIL occurring after the initial pre-oxidation circuit but prior to the stream being routed to the Pumpcell circuit.

Most of the deposits contain some extent of free native gold, which means it is large enough to recover via a density separating step, which is performed with Knelson gravity concentrators during the milling cycle.

The processing plant rated throughput is 3.6 Mtpa of soft oxide rock ore through the oxide circuit and 3.6 Mtpa of primary sulphide rock ore through a parallel sulphide circuit. Once the plant is sulphide only, the capacity is 7.2 Mtpa of sulphide ore. Kibali’s operational performance has demonstrated that the process plant is fully capable of its design capacity, and further modifications to the crushing circuit with finer F80 (38mm) coupled with a decreased inlet trunnion size has allowed for an even greater power draw and hence higher throughputs.

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The constant improvement in terms of the plant utilisation and availability is mainly driven by regular planned maintenance coupled with good performance of process operations. Plant utilisation and availability from 2013 to 2021 is presented on Table 17-1.

Table 17-1 Plant Availability and Utilisation



Years	2013	2014	2015	2016	2017	2018	2019	2020	2021

Availability (%)	74.9	87.0	93.6	94.7	96.4	95.2	95.6	94.9	95.4

Utilisation (%)	64.9	93.1	95.9	98.0	98.6	98.8	98.8	99.5	99.1

The process plant specific energy consumption has been reduced from 28.5 KWh/t in 2015 to 25.8 KWh/t in 2021 (Figure 17-2) as a result of crushing the KCD UG plant feed to P₈₀ below 50 mm. This is forecast to continue at these levels for the remaining LOM based upon Mineral Reserves. The Kibali hydropower system has a peak capacity of 42.8 MW and 43 MW of thermal generation full redundancy. Actual hydro generation capacity is season dependent. The total load demand of the mine is not constant, power demand at full production ranges between 39 MW and 43 MW, currently averaging approximately 41 MW, which in the QP’s opinion is well resourced to cater for the LOM process plant power demands based upon Mineral Reserves.

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Figure 17-2 Kibali Processing Plant Specific Energy Consumption 2015 to 2021

The plant water demand has been stabilised at an average of 9.9 Mm³ per annum for the past three years with a specific water consumption of 1.30 m³/t (Figure 17-3 and Figure 17-4). The remaining LOM is forecast to continue at these levels based upon Mineral Reserves.

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Figure 17-3 Kibali Processing Plant Water Demand 2013 to 2021

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Figure 17-4 Kibali Processing Plant Specific Water Consumption 2013 to 2021

The processing plant has a total of 335 employees, assisted by 164 contractors, whose responsibilities are split as per below:

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Paragon: Tailings Storage Facilities (TSF) management.

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Lutula & Munguleni: Plant cleaning.

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Air Liquide: Oxygen plant production and maintenance.

The plant has the capacity to make the stated throughput based on historic results for oxides and sulphides. Figure 17-5 attests to the continued improvement and consistently stable throughput that is well beyond design capacity. No fatal flaws have been determined.

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Figure 17-5 Kibali Processing Plant Performance Tonnes Treated 2013 to 2021

The oxide circuit has the following processes:

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Primary crushing.

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An optional secondary hybrid roll type crusher for the harder transitional and free-milling sulphide ores.

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Milling.

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Cyclone classification.

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Gravity concentration.

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Flash flotation – runs optionally when the feed blend is predominantly free milling fresh ore.

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CIL.

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Tailings disposal.

The sulphide circuit has the following processes.

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Primary and secondary crushing.

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Milling.

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Cyclone classification.

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Gravity concentration.

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Flash flotation.

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Conventional flotation.

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Ultra-fine grinding of the concentrates.

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Pre-oxidation circuit.

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Extended intensive oxygenation assisted leach.

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Pumpcell adsorption circuit to recover gold from the concentrates.

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Tailings disposal.

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The loaded carbon from the Pumpcell circuit, that is, from the concentrate leach and CIP, together with carbon from whole-ore leach, are treated in independent elution circuits, followed by electro-winning of gold eluate.

Once the oxide, transition and free-milling ore sources have been depleted, the existing oxide plant can be converted to a parallel sulphide circuit, which will necessitate the expansion of the concentrate handling and Pumpcell circuits. There are two flotation circuits already present in the plant.

Kibali Goldmines further expanded the original fine-grind section in the 2017 sulphide expansion project by adding an additional four ultra-fine-grind mills, making eight in total. The current Kibali feed plan allows for an oxide – sulphide campaign for thirty percent of the year, with the remainder of the year treating full sulphide ores.

Sulphide Crushing and Screening

Two primary jaw crushers (two Sandvik CJ815:200 kW, CSS:16 0 mm) are used targeting 1,300 tph and feeding two secondary crushers (two Sandvik CS660; 250 kW, CSS:45 mm) via a coarse ore stockpile (COS).

ROM sulphide ore, received from trucks, is treated in a primary crushing circuit comprising of a ROM bin, apron feeder, and primary jaw crusher (Sandvik CJ815) operated in open circuit at 1,300 tph target. This primary crushed product is then conveyed to a primary crushed stockpile or COS with a 5,000 t live capacity. The sulphide ore from underground has already been crushed underground and is also conveyed to this stockpile.

Apron feeders under the stockpile are used to combine the sulphide ores from the two sources, before it is conveyed to the secondary crushing circuit that has two secondary cone crushers (Sandvik CS660) operating in parallel (running/standby) in open circuit to produce a crushed product stream with a P80 of 45 mm. The secondary circuit was commissioned in May 2014.

When sulphide ore is being treated, secondary crusher product is fed onto a fine ore stockpile (FOS) via a conveyor system. The FOS serves as a common mill stockpile to both the mills and has a live capacity of 11,700 t of sulphide ore to each mill. The mill is fed from the mill feed stockpile using apron feeders that feed directly onto the mill feed conveyor.

When oxide ore is being treated through its circuit, the primary crusher product (Sandvik CJ815) is fed directly to the mill feed conveyor and not via the secondary crushing circuit but may be subject to an in-line hybrid crushing stage, if deemed necessary, at least on one of the two parallel streams.

The design of the crushing circuit includes provision for the installation of a tertiary crushing circuit.

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Oxide Crushing and Screening

ROM ore received from trucks is treated in a primary crushing circuit comprising of a ROM bin, an apron feeder, and a single toggle jaw crusher. This primary crushed product (at 450 tph) can either be diverted to the primary mill feed conveyor when oxide ore is treated, or alternatively conveyed to a common 5,000 t live primary crushed stockpile when sulphide ore is treated.

Sulphide Milling and Oxide Milling

A ball milling circuit comprising two Polysius ball mills, each operating independently in parallel, treats ore at a total feed rate of 900 tph dry solids.

When treating oxide ore, the primary crusher product will feed directly onto the mill feed conveyor and into the milling circuit. When sulphide ore is treated, the mill will be fed from the mill feed stockpile.

Each ball mill is operated in closed circuit with a cyclone cluster used to produce a target grind of 80% passing 75 µm on sulphide and 80 µm on oxide. The mill feed consists of fresh crushed ore, a portion of the cyclone underflow, gravity concentrator scalping screen oversize and flash flotation cell high density tailings. Ground ore from the mill reports to the mill discharge sump where it combines with gravity concentrator tailings, flash flotation low density tailings and Gekko Inline Leach Reactor (ILR) tailings, before being pumped to the cyclone cluster.

When treating sulphide ore, the cyclone overflow is gravity fed to a rougher flotation circuit, whilst the overflow from the cyclones when treating oxide ore is directed to the CIL circuit. Cyclone underflow is split into three streams:

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Gravity concentration circuit.

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Flash flotation circuit.

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Remainder of the cyclone underflow is re-cycled to the mill feed.

Gravity concentrator tailings gravitates to the mill discharge sump, while the concentrate reports to a batch ILR circuit.

The flash flotation cell produces a concentrate and a high-density tailings stream. The concentrate is directed to the concentrate handling circuit, alternatively it can be deposited into the feed of the gravity recovery pre-screening to enhance gravity recovery, while the high-density tailings are circulated back to the mill feed. Frother, collector, and promotor are added to the flash flotation cell for recovery of flash flotation concentrates. The required copper sulphate conditioner is added into the mill feed.

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Flotation

Cyclone overflow from the primary milling circuit is routed to either the rougher flotation cells or bypasses the circuit to the rougher tailings tank before being pumped to the CIL circuit (if oxide or free-milling material is being treated).

Two separate parallel banks of six 70 m³ Outotec forced air rougher flotation cells in series are used for flotation; however, when one mill is processing oxide ore, then only one bank is required. Frother, collector, activator, and promoter are added to the slurry stream of sulphide ore for the recovery of a flotation concentrate. Rougher flotation concentrates from the first three tank cells in each bank are pumped to the concentrate handling circuit where it can combine with the flash flotation concentrates. The concentrate from the last three tank cells in each bank is recycled to the flotation feed. The reason for this split of concentrates is to not only reduce the overall mass pull for the limited capacity downstream concentrate treatment circuits, but also maximise potential flotation recovery by ensuring the limited mass pull comprises the higher grade concentrate emanating from the first three cells, whilst the lower grade concentrate from the last three cells is not lost as it is simply recycled for further processing. The rougher flotation tailings stream is pumped to the flotation tailings thickener. Occasionally, when warranted, either for reasons of maintaining stability of parameters within the main CIL circuit, or in the event of a low flotation recovery, the float tail is routed for leaching in the main CIL circuit, thereby facilitating further recovery of the residual gold in the CIL circuit.

Inline Leach Reactors (ILR)

Gravity concentrator concentrate is gravity fed into the reaction drum of the Gekko ILR (Inline Leach Reactor) from the feed cone. Sodium hydroxide, sodium cyanide, and oxygen are added in high concentrations to the reactor to place the gold into solution. At the completion of the leach, the pregnant leach solution is pumped to the ILR electro-winning pregnant solution tank.

Oxygen is supplied from a 30 tpd oxygen plant being upgraded to a 40 tpd pressure swing adsorption (PSA) plant operated by Air Liquide.

Concentrate Handling and Ultra-Fine Grinding

Flotation concentrates, together with gold-room waste, report to the concentrate thickener. Thickener underflow is fed to the ultra-fine milling circuit. The UFG consists of eight VXP2500 FL Smidth (originally Deswik) ceramic bead mills in parallel, where circuit feed material of 80% passing 40 µm is treated to achieve a target grind of 80% passing 23 µm.

The ultra-fine milling products are pumped to the pre-oxidation and pre-leach circuit followed by the Pumpcell circuit.

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Pre-Oxidation and Pre-leach

Slurry from the UFG circuit is fed to the Pumpcell circuit via the pre-oxidation and pre-leach circuits.

The pre-oxidation circuit consists of two agitated tanks operated in series. Flexibility exists to operate as the ore demands. Each tank is fitted with four (two duty and two standby) Aachen REA450 reactors, through which slurry is circulated and contacted with oxygen. Both pre-oxidation tanks also have three oxygen sparge units. Lime and lead nitrate are dosed in both pre-oxidation tanks. Hydrogen Peroxide dosage is also available. The product stream from the second pre-oxidation tank overflows to an agitated pre-leach tank, fitted with one Aachen REA400 reactor, with a dedicated Aachen reactor pump. The tank is also equipped with oxygen sparge units. Cyanide and lime are added to maintain the leach pH and hydrogen peroxide can also be dosed if required. A diesel dosing facility, originally installed to deal with preg-robbing from carbonaceous material has since been decommissioned and removed from the circuit. The pre-leach slurry product is pump fed to two 2,100 m³ leach tanks for extended residence time in an Aachen assisted leach environment.

Pumpcells

The pre-oxygenated and pre-leached product stream overflows to eight 100 m³ Kemix Pumpcell tanks operated in series. Before the concentrate expansion project, there were only six tanks. Eight tanks have proved sufficient for twin-stream sulphide operation, both on account of the mass-pull reduction initiatives introduced to the flotation circuit, but also because the Pumpcell circuit residue stream can be blended into the main CIL circuit to mop up any residual gold. The Pumpcell tanks are operated in a carousel mode, with counter current flow of carbon relative to slurry. Slurry is moved between tanks with MPS(P) screens, which both transfer slurry to the next sequential tank and screens out the carbon which remains in the original tank. The carbon in a Pumpcell circuit always remains in the same tank but the position of the tank in the circuit is changed. Two tanks are isolated each shift and the entire content of the tank is pumped to the elution circuit. It is then re-introduced to the circuit as the last tank and receives a fresh batch of carbon. This results in a loaded carbon batch size of 5 t/day. This process is also flexible where advantage can be taken of two tanks being harvested in a single day to completely fill an elution column.

Loaded Carbon and Tailings

Loaded carbon from the Pumpcell tank taken offline is pumped to a vibrating screen. The screen oversize reports to the elution circuit acid wash tank, while the undersize is routed to CIL Tank 3.

Tailings from the Pumpcell circuit exiting the last Pumpcell tank in the carousel are pumped to the CIL circuit Tank 2.

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CIL

Pre-Oxidation

Oxide material, bypassing the flotation cells, is pumped to CIL tank No.1 for pre-oxidation via dedicated Aachen reactors.

Lime is added to increase the slurry pH before cyanide is added. The leach pH level must be maintained at 10.5 or higher. Lead nitrate is also added. Hydrogen peroxide can also be used, if required. Cyanide is added in the circuit tank 2, which is Aachen assisted for locking on prior to gravitating or overflowing into the first carbon adsorption tank (CIL Tank 3).

The last two tanks in the CIL train have been dedicated to the extended pre-oxidation and pre-leaching of the concentrate stream prior to being routed to the Pumpcell circuit.

CIL Tanks

Slurry and activated carbon flow counter current to each other through CIL tank numbers one to nine. Slurry from the pre-oxidation tank flows by gravity to CIL Tank 2. This tank was converted into a pre-leach Aachen assisted leach tank, where cyanide is added prior to slurry overflowing into the first adsorption tank. Pumpcell tailings also report to the CIL circuit with the flexibility of being treated in a selection of CIL tanks as deemed appropriate for the current ore feed blend and configuration, this after passing through the Pumpcell tailings samplers. The slurry in Tank 1 is routed to Tank 2, and thereafter each subsequent tank, using MPS(P) interstage screens. The MPS(P) screens only transfers the slurry while retaining the carbon in the tank.

Fresh/Re-gen carbon is added to the final tanks in the train and is pumped from each tank to the preceding tank in the sequence using carbon transfer pumps. A total of 24 t of carbon is transferred daily from each tank. Carbon and slurry are transferred by a carbon transfer pump from CIL Tank 3 to the loaded Carbon screen at the top of the elution columns, where the carbon is screened out for elution. The screen underflow returns to CIL Tank 3.

The following reagents are dosed in the CIL tanks – oxygen, cyanide, lime, lead nitrate, and hydrogen peroxide (used when oxygen is not available).

Detox Tanks

Two detox tanks, each with two Aachen REA450 reactors, were designated to be used for cyanide destruction. Cyanide must be destroyed to below 50 ppm WAD at the point of discharge in the lined tailings dam as per ICMC requirements. The pulp in the detox circuit is typically contacted with activated carbon, which serves as a catalyst for the oxidation of cyanide to cyanate. These tanks proved to be redundant both in terms of their ineffectiveness coupled with an alternative and superior method of controlling the cyanide concentrations within the CIL circuit via a cyanide analyser and controller. Thus, they were transformed into normal CIL tanks, effectively lengthening the CIL train and hence extending the residence time, making up the end tanks of the main CIL circuit, well suited also for the dedicated treatment of extended pre-oxidation and pre-leaching of concentrate prior to being routed to the Pumpcell circuit.

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Elution

The CIL and Pumpcell carbon are batch treated in the Anglo American Research Laboratories (AARL) elution circuit separately through two identical circuits. The duplicate 12 t AARL columns share a common heater facility capable of running both columns simultaneously. The CIL carbon is to be treated in 12 t batches twice every 24 hours, while carbon from the Pumpcell circuit will be treated in 10 t batches every 48 hours or more frequently if desired. The elution heaters are electric whilst the regeneration kilns are diesel fired. With the improvement in the elution heating system changed from electrical to diesel and the need to improve carbon management in the circuit, both columns are running two 12 t batches every 24 hours. This has also taken away the need to discretely treat the CIL and Pumpcell carbon separately.

Loaded carbon is collected in an elution circuit acid wash tank. Carbon that has been acid washed is then loaded into a 12 t AARL elution column by gravity. The carbon has a clear eluate solution (1% NaCN and 3% NaOH at 125°C) pumped through it that desorbs the gold from the carbon and places into solution to form the loaded solution, which is then pumped to the elution electro-winning circuit feed tank/pregnant solution tank.

Barren carbon is removed from the elution column and reports to either of two carbon regeneration kilns.

Electro-Winning and Gold Room

Pregnant solution from the ILR circuit is circulated through a single electro-winning cell and steady head tank. Gold is deposited on the cathodes as sludge and the solution circulated until the desired barren gold concentration is achieved, tracked with continuous sampling and analysis of the circulating pregnant solution between electrowinning cells and the pregnant solution holding tanks.

Pregnant solutions from the CIL elution circuit and the Pumpcell elution circuit are treated in the same way, except that there are six electro-winning cells operating in parallel for each stream.

Loaded cathodes are periodically removed from the cells, the gold sludge is washed off using a high-pressure washer and the washed mixture is then decanted. The gold sludge left behind is calcined in two electric calcination furnaces. The calcined sludge is then mixed with fluxes and loaded into an induction smelting furnace. After smelting, the furnace crucible contents are poured into cascading moulds to produce gold bullion and slag.

General

The plant has two distinct processing streams that are largely separate in most areas. Hence a loss area in one will not affect all the output of gold in most loss scenarios. This includes the substations where, for example, each ball mill has its own substation. However, once the elution and carbon handling areas are reached, the production of gold is concentrated into single work areas.

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In general, there is a satisfactory level of protection against collision damage in the structures that support elevated structures. Bypass arrangements for tanks mean that if one unit is offline, the process flow does not stop. There are only minor impacts on recovery efficiency if only one tank or flotation cell is offline.

The process plant has been built on an existing site. The site has been cut largely from the top of hills to accommodate the present structures. The crusher, bin and mill were built on the old leach pad area, while the CIL tanks reside in the old pregnant solution pond area. Geotechnical boreholes and test pits were completed to design the terraces.

In summary, this is a new process plant that is meeting its throughput and recovery efficiency expectations. It can be considered as having matured to steady stated operations.

17.2

Processing Recovery

The actual process plant gold recovery in 2021 varied monthly from 87.8% to 90.5% (Figure 17-6 and Table 17-2). The average gold recovery in 2021 was 89.8%. Recovery for 2022 is expected to be 89.8%, averaging 89% for the LOM.

High GRG contribution of 24.17% compared to the forecast of 23% mainly driven the high GRG from Gorumbwa pit ores fed compared to previous years. The October low recovery of 87.8% was mainly due to the high residues emanating from the circuit changes between full sulphide and sulphide/oxide campaign treatment. The changeover often results in flashing out of high residues that build up in the CIL circuit coupled with reduced residence time during the changeover period.

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Figure 17-6 Kibali Processing Plant Overall Gold Recovery in 2021

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Table 17-2 Kibali Processing Plant Overall Gold Recovery in 2021 by Month



Item	Jan	Feb	Mar	Apr	May	Jun	Jul	Aug	Sep	Oct	Nov	Dec	2021 Total

Tonnes Treated (Dry) (kt)	678	622	686	642	695	654	656	638	644	633	611	625	7,783

Plant Head Grade (g/t Au)	3.32	3.27	3.41	3.59	3.35	3.63	3.85	3.77	3.58	3.97	4.00	3.77	3.62

Recovery (%)	90.3	89.7	90.2	89.9	90.3	89.6	89.2	90.5	90.4	87.8	90.3	89.5	89.8

The Kibali processing facility has largely seen improvements in its operational performance on a year-by-year basis. This performance extends to overall gold recovery (Figure 17-7) and throughput (Figure 17-8).

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Figure 17-7 Kibali Processing Plant Recovery

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Figure 17-8 Kibali Processing Plant Pumpcell Residue and Throughput

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17.3

Production History

The Kibali processing plant has been operating since 2013. The production history is summarised in Table 17-4.

Table 17-3 Kibali Processing Plant Production History



Year	Tonnes Milled (kt)	Grade (g/t Au)	Contained Gold (oz Au)	Recovery (%)

2013	808	3.87	88,199	91.5

2014	5,546	3.81	526,627	79.0

2015	6,833	3.55	642,720	83.8

2016	7,299	3.10	586,530	79.8

2017	7,621	2.87	596,226	83.6

2018	8,218	3.45	807,251	88.6

2019	7,513	3.80	814,027	88.7

2020	7,632	3.68	808,134	89.5

2021	7,783	3.62	812,152	89.8

Total	59,254	3.48	5,681,866	85.7

17.4

Processing Costs

Operating Costs (OPEX)

The 2019, 2020 and 2021 actual processing operating costs can be found in Table 17-4.

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Table 17-4 Actual Process and Plant Engineering Operating Costs for 2019, 2020 and 2021



Cost	Units	2019	2020	2021

Fixed Cost

Consultants	$ 000	537	307	256

Contractors – Assays	$ 000	1,676	1,639	1,680

Contractors – Oxygen	$ 000	109	193	-52¹

Equipment Hire	$ 000	2,597	1,898	1,815

General Costs	$ 000	10,396	11,038	13,789

Gold Refining	$ 000	3,443	5,817	5,870

Labour	$ 000	7,796	8,710	8,695

Stores – Other	$ 000	1,730	1,544	1,824

Total Fixed	$ 000	28,284	31,148	33,878

Tonnes Processed	kt	7,513	7,632	7,783

Total Fixed	$/t	3.76	4.08	4.35

Variable Costs

Power	$/t	3.27	1.93	2.12

Reagents – Cyanide	$/t	2.71	2.60	2.69

Reagents – Lime	$/t	1.03	0.65	0.65

Good Issues – Caustic Soda	$/t	0.44	0.41	0.43

Good Issues – Activated Carbon	$/t	0.12	0.11	0.11

Reagents – Other	$/t	1.36	1.39	1.33

Stores – Grinding Media	$/t	0.81	0.90	0.97

Stores – Liners	$/t	0.54	0.55	0.48

Stores – Screens and Panels	$/t	0.05	0.01	0.03

Total Variable	$/t	10.34	8.56	8.81



Total	$/t	14.11	12.64	13.16



Plant Engineering	$/t	3.11	3.22	3.31



Combined Process and Plant Engineering	$/t	17.21	15.86	16.47

Notes:

Included in this amount is a cumulative catchup amount related to an IFRS (International Financial Reporting Standards) 16 adjustment processed to account for the extension of a lease agreement, with the corresponding right of use asset being unwound in the income statement. The impact of the adjustment is not considered material.

LOM processing costs are modelled to be $17.49/t (which includes Plant Engineering cost). The actual costs for 2021 were $16.47/t, with the key improvements over the LOM as a result of:

Power cost due to the installation of a grid stabiliser and diesel heaters in the elution circuit. The first one has improved the hydro blend, while the second one has reduced plant power consumption from the elution circuit.

Lime cost due to the change from hydrated lime to quick lime, which has lower price and consumption over the LOM.

Optimisation of cyanide and caustic consumptions.

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However, the improvement in 2020 and 2021 has been partially offset by the increase in gold refining fees due to the lack of direct commercial flights from Nairobi to South Africa following Covid-19 travel restrictions.

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18	Project Infrastructure

18.1

Mine Roads

The Mine is located in the NE of the DRC. Access to site by road is via Uganda and the Ugandan border town of Arua. The road from the Ugandan border at Arua has been upgraded by the company to accommodate the Project and on-going operations traffic. Maintenance of this road is carried out by the Kibali Goldmines.

The local road infrastructure was developed during the exploration drilling programmes and upgraded during the construction of the mine. Internal roads provide access to various infrastructure areas, including roads to the TSF, Explosives Storage, Land Fill Site, Mine Villages, Central Mine Offices, Shaft Collar Area, Open Pit Mining Central Operations Area, general mining operations areas, new exploration areas, various water boreholes, and overhead line routes.

All roads are constructed by layered rock/gravel/laterite varying in specification according to traffic expectations.

18.2

Supply Chain

Since the Project’s inception, the majority of Kibali’s imports are shipped into the port of Mombasa, Kenya, and thereafter trucked through the Northern Corridor Road route that links Mombasa to the landlocked countries in Eastern and Central Africa. The cargo initially moves through Kenya and Uganda into Eastern DRC (to Kibali). Up to the Uganda / DRC border, the trucks use a two-way tarmac road considered to be the main route from the port of Mombasa to East and Central Africa. The final 200 km of the trip from the DRC border to Kibali is on laterite roads.

The primary ports for mining spares and consumables are Durban and Antwerp. Reagents, such as cyanide, steel balls, peroxide, hydrochloric acid, and other flotation reagents are shipped from a variety of different ports worldwide. The shipping terms for the mining consumables and reagents are typically Ex-Works or Free on Board and Cost, Insurance and Freight, respectively.

The costs associated with 20 ft and 40 ft containers, for both sea-freight and inland transport (Mombasa to the Mine site), are calculated on a cost-plus basis. This is a fully transparent exercise with shipping/freight invoices being sent through for verification.

Estimated port to port transit times for Kibali’s most frequent sailings:

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South Africa = 10 days

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Europe = 35 days

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China = 45 days

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USA = 65 days

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Procurement for Kibali Goldmines is carried out by its supply chain partner, Tradecorp Logistics.

18.3

Surface Water Management

Kibali lies within the northern tropical climatic region of the DRC. The area has a distinct rainy and almost dry season. The rainy season extends from March to November and the dry season from December to mid-February.

The Kibali River dominates the drainage of the Project area and flows along the southern boundary of the Project area. The Nzoro River flows into the Kibali River approximately 30 km downstream of the Mine site. Numerous springs exist in the area and the spring flows remain near constant throughout the dry season.

The significant sources of water that can affect the operations include rainfall directly into the open pits, rainfall surface run-off and groundwater entering the pits from the surrounding rock masses.

Surface run-off is high, due to high intensity rainfall events and an undulating landscape. A system of bund walls and dewatering trenches has been established prior to mining of each of the pits, which prevents inflow of surface water to the pit areas. The network of drainage channels is used to discharge water intercepted by the perimeter drains to the Kibali River via a series of settling ponds.

All the deposits are characterised by the presence of a near-surface groundwater table with the potential for high groundwater into the pits. The possible impacts of ingress of groundwater are investigated prior to mining and during the mining activities. Dewatering well systems are installed for all pits to lower the groundwater level prior to mining. A system of dewatering trenches is procedurally established prior to the commencement of mining in each of the pits, preventing the inflow of any surface water to the active mining areas.

The rainfall that falls within the pit perimeter is directed out of the pit, if this is possible, particularly in the upper levels. The water that cannot be directed outwards flows to the sump at the pit bottom from where it is pumped.

Figure 18-1 presents an overview of the Kibali Water Management Plan.

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18.4

Water Supply

Raw water is collected and stored in the raw water dam (RWD), which has a storage capacity of 16,000 m³. The primary sources of raw water are rain, spring water, Kibali River, and water from pit dewatering.

The processing plant requires 25,000 m³ of water per day, of this 75% is recycled water (from the TSFs) and 25% is from the Return Water Dam (RWD).

The plant process water circuit consists of a 25 m diameter process water clarifier and process water dam with a capacity of 4,600 m³.

The Process Plant has a Water Treatment Plant unit that produces soft water. That soft water is used in some strategic areas (Elution circuit, Laundries, Flocculent make-up, Firefighting system and Metallurgy Laboratory section). The operational camp has an independent water purification plant and storage facility.

18.5

Tailings Facilities

The cyanide containing TSFs comprise CTSF1 and CTSF2 for the CIL tails and the FTSF is dedicated to flotation tails.

Epoch Resources (Pty) Ltd (Epoch) designed the TSFs and has prepared a LOM Strategy for the CTSF and FTSF. The facilities are managed and operated by Paragon Tailings, as an onsite contractor. The latter reports to both the Engineer of Record and the site Process Plant Manager.

A large volume of the tailings generated by the plant is used for underground backfill. Currently, it is estimated that up to 40% of the flotation tailings is used for paste backfill.

The CTSF comprises two full containment, HDPE lined facilities (CTSF1 and CTSF2) that have a continuous surrounding embankment and share a common internal wall. CTSF1 and CTSF2 footprints have been merged into a single footprint (CTSF 1^stLift) by raising the embankment walls and sacrificing the common internal wall. Currently, the CTSF 1^st lift is at 80% of its capacity in terms of available storage capacity. The embankment walls of the CTSF 1^st lift are currently being raised by 3.5 m using a downstream construction (CTSF 2^nd Lift) method. The CTSF 2^nd Lift is expected to be completed by March 2022 and will give an additional dam capacity of approximately 6.5 Mt that covers the tailings deposition plan up to 2026.

The current minimum vertical freeboard for CTSF 1^st lift is 1.9 m below the emergency spillway invert level.

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CTSF Phase 3 is currently envisaged to be constructed north of the previous CTSF Phase 2 footprint, however, this is still conceptual. Based on the current LOM plan, deposition into the Phase 3 facility would commence in Q1 2027 and continue until the end of the expected mine life.

The FTSF will require a LOM capacity of approximately 57.8 Mt. Phase 1 of the facility is an unlined valley impoundment, formed behind an embankment that traverses the valley. Phase 2 is currently in operation and is being operated as a full ring-dyke impoundment as a self-raising facility where the TSF is being raised by paddock deposition. Phase 3 of the TSF consists of buttressing the downstream slope of the embankment walls and paddocks with waste rock to ensure the TSF has adequate stability under post liquefaction conditions. The first phase of the buttressing construction is scheduled to start in December 2021 and continue until Q3 2022.

In unlined RWD captures and stores return water from the FTSF. An unlined catchment dam captures water released from the RWD into the main storm water diversion channel, and it captures all the run-off accumulated from within the mine footprint.

18.6

Power Supply

Since there is no grid power available in the region, Kibali needs to be self-sustaining and indeed possesses considerable thermal power generation capacity to do so. Diesel generated power comes from three banks of on-site high-speed diesel generators, each bank consisting of twelve x 1500 kVA, 400V CAT 3512B generators. To mitigate the running costs of this facility, three hydropower plants have been installed.

These are as follows:

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Nzoro 2 Four x 5.5 MW turbines – Total installed 22 MW

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Ambarau Two x 5.3 MW turbines – Total installed 10.6 MW

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Azambi Two x 5.1 MW turbines – Total installed 10.2 MW

Separately, the pre-existing Nzoro 1 facility is of low capacity (i.e., less than 1 MW). It was previously refurbished and represents a historical legacy comprising equipment dating from the 1930s. This power is dedicated to local communities.

The Nzoro 2 hydropower station was optimised during 2015 and reached its design power supply (22 MW) by the start of 2016. Commissioning of the second new hydropower station, Ambarau, was achieved in 2017 and the completion of a third station, Azambi, was achieved in 2018.

The long-term power supply strategy for the operation is aimed at generating the maximum amount of power from hydro sources. Diesel generators will remain available as back up and as a spinning reserve for peak loads from the shaft hoist. Further improvement was made by installing a 9MW battery bank that was commissioned in 2020. The running generators have been reduced by half during wet season. This has a marked effect on reducing unit power operating costs. Wet seasons with high river flows allow for more beneficial hydro operating

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conditions, however the beneficial effect is not seen in the lower rainfall months. This effect is evident in Figure 18-2 which shows the power supply mix to the end of 2021.

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Figure 18-2 Kibali Electrical Supply Mix

The total installed hydroelectric power capacity is 42.8 MW, which currently covers most of the mine power demand. Full power demand at full production is currently between 39 MW and 43 MW. With the full implementation of the hydro strategy in addition to the battery storage system and grid stabiliser, the diesel generators only supply 4.8 MW, with the remainder being provided by the hydro stations.

Therefore, the system has a potential capacity of 42.8 MW of Hydropower (at peak) and 43 MW of thermal generation. Actual hydro generation capacity is season dependent:

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Maximum Capacity (43 thermal generation + 42.8 hydropower) MW.

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Minimum Capacity (32 thermal generation + 10 hydropower) MW.

The load demand of the mine is not constant, power demand at full production is currently between 39 MW and 43 MW, averaging approximately 41 MW.

Electrical power at 66 kV is supplied by the hydropower stations connected to a main grid supply.

The hydro-generated power is reticulated to the site by means of 66 kV overhead lines from the hydropower plants to a switchyard located at the mine. The voltage is be stepped down from 66 kV to 11 kV, feeding the 11 kV consumer substation.

Diesel generation supplies power to the mine at 400V, which is stepped up to 11kV for distribution.

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18.7

Site Infrastructure

Operational Camp (Village)

The operational camp provides accommodation for single and married staff and incorporates all the required facilities in terms of accommodation, ablution, catering, and messing facilities.

The camp comprises two villages to accommodate the mine employees; a large single status camp near the mine operations and a married-quarters camp that was opened in 2015.

A single kitchen and dining room is provided for residents of the camp. A further kitchen and dining area is available at the social club that could be used if the camp kitchen was destroyed. Each of the major contractors operates their own camp and kitchen facilities nearby to their base of operation.

Offices, Stores, and Workshops

A central administration area office complex accommodates senior and administrative personnel as well as discipline functions not located specifically in the process plant or mine operations offices.

The plant area includes the necessary buildings for the operations personnel related to the process operation including a gate house, control room containing the plant server and SCADA equipment, engineering room, UPS rooms, engineering offices, laboratory including carbon room, metallurgical laboratory, wet laboratory, bullion room, balance room, environmental laboratory, receiving area, sample preparation and grade control preparation, and a maintenance workshop and offices.

The Central Mine Facilities Area is located adjacent to the processing plant and includes large stores facilities related to spares and engineering consumables for mining, processing, and general operations.

There are four large buildings to hold most of the stores stock, most of which is spares for machinery, and the remainder is consumables, such as personal protective equipment. There is sufficient covered space for spares and consumables.

The buildings are all steel framed and clad with steel sheeting. Floors are reinforced concrete.

The shaft collar area provides an office building, change house, security gate house, and a workshop for the underground mining operation.

The open pit mining central operations area includes a large workshop for the maintenance of the mining fleet, an office building, a change house, and a security gate house.

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Emergency Response and Medical Facilities

There are two mine rescue teams on-site with a total of seventeen active members of which ten are on-site at all times. These people all have positions on the mine with the mine rescue being an additional responsibility.

Emergency situations will be communicated by radio on a dedicated channel. A stench gas system is available.

A fire truck and trailer are available for rescue teams.

Medical staff on site includes two doctors, six nurses, and laboratory technicians. There are three ambulances on site and four first aid rooms together with a health clinic.

The nearest hospital with good facilities is in Kampala. In the event of a need for medivac, arrangements with the air charter company would be made.

Kibali Goldmines runs a malaria prevention programme involving bush clearance and spraying, and a campaign to improve awareness. This has resulted in a significant reduction in malaria cases. The programme is ongoing. In addition, there is a continued HIV AIDS campaign including voluntary counselling and testing.

Fuel Storage

The fuel storage installation includes three separate fuel farms.

Daily consumption is approximately 180,000 litres during the wet season and 200,000 litres during the dry season. Approximately 65 to 70% of the consumption is used by the diesel generators at the thermal power station, 20% is used by mining and the remaining 10% is general use.

The largest fuel farm is located in the central mine facilities area. The main fuel farm for the mine has three one million litre tanks and six 100,000 litres tanks, giving a total storage capacity of 3.6 Ml. Diesel is filtered before it is pumped into the main tanks and after it leaves the secondary tanks.

Extensive fire protection is provided for the main fuel farm and includes a series of foam generators located around the perimeter of the containment bund and cooling rings on the tanks. The water for these fire protection systems is supplied from two dedicated tanks and two fire pumps located at the process plant.

Two other fuel farms have been built at the open pit and underground operations and have a capacity of 1,200 m³ each with similar dispensing facilities.

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Airstrip

Access by air to Kibali involves a commercial flight to Entebbe in Uganda followed by a charter flight to Doko airport, situated on the mine property. The Doko airstrip was upgraded by Kibali Goldmines and is equipped with runway lights and precision approach path indicator lights.

Charter flights to site are arranged by Kibali Goldmines on a regular schedule at frequencies dictated by operational requirements.

18.8

Communication and Information Technology

The mine wide voice and data backbone with satellite fibre optic link(s) provides cellular for voice and internet connections via wireless Local Area Network (LAN). Voice communication is supplemented by two-way radio.

Fibre optics on overhead lines provide for communication between the various operations sites.

18.9

Security

There is comprehensive security infrastructure at the site, with controlled access to the operations. The Security Manager reports directly to the Kibali Goldmines General Manager.

The Kibali property is surrounded with a high fence and a security access road running along the perimeter.

The plant area is fenced with security at the main gate and additional electronic access systems and security at higher values areas within the plant.

The spares and materials storage sites are fenced, and access gates are kept locked, and access controlled by security staff.

Gold doré produced at the mine site is shipped from site on security escorted private charter flights.

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19	Market Studies and Contracts

The principal commodity produced at Kibali is gold, which is freely traded at prices that are widely known, so that prospects for sale of any production are virtually assured. Prices are usually quoted in US dollars per troy ounce.

19.1

Markets

Gold doré produced at the mine site is shipped from site under secured conditions and sold under agreement to Rand Refinery in South Africa. Under the agreement, Kibali Goldmines receives the ruling gold price on the day after dispatch, less refining and freight costs, for the gold content of the doré gold. Kibali Goldmines has an agreement to sell all gold production to only one customer. The “customer” is chosen periodically on a tender basis from a selected pool of accredited refineries and international banks to ensure competitive refining and freight costs. Gold mines do not compete to sell their product given that the price is not controlled by the producers.

19.2

Contracts

It is Kibali Goldmines’ strategy to outsource open pit mining activities to contractors and, in all instances, the contracts are such that the equipment can be purchased by Kibali Goldmines at the end of the contract period at its depreciated price or should the contractor default at a predetermined pricing mechanism.

Prior to start-up all major mining contractors are requested to tender and the most appropriate tender is accepted thereby ensuring that the best competitive current pricing is achieved. Care is taken at the time of finalising contracts to ensure that the rise and fall formula is totally representative of the build-up of the quoted price per unit. At the time of award, prices quoted are compared to benchmark prices of other owner miner operations.

The contract mining costs are dependent on when tenders are issued as the price of major equipment varies dependant on demand as well as the cost of finance. Rise and fall can be negatively affected by currency fluctuations as well as price squeeze due to scarcity.

The mine produces doré bars which are sent to an accredited gold refinery for refining. Refining prices are subject to fluctuations in the cost of transport as well as insurance costs.

Other contracts that are put in place include assay facilities, oxygen supply, catering services,

The QPs note that all material contracts discussed above are currently in place and the terms contained within the sales contracts are typical and consistent with standard industry practice and are similar to contracts for the supply of doré elsewhere in the world. All contract terms, rates and charges are within the norms of Barrick’s regional benchmarks, which are generally within the lower half of industry wide standards.

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The QPs note that metal prices used in this Technical Report have been set by Barrick on behalf of Kibali Goldmines and are appropriate to the commodity and mine life projections, fuel supply, explosive supply, and security.

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20	Environmental Studies, Permitting, and Social or Community Impact

20.1

Environmental and Social Management

Summary

Kibali is located in the Haut-Uélé Province and within the administrative area of Watsa Territory. The border towns of Aru in the DRC and Arua in Uganda are located 150 km east of the Project and are on the main road servicing the project area. The capital city of DRC, Kinshasa, is approximately 1,800 km SW of Kibali. The town of Durba is immediately adjacent to the southern boundary of Kibali; nearby villages in the Surur Secteur include Renzi approximately 3 km south, Kotamalembe 5 km east, and Kokiza 3 km west.

The Project’s deposits (outlined in Section 14 in this Technical Report) are mined using open pit and underground methods. While first gold was poured by Kibali Goldmines at Kibali in 2013, the Project area has been previously mined by various companies and numerous artisanal miners.

Waste rock is disposed on Waste Rock Dumps (WRDs) that are located adjacent to the open pits and underground shaft where it was mined. Oxide and sulphide ore is trucked to a central processing plant for processing using crushing, grinding (including an ultra-fine grind), gravity concentration, flotation and CIL followed by smelting to produce doré bullion. Tailings are disposed of in two tailings facilities; the first unlined facility is used for the flotation tails (FTSF), and two lined facilities (for the concentrate tails (CTSF1 and CTSF2)), which have been merged into one facility. The concentrate tails are acid producing and contain cyanide residues and arsenic containing materials. A portion of the flotation tailings are used for paste backfill in the KCD underground mine (refer to Section 16 of this Technical Report).

The Kibali River dominates the drainage of the Kibali Moratorium Zone. The Kibali River flows along the southern boundary of the mine and then flows northwards to a confluence with the Nzoro River approximately 40km downstream; the Kibali River then flows into the Uélé River approximately 200 km downstream from the mine. The Mengu, Marakeke, Indi, Doko, and Renzi rivers are all tributaries of the Kibali River catchment within the Moratorium Zone. The flow and volume of the Kibali River increases during June and July and experiences a larger rise during September and October due to seasonal rainfall. Three hydropower plants were built by the company: Nzoro 2 in 2015, Ambarau in 2017 and Azambi in 2018. A battery energy storage system was incorporated in 2020 to improve power stability. The hydropower plants and battery energy storage system significantly reduce dependence on diesel fired power plants and thus also reduce operational greenhouse gas emissions (GHG). In addition, the existing Nzoro 1, hydrostation was refurbished and is exclusively used to provide power to the local community.

Physical and economic resettlement has been completed for project-related land acquisition. Approximately 36,500 people have been resettled to date. The largest resettlement was associated with the KCD mining operations and involved 20,000 people. In preparation for the

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influx due to mining and mine construction, to house a number of the resettled people, the Kokiza village was established. Economic displacement (loss of land, crops, trees, and access to natural resources/ecosystem services) has taken place across the 3,150 ha Kibali Moratorium Zone.

A RAP is being developed and will be implemented for the Kalimva-Ikamva deposits (see Section 20.3).

Environmental Assessment and Permitting

Three ESIAs, and two ESIA updates have been completed for the Project. All ESIAs were undertaken in compliance with DRC legislation and the applicable IFC PS (2006); ESIA updates were compliant with DRC Legislation and IFC PS (2012). The following list identifies the ESIAs and EISA updates completed since 2010:

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An ESIA completed by independent consultants (Digby Wells, 2011) as part of the feasibility study during 2010 and 2011. The ESIA report was submitted to the authorities in 2011 and approval was received in 2011.

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An ESIA was completed in June 2011 for the new Nzoro 2 hydropower station, and refurbishment of the Nzoro 1 hydropower station adjacent to the Kibali and Nzoro Rivers, respectively (Digby Wells, 2011). This ESIA included details of the upgrade of the existing powerlines from the Nzoro 1 station, construction of new powerlines from Nzoro 2 and the construction of a diversion canal from the Nzoro River to the Nzoro 2 station.

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An ESIA was completed in 2012 for the Ambarau and Azambi hydropower plants located on the Kibali River (Digby Wells, 2012).

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ESIA Updates for the Mine in 2015 (Digby Wells, 2015) and 2020 (Digby Wells, 2020).

The Project is predominantly governed by the DRC Mining Code (2002) and associated Mining Regulations. Decree No. 038/2003 of 26 March 2003 relating to the Mining Regulations as modified and completed by Decree No. 18/024 of 08 June 2018 contain provisions regarding ESIAs and environmental management, public consultation, and compensation for loss of access to land. Articles 127 and 128 of the Mining Regulations (2018) sets out the contents of the EIS and the EMP and Article 452 establishes the objectives of management measures and standards of the EMP. Public consultation of the Project was achieved in accordance with Articles 451 and 478 of the Mining Regulations (2018) and with the IFC PS.

Under the DRC Mining Code (2002), mining operations must be covered by an EAP, which must be approved by the DPEM. The EAP must give an overview of the environmental conditions of the areas covered by the relevant mining title and to describe any measures that have been or will be taken to protect the environment. In practice, the EAP covers what is normally required in an EIS and an EMP (collectively referred to herein as the ESIA).

The review of the environmental impact studies and the environmental management plans presented in the Kibali EAP was completed by the Standing Committee of Evaluation (CPE) comprising 14 members and directed by the Director of the DPEM. The EAP was approved by the CPE, required under Articles 455 and 456 of the Mining Regulations (2003) and included the following conditions:

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Adequate management of social aspects around the mine.

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Respect of air quality requirements.

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Water management and effluents to be in line with the legal limits before any discharge from the mine.

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Waste management and hazardous waste management in line with legislation.

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Flora and fauna promotion and conservation.

Copies of the EAP were submitted to the Mining Registry Office as requested under Articles 69, 92, 103, 154 of the DRC Mining Code (2002) and Article 454 of the Mining Regulations (2003).

Subsequent to the 2011 ESIA, amendments were made to the ESIA as summarised below. These were approved by the General Secretary of the Ministry of Environment who chaired the CPE to assess the ESIA and management measures:

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In 2014 and 2015, the three ESIAs were consolidated into one document and the impact assessment and management plans were updated in the document. It was approved in 2016.

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In 2020, the ESIA was revised to incorporate Kalimva-Ikamva, and to comply with the Mining Regulations (2018) that stipulates a mine’s ESIA is to be updated every five years (Article 463). This allows for a re-examination of the management processes and responsibilities and assists the mine in managing its environmental and social impacts on an ongoing basis.

The 2020 ESIA update complied with DRC laws and regulations and conformed with the IFC PS (2012). Mitigation and rehabilitation measures and financial provision for planned Project closure have been included in the ESIA update. Kibali undertakes concurrent rehabilitation of disturbed areas. Pakaka, Kombokolo, Rhino, Mofu and Mengu pits have been fully, or partially rehabilitated and environmental monitoring of these areas is ongoing. Some closed pits may be subject to future mining and/or underground mining.

All environmental permits are in place for the Kibali processing plant, open pits and underground operations, the hydropower stations, and a permit register forms part of the EMP. Permits include:

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ESIA approbation – letter for approval of the environmental impacts assessment (valid for 5 years and subject to ESIA Updates).

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Certifcate environnemental (valid as long as taxes are paid).

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Permit to export used oil (1 year licence subject to annual renewal).

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Permit d’exploitation (25 years).

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Authorisation for owning the hydropower plants (25 years).

Other project permits and licences in place include an import and export licence, permit for the construction of infrastructure at Kokiza, authorisation to import explosives, demolition permit, authorisation to resettle people, authorisation for exhumation (so that graves can be relocated out of the mining zone), and title deeds for all people resettled in Kokiza.

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Environmental and Social Management and Monitoring

Personnel of the Kibali Goldmines Environmental and Social Departments report to the General Manager on site. Functional reporting is to the Africa and Middle East (AME) Regional Sustainability Manager and Group Sustainability Executive, who provides technical support as needed. Departments develop budgets and programmes, in consultation with the General Manager, AME Sustainability Manager and Group Sustainability Executive. Departments are responsible for performance, and compliance/conformance through carrying out audits, inspections, and monitoring. The mine ESMS is ISO 14001:2015 certified and audited by independent consultants. Audits are also carried out to gauge conformance with the ICMC certification and construction of a cyanide detox plant for the tailings stream is planned to commence in 2022.

A consolidated ESMP is in place which covers all aspects of the operation and was updated as part of the 2020 ESIA. The ESMP includes current, future planned and proposed activities and a rehabilitation plan. The ESMP includes an Environmental and Social Monitoring Plan as approved by the regulators and comprises the following:

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Air quality and dust.

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Water sampling and analysis of:

TSF seepage water and tails streams (particular focus on arsenic and WAD cyanide which can be analysed on site).

Potable water.

Groundwater.

Surface water

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Terrestrial and aquatic biodiversity/habitats.

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Noise and blasting.

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Soil.

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Community relations and grievances.

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Energy use.

In 2020, more than 80% of the energy consumed by Kibali was provided by the hydropower plants. Waste is segregated and managed by adopting the waste hierarchy (avoid-reuse-recycle-landfill); some incineration takes place on site at the installed Macrotech incinerator V70. In 2020, a total of 650 t of waste was incinerated at the onsite incinerator, 3,400 t of waste reused or recycled, and a further 1,900 t to landfill. New opportunities are being sought for reusing or recycling waste to further reduce waste to landfill.

Environmental performance objectives are established by senior management, including the General Manager and AME Regional Sustainability Manager, and communicated within the operations. All the strategic inputs are formalised, and close tracking is carried out through the ESMS.

Environmental incidents are recorded in a register which forms part of the ESMS. Causes and responses are identified, and incidents closed out once investigations are completed. A total of

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11 environmental incidents were recorded in 2021; all were classified as Class 3 environmental incidents, which are defined as minor incidents that do not pose any adverse impacts or risks to human health or the environment. Most Class 3 incidents were minor oil and fuel spills. Investigations were undertaken for events to ensure appropriate maintenance and pre-work inspections are carried out to reduce the risk of recurrence.

20.2

Environmental Considerations

Water

Streams that flow through communities are used for domestic purposes including washing or fishing, and riparian vegetation is used for building material and implements. Deep boreholes were established by Kibali, powered with solar energy, as water points for surrounding communities to provide cleaner potable water for domestic use at sites where communities have been moved to or as part of community development projects.

A Water Quality data review was carried out as part of the 2020 ESIA update (Digby Wells, 2020), which examined historical and current water quality results collected to date. There are a total of 39 water monitoring locations at Kibali, comprised of 38 surface water and 13 groundwater sampling locations, and a further 22 boreholes (11 pairs of shallow and deep boreholes) associated with the TSFs monitoring. The water quality review had the following conclusions:

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Groundwater:

The main groundwater types are Ca-HCO3 and Mg-HCO3, indicating mostly fresh, unpolluted groundwater.

TSF monitoring wells indicate predominantly moderately acidic to neutral conditions (pH between 5.5 and 7).

Arsenic was detected in most of the boreholes at Kibali, however, only a few exceeded the regulatory guideline limits (in March 2019 and only one borehole in September 2019).

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Surface water:

In 2018, the catchment dam, downstream of the ROM pad, had elevated levels of Total Dissolved Solids (TDS), sulphate and arsenic against the World Health Organisation (WHO) Drinking Water Guidelines (2017).

Water quality at sampling locations within the Kibali Moratorium Zone predominantly conform with IFC Environmental Health and Safety Mining Guidelines (2007), with several spatial and temporal exceptions, notably TDS and sulphate levels adjacent to the FTSF and at the Return Water Dam.

Arsenic levels were elevated at sampling locations adjacent to operational areas, namely Sessenge and Pakaka. These waste dumps were rehabilitated and encapsulated to prevent leaching.

The pH across the monitored sites were within recommended IFC EHS Mining Guidelines (2007) values (pH 6 to 9), the exceedances were for samples associated with KCD in 2014 and within the CTSF and FTSF in 2018.

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The gold deposits have variable amounts of arsenic; these are identified during mining and any high arsenic ore is stockpiled and blended with low arsenic ore through the plant to provide steady and consistent throughput. Arsenic has also been identified within the mining and process streams during routine water quality monitoring and led to a more detailed assessment of the FTSF seepage.

A comprehensive water balance model has been compiled for the site, which models flows, inputs and losses throughout the operations (i.e., the open pits, underground workings, process plant, TSFs, water management structures, offices, camp, and sewage treatment facilities). The model includes inputs regarding river water use (e.g., discharges, gains, and losses, volumes of potential savings/recycling opportunities), which would reduce abstraction rates from the river. Opportunities to reuse water within operations has significantly reduced the volume of freshwater abstracted from the Kibali River. In Q3 2021, abstraction from the Kibali River had reduced to an average of 22, 000 m³ per month, from a capacity potential of 210,000 m³ per month.

Geochemistry

A summary of the geochemical characteristics for Kibali was provided in the 2018 Technical Report (Randgold, 2018). Waste material was found to have moderate to high acid neutralising capacity for the majority of lithologies tested. Acid base accounting showed that 5% of the samples were classified as being potentially acid forming, which is considered to be a very low proportion of the total waste rock. Waste samples were found to be enriched in a limited number of elements with arsenic and antimony being the most highly enriched elements. Water extraction testing found that most of the enriched elements had low solubilities under certain neutral pH conditions. Arsenic, molybdenum, barium, cobalt, zinc, nickel, and selenium were found to be above the guideline water quality values in one or more samples, with arsenic and molybdenum the most commonly elevated elements.

Since 2018, geochemical assessments have been undertaken by Digby Wells for the following deposits: Pakaka, KCD Open Pit and Underground, Sessenge (2019), Pamao (2020), Kalimva-Ikamva (2020) and Aerodrome (2021). All other deposits were all assessed prior to 2018. The geochemical assessments included X-Ray Diffraction, Aqua regia digestion, Acid Base Accounting, Net Acid Generation (NAG) and Sulphur speciation, and water leachate tests at external independent accredited laboratories in South Africa total of 122 samples were undertaken for the assessments since 2018.

Only seven of the 122 samples demonstrated to be Potentially Acid Forming (three at Pakaka, two at Kalimva, and one each at Sessenge and Gorumbwa), the remaining samples being classified as Net Neutralising Potential (NNP). Collectively the risk for Acid Mine Drainage (AMD) is low due to the neutralising potential of the majority of the waste rock. Arsenic was the leachate of concern identified in all ore bodies under conservative conditions. Any waste rock dumps with the potential for leachate generation are placed within designated areas and water runoff is separated, collected, and monitored for water quality. Dirty water is contained within operational areas and circulated back to the plant for use. Clean water is diverted away from operational areas into the environment. Sessenge and Pakaka water quality showed elevated arsenic and the waste dumps were encapsulated to prevent leaching.

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Waste rock is used to build various infrastructural platforms on site, while the remainder is stockpiled on surface or deposited in stopes as backfill. Acid Rock Drainage (ARD) has not been detected at the site.

Biodiversity

The biodiversity description was sourced from the ESIA Update (Digby Wells, 2020).

The Project lies within the Northern Congolian Forest Savanna Mosaic ecoregion that includes the northernmost savanna woodlands in Africa (White, 1983). This narrow transition zone marks an abrupt habitat discontinuity between the extensive Congolian rain forests to the south and the Sudanian/Sahelian grasslands to the north. This forest-savanna mosaic represents the eastern half of the Guineo-Congolian/Sudanian phytogeographical regional transition zone.

The Northern Congolian Forest Savanna Mosaic ecoregion is typified by a combination of gallery forest, woodland and secondary grassland and is controlled by annual precipitation, duration of water stress, and the severity of dry-season fires and human activity. A vegetation gradient can be observed from the savanna hills with sporadic tree species, through to the moist savanna on the flatter lower-lying areas to the relatively dense riparian woodland. The changes in vegetation are due to inter-relationships between land use and topography, changes in elevation, and soil moisture. Intact open savanna occurs on rocky hills which is not used for cultivation.

The natural vegetation in the concession and area of influence, has been largely transformed through human activity. Site clearance for the establishment of infrastructure, together with anthropogenic activities has occurred across all vegetation habitat types. Alien invasive plant species occur throughout all habitat types. As the local population has increased, traditional agriculture has become less successful due to loss of soil fertility; the average fallow period has decreased from 20 years to as little as three, reducing the ability of soils to regain fertility. The number of trees has reduced due to informal harvesting and created extensive sparsely wooded grasslands.

The landscape can be described as largely disturbed with pockets of natural vegetation. Most protected plants species remain within gallery forests (Digby Wells, 2015) that are associated with drainage lines and water courses. Protected plant species identified within these forests include:

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Albizia ferruginea is considered to be of conservation significance. It is listed as being Vulnerable by the International Union for Conservation of Nature (IUCN). It is described as a widespread and often common timber species, which has suffered heavy exploitation. The major threat is from timber exploitation. This tree was encountered in gallery forest and secondary forest only.

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Guarea cedrata (also called Light Bossé or Scented Guarea) is a species of tree in the Meliaceae family; it is exploited in the timber industry and has therefore been listed as Threatened by the IUCN. This tree species was only encountered in the gallery forest vegetation type.

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Pterygota bequaertii is a species of flowering plant in the Sterculiaceae family; it is listed as being Vulnerable by the IUCN. This tree species was only encountered in the gallery forest.

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The animals expected to occur within the Project area are predominantly savanna and forest dwelling species. Hunting for bushmeat is common practice and certain primate species such as Patas monkeys have been rescued by mine officials from villages in the area. A total of 26 mammal species have been identified in the region since 2006, the largest numbers occurring within gallery forests. Three IUCN Red Data species have been observed within the area since 2006, namely the African Golden Cat, Hippopotamus, and African Elephant, although these species have not been identified within the last decade.

The gallery forest along the Kibali River functions as an important ecological corridor, while the river itself is utilised by several species of birds. No bird or reptile species of conservation concern have been identified to date.

With an emphasis on the poor level of taxonomic exploration, there are very few studies relating to the biological and/or hydrological knowledge of the fish expected within the study area. Nonetheless, at least 136 different fish species have been formally documented within the ecoregion with largely diverse groups. Biannual biomonitoring is undertaken by independent consultants, and the 2020 results (Digby Wells, 2020) led to the conclusion that instream habitat availability was regarded as adequate to good. Fish assemblages showed a marginal increase from previous assessments since biannual biomonitoring commenced in 2017.

Additional biodiversity monitoring is ongoing, such as the use of camera traps to detect fauna within the concession. Kibali Goldmines has a Biodiversity Action Plan (BAP) and Biodiversity Management Plan which are updated to reflect new information gathered during specific ecosystems monitoring:

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Aquatic biomonitoring conducted at least annually.

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Wetlands and amphibians assessment annually.

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Automatically monitoring with cameras, drones, and satellites / aerial imageries (on weekly basis).

The mine site lies approximately 65 km south of the Garamba National Park, which is on the border with South Sudan. A partnership between Kibali Goldmines and the Garamba National Park has been established; Kibali Goldmines provides technical and financial assistance and support to Garamba, including provision of scientific survey information (such as aquatic biodiversity surveys), support for collaring and tracking of elephants, supplies fuel to anti-poaching teams, supports the Kordofan giraffe programme, and building road infrastructure such as bridges within the Garamba National Park to assist the rapid response of ranger patrols. This partnership provides a wider strategic support for game protection from poachers from the north, and connections with local enforcement networks.

Mine Rehabilitation and Closure

Mine rehabilitation is designed to restore the biophysical environment (e.g., chemical, biological quality of air, land, and water regimes) and prepare the concession for post-mining land use. Concurrent rehabilitation opportunities are limited as some mined pits and inactive WRDs are being assessed for potential future open pit expansion or underground operations. Inactive pits and WRDs remain within Kibali’s environmental monitoring programmes to analyse potential

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impacts. Inactive pits are access restricted and are located within the Moratorium Zone and have security posts to ensure controlled entry.

The aim of mine closure is to (a) develop a passive system that is a self-sustaining natural ecosystem or (b) prepare the concession for alternate land use that stakeholders agree to, and the authorities are willing to sign off. Once the post-monitoring period (at least 5 years) has established that the site is stable, the authorities will sign off and end the company’s liabilities for the concession.

A framework closure plan was developed as part of the 2011 ESIA and has been updated as part of the 2020 ESIA to reflect changes in mine development, operational planning, and the environmental and social status quo. The framework closure plan addresses the following:

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The regulatory framework for mine closure.

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Methods used to close all mine components.

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The overall closure objectives for all components of the Project.

Mine closure costs are updated each year. The current cost as of 31^st December 2021 for rehabilitation and closure of the mine is $23.67 million (Digby Wells, 2021).

Allowance has been made for the shaping of the open pit edges and WRDs to a safe and sustainable angle. Rehabilitation of the ROM pads, demolition and management of physical infrastructure, creation of a free-draining topography, replacement of soil, re-vegetation, and general surface rehabilitation of all the disturbed areas within Kibali has also been calculated.

At closure the CTSF will be rehabilitated by covering the facility with a 300 mm saprolite layer (breaker layer) followed by a 300 mm layer of topsoil. The top and side walls will be vegetated to stabilise the tailings against wind and water erosion, and to reduce water ingress into the tailings. Surface water diversion and management measures will be left in a state such that they can continue to control runoff from the TSF and to divert clean water around it. All upslope water will be permanently diverted around the facility; water falling on the TSF will be encouraged to discharge from it to avoid pools forming on the surface.

The FTSF will be rehabilitated in the same manor but will exclude the 300 mm saprolite breaker layer since the FTSF material has less contamination potential and it is therefore assumed a breaker layer will not be required.

Infrastructure at the Doko airport, mine camp and mine offices could have uses for the community after mine closure and will be discussed with stakeholders as post-mining land use plans are developed. Extension areas such as laydown areas will be rehabilitated as per the contractor’s agreement with Kibali Goldmines. The cost of demolishing and rehabilitating these areas has therefore been excluded from the closure cost assessment. A contingency of 10% has been included. A 12% allowance has been included for operator or contractor project management fees of the airport.

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The total includes costs for bi-annual aquatic biomonitoring, surface and groundwater monitoring for five years after mine closure, monitoring and maintaining re-vegetated areas for three years after mine closure, hydro-carbon clean-up, and cyanide decontamination.

20.3

Environmental Studies, Permitting, and Social or Community Impact

Employment and Procurement

Kibali Goldmines complies with the labour laws of the DRC, which govern the following:

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Salary and remuneration.

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Job classification and competencies.

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Annual leave system.

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Ratios of expat to national workforce.

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Representation by unions.

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Employee code of conduct and disciplinary measures.

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Mine Level Agreement (MLA).

Kibali Goldmines has an internal in-reach programme which is a platform where both employer and employee are able to actively engage with each other regarding operational updates in addition to social and community matters.

Kibali Goldmines employment policy gives priority to DRC nationals who have the required skills and experience. Identifying skilled nationals involves advertising and searching in the nearby communities before extending the recruitment process to other regions of the country. Where there is a lack of skills, expatriates with specific skills are employed with the primary aim of training nationals. A timeframe is developed for training nationals to take over from the skilled expatriates. Development plans are in place to facilitate skills development and succession planning.

Kibali Goldmines has a local procurement policy, and this extends to procurement through contractors. Kibali Goldmines procured in excess of $110 million of goods and services from DRC suppliers in 2021. This includes produce from agribusinesses (e.g., producers of eggs, pork, maize) which is purchased for use in the mine canteens.

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Resettlement

As indicated in the 2012 RAP report (RADS and Digby Wells, 2012), Kibali Goldmines is committed to the following resettlement principles:

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To apply international standards (specifically, IFC PS 5 – Land acquisition and involuntary resettlement)

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DRC legislation

International standards require that host-country laws are complied with, by default. Kibali Goldmines reviews legislation and international standards and adopts the most stringent requirements.

Kibali Goldmines follows a resettlement and compensation process that will leave PAPs in the same or better off position than before the Project intervention, which is in conformance with IFC PS.

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To follow in-kind compensation where possible and limit cash compensation as far as possible, especially where the affected community’s livelihoods are at stake.

As a result of the construction of the Project and establishment of the Moratorium Zone, it was necessary to resettle approximately 36,700 people, from 7,504 households, since 2012 to date (RADS and Digby Wells, 2012; Digby Wells, 2015; Digby Wells, 2016; Digby Wells, 2020). The Project also displaced around 134 items of community infrastructure, including 13 communal agricultural projects, five communal business/commercial facilities, 12 education facilities, 19 health facilities, nine recreational/community facilities, 39 religious facilities, and 41 water sources.

The first RAP was initiated in 2012 for the establishment of the mine’s Moratorium Zone (Zone A) and completed during 2013 and involved 20,000 people from 4,000 households in 14 villages. Where PAPs insisted on cash compensation, Kibali Goldmines put processes in place to make sure funds were used appropriately and that recipients receive materials and goods that were provided for in the budget (e.g., if people decide to build infrastructure themselves, payments were made in instalments and full payment was only made upon completion of construction).

A Resettlement Working Group (RWG) was established as the primary consultation forum to develop and implement a RAP. The RAP process was carried out by independent consultants. All primary stakeholders are represented on the RWG.

The RAP included construction of water, energy, and road infrastructure. Guidance was provided by Congolese town planners, as well as the RWG, for a town plan outlining the development of the Kokiza village that improved the provision of basic services and social infrastructure whilst still being maintainable, considering the overall remoteness of the area. The following was constructed at Kokiza:

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Catholic Church.

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4,077 houses.

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14 schools and recreation facilities.

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Five clinics/health centres.

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11 churches.

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55 boreholes.

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Upgraded electricity

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Four new cemeteries.

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Two communal markets.

A RAP completion audit was carried out in 2013 by an independent, third party (Digby Wells, 2013). At that time, nine of the 14 affected villages had been relocated to Kokiza, the site identified by the RWG. The audit comprised assessing the implementation of the RAP against commitments set out in the RAP report.

The objectives of the audit were to identify any social/community issues or risks associated with the Project, focusing on the following four Project components: Mine site, Resettlement host site, Doko-Aru Road upgrade, and Nzoro hydropower station and power line. Since completion of the audit, all recommendations have been implemented.

The Gorumbwa RAP was initiated in 2016 to allow for the mining of the Gorumbwa Pit and consisted of 1,482 households. The resettlement was applied to households situated within the mining perimeter and was based on compensation and construction of a replacement property. Additionally, eight boreholes were drilled at Kokiza and handed over to the community as well as all community infrastructures.

The Pamao RAP initiated in 2020 includes Pamao North and Pamao South as expansion areas to the Moratorium Zone A to allow mining activities of the Pamao pit. It involves 628 households from two villages, whereby 222 households are physically displaced and 406 economically displaced, who were engaged in faming activities within the affected zone but did not reside there. An additional 250 households were affected by Pamao Diversion Road and Gatanga-Surur Diversion Road, which are both deviating the RN26 National Road whose section is affected by the Pamao North Zone. The physically affected households will be resettled at the Avokala host site, along with the Kalimva-Ikamva PAPs.

The Kalimva-Ikamva RAP was initiated in 2019 and is still under development. It involves 1,888 households from six villages, whereby 1,141 households are physical displaced and 747 economically displaced (Digby Wells, 2020). An additional 232 households are affected by the host site work at Avokala, and two diversion roads created heading to the host site. Through the RWG consultation, Kibali Goldmines has made funds available in the event that PAPs decide to build infrastructure themselves. In such cases, payments are made in three instalments and full payment will only be made upon completion of construction.

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Stakeholder Engagement

A Stakeholder Engagement Plan (SEP) was developed in 2015 (Digby Wells, 2015). The Kibali Goldmines Social and Community department updates the SEP annually and has also developed a Social Licence Strategy and a Community Development Plan. The SEP outlines the Mine’s stakeholder groups, engagement governance and objectives, communication methods and frequencies. In 2017, Kibali Goldmines reinforced its relationship with the Provincial Government through the collaboration on various projects e.g., the Master Plan for the Development of the Province of Haut-Uélé; the evaluation of stakeholder engagement activities, creation of the steering committee for the construction of the General Hospital of Watsa and others.

Kibali Goldmines has built strong relations with the community through reinforced and continuous stakeholder engagement which includes regular meetings with a range of stakeholders and regular radio broadcasts targeting key issues pertinent to the community.

A community development committee is in the process of being established in line with new legislation, the grievance mechanism was revisited and readjusted (Kibali, 2021) and a full programme of stakeholder engagement meetings takes place with key stakeholders from the communities. The Mine has reintroduced the community monthly and quarterly meetings, media, Civil Society, Watsa-Faradje-Aru Youth, and ASM bi-monthly meetings, one-on-one meetings, and radio broadcasts (Kibali, 2021).

Recent development initiatives by Kibali Goldmines include developing a cahier de charge, which is a five-year community development plan and will include investment in agriculture and livestock projects (e.g., maize, pork, poultry, and fish, and undertaking feasibility studies for cocoa, bananas, and palm oil plantations).

At the end of October 2021, a government-led eviction was initiated of illegal occupants of the mine at Moratorium Zone B took place. The land had previously been cleared by Kibali Goldmines and the residents were compensated in accordance with the RAP, but some illegal settlers re-populated the area. The government has since initiated eviction of the displaced illegal settlers. Kibali Goldmines has offered assistance in relocation, the provision of basic materials and town planning activities for alternative settlement, as a gesture of goodwill.

No major issues were reported in 2021 and most of complaints and grievances brought by the community were related to the current and past RAP. Out of 247 grievances received, 209 have been resolved peacefully and 38 remain to be closed (Kibali grievance register, 2021).

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Community Development/Corporate Social Responsibility

2021 community development projects included a tar road through Durba; sponsoring youth education; sporting activities, and renovating infrastructures. Sport and cultural activities were central to the Kibali Goldmine’s corporate social responsibility activities in 2021.

A total of $2.7 million was spent on community development projects in 2021, over and above the 0.3% of Annual Gross Revenue designated for community spend in line with the Mining Code (2018).

In 2021, between 70% and 100% of goods were sourced from local suppliers.

Community healthcare initiatives managed by Kibali Goldmines include Covid-19 awareness and vaccination campaigns, support to Watsa hospital (development of a business plan), Central Hospital Kibali (CHK) hygiene and housekeeping (CHK has contracted a local company to perform daily housekeeping for three months with possibility of renewing; progress is monitored on a daily basis by Kibali Goldmines doctors), and support to the Watsa Orphanage, where health improvements have been observed as a direct result of the involvement of the Kibali Goldmines clinic. Health facilities are supported by Kibali Goldmines and these provide services to all local residents whether they are employees or not.

The economy has boomed in the area adjacent to the Mine as observed by increased population, development of social and economic facilities, and the large amount of employment created by the secondary economy and local spend by the mine.

Influx

As with all large-scale commercial mining developments, the influx of people into the Project area in search of jobs or to take advantage of the economic growth during construction and operations was an expected outcome of the Project construction and ongoing operations (Digby Wells, 2011). As indicated in Figure 20-1, there has been substantial growth in the local community which now has an established infrastructure and services such as banking. The region’s isolation from other parts of the DRC, particularly to the south, where the Congo Basin extends to the western edge of Rift Valley and the borders with Rwanda and Burundi, has meant that this settlement has become a regional hub which attracts opportunistic job seekers as well as artisanal miners from other parts of the DRC, and other parts of Africa. The existing ASM activities in the Project area have resulted in social ills already being present in the Project area.

Durba town, which is adjacent to the mine, received most of the population influx. The mine opened large roads and built key infrastructures such as markets, water points and sports facilities to manage the influx. The original local community was still given preference to ensure they are empowered and maintain leadership in their area of development.

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LOGO

Source: Google Earth images, Kibali Mine Resources

Figure 20-1 Time Series Satellite Imagery of the Kibali Region Indicating Land Transformation from 2003 to

2018

Artisanal and Small-Scale Mining

Despite the potential boost to the local economy, the Project has resulted in a loss of economic activity and livelihood for some people, particularly those previously benefiting from the artisanal and ASM economy within the moratorium zone, including those people employed by SOKIMO (Digby Wells, 2011). Although the ASMs are bound, in terms of the DRC legislation, to make way for the Project as the legally entrenched industrial mining enterprise, they had the potential to be a formidable obstacle to orderly resettlement and development. The closure of ASM pits not only affects people directly involved in ASM activity, but it also has a knock-on effect on small businesses and farmers who sell goods to the ASM miners in the village. This impact was most experienced by villages that relied on ASM pits that were closed by the Project, but who resided some distance away from the Moratorium Zone and were consequently, not ideally positioned to benefit from the economic opportunities of the Project.

For many of the resettled communities, ASM represented a significant proportion of their household income, and households spent more time mining than farming. Livelihood replacement with non-mining activities has therefore been a focus of the RAP activities.

Kibali Goldmines is working with the Provincial authorities to eliminate ASM within the Exploitation Permit. They have been observed to have expanded and are close to the R26 road, which constitutes a threat to the main access road for the Province. The Provincial authorities have signed a number of orders for a moratorium, and a programme for the cessation of ASM in compliance with the DRC Mining Code (2002).

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21	Capital and Operating Costs

Capital and operating costs for Kibali Goldmines are based on extensive experience gained from 8 years of operating this mine and an extensive number of years operating other gold mines situated within Africa. Sustaining (replacement) capital costs reflect current price trends. Operating costs are in line with historical averages. Any potential non-capitalized exploration expenditure has not been included in the economic forecasts.

21.1

Capital Costs

Basis of Estimate

Kibali is an on-going combined open pit and underground mining operation with the necessary facilities, equipment, and manpower in place to produce gold.

The basis for the combined LOM plan is the Proven and Probable Mineral Reserves estimate described in Section 15 of this Technical Report.

In the QP’s opinion, the open pit and underground LOM and cost estimates have been completed in sufficient detail to be satisfied that economic extraction of the Proven and Probable Mineral Reserves is justified.

The majority of the capital cost estimates contained in this report are based on quantities generated from the open pit and underground development requirements and data provided by Kibali Goldmines.

Capital expenditure over the remaining LOM is estimated to be $715 million (from 2022) based on Mineral Reserves, consisting of the following allocation of costs (as defined in Table 21-1):

Grade Control Capital

Grade control capital cost is related to resource conversion and reserve replacement.

Capitalised deferred Stripping

Capitalised deferred stripping covers open pit waste stripping.

Underground Capital Development and Drilling

This category covers the cost of on-going LOM capital ore and waste development. Capital development costs are based on a calculated average cost per metre for development including development of declines, inclines, stockpiles, ventilation drives, grade control platforms, level access drives and ventilation raises.

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RAP Growth Capital

Growth capital include the cost of RAP for Kalimva-Ikamva, Pamao, and Megi-Marakeke.

Other Sustaining Capital

Sustaining capital costs include all sustaining capital, mainly the CTSF and FTSF projects, the cyanide recovery project, and underground sustaining capital (mobile fleet).

A summary of capital requirements anticipated over the LOM based on Mineral Reserves (from 2022) is summarised in Table 21-1.

Table 21-1 LOM Capital Expenditure Based on Mineral Reserves



Description		Value ($M)

Grade control drilling		41

Capitalised deferred stripping		35

Underground Capital Development and Drilling		185

RAP growth capital		18

Drilling Capitalised		6

Other Sustaining capital		430

Total LOM Capital Expenditure		715

The QPs note that all material contracts discussed above are currently in place and the terms contained within the sales contracts are typical and consistent with standard industry practice, and are similar to contracts for the supply of doré elsewhere in the world. All contract terms, rates and charges are within the norms of Barrick’s regional benchmarks, which are generally within the lower half of industry wide standards.

In the opinion of the QPs, the projected capital costs at Kibali are reasonable and are comparable with those of other operations within the Africa & Middle East region.

21.2

Operating Costs

Basis of Estimate

The open pit mining operation is contractor-run by KMS, while underground mining has been owner-operated by Kibali Goldmines since 2018.

The basis for the combined LOM plan is the Proven and Probable Mineral Reserves estimate described in Section 15 of this Technical Report.

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The mining costs used for the 2021 pit optimisations were derived from the KMS 2020 budget unit plan (BUP) and Long-Term Review (LTR) pricing for the Kibali open pit operations. Owner’s costs were also added.

Labour costs for national employees were based on actual costs. Local labour laws regarding hours of work etc. were also considered and overtime costs included.

During 2021, costs for processing and G&A were updated based on actuals adjusted for the latest forward estimates, production profiles and personnel levels. Customs duties, taxes, charges and logistically costs are included.

LOM Operating Costs

Unit costs used to estimate LOM operating costs based on Mineral Reserves (from 2022) are summarised in Table 21-2. The annual fluctuation in production levels is relatively low, such that the effect of fixed versus variable expenses is minimised.

Table 21-2 LOM Operating Unit Costs Based on Mineral Reserves



Activity	Units	Value

Open Pit Mining – Kibali	$/t mined	3.44

Open Pit Mining – Kibali	$/t ore mined	33.00

Underground Mining	$/t mined	36.16

Underground Mining	$/t ore mined	37.95

Processing	$/t milled	17.49

G&A	$/t milled	9.35

Mining Total	$/t milled	35.60

Total LOM Net OPEX	$/t milled	62.44

Notes:

Total LOM Net of Opex in this table, represents the total amount, before capitalised cost and royalty costs of 4.7% based on the total revenue

Kibali Goldmines has used the unit costs to estimate LOM operating costs based on Mineral Reserves (from 2022). Operating costs for the LOM plan are shown in Table 21-3.

Table 21-3 LOM Operating Total Costs Based on Mineral Reserves



Description		LOM Operating Total Cost ($M)

Open Pit Mining		1,219

Underground Mining		1,739

Processing		1,453

Stockpile		13

G&A		776

Total Operating Cost		5,189

Notes:

Total LOM Net of Opex in this table, represents the total amount, before capitalised cost and royalty costs of 4.7% based on the total revenue

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Cost inputs have been priced in real Q4 2021 dollars, without any allowance for inflation or consideration for changes in foreign exchange rates.

The QPs consider the operating cost estimates in the LOM plan to be reasonable and consistent with historical performance.

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22	Economic Analysis

The QP has verified the economic viability of the Mineral Reserves via cash flow modelling, using the inputs discussed in this report.

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23	Adjacent Properties

The Kibali South Exploration Permit is located 2.5 km SW of the KCD pit in an exclusion zone surrounded by the Kibali Exploitation Permit. Kibali South is currently owned by SOKIMO, however, Kibali South was previously owned by Kibali Goldmines and was transferred to SOKIMO in December 2012. The mineralisation is an up-plunge projection of mineralisation below the KCD 9000 lodes and is refractory in nature (Randgold, 2009).

The QP has not independently verified this information and this information is not necessarily indicative of the mineralisation at Kibali.

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24	Other Relevant Data and Information

This section is not relevant to this report. No additional information or explanation is necessary to make this Technical Report understandable and not misleading.

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25	Interpretation and Conclusions

Based on the total synthesis of the above work, the QPs offer the following conclusions:

25.1

Geology and Mineral Resources

QA/QC

Mineral Resources

In the QP’s opinion, the Kibali Mineral Resources top capping, domaining and estimation approach are appropriate, using industry accepted methods. Furthermore, the constraint of underground Mineral Resource reporting to use optimised mineable stope shapes has been deemed to reflect best practice by external project audits. The QP considers the Mineral Resources at Kibali as appropriately estimated and classified.

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25.2

Mining and Mineral Reserves

The QP considers the modelled recoveries for all ore sources and combined process and plant engineering unit costs, used within the Mineral Resource and Mineral Reserve process to be acceptable.

25.3

Processing

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The QP considers the modelled recoveries for all ore sources and the process and plant engineering unit costs applied to the Mineral Resource and Mineral Reserve process to be acceptable.

25.4

Infrastructure

Kibali is a mature operation that has all necessary support infrastructure already in place.

25.5

Environment and Social Aspects

All permits are in place and an Environmental Adjustment Plan has been approved by the DPEM.

Stakeholder engagement is ongoing, and all senior management are involved in regular meetings with the community.

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ASM remains a concern in the Kibali Exploitation Permit area and the mine is working with provincial authorities to prevent and relocate artisanal miners within the permit area.

The QP considers the extent of all environmental liabilities, to which the property is subject, to have been appropriately met.

25.6

Risks

Kibali Goldmines has undertaken analysis of the project risks as summarised in Table 25-1; together with the QPs assessment of the risk degrees and consequences, as well as ongoing/required mitigation measures. The QPs, however, note that the degree of risk refers to their subjective assessment as to how the identified risk could affect the achievement of the Project objectives. Kibali has been in production for nine years and is a mature operation

Risk Analysis Definitions

The following definitions have been employed by the QPs in assigning risk factors to the various aspects and components of the Project:

●

Major – Risks that have a definite, significant, and measurable impact on the economics. This may include basic errors or substandard quality in the basis of estimate studies or

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project definition. These risks can be mitigated through further study and expenditure that may be significant. Included in this category may be environmental/social non-compliance, particularly regarding Equator Principles and IFC PS.

●

In addition to assigning risk factors, the QPs provided an opinion on the probability of the risk occurring during the LOM. The following definitions have been employed by the QPs in assigning probability of the risk occurring:

●

Rare – The risk is very unlikely to occur during the Project life.

●

Unlikely – The risk is more likely not to occur than occur during the Project life.

●

Possible – There is an increased probability that the risk will occur during the Project life.

●

Likely – The risk is likely to occur during the Project life.

●

Almost Certain – The risk is expected to occur during the Project life.

Risk Analysis Table

Table 25-1details the Kibali Risk Analysis as determined by the QPs.

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Table 25-1 Kibali Risk Analysis



Issue	Likelihood	Consequence Rating	Risk Rating	Mitigation

Geology and Mineral Resources – Confidence in Mineral Resource Models	Unlikely	Minor	Low	Additional scheduled infill drilling. Resource model updated on a regular basis using production reconciliation results.

Mining and Mineral Reserves – Open Pit Slope Stability	Unlikely	Moderate	Minor	Continued in-pit monitoring, geotechnical drilling, instrumentation, and continued updating of geotechnical and hydrology models.

Mining and Mineral Reserves – Underground Recovery and Dilution	Possible	Moderate	Low	Change in drilling and blasting practices and paste filling binder to reduce dilution and increase recovery.

Processing - Salts build up in the process water - leading to carbon fouling in the CIL and elution circuits	Possible	Moderate	Low	A full salt and water balance has been completed and tracked in the plant to ensure that correct water dilution into the critical streams of elution is managed with minimum impact on carbon fouling and gold recovery.

Environmental - Groundwater contamination (arsenic) -Tailings failure and Waste Rock	Possible	Major	Low	Manage As levels through feed profile and capture runoff. All high arsenic feed reports to lined tailings facility. Encapsulate and rehabilitate waste dumps. Continuing monitoring and external or third-party audits.

Social – Social License to Operate	Possible	Moderate	Moderate	Dedicated community engagement by company social and sustainability department. Accessible Grievance Mechanism

Country & Political – Security – Governmental	Possible	Major	Moderate	Dedicated government liaison team in Kinshasa. Government participation/ownership.

Capital and Operating Costs	Unlikely	Moderate	Low	Continue to track actual costs and LOM forecast costs, including considerations for inflation and foreign exchange.

Fiscal Stability	Possible	Moderate	Moderate	Dedicated government liaison team in Kinshasa Government participation/ownership

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26	Recommendations

Based on the total synthesis of the above work, the QPs offer the following recommendations:

26.1

Geology and Mineral Resources

●

Investigate the potential to transfer from explicit strings to implicit lithological modelling.

●

Pamao South domaining generates poor grade distributions and is challenging to link up mineralised wireframes from one section to the next. Extra geological observation and adjustments to short term GC models will be required in 2022 with added GC drilling.

●

Re-establish the regular use of blast movement monitoring in the open pits (OrePro3D) to adjust dig polygons and reduce dilution.

●

Address all outstanding recommendations from the RSC independent audit, as outlined in 14.16 (External Resource Audits).

●

Address each low-risk recommendation and value add comment from the RSC independent audit and collate findings in a presentation and review results at year end 2022.

26.2

Mining and Mineral Reserves

●

Monitor the 90/10 slag cement binder performance over time and ensure that witness QA/QC samples are kept beyond the normal testing period and break at the time of drilling the adjacent stope.

●

Due to the high variability in the 9101 lode stope, ensure that a ratio of no more than 30% to 35% of the 9101 Lode is mined and fed at any given time.

●

Implement bigger chamfer for mining the 9101 and 9101 upper transverse advancing face stope in order to optimize the bogging and improve the stope recovery.

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26.3

Processing

●

Implementation of the cyanide recovery process must be completed to realise process efficiencies in cyanide consumption.

●

Continuous process improvement and geometallurgical work on new satellite orebodies must remain in place to ensure that the plant performance remains optimal for both sulphide and free milling ores.

26.4

Infrastructure

●

26.5

Environment and Social Aspects

●

Continued stakeholder engagement and re-enforcement of the accessibility of the grievance mechanism.

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27	References

Allibone, A., Vargas C., Mwandale E., Kwibisa J., Jongens R., Quick S., Komarnisky N., Fanning M., Bird P., MacKenzie D., Turnbull R., and Holliday J. (2020). Orogenic Gold Deposits of the Kibali District, Neoarchean Moto Belt, Northeastern Democratic Republic

Allibone, A.H., and Vargas, C.A. (2017a). Geological setting and prospectivity of the KZ trend, Moto belt, DR Congo: St. Helier, Jersey, Channel Islands, Randgold Resources Ltd., Unpublished report, 124 p.

ALS Metallurgy (2016). Metallurgical Test Work conducted upon samples from the Gorumbwa Project for Kibali. Report No. A16184

Amtel (2013). Deportment of Gold in Mengu Hill feed and flotation products. Report 12/55

Amtel (2014). Gold deportment analysis of Pakaka major ore types. Report 14/14

Amtel (2014). Gold Deportment in Gorumbwa ores by CN leach. Report 14/42

Amtel (2016). Deportment of Gold in Kibali Sessenge ores. Report 16/38

Amtel (2019). Deportment of Gold in Kalimva & Ikamva ore - 2019. Report 19/39

Amtel (2019). Deportment of Gold in Megi-Marakeke-Sayi ore - 2019. Reports 20/50 and 20/51

Amtel (2020). Deportment of Gold in 3000 lode & 5000 lode DP ore - 2020 -Report 20/41

Amtel (2021). Ore characterisation – Pamao extension & low recovery zone - 2021-Report 21/51

Applied Geomechanics Consulting (2020). Life-of-mine deformation & stability assessment for Kibali.

Beck Engineering (2014). Kibali Numerical Modelling Base Case Simulation. Letter Report dated 16 December 2014.

Beck Engineering (2015). Numerical Simulation of Kibali MHS. Report dated 10 September 2015.

Beck Engineering (2017). Global Deformation Modelling at Kibali. Report dated 22 January 2017.

Beck Engineering (2018a). Life of Mine Deformation & Stability Assessment for Kibali. Report dated 04 July 2018.

Beck Engineering (2018b). Assessment of Revised Life of Mine Sequence for Kibali. Letter Report dated 04 July 2018.

Bird, P.J., 2016, Evolution of the Kibali granite-greenstone belt, northeast Democratic Republic of the Congo, and controls on gold mineralisation at the Kibali gold deposit: Unpublished Ph.D. thesis, London, Kingston University, 307 p.

CIM (2014). CIM Definition Standards for Mineral Resources and Mineral Reserves, adopted by the CIM Council on May 10, 2014.

CIM (2019). CIM Definition Standards for Mineral Resources and Mineral Reserves (MRMR) Best Practice Guidelines 2019.

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Coffey Mining (2013). Kibali Gold SA 3D Mine Wide Numerical Modelling – Stage 1. Report dated 11 March 2013.

Coffey Mining (2014). Kibali Gold SA 3D Mine Wide Numerical Modelling – Stage 2. Report dated 31 March 2014.

Cube Consulting Pty Ltd (2009), Amended and Restated Technical Report (Ni 43-101), Moto Gold Project Democratic Republic of Congo for Moto Goldmines Ltd. April 2009

Dempers & Seymour (2012). Kibali Project Mining Rock Mass Model. Report dated November 2012.

Dempers & Seymour (2014). Kibali Project Mining Rock Mass Model Update. Report dated November 2014.

Dempers & Seymour (2015). Kibali Project Mining Rock Mass Model Update. Report dated March 2015.

Dempers & Seymour (2017). Kibali Project Mining Rock Mass Model Update. Report dated March 2017.

Dempers & Seymour (2017). Kibali Project Mining Rock Mass Model Update. Report dated September 2017.

Dempers & Seymour (2018). Kibali Project Mining Rock Mass Model Update. Report dated December 2018.

Digby Wells Environmental (2011). Kibali ESIA 2011.

Digby Wells Environmental (2013). Resettlement Action Plan Audit Report March 2013.

Digby Wells Environmental (2015). Kibali Stakeholder Engagement Plan July 2015.

Digby Wells Environmental (2015). Resettlement Policy Framework and Action Plan for the Ambarau Hydropower Project. Kibali Gold Mine - Ambarau Hydropower Project.

Digby Wells Environmental (2016). Kibali ESIA Megi Update 2016.

Digby Wells Environmental (2016). Site Report - Gorumbwa Resettlement Review, and Monitoring and Evaluation Framework Development.

Digby Wells Environmental (2017). Annual Closure Cost assessment, 2017

Digby Wells Environmental (2020). Kibali Gold Mine Project Update and Assessment of the Kalimva-Ikamva Satellite Pits, Environmental and Social Impact Assessment Report. Report dated October 2020.

Digby Wells Environmental (2021). Annual Closure Cost Assessment.

Hutchinson, D.J. and Diederichs, M.S. (1996). Cablebolting in Underground Mines. Bi Tech Publishers, Richmond.

ISO 14001:2015 (EMS) Certificate, Feb 2018.

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Jongens, R., Allibone, A.H., and Vargas, C.A. (2016). Geology and controls on the location of the 5000 and 9000 series lodes, KCD gold deposit, Kibali: St. Helier, Jersey, Channel Islands, Randgold Resources Ltd., Unpublished report, 53 p.

JORC (2004). Australasian Code for Reporting of Exploration Results, Mineral Resources and Ore Reserves, Joint Ore Reserve Committee of The Australian Institute of Mining and Metallurgy, Australian Institute of Geoscientists and Mineral Council of Australia (JORC), 2004.

Kibali Goldmines (2014). Internal Report-Kibali Goldmines Internal Review and Summary of all tests conducted

Kibali Goldmines (2017). Internal Report - Metallurgical Test Work – Pamao_2017

Kibali Goldmines (2017). Internal Report - Pamao BRT and Arsenic Distribution

Kibali Goldmines (2017). Internal Report -Kibali Goldmines Internal Review and Geomet Report

Kibali Goldmines (2018). Kibali Geomet Internal Test Work and Review Report

Kibali Goldmines (2019). Internal Report - Metallurgical Test Work – Kalimva-Ikamva_2019

Kibali Goldmines (2020). Internal Report - Metallurgical Test Work – Aerodrome_2020

Kibali Goldmines (2020). Internal Report - Metallurgical Test Work_Megi-Marakeke-Sayi_2020

Kibali Goldmines (2020). Internal Report - Metallurgical Test Work – Sessenge-KCD Gap_2020

Kibali Goldmines (2020). Internal Report - Metallurgical Test Work – 3000 lode & 5000 lode DP_2020

Kibali Goldmines (2021). Internal Document. Stakeholders register

Kibali Goldmines (2021). Internal Document: Environmental Incident Register 2021.

Kibali Goldmines (2021). Internal Document: Grievance Mechanism Procedure.

Kibali Goldmines (2021). Internal Document: Grievance Register 2021.

Kibali Goldmines (2021). Metallurgical Test Work – Pamao extension & low recovery zone_2021

KSCA Geomechanics Pty Ltd (2012). A Review of the SRK Consulting Kibali Underground Geotechnical Feasibility Study Report (Rev0) dated February 2012.

KSCA Geomechanics Pty Ltd (2016). Kibali Stope Performance Database. Excel spread sheet dated 03 September 2016.

KSCA Geomechanics Pty Ltd (2017). Kibali Gold Mine Stope Performance (Stability Graphs). Report dated April 2017.

KSCA Geomechanics Pty Ltd (2018). Kibali Gold Mine Stope Performance (Stability Graphs). Report dated January 2018

Lawrence, D.M. (2011). Gold deportment studies on the Karagba-Chauffeur-Durba deposit, Kibali, NE Democratic Republic of Congo: St. Helier, Jersey, Channel Islands, Randgold Resources Ltd., Unpublished report, 44

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Maelgwyn Mineral Services Africa (2018). Processing of three samples from the Kibali – Sessenge Pit according to the current Kibali flowsheet - Report 8-008

Maelgwyn Mineral Services Africa (2019). Metallurgical Test Work – Kalimva-Ikamva_2019 - Report N0. 19-059

Maelgwyn Mineral Services Africa (2020). Metallurgical Test Work_Megi-Marakeke-Sayi_2020 - Report N0. 20-197

Maxwell GeoServices Pty Ltd (2020). Kibali Gold Mines Data Validation Audit and Review – July 2020

Mining One (2017). Paste Fill exposure facility analysis for the Kibali Gold Mine, Report dated 6 January 2017.

Moto Goldmines Ltd (2008). Moto Gold Project, Democratic Republic of Congo. Independent Technical Report.

Moto Goldmines Ltd (2009). Amended and Restated Technical Report (NI 43-101). Moto Gold Project, Democratic Republic of Congo.

Optiro Pty Ltd (2017). Memorandum. Mining Observations and Comments – August 2017.

Optiro Pty Ltd (2018a). Kibali Mineral Resource and Ore Reserve process review – August - December 2017.

Optiro Pty Ltd (2018b). Kibali Mineral Resource Validation – December 2017.

Orway Mineral Consultants (2011). Kibali Gold Project – Metallurgical Testwork Report, Kibali joint Venture. Report No. 8541 – 03 Rev A, September 2011.

Orway Mineral Consultants (2012). Mengu Hill Test Work Summary - Report No. 8888 Rev 1

Outotec Research Finland (2016). Metallurgical Performance of the Pakaka Feed Blends in the CIL – Review Relative to Feasibility and Geomet Arsenic domains -Report 15142-ORC-T

Peacocke & Simpson (2017). Pamao Gravity Test Work- Report PS394A to F

Quantitative Group (2013). Mineral Resource Review. Kibali, Democratic Republic of Congo. March 2013.

Randgold Resources Ltd (2009). Technical Report NI 43-101 Kibali Gold Project in the Democratic Republic of Congo.

Randgold Resources Ltd (2010). Technical Report NI 43-101 Kibali Gold Project in the Democratic Republic of Congo.

Randgold Resources Ltd (2018). Technical Report NI 43-101 Kibali Gold Project in the Democratic Republic of Congo.

Resettlement and Development Solutions (RADS) and Digby Wells Environmental 2012. Resettlement Action Plan for the Kibali Gold Mine Project

RSC Ltd (2021). Resource and Reserve Audit – Kibali Gold Mine, DRC. December 2021.

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SRK Consulting (2011). Kibali Underground Geotechnical Feasibility Study report (Rev0) dated November 2011.

SRK Consulting (2021), Dewatering strategy for 3000 & 5000 lode exploration holes, Report dated May 2021.Stenhouse, P., 2020. Kalimva-Ikamva project review, geology and exploration potential. Polyphase Consulting, Unpublished report, 77 p.

Turnbull, R., Allibone, A.H., Fanning, C.M., and Matheys, F., 2017. Geochronology, isotope chemistry, and relative prospectivity of Archean rocks in the northeast Democratic Republic of Congo, central Africa: St. Helier, Jersey, Channel Islands, Randgold Resources Ltd, Unpublished report, 79 p.

Western Australian Department of Mines and Petroleum (1995). Western Australian Mines Safety and Inspection Regulations 1995

Western Australian School of Mines (2011). Stress measurements from oriented core using the acoustic emission method – Kibali Gold Mine. Report to Randgold Resources, April 2011.

Western Australian School of Mines (2012). Stress Measurements from Oriented Core using the Acoustic Emission Method. Report to Dempers & Seymour, December 2012.

Western Australian School of Mines (2020). Stress Measurements from Oriented Core using the Acoustic Emission Method. Report to Kibali Gold Mines SPRL, May 2020

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28	Date and Signature Page

This report titled “Technical Report on the Kibali Gold Mine, Democratic Republic of the Congo” with an effective date of 31 December 2021 and dated 18 March 2022 was prepared and signed by the following authors:


		(Signed) *Rodney B. Quick*

Dated at St. Helier, UK		Rodney B. Quick, MSc, Pr. Sci.Nat
18 March 2022		Mineral Resource Manager and Evaluation
		Executive
		Barrick Gold Corporation

		(Signed) *Simon P. Bottoms*

Dated at London, UK		Simon P. Bottoms, CGeol, MGeol, FGS, FAusIMM
18 March 2022		Senior Vice President, Africa and Middle East,
		Mineral Resource Manager
		Barrick Gold Corporation

		(Signed) *Christopher B. Hobbs*

Dated at St. Helier, UK		Christopher B. Hobbs, CGeol, MSc, MCSM,
18 March 2022		FAusIMM
		Group Resource Geologist
		Barrick Gold Corporation

		(Signed) *Graham E. Trusler*

Dated at Johannesburg, ZA		Graham E. Trusler, MSc, Pr Eng, MIChE, MSAIChE
18 March 2022		CEO
		Digby Wells and Associates Pty Ltd.

		(Signed) *Thamsanqa Mahlangu*

Dated at St. Helier, UK		Thamsanqa Mahlangu, Pr. Eng, PhD
18 March 2022		Head of Metallurgy, Africa and Middle East
		Barrick Gold Corporation

		(Signed) *Shaun Gillespie*

Dated at St. Helier, UK		Shaun Gillespie, Reg Eng Tech, FAusIMM
18 March 2022		Group Planning Manager, Africa and Middle East
		Barrick Gold Corporation

18 March 2022

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		(Signed) *Ismail Traore*

Dated at St. Helier, UK		Ismail Traore, MSc, FAusIMM, M.B. Law, DES
18 March 2022		Group Underground Planning Manager, Africa and
		Middle East
		Barrick Gold Corporation

18 March 2022

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29	Certificate of Qualified Persons

29.1

Rodney B. Quick

I, Rodney B. Quick, MSc, Pr. Sci.Nat, as an author of this report entitled “Technical Report on the Kibali Gold Mine, Democratic Republic of the Congo” (the Technical Report) with an effective date of 31 December 2021 and dated 18 March 2022 prepared for Barrick Gold Corporation, do hereby certify that:

1.	I am Mineral Resource Manager and Evaluation Executive with Barrick Gold Corporation, of the 3^rd floor, Unity Chambers, 28 Halkett Street, St. Helier, Jersey, Channel Islands, UK, OJE2.

2.	I am a graduate of the University of Natal Durban, South Africa in 1993 with a Bachelor of Science Honours degree in Geology, and of Leicester University, UK in 2000 with a Master of Science degree in Geology.

3.	I am registered as a Professional Natural Scientist (400014/05) with the South African Council for Natural Scientific Professions (SACNASP). I have worked as a geologist for 28 years since my graduation. My relevant experience for the purpose of the Technical Report is:

●

Experience in the exploration, evaluation, construction, and production phases of all of Randgold Resources projects since Morila. Leading all project development and evaluation for Randgold since 2009 and for Barrick globally since 2019.

I have read the definition of “qualified person” set out in National Instrument 43-101 (NI 43-101) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfil the requirements to be a “qualified person” for the purposes of NI 43-101.

5.	I visited the Kibali Gold Mine most recently on 06 to 09 July 2021.

6.	I am responsible for Sections 1.1, 1.2, 1.3, 1.10, 2, 4 to 6, 19, and 23 and contributions to Sections 1.14 (Risks), 25.6, and 27 of the Technical Report.

7.	I am not independent of the Issuer applying the test set out in Section 1.5 of NI 43-101, as I have been a full-time employee of Barrick Gold Corporation (previously Randgold Resources) since 1996.

8.	I have had prior involvement with the property that is the subject of the Technical Report, as the Mineral Resource Manager and Evaluation Executive for Barrick Gold Corporation and as a Qualified Person for a NI 43-101 on the property dated 18 September 2018.

9.	I have read NI 43-101, and the sections of the Technical Report for which I am responsible have been prepared in compliance with NI 43-101 and Form 43-101F1.

10.	At the effective date of the Technical Report, to the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 18^th day of March, 2022

(Signed) Rodney B. Quick,

Rodney B. Quick, MSc, Pr. Sci.Nat

18 March 2022

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29.2

Simon P. Bottoms

I, Simon P. Bottoms, CGeol, MGeol, FGS, FAusIMM, as an author of this report entitled “Technical Report on the Kibali Gold Mine, Democratic Republic of the Congo” (the Technical Report) with an effective date of 31 December 2021 and dated 18 March 2022 prepared for Barrick Gold Corporation, do hereby certify that:

1.	I am Senior Vice President, Africa and Middle East, Mineral Resource Manager, with Barrick Gold Corporation, of the 1^st Floor, 2 Savoy Court, Strand, London, WC2R 0EZ, United Kingdom.

2.	I am a graduate of the University of Southampton, UK in 2009 with a Masters of Geology degree.

I am registered as a Chartered Geologist registered (1023769) with the Geological Society of London. I am a current Fellow of the Australasian Institute of Mining and Metallurgy (313276). I have worked as a geologist continuously for 13 years since my graduation from University My relevant experience for the purpose of the Technical Report is:

●

Leading Mineral Resource estimation, mine geology Mineral Reserve estimation and mine planning for all operations within the Barrick Africa & Middle East Region since 2019. Including evaluation of mine projects from preliminary economic assessments to pre-feasibility and feasibility studies across multi-commodity operations, spanning both underground and open-pit production. Practical experience in development, construction and operational management of mine operations. Previously, held positions in exploration and mine geology across Africa, Central Asia, Russia and Australia.

5.	I visited the Kibali Gold Mine most recently on 07 to 09 October 2021.

6.	I am responsible for Sections 1.4, 1.12, 1.13, 1.15 (Geology), 3, 7 to 9, 21, 22, 24, and 26.1 (Geology), and contributions to Sections 1.14 (Risks), 25.6, and 27 of the Technical Report.

7.	I am not independent of the Issuer applying the test set out in Section 1.5 of NI 43-101, as I have been a full-time employee of Barrick Gold Corporation (previously Randgold Resources) since 2013.

8.	I have had prior involvement with the property that is the subject of the Technical Report, as Senior Vice President, Africa and Middle East, Mineral Resource Manager of Barrick Gold Corporation and as a Qualified Person for a NI 43-101 on the property dated 18 September 2018.

9.	I have read NI 43-101, and the sections of the Technical Report for which I am responsible have been prepared in compliance with NI 43-101 and Form 43-101F1.

10.	At the effective date of the Technical Report, to the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 18^th day of March, 2022

(Signed) Simon P. Bottoms

Simon P. Bottoms, CGeol, MGeol, FGS, FAusIMM

18 March 2022

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29.3

Christopher B. Hobbs

I, Christopher B. Hobbs, CGeol, MSc, MCSM, FAusIMM, as an author of this report entitled “Technical Report on the Kibali Gold Mine, Democratic Republic of the Congo” (the Technical Report) with an effective date of 31 December 2021 and dated 18 March 2022 prepared for Barrick Gold Corporation, do hereby certify that:

1.	I am a Group Resource Geologist with Barrick Gold Corporation, of the 3^rd floor, Unity Chambers, 28 Halkett Street, St. Helier, Jersey, Channel Islands, UK, OJE2.

2.	I am a graduate of Cardiff University, UK in 2004, with a Bachelor of Science Honours degree in Exploration and Resource Geology. I am also a post-graduate of Camborne School of Mines, Exeter University, UK in 2005, with a Master of Science degree in Mining Geology.

I am registered as a Chartered Geologist (1012989) with the Geological Society of London. I am a current Fellow of the Australian Institute of Minerals and Metals (321498). I have worked as a geologist continuously for 16 years since my graduation. My relevant experience for the purpose of the Technical Report is:

●

Mineral Resource estimation, pre-feasibility and feasibility studies. Corporate/consultant review and audit of exploration projects and mining operations. Previous Technical Reports for iron, gold and copper worldwide. Held various advanced exploration and mine geology positions across West Africa and in Western Australia.

5.	I visited the Kibali Gold Mine most recently from 07 to 09 October 2021.

6.	I am responsible for Sections 1.5, 1.14 (Mineral Resources), 1.15 (Mineral Resources), 10 to 12, 14, 25.1, and 26.1 (Mineral Resources) and contributions to Sections 1.14 (Risks), 25.6, and 27 of the Technical Report.

7.	I am not independent of the Issuer applying the test set out in Section 1.5 of NI 43-101, as I have been a full-time employee of Barrick Gold Corporation (previously Randgold Resources) since 2018.

8.	I have had prior involvement with the property that is the subject of the Technical Report, as a Group Resource Geologist for Barrick Gold Corporation.

9.	I have read NI 43-101, and the sections of the Technical Report for which I am responsible have been prepared in compliance with NI 43-101 and Form 43-101F1.

10.	At the effective date of the Technical Report, to the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 18^th day of March, 2022

(Signed) Christopher B. Hobbs

Christopher B. Hobbs, CGeol, MSc, MCSM, FAusIMM

18 March 2022

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29.4

Graham E. Trusler

I, Graham E. Trusler, MSc, Pr Eng, MIChE, MSAIChE, as an author of this report entitled “Technical Report on the Kibali Gold Mine, Democratic Republic of the Congo” (the Technical Report) with an effective date of 31 December 2021 and dated 18 March 2022 prepared for Barrick Gold Corporation, do hereby certify that:

1.	I am CEO of Digby Wells and Associates Pty Ltd. of Turnberry Office Pk, 48 Grosvenor Rd, Bryanston, Johannesburg, South Africa 2191.

2.	I am a graduate of the University of KwaZulu-Natal, South Africa in 1988 with a Master of Chemical Engineering degree.

I am registered as a Professional Engineer (920088) with the Engineering Council of South Africa. I am also registered as a Member of the Institution of Chemical Engineers (SAIChE) since1994. I am also registered as a Chartered Chemical Engineer with the Institution of Chemical Engineers, as a fellow of the Water Institute of South Africa, and a lifetime member of the American Society of Mining and Reclamation. I have worked as an engineer for a total of 30 years since my graduation. My relevant experience for the purpose of the Technical Report is:

●

Over 30 years of experience within the mining industry in metallurgical production, research and environmental issues. Working on environmental matters affecting the mining industry for more than 29 years. Having conducted numerous projects and managed processes related to the needs of Kibali Gold Mine through feasibility, construction and operations.

5.	I visited the Kibali Gold Mine most recently on 19 to 23 July 2021.

6.	I am responsible for Sections 1.11, 1.14 (Environment), 1.15 (Environment), 20, 25.5, and 26.5 and contributions to Sections 1.14 (Risks), 25.6, and 27 of the Technical Report.

7.	I am independent of the Issuer applying the test set out in Section 1.5 of NI 43-101.

8.	I have had prior involvement with the property that is the subject of the Technical Report as a Qualified Person for a NI 43-101 on the property dated 18 September 2018.

9.	I have read NI 43-101, and the sections of the Technical Report for which I am responsible have been prepared in compliance with NI 43-101 and Form 43-101F1.

10.	At the effective date of the Technical Report, to the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 18^th day of March, 2022

(Signed) Graham E. Trusler

Graham E. Trusler, MSc, Pr Eng, MIChE, MSAIChE

18 March 2022

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29.5

Thamsanqa Mahlangu

I, Thamsanqa Mahlangu, Pr. Eng, PhD, as an author of this report entitled “Technical Report on the Kibali Gold Mine, Democratic Republic of the Congo” (the Technical Report) with an effective date of 31 December 2021 and dated 18 March 2022 prepared for Barrick Gold Corporation, do hereby certify that:

1.	I am Head of Metallurgy, Africa and Middle East, with Barrick Gold Corporation, of the 3^rd floor, Unity Chambers, 28 Halkett Street, St. Helier, Jersey, Channel Islands, UK, OJE2.

2.	I am a graduate of the University of Zimbabwe in 1993 with a Bachelor of Science (Honours) degree in Metallurgical Engineering and in 2002 with a PhD in Metallurgical Engineering.

I am registered as a Professional Engineer (Pr. Eng) with the Engineering Council of South Africa (ECSA) (Reg. F20070233). I have worked as both a Researcher and Operations/Projects Metallurgist for a total of 28 years since my graduation. My relevant experience for the purpose of the Technical Report is:

●

Experience as a Projects and Operations Metallurgist on various gold plant feasibility, commissioning, and optimisation projects. Experience leading metallurgical studies for preliminary economic assessments, pre-feasibility and feasibility studies for geo-metallurgically complex ore sources in support of operations.

5.	I visited the Kibali Gold Mine most recently on 07 to 09 October 2021.

6.	I am responsible for Sections 1.8, 1.9, 1.14 (Processing), 1.14 (Infrastructure), 1.15 (Processing), 1.15 (Infrastructure), 13, 17, 18, 25.3, 25.4, 26.3, and 26.4 and contributions to Sections 1.14 (Risks), 25.6, and 27 of the Technical Report.

7.	I am not independent of the Issuer applying the test set out in Section 1.5 of NI 43-101, as I have been a full-time employee of Barrick Gold Corporation (previously Randgold Resources) since 2011.

8.	I have had prior involvement with the property that is the subject of the Technical Report, as the Head of Metallurgy, Africa and Middle East for Barrick Gold Corporation.

9.	I have read NI 43-101, and the sections of the Technical Report for which I am responsible have been prepared in compliance with NI 43-101 and Form 43-101F1.

10.	At the effective date of the Technical Report, to the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 18^th day of March, 2022

(Signed) Thamsanqa Mahlangu

Thamsanqa Mahlangu, Pr. Eng, PhD

18 March 2022

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29.6

Shaun Gillespie

I, Shaun Gillespie, Reg Eng Tech, FAusIMM, as an author of this report entitled “Technical Report on the Kibali Gold Mine, Democratic Republic of the Congo” (the Technical Report) with an effective date of 31 December 2021 and dated 18 March 2022 prepared for Barrick Gold Corporation, do hereby certify that:

1.	I am Group Planning Manager, Africa and Middle East with Barrick Gold Corporation, of the 3^rd floor, Unity Chambers, 28 Halkett Street, St. Helier, Jersey, Channel Islands, UK, OJE2.

2.	I am a graduate of the University of Johannesburg, South Africa in 1987 with a NHD (National Higher Diploma) Mining.

3.	I am registered as a Fellow of the Australian Institute of Mining and Metallurgy (333703) and have worked as a mining engineer for a total of 34 years since my graduation. My relevant experience for the purpose of the Technical Report is:

●

Leading open pit Mineral Reserve estimation and mine planning for all operations within the Barrick Africa & Middle East Region since 2019 and within Africa for Randgold since 2013. Including evaluation of open pit mine projects from preliminary economic assessments to pre-feasibility and feasibility studies across multi-commodity operations. Previously, held positions in technical consulting and mine operations management.

5.	I visited the Kibali Gold Mine most recently from 07 to 09 October 2021.

I am responsible for Sections1.6 (Open Pit (OP) and Stockpiles), 1.7 (OP), 1.14 (Mining and Mineral Reserves – OP and Stockpiles), 1.15 (Mining and Mineral Reserves – OP and Stockpiles), 15.1 to 15.4 (OP and Stockpiles), 15.6 to 15.8 (OP), 16.1 (OP and Stockpiles), 16.2, 16.6 (OP), 25.2 (OP and Stockpiles), and 26.2 (OP and Stockpiles) and contributions to Sections 1.14 (Risks), 25.6, and 27 of the Technical Report.

7.	I am not independent of the Issuer applying the test set out in Section 1.5 of NI 43-101, as I have been a full-time employee of Barrick Gold Corporation (previously Randgold Resources) since 2013

8.	I have had prior involvement with the property that is the subject of the Technical Report, as the Group Planning Manager, Africa and Middle East for Barrick Gold Corporation.

9.	I have read NI 43-101, and the sections of the Technical Report for which I am responsible have been prepared in compliance with NI 43-101 and Form 43-101F1.

10.	At the effective date of the Technical Report, to the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 18^th day of March, 2022

(Signed) Shaun Gillespie

Shaun Gillespie, Reg Eng Tech, FAusIMM

18 March 2022

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29.7

Ismail Traore

I, Ismail Traore, MSc, FAusIMM, M.B. Law, DES, as an author of this report entitled “Technical Report on the Kibali Gold Mine, Democratic Republic of the Congo” (the Technical Report) with an effective date of 31 December 2021 and dated 18 March 2022 prepared for Barrick Gold Corporation, do hereby certify that:

1.	I am Group Underground Planning Manager, Africa and Middle East with Barrick Gold Corporation of the 3^rd floor, Unity Chambers, 28 Halkett Street, St. Helier, Jersey, Channel Islands, UK, OJE2.

2.	I am a graduate of the Colorado School of Mines, USA in 2013 with a Master of Science in Mining and Earth Systems Engineering.

3.	I am registered as a Fellow of the Australian Institute of Mining and Metallurgy (334992) and have worked as a mining engineer for a total of 12 years since my graduation. My relevant experience for the purpose of the Technical Report is:

●

Multiple mine planning, mine operations and mine management roles. This includes over eight years’ experience in mine planning of underground gold mines. Technical services manager from 2016 to 2021 at Kibali Gold Mine, Project Manager, and Senior Mine planning and Technical Services Engineer from 2014 to 2016 at Loulo Gold Mine.

5.	I visited the Kibali Gold Mine most recently from 15 to 20 July 2021

I am responsible for Sections 1.6 (Underground (UG)), 1.7 (UG), 1.14 (Mining and Mineral Reserves – UG), 1.15 (Mining and Mineral Reserves – UG), 15.1 to 15.3 (UG), 15.5, 15.6 to 15.8 (UG), 16.1 (UG), 16.3 to 16.5, 16.6 (UG), 25.2 (UG), and 26.2 (UG) and contributions to Sections 1.14 (Risks), 25.6, and 27 of the Technical Report.

7.	I am not independent of the Issuer applying the test set out in Section 1.5 of NI 43-101, as I have been a full-time employee of Barrick Gold Corporation (previously Randgold Resources) since 2014.

8.	I have had prior involvement with the property that is the subject of the Technical Report, as the Group Underground Planning Manager, Africa and Middle East for Barrick Gold Corporation.

9.	I have read NI 43-101, and the sections of the Technical Report for which I am responsible have been prepared in compliance with NI 43-101 and Form 43-101F1.

10.	At the effective date of the Technical Report, to the best of my knowledge, information, and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 18^th day of March, 2022

(Signed) Ismail Traore

Ismail Traore, MSc, FAusIMM, M.B. Law, DES

18 March 2022

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