EX-99.68 69 exhibit99-68.htm EXHIBIT 99.68 B2Gold Corp. - Exhibit 99.68 - Filed by newsfilecorp.com

 



Otjikoto Gold Project
NI 43-101 Technical Report
 
 

Title Page

Project Name: Otjikoto Gold Project
   
Title: NI 43-101 Technical Report Feasibility Study
   
Location: Province of Otjozondjupa, Republic of Namibia

Effective Date of Technical Report: February 25, 2013
   
Effective Date of Mineral Resources: August 2012
   
Effective Date Drilling Database: June 6, 2012

Qualified Persons:

  • Bill Lytle, P.E, M.Sc. (Civil Engineering), B.Sc. (Chemical Engineering), VP – Country Manager, Namibia, B2Gold Corp., responsible for compilation of the NI 43-101 Technical Report Feasibility Study
  • Tom Garagan, P.Geo, BSc. (Hons) Geology, Senior Vice President, B2Gold Corp. responsible for the mineral resource estimate as well as exploration, drilling, sample preparation and analysis and data verification.
  • Alan Naismith Pr.Eng, M. Eng. (Geological Engineering) working for SRK (SA) (Pty) Ltd as a Partner and Principal Rock Engineer is responsible for geotechnical assessments
  • Hermanus Kriel, Pr. Eng., B. Eng., MBL, CP (Mining), CEO and Principal Mining Engineer, working for VBKom – Consulting Engineers (Pty) Ltd, responsible for the mining engineering design and mineral resource to mineral reserve conversion
  • Glenn Bezuidenhout, Pr. Eng, Dip Extractive Metallurgy, FSAIMM, Process Director of DRA Mineral Projects (Pty) Ltd, responsible for the process plant design
  • Graham Smith, Pr. Eng, Bsc Eng (Civil), MD Holly & Associates (Pty) Ltd., Consulting and Structural Engineers, responsible for the civils and infrastructure design
  • Guy Wiild, Pr. Eng, MSc (Civil), Bsc (Civil), Director of Epoch Resources (Pty) Ltd (S. Africa) responsible for the tailings storage facility design
  • Werner Petrick, Certified Environmental Practitioner, Bsc Eng (Civil), M.Env Mgt., Environmental Assessment Manager, SLR Environmental Consulting (Pty) Ltd, compiled and project managed the environmental impact analysis, environmental management plan and mine closure framework.

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Table of Contents

1 SUMMARY 27
  1.1 LOCATION, PROPERTY DESCRIPTION, AND OWNERSHIP 27
1.2 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY 28
  1.3 HISTORY 29
  1.4 GEOLOGICAL SETTINGS AND MINERALIZATION 29
  1.5 EXPLORATION 29
  1.6 DRILLING 30
  1.7 SAMPLE PREPARATION, ANALYSIS AND SECURITY 30
  1.8 MINERAL PROCESS AND METALLURGICAL TESTING 30
  1.9 MINERAL RESOURCE ESTIMATE 31
  1.10 MINERAL RESERVE ESTIMATE 32
  1.11 MINING METHOD 32
  1.12 RECOVERY METHODS 34
  1.13 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL IMPACT 35
  1.14 CAPITAL AND OPERATING COST 36
  1.15 ECONOMIC ANALYSIS 37
  1.16 INTERPRETATION AND CONCLUSIONS 38
  1.17 RECOMMENDATIONS 38
2 INTRODUCTION 39
  2.1 TERMS OF REFERENCE 39
  2.2 SOURCES OF INFORMATION 39
  2.3 PERSONAL INSPECTIONS 42
  2.3.1 Geology and Resources 42
  2.3.2 Geotechnical 43
  2.3.3 Mining and Reserves 43
  2.3.4 Process and Metallurgy 43
  2.3.5 Tailings Facility Site 44
  2.3.6 Infrastructure 44
  2.3.7 Financial Evaluation 45
  2.3.8 Environmental 45
3 RELIANCE ON OTHER EXPERTS 47

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4 PROPERTY DESCRIPTION AND LOCATION 48
  4.1 LOCATION 48
  4.2 TENURE 48
  4.3 SURFACE RIGHTS 49
  4.4 ROYALTIES AND OTHER 51
  4.5 ENVIRONMENTAL LIABILITIES 52
  4.6 PERMITS 52
  4.7 SIGNIFICANT FACTORS AND RISK 53
5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY 55
  5.1 ACCESSIBILITY 55
  5.1.1 Means of access to the property 55
  5.1.2 Surface Rights for operations 55
  5.2 CLIMATE 55
  5.2.1 Wind Speed and Direction 56
  5.2.2 Temperature 57
  5.2.3 Rainfall 58
  5.3 MINING PERSONNEL 59
  5.4 LOCAL RESOURCES 60
  5.4.1 Water Supply and Availability and source of water 60
  5.4.2 AVAILABILITY AND SOURCE OF ELECTRICAL POWER 60
  5.5 INFRASTRUCTURE 61
  5.5.1 Tailings storage area 61
  5.5.2 industrial Waste disposal 62
  5.5.3 mine waste Rock disposal area 62
  5.5.4 Processing plant site 63
  5.5.5 Construction camp 64
  5.5.6 Proximity of the property to a population centre 64
  5.6 PHYSIOGRAPHY 65
  5.6.1 Topography and Elevation 65
  5.6.2 Vegetation 65
6 HISTORY 67
  6.1 PRIOR OWNERSHIP 67

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  6.2 PREVIOUS EXPLORATION 68
  6.2.1 Pre-Avdale 68
  6.2.2 Avdale 68
  6.3 HISTORIC MINERAL RESOURCES AND RESERVES 69
  6.4 HISTORIC PRODUCTION 70
7 GEOLOGICAL SETTING AND MINERALIZATION 71
  7.1 GEOLOGICAL SETTING 71
  7.2 REGIONAL GEOLOGY 74
  7.2.1 Regional Stratigraphy 74
  7.2.2 Regional Structure 76
  7.2.3 Metamorphism 78
  7.3 LOCAL AND PROPERTY GEOLOGY 79
  7.3.1 Lithologies 79
  7.3.2 Alteration 85
  7.3.3 Local Structural Geology 85
  7.4 MINERALIZATION 87
8 DEPOSIT TYPES 97
9 EXPLORATION 100
  9.1 GEOPHYSICAL SURVEYS 100
  9.1.1 Airborne Surveys 100
  9.2 GEOCHEMICAL SAMPLING (SOIL SURVEYS) 106
10 DRILLING 107
  10.1 INTRODUCTION/SUMMARY 107
  10.2 DRILLING METHODS AND EQUIPMENT 107
  10.2.1 Reverse circulation drilling 112
  10.2.2 Diamond core drilling 112
  10.3 DRILLING PROCEDURES 112
  10.3.1 Drilling Orientation AND Spacing 112
  10.3.2 RECOVERY 113
  10.3.3 CORE HANDLING 113
  10.3.4 SURVEY CONTROL 113
  10.4 LOGGING PROTOCOLS 114

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  10.4.1 RC Drill Chip Logging 114
    10.4.2 Diamond drill core logging 114
  10.5 DRILL PLAN AND SECTIONS 115
  10.6 INFILL DRILLING 115
  10.7 METALLURGICAL DRILLING 115
  10.8 GEOTECHNICAL DRILLING 115
  10.9 CONDEMATION DRILLING 116
  10.10 EXPLORATION DRILLING (POTENTIAL RESOURCE EXPANSION) 128
  10.10.1 Down Plunge and East 128
  10.10.2 Main Magnetite zone 128
  10.10.3 Wolfshag Zone 128
  10.11 NEW HOLES AND INFORMATION AFTER DATABASE CLOSE-OFF 131
  10.12 COMMENT ON SECTION 10 131
11 SAMPLE PREPARATION, ANALYSIS AND SECURITY 132
  11.1 SAMPLE COLLECTION 132
  11.1.1 Reverse circulation (“RC”) Samples Collection 132
  11.1.2 Drill Core Samples 134
  11.1.3 Specific Gravity and Bulk Density Measurements 136
  11.1.4 SAMPLE INTEGRITY (Chain of Custody and Security) 136
  11.2   SAMPLE PREPARATION AND ANALYTICAL METHODOLOGY 137
  11.2.1 Gold Fire Assay 137
  11.2.2 Multi-element analysis 139
  11.2.3 Sulphur analysis 140
  11.3 ANALYTICAL AND PREPARATION LABORATORIES ACCREDITATION 140
  11.4 DATA MANAGEMENT 141
  11.4.1 DATABASES 141
  11.4.2 Data Entry and Data Management 142
  11.4.3 Hardcopy Data Organization and Filing 143
  11.5 QUALITY ASSURANCE AND QUALITY CONTROL 143
  11.5.1 Blanks 144
  11.5.2 Certified Reference Materials (Analytical Standards) 145
  11.5.3 DUPLICATES 147
  11.5.4 External Check Laboratories 150

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  11.6 COMMENTS ON SECTION 11 150
12 DATA VERIFICATION 152
  12.1 HISTORIC DATA VERIFICATION PROCEDURES 152
  12.2 DATA VERIFICATION PROCEDURES 152
  12.2.1 SITE VISITS 152
  12.2.2 RESOURCE ESTIMATION REVIEW 153
  12.3 LIMITATIONS 154
  12.4 COMMENTS 154
13 MINERAL PROCESSING AND METALLURGICAL TESTING 155
  13.1 GRINDING AND METALLURGICAL RECOVERY TEST WORK AND SCOPE 155
  13.2 TEST SAMPLES 159
  13.3 TEST WORK RESULTS 163
  13.3.1 Ore Chemical Analysis and Mineralogy 163
  13.3.2 Materials Handling Test Work 169
  13.3.3 Comminution Test Work and Grinding Modelling Results 169
  13.3.4 SGS Lakefield Gravity Concentration and Leach Test Work on Composite Samples 172
  13.3.5 CANMET Leach Optimization Test work 176
  13.3.6 Knelson Gravity Concentration and Intensive Leach Test work and Modelling/Design 178
  13.3.7 Lakefield Recovery Variability Test Work 181
  13.3.8 Carbon Adsorption Test work and Modelling 182
  13.3.9 Settling and Rheology Test Work for Thickener Sizing 183
  13.3.10 SO2/Air Cyanide Destruction Test work and Design 184
  13.3.11 Sampling Test Work– Pitard Test 187
  13.3.12 Tailings Test Work 187
  13.3.13 Plant Recovery 188
  13.3.14 Estimation of Otjikoto Plant Recovery 192
14 MINERAL RESOURCE ESTIMATES 198
  14.1 SUMMARY 198
  14.2 DATABASE 198
  14.3 SOFTWARE 199
  14.4 GEOLOGY 199
  14.5 EXPLORATORY DATA ANALYSIS (EDA) 200

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  14.6 INTERPRETATION AND WIREFRAMING 203
  14.7 DOMAINS 204
  14.8 CAPPING AND COMPOSITING 204
  14.9 BLOCK MODEL PARAMETERS 206
  14.10 LITHOLOGY MODEL 206
  14.11 DENSITY 207
  14.12 METALLURGICAL DOMAINS 207
  14.13 GRADE ESTIMATION 208
  14.13.1 Variography 208
  14.13.2 Search Criteria 209
  14.13.3 Estimation plan 210
  14.14 CATEGORIZATION 210
  14.15 BLOCK MODEL VALIDATION 214
  14.15.1 Visual comparison 214
  14.15.2 Composite and block model statistics 214
  14.15.3 Comparison of estimation methods by easting, northing and elevation 215
  14.16 MINERAL RESOURCE STATEMENT 217
  14.17 COMMENTS ON SECTION 14 220
15 MINERAL RESERVE ESTIMATE 221
  15.1 MINERAL RESERVE CLASSIFICATIONS 221
  15.2 PIT OPTIMIZATION METHODOLOGY 221
  15.3 GEOTECHNICAL PARAMETERS FOR PIT DESIGN 224
  15.4 ULTIMATE PIT DESIGN 224
  15.5 COMPLIANCE WITH THE WHITTLE SHELL 225
  15.6 CUT-OFF GRADE CALCULATION 226
  15.7 OTJIKOTO PROBABLE MINERAL RESERVE 227
16 MINING METHODS 229
  16.1 PIT SLOPE AND BENCH DESIGN PARAMETERS FOR PIT DESIGN 229
  16.2 BENCH HEIGHT AND MINING DIRECTION 230
  16.3 HAUL ROAD DESIGN 230
  16.3.1 Haul road gradient 230
  16.3.2 Haul road width 231
  16.3.3 Surface Haul Roads 232

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  16.4 ULTIMATE PIT DESIGN 232
  16.4.1 Grade-tonnage curve 236
  16.5 ANNUAL PIT DESIGNS 237
  16.6 MINE ROCK STORAGE FACILITY 238
  16.7 MINE EQUIPMENT SELECTION AND SCHEDULE 239
  16.7.1 Blast hole Drilling Equipment 240
  16.7.2 Ore and Waste Loading and Hauling 241
  16.7.3 Secondary and Support Equipment 245
  16.7.4 Surface and Pit Hydrology 246
  16.8 PRODUCTION SCHEDULE 247
  16.9 MINE PERSONNEL SCHEDULE 249
17 RECOVERY METHODS 251
  17.1 PROCESS DESIGN CRITERIA 251
  17.2 PROCESS DESCRIPTION 252
  17.3 LABORATORY 257
18 PROJECT INFRASTRUCTURE 259
  18.1 SITE LAYOUT AND PREPARATION 261
  18.2 INTERNAL AND ACCESS ROADS 261
  18.3 BUILDINGS AND FACILITIES 262
  18.4 TAILINGS MANAGEMENT 264
  18.4.1 Engineering Analysis 266
  18.4.2 Rehabilitation 267
  18.5 HYDRAULIC DESIGNS 267
  18.5.1 Storm Water Dam 268
  18.5.2 Water Storage Dams 268
  18.5.3 Groundwater Hydrology and Dewatering 268
  18.6 WASTE AND WASTEWATER MANAGEMENT 273
  18.7 POWER SUPPLY AND DISTRIBUTION 274
  18.7.1 Main Substation 274
  18.8 DORÉ TRANSPORT 275
  18.9 HEALTH, SAFETY, ENVIRONMENT, AND SECURITY 275
  18.9.1 Industrial Hygiene 275
  18.9.2 Security 275

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18.9.3 First Aid 276
18.9.4 Training 276
18.9.5 Environmental 276
18.10 GENERAL AND ADMINISTRATIVE 276
19 MARKET ANALYSIS 277
19.1 MARKET OVERVIEW 277
19.1.1 Late 2000s Financial Crisis 277
19.1.2 HISTORIC GOLD PRICES 278
19.2 A FEW WORDS ON GOLD 279
19.2.1 Market Volatility 279
19.3 MACRO OUTLOOK 280
Many Reasons to Buy the Dips after US Fiscal Cliff is Settled for the Longer-Term 281
19.4 IMPACT OF OTJIKOTO PROJECT ON GLOBAL GOLD SUPPLY 291
19.4.1 Selection of Gold Price and Exchange rate 291
20 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACTS 294
20.1 INTRODUCTION 294
20.2 PERMITTING 295
20.3 ENVIRONMENTAL IMPACT ASSESSMENT PROCESS 296
20.4 SUMMARY OF ENVIRONMENTAL BASELINE CONDITIONS 297
20.4.1 Climate 297
20.4.2 Surface water 297
20.4.3 Groundwater 298
20.4.4 Air quality 299
20.4.5 Biodiversity 299
20.4.6 Socio-Economic 301
20.4.7 Archaeology 302
20.4.8 Visual 302
20.4.9 Noise 302
20.5 SUMMARY OF KEY ENVIRONMENTAL IMPACTS 303
20.5.1 Topography 303
20.5.2 Surface Water 303
20.5.3 Groundwater 303

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    20.5.4 Air 304
    20.5.5 Noise 304
    20.5.6 Biodiversity 305
    20.5.7 Visual 305
    20.5.8 Archaeology 306
    20.5.9 Traffic 307
    20.5.10 Economic 307
    20.5.11 Social 308
  20.6 ENVIRONMENTAL MANAGEMENT PLAN 309
    20.6.1 Stakeholder consultation/communications 310
    20.6.2 Safety and Security 310
    20.6.3 Surface Water Management Program 311
    20.6.4 Groundwater 311
    20.6.5 Air quality 312
    20.6.6 Noise and vibration 312
    20.6.7 Biodiversity Management 312
    20.6.8 Visual management program 313
    20.6.9 Archaeology Management Program 313
    20.6.10 Traffic Control 313
    20.6.11 Socio-economics Management Program 313
    20.6.12 Resources 314
    20.6.13 Soils 314
    20.6.14 Waste Management 314
  20.7 RECLAMATION 315
21 CAPITAL AND OPERATING COSTS 316
  21.1 CAPITAL COST ESTIMATE 316
    21.1.1 SUMMARY 316
    21.1.2 Basis of Estimate 317
    21.1.3 Direct Costs 318
    21.1.4 Indirect Cost 319
    21.1.5 Working Capital 319
    21.1.6 Contingency 319
    21.1.7 Qualifications and Assumptions 320

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  21.1.8 Pricing 321
  21.1.9 Taxes 321
  21.1.10 Project Currency, Estimate Base Date and Foreign Exchange 321
  21.1.11 Accuracy 322
  21.1.12 Project Implementation 322
  21.1.13 Exclusions 322
  21.2 PROJECT DIRECT AND INDIRECT COSTS 323
   21.2.1 Quantities and Unit Pricing 323
  21.2.2 Earthworks 323
  21.2.3 Concrete, Formwork and Reinforcing Steel 323
  21.2.4 Structural Steel 324
  21.2.5 Mechanical Equipment 324
  21.2.6 Plate Work, Linings and Tanks 324
  21.2.7 Piping and Valves 325
  21.2.8 Electrical, Control and Instrumentation 325
  21.2.9 Buildings 325
  21.2.10 Belt Conveyors 326
  21.2.11 Freight 326
  21.2.12 Engineering, Procurement and Construction Management 326
  21.2.13 First Fills and Spares 327
  21.2.14 Preliminary and General 327
  21.3 SURFACE MINE 327
  21.4 PROCESS FACILITIES 329
  21.5 ELECTRICAL & POWER GENERATION 329
  21.6 TAILINGS STORAGE FACILITY 330
  21.7 CONSTRUCTION EQUIPMENT 331
  21.8 EARTHWORKS 333
  21.9 MINE INFRASTRUCTURE, MINE BUILDINGS 333
  21.10 BUILDINGS AND ANCILLARY FACILITIES 334
  21.11 MECHANICAL & ELECTRICAL SPARES 335
  21.12 OWNER’S COSTS 336
  21.13 OWNERS CONSTRUCTION MANAGEMENT 338
  21.14 ENGINEERING, PROCUREMENT AND CONSTRUCTION MANAGEMENT 338

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21.15 CONTINGENCY 339
21.16 PROCESS PLANT OPERATING COSTS 345
21.16.1 Process Plant Operating Cost Risk Analysis 345
21.16.2 Site General Operating Costs 354
22 ECONOMIC ANALYSIS 356
22.1 CASH FLOW BASIS 357
22.1.1 Mineral Reserves 357
22.1.2 Mine Permitting and Development Schedule 358
22.1.3 Mine Plans and Schedules 358
22.1.4 Metals Production 358
22.1.5 Transport, Treatment and Refining Costs 359
22.1.6 Royalties and Taxes 359
22.1.7 Operating Costs 360
22.1.8 Summary of Parameters 360
22.2 PROJECT CASH FLOW 362
22.3 CASH FLOW ANALYSIS 376
22.3.1 Economic Result Summary 376
22.4 SENSITIVITY ANALYSIS 376
23 ADJACENT PROPERTIES 379
23.1 MINERAL OCCURRENCES 379
23.2 OKORUSU FLUORSPAR MINE 380
23.3 KOMBAT COPPER-LEAD MINE 380
23.4 TSUMEB (ONGOPOLO) LEAD-COPPER-ZINC-SILVER MINE 381
24 OTHER RELEVANT DATA AND INFORMATION 382
25 INTERPRETATION AND CONCLUSIONS 383
26 RECOMMENDATIONS 385
27 REFERENCES 387
28 DATE AND SIGNATURE PAGES 390

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LIST OF FIGURES

Figure 1.1: Otjikoto ultimate pit design 33
Figure 1.2: Otjikoto ultimate pit and internal phase designs 34
Figure 4.1: General Location Map of Otjikoto Project 49
Figure 4.2: General Location map of Otjikoto Project showing location of mineral licenses 50
Figure 4.3: Legal Survey diagram of Otjikoto Mining License ML 169 54
Figure 5.1: Average Wind Frequency and Velocity Over the Period 8 June 2007 to November 2011 57
Figure 5.2: Monthly Average Temperatures for the Otjikoto Project for the Period 8 June 2007 To November 2011 58
Figure 5.3: Total Monthly Rainfall as recorded at the Otjikoto Project for the Period 8 June 2007 to November 2011 59
Figure 5-4: Otjikoto Mine Rock Storage Facility 63
Figure 7.1: Pan-African mobile belt and location of major mineral deposits 71
Figure 7.2: Geological Map of Namibia (from Geological Survey of Namibia) 72
Figure 7.3: Locality plan of Otjikoto deposit area (indicated by red block) within the tectonostratigraphic Northern and Northern Central Zones 73
Figure 7.4: Stratigraphic column of the northern Damara Orogen lithologies, indicating the positions of the Otjikoto and Navachab gold mineralization 75
Figure 7.5: Regional Geology of the Otjikoto Deposit area (B2Gold licenses outlined in red) 77
Figure 7.6: General outline of inferred “Otjikoto” batholith (thick white dashed lines) (Corner, 1996) and smaller dome complexes (thin white dashed lines) in the Otjikoto region (backdrop First Vertical Derivative Magnetics) 78
Figure 7.7: Local Stratigraphy of the Otjikoto Deposit area 80
Figure 7.8: Local Geology of the Otjikoto Deposit 82
Figure 7.9: Examples of some of the principal lithologies in the deposit 84
Figure 7.10: Schematic section looking north through main deposit showing distribution of principal lithologies 86
Figure 7.11: Summary of faults in the Otjikoto Deposit area 88
Figure 7.12: Analytical Signal airborne magnetics showing strong regional anomaly associated with the Otjikoto deposit area 89
Figure 7.13: Simplified cross-section through the Otjikoto mineralization hosted by sheeted sulphide- quartz veins within the Okonguarri Formation 89
Figure 7.14: General location of the mineralized zones and principal mineralogy of the veins 90
Figure 7.15: Location of ore shoots at Otjikoto 92
Figure 7.16: Summary Detailed log through the Otjikoto mineralization. 93
Figure 7.17: Schematic longitudinal section showing continuity of mineralization (and gold grades) down the West Shoot 94
Figure 7.18: Examples of principal vein mineralogy assemblages: 95
Figure 7.19: Plates of examples of Visible Gold occurrences (Po – pyrrhotite; Alb – albitite; Grt – garnet; carb –carbonate; Qtz – quartz) 97
Figure 9.1: High resolution NRG magnetics – Analytical Signal 101

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Figure 9.3: SPECTRUM AEM channel 5 and EM conductor picks 104
Figure 9.4: Ground-induced polarization (IP) survey – Chargeability image 105
Figure 10.1: Location of Otjikoto RAB Holes (all pre B2Gold) 109
Figure 10.2: Location of Otjikoto Diamond Drill Holes (“DDH”) 110
Figure 10.3: Location of Otjikoto RC Holes 111
Figure 10.4: Drill plan map 117
Figure 10.5: Schematic Longitudinal Section 4675E, showing Bloy (Nicholls, 2011) 0.4 g/t gold grade shell (orange and blue lines indicate 2011 pit shell iterations) (All holes are pre B2Gold) . 118
Figure 10.6: Schematic Longitudinal Section 4725E, showing Bloy (Nichols, 2011) 0.4 g/t gold grade shell (orange and blue lines indicate 2011 pit shell iterations) (All holes are pre B2Gold) . 119
Figure 10.7: Schematic Longitudinal Section 4850E, showing Bloy (Nichols, 2011) 0.4 g/t gold grade shell (orange and blue lines indicate 2011 pit shell iterations) (All holes are pre B2Gold) . 120
Figure 10.8: Schematic Cross Section Line 7050N, showing Bloy (Nicholls, 2011), 0.4 g/t gold grade shell (orange and blue lines indicate 2011 pit shell iterations) (All holes are pre B2Gold) . 121
Figure 10.9: Schematic Geological Cross Section Line 7250N 122
Figure 10.10: Schematic Cross Section Line 7250N, showing Bloy (Nicholls, 2011) 0.4 g/t gold grade shell (orange and blue lines indicate 2011 pit shell iterations) (All holes are pre B2Gold) . 123
Figure 10.11: Schematic Geological Cross Section Line 7400N 124
Figure 10.12: Location Map of 2012 Metallurgical test sample drill holes 125
Figure 10.13: Location Map of 2012 Pit Stability Study Geotechnical drill holes 126
Figure 10.14: Location Map of 2012 Condemnation drill holes (backdrop DRA site layout L002- SHT_180) 127
Figure 10.15: General Location Map of Wolfshag (K2) Zone 129
Figure 10.16: Wolfshag (K2) Zone Section 7900N (Values after Auryx Gold news release Nov 23, 2011 and B2Gold news release Dec 11, 2012) 130
Figure 11.1: RC Sampling flow chart 133
Figure 11.2: QQ plots of Fire Assay (FA) versus Screen analysis results (gold g/t) 139
Figure 11.3: Genalysis Blank Performance 2011 and 2012 144
Figure 11.4: Example of CRM charting (2012) – “Low Grade” CRM OREAS 15g 145
Figure 11.5: Example of CRM charting – “High” Grade CRM’s OREAS 17c 146
Figure 11.6: Example of CRM charting – Low Grade CRM’s 147
Figure 11.7: Scatter plots of various duplicates 148
Figure 11.8: Thompson Howarth plot of Otjikoto duplicates 148
Figure 11.9: Check sample analyses of historic samples 150
Figure 13.1: Metallurgical Test Work Programme Flow Chart 158
Figure 13.2: Plan Map of sulphide species in the Otjikoto deposit 161
Figure 13.3: Metallurgical Recovery Composites Test Work Flowsheet 173
Figure 13.4: Evaluation of reproducibility of leach kinetics and gold extraction for the selected optimum leach conditions for the XR3 ore sample 177
Figure 13.5: Overall Gold Recovery Contributions from Variability Test Work 189
Figure 13.6 Total GRG Recovery as a function of head grade from Variability Test Work 190
Figure 13.7: XR1 Monte Carlo Probability Distribution of Variability Test Work Recovery 191
Figure 13.8: XR2 Monte Carlo Probability Distribution of Variability Test Work Recovery 191
Figure 13.9: XR3 Monte Carlo Probability Distribution of Variability Test Work Recovery 192

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Figure 13.10: Estimated Full Scale Plant Recovery Distribution for XR1, XR2 and XR3 193
Figure 13.11: Overall Gold Recovery as a function of Head Grade for the XR1 Variability Composites 194
Figure 13.12: Overall Gold Recovery as a function of Head Grade for the XR2 Variability Composites 194
Figure 13.13: Overall Gold Recovery as a function of Head Grade for the XR3 Variability Composites 195
Figure 13.14: Calculated Au Recovery based on Empirical Model Fit and Mine Production Schedule for XR1 196
Figure 13.15: Calculated Au Recovery based on Empirical Model Fit and Mine Production Schedule for XR2 196
Figure 13.16: Calculated Au Recovery based on Empirical Model Fit and Mine Production Schedule for XR2-3 (XR4) 197
Figure 13.17: Calculated Au Recovery based on Empirical Model Fit and Mine Production Schedule for XR3 197
Figure 14.1: Summary geological section 6700N looking northeast 200
Figure 14.2: Sample long section 4750E looking southeast 200
Figure 14.3: EDA – Assays by lithology 201
Figure 14.4: Albitite with gold in vein bins 202
Figure 14.5: Hornfels with gold in vein bins 202
Figure 14.6: Schematic showing general location of domains (looking north) 204
Figure 14.7: Composite statistics by domain 205
Figure 14.8: Metallurgical domains 208
Figure 14.9: Resource categorization 213
Figure 14.10: Swath plot comparison of models 216
Figure 14.11: Grade tonnage curves 219
Figure 15.1: Plan view of the Otjikoto pit design 225
Figure 16-1: Two-lane traffic haul road configuration 232
Figure 16-2: Otjikoto ultimate pit design 234
Figure 16-3: Otjikoto ultimate pit and internal phase designs 235
Figure 16.4: Otjikoto Pit Grade-Tonnage Curve 237
Figure 16.5: Otjikoto pit progression map, year 3 238
Figure 16.6: Otjikoto pit progression map, year 6 238
Figure 16.7: Otjikoto Mine Rock Storage Facility 239
Figure 16.8: Otjikoto Primary Fleet 245
Figure 16.9: Secondary Fleet 246
Figure 16.10: Otjikoto pit pushback schedule 247
Figure 16.11: Otjikoto mine labour over the life of mine 249
Figure 17.1: Otjikoto Mill Flowsheet 254
Figure 18.1: General facilities arrangement 259
Figure 18.2: Construction Schedule 260
Figure 18.3: Processing and Administration facility 263
Figure 18.4: Mine facilities 265
Figure 18.5: Tailings Storage facility 267

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Figure 19.1: 15-year and 12-month historical gold prices (source: www.goldprice.org) 278
Figure 19.2: Gold Price Range Bound (Price in $/oz) 279
Figure 19.3: Commodities Follow Global Growth – Moving up 281
Figure 19.4: China Biggest Incremental Consumer 282
Figure 19.5: China Real Estate Turning a Corner 283
Figure 19.6: Holding on Support for now 284
Figure 19.7: Inflation Expectations to Continue to Lift Gold Higher in 2013 285
Figure 19.8: Broad USD Index Short term Issue for Gold – Big Boost longer-term 286
Figure 19:9: Gold ETF Holdings 286
Figure 19.10: More Fed Bond Buying Spells Higher Gold Price, eventually 287
Figure 19.11: Lending/Borrowing Appetite 277
Figure 19.12: Gold Reserves in Emerging Markets 287
Figure 19.13: 24-month historical gold price (source: www.goldprice.org) 291
Figure 19.14: 24-month historical ZAR: USD exchange rate (source: www.oanda.com) 292
Figure 20.1: Maximum, minimum and average monthly temperatures. 297
Figure 21.1: Estimated Plant Opex based on mill power and reagent consumptions from variability test work 346
Figure 21.2: Estimated Plant Opex Variability for XR1 347
Figure 21.2: Factors Affecting Plant Opex Variability for XR1 347
Figure 21.3: Estimated Plant Opex Variability for XR2 348
Figure 21.4: Factors Affecting Plant Opex Variability for XR2 348
Figure 21.5: Estimated Plant Opex Variability for XR3 349
Figure 21.6: Factors Affecting Plant Opex Variability for XR3 350
Figure 21.7: Estimated Plant Opex Variability for XR2-3 (XR4) 350
Figure 22.1: Sensitivity of Cash Flow to Gold Price, Diesel Fuel Price, Heavy Fuel Oil Price, Exchange Rate, National Labor Cost 378
Figure 23.1: Mineral Occurrence map of region (after data from MME) 379

LIST OF TABLES

Table 1.1: Climate Data 28
Table 1.2: Mineral Resource Estimate, August 2012 31
Table 1.3: Probable Mineral Reserve Estimate 32
Table 1.4: Pre-Production Capital Costs Summary 36
Table 1.5: Life of Mine Cost of Production Unit Operating Cost Summary 37
Table 1.6: Life of Mine Average Operating Cost per Gold Ounce Produced 37
Table 4.1: Otjikoto license status 50
Table 4.2: Farm Registration 50
Table 5.1: Climate Data 56
Table 5.2: Minimum, Maximum and Average Temperatures for the Otjikoto Gold Project for the Period 8 June 2007 to November 2011 58
Table 5.3: Total Monthly Rainfall for the Otjikoto Project for the Period 8 June 2007 to November 2011 59

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Table 5.4: Rock dump capacities 63
Table 6.1: Summary of previous exploration in the region before 1995 68
Table 6.2: 2005 Inferred Mineral Resources 69
Table 6.3: Summary of previous mineral resources for the Otjikoto deposit 70
Table 9.1: Summary of Airborne Surveys and Imagery 100
Table 10.1: Summary of drilling on the Otjikoto Deposit 108
Table 13.1: Summary of Metallurgical Test work Scope 156
Table 13.2: Metallurgical Testing Sample Requirements 160
Table 13.3: Whole Rock Analysis and Head Assays 164
Table 13.4: Head Analysis of XR1, XR2 and XR3 Metallurgical Composites 165
Table 13.5: Elements ICP-Scan XR1, XR2 and XR3 Metallurgical Composites 166
Table 13.6: Screen Metallics Assay Results for the Metallurgical Recovery Composites 167
Table 13.7: Grindability Test Summary and Overall Statistics 170
Table13.8: Grindability Data Set used for Mill Selection 171
Table 13.9: Summary of the gravity concentration results on composite samples 174
Table 13.10: Results of the intensive cyanidation of gravity concentrates 174
Table 13.11: Results of the Cyanidation tests on Gravity Concentration Tailings 175
Table 13.12: CANMET Leach Optimization Test Conditions 176
Table 13.13: Summary of Optimized Leach Conditions from CANMET Testing 178
Table 13.14: Overall E-GRG and Cyanidation Test Work Recovery for XR1, XR2 and XR3 Composites 179
Table 13.15: Gravity Circuit Modelling Results 180
Table 13.16: Thickener Design Parameters 184
Table 13.17: CND Results of the Gravity Tailings Bulk Leach Product Samples 185
Table 13.18: CND circuit design criteria based on the gravity tailings bulk leach composite samples 186
Table 13.19: CND circuit design criteria based on the gravity tailings variability samples 187
Table 14.1: Otjikoto Mineral Resource Estimate, August 2012 198
Table 14.2: Summary by year and type of holes used in resource database 199
Table 14.3: Resource domains 204
Table 14.4: Capping by Domain 205
Table 14.5: Block model parameters 206
Table 14.6: Lithology Model codes 206
Table 14.7: Lithology Model codes 207
Table 14.8: Density values applied to model 207
Table 14.9: Variogram parameters 209
Table 14.10: Search criteria 209
Table 14.11: Resource categorization criteria 211
Table 14.12: Comparison of composite and block model statistics 215
Table 14.13: Undiluted Mineral Resources (unconstrained by optimized pit) (*base case highlighted in pink) 217
Table 14.14: Undiluted resource within a $US1,350/oz. pit (*base case highlighted in pink) 217
Table 14.15: Whittle pit key parameters for constrained in-pit resource 218
Table 15.1: Pit optimization input parameters 222

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Table 15.2: Otjikoto Pit slope architecture 224
Table 15.3: Otjikoto pit design operational constraints 224
Table 15.4: Comparison between the ultimate pit design and the optimal Whittle shell 226
Table 15.5: Break-even cut-off grades 227
Table 15.6: Probable Mineral Reserve, effective date December 2012 228
Table 16.1: Otjikoto Pit slope architecture 229
Table 16.2: Otjikoto ultimate pit design ore and waste content - indicated material 236
Table 16.3: Rock dump capacities 239
Table 16.4: Blasthole Drilling Penetration Rates 240
Table 16.5: Global Equipment Parameters 242
Table 16.6: Typical Haul Cycle Detail 244
Table 16.7: Mine Production Schedule 248
Table 16.8: Patterson job grades with associated “cost to company” bands 250
Table 17.1: Otjikoto Major Process Design Criteria 252
Table 17.2: Mill Reagent Consumptions by Ore Type 256
Table 18.1: Monthly Rainfall 269
Table 18.2: Water balance with pit de-watering 272
Table 18.3: Water balance without pit de-watering 273
Table 18.4: General & Administrative Staffing 276
Table 19.1: Commodity Price Forecasts 290
Table 20.1: Waste Inventory list: 315
Table 21.1: Pre-Production Capital Costs Summary 317
Table 21.3: Surface Mine Mobile Equipment Capital Purchase Schedule 328
Table 21.4: Process Plant Initial Capital Purchase 329
Table 21.5: Electrical and Power Generation Initial Capital 330
Table 21.6: Tailings Storage Facility Initial Capital 331
Table 21.7: Construction Equipment Capita Purchase Schedule 332
Table 21.8: Earthworks Capital 333
Table 21.9: Mine Infrastructure and Mine Buildings Capital 334
Table 21.10: Buildings and Ancillary Facilities Capital 335
Table 21.11: Mechanical and Electrical Spares Capital 335
Table 21.12: Owners Costs 337
Table 21.13: Owners Construction Management 338
Table 21.14: Engineering, Procurement and Construction Management (EPCM) 339
Table 21.15: Contingency 340
Table 21.16: Life of Mine Sustaining Capital Costs 341
Table 21.17: Surface Mine Life of Mine Sustaining Capital Costs 341
Table 21.18: Life of Mine Cost of Production Unit Operating Cost Summary 341
Table 21.19: Life of Mine Average Operating Cost per Gold Ounce Produced 342
Table 21.20: Surface Mine Annual Operating Costs 343
Table 21.21: Surface Mine Annual Production 343
Table 21.22: Life of Mine Unit Operating Costs by Activity 344
Table 21.23 a: Life of Mine Unit Operating Costs by Consumable Category 344
Table 21.23 b: Life of Mine Site General Unit Operating Costs by Activity 344

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Table 21.24: Processing Annual Operating Costs 351
Table 21.25: Processing Annual Production 352
Table 21.26: Life of Mine Unit Processing Operating Costs by Activity 353
Table 21.27: Life of Mine Processing Unit Operating Costs by Consumable Category 354
Table 21.28: Site General Annual Operating Costs 355
Table 21.29: Site General Annual Manpower 355
Table 21.30: Life of Mine Site General Unit Operating Costs by Activity 356
Table 22.1: Sensitivity Analysis Parameters 357
Table 22.2: Life of Mine Gold and Silver Production 359
Table 22.3: Typical Transportation and Refining Terms 359
Table 22.4: Life of Mine Cost of Production Unit Operating Cost Summary 360
Table 22.5a: Base Case Parameters Used for the Cash Flow Analysis 361
Table 22.5b: Base Case Parameters Used for the Cash Flow Analysis (continued) 362
Table 22.6: Pro forma Income Statement 365
Table 22.7: Production Summary 366
Table 22.8: Net Production Value 367
Table 22.9a: Summary of Surface Mining Costs 368
Table 22.9b: Summary of Surface Mining Costs (continued) 369
Table 22.10a: Summary of Processing Costs 370
Table 22.10b: Summary of Processing Costs (continued) 371
Table 22.11: Summary of Site General Costs 372
Table 22.12: Summary of Distributed Account Costs 372
Table 22.14: Other Costs 374
Table 22.15: Summary of Reclamation and Closure Costs 374
Table 22.16: Other Operating Costs 375
Table 22.17: Cash Flow Analysis Cases 376
Table 22.18: Economic Results for Gold Price Cases 376
Table 22.19: NPV at 5% Discount - Sensitivity to Net Production Value, Operating Cost and Pre- Production Capital 377
Table 22.20: Sensitivity of Cash Flow to Gold Price, Diesel Fuel Price, Heavy Fuel Oil Price, Exchange Rate, National Labor Cost 377

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ACRONYMS and ABBREVIATIONS

$ United States Dollar
+_ Plus minus
< Less than
= equal to
> Greater than
0 Degrees
0C Degrees Celcius
2D Two Dimentional
3D Three Dimentional
A Amp
AA Atomic Absorption
AAC Anglo American Corporation
AARL Anglo American Research Laboratories
ac alternating current
AEM Airborne Electromagnetic
Ag Silver
AGES AGES (South Africa)
Al Aluminum
AMR Annual Monitoring Report
ANFO Ammonium Nitrate Fuel Oil (blasting agent)
Anglovaal Anglovaal Mining Ltd.
ARM African Rainbow Minerals
ASTM American Society for Testing and Materials
Au Gold
Auryx Auryx Gold Namibia (Pty) Ltd
Avdale Avadale Namibia (Pty) Ltd. (registration number 93/613)
B1 Main National Road in Namibia
Ba Barium
BC Ltd 0824239 BC Limited
Bi Bismuth
BMRE Bloy Mineral Resource Evaluation
Cd Cadmium
Cenored Regional electrical distributor
Cfm Cubic feet per minute
CIL Carbon-in-Leach
CIP Carbon-in-Pulp
cm Centimetre
CND Cyanide Destruction
CNwad Weak acid dissociable cyanide
Co Cobalt
CO2 Carbon Dioxide

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COA Certificate of Analysis
CRM Certified Reference Material
CS Carbon and Sulphur
CSAMT Controlled Source Audio-Frequency Magneto Telluric
CSR Corporate Social Responsibility
csv csv file format
Cu Copper
CV Coefficient of variation
D1 Deformation stage 1
D2 Deformation stage 2
D2808 District road through mine property
D3 Deformation stage 3
DB Database
dBA decibels, A – weighted
dc Direct Current
DDH Diamond Drill Holes
DFS Definitive Feasibility Study
DO Dissolved Oxygen
DRA DRA Mineral Projects (Pty) Ltd
DRC Democratic Republic of the Congo
DTH Down the Hole (drilling)
E East
ECB European Central Bank
ECC Environmental Clearance Certificate
EDA Exploratory Data Analysis
EDS Electronic Data System
EGL Effective Grinding Length
E-GRG Extended Gravity Recoverable Gold
EIA Environmental and Social Impact Assessment
EM Electromagnetic
EMP Environmental Management Plan
EMT Emergency Medical Trauma
EMV Earth Moving Vehicle
EPCM Engineering Procurement and Construction Management
EPL Exclusive Prospecting License
EPOCH Epoch Resources (Pty) Ltd.
ERL Exclusive Reconnaissance Licenses
et al and associates
ETF Exchange Traded Funds
EV Expected mean value
EVI EVI Mining Corporation
EZ Euro Zone
FA Fire Assay
FLS FLSmisth Salt Lake City Inc.

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Ft foot
FW Footwall
G&A General and Administration
g/t grams per tonne
GLC Ground level concentrations
GNI Gross Namibian Income
GPS Ground Positioning System
GRG Gravity Recoverable Gold
HDPE High density polyethylene
HFO Heavy Fuel Oil
HIV Human Immunodeficiency Virus
HLS Heavy Liquid Separation
HQ HQ diamond drill core size (63.5 mm)
HW Hanging wall
ICP Inductively Coupled Plasma
ICP-AES Inductively Coupled Plasma – Atomic Emission Spectrometry
ICP-MS Inductively Coupled Plasma Mass Spectrometry
ID2 Inverse distance squared
ID3 Inverse Distance to the power of three
IEC International Electrotechnical Commission
IP Induced Polarization
IRR Internal Rate of Return
ISO International Organization for Standardization
JJL Jenike & Johanson Limited
JSE Johannesburg Stock Exchange
JV Joint Venture
k kilo – 1 thousand
K Potassium
kg Kilogram
kg/t kilogram per tonne
km Kilometer
KOP Key observation point
kV kilo Volt
KVA kilo Volt Amps
kW Kilowatt
kWh/t kilowatt hours/tonne
kWhr kilo watt hour
L Left
l Litre
L/m Liters per minute
LA, 90 Measure of noise level with background noise lower than the LA level 90% of
the time
LECO LECO Corporation carbon and sulphur analyzer
LG Low Grade

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LIDAR Light Detection and Ranging system for topographic surveys
LM2 and LM5 Labtech Essa pulverizers (2 kg and 5 kg bowl size respectively)
LoM Life of Mine
M Mega – 1 million
m metre
m/s meters per second
Ma Million years
mamsl metres above mean sea level
MAWF Ministry of Agriculture and Forestry
MBS Mortgage – Backed Security
MCC Motor Control Center
MCF Mine Closure Framework
MET Ministry of Environment and Tourism
Mg Magnesium
ML Mining License
mm millimeter
mm millimeter
Mm3/ Mega cubic meters per annum
MME Ministry of Mines and Energy
MMII Metso Mineral Industries Inc.
Mo Molybdenum
MP Management plan
MPs Management Programs
msec millisecond
Mtpa Million tonnes per annum
Mtph metric tonnes per hour
Mtpy Mega tonnes per year
MW Megawatt
MWT Ministry of Works and Transport
N North
N/A Not Applicable
NaCN Sodium Cyanide
NAD Namibian Dollar
NATA National Association of Testing Authorities
NCZ Northern Central Zone
NI National Instrument
Ni Nickel
NN Nearest Neighbour
NOx Nitrous Oxides
NP Northern Platform
NPV Net Present Value
NQ NQ diamond drilling core size (47.6 mm)
NRG New Resolution Geophysics airborne survey contractors
NSAMT Natural Source Audio Magnetotellurics

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NZ Northern Zone
Ø Diameter
OEM Original Equipment Manufacturer
ohm electrical resistance measurement
OK Ordinary Kriging
OREAS ORE Research and Exploration
oz. Ounce
P Phosphorus
P&G Preliminary and General
PBoC Peoples Bank of China
Pd Palladium
PGMs Platinum Group Metals
PM Particulate matter
Po Pyrrhotite
ppb parts per billion
PPE Personal Protective Equipment
ppm parts per million
PQ PQ drill core size (85 mm)
Pt Platinum
PVC Polyvinyl chloride
Py Pyrite
QA/QC Quality Assurance and Quality Control
QE Quality Evaluation
QEMSCANTM Quantitative Evaluation of Materials by Scanning Electron Microscopy
QP Qualified Person
R Right
RA Roads Authority
RAB Rotary Air Blast
RC Reverse Circulation
REE Rare Earth Element
ROI Return on Investment
ROM Run of Mine
RRR Reserve Requirement Ratio’s
RSG RSG Mining (now Coffey Mining group)
S South
S Sulphur
SAB SAG Mill/Ball Mill
SABC SAG Mill/Ball Mill/Pebble Crusher
SABLE Standardized Approach Borehole Log Evaluation
SAG Semi-Autogenous grinding
SANAS South African National Accreditation System
SANS South African National Standards
Sb Antimony
scfm square cubic foot per minute

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SD Standard Deviation
SEP Stakeholder Engagement Plan
SG Specific Gravity
SHE Safety, Health and Environment
SLR SLR Consulting Environmental Inc.
SMBS Sodium Metabisalphite
SMP Scorpion Minerals Processing
SMU Selective Mining Unit
Sn Tin
SO2 Sulphur Dioxide
SRK Steffan Robertson & Kirston (Pty) Ltd
SWA South West Africa
t/h tonne/hour
TB Tuberculosis
TCLP Toxicity Characteristic Leaching Procedure
TDF Tailings Dam Facility
Te Tellurium
Teal Teal Exploration and Mining
TNW TNW diamond drill core size (60.8 mm)
TOF Table of Failures
Tova Tova Ventures Inc.
TSF Tailings Storage Facility
TSX Toronto Stock Exchange
UCS Unconfined Compressive Strength
USD United States Dollar
V Vanadium
V Volt
VAT Value Added Tax
VBKOM VBKOM Consulting Engineers Namibia
W Tungsten
W West
w/w Weight by weight concentration
WHO World Health Organization
XR2 Pyrite (+marcasite) dominant mineralization (Transition Zone)
XR3 Pyrrhotite dominant mineralization (Deeper Sulphide Zone)
XRD X-Ray Diffraction
XRF X-Ray Fluorescence Spectrum
XRI Oxide Zone
ZAR South African Rand
Zn Zinc
μm micron

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1 SUMMARY

DRA Mineral Projects (“DRA”) was appointed by B2Gold Corp. (“B2Gold”) to complete a Definitive Feasibility Study (“DFS”) for the Otjikoto Gold Project (the “Otjikoto Project”), located near Otjiwarongo in Namibia. The major sections of the report were compiled by the following consultancies and companies:

  • B2Gold: Exploration, Mineral Resource, Project History and Financial Analysis
  • VBKOM: Mineral Reserves, Mine Planning, Mine Equipment, and Mine Capital and Operating Costs
  • SRK: Open Pit and Civil Geotechnical Studies
  • DRA : Process Plant Design, Infrastructure, Capital Costs and Operating Costs
  • EPOCH: Tailings Storage Facility
  • SLR: Environmental Impact Analysis
  • AGES: Hydrological Studies

The DFS was initiated to develop an appropriate single-option process plant and mine solution to confirm the business case for the selected solution. The DFS commenced in April 2012 and was completed in December 2012.

1.1 LOCATION, PROPERTY DESCRIPTION, AND OWNERSHIP

The Otjikoto Project is located approximately 70 kilometres from the town of Otjiwarongo and 50 kilometres from the town of Otavi within the Province of Otjozondjupa in the north-central part of the

Republic of Namibia. The project site is 300 kilometres north of Windhoek, the country’s capital, and centred on the farms Otjikoto and Felsenquelle at approximately 718,900 East and 7,787,100 North in WGS84 Zone 33 South coordinate system (20°00’ South (Longitude) and 17°05’ East (Latitude)).

In 2011, the farms Wolfshaag, Otjikoto, Gerhardshausen and Okaputa Nord I were purchased and consolidated by Auryx Properties Holdings (Pty) Ltd. On July 19, 2012, Auryx Property Holdings (Pty) Ltd. was legally renamed to B2Gold Namibia Properties (Pty) Ltd. The Mining License and all proposed infrastructure are situated on the B2Gold Namibia Properties (Pty) Ltd. farms.

The mining license ML 169 is situated within the Exclusive Prospecting License (“EPL”) 2410. EPL 2410 covers an area of 54,125 ha (inclusive of ML 169) and is in good standing, with renewal for an additional two years granted by the Ministry of Mines and Energy (“MME”) on September 14, 2012. Exploration is conducted under the terms of an Environmental Clearance Certificate (“ECC”) issued by the Ministry of Environment and Tourism (“MET”) on June 20, 2002. A renewal ECC application was submitted to the MET on January 31, 2013.

B2Gold Corp. indirectly holds a 92% interest in the Otjikoto Project with EVI Mining Company Ltd. (“EVI”), a Namibian black empowerment company, holding the remaining 8%.

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1.2 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

Vehicle access to the mine will be off the main B1 road, a primary paved road. Mine related traffic will travel from both the Otavi and Otjiwarongo directions on a daily basis with all traffic using the B1 road to access the site.

The Otjikoto Project is located approximately 300 km inland from the west coast of Namibia. The area is characterized by low rainfall with extreme temperature ranges and unique climatic factors influencing the natural environment and biodiversity. In general, the climatic conditions at the Otjikoto Project site allow for year-round construction and mine operations. A summary of the key climatic data is shown in Table 1.1.

Table 1.1:        Climate Data

SITE DATA    
Approx. Elevation / Altitude mamsl 1500
Temperature    
                   Min Annual Temp °C -5
                   Max Annual Temp °C 37.9
Wind    
                   Design Peak Velocity m/s 6
                   Wind Direction N/A ESE (East South East)
Rainfall    
                   Average Annual mm 350-450
Evaporation    
                   Average Annual Rate mm 2400-2600

Water for construction activities will be supplied from boreholes (pre-sunk and equipped prior to construction). Borehole water will be supplied for ablution facilities. Bottled water will be supplied for drinking during the period when the potable supply facility is still not commissioned and in operation. Water for production activities will be supplied from boreholes and the design of the water supply system is based on a geohydrological technical report.

Power for the purposes of construction will be provided by the national grid. Following a trade-off study, it was concluded that due to escalating costs and supply constraints on the side of Nampower (the national energy supplier), power will be produced on site during operations.

The local topography is flat with a gentle slope towards the north-west with freely draining soils. The site is located at an elevation of 1,500-1,510 mamsl, just north of a local surface water divide. There are no well-defined surface water drainage features on the site and no major surface water flow or defined channel flow is expected other than local events after heavy rainfall.

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Vegetation in seasonally wet areas also varies with the degree of water logging. The greater part of the Otjikoto Project area falls within the Tree and Shrub savannah zone, which is listed as the dominant vegetation type in central Namibia.

1.3 HISTORY

The deposit was discovered by Avdale Namibia (Pty) Ltd. (“Avdale”) in 1999 as the result of a base metal exploration program initiated by Anglovaal Mining Ltd (“Anglovaal”) in 1995. Between 1999 and 2011, a series of operators completed numerous airborne and ground geophysical and geochemical surveys and drilled 305 rotary air blast (“RAB”), 458 reverse circulation (“RC”) and 624 diamond drill holes totalling 173,156 metres on the property. Prior to B2Gold acquiring the property in December 2011, a number of previous mineral resource estimates have been reported for the property and are documented in Tables 6.2 and 6.3 of this report.

1.4 GEOLOGICAL SETTINGS AND MINERALIZATION

The Otjikoto Project is located within the Neoproterozoic Damara Mobile Belt, which forms part of the Pan–African Mobile Belt system. The deposit falls under the general classification of an orogenic gold deposit and occurs in a similar stratigraphic position as the Navachab gold deposit, Namibia’s only current operating gold mine.

The deposit is hosted within amphibolite grade metasediments of the Okanguarri Formation. The OTC albitite-hornfels unit hosts most of the mineralized vein system and is underlain by a distinct marker horizon, the un-mineralized OTB calcitic marble.

Gold in the main Otjikoto deposit is hosted by a north-northeast (NNE) striking sheeted sulphide (+ magnetite) - quartz+carbonate vein system. The sheeted veins and related gold mineralization occur in a series of distinct en-echelon zones oriented at approximately 010° to 020° NNE and plunging at 10-15° (average 12°) to the SSW. The bulk of the mineralization occurs in the Central and West shoots. Vein concentrations range from 1 to 30 veins per metre with higher vein concentration within the shoots. The mineralized system has been traced over a strike length of 2.3 kilometers and to a depth of 475 metres below surface.

1.5 EXPLORATION

The bedrock geology within the deposit area is covered by 10 to 15 metres of calcrete, with only sporadic outcrops of Karibib Marble. The exploration program therefore had to rely on a combination of airborne and ground geophysical surveys to map bedrock geology and identify exploration targets for drill testing. Systematic drilling of the geophysical anomalies led to the discovery of the Otjikoto gold deposit. Geochemical sampling over the deposit was able to locate erratic anomalous values of gold but no coherent anomalies.

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1.6 DRILLING

Drilling at the Otjikoto Project started in 1998 with a shallow RAB drill program testing geophysical targets. The initial RAB program resulted in the discovery of gold mineralization and was followed up with additional RAB drilling and the first diamond drill core hole program in 1999. The third hole of the 1999 program, OT3, is considered the discovery hole of the deposit. Drilling of a more regional nature was followed by delineation drilling to determine the distribution of mineralization and grade continuity within the zones. In the shallower levels of the deposit the majority of drilling was completed with RC. Drill hole spacing in the shallower portions of the deposit is 25 metres x 25 metres. Deeper level spacing ranges from 25 metres x 50 metres and 50 metres x 100 metres. A total of 1,554 holes totalling 195,717 metres have been drilled on the Otjikoto and Felsenquelle farms as of October 2012.

1.7 SAMPLE PREPARATION, ANALYSIS AND SECURITY

RC chip samples and drill core are delivered by B2Gold Namibia personnel from the field to the B2Gold Namibia core yard. Following documentation and splitting the samples are shipped via a registered courier to the Genalysis facility in Walvis Bay, Namibia for preparation. Analysis of samples was performed at Genalysis’ laboratory in Johannesburg, South Africa and in the latter half of 2012 at

ALS Minerals laboratory in Vancouver, B.C., Canada.

An extensive quality assurance and quality control (“QA/QC”) program has been in place since the start of the work on the exploration program. This program involves the use of certified reference material (“CRM”) to monitor accuracy, field, preparation and pulp duplicates to monitor precision and blanks to monitor possible contamination and/or sample mix ups. The QA/QC is monitored on a continuous basis and reported in a table of failures. Bureau Veritas is used as the referee laboratory.

1.8 MINERAL PROCESS AND METALLURGICAL TESTING

Feasibility metallurgical testing was performed on Otjikoto drill core samples representing the major ore types to finalize the gravity/whole ore leach process design criteria and flow sheet, and to evaluate ore variability across the deposit. The gravity/whole ore leach process was selected as the preferred mill flow sheet over gravity/flotation/concentrate leach in a trade-off study (SMP, Mining Licence Application) prior to the completion of the DFS.

The following metallurgical test programmes were completed:

  • Characterization of chemical and mineralogical composition of major ore types.
  • Characterization of materials flow properties to determine the stockpile and ore reclaim system design parameters.

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  • Leach and gravity concentration test work on XR1 (oxide), XR2 (pyrite dominant) and XR3 (pyrrhotite dominant) composites of the three major ore types to establish the optimum grind size and produce samples for downstream testing.
  • Grinding test work and modelling/simulation to determine power requirements and sizing for the SAG and ball mills and evaluate variability of ore grind ability.
  • Leach optimization testing to establish optimum leach process parameters for the XR1, XR2 and XR3 ore types.
  • Evaluation of ore variability with optimum leach conditions for recovery modelling and determination of reagent consumptions.
  • Thickener and rheological testing to establish final pre-leach and tailings thickener sizing.
  • SO2 /Air cyanide destruction test work to establish design criteria and reagent consumptions for the cyanide destruction circuit.
  • Carbon adsorption test work and CIP modelling to confirm the CIP circuit design parameters.
  • Sampling test work to determine the design requirements for the mill sampling stations and assay laboratory.
  • Tailings characterization testing.

1.9 MINERAL RESOURCE ESTIMATE

The current resource model represents an update of previous models prepared by RSG Global (van der Merwe et al., 2005), SRK (Wanless et al., 2007; Wanless et. al., 2009; McDonald et. al., 2011) and Bloy Resource Evaluations (“BRME”) (Nicolls, 2011 and 2012).

In the current model, mineralized zone outlines were based on logged lithology, vein percentage and gold grades. The outlines define a nominal 0.4 g/t gold cut-off. Gold grades were estimated into a 3D block model using a mix of ID3 (inverse distance to the power of three) and OK (ordinary Kriging) estimation methods. A single indicator was used to better define high grade mineralization at a threshold of approximately 0.8 g/t gold. The total mineral resource estimate for the Otjikoto Project at a 0.4 g/t gold cut-off within a $1,350/oz. optimized pit is shown in Table 1.2.

Table 1.2:        Mineral Resource Estimate, August 2012

  INDICATED     INFERRED    
Weathering kTonnes Au g/t kOunces kTonnes Au g/t kOunces
Ox/Trans 3,049 1.31 129 93 0.78 2
Sulphide 25,899 1.53 1,276 55 1.83 3
Total 28,949 1.51 1,405 149 1.17 6

Notes:
1) Mineral resources that are not mineral reserves do not have a demonstrated economic viability.
2) Due to the uncertainty that may be attached to inferred mineral resources, it cannot be assumed that all or any part of an inferred mineral resource will be upgraded to an indicated or measured mineral resource as a result of continued exploration.

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3) Mineral resources are inclusive of mineral reserves.
4) The Qualified Person for the mineral resource estimate is Mr. Tom Garagan, P.Geo.

1.10 MINERAL RESERVE ESTIMATE

The mineral reserve for the Otjikoto Project was generated as part of the DFS completed on the

Otjikoto Project in December 2012 that is summarized herein in the Technical Report “Otjikoto Gold Project NI 43-101 Technical Report Feasibility Report” dated February 25, 2013. The mineral reserve is based on a block model resource and mine plan that envisions open pit mining using conventional hard rock mining techniques. The Qualified Person for the mineral reserve estimate is Mr. Manie Kriel, Pr. Eng., B. Eng., MBL.

At the time of this report there is no known permitting, metallurgic, environmental or socio-economic conditions that would have a material effect on the mineral reserve estimate for the Otjikoto Project.

The probable mineral reserves for the Otjikoto Project are provided in Table 1.3 below. The mineral reserves have been shown to be economic and are reasonable for the statement of probable reserves. The final probable mineral reserves for the Otjikoto Project are 29.4 million tonnes of ore at a diluted grade of 1.42 g/t resulting in 1.3 million ounces (39.3 million grams) of contained gold at a stripping ratio of 5.59:1.

Table 1.3:        Probable Mineral Reserve Estimate

Parameter Unit Value
Total Ore
Total Waste
Stripping Ratio
Grade
Metal Content
tonnes
tonnes

g/t
ounces
29, 405, 338
164, 266, 053
5.59
1.42
1, 341, 151

1.11 MINING METHOD

The following methodology was followed during the design process:

  • Use the selected optimal pit shells derived from the pit optimization as the design limit;
  • Use the latest block model to show the ore distribution; and
  • Apply the pit design criteria and geotechnical parameters as discussed above.

The available pit room was used for haul roads wherever possible instead of expanding pit walls, and haul road width was reduced at the lower levels of the pit to minimise waste stripping as much as possible. The design work was performed in Gemcom’s SurpacTM and MicromineTM mine design software. Three (3) pushbacks were designed based on the selected interim pit shells and the designs were used to evaluate the tonnage and grades of the various material types which in turn were applied to the production scheduling.

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Ramp positioning within the overall pit design is an integral component of mine design as it influences the stripping ratio on the overall design, the performance of the equipment, and the operating costs due to direct impact of the ramps on the hauling profiles. The exit positions of the ramps were determined based on the proposed positions of the primary crusher and the waste dump.

The ultimate pit is shown in Figure 1.1 below and the internal phases are shown in Figure 1.2 below. The ultimate pit will have dual access established along the final limits as indicated in Figure 1.1 but not in all associated stages. The access on the north will be used for both ore and waste hauling while the access on the south will mainly be used for waste hauling.

Figure 1.1:        Otjikoto ultimate pit design

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Figure 1.2:        Otjikoto ultimate pit and internal phase designs

1.12 RECOVERY METHODS

The design of the Otjikoto Project mill is based on a gravity/whole ore leach flow sheet with a nominal treatment rate of 2.5 Mtpa.

Gold will be recovered by gravity concentration/intensive leaching and by a cyanide leach/CIP process for treatment of gravity tailings. The Otjikoto mill design is robust and will be able to process the 3 major ore types (XR1 – Oxide, XR2 – Pyrite dominant, XR3 – Pyrrhotite dominant) over the range of ore grades mined, and with variable materials handling characteristics. The process flow sheet is comprised of the following:

  • Crushing
  • Grinding
  • Gravity concentration and intensive leaching
  • Cyanide leaching of gravity tailings
  • Carbon-in-Pulp (“CIP”)
  • Cyanide destruction
  • Tailings disposal

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  • Acid wash and Elution
  • Electrowinning and Gold Room
  • Carbon regeneration
  • Reagents make-up and distribution
  • Air services and Plant water service

1.13 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL IMPACT

An Environmental and Social Impact Assessment (“EIA”) complete with an Environmental Management Plan (“EMP”) and Mine Closure Framework (“MCF”) was completed by SLR (Namibia) for the Otjikoto Project. The report details:

  • the EIA process followed (including legal requirements);
  • baseline studies completed;
  • process design considerations critical to community health & safety, environment, and social issues;
  • identifies the impact of the project on the environment, proposes monitoring programs for all phases of project development; and,
  • demarcates mitigation strategies.

The permitting process in Namibia is similar to many other jurisdictions. A full Mining License Application must be completed and approved. As part of the application, a full environmental impact assessment (complete with regulated public disclosure) must be completed. Once the EIA is granted environmental clearance, the MME can issue approval for a Mining License. After the Mining License is released, significant changes (to include the proposed change from power from the National grid to on-site power generation with heavy fuel oil (“HFO”) to the project have to be assessed by MME and MET to ensure that the changes comply with regulatory requirements. The Mining License contains a requirement to engage the MET in a contract to ensure compliance with the Environmental Management Program.

The MET approved the EIA and granted an Environmental Clearance Certificate (“ECC”) for the Otjikoto Project in August 2012. In conjunction with this the MME granted the Otjikoto Project Mining License on December 5, 2012.

Namibia has one of the healthiest economies in southern Africa. Its small population, significant natural resources, sound governance, and human capital have allowed peace and prosperity to prevail in the 23 years since independence. Low rainfall and poor soils mean that agricultural potential is low, restricting most agricultural development to the extensive use of rangelands for livestock production. The use of minerals has been the backbone of the natural resource-based economy and economic rents from minerals have to a certain extent been reinvested with the aim of ensuring sustainable growth and development. The tourism sector has also performed well with a significant nature-based tourism component showing good growth. The Namibian economy is closely linked to the much larger South African economy, which has also been growing well in the last decade. Namibian per capita income is among the highest in southern Africa and Africa as a whole, which belies marked inequalities in wealth. Significant growth in mineral income is expected in Namibia in the next few years as a result of major investments in new mining projects.

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1.14 CAPITAL AND OPERATING COST

The capital cost estimate consists of these components: direct costs, indirect costs and contingency, as described below. Owner’s costs were estimated separately by B2Gold.

The capital cost estimate was developed for the DFS and this report by:

•        DRA Mineral Projects (Pty) Ltd:                       Process Plant, Infrastructure
•        VBKOM Mining Engineers Namibia:               Mine equipment and mine development
•        Epoch Resources (Pty) Ltd:                               Tailings storage facility
•        B2Gold:                                                                  Owner’s costs

The process plant estimate is based on a 2.5 Mtpa capacity that was designed to feasibility study level by DRA and the Otjikoto Project’s management team.

The capital cost estimate for the Otjikoto Project is approximately US$244.2 million as of December 2012, subject to qualifications, assumptions and exclusions. The pre-production capital cost summary and distribution are shown in Table 1.4.

Table 1.4:        Pre-Production Capital Costs Summary

Item US $ Millions
Direct Costs  
       Earthworks 14.2
       Tailings Storage Facility 25.3
       Process Plant 106.3
       Electrical & Power Generation (leased) 8.4
       Mining Equipment (leased) 18.0
       Construction Equipment (leased) 2.7
       Mine Infrastructure, Mine Buildings 7.1
       Site Buildings, Ancillary Facilities 7.6
Total Direct Costs 189.6
Indirect Costs  
       Owner’s Costs 19.8
       Mechanical & Electrical Spares 3.3
       Owner’s Construction Management 2.4
       EPCM 13.6
Total Indirect Costs 39.1
Total Direct and Indirect Costs 228.7
Contingency 15.5
Total Pre-Production Capital 244.2

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Life of mine site unit operating cash costs, net of capitalized pre-stripping and other pre-development costs, are US$ per tonne milled, as summarized in the Table 1.5, these costs were developed in the fourth quarter of 2012.

Table 1.5:        Life of Mine Cost of Production Unit Operating Cost Summary


Life of Mine Unit Operating Costs
Total Cost
(USD$)
Surface Mining - Otjikoto 15.21
Processing 14.02
Site General 2.47
Total (USD$) USD 31.69

Life of mine average operating cash cost per gold ounce produced is US$689 per ounce. A summary of the cost breakdown is listed in Table 1.6.

Table 1.6:        Life of Mine Average Operating Cost per Gold Ounce Produced


Life of Mine Average Operating Cost per Gold Ounce Produced
Full Years1-5
USD$ / Gold Ounce Produced
Life Of Mine
USD$ / Gold Ounce Produced
Surface Mining - Otjikoto 313 375
Surface Mining - Otjikoto Prestripping (2) (73)
Processing 246 323
Site General 45 64
Change in Stockpiled Ore (80) (0)
Silver Sales (3) (3)
Refinery Charge 3 3
Dore Transportation, Security, Insurance 1 1
Total (USD$) USD 524 USD 689

1.15 ECONOMIC ANALYSIS

Economic analysis of the Otjikoto Project was performed to assess the economic viability of constructing and operating the project as designed. The analysis was based on mine plans and production schedules derived from the most current resource estimates. The average annual life of mine metal production averages approximately 112,000 ounces of gold over the 12 years of full production from the process plant.

The pre-tax cash flow includes: net production value at a gold price of US$1,550 per ounce, surface mining cost inclusive of pre-stripping, processing cost, site general cost, other income (expenses), reclamation and closure, change in supplies inventory pre-production and sustaining capital expenditures. The pre-tax Net Present Value (“NPV”), at 5% discount rate, and Internal Rate of Return (“IRR”) are calculated to be US$402 million and 30.4% respectively. The initial capital costs are US$244.2 million with a simple payback of less than three years.

The after-tax cash flow includes: net production value at a gold price of US$1,550 per ounce, surface mining cost inclusive of pre-stripping, processing cost, site general cost, royalties, taxes, other income (expenses), reclamation and closure, change in supplies inventory pre-production and sustaining capital expenditures. The after-tax NPV at 5% discount rate and IRR are calculated to be US$243 million and 23.6% respectively. The initial capital costs are US$244.2 million with a simple payback of less than three years.

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1.16 INTERPRETATION AND CONCLUSIONS

Based on the robust project economics, the B2Gold Board of Directors has accepted the DFS and has instructed management to implement the study’s recommendations to develop and bring the Otjikoto Project to commercial production.

The Otjikoto Project has excellent exploration potential. An aggressive exploration drilling program continues on the success of the high grade Wolfshag zone discovered in late 2011, that is adjacent to the planned Otjikoto pit. These positive results indicate significant exploration upside and the potential to outline additional resources which could lead to the expansion of through-put capacity and increased annual average gold production.

1.17 RECOMMENDATIONS

The Otjikoto Project should be advanced to commercial production as soon as possible and this process has commenced. Sufficient testing and trade-off studies have been completed to confidently select the recovery process, establish the design criteria, and estimate reliable capital and operating costs. Required government permits and licenses for construction have been received. Construction activities have begun, and B2Gold estimates that the Otjikoto Project will produce gold prior to the end of 2014.

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2 INTRODUCTION

DRA was appointed by B2Gold and its subsidiary B2Gold Namibia (Pty) Limited (previously Auryx Gold (Namibia) (Pty) Limited) to complete a DFS for the Otjikoto Project, located near Otjiwarongo in the Republic of Namibia (“Namibia”).

2.1 TERMS OF REFERENCE

This report is prepared in accordance with guidelines for disclosure pursuant to NI 43-101 and Form 43-101F1 in support of the results of the DFS, updated mineral resource and initial publication of a mineral reserve estimate for the Otjikoto Project as disclosed in the B2Gold news release dated January 10, 2013.

The scope of the study was to prepare a bankable feasibility study for the Otjikoto Project to assess the economic viability of the project and related decision to commence construction of the Otjikoto Gold Mine.

2.2 SOURCES OF INFORMATION

This Technical Report summarizes the detailed work presented in the DFS prepared by DRA (DRA, 2013).

The following companies have undertaken work in preparation of this Technical Report:

  • DRA Mineral Projects (“DRA”) provided overall management and preparation of the Feasibility
    Study with emphasis on the process plant and infrastructure design;
  • VBKom Consulting Engineers Namibia (Pty) Ltd (“VBKOM”) completed the mine planning and mineral reserve estimation;
  • Epoch Resources (Pty) Ltd (“Epoch”) were responsible for the tailing impoundment design;
  • SLR (Namibia) (“SLR”) completed the EIA, EMP and the Mine Closure Framework (“MCF”).
    Geochemical analysis was also undertaken;
  • AGES (S. Africa) completed a hydrological study of the area; and
  • SRK (S. Africa) completed a geotechnical assessment of the plant and mine areas.

Contributors associated with these and other companies which have contributed to the Technical Report are summarized in Table 2.1.

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Table 2.1:        Summary of Contributor Details

Discipline Responsible Party Subject Matter Expert
Geology and mineral resources B2Gold Corp. Tom Garagan, QP, P. Geo, Senior Vice President Exploration, Qualified Person, Mineral Resources
Hugh MacKinnon, P. Geo Exploration Manager
Tyler Mckinnon, Senior Resource Geologist
Anton Lombard, Pr.Sci.Nat, Senior Geologist
Mining Geotechnics SRK Consulting Alan Naismith, QP, Pr. Sci. Nat., Partner and Principal Rock Engineer
Robert Armstrong Pr.Sci.Nat
Ohveshlan Pillay Can.Sci.Nat
Edrie du Toit Can. Eng.
Mineral Reserves
VBKom Namibia Consulting
Engineers (Pty) Ltd.
Hermanus Kriel, QP, Pr Eng, Bsc Eng, MBL, MD, Princepal Mining Engineer.
Werner K Moeller, Director, Senior Mining Engineer – Lead Mining Engineer – Pit Optimisation, Pit Design, Production Schedule, OPEX & CAPEX Estimate
Bertha M Iitana, Mining Engineer,– Mining Engineer, Pit Design
Barend D Human, Mining Engineer, Mining Engineer, Production Schedule, OPEX & CAPEX Estimate
Strauss Oosthuizen, Industrial Engineer, Project & Risk Manager
Process and Metallurgy DRA Minerals Processing Glenn Bezuidenhout,QP, Process Director, Diploma in Extractive Technology, FSAIMM
Mark Townsend, Project Manager, B. Sc. Mechanical Engineering
Val Coetzee, P. Eng, Sr. Process Engineer, B. Chem.

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Discipline Responsible Party Subject Matter Expert
      Eng.
Tailings Facility Design Epoch Resources (Pty) Ltd. Guy Wiid, QP, Pr. Eng, Bsc Eng (Civil), Msc Eng (Civil)
George Papageorgiou,Pr Eng, PhD, Msc, Bsc Eng (Civil)
Drian Roos, Bsc Eng (Civil)
Geotechnical Site Investigation of the Tailings Facility Site SRK (Pty) Ltd Alan Naismith, QP, Pr. Sci. Nat., Partner and Principal Rock Engineer
Robert Armstrong, Pr. Sci. Nat., Bsc Hon.
Hydrological Study Of the Tailings Facility Site Peens and Associates Hendrik Peens
Environmental Impact Study of the Tailings Facility Site SLR Namibia Alexandra Speiser, Arnol Bitner and Auriol Ashby
Infrastructure Holley Associates
B2Gold Corp.
Graham Smith, QP, Pr. Eng, Bsc Eng (Civil), Managing Director
Pardon Mukamba
Capital and Operation Costs B2Gold Corp.
DRA Minerals Processing
Doug Wollant,B. Sc. Mining Engineer
Glenn Bezuidenhout, Process Director, Diploma in Extractive Technology
Economic Analysis B2Gold Corp. Bill Lytle, QP, Pr. Eng, M. Sc. Civil Engineering, B. Sc. Chemical Engineering
Doug Wollant, B. Sc. Mining Engineering
EIA Project management SLR Consulting Namibia (Pty) Ltd Werner Petrick, QP, Pr Eng, B.Eng (Civil), M.Env Mgt)
Geohydrology


AGES Gauteng


Stephan Meyer, B.Sc. Hons Geohydrology
Koos Vivier, Ph.D Environmental Management, M.Sc Hydrogeology, Pr.Sci.Nat
Megan Hill, B.Sc. Hon. Geology
 
Geochemical Assessment AGES Gauteng Robert Hansen, M.Sc: Geology Pr. Sci. Nat
Socio-economist and stakeholder consultations Ashby Associates cc A.M. Ashby - BSc (Hons)
Socio-economist and stakeholder Ashby Associates cc A.M. Ashby - BSc (Hons)

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Discipline Responsible Party Subject Matter Expert
consultations      
Noise impact assessment M2ENCO Morne de Jager - B. Ing (Chem)
Visual impact assessment Visual Resource Management Africa cc Stephen Stead - BA (Psychology and Geography, BA (Hons) Human Geography and Geographic Information Management Systems
Visual impact assessment Visual Resource Management Africa cc Stephen Stead - BA (Psychology and Geography, BA (Hons) Human Geography and Geographic Information Management Systems
Economic impact assessment Independent consultant Jon Barnes PhD - BCom, BSc (Hons), BSc
Biodiversity: Project leader/ecologist Gobabeb Desert Research Foundation Joh Henschel
Biodiversity: Vegetation assessment Independent consultant Coleen Mannheimer - M.Sc. Botany, Plant Taxonomy, BSc (cum laude)
Biodiversity: Fauna assessment Independent consultants Michael Griffin - BSc (Hons), Pr.Sci.NatCompany
Biodiversity: Avifauna assessment Independent consultant Dr Chris Brown
Groundwater assessment AGES WJ Meyer - B.Sc (Hons) Geohydrology

2.3 PERSONAL INSPECTIONS

2.3.1 GEOLOGY AND RESOURCES

The following Qualified Person conducted a personal inspection of the Otjikoto Project property: Tom Garagan, P.Geo, Senior Vice President, B2Gold Corp. Tom Garagan visited the Otjikoto Project site and core logging and exploration office facilities in Otjiwarongo on the following dates:

  • January 7th to 12th, 2012
  • February 3rd to 5th, 2012
  • August 21st to 24th, 2012
  • November 20th to 22nd, 2012

During the visits to drilling operations, RC sampling and core handling were observed in addition to logging, core cutting and sampling procedures. Detailed discussions were held with the B2Gold Namibia staff on the geology and mineralization of the deposit.

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2.3.2 GEOTECHNICAL

The following Professional Engineer conducted a personal inspection of the Otjikoto Project property starting on Monday, 16, April, 2012, for the duration of 5 days:

  • Robert Armstrong, Pr.Sci.Nat., Senior Geologist, SRK Consulting

The site visit was conducted in order to review the mining geotechnical drilling and data collection that was being undertaken on site. The drilling equipment was inspected and geotechnical logging procedures were reviewed.

In addition to the above site visit, Ohveshlan Pillay BSc. Hons, was present on site on April 11 through to May 14, 2012. Geotechnical drilling was supervised and geotechnical and structural logging was conducted as the geotechnical holes were drilled.

2.3.3 MINING AND RESERVES

In accordance with the NI 43-101 guidelines, the following Qualified Persons conducted a personal inspection of the Otjikoto Project property conducted on November 14, 2012, for the duration of one day:

  • Hermanus J. Kriel, CEO, Senior Mining Engineer, Pr. Eng – Qualified Person of VBKom Namibia Consulting Engineers (Pty) Ltd
  • Werner K Moeller, Director, Senior Mining Engineer – Lead Mining Engineer of VBKom Namibia Consulting Engineers (Pty) Ltd

The purpose of the visit was to inspect the core, interact with management and assess the conditions on the Otjikoto Project site. The site visit commenced in the morning at the exploration offices and core shed of the Otjikoto Project in Otjiwarongo, Namibia where the group was introduced to Louis Brown (Exploration Operations Manager) and Hugh MacKinnon (Exploration Manager). The group was allowed to peruse the core shack and discuss the geology of the project site with the staff.

The tour then commenced just before lunch time to the actual project location at Exclusive Prospecting License (“EPL”) 2410. The group was shown the surface locations of the deposit area and where the main components of the waste rock dump, tailings storage facility, process plant and mine offices, primary access road and power line would be located. During the site visit a contractor was busy de-bushing and clearing the proposed process plant footprint. The tour ended in late afternoon.

2.3.4 PROCESS AND METALLURGY

The following Qualified Person conducted a personal inspection of the Otjikoto Project on March 22, 2012 for the duration of 1 day:

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  • Glenn Bezuidenhout, Process Director, DRA Minerals Processing

The purpose of the site visit was to conduct a visual inspection of the Otjikoto process plant area and to confirm the conditions (as they were visually present).

2.3.5 TAILINGS FACILITY SITE

The following Qualified Person conducted a personal inspection of the Otjikoto Project property on September 3, 2012 for the duration of 1 day:

  • Drian Roos, BEng (Civil), Geotechnical Engineer, Epoch Resources (Pty) Ltd.

The purpose of the site visit was to conduct a visual inspection of the Otjikoto tailings storage facility area and to confirm that there are no visible impediments for the construction of the tailings storage facility (“TSF”). Photographs were taken and kept as visual evidence of the visit.

2.3.6 INFRASTRUCTURE

The following Qualified Persons conducted a personal inspection of the Otjikoto Project on Thursday, March 22, 2012, for the duration of 1 day;

  • Bill Lytle, P. Eng, M. Sc. Civil Engineering, B. Sc. Chemical Engineering

The following Qualified Persons conducted a personal inspection of the Otjikoto Property on Thursday, 22nd March 2012, for the duration of 1 day;

  • Graham Smith, Holley & Associates, Director – Infrastructure
  • Danny Russell, Senior Study Manager, DRA
  • Glen Bezuidenhout, Director – Process Engineering, DRA
  • Ron Liddiard, Electrical Consultant, Dra

The inspection comprised of a visit to the Geological Core Storage and Offices in Otjiwarongo where the drill cores were inspected and the geological conditions of the mine site were demonstrated. Thereafter, an inspection of the entire site was conducted. This included the following:

  • The B1 National highway along the mine property
  • The rail line parallel to the highway and the mine property and the rail siding
  • The fenced areas including the existing house and store in the southern portion of the property
  • The original airfield area
  • The proposed mine pit area
  • The proposed plant site areas

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  • The proposed overburden dump area
  • Some of the relevant borehole positions
  • The proposed site access road location
  • The existing D2808 District road
  • The existing house in the northern portion which is proposed to be used for the site accommodation and construction camp area
  • The proposed location of the proposed re-routed D2808 District road
  • General view of the site including vegetation

2.3.7 FINANCIAL EVALUATION

The following Qualified Person conducted a personal inspection of the Otjikoto Project property on numberous occasions in 2012 for multiple days:

  • Bill Lytle, P. Eng, M. Sc. Civil Engineering, B. Sc. Chemical Engineering

2.3.8 ENVIRONMENTAL

The following persons conducted a site visit in May 2012 as part of the EIA process.

  • Werner Petrick, B.Eng (Civil), M.Env Mgt, SLR Environmental Consulting Namibia (PTY) Ltd.
  • Chris Herbert, BA (Planning Studies), Diploma (Town Planning), MRTPI (Member of the Royal Town Planning Institute, SLR Environmental Consulting.

A site visit was conducted to get familiarized with the site, the existing environment and to understand where the proposed facilities, mining etc. will take place on the ground.

Werner Petrick also visited the Otjikoto Project with A.M. Ashby, (BSc Hons), of Ashby Associates cc, on June 21, 2012 for the duration of 2 hours:

They visited the proposed mine site and proceeded to other parts of the Otjikoto Project to view existing farm dwellings which are planned to be converted into an environmental education centre.

In addition to the site visits mentioned above, the following people (part of the EIA Team) conducted site visits as follows:

  • Alex Speiser, M. Sc (Geology/Palaeontology), MPhil (Environmental Science), SLR Consulting Namibia (PTY) Ltd. Site Visit: June 2007
  • A.M. Ashby, (B. Sc Hons), Ashby Associates cc, Site Visit: June 2012
  • Hanlie Liebenberg-Enslin, M. Sc (University of Johannesburg), Airshed Planning Professionals. Site Visit: November 2010
  • Morne de Jager, B. Ing (Chemical), M2ENCO. Site visit: May 2012

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  • Stephen Stead, BA (Psychology and Geography, BA (Hons) Human Geography and Geographic Information Management Systems, Visual Resource Management Africa cc. Site Visit: Oct 2010
  • John Kinahan, Ph.D MSAIE&ES, Quaternary Research Services. Site Visit: May 2012
  • Coleen Mannheimer, M. Sc. Botany, Plant Taxonomy, BSc (cum laude), Independent consultants. Site Visit: April 2007 and February 2008
  • Michael Griffin, B. Sc (Hons), Pr.Sci.NatCompany, Independent consultants. Site Visit: 3 April to 10 April, 2007.
  • Peter G Hawkes, B. Sc. (Hons) Pr.Sci.Nat, Independent consultants. Site Visit: June 2012
  • Dr Chris Brown, Independent consultants. Site Visit: August 2007 & February 2008

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3 RELIANCE ON OTHER EXPERTS

Other individuals have provided input to this report who technically would not be considered Qualified Persons under NI 43-101 guidelines, but who have the necessary qualifications and experience to provide input and opinions incorporated into the report.

A detailed summary of these contributors is depicted in Section 2 “Introduction” of this report in Table 2.1.

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4 PROPERTY DESCRIPTION AND LOCATION

4.1 LOCATION

The Otjikoto Project is located approximately 70 kms from the town of Otjiwarongo and 50 kms from the town of Otavi within the Province of Otjozondjupa in the north-central part of the Republic of Namibia (Figure 4.1) . The project site is 300 kms north of Windhoek, the country’s capital, and centred on the farms Otjikoto and Felsenquelle at approximately 718,900 East and 7,787,100 North in WGS84 Zone 33 South coordinate system (20°00’ South (Longitude) and 17°05’ East (Latitude)).

4.2 TENURE

In Namibia, all mineral rights to the property are vested in the State. The Minerals (Prospecting and Mining) Act of 1992 (the “Minerals Act”) regulates the mining industry in the country.

On December 5, 2012 the Namibian MME granted Auryx Gold Namibia (Pty) Ltd (“Auryx Gold Namibia”) the Otjikoto mining license, ML 169. Auryx Gold Namibia was legally renamed B2Gold Namibia Pty Ltd. (“B2Gold Namibia”) on July 19, 2012, but the change of name registration is still pending with the MME. B2Gold Namibia is owned indirectly 92% by B2Gold and 8% by EVI, a Namibian black empowerment company. The mining license (“ML”) was granted in accordance with the Minerals Act and covers an area of 6,933.99 hectares (Figure 4.2) . The license is valid for a term of 20 years with expiry of December 4, 2032. The license can be renewed for a further 20 years upon application to the MME. The mining license requires an annual fee, development of a works program, environmental compliance, commitment to seek local suppliers for fuel and lubricants, approval of the product take-off agreement, and payment of taxes by permanent employees in Namibia.

The mining license ML 169 is situated within EPL 2410 (Figure 4.2) . EPL 2410 covers an area of 54,125 ha (inclusive of ML 169) and is in good standing, with renewal for an additional two years granted by the MME on September 14, 2012. An annual fee of N$6,000 and filing of quarterly exploration reports with the MME and bi-annual environmental reports with the MET are required to keep the license in good standing. Exploration is conducted under the terms of an ECC issued by the MET on June 20, 2002. A renewal ECC application was submitted to the MET on January 31, 2013, in accordance with requirements of 2012 changes to the environmental laws of Namibia.

Three additional EPL’s are held by B2Gold Namibia in the area (Figure 4.2) . These are summarized in Table 4.1.

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4.3 SURFACE RIGHTS

In 2011, the farms Wolfshaag, Otjikoto, Gerhardshausen and Okaputa Nord I were purchased and consolidated by Auryx Properties Holdings (Pty) Ltd. (“Auryx Properties Holdings”) (Table 4.2) . On July 19, 2012, Auryx Property Holdings was legally renamed to B2Gold Namibia Properties (Pty) Ltd. (“B2Gold Namibia Properties”). The Mining License and all proposed infrastructure are situated on the B2Gold Namibia Properties farms.

Figure 4.1:        General Location Map of Otjikoto Project (modified after Namibian Ministry of Fisheries and Marine Resources map)

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Table 4.1:        Otjikoto license status

TYPE NUMBER NAME COMPANY ORIGINAL
AREA ha
EXPIRY
DATE
ANNUAL
FEE N$
EPL 2410 AREA 9 (OTJIKOTO) Auryx Gold Namibia (Pty) Ltd 54125 14/09/2014 6000
EPL 3177 AREA14 Auryx Gold Namibia (Pty) Ltd 26146 23/06/2013 3000
EPL 4309 AREA 16 (RABBIT EARS) Auryx Gold Namibia (Pty) Ltd 97881 14/06/2014 10000
EPL 3765 HOMESTEAD Auryx Minerals Exploration (Pty)Ltd 97113 25/01/2013* 10000
ML 169 OTJIKOTO MINING LICENCE B2Gold Namibia (Pty) Ltd 6933.988 04/12/2032 5000

* Renewal pending MME approval
 
Figure 4.2:        General Location map of Otjikoto Project showing location of mineral licenses

Table 4.2:        Farm Registration

Farm Name & No Deed of Transfer
(Registration Division “B” Otjozondjupa Region)
Area
(ha)
Owned by
Comments
Otjikoto 6,789.6341 Auryx Property Holdings Consolidation of and

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No 2035
T 985/2011
   (Pty) Ltd Otjikoto No.573 by Certificate of Consolidation Title No T 3556/2009.
farms Wolfshaag No.574
Felsenquelle
No 2036
T 984/2011
8,008.1867 Auryx Property Holdings (Pty) Ltd Consolidation of farms Gerhardshausen No.572,
Okaputa Nord I No.571 and the remainder of Okaputa II No.570 by Certificate of Consolidation
Title No T 4268/2009.

4.4 ROYALTIES AND OTHER

The Otjikoto Project is not subject to any back-in rights payments, agreements or encumbrances. The corporate tax rate in Namibia for mining properties is a flat rate of 37.5% . Corporate tax of 34% applies to taxable income from non-mining activities. Mine production is subject to royalties at 3% of net market value payable to the Namibian State. Additionally, production is subject to the provisions of Section 114 of the Minerals Act. Detailed quarterly and annual reports on all relevant aspects of operations must be submitted.

Allowable tax deductions for mining companies are as follows:

  • All pre-production exploration expenditures are fully deductible in the first year of production; to the extent that this deduction exceeds income from mining operations for the year concerned, it will create an assessed loss for carry forward or set off against other income of the taxpayer;
  • Subsequent exploration expenditures are not ring fenced and are fully deductible in the year they occur, so that profits from existing operations can be used to fund exploration in any part of the country; and
  • Initial and subsequent development costs (include all operating and capital expenditure incurred in connection with the development operations) are fully deductible in equal installments over three years commencing in the year the mine starts production.

The Minerals Act also makes provision for a penalty royalty (for the failure of beneficiating minerals in Namibia, where such beneficiation is possible, transfer pricing arrangements and excessive brokerage fees) as well as for a windfall royalty.

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4.5 ENVIRONMENTAL LIABILITIES

A small, shallow quarry is present in the northern eastern part of ML 169. This calcrete aggregate quarry dates from the construction of the D2808 regional road that runs through the property. There are no environmental liabilities associated with this quarry and no other known environmental liabilities on the property.

All work conducted on the licenses must be in compliance with the MME and MET regulations as outlined on the ML, EPL and ECC documents.

4.6 PERMITS

The Otjikoto Project has all the permits required to allow for the start of construction. Additional details are covered in Section 20 “Environmental Studies, Permitting and Social or Community Impacts” of this report.

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4.7 SIGNIFICANT FACTORS AND RISK

B2Gold is not aware of any significant factors or risks that may affect access, title or the right or ability to perform work on the property.

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Figure 4.3:        Legal Survey diagram of Otjikoto Mining License ML 169

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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 ACCESSIBILITY

5.1.1 MEANS OF ACCESS TO THE PROPERTY

The mine will have access off the main B1 road, a primary paved road. Mine related traffic will travel from both the Otavi and Otjiwarongo directions on a daily basis with all traffic using the B1 to access the property.

Otjiwarongo lies some 70 km southwest from the Otjikoto Project area and Otavi is approximately 50 km to the northeast and both are situated on the B1. Internal access to the project area is via a well-maintained network of secondary roads and farm tracks. Given the arid climate of the area, these roads are generally passable all year-round.

The fence around the construction site will be the final fencing required for the plant. It will be 2.4 metres high, with barbed wire at the top. Site access will be controlled via a boomed entrance, which will also be locked outside of normal working hours.

5.1.2 SURFACE RIGHTS FOR OPERATIONS

The surface rights of the farms on which the mining will take place are owned by B2Gold Corp. through its subsidiary B2Gold Namibia Properties.

B2Gold Properties was previously Auryx Properties, but following the acquisition of the project by B2Gold the name of the company was changed. The existing EPLs are held in the name of Auryx Gold Namibia, and a change of name registration to B2Gold Namibia is still pending with the MME.

The existing surface rights cover four farms which have been consolidated into two farms. There is more than sufficient surface area for the mine, waste dumps, plant, tailings pond, associated infrastructure and any other requirements for construction and operations.

5.2 CLIMATE

The Otjikoto Project is located approximately 300 km inland from the west coast of Namibia. The area is characterized by low rainfall with extreme temperature ranges and unique climatic factors influencing the natural environment and biodiversity. In general, the climatic conditions at the Otjikoto Project site allow for year-round construction and mine operations.

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Detailed information can be obtained from Appendix 10-1 “Environmental Impact Assessment” in the DFS. A summary of key data is shown in Table 5.1.

Table 5.1:        Climate Data

SITE DATA    
Approx. Elevation / Altitude mamsl 1500
Temperature    
                   Min Annual Temp °C -5
                   Max Annual Temp °C 37.9
Wind    
                   Design Peak Velocity m/s 6
                   Wind Direction N/A ESE (East South East)
Rainfall    
                   Average Annual mm 350-450
Evaporation    
                   Average Annual Rate mm 2400-2600

The site experiences rainfall predominately in the summer (December to March) in the form of thunderstorms and showers. Rainfall peaks in January and February. The Otjiwarongo area falls within the higher rainfall band of Namibia where average annual rainfall varies between 350 mm – 450 mm per year. Rainfall measurements on site have recorded a maximum annual rainfall of 652 mm in 2011. The previous two years were noted as drier. Frost can occur on site but is only likely to be experienced for 1-5 days per year. The site falls in an area of Namibia which experiences the highest solar radiation. Solar radiation averages between 6.2 - 6.4 kWh per m² per day. The prevailing wind direction is from the ESE, with peak wind speeds increasing from the middle of winter through into spring (July - October). Wind speeds are primarily below 5 m/s with peaks greater than 6 m/s recorded for only 0.1% of time.

5.2.1 WIND SPEED AND DIRECTION

Wind speed and direction as recorded, by the on-site weather station, were verified by purchasing weather data from the United States MM5 database for the years 2008 and 2009. The MM5 data were extracted for the same location as where the Otjikoto weather station is located.

The MM5 and on-site weather data show similar prevailing wind directions to be from the E and ESE with occasional airflow from the NE and SE. MM5 wind speeds are much higher in general than the on-site data, ranging primarily between above 4 m/s whereas the on-site data have wind speeds primarily below 5 m/s.

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Measured data are preferred above simulated meteorological data and since the prevailing wind directions are mainly similar between the two datasets, the on-site weather data can be regarded as representative of the on-site conditions. From Figure 5.1 it is evident that the wind speeds are primarily below 5 m/s with strong winds of more than 6 m/s occurring for less than 0.1% of the time. Moderate wind speeds between 2 m/s and 4 m/s occur for 26.8% with those between 1 m/s and 2 m/s occurring for 24.5% .

Figure 5.1:        Average Wind Frequency and Velocity Over the Period 8 June 2007 to November 2011.

5.2.2 TEMPERATURE

The minimum, maximum and mean temperatures recorded on-site are shown in Table 5.2. Monthly average diurnal temperatures for the Otjikoto Project site are provided in Figure 5.2.

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Table 5.2:        Minimum, Maximum and Average Temperatures for the Otjikoto Gold Project for the Period 8 June 2007 to November 2011.

  Jan Feb Mar Apr May Jun Jul Aug Sept Oct Nov Dec
Maximum 36.7 34.1 33.8 29.9 29.8 28.6 27.7 31.5 34.8 37.4 37.9 37.4
Minimum 9.8 9.6 7.6 7.8 0.6 -5 -1.6 -2.1 0.1 1.4 5.6 7.8
Average 22.0 21.3 20.6 19.4 16.9 13.9 14.1 17.9 22.4 23.4 23.2 23.4

Figure 5.2:        Monthly Average Temperatures for the Otjikoto Project for the Period 8 June 2007 To November 2011.

5.2.3 RAINFALL

The total rainfall for the project area is shown in Table 5.3 and Figure 5.3.

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Table 5.3:        Total Monthly Rainfall for the Otjikoto Project for the Period 8 June 2007 to November 2011

  Jan Feb Mar Apr May Jun Jul Aug Sept Oct Nov Dec Total
2007 N.D N.D N.D N.D N.D 0 0 0 0 12.8 14.8 8.2 35.8
2008 85.40 N.D N.D N.D N.D N.D N.D N.D N.D N.D 8 1 95.0
2009 9.8 8.2 51.4 11.8 18.4 0.4 0 0 10.6 61.4 57.6 88.6 318.2
2010 168.4 82.2 120.4 0 0 0 0 0 0 0 94 91.8 556.8
2011 215.2 133.2 129.8 150.4 1.2 0 0 0 0.2 6.2 16.2 0 652.4

Figure 5.3:        Total Monthly Rainfall as recorded at the Otjikoto Project for the Period 8 June 2007 to November 2011

5.3 MINING PERSONNEL

Mining operations will also commence during the construction phase and the number of personnel involved with mining operations during year one will be approximately 190. This number will gradually increase to around 300 at the peak of production. Most of the permanent employees will be housed either in Otavi or Otjiwarongo.

Process and management staff will also be recruited during the construction phase and the numbers will increase from around five initially to around 135 at full production. These employees will also be housed at either Otavi or Otjiwarongo.

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During the construction period the majority of the workers will be housed in a temporary construction camp with some workers being housed in Otavi and Otjiwarongo. Both the workers housed on site in the camp and off site will be transported by bus on a daily basis to their assigned work areas.

Recruitment of skilled workers is already underway, with offices being set up in both Otavi and Otjiwarongo. Namibia has limited resources of skilled workers, but does have established mining and construction industries. Extensive training programs will be required to provide sufficient trained personnel for mine operations. This will be supported is some areas (such as the mine) with vendor support for maintenance.

5.4 LOCAL RESOURCES

5.4.1 WATER SUPPLY AND AVAILABILITY AND SOURCE OF WATER

Water for construction activities will be supplied from boreholes (pre-sunk and equipped prior to construction) and pumped to a temporary tank with a polyvinyl chloride (“PVC”) piping distribution network throughout the construction site. Borehole water will be supplied for showers and ablution facilities. Bottled water will be supplied for drinking during the period when the potable supply facility is still not commissioned and in operation.

Water for production activities will be supplied from the boreholes and the design of the water supply system is based on the geohydrological AGES technical report, G12/056_21-05-12 (Appendix 10-5 of the DFS, and additional information sourced from AGES).

The HDPE water delivery lines will be laid on surface with no concrete plinths. The pipe sizes will range from diameters of 110 mm to 250 mm depending on the design flows.

5.4.2 AVAILABILITY AND SOURCE OF ELECTRICAL POWER

5.4.2.1 Construction power

Power for the purposes of construction will be taken from a regional power distributor (Cenored) 11kV 500kVA supply point, located on the south western boundary of the plant site. This power will be provided to the construction area (to include the laydown area) and groundwater wells that will be used for dust suppression during construction. The various contractors will be responsible for the supply and installation of their own 400 V outdoor distribution boards and the cabling to the nearest minisub.

The construction camp will also be supplied by Cenored via a 1000 kVA 33 kV/420 V transformer. A 400 V distribution boards housed in a dedicated DB room will supply the various loads in the camp.

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Backup generators will be purchased to assure that there is sufficient power in all areas during the construction phase of the project.

5.4.2.2 Power Supply for Operational Activities

Following a trade-off study, it was concluded that due to escalating costs and supply constraints on the side of Nampower (the national energy supplier) the plant would be self-generating.

The generators will supply power to the plant at 6.6 kV. This voltage level was preferred to the more widely used 11 kV in order to permit a bypass facility around the SAG mill variable speed drive in the event of failure of the latter, and to dispense with the otherwise needed 11/6.6 kV transformers to supply the mills.

Based on the motor list and the estimated power consumption of the mills, the power plant will be constructed with 15 MW (minimum) of installed generating capacity at site conditions, plus the capacity to start both mills.

The generators will be fuelled with HFO, which is available in Namibia. A storage facility with minimum ten days fuel supply has been allowed to support the generators. The necessary fuel storage capacity is subject to revision once the logistics of fuel delivery have been established.

5.5 INFRASTRUCTURE

5.5.1 TAILINGS STORAGE AREA

During the DFS, test work on the tailings stacking option and cyanide destruction options were conducted with the final results of some of this work are still pending. The results of this work will be used to finalize the tailings facility design prior to the start of construction. The dam will conform to international best practice.

The TSF will accommodate tailings generated in the leaching section of the processing plant, which is estimated at 200,000 tonnes per month for a design life of 12 years. This lined facility will be located south of the process plant site and has been designed utilizing South African “slimes dam” technology resulting in low water storage levels and a maximum return of process water to the plant. There is additional space around the designed facility for future expansions if needed. It will be required to establish a water inventory in the ponds associated with this facility prior to start-up to assure sufficient process water is available in the first months of operation. The process water well system is designed to supply sufficient make-up water during the dry season at full production levels, but is not capable of supplying 100% of the process water needs until reclaim water is available from the tailings facility.

Details of this facility can be viewed in Section 8 of the DFS and the associated appendixes.

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5.5.2 INDUSTRIAL WASTE DISPOSAL

Waste will be separated at source, stored in a manner that there can be no discharge of contamination to the environment and either recycled or reused where possible. The remainder of Hazardous waste will be transported off site to appropriate recycling or disposal facilities (Otavi or Otjiwarongo for general waste and Walvis Bay for hazardous waste). General waste will be stored on site in an approved facility. A single waste management contractor will service the entire site. The types of non-mineralised waste expected to be generated during the construction phase include:

  • General waste (domestic waste and other non-hazardous waste);
  • Industrial waste; and
  • Hazardous waste.

Waste is generated during all phases of the mine. Details of the Management Plan that deals with solid waste management can be found in Appendix 10-2 “Environmental Management Plan” of the DFS.

5.5.3 MINE WASTE ROCK DISPOSAL AREA

Waste dumps for the mine waste will be located on the west side of the open pit, between the pit and the B1 highway. The waste will be used as buffer for visual, noise and dust, but the height will be limited to reduce the visual impact from the B1 highway.

The waste dump configuration is such that the main dumping area will be surrounded by a berm initially to ensure that dust and noise are contained within the dumping area and therefore does not become a nuisance to the general public. The perimeters berm construction methodology of the waste dump also allows for early visual impact minimization because these slopes can be vegetated as soon as possible. This implies the berm will first have to be created before the dumping continues to the rest of the dumping area. The outer edge of the berm will be contoured to give it a more attractive, natural look, especially after re-vegetation. Table 5.4 below shows the phased and total capacity and Figure 5.4 below shows the waste dump configuration and progression.

There is sufficient area available for the waste dumps as designed and for potential future mine expansions to the north.

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Table 5.4:        Rock dump capacities

Dumping Phases In-situ volume Tonnes
Phase 1
Phase 2A
Phase 2B
Phase 3
15 995 233
13 239 658
10 544 134
44 047 467
31 648 854
26 196 555
20 863 075
87 154 206
Overall Dump 83 826 492 165 862 691

Figure 5-4:        Otjikoto Mine Rock Storage Facility

5.5.4 PROCESSING PLANT SITE

The primary crusher will be located on the west side of the pit, and ore will be conveyed to the process plant site located south of the waste rock facility and south west of the pit. There is sufficient area near the primary crusher for a run-of-mine stockpile, and multiple stockpiles if needed. The primary crusher has been designed for direct dumping of trucks, so the available stockpile area is more than sufficient.

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The process plant area is relatively flat and has sufficient area for all facilities as well as any future expansions. The HFO generators, mine shops, administrative buildings and other maintenance facilities are in the same general area as the process plant. The footprints of these areas have been optimized for security, constructability and cost, but and sufficient area has been provided is available for the potential future expansion of all facilities.

5.5.5 CONSTRUCTION CAMP

The construction camp will be located northeast of the pit near an existing farmhouse. This site was chosen as it is of sufficient distance from the construction site and mines to provide quiet living conditions, has developed water wells, and is serviced by an existing Cenored power line. There is sufficient area for all required camp facilities.

5.5.6 PROXIMITY OF THE PROPERTY TO A POPULATION CENTRE

The Otjikoto Project area can be reached from the capital city, Windhoek, some 300 km to the south along the B1 national road, which is a surfaced road in excellent condition. Windhoek is connected by direct commercial air travel from several European countries, South Africa and other African countries. The large regional town of Tsumeb, a mining centre in its own right, is located some 110 km north of the project area, also along the B1 road. A domestic airport is located at Tsumeb, currently with both scheduled and charter flights connecting with Windhoek. The smaller town of Otavi lies some 50 km north from the Otjikoto Project area near the cross roads of the main north-south National Road B1, the road to Grootfontein and the road to Outjo. Otjiwarongo is the nearest town to the site and is approximately 70 km away.

5.5.6.1 Transport

The Ministry of Works and Transport is tasked to provide effective transport infrastructure and specialized services. This includes the National Roads Authority (“RA”) which manages the national road network with a view to support economic growth.

5.5.6.2 Construction Transport

During the construction of the project, there will be a limited increase in the number of people carrying vehicles travelling to and from site and there will be an increase in heavy vehicles (approximately 10 to 20 a day) supplying input materials. While most people will be based on site during construction, there will be a regular bus service (approximately 4 trips a day) from Otavi and Otjiwarongo to site.

5.5.6.3 Product Transport

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Doré bullion will be packed in plastic boxes with a maximum weight of 25 kg per box. The transporting of the doré bullion bars from the gold vault to the refinery will be outsourced to the selected refinery.

5.5.6.4 Material & Supply Transport

Materials and supplies required during both the operations and construction phases of the project will be delivered by truck utilizing the B1 highway which connects to all major supply areas in Namibia. Large equipment, construction materials, diesel fuel, HFO, process reagents and other bulk supplies will most likely come from the port at Walvis Bay by truck. There exists a rail line along the B1 that connects directly to Walvis Bay, but it is not currently in a state of repair for dependable service. It does however represent a potential future asset for material and supply transport.

5.6 PHYSIOGRAPHY

5.6.1 TOPOGRAPHY AND ELEVATION

The local topography is flat with a gentle slope towards the north-west with freely draining soils. The site is located at an elevation of 1,500 – 1,510 mamsl, just north of a local surface water divide. There are no well-defined surface water drainage features on the site and no major surface water flow or defined channel flow is expected other than local events after heavy rainfall.

5.6.2 VEGETATION

Two site visits were conducted to investigate the flora of the proposed Otjikoto Project area. The first one was carried out in April 2007 during the dry season focusing on the general vegetation of the area, while the second visit was conducted at the beginning of February 2008 after good rains had been received in the area investigating specifically geophytes, which often are only seen early in the rainy season. Vegetation in seasonally wet areas also varies with the degree of water logging. The greater part of the project area falls within the Tree and Shrub savannah zone, which is listed as the dominant vegetation type in central Namibia. The specific area affected by the proposed mine includes two habitat types. On the farms Otjikoto and Gerhardshausen there is a large ephemeral pan surrounded by woody species, and the rest of the area is composed of thorn bush thicket. In all, 107 plant species (including infraspecific taxa) were found on the Otjikoto study site. No red data species have been recorded for the area. However, eight endemic, seven near-endemic and thirteen protected species appear in this list.

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There are no plant species of sufficient conservation concern in any of the above habitats and due to the relatively low sensitivity of the vegetation present no special mitigation measures are necessary. Permits have been obtained and removal of vegetation is in progress in the proposed areas for the construction camp, designed roads, open pit, wasted dump area, tailing pond area, process plant site, mine shop area and other infrastructure.

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6 HISTORY

6.1 PRIOR OWNERSHIP

Historical exploration within the Otjikoto Project area has been reported to the MME by means of the Energy Grant and EPL license system. This data is on open file within the Geological Survey of Namibia national archive. Original reports are referenced from the Grant or EPL numbers.

Prior to Avdale’s involvement in the region, only a limited portion of the licenses were held and explored by other companies.

Avdale was incorporated as a private limited liability company in 1993 as a wholly-owned subsidiary of Anglovaal. In 1995, Anglovaal’s project generation team initiated compilation work of the area. In 1997, Exclusive Reconnaissance License (“ERL”) were granted to Avdale, a portion of which covered the current property area. In 1998, a series of EPL’s covering 1,061,438 hectares were staked by Avdale and the ERL’s relinquished.

The EPL’s were collectively called the Otavi Exploration Project and a 50:50 JV was formed with Rio Algom in 1998 to explore the EPLs. From 1998 to 2005 work was conducted on the property by Avdale. In 2003, Anglovaal, and thus Avdale, was acquired by ARM. With the acquisition of Rio Algom by BHP-Billiton in 2003 the JV on the Otavi Exploration project was terminated with Avdale retaining a 100% interest and BHP-Billiton having no back-in rights.

In 2005, Teal went public on the TSX and JSE with the Otavi Exploration Project, as well as properties in Zambia and the DRC, forming the basis of the new company. At the time of listing, ARM had a direct 64.9% interest in Teal with investors having the remaining 35.1% . In 2007, EVI purchased an 8% interest in Avdale; effectively an 8% interest in the Otjikoto Project.

In 2009, Vale and ARM formed a 50:50 joint venture by delisting Teal from the JSE and TSX, whilst at the same time Vale purchased a 15% interest in Teal from ARM, thereby creating a 50:50 joint venture for the purposes of developing Teal’s Zambian copper assets. The Namibian gold assets were deemed non-core at the time and the Vale/ARM joint venture decided to dispose of them. In early 2010, BC Ltd, a private company incorporated in British Columbia, Canada, acquired the Otjikoto Project from Teal. The Zambian and DRC assets of Teal were not included in the acquisition. In order to become a publicly traded company in Canada, BC Ltd agreed to merge with the publicly listed company Tova. On June 25, 2010 Tova completed the acquisition of Teal’s 92% interest in the Otjikoto Project. On June 21, 2010 Tova was renamed Auryx. From June 25, 2010 to December 22, 2011 Auryx owned and operated the Otjikoto Project. On December 22, 2011, B2Gold acquired 100% of the shares of Auryx and Auryx became a wholly-owned subsidiary of B2Gold. B2Gold now indirectly holds a 92% interest in the Otjikoto Project with EVI retaining its 8% interest.

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6.2 PREVIOUS EXPLORATION

6.2.1 PRE-AVDALE

Historic work in the region from the late 1960’s through mid-1990’s was focused on base metal exploration and there was little or no known exploration conducted for gold prior to the discovery of gold in 1999. Due to the extensive calcrete cover, very little exploration was conducted in the immediate area of the deposit. Historic exploration in the region is summarized in Table 6.1 and in McDonald et. al, 2011. Of principal note is the work conducted by Anglo American Corporation (“Anglo American”) as it was focused on the farm Eque in the northern part of EPL 2410.

Table 6.1:        Summary of previous exploration in the region before 1995

Company Grant/EPL Ref. Target Commodity Area (ha) Date of Final Report
Kennecott Exploration 196 Cu 37,732 Dec-68
Kennecott Exploration 184 Cu/Pb/Zn 48,482 Dec-68
Kennecott Exploration 154 Cu/Pb/Zn 10,351 Aug-69
Falconbridge of SWA 454 Cu/Pb/Zn 10,527 Sep-73
Falconbridge of SWA 423 Cu/Pb/Zn 20,626 May-73
Falconbridge of SWA 310 Cu/Pb/Zn 33,161 Feb-74
Tsumeb Corporation 720 Cu 1,325 Oct-75
Anglo American 1070 Base metals 25,486 Apr-85
Gold Fields Prospecting 1654 Cu 39,568 Mar-91
Tsumeb Corporation 1655 Base metals 57,122 unknown

Anglo American prospected Grant 1070, which overlaps historical Grant 184, during the early 1980s for base metals. Anglo American specifically targeted the quartz-hematite “plug” on the farm Eque 578, located approximately 15 kms north of the Otjikoto deposit. Soil samples were collected and analysed, and magnetic and gravity ground surveys were carried out. The target area was geologically mapped at a scale of 1:500. Three diamond holes were drilled on the farm Eque 578. Two were sited to investigate a gravity low, and another to investigate the quartz-hematite filled fault zone to the southeast of the “plug”. Anglo American concluded that the quartz/haematite “plug” was actually a fault filled feature.

6.2.2 AVDALE

In 1995, Avdale initiated a reconnaissance program in the area targeting base metals. Initial work involved the compilation of all available data, including: water bore hole data, photo-geological interpretation, government airborne data, published geological maps, and reconnaissance mapping (Wilton et al. 2001). The compilation work resulted in the staking of a group of ERL’s followed by EPL’s. In 1996 to 1998, Avdale contracted a number of airborne magnetic surveys over the area. After merging with the government airborne data, 12 high priority targets were identified for testing (Corner, 1997). One of the targets was an intense, NE-SW trending, 9 km long linear magnetic feature centred on the farm Otjikoto (Corner, 1997; Wilton et al., 2001). Ground magnetic, electromagnetic and induced polarization surveys were conducted over the anomaly, followed by a program of shallow (up to 20 metres) RAB drilling in 1998. RAB holes AV109 and AV110 bottomed in mineralization, with visible gold present in these holes. In 1999, a follow up core diamond drilling program lead to the discovery of the deposit. The discovery was followed by resource delineation and definition by Avdale, Teal and Auryx.

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Additional details are outlined in Sections 9 and 10, “Exploration” and “Drilling”, respectively, of this report.

6.3 HISTORIC MINERAL RESOURCES AND RESERVES

In 2005, the first NI 43-101 compliant “maiden” resource for the Otjikoto deposit was prepared by P. van der Merwe (Pr.Sci.Nat.), an employee of ARM, and the results presented in a Technical Report by RSG Global (A.van der Merwe, al. 2005) (Table 6.2) .

Table 6.2:        2005 Inferred Mineral Resources

Inferred Mineral Resource estimate for the Otjikoto Project
Mineralised Zone Tonnes (‘000t) Gold Grade (g/t)
Upper 4,410 1.43
Middle 2,620 0.31
Lower 18,551 1.08
Total 25,581 1.06

Updates to this initial mineral resource are tabulated below in Table 6.3.

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Table 6.3:        Summary of previous mineral resources for the Otjikoto deposit

Report Date Report Resource Class Tonnes
(Mt)
Grade
(g/t Au)
Gold content
(Koz)
January 23, 2007 Wanless and van der Merwe, 2007 Inferred 32.2 1.26 1,300
           
September 17, 2007 Wanless, Minzar and van der Merwe, 2007 Indicated 11.82 1.21 459.2
    Inferred 32.1 1.28 1,320.40
           
August 13, 2009 Wanless and Crisp, 2009 Indicated 28.42 1.34 1,221.50
    Inferred 17.17 1.18 708.4
           
October 25, 2011 McDonald et. al, 2011 Indicated 25.12 1.44 1,160
    Inferred 15.58 1.31 660
           
April 25, 2012 B2Gold Corp. News Release Indicated 24.93 1.74 1,392.69

Mineral Resources reported prior to 2012 and summarized in this Section 6 were the subject of previously released technical reports prepared in accordance with NI 43-101, as referenced above. The Qualified Person has not verified the previous resources and the previous resources reported here are not treated as current mineral resources. Section 14 “Mineral Resource Estimate” of this report documents the updated mineral resources for the Otjikoto Project.

There are no reported historical mineral reserves for the property.

6.4 HISTORIC PRODUCTION

There is no known historic gold or base metal production from the property. Several small-scale amethyst quarries are present on the property but not in the immediate area of the main deposit.

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7 GEOLOGICAL SETTING AND MINERALIZATION

7.1 GEOLOGICAL SETTING

The Otjikoto deposit is located within the Damara Mobile Belt, which forms part of the Pan–African Mobile Belt system (Figure 7.1 & Figure 7.2) . The Damara Mobile Belt consists of two branches, one running approximately parallel to the present Namibian coastline, while the second branch strikes north-eastwards and is referred to as the “Intracratonic Branch”. Otjikoto is located within the northern portion of the Intracratonic Branch (Figure 7.3) .

The Otjiwarongo-Otavi regional area is located in the Northern Central Zone (“NCZ”) and Northern Zone (“NZ”) of the Damara tectonostratigraphic zones (Figure 7.3) . The Otjikoto exploration properties lie predominantly within the NZ. The edge of the Northern Platform (“NP”, also known as the “Otavi Platform”) is to the north of the property in the vicinity of Kombat Mine.

Figure 7.1:        Pan-African mobile belt and location of major mineral deposits

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Figure 7.2:        Geological Map of Namibia (from Geological Survey of Namibia)

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Figure 7.3:        Locality plan of Otjikoto deposit area (indicated by red block) within the tectonostratigraphic Northern and Northern Central Zones

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7.2 REGIONAL GEOLOGY

7.2.1 REGIONAL STRATIGRAPHY

A new, more concise stratigraphic setting for the Otavi area has been proposed by Hoffmann et al. (1996), as shown in Figure 7.4. The stratigraphic column B, Outjo, is applicable to the Otjikoto area.

The Otjikoto area is predominantly underlain by lithologies belonging to the Neoproterozoic Swakop Group of the Damara Orogen. The “basin facies” Swakop Group occurs above the “rift facies” Nosib Group (Figure 7.4) . The Okonguarri Formation, of the Swakop Group, hosts the gold mineralization and is overlain and underlain by distinctive glacial diamictite horizons, the Ghaub and Chuos formations, respectively. The Okonguarri Formation is principally composed of thick units of dark grey carbonaceous marble, biotite-schist, graphitic schist and calc-silicate horizons.

The Ghaub Formation diamictite is poorly exposed about 10 metres below the Karibib Formation marble. The diamictite is overlain by the Cap carbonate unit, which marks the end of the glaciation. Chuos Formation diamictite is exposed in the western and northern portion of the region. The presence of silicate and oxide iron formation distinguishes the Chuos from the Ghaub.

The lower Karibib Formation consists of a 10 to 20 metres wide schistose horizon, which contains minor carbonate bands. The schistose unit is overlain by calcitic marble which forms the bulk of the Karibib Formation. The Karibib marbles form the majority of the Damara outcrop in the region.

The poorly exposed rocks immediately overlying the Karibib Formation have been assigned geophysically to the Kuiseb Formation of the Khomas Subgroup. The Kuiseb Formation occupies topographically flat, depressed areas in the axial zones of the Karibib Marble synclines and consists of schistose quartz-feldspar-mica metagreywacke, metapelite and some metavolcanics.

Domal structures in the southern part of the region have local exposures of pegmatite’s and intrusive rocks. These structures are very well defined by magnetic mapping of the region.

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Figure 7.4:        Stratigraphic column of the northern Damara Orogen lithologies (after Hoffmann et al., 2004), indicating the positions of the Otjikoto and Navachab gold mineralization

The Cretaceous Okorusu alkaline-carbonatite intrusive complex, which hosts a fluorite deposit, is exposed approximately 40 km west of Otjikoto (Figure 7.5) .

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Most of the region is covered by a thick layer of calcrete, transported sand (Kalahari Group) and soil.

The stratigraphic position of the Otjikoto deposit is essentially the same as that of the Navachab gold mine (Figure 7.4) .

7.2.2 REGIONAL STRUCTURE

Depositional patterns within the Damara Sequence were controlled by early rifting followed by spreading. Subduction of the Kalahari Craton beneath the Congo Craton triggered the onset of the Damara orogenic activity at some time between 700 Ma and 650 Ma (Miller, 1983). Geological structures and metamorphism were developed during the subsequent Damara compressional events.

D1 recumbent folding, pre-dating the deposition of the Mulden Group, has been recognized, but major folds can all be related to the D2 deformational event. This event led to the development of a series of major doubly plunging anticlines and synclines throughout the area and across the Northern Platform to the north. The folds vary in closure from open to tight and in attitude from upright to overturned, occasionally recumbent. A fairly abrupt swing in axial trend of D2 folds from NE-SW to more or less E-W roughly north of latitude 20°S has regional significance within the Otjiwarongo-Otavi area. The major D2 folds are internally quite complex structures with numerous parasitic folds on the flanks of the major structures.

A D3 doming event has evidently played a major role in disrupting and refolding D2 fold structures, particularly in the area south of Otjikoto. Several Nosib-cored domes or relatively short, northeast-trending, doubly plunging anticlinal structures can probably be ascribed to this event. Several domal structures in the Otjikoto area have been investigated and found to be related to the intrusion of circular granitic bodies (Figure 7.6) .

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Figure 7.5:        Regional Geology of the Otjikoto Deposit area (B2Gold licenses outlined in red)

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Figure 7.6:        General outline of inferred “Otjikoto” batholith (thick white dashed lines) (Corner, 1996) and smaller dome complexes (thin white dashed lines) in the Otjikoto region (backdrop First Vertical Derivative Magnetics)

Most of the Damara faults so far discovered and mapped in the Otjiwarongo-Otavi Region appear to be related to the D2 deformational event. A number of closely spaced strike-slip faults are interpreted from photo geology and geophysics. These are probably related to the earlier phase of the D2 event. Late brittle faults, particularly of NW and EW orientation are interpreted to be late, possibly of Karoo age (Permian to Jurassic). Some of these faults remain seismically active.

The region is bounded by several major thrusts, which in part define Terrane boundaries. The Waterberg thrust occur to the east of Otjikoto and thrusts younger Karoo formations over older Damara sequence lithologies. The Eckberg and Otjihorongo thrusts have been defined through mapping of repetition of Karibib marble and occur to the east and west of the deposit respectively. West of Otjikoto the repetition of Chuos Formation suggests additional thrusts, as well as tight fold repetition.

Most of the thrusts have some component of strike slip movement.

7.2.3 METAMORPHISM

Goscombe et al. (2004) investigated the structures and metamorphic grade along the northern margin of the Damara Orogen. In the central part of the region he found the metamorphic grade to be middle amphibolite-facies. The local presence of kyanite and sillimanite suggests relatively high pressures and moderate temperatures.

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7.3 LOCAL AND PROPERTY GEOLOGY

Most of the Otjikoto area is covered by calcrete, with the exception of Karibib marble outcrops. The geological knowledge of the area has therefore been principally derived from a combination of geophysical work and drilling (Figure 7.8) . Interpretation of the airborne EM (Electromagnetic) SPECTREM survey, in particular, mapped out the non-conductive marble of the Karibib Formation and the marble units hosted within the Okonguarri Formation. High resolution airborne magnetics and ground surveys have provided additional details on distribution of the main stratigraphic units.

The local stratigraphy is summarized in Figure 7.7.

7.3.1 LITHOLOGIES

Soil is present as a relatively thin 10 to 20 centimetres veneer over the deposit, but commonly ranges to a maximum depth of 2 metres. Aeolian soils and sands of the Kalahari Group have been transported into the area and generally fill local depressions. Due to underlying calcrete, the soil horizon is poorly developed and commonly organic rich and/or mixed with the transported sediments.

Examples of the principal lithologies are shown in Figure 7.9 and briefly described below.

Surficial deposits of leached accumulations of calcium carbonate and calcrete occur over all of the deposit area and range in depth from 2 to 15 metres with thicker development in local topographic depressions and fault zones. The calcrete varies from hardpan to powdered calcrete. The former is hard and commonly has a conglomeratic appearance with rounded to angular fragments of partially to completely calcretized bedrock fragments within a calcrete matrix. The latter represents zones of powdery calcite/clay and is normally found in discontinuous lenses and/or as lining in dissolution cavities within the hardpan calcrete.

A transitional zone below the calcrete consists of weathered, weak to moderately calcretized bedrock that is cut by calcrete bands and fracture related seams of calcrete of variable orientation.

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Figure 7.7:        Local Stratigraphy of the Otjikoto Deposit area (modified after Pesce, 2005)

The oxidation depth commonly ranges from 20 to 40 metres with an average of around 35 metres. The transition to fresh rock is generally fairly rapid, within 1 to 4 metres. As with the calcrete there are “wells” of deeper oxidation commonly associated with fracture and/or fault zones.

Karibib marbles occur as ridge-forming areas of outcrop in the eastern and northern portion of the property. The marbles commonly are white to light grey, medium to coarse grained, massive to banded and predominantly composed of calcite. Locally they are intercalated with biotite schist bands towards the base of the Karibib Formation. The Karibib marbles in the region have thicknesses of 50 to >100 metres. Karibib schist units of variable thickness have been intersected below the marble, principally to the west and south of the main deposit. The schist is commonly dark grey, fine grained, massive to weakly foliated metagreywacke biotite schist.

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Cap Dolomite was intersected above the Ghaub diamictite and is a thin (10 centimeters to 1 metre), commonly cream colored and well laminated unit. The Ghaub Formation was intersected to west the west, east and south of the main deposit and ranges in thickness from 10 to 25 metres. The unit is light grey colored and readily identifiable by the presence of elongated clasts of variable composition ranging in size from <1 centimeter to +5 centimeters. The clasts represent glacial derived drop stones and thus the unit is a

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Figure 7.8:        Local Geology of the Otjikoto Deposit

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diamictite. The Ghaub is strongly sheared to the west of the deposit. A dark grey to black, sometimes graphitic, shale/schist unit underlies the Ghaub. West of the deposit this unit contains disseminations and lenses of “vuggy” fine grained pyrite. Thick successions of biotite-schist and garnet-biotite schist of the Okonguarri Formation occur below the black schist.

The Okonguarri schist units, which form the hanging and footwall lithologies of the deposit, are variably composed of a mix of biotite and feldspar (plagioclase dominant) with lesser quartz, dolomite, calcite and garnet. They have poor to well-developed schistosity with foliation commonly parallel or sub-parallel to primary bedding or metamorphic banding. Graded bedding is preserved in some of the schist. The schist units are derived from semi-pelitic, pelitic, marl and psammitic units in a turbiditic sedimentary package. The schist tends to be quiet competent but of variable hardness, usually directly related to the biotite content and/or degree of schistosity. Three main marble bands (hangingwall, OTB and footwall) occur in the Okonguarri Formation. The hanging wall marble is a discontinuous coarse grained marble horizon(s) which is believed to have been boudinaged during folding or thrusting. The OTB marble forms the structural footwall of the main mineralized zone and represents the most important, and distinctive, marker horizon in the deposit area. Most drill holes within the deposit bottomed in the OTB marble. The OTB consists of weakly banded, white to light grey, medium to coarse grained calcite crystals. The thickness of the OTB is relatively uniform (averaging 10 metres) throughout the deposit area. It forms an essentially flat tabular unit dipping at approximately 25 degrees to the east-southeast (125° azimuth). The footwall marble is located approximately 20 to 30 metres below the OTB marble in the main deposit area. The footwall marble is similar in appearance to the OTB but tends to have more biotite crystals present, forming weak banding. It is commonly 15 to 20 metres thick.

Biotite and biotite-garnet schist units of the Middle Okonguarri Formation occur below the footwall marble.

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Figure 7.9:        Examples of some of the principal lithologies in the deposit (modified after Pesce 2005)

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7.3.2 ALTERATION

Albitization of all the lithologies in the Otjikoto area is widespread. Chlorite, amphibole, carbonate and biotite alteration are also common. Depending on the intensity and type of alteration and original host rock composition, the following alteration “lithologies” have been systematically logged: albitite, biotite-albitite schist, garnet-albitite-biotite schist, amphibole bearing albitite, garnet-amphibole metasediment and banded biotite-amphibole metasediment. Calcareous biotite schist is also present, with much of the carbonate believed to be introduced.

The amphibole bearing units have been grouped together in what is called a HORN “unit” and the albite altered beds/bands in the ALB “unit”. Both the HORN and ALB units are more competent and harder than the SCHIST “unit”. Where garnets are present the schist lithologies are also abrasive.

The albitization throughout the sedimentary sequence indicates widespread sodium metasomatism. The intense albitization of the granofels and schist observed at Otjikoto is unusual in the Northern Zone of the Damara Orogen. The albitization may have occurred relatively early in the tectonothermal history and may be unrelated to the gold-mineralizing event. However, albitization probably changed the rheology of the host rocks and made them prone to fracturing during deformation.

Figure 7.10 presents a schematic cross sectional representation of the general distribution of the principal rock types present in the deposit area.

7.3.3 LOCAL STRUCTURAL GEOLOGY

The rocks drilled in the Otjikoto area have experienced at least three phases of moderate to tight folding and some thrusting. They have also been affected by extensive metasomatism, followed by prograde regional metamorphism in the upper greenschist to lower amphibolite facies. Nevertheless, primary sedimentary layering is often well preserved, with graded bedding on the millimeter- and centimeter-scale indicating that the stratigraphy is generally the right way up.

The dominant foliations in the deposit area are parallel to sub-parallel to the bedding (S0) and/or metamorphic banding (S1) and strike 035° with dip at 25° to the southeast.

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Figure 7.10:        Schematic section looking north through main deposit showing distribution of principal lithologies (Pink area shows rough outline of mineralized zone; Yellow is the calcrete and transitional zone)

It is not known if the foliation represents transposed bedding or a foliation developed in response to tight isoclinal folding. A second set of foliation (S2 or S3?) is rarely developed, usually in association with “S”, “Z” or “W”/”M” folds.

Rotation of garnets and other local shear fabrics indicate the presence of ductile shear zones within the deposit area. The exact orientation of these shear systems is still under review. The shears are believed to be of Damara age.

Pesce et al. (2010) defined a mineral lineation with average trend of 010° and plunge of 11° degrees to the southwest. In some cases it was observed that the sulphides follow the same lineation direction.

A broad open synform, east of the deposit, is defined by the Karibib marble geometry and is cored by Kuiseb schist. West out of the synform the Okonguarri has been partially sheared and compressed with tight isoclinal, recumbent fold geometries present. The east hanging wall of the deposit is defined by a recumbent antiform with parasitic folding along the limbs. Further to the west the Ghaub and Karibib are repeated in a tight, attenuated, recumbent synform. A major thrust in the footwall of the deposit, which repeats the stratigraphy, is postulated to accommodate the geometry of the area. It is also presumed that there may be a thrust or shear zone below the Karibib/Ghaub on the east side of the deposit area owing to the change in fold geometries across this contact.

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Detailed structural studies of the Otjikoto deposit by Dr. Karl Kasch and John Fedorowich, have documented complex folding in the hanging wall and footwall of the deposit. In the northern part of the deposit the footwall marble is folded into a complex Z fold geometry. To the south the work has identified a series of opposing Z (eastern side of deposit) and S (central portion of deposit) fold geometries in the hanging wall schist. It is suspected that sheath folds have developed corresponding to these geometries. While these folds follow the general plunge (10-15 degrees to the southwest) and orientation (azimuth 020° to 035°) of the mineralized zone, no direct link between these fold geometries and mineralization has been identified to date.

Several generations of late brittle faults cross cut the deposit area. The fault locations are based on a combination of photo-geological interpretation, topography, drill core logging and geophysical interpretation. A summary map is provided in Figure 7.11.

There are two principal directions of late brittle faults in the deposit area: NNW and EW-ENE. These structures are inferred to be Karoo to post-Karoo age structures. Some of the NNW structures remain seismically active. The NNW faults are commonly steep westerly dipping. Displacement along the faults within the proposed pit area is generally less than 10 metres.

7.4 MINERALIZATION

Gold in the main Otjikoto deposit is hosted by a NNE striking sheeted sulphide (+ magnetite) - quartz+carbonate vein system. The system has been traced over a strike length of 2.3 kilometres, to a depth of 475 metres below surface. The deposit shows up as a very pronounced regional magnetic high anomaly (Figure 7.12) due to the magnetite and pyrrhotite content. The vein swarm is lying at an angle of 20° to 30° to deep-seated NNE trending linear structures, which could be an extensional array that formed as a result of late Damara dextral movement (Steven, 2001). All significant hydrothermal gold deposits, as well as most leucogranite-hosted uranium mineralization in central Namibia that have been discovered to date, are hosted within these NNE trending corridors (Steven, 2001).

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Figure 7.11:        Summary of faults in the Otjikoto Deposit area. (Major faults are drawn as thicker lines, inferred faults and/or fracture zones as dash lines; dark blue lines indicate water bearing faults; light blue inferred water bearing faults or fracture zones)

The gold occurs in a series of thin (commonly <10 centimetres) sheeted veins in the schist and granofels (meta-marls) of the Upper and Middle Okonguarri Formation. The majority of the veins lie parallel to an S0/S1 transposition foliation which approximates bedding, and the term “bedding” is used in the rest of this report for simplicity.

The Otjikoto gold deposit lithology has been divided into three lithostratigraphic units. The OTC albitite-hornfels unit hosts most of the mineralized vein system and is underlain by the 6 metres to 10 metres thick un-mineralized OTB calcitic marble. The albitized OTA fels (~30 metres thick), which hosts minor bedding-parallel veins with irregularly distributed gold values, occurs between the OTB marble and the footwall marble (~20 metres thick). The OTA fels and the OTB marble are part of the Middle Okonguarri Formation and the OTC is the basal unit of the Upper Okonguarri Formation (Figures 7.13, 7.14 and 7.16) .

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Figure 7.12:        Analytical Signal airborne magnetics showing strong regional anomaly associated with the Otjikoto deposit area

Figure 7.13:        Simplified cross-section through the Otjikoto mineralization hosted by sheeted sulphide-quartz veins within the Okonguarri Formation (the vertical lines represent drill holes)

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Figure 7.14:        General location of the mineralized zones and principal mineralogy of the veins

The sheeted veins and related gold mineralization occur in a series of distinct en-echelon zones oriented at approximately 010° to 020° NNE and plunging at 10-15° (average 12°) to the SSW (Figures 7.15 and 7.17) . The bulk of the mineralization occurs in the Central and West shoots. Mineralization in the West 1 and Main Magnetite zones is not considered of significance for the resource at this point in time. The newly discovered Wolfshag zone is described in Section 10 “Drilling” of this report.

Vein concentrations range from 1 to 30 veins per metre with higher vein concentration within the shoots.

The sheeted vein system at Otjikoto is sulphide rich with up to 100% sulphides present in some veins, but averaging 20-30% sulphides (Figure 7.18) . In the shallower and northern portions of the deposit the veins are pyrite rich while in the southern and deeper levels of the deposit the veins are pyrrhotite dominant. Veins proximal to the OTB marble, the bottom or footwall veins tend to be carbonate rich while those in the rest of the deposit contain a mix of quartz - calcite and iron carbonate. Pyrite within the pyrite dominant veins tends to have a “vuggy” texture.

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Magnetite represents an important constituent of the veins, ranging from a fraction of a percent to 100% of individual veins. Granular to massive magnetite is more common than bladed “feathery” magnetite. The granular magnetite veins are commonly hosted by the intensively albitized metasediments.

Quartz, calcite, dolomite, siderite and ankerite are the main gangue minerals. Other important gangue minerals in the veins include garnet, amphibole and to a lesser extent chlorite and muscovite/sericite. Silicate + sulphide vein assemblages show very variable proportions of the constituent minerals. Garnets within and on the haloes of veins tend to be larger than the host rock garnets. Chalcopyrite is infrequently present and rare and minute (<10 µm) grains of maldonite (Au2Bi) were observed in the metallurgical test work of the pyrite dominant veins. Marcasite is also present in the shallower, pyrite-dominant vein systems. Within the oxide zone the sulphides commonly weather to hematite and/or limonite.

Gold occurs within the vein system as coarse native gold with a size variation from 5 µm to 400 µm, with the median at about 100 µm. No specific location for gold has been noted. It has been observed adjacent to and within sulphides, along fractures, adjacent to and within garnets, within magnetite, on the edges of amphiboles and chlorite, and as free gold in quartz and carbonate (Figure 7.19) .

As a general rule:

  • More visible gold is present in the pyrrhotite dominant mineralization;
  • The more complex the vein mineralogy the more likely it is to have significant gold values;
  • Thicker veins tend to be lower grade;
  • Higher intensity of sheeted thin veins generally yields better gold grades;
  • Quartz only veins do not commonly carry significant gold; and
  • Gold is more common on the margins of veins.

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Figure 7.15:        Location of ore shoots at Otjikoto (Pink = 2012 Leapfrog 0.8 g/t Au wireframe) (Inset: Shoots relative to mineral lineations (after Pesce et al. 2010))

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Figure 7.16:        Summary Detailed log through the Otjikoto mineralization. Of note is the concentration of the gold mineralization (see column on the right) within quartz veins hosted by albitites and hornfels (amphibole granofels), with the highest values attained in the unit (OTC) above the OTB marble.

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Figure 7.17:        Schematic longitudinal section showing continuity of mineralization (and gold grades) down the West Shoot. (After Auryx Gold presentation, 2011).

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Figure 7.18:        Examples of principal vein mineralogy assemblages:

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Figure 7.18 cont’d:        Examples of principal vein mineralogy assemblages

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Magnetite Rich Zones

The Main Magnetite zone to the south of the deposit has horizons or lenses of massive magnetite up to several metres thick. The magnetite is replaced by massive hematite in oxidized zones. Gold is present as sporadic fracture fill in the Main Magnetite zone.

Figure 7.19:        Plates of examples of Visible Gold occurrences (modified after Pesce, 2010) (Po pyrrhotite; Alb albitite; Grt garnet; carb carbonate; Qtz quartz)

8 DEPOSIT TYPES

Historic exploration in the Otjikoto Project area has concentrated on the search for base metal deposits. The focus was on either Tsumeb or Berg-Aukas styles of deposits, owing to the relative proximity to the Tsumeb or Kombat and Berg-Aukas base metal mines. More recently, the recognition of cratonic margin and basinal settings and thus potential for Kupferscheifer style base metal deposits has been used as an exploration model in the region.

With the discovery of the Navachab gold mine, near Karibib, in 1984 and Otjikoto in 1999, some of the exploration efforts have shifted to gold. Coincidentally both these gold deposits were discovered while companies were exploring for base metals. Navachab is the only significant producing gold mine in Namibia. It has received a relatively high degree of study but the genesis of the deposit is still in debate. Neither Navachab nor Otjikoto fall within any “typical” ore deposit classification.

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The key features of the Navachab deposit are as follows (Steven et al., 2008 and Wulff et al 2010):

  • Sediment hosted with principal host rocks marbles, biotite schist’s and calc-silicates;
  • Amphibolite to granulite metamorphic facies;
  • Two distinct styles of mineralization:

1) Main “high grade” shoot is a semi-massive sulphide replacement body, the MC skarn
2) Strong sheeted quartz-sulphide vein system;

  • The MC skarn shoot has a pipe like geometry that plunges at 020° azimuth to NNE in excess of 2,000 metres;
  • A second similar shoot sits structurally above and to the northeast of the MC skarn;
  • The sheeted vein zone is a large, lower grade, shear derived, extensional vein array;
  • Skarn assemblages: garnet+tremolite/actinolite+clinopyroxene+Kfeldspar+quartz depending on host lithology;
  • The main ore mineral is pyrrhotite, with minor chalcopyrite, sphalerite, pyrite, arsenopyrite and Bismuth (Bi) minerals such as maldonite;
  • Gold is most commonly free with lower amounts locked within maldonite;
  • Ore body is cut by lamprophyre, aplite and pegmatite’s dykes;
  • Damara age, structurally controlled mineralization related to a combination of doming, dextral shearing and thrusts; and
  • Low magmatic component to hydrothermal fluids but possible high heat flow due to intrusions.

Similarities of Navachab and Otjikoto include:

  • Nearly identical relative stratigraphic position;
  • Similar ore shoot direction and shallow plunge but different plunge direction (south versus north);
  • Gold is free;
  • Pyrrhotite dominant in deeper levels; and
  • En-echelon vein systems and shoots extending to depth.

Given the similarities between the deposits, exploration in the Otjikoto region should partially use a Navachab style model as a tool. However, there are some significant differences between the deposits as well at Otjikoto. These include:

  • Lack of intrusive association;
  • High magnetite content of veins;
  • Low quartz but high sulphide veins;
  • No direct skarn or calc-silicate association;

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  • Strong sodic enrichment in alteration (which may in part have played a role in creating a more brittle host rock);
  • Very low bismuth and low arsenic;
  • Generally higher free gold content (more nuggety); and
  • Zonation from pyrite to pyrrhotite dominant domains.

Both the Navachab and Otjikoto gold deposits are believed to fall under the general classification of orogenic lode gold deposits. In both cases the systems are open to depth and show both lateral and vertical continuity down plunge on shoots, which is typical of Archean and Neoproterozoic orogenic systems.

Sheepers (2000), and more recently Fedorowich (2011), Gasparotto (2012) and Buttner (2012), have documented evidence for shear strain in the vein systems and host lithologies at Otjikoto. Early in the exploration history at Otjikoto, Steven (2001) provided evidence for a dextral shear array model for the vein development. Pesce et al. (2005 and 2010) have noted the presence of a mineral lineation parallel to the shoots and that mineralization post-dates the development of the lineation. Early and more recent work by Kasch (2004 and 2012, personal communication.) and Fedorowich (2011) has shown a complex fold history for the host metasediments, which may have played a role in localizing the ore shoots and or refolding earlier veins. In addition, stratigraphic work based on drilling, mapping and geophysics indicates that there is repetition of stratigraphy by thrusting in addition to the folding. This work indicates that mineralization may in part be related to these thrusts. While the exploration model is still a work in progress, the above ideas can be used as tools to help guide exploration in the region.

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9 EXPLORATION

The bedrock geology within the deposit area is covered by 10 to 15 metres of calcrete, with only sporadic outcrops of Karibib Marble. The exploration program therefore had to rely on a combination of airborne and ground geophysical surveys to map bedrock geology and identify exploration targets for drill testing. Systematic drilling of the geophysical anomalies led to the discovery of the Otjikoto gold deposit. Geochemical sampling over the deposit was able to locate erratic anomalous values of gold but no coherent anomalies.

9.1 GEOPHYSICAL SURVEYS

9.1.1 AIRBORNE SURVEYS

A number of airborne geophysical surveys have been flown over the deposit area and high resolution government airborne geophysical data (Geological Survey of Namibia (MME)) and ASTER imagery was purchased (Table 9.1) . These datasets were merged and interpreted by Corner Geophysics (Dr. Branco Corner) and more recently by EarthMaps (Klauss Knupp).

The Otjikoto deposit is indicated by a distinct magnetic high in the analytical signal of the processed high resolution airborne magnetic data covering the Otjikoto area (Figure 9.1) . This anomaly can be ascribed to the magnetite and pyrrhotite content of the mineralization.

The first vertical derivative calculated from the high resolution geophysical magnetic data also delineates the Otjikoto deposit (Figure 9.2) . This signature is not as striking, but it shows more detail. The complex structures in the area are clearly mapped by first vertical derivative plots.

Table 9.1:        Summary of Airborne Surveys and Imagery

Survey Year Type Line spacing (m) Line
kilometres
Fugro Geodass 1998 Magnetics & radiometrics 200 2,500
GEOTEM AEM 2001 Electromagnetic 400 1,290
Namibian MME magnetics 2004 Magnetics & radiometrics 200  
Aster   Aster Satellite Imagery    
Spectrum AEM 2005 Electromagnetic 200 3,655
NRG High Resolution Helicopter 2011 Magnetics & radiometrics 50 8,650

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Figure 9.1:        High resolution NRG magnetics Analytical Signal

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Figure 9.2:        High resolution NRG magnetics First Vertical Derivative (1VD)

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A strong electromagnetic signature is also present over the deposit area, reflecting the conductive sulphide rich sheeted veins. The airborne electromagnetic (“AEM”) survey also provides a valuable tool for bedrock mapping of the area; in particular the marble horizons are indicated by areas of low conductivity (Figure 9.3) .

9.1.1.1 Ground Geophysical Surveys

Various ground geophysical surveys were conducted over the deposit area to follow-up airborne geophysical anomalies. Ground geophysical surveys were done at grid line spacing ranging from 50 to 200 metres.

The following ground geophysical techniques were completed by previous operators:

  • Ground magnetics (468.7 line kms);
  • Gradient Array Induced Polarization (“IP”) (286.7 line kms);
  • Fixed loop Time Domain EM;
  • Controlled Source Audio Magnetotellurics (“CSAMT”) – (single 4 km line);
  • Natural Source Audio Magnetotellurics (“NSAMT”) (18.0 line kms); and
  • MAXMIN (Frequency Domain EM) (96.7 line kms).

9.1.1.2 Ground Magnetics

Previous operators found detailed ground magnetics surveys to be a useful tool to aid in defining the mineralized zones; however, shallow, magnetite rich bands, particularly in the southern and eastern portions of the deposit, partially mask the more important pyrrhotite rich sheeted vein mineralization. The 2011 NRG airborne magnetic survey provided as good or better resolution than the ground magnetics and is now a key tool in exploration targeting.

9.1.1.3 Induced Polarization

The entire Otjikoto grid was surveyed with a gradient array (a=50 metres) using a number of electrode setups. The central portion of the survey area displays two sharp, fairly linear and very strong (30 to 40 m/sec) chargeability responses (Figure 9.4) . IP response is stronger over the magnetite and pyrrhotite rich domains than the pyritic domains.

The extreme northern portion of the grid contains a large and more diffuse chargeability response, which still displays amplitudes in the range of 34 to 38 msec.

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Figure 9.3:        SPECTRUM AEM channel 5 and EM conductor picks

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Figure 9.4:        Ground-induced polarization (IP) survey Chargeability image

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The change in the shape of the response could be due either to the presence of a broader mineralized system or to a change in attitude (dip) of the anomalous unit.

The entire survey displays a trough of low resistivity (<150 ohm m) surrounded on both sides by a relatively high resistivity host (>2,000 ohm m). Within this broad zone of low resistivity, a number of discreet anomalous zones can be recognized. These appear to correlate well with the main chargeability zones. The large magnetic responses are also associated with some of these low resistivity zones.

In 2012, a 37.1 line km Pole-Dipole array IP survey was conducted over the deposit and proposed (January 2012) infrastructure areas as a tool to define targets for sterilization drilling. The pole-dipole survey proved useful at defining zones with disseminated pyrite and thin lenticular pyrite veins in schist and black shale.

9.1.1.4 Ground Electromagnetics, NSAMT and CSAMT

Previous operators conducted a number of fixed horizontal loop frequency domain Max-Min electromagnetic (EM-34) surveys over the entire area of interest, with a large emphasis in the area of strong magnetic/IP responses. In most instances, relatively good conductors were located at fairly shallow depths, except in the southern portion of the farm Otjikoto where depths in excess of 200 metres were interpreted. It is worth noting the decrease in conductivity in both the northern and southern portions of the survey area despite the presence of good chargeable material. This indicates a change in the nature of the mineralization away from the main magnetic responses.

NSAMT and CSAMT were tested at Otjikoto by previous operators but it was found that the SPECTRUM AEM data provided similar or better results.

9.2 GEOCHEMICAL SAMPLING (SOIL SURVEYS)

Avdale completed an initial orientation survey over the Otjikoto deposit in 2001 which found the -180 fraction collected at the calcrete soil interface was the best medium to sample. In general, the soil horizon is poorly developed over the 5 to 15 metre thick calcrete regolith. The coarse gold which is characteristic of the deposit makes it difficult to get a representative sample at any sample site. The initial soil survey thus identified spot anomalies of gold but no well-defined anomaly over the mineralization.

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10 DRILLING

10.1 INTRODUCTION/SUMMARY

Drilling at the Otjikoto Project started in 1998 with a RAB shallow drill program testing geophysical targets. The initial RAB program resulted in the discovery of gold mineralization and was followed up with additional RAB drilling and the first diamond drill core hole program in 1999. The third hole of the 1999 program, OT3, is considered the discovery hole of the deposit. Drilling of a more regional nature was followed by delineation drilling, to determine the distribution of mineralization and grade continuity within the zones. In the shallower levels of the deposit the majority of drilling was completed with RC. Drill hole spacing in the shallower portions of the deposit is 25 metres x 25 metres. Deeper level spacing ranges from 25 metres x 50 metres and 50 metres x 100 metres. A total of 1,554 holes totalling 195,717 metres have been drilled on the farms Otjikoto and Felsenquelle to October 2012 (Table 10.1; Figures 10.1 to 10.2) .

10.2 DRILLING METHODS AND EQUIPMENT

The following description is modified after McDonald, et al. (2011). Three drilling techniques have been used on the Otjikoto Project at various stages of exploration and through various phases of work:

  • RAB percussion drilling was used in the early exploration phases to recover geochemical samples below the calcrete cover and allow for basic geological and regolith assessments.
  • RC drilling was used to test exploration targets and as part of the advanced exploration, prefeasibility and feasibility evaluation of Otjikoto. This method generates a large sample mass and provides improved recovery in the near surface, often fractured oxide mineralized zone.
  • Diamond drilling with HQ (and TNW) and NQ core diameters to test exploration targets, to evaluate geological environments and to provide geological and structural information at Otjikoto. Diamond drilling was used during the exploration and mineralization delineation and definition phases. PQ size core was drilled in a limited number of holes for the collection of metallurgical test samples.

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Table 10.1:        Summary of drilling on the Otjikoto Deposit (farms Felsenquelle and Otjikoto (with farms Wolfshaag & Gerhardshausen)

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Figure 10.1:        Location of Otjikoto RAB Holes (all pre B2Gold)

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Figure 10.2:        Location of Otjikoto Diamond Drill Holes (“DDH”)

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Figure 10.3:        Location of Otjikoto RC Holes

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10.2.1 REVERSE CIRCULATION DRILLING

Drillcon Africa (Pty) Ltd, Ferrodrill Namibia (Pty) and Xplor Drilling (Pty) Ltd have been employed to carry out the RAB and RC drilling on the project. RC drilling utilized a 4¾ inch (~120 mm) bit, giving a drill hole volume of approximately 11 l/m which relates to a theoretical sample weight of 30 kg/m, assuming 100% recovery and a bulk density of 2.8 t/m3. Experience gained at Otjikoto indicates that 26 kg to 28 kg/m is an average sample mass obtained by RC drilling.

10.2.2 DIAMOND CORE DRILLING

Two drilling contractors, Drillcon Africa (Pty) Ltd and Ferrodrill Namibia (Pty) Ltd, have provided the core drilling services over the length of the project. The contractors supply all required drilling equipment, trained personnel and safety equipment. HQ (63.5 millimetres) is commonly used to ‘case’ through the calcrete and transitional weathered zone to a depth between 30 and

50 metres, after which depth the hole is continued with NQ (47.6 millimetres). HQ, TNW (60.8 millimetres) and/or PQ (85 millimetres) sized core was used for the full depth of the hole on selected holes and holes drilled for metallurgical test sampling.

Drill hole depths and core recoveries are calculated on a run-by-run basis from measured stick-up of the quill rod and recorded in the field in the daily drilling reports. Any gains or losses are noted in the field and depth corrections made as needed and checked against the rod count and stick-up.

10.3 DRILLING PROCEDURES

10.3.1 DRILLING ORIENTATION AND SPACING

Initial exploration holes at Otjikoto were angled (10 holes), but owing to the relatively shallow dip of the mineralization and structural fabric it was decided that the deposit could be delineated with vertical holes. All of the RC holes were drilled vertically. Condemnation and geotechnical holes were drilled as angled holes during the prefeasibility and feasibility drilling.

Drill sections orientation (305° Azimuth) was originally set up perpendicular to the geophysical targets and structural fabric of the area. The mineralization shoot direction was recognized at a much later stage and therefore the drilling sections are slightly oblique to the shoots, but not to the overall trend of the deposit.

The shallower portions of the deposit were drilled on a 25 metres x 25 metres drill grid and the deeper levels at 25 metres x 50 metres and 50 metres x 100 metres drill holes spacing. Infill holes were usually stopped just before or within the OTB marble marker horizon.

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Condemnation drilling consisted of the drilling of fences of holes across areas of proposed infrastructure and the drilling of selected geochemical, geophysical and geological targets within each area. Metallurgical holes twinned existing RC or diamond drill holes.

10.3.2 RECOVERY

Drilling quality for both RC and diamond drilling was generally very good. The recovered sample masses are recorded by field geologist and recorded on appropriate field logging sheets. The overall average RC mass recovery is 87.45% . The average RC recovery within the mineralized zones above 40 metres is 80.85% and the average recovery below 40 metres is 89.72% . The relatively lower percentage recovery above 40 metres is related to the oxidized nature of both host rock and mineralization.

Core recovery is documented at the drill by the drilling contractor and again during the geotechnical logging and for every sample during detailed logging. Apart from the very weathered near-surface core, the diamond core recoveries were very good, averaging 99.88% within the mineralized zone.

In the opinion of the author there are no sampling or recovery factors which could materially impact the accuracy and reliability of the results used for the resource model.

10.3.3 CORE HANDLING

Core was obtained using wire-line methods and unloaded from the core barrel, as drilled, into a 3 metres “V” groove core stand, washed and fitted together. Each run is then measured and recoveries calculated. Core is then systematically placed into wooden (or aluminium) core trays in the same orientation as it came out of the core barrel.

Blocks were placed after each core run, indicating depth and recoveries for the run. A written record of each run is maintained by the drilling contractor. Core boxes were marked with the drill hole identity number, the intersection interval (start and final depths in that box), an arrow indicating which side is down-the-hole, and the box number. The driller, drilling supervisors and drill geologist ensure the correct placement of the drill core in the core boxes.

Core boxes are collected in the field by B2Gold staff and delivered to the B2Gold core logging and storage facility in Otjiwarongo.

10.3.4 SURVEY CONTROL

Holes were initially spotted on cut grid lines, with stations and lines based on a combination of theodolite and measuring tape. For angled holes the set-ups made use of siting pickets for drill orientation line up. The line-up was checked with compass and head inclination set with Brunton compass or degree rule device.

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All drill holes were surveyed using either a single or multi-shot down hole survey instrument (e.g. Reflex Ez-shot) which records the azimuth (magnetic) and dip. Prior to 2012, downhole survey measurements were taken at 50 to 100 metre intervals and in 2012 at 25 metre intervals within the main deposit. Corrections to the magnetic azimuth were made in accordance to the magnetic field declination in the area. Upon completion of a set of drill holes, the holes were surveyed by a contract professional land surveyor.

10.4 LOGGING PROTOCOLS

10.4.1 RC DRILL CHIP LOGGING

All RC drill holes were logged at the drill site during drilling with logging based on examination of a representative sample of washed chips collected from the sample split. RC geological logs capture the following:

  • UNIT CODE: a summarized geological classification of the depth of oxidation and major identifiable units.
  • ROCK TYPE: an observation of the predominant rock type in each metre sample.
  • VEIN (RC): a visual estimate of the total percentage of vein material per metre sample and the relative proportion of important economic and gangue minerals.

A representative sample of chips from all RC holes is retained in small plastic jars for future reference.

10.4.2 DIAMOND DRILL CORE LOGGING

B2Gold has standard procedures documentation for core yard operations. In summary, the following are the principal steps involved in core documentation:

  • Checking of core blocks, realignment of core and placement of metre marks on core followed by;
  • Confirmation, then marking of start and end points on core boxes;
  • Geotechnical logging of core;
  • Orienting of core and placement of “low line”;
  • Geological logging of core, including markup of samples;
  • Placement of “sample” line on core; and
  • Photographing of core.

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“Low line” and reference lines are indicated on the core according to standard procedures before geological logging is commenced. The “low line” is positioned as the line of maximum apparent dip of the fabric in the core. The placement of the low lines allows for consistent unbiased sampling and also was originally used as an aid in the measurement of structures.

Historically the core logging was conducted by recording information in two log types for each borehole:

  • Geological log of main units and rock; and
  • Vein log including each individual vein and its mineralogy.

In 2012, the core logging format was changed to allow for a capture of all crucial geological data on a single form. In addition, the change in format allowed for the entry of alteration information and structural data, which had previously not been captured. To aid in the resource modelling (Exploratory Data Analysis “EDA” and related indicator studies) the data for certain parameters was simplified through “binning”.

10.5 DRILL PLAN AND SECTIONS

A plan map and several representative longitudinal and drill sections are presented in Figures 10.4 to 10.11.

10.6 INFILL DRILLING

29 “in-pit” infill holes totaling 4,267 metres were drilled in 2012 by B2Gold Namibia. Results for these holes were similar to the adjacent holes drilled by previous operators.

10.7 METALLURGICAL DRILLING

To provide sample material for testing of the different ore types, 38 new metallurgical test sample drill holes totalling 2,663.69 metres were drilled in 2012 and an additional 27 “historic” holes were sampled (Figure 10.12) . All 2012 metallurgical holes were twins of previous RC and/or diamond drill holes. Additional details are provided in Section 13 “Mineral Processing and Metallurgical Testing” of this report.

10.8 GEOTECHNICAL DRILLING

Eight holes totalling 1,516.02 metres were drilled in 2012 for open-pit stability studies (Figure 10.13) . An additional ten shallow drill holes were drilled for civil studies of the proposed process plant and crusher sites. The detailed geotechnical logging and sampling of these holes was completed by personnel from SRK, Johannesburg, South Africa. Results of the geotechnical study are presented in Section 16 “Mining Methods” of this report.

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10.9 CONDEMATION DRILLING

49 holes were drilled in 2012 as condemnation of the proposed infrastructure locations (Figure 10.14) . No significant zones of mineralization were intersected under the proposed areas of infrastructure.

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Figure 10.4:        Drill plan map

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Figure 10.5:        Schematic Longitudinal Section 4675E, showing Bloy (Nicholls, 2011) 0.4 g/t gold grade shell (orange and blue lines indicate 2011 pit shell iterations) (All holes are pre B2Gold)

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Figure 10.6:        Schematic Longitudinal Section 4725E, showing Bloy (Nichols, 2011) 0.4 g/t gold grade shell (orange and blue lines indicate 2011 pit shell iterations) (All holes are pre B2Gold)

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Figure 10.7:        Schematic Longitudinal Section 4850E, showing Bloy (Nichols, 2011) 0.4 g/t gold grade shell (orange and blue lines indicate 2011 pit shell iterations) (All holes are pre B2Gold)

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Figure 10.8:        Schematic Cross Section Line 7050N, showing Bloy (Nicholls, 2011), 0.4 g/t gold grade shell (orange and blue lines indicate 2011 pit shell iterations) (All holes are pre B2Gold)

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Figure 10.9:        Schematic Geological Cross Section Line 7250N (from Teal 2009 internal report)

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Figure 10.10:        Schematic Cross Section Line 7250N, showing Bloy (Nicholls, 2011) 0.4 g/t gold grade shell (orange and blue lines indicate 2011 pit shell iterations) (All holes are pre B2Gold)

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Figure 10.11:        Schematic Geological Cross Section Line 7400N (from Teal 2009 internal report)

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Figure 10.12:        Location Map of 2012 Metallurgical test sample drill holes

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Figure 10.13:        Location Map of 2012 Pit Stability Study Geotechnical drill holes

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Figure 10.14:        Location Map of 2012 Condemnation drill holes (backdrop DRA site layout L002-SHT_180)

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10.10 EXPLORATION DRILLING (POTENTIAL RESOURCE EXPANSION)

10.10.1        DOWN PLUNGE AND EAST

Exploration drilling down plunge on the main shoots and to the east has indicated that the main Otjikoto deposit remains open to depth down plunge on the principal shoots. As an example, hole OT126, drilled by Teal in 2006 on section 6100N on the Western shoot, returned 3.0 metres at 45.84 g/t gold and 19.0 metres at 5.01 g/t gold starting at depths of 229 and 246 metres down hole, respectively. Significant albitic alteration at depth to the east suggests the hydrothermal system was quite “robust” and potential remains to define more mineralization.

10.10.2        MAIN MAGNETITE ZONE

Detailed RC drilling in the Main Magnetite zone to the south of the main deposit located erratic gold values within the “magnetite” zones. At the present time this area is not reported in any resource but potential remains to define resources in this area with further drilling. Additionally, approximately two kilometers further to the south a zone of very strong albitization with limited veining has been identified in several drill holes (e.g. GH12-016 and GH12-017) on the farms Gerhardshausen and Felsenquelle. Gold grades in these holes are low but the alteration suggests potential for this area.

10.10.3        WOLFSHAG ZONE

Following structural interpretation work in the winter (August – September) of 2011, a series of holes were drilled, and several old holes deepened on the farms Wolfshag and Otjikoto, starting at the northeast end of the proposed open pit (Figure 10.15) . Significant mineralization was intersected over 800 metres, from section 7900N to 8700N, within a zone of interpreted tight folding. Better results from the new Wolfshag (K2) zone include: OTG10D with 7.55 g/t gold over 20.0 metres (Figure 10.16) and WH31 with 4.44 g/t gold over 20 metres (Auryx Gold news release dated November 23, 2011).

The Wolfshag zone has been drilled on 200 metre sections with drill holes spaced at 25 to 50 metres on section. Insufficient drilling has been completed to define a resource on this zone but apparent continuity appears to be quite good and, given the close proximity to the existing resource, this zone represents the best current exploration target on the property.

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Figure 10.15:        General Location Map of Wolfshag (K2) Zone

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Figure 10.16:        Wolfshag (K2) Zone Section 7900N (Values after Auryx Gold news release Nov 23, 2011 and B2Gold news release Dec 11, 2012)

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10.11 NEW HOLES AND INFORMATION AFTER DATABASE CLOSE-OFF

A total of 58 holes were drilled on EPL 2410 following the close-off date for information for the database. Of these holes, only 10 were within the proposed pit area. These holes were drilled as infill for geostatistical crosses for future resource updates.

Assay results for a series of infill holes drilled prior to the database close-off were received after close-off. The resource is based solely on the information received prior to the database close-off. Although the newer assays within the resource model will change the grades locally, overall the new assays and drilling should have a minimal effect on the average grade of the model, and should increase confidence for some blocks.

10.12 COMMENT ON SECTION 10

The quantity and quality of the DDH and RC drilling conducted on the property is sufficient to meet the requirements for support of a mineral resource estimate. The collar, and down-hole surveys and lithological data collection and methodologies meet or exceed industry standards.

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11 SAMPLE PREPARATION, ANALYSIS AND SECURITY

11.1 SAMPLE COLLECTION

11.1.1 REVERSE CIRCULATION (“RC”) SAMPLES COLLECTION

The following description is modified after McDonald et al. (2011).

RC drilling was employed for the Otjikoto deposit evaluation sampling as part of the dataset used for mineral resource estimation. RC sampling flow is shown in Figure 11.1.

RC sample material was routed from the bit to the drill rods’ inner-tube and went via a hose to a cyclone. Samples were collected in 50 kg plastic bags at the cyclone outlet. Each one metre sample was weighed before splitting on site. This allowed for immediate monitoring of sample recovery whilst drilling. If sample recovery dropped below ~85%, drilling would be suspended and equipment checked.

The one metre samples were split in half in a two-step process through a large riffler to achieve homogenization and Left (“L”) and Right (“R”) samples obtained. Each of these samples was again split in half through two smaller rifflers, producing four sub-samples (i.e. L1, L2 and R1, R2). All rifflers are cleaned with compressed air after splitting of each one metre sample.

The L1 and R1 samples are bagged in separate A3 size thick polyurethane bags with identification tickets on the inside and outside of the bags. Average sample mass is approximately 6 kg to 7 kg. Bags containing material from every 5 metres of L1 and R1 samples are packed in separate 50 kg bags and sample numbers marked on the outside. The large bags containing the L1 and R1 samples are transported to the core yard facility.

The L2 sample is dry screened using a 2 mm sieve and the +2 mm sample placed in a clearly labelled 500 ml plastic bottle, which is transported to the core yard for additional detailed geological logging or retained as a reference sample. In the field, the R2 sample is wet screened using a 2 mm sieve and the +2 mm fraction logged for drilling control and geological information.

At various stages of drilling, check bypass material was collected and assayed to compare with the grade of the RC material. No significant discrepancies were noted.

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Figure 11.1:        RC Sampling flow chart

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All further sample processing is carried out in the core yard facility. Splitting and bagging tasks are carried out in batches of ten samples. All the handled samples are checked again for possible labeling errors. The riffling equipment is then cleaned with compressed air between every sample and the work place cleaned after every batch of ten samples.

The L1 and R1 samples from the field are homogenized, composited and split again by feeding both the samples through a large riffler. Two samples with similar masses are produced. These samples are again split in half, one half from each sample is bagged and labeled – this is referred to as the M sample (this material was for metallurgical test work). The M samples are stored in 50 kg bags with each of these bags containing five samples. M samples have a mass of approximately 6 kg to 7 kg each.

The second phase of sample preparation involves the riffling of remaining samples into four subsamples.

Two of these sub-samples were combined to produce the R sample (the reference sample intended for future re-sampling and possible auditing). The remaining two samples were bagged as A and B samples for assay. These samples typically range from 1.5 kg to 1.8 kg. All A samples and every twentieth B sample (full field duplicate) were submitted for assay.

A summary of samples produced from each metre of RC drilling is given below:

  • A Sample: 1.5 kg to 1.8 kg assay sample;
  • B Sample: 1.5 kg to 1.8 kg complete field duplicate of a sample. Every 10th sample submitted for assay;
  • M Sample: 6 kg to 7 kg retained and submitted as required for metallurgical test work; and
  • R Sample: balance of sample material retained for reference material, re-sampling and/or sample audits.

All sampling is supervised by B2Gold personnel both at the drill rig and in the core yard. Only authorized personnel are allowed at the drill sites and in the core yard.

Reference RC chip samples and split rejects are retained in a secure storage facility in Otjiwarongo, Namibia. Additional laboratory coarse rejects are retained in secure storage facilities in Johannesburg, South Africa and Walvis Bay, Namibia.

11.1.2 DRILL CORE SAMPLES

The core is oriented and a low point-line placed on the maximum dip of the prevalent dip of the fabric. A second reference line is also placed down the entire length of the core to ensure that a standard half (the top half) of the core is always sampled. Metre depth marks are placed on the core. The core is then geologically logged. Minimum sample length is 30 cm for HQ and 40 cm for NQ sized core. The majority of the sampling on the project was done at one metre sample intervals with samples labeled according to hole number and depth of end of sample. In 2012, the protocols were revised with the sampling done based on geology and a numeric sample tag system was started with information on each sample marked in the detailed logs and the tag books, in addition to on the core and boxes, as a further check on sampling. Three to five metres of material is sampled above and below the mineralized zones and sampled continuously. In narrow mineralized zones that are separated by more than three metres a gap in the sampling is allowed. Sample start and end points are marked on the core and on the core boxes adjacent to the samples.

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Prior to sampling, high quality photos are taken of the core with a reference header indicating hole number, box numbers and “from - to” depths. All drill cores were systematically photographed dry. In 2012, the procedure was modified to include photographing of core both dry and wet.

The orientated core is split in half along the low point line with a core cutting saw. Fresh water is used at all times while cutting and the diamond saw blades cleaned with a blank medium (such as concrete) on a regular basis to prevent possible contamination. A quarter split of core is done for field duplicates.

Prior to collection of the cut samples and placement into sample bags, the core is checked to confirm the core was placed back in the box correctly after cutting and that the correct sample information is marked on the core, box and sample bag. Samples are checked against the original log information at time of sampling and recorded on sampling sheets. Sample tags with matching information are placed on the inside and outside of each A3 sized thick polyurethane sample bag and the top half of the core placed in the sample bag and the sample bag secured with a cable tie. The remaining (reference) core in the box is reoriented, flat side up, and the sample locations marked on the core, as per the original sample marks, as a permanent record.

Samples are organized into shipments and QA/QC samples are then inserted. All samples are then placed in 50 kg “rice” bags secured with cable ties. All sample shipments are documented and controlled by senior geotechnical staff and the database manager.

Core boxes are stored in secure core storage facilities in Otjiwarongo.

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11.1.3 SPECIFIC GRAVITY AND BULK DENSITY MEASUREMENTS

Over the course of the project a total of 11,243 measurements have been collected of the specific gravity (“SG”) or bulk density of the various rock types of the deposit. The SG data was collected utilizing two different methodologies:

  • Pycnometer laboratory determinations on RC and drill core pulp samples; and
  • Immersion (“Archimedes”) methodology either on whole or half core with 2,633 measurements taken with wax coated and 3,242 on un-waxed.

Initial comparisons between the different methodologies indicated that results were quite similar (Wilton, 2006) so a program of systematic pychnometer determinations was initiated. In 2012, a review of the pychnometer data versus the Archimedes data found that the pychnometer on pulp data over-estimated the SG in the oxide zone by 6 to 12 % and is 2.5% high in the fresh mineralized zone. Consequently, the pycnometer data was not used in the 2012 resource model.

Archimedes SG samples were collected for the complete hole for numerous holes. In 2012, a program of systematic sampling was undertaken whereby representative samples were taken of all lithologies, and at regular intervals (every 25 metres), from a series of holes scattered throughout the deposit. “Composite” samples were also collected, consisting of the measurement of all cores within the 1 to 2 metre mineralized zones.

Given the arid conditions in this part of Namibia, all samples were air dried prior to measurement of the density versus oven dried. Core samples were first weighed dry then coated with several thin layers of paraffin wax. Air bubbles were purged and the core resealed with wax as needed. Samples were allowed to cool for approximately 30 minutes prior to weighing again dry and then within a water bath utilizing a suspended basket balance set up. The SG calculations for each sample were completed via a set of automatic formulas in the EXCEL data entry sheet. Unusual SG sample values were checked to confirm validity.

A QA/QC program for the SG measurements was initiated in 2012 with the use of borosilicate glass cylinders of known SG as reference standards and the monitoring and control of the immersion bath temperature. The bath temperature was kept at approximately 22°C, as per ASTM C97-96 and ASTM C914-95. The standards were measured every ~20 samples. The SG of the standards was plotted regularly as a check on the quality of the overall testing procedures.

11.1.4 SAMPLE INTEGRITY (CHAIN OF CUSTODY AND SECURITY)

Only authorized drill and B2Gold personnel are allowed at the drilling sites.

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All RC samples are collected at the RC rig by B2Gold personnel and transported direct to the B2Gold core yard in Otjiwarongo.

Unloading of the core tube is controlled by the driller and site geologists. Checks are done at site to ensure all core is placed in the boxes correctly prior to shipping. The drill geologist and senior personnel sign-off on the detailed daily drill reports at site and take possession of the core at that time. Core is transported direct to the secure Otjiwarongo core yard by B2Gold personnel. The Otjiwarongo core yard is surrounded by a security fence with the office and complex alarmed and monitored by a local independent security firm.

Sample shipments are controlled by B2Gold Exploration Operations and Database Managers. Transportation to the laboratory is done by an independent bonded courier company (ACT Logistics) with appropriate sign-off documentation accompanying each shipment at both shipping and receiving. Sample shipment damage, if any, is noted by the laboratory upon reception and B2Gold personnel immediately notified. Additionally, the laboratory immediately notifies B2Gold of any discrepancies between sample submittal information, shipment weights and samples received by the laboratory. Any issues are addressed before preparation of the samples start.

All logged and sampled drill core is kept in the core yard or secure storage facilities in Otjiwarongo. Representative core intervals are missing for portions of holes used for metallurgical and geotechnical testing.

11.2 SAMPLE PREPARATION AND ANALYTICAL METHODOLOGY

11.2.1 GOLD FIRE ASSAY

During the initial phase of exploration of the Otjikoto deposit, the presence of coarse particulate gold was noted. In order to resolve the high nugget effect within the sampling results, discussions were held with numerous analytical laboratories regarding the optimal assay procedure to accurately establish the gold content of the sample material. Samples originating from both inside and outside of the mineralized zone were selected and an orientation screen fire assay program at SGS Lakefield (referred to as ‘pulps’ and ‘metallics’) was undertaken. An orientation batch of 128 RC samples was submitted to confirm the applicability of this method. The results of the orientation samples found that 39.4% of the gold occurs in the +106 μm and 60.6% in the -106 μm fraction (Avdale, 2004) and that the one screen fire assay is equivalent to 14 individual fire assays (A.Lombard, personal communication). This confirmed that significant coarse gold is present in the deposit and the screen fire assay methodology provides a more representative analysis of the gold content of the samples. The “metallic” screen methodology was utilized for the remainder of the project.

The original metallic screens methodology used by SGS Lakefield was as follows:

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  • Crush complete sample to -2 mm
  • Pulverize entire sample (3 minutes in LM5 “ring and puck” mill)
  • Screen with 106 μm sieve
  • Accurate weight determined for both fractions (after pulverizing mass)
  • +106 μm assayed (fire assay) to completion (typically 10g to 50g)
  • -106 μm split with a rotary splitter (initially, now with a riffle splitter)
  • -106 μm sub-sample fire assayed (50 g)
  • Resultant gold grades combined using weighted (mass) calculation

Gold grades are determined utilizing the standard fire assay methodology with either an atomic absorption (“AA”) (<10ppm gold) or gravimetric finish (>10 ppm gold). The method codes for each certificate of analysis are preserved in the database.

Sample material submitted to Intertek Genalysis (and ALS Minerals), closely followed this method except that two 25 g aliquots (GEN1 and GEN2) from the -106 μm fraction were combined for each sample to determine the gold grade of the -106 μm fraction.

Au(g/t) = (Au(g/t) +106μm x weight of +106μm) + (wt ave of two Au(g/t) -106μm x weight of 106μm)
                                                            total weight of sample after pulverizing

In 2012, the preparation was changed slightly with an 800 to 1,000 g split taken off of the crush and pulverized in an LM2 (B2000 bowl) in place of the LM5. The principal reasons for the change were to reduce the risk of cross sample contamination and ensure all sample was retrieved from bowl.

In 2012, a separate internal study was completed to look into the viability (operationally) of using a straight fire assay versus the fire screen methodology. Screen fire assays consistently report higher in the 0.4 to 3.5 g/t gold range but fire assays are similar, to slightly higher, in the greater than 3.5 g/t gold range (Figure 11.2) . This has implications for defining ore-waste boundaries around the cut-off range.

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Figure 11.2:        QQ plots of Fire Assay (FA) versus Screen analysis results (gold g/t)

11.2.2 MULTI-ELEMENT ANALYSIS

For a better understanding of the geochemistry of the deposit and deposit area, samples of core, RC and grab samples were sent for multi-element analysis. The most frequently used methods are ICP-MS, ICP-AES and XRF for the determination of precious metal, base metals and multi-element content. Typically the analytical packages include the following elements: Au, Pd, Pt, Ag, Al, Ba, Cd, Co, Cu, K, Mg, Mo, Ni, P, S, Sb, Sn, Te, V, W, Zn and Zr. The analysis was completed at Intertek Genalysis in Perth, Australia, and ALS Minerals, Vancouver, Canada.

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11.2.3 SULPHUR ANALYSIS

To aid in the determination of the total sulphide content of the host lithologies and mineralization in the deposit, sulphur analysis was performed on select groups of samples using either a LECO or similar Carbon and Sulphur (“CS”) analyzer. The pulped sample is weighed out and placed in a ceramic dish or boat. An accelerant is added to act as a flux and improve fluidity and oxidation of the carbon and sulphur. Heating is accomplished in a high frequency induction furnace as this provides both speed and accuracy. Any sulphur or carbon is converted to SO2 or CO2, respectively. These gases absorb infrared radiation at specific wavelengths which is proportional to the concentration of the carbon or sulphur in the sample. Any water in the sample is removed by passing the gases produced through magnesium perchlorate as water interferes with the analysis. Calibration is effected by standards of known carbon or sulphur concentration.

The LECO and CS analyses were completed by Intertek Genalysis in Johannesburg, South Africa and ALS Minerals, Vancouver, Canada.

11.3 ANALYTICAL AND PREPARATION LABORATORIES ACCREDITATION

Various laboratories were utilized for the various stages of the exploration program. All laboratories used were independent of B2Gold Corp. and previous operators.

During the initial regional exploration, geologically identified zones of interest in the drilling core were sampled and sent to at Chemex Laboratories, Toronto, Canada (now ALS Chemex, ISO 9001:2008) and Anglo American Research Laboratories, Johannesburg RSA (ISO/IEC 17025:2005 accredited for Au analysis with Fire assay with an ICP finish through the South African National Accreditation System (“SANAS”) Accreditation number T0051) for analyses.

During definition drilling the following laboratories have been used:

SGS Lakefield Research Africa (Pty) Ltd is accredited with the South African National Accreditation System for “chemical testing” of inter alia “determination of Au by lead fusion followed by Atomic Absorption analysis or gravimetry”. The same certificate states that the laboratory complies with ISO/IEC 17025.

Moruo Analytical Services is accredited with the South African National Accreditation System for “chemical and physical testing” of inter alia “fire assay methods for gold, silver and platinum group metals”. The same certificate states that the laboratory complies with ISO/IEC 17025.

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Intertek Genalysis, Walvis Bay, Namibia (sample preparation), Intertek Genalysis, Johannesburg, South Africa and Perth, Australia, (Primary Analytical) is accredited with ISO 9001: 2000; ISO/IEC 17025 and NATA certification.

The following laboratories are currently being used for the project:

Primary Analytical Laboratory: ALS Minerals, Vancouver, Canada, with ISO 9001:2008 & ISO/IEC 17025:2005; and NATA (Accreditation No. 3244) certification.

Referee Laboratory: Bureau Vertitas, Swakopmund, Namibia (ISO 9001:2008 certified).

Sample Preparation: Intertek Genalysis, Walvis Bay, Namibia

The preparation laboratories and analytical laboratories have been visited and audited on a regular basis by Avdale/Auryx Gold (pre-B2Gold), B2Gold personnel and/or independent auditors (e.g. Du Preez, D., 2011).

11.4 DATA MANAGEMENT

11.4.1 DATABASES

Up until 2012, the Standardized Approach Borehole Log Evaluation (“SABLE”) database management was used to store and verify all digital data. From 2004 to 2011 the management of the database was handled by Dr. Stephen Frindt in direct consultation with the SABLE software developers. Prior to 2004, the database and QA/QC management was handled by an external consulting group.

The SABLE database provided an export format which could be directly exported into Datamine or other resource modelling software.

In 2012, all SABLE data was moved over to a Microsoft ACCESS based B2Gold database. The conversion of the SABLE database to ACCESS was handled by Mike Glover, B2Gold’s

International Database Manager and Liisa Kawali, Otjikoto Exploration Database Manager. The ACCESS database was checked against the SABLE database for errors in conversion. The original SABLE database has been preserved for reference and all original “SABLE” tables are retained in the new database.

A master of the databases is maintained by the database manager. The database is copied and backed up on the B2Gold servers in Otjiwarongo and Vancouver, Canada on a regular basis.

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As with SABLE, the ACCESS database is formatted to be compatible with the needs of all end users, including the resource modeller’s.

For the DFS mineral resource estimate, the primary Drillhole database was issued on May 31, 2012 (Otjikoto DDH DB May 31 2012 LK V2.accdb). Certain key tables including assays, lithology and collars were updated with the June 6, 2012 version (Otjikoto DDH DB Jun 06 2012 LK.accdb). As there are no trenches or surface sampling material to the resource estimation, the surface database was not included in the DFS.

11.4.2 DATA ENTRY AND DATA MANAGEMENT

Prior to 2012, all logged data was captured on paper and entered by the geologist and/or the database personnel into the SABLE “Sputnik” remote data entry module. Once entered into Sputnik, the field data was checked and verified by the logging geologist and signed-off on prior to entry into the database by the database manager. Within SABLE a series of checks/data verifications are performed which examines the data for issues such as overlapping samples. A report is produced by SABLE and any errors are identified, addressed and corrected in the original data and the final database.

As part of the previous Resource modeling exercises, all the SABLE drill data was validated and any errors corrected. The following independent audits were conducted of the Otjikoto Project and related SABLE database in conjunction with the Resource Modeling:

  • In 2005, RSG prepared an Independent Technical report on the Otjikoto Project. (van der Merwe and Jones, 2005);
  • In 2007, SRK completed an independent review of the Mineral Resource at the Otjikoto Project; (van der Merwe and Wanless, 2007; and Wanless et al 2007);
  • In 2009, SRK reviewed the Mineral Resource estimate at the Otjikoto Project (Wanless and Crisp, 2009); and
  • In 2011, BMRE was contracted to evaluate the drill hole data for Auryx Gold and provide an updated resource model. (Nicholls, 2011 and 2012).

In 2012, all logged data was entered into EXCEL templates by data entry personnel. Inputs in various EXCEL columns are locked in a “data validation” format which only allows correct coding to be entered within the EXCEL spreadsheet. The entered data is then printed out and a 100% check is done against the original logged data or other original information (e.g. downhole survey data). A further check is completed by the logging geologist prior to the data being entered into the ACCESS database by the database manager.

Downhole survey data is received as written pieces of paper from the drill foremen and entered by hand into an EXCEL file and checked in a similar fashion as above. As a further check the data is occasionally checked against the actual instrument data in the field and downloaded from the instrument (Reflex Ez-shot) and checked against the written information. Obvious errors in drill hole downhole survey data are flagged as do-not-use, “N”, in the database. As a further check a plot is completed of drill holes and any sudden jogs in the drill holes are trace reviewed for validity.

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Collar survey data received from the professional surveyors was loaded directly into the database and the original copies retained and kept with each drill hole file. Temporary field GPS drill collar coordinate data is removed from the database and replaced with the final survey data.

Validation queries were run within the database to ensure the integrity of the database.

Prior to 2012, all assay data was received from the laboratories in csv files and directly imported into the SABLE database. In 2012, csv files received from the laboratories required a reformat prior to importation into ACCESS. The reformatting was completed by a macro(s), reducing the chance of human error. The original analytical information is not tampered with and the raw data imported into the database. The original and digital assay certificates are retained in separate files for reference and for verification against the database information. As a quick review for obvious errors, a visual check of each Certificate of Analysis (“COA”) is completed prior to import into the database.

In all cases of data falling below the detection limit the values are reported as half the detection limit (e.g. for gold <0.01 is reported as 0.005) .

11.4.3 HARDCOPY DATA ORGANIZATION AND FILING

All original data is stored in a fire resistant storage room in the Otjiwarongo Exploration Office. All data is very well organized with all original documentation keep in individual drill hole files, which include all logged information, original assay certificates and survey information.

11.5 QUALITY ASSURANCE AND QUALITY CONTROL

QA/QC procedures have been in place since the start of the project and are clearly documented in the procedure manuals. The procedure manuals have been updated on a regular basis. McDonald et. al. (2011) provides a detailed report of the historic QA/QC procedures used on the project to 2011.

During the life of the project, the following external (geological) controls samples have been routinely inserted:

  • BLANKS – for monitoring of contamination and sample mix ups.

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  • CERTIFIED REFERENCE MATERIAL (CRM’s) – to monitor the accuracy of the laboratory.
  • DUPLICATES – To monitor laboratory precision.

In addition to the geological QA/QC samples inserted and evaluated during the course of the project, the individual laboratories provide their internal QA/QC information with each COA and, in the case of Genalysis, also as a laboratory quarterly summary QA/QC report.

Each COA is vetted by the database manager in combination with SABLE (up until 2011) and ACCESS (2012). The result of the 2011 and 2012 vetting of all COA’s is documented in a table of failures (“TOF”). The TOF shows the results of the QA/QC, any follow up action and final decision to include or not include COA in the resource database. All COA’s are retained in the database but only COA’s which pass the vetting process are included in the resource. Laboratories are requested to review and/or re-run submittals until the submittal passes QA/QC.

Monthly QA/QC reports are prepared documenting the laboratory performance.

11.5.1 BLANKS

Dolomite 50 mm aggregate from a local source in Otjiwarongo has been used as the blank material for the project. Repeat assaying has shown that this material has no significant values and is a good blank medium. Blanks are inserted at a frequency of 1:20 samples. The failure threshold is deemed to be 10 x the detection limit, i.e. >100 ppb. Plots of the 2011 and 2012 blanks indicate that the Genalysis laboratory is doing a good job of minimizing sample contamination (Figure 11.3) . Initial blank failures mostly correlate with blank “wash” samples inserted after samples with visible gold to minimize the potential carry-over of gold into the next samples.

Figure 11.3:        Genalysis Blank Performance 2011 and 2012

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11.5.2 CERTIFIED REFERENCE MATERIALS (ANALYTICAL STANDARDS)

A group of “low”, “medium” and “high” gold grade CRM Standards has been used on the project to monitor the accuracy of the laboratory analysis. The standards were purchased from OREAS, Australia, and CDN Resource Laboratories, Canada. The CRM insertion frequency into the sample stream was based on the fusion oven charge: for Genalysis 1:20; and for ALS Minerals 1:38 (½ charge size of 78; i.e. two CRM’s per fusion batch).

Within the SABLE QA/QC vetting query, the failure of CRM’s was based on + 20% of the expected mean value (EV) of the CRM and warning at 10% of EV. In 2012, the limits were set to B2Gold QA/QC standard of the greater of +10% or + 3 Standard Deviations (3SD) of CRM mean. Warning limits were set at +2SD, with failure based on two or three consecutive 2SD CRM’s warnings falling on the same side of the warning limit, suggesting a laboratory bias may be present.

Examples of CRM plots are presented in Figures 11.4 to 11.6.

Figure 11.4:        Example of CRM charting (2012) “Low Grade” CRM OREAS 15g

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Figure 11.5:        Example of CRM charting “High” Grade CRM’s OREAS 17c

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Figure 11.6:        Example of CRM charting Low Grade CRM’s

11.5.3 DUPLICATES

Precision of the various laboratories was monitored by the use of field duplicate samples. Field duplicate samples were collected at a frequency of 1:20. Starting in 2012, in addition to the field duplicates, preparation and pulp duplicates were requested to ensure precision was monitored after all phases of the preparation. Duplicate results are presented in a series of charts starting with simple scatter plots and ending up with Thompson Howarth precision charts (Figures 11.7 and 11.8) . The overall precision of the various laboratories is poor, about 20% for pulp duplicates. This is due to the high nugget component of the gold mineralization. The main potential impact of the poor precision is in grade control (ore-waste boundary) around the cut-off grade.

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Figure 11.7:        Scatter plots of various duplicates

Figure 11.8:        Thompson Howarth plot of Otjikoto duplicates

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11.5.4 EXTERNAL CHECK LABORATORIES

Pulp samples were collected and sent to an umpire/referee laboratory on a regular (quarterly) basis for routine checks of the quality of the primary laboratory. Given the nuggety nature of the mineralization, the check assay results with the umpire laboratory correlate reasonably well with the primary laboratory.

As an additional check on historic analysis, 14% of pre-2012 RC and drill core samples were sent for check analysis to ALS Minerals. These samples were sent in representative grade bins rather than at random. A scatter plot of the comparison is presented in Figure 11.9. There is no apparent bias with the historic data.

Figure 11.9:        Check sample analyses of historic samples

11.6 COMMENTS ON SECTION 11

In the opinion of the Qualified Person, Mr. Tom Garagan P.Geo, sample preparation, security and analytical procedures are within industry and best practice for the following reasons:

  • Sampling follows set company protocols and is compliant with industry best practice standards;
  • Chain of custody and security is adequate to ensure integrity of samples;
  • Sample preparation and analytical procedures address the problem of nuggety gold and meets or exceed industry standards;

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  • Quality control systems have been in place since the start of the project to ensure the data included in the resource estimation has been verified; and
  • Table of failures documents QA/QC actions taken to monitor lab and ensure validity of data included in resource database.

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12 DATA VERIFICATION

Tom Garagan, P. Geo of B2Gold Corp., acted as the Qualified Person for the data verification and mineral resources reported in this Report. As a result of the data verification summarized in this section, Mr. Garagan concludes that the database used to create the mineral resources and mineral reserves is reliable and valid.

12.1 HISTORIC DATA VERIFICATION PROCEDURES

As indicated in Section 11.4.2, a series of independent companies have reviewed and audited the data, procedures and historic resources for the property and provided documentation to that effect in NI 43-101 Technical Reports. The most recent review by Bloy Resource Evaluation (Nicolls, 2012) included checking of collar co-ordinates, down hole survey data, geological and mineralization logging, and assay and density data. Minor errors were noted and corrected, and the dataset deemed suitable for use in Mineral Resource estimation.

The historic QA/QC program, while providing for vetting of COA, used failure limits higher than normal practice. Referee laboratory checking of a large number of the historic data indicates that the original data is within acceptable levels given the overall precision of the deposit and is therefore suitable for inclusion in the resource estimation.

12.2 DATA VERIFICATION PROCEDURES

12.2.1 SITE VISITS

Visits to the Otjikoto Project site and Otjiwarongo core logging and office facility were made by Mr. Garagan on the following dates:

  • January 7th to 12th, 2012
  • February 3rd to 5th, 2012
  • August 21st to 24th, 2012
  • November 20th to 22nd, 2012

During the series of visits, Mr. Garagan verified the following:

  • RC drilling and sampling procedures at the rig during drilling;
  • Diamond drilling at various drills and the core retrieval and handling procedures;
  • A comparison of DDH versus RC drilling results and a review of assays from twinned RC and DDH holes;
  • RC sample splitting procedures;
  • Core metre and low line marking and geotechnical assessment procedures;
  • Core logging procedures, protocols and geological control;

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  • Core photography procedures and quality;
  • Core cutting and sampling procedures;
  • Core storage and security;
  • SG measurement and SG QA/QC procedures;
  • Sample shipping and chain of custody procedures;
  • Data entry and data verification procedures;
  • Data filing and organization1;
  • Database management procedures; and
  • Geological interpretations accuracy.

Laboratory performance was reviewed by Mr. Garagan through examination of monthly QA/QC reports prepared by the Database Manager, Liisa Kawali, and B2Gold’s International Database Manager, Mike Glover. These reports provide:

  • Documentation of the vetting of every COA and actions taken (via TOFs);
  • Tracking of the laboratory performance, through charting of the blanks, certified reference material (standards) and field, preparation and pulp duplicates; and
  • Verification of primary laboratory quality (biases) through comparison of external referee data.

QA/QC for the Otjikoto Project is overseen on site by the Exploration Manager, Hugh MacKinnon, P.Geo.

12.2.2 RESOURCE ESTIMATION REVIEW

Work of the resource modeller was verified by review of the following:

  • Database verification methodology;
  • Geologic interpretation of the Otjikoto model wireframes including oxide surfaces, lithology model, mineralization wireframes on section and level;
  • Geostatistical procedures followed during exploratory data analysis to determine capping levels, composite lengths and geologic model tagging;
  • Comparison of grade in drill holes and adjacent blocks in model;
  • Comparison of final block model resource with previous resource models; and
  • Comparison of final grade estimation model with different techniques of estimation.

____________________________________
1
While the author did not check every file, the author verified that the hardcopy data files contain the original information (collar survey; geological log; geotechnical log; core box inventory; down hole survey information; contractor drill reports; and, original assay certificates) for each drill hole plus the support of the verification of data within the database

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12.3 LIMITATIONS

Topographic surveys for the Otjikoto Project are derived from the 2008 LIDAR airborne survey and as such were not verified by Mr. Garagan; however the collar data as presented in the database closely matches the LIDAR digital elevation model.

12.4 COMMENTS

Mr. Garagan for Section 12 has reviewed:

  • The geological data for the project and believes that the data represents the geology and mineralization of the deposit;
  • The sampling density (RC and DDH) and it reflects the distribution of mineralization in the deposit within the confidence levels indicated in this report;
  • Drill collar and downhole survey information and it accurately represents the location of the data collected for the resource estimation;
  • The current sampling and analytical procedures, and related QA/QC program and believe these meet or exceed industry best practice standards;
  • SG determination procedures and data and believes they provide an accurate estimation of the SG of the mineralization and host rocks sufficient for inclusion in the resource estimation and mine planning;
  • The data in the database and believes it reflects the original geological and analytical data;
  • The resource model and believes it provides an accurate estimation of the resource of the Otjikoto deposit and is suitable for purposes of mine planning.

The Qualified Person responsible for this Section, Mr. Tom Garagan, P. Geo, believes that a better understanding of the geological controls on the Otjikoto mineralization will lead to improved resource estimation.

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13 MINERAL PROCESSING AND METALLURGICAL TESTING

Feasibility metallurgical testing was performed on Otjikoto drill core samples representing the major ore types to finalize the gravity/whole ore leach process design criteria and flowsheet, and to evaluate ore variability across the deposit. The gravity/whole ore leach process was selected as the preferred mill flowsheet over gravity/flotation/concentrate leach in a trade-off study(SMP, Mining Licence Application) prior to the start of the December 2012 Otjikoto Gold Definitive Feasibility Study.

The following metallurgical test programmes were completed:

  • Characterization of chemical and mineralogical composition of major ore types.
  • Characterization of materials flow properties to determine the stockpile and ore reclaim system design parameters.
  • Leach and gravity concentration test work on XR1 (oxide), XR2 (transition) and XR3 (deeper sulfide) composites of the three major ore types to establish the optimum grind size and produce samples for downstream testing.
  • Grinding test work and modelling/simulation to determine power requirements and sizing for the SAG and ball mills and evaluate variability of ore grindability.
  • Leach optimization testing to establish optimum leach process parameters for the XR1, XR2 and XR3 ore types.
  • Evaluation of ore variability with optimum leach conditions for recovery modelling and determination of reagent consumptions.
  • Thickener and rheological testing to establish final pre-leach and tailings thickener sizing.
  • SO2/Air cyanide destruction test work to establish design criteria and reagent consumptions for the cyanide destruction circuit.
  • Carbon adsorption test work and CIP modelling to confirm the CIP circuit design parameters.
  • Sampling test work to determine the design requirements for the mill sampling stations and assay laboratory.
  • Tailings characterization testing.

13.1 GRINDING AND METALLURGICAL RECOVERY TEST WORK AND SCOPE

The metallurgical scope of the feasibility study was extensive and designed to meet the objectives of completing all test work required to finalize the process design criteria and the mill flowsheet, size mechanical equipment and determine plant capital and operating costs. Table 13.1 presents a summary of the test work scope including the metallurgical laboratory used, samples tested and brief description of the scope of work for each specific test program. A flow chart of the feasibility grinding and metallurgical test programs is provided in Figure 13.1.

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Table 13.1:        Summary of Metallurgical Test work Scope

Test Work Description Laboratory Samples Scope
Materials Handling Test work




Jenike & Johanson




3 x Composites:
1 x XR1
2 x XR2
3 x XR3

Particle density determination
Compressibility tests
Loose and compacted bulk density tests
Flow function tests
Wall friction determination
Critical chute angle determiatnion
Comminution Test Work







SGS Lakefield







3 x Composites:
1 x XR1
1 x XR2
1 x XR3

17 x variability Samples:
3 x XR1
5 x XR2
9 x XR3
JK drop-weight test
Bond low-energy impact test
Bond rod mill grindability test
Bond ball mill grindability test
Bond abrasion test
SMC test
Bond rod mill grindability test
Bond ball mill grindability test
Bond abrasion test
UCS Test Work

Rocklab

20 x Variability Samples:
10 x XR2
10 x XR3
Unconfined (Uni-axial) compressive strength with
elastic modulus ( UCM)
Chemical Composition and
Mineralogy

SGS Lakefield


3 x Composites:
1 x XR1
1 x XR2
1 x XR3
Comprehensive Chemical Analysis and Assay Work
Gold Deportment Study
QEMSCAN Study
Leach and Gravity Test Work to
Determine Optimal Grind and
Produce Samples for
Downstream Testing
SGS Lakefield 3 x Composites:
1 x XR1 - 60kg
1 x XR2 -60kg
1 x XR3 -60kg
Sample Preparation and Head Analysis
Gravity Concentration
Intensive Cyanidation of Gravity Concentrate
Direct Cyanidation on gravity tails at three grind sizes: 100 mesh, 150 Mesh and200 Mesh.
CIP/CIL Tests on gravity tails at the optimum grind size
Bulk Cyanidation on bulk composites at the optimum grind size to produce feed material for cyanide destruction test work
Thickener and Tailings Characterization Testing
Leach Optimization Test Work CANMET 3 x Gravity tailings samples:
1 x XR1 -10kg
1 x XR2 - 10kg
1 x XR3 -20kg
Leach Optimization test work on XR3 to evalute the effect of each of the following parameters on overall recovery and leach kinetics:
The inclusion of a pre-treatment stage ( 0.5hr ,4hr, 7hr,8hr)
Lead nitrate addition (50g/t, 100g/t, 250g/t)
Cyanide Concentration (200ppm,300ppm,400ppm,500ppm)
Pretreatment pH ( 8, 10.5)
Oxygen addition instead of air to acheive a DO >16ppm

Leach Optimization test work on XR1 and XR2 based on the optimum leach conditions for XR3 to evalute the effect of each of the following parameters on overall recovery and leach kinetics:
Cyanide Concentration (200ppm,300ppm,400ppm,500ppm)
Leach Variability Test Work SGS Lakefield 46 variability Samples:
8 x XR1
15 x XR2
3 x XR4
20 x XR3
Sample Preparation and Head Analysis
Gravity Concentration
Intensive Cyanidation of Gravity Concentrate
Direct Cyanidation on gravity tails with optimum leach conditions
Carbon Adsorption test Work and Modelling SGS Lakefield 3 gravity tailings samples
1 x XR1
1 x XR2
1 x XR3
Leach and carbon adsorption kinetic tests
Determiantion of carbon loading isotherms
Develop a model to simualte the performance of a full-scale CIP circuit

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Test Work Description Laboratory Samples                                                Scope
Cyanide Destruction Test Work



SGS Lakefield



3 x Bulk cyanidation,
cyanide leach pulp samples
1 x XR1
1 x XR2
1 x XR3
Batch SO2 /Air caynide destruction tests
Continuous SO2 /Air caynide destruction tests


Gravity Concentration and
Intensive Leach Test work

FLS-Knelson


3 x Gravity tailings samples:
1 x XR1 -50kg
1 x XR2 - 50kg
1 x XR3 -50kg
E-GRG test work
Intensive cyanidation of gravity concentrates

Sedimentation and Rheological
Test work






FLS - Salt Lake City
conducted at
SGS Lakefield
Ontario




3 x Gravity tailings samples:
1 x XR1 -5kg
1 x XR2 - 5kg
1 x XR3 -5kg

3 x CND tailings samples:
1 x XR1 -10kg
1 x XR2 - 10kg
1 x XR3 -10kg
Bench scale sedimentation tests that included
the following:

Flocculant screening,
Determination of optimum feed solids dilution
Settling tests
Thickener underflow rheology measurements

Sampling Test Work



SGS Lakefield



1200 kg of XR3



Sample Preparation and Head Analysis
Gravity Concentration
Intensive Cyanidation of Gravity Concentrate
Bulk Cyanidation on gravity tails
Overall Gold Balance
Environmental and Geotechnical
Testing












SGS Lakefield
Amec












XR1, XR2 and XR3 CND
Tailings Samples












Comprehensive solution and solids analyses
Modified acid base accounting
Net acid generation testing
Deionised water leach
TCLP Testing
Whole rock analyses
Specific Gravity determination
Determination of particle size distribution
Settling Tests
Drained Settling Tests
Standard Proctor Tests
Determination of Atterberg Limits
Air Drying Tests
Rowe cell consolidation with hydraulic conductivity (Amec)
Triaxial Tests (Amec)

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Figure 13.1:        Metallurgical Test Work Programme Flow Chart

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13.2 TEST SAMPLES

The feasibility metallurgical sampling program was designed by B2Gold-Auryx exploration and B2Gold Corporate to provide sufficient material for the following principal test programs:

  1.

Grinding Test Work

  2.

Metallurgical Recovery Test Work

  3.

Sampling Test Work - “Pitard” Test Sample

The Grinding and metallurgical test work programmes were performed on both composite and variability samples. The composite samples represented the three major ore types found in the Otjikoto deposit designated as XR1, XR2 and XR3 and were used to evaluate the grade/recovery relationship and grindability for each particular zone. The ore variability samples represented various spatial locations and zone distribution throughout the Otjikoto deposit and were also categorized as XR1, XR2 or XR3.

The sample selection was based on three principal styles of mineralization (ore types) present in the deposit:

  • XR1 – Oxide
  • XR2 – Pyrite (+ marcasite) dominant mineralization
  • XR3 – Pyrrhotite dominant mineralization

A fourth ore type was identified during the course of the feasibility metallurgical investigations and was included in the recovery variability testing. The fourth ore type has been designated as XR4 and is a mixture of XR2 and XR3 sulphide mineralization.

Minimum composite sample weight and/or sample number requirements were defined and these requirements are detailed in Table 13.2 below:

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Table 13.2:        Metallurgical Testing Sample Requirements

In general, the Otjikoto deposit mineralization is distributed as described below and indicated in the plan map of Figure 13.2:

  • Strong to transitional oxidation below the calcrete (hardpan) to 30 to 40 metres below surface with local deeper “wells” of oxidation associated with fault and or fracture zones.
  • Pyrite dominant mineralization in the northern and shallower portions of the deposit
  • Pyrrhotite dominant mineralization down plunge in both the Central (Main) and Western shoots but more prevalent in the Western shoot.
  • More magnetite present in the southern portion of the deposit, in particular upper levels of southern part of the deposit
  • Mixture of pyrite and pyrrhotite in various proportions.

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Figure 13.2:        Plan Map of sulphide species in the Otjikoto deposit

Thirty-eight new metallurgical test sample drill holes totalling 2,663.69 metres were drilled in 2012 and an additional 27 “historic” holes were sampled to provide material for the feasibility test programs. Some of the “historic” holes represent holes drilled in 2011, or in previous programs, but not previously sampled. For the feasibility testing a total of 1,256 samples weighing a total of 8,029.5 kilograms were shipped to SGS Lakefield in Ontario, Canada (“SGS Lakefield”), for the various test programs.

Holes were selected for drilling and/or sampling based on the composition and grade of the original reverse circulation (“RC”) hole or diamond drill hole. Each historic hole was twinned, stepping 2 to 4 metres away from the original collar location.

New holes were drilled for all the XR1, oxide mineralization samples. When historic holes were sampled, locations with limited to no oxidation were used. In general the level of oxidation of old core is low with surface oxidation “tarnish” common but only a few marcasite bearing samples more strongly weathered, in part to iron sulphates. Only new core was used for the comminution and grinding variability studies.

For the Metallurgical Recovery and Comminution test programmes samples were composited from different geographic areas to provide a “bulk” sample with grades and geological characteristics representative of the main mineralization domains. Metallurgical Recovery Variability samples were selected to be representative of the principal ore zones (Central Shoot (Main), West Shoot, HW and FW), a variety of vein mineralogy’s, various intensities of veining, different host rock lithology, a range of spatial distribution, a range of gold grades and different depth levels within the deposit. Grinding Variability test samples were selected similar to above but with emphasis on veining intensity and difference in host rock lithologies. Drill-hole and related sample locations were selected prior to the completion of the current resource model wireframes.

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Given the sheeted nature of the vein system no specific dilution was added to the samples in either the footwall (“FW”) or hanging wall (“HW”) with consideration given more to average vein intensity and thus “average” host rock “dilution”. In order to create the final composites lower grade samples were on occasion incorporated, sometimes in hangingwall or footwall, to bring the average grade of the composite sample down into the selected range. Due to the nuggety nature of the mineralization, particularly in the pyrrhotite dominant mineralization, XR3, overall assay precision was low. An effort was made to composite continuous samples within an individual drill hole to be more representative of the average characteristics of a zone. For the Metallurgical Recovery and Metallurgical Recovery Variability the final composites were based on the assay information provided by SGS Lakefield.

The Pitard sample was created from a mix of old and new drill holes with a combination of known and unknown (metallurgical holes) grade. Where the grade was not known a ¼ or ½ split of the core was provided to SGS Lakefield in order to complete a screen fire assay of the sample. Drill holes were in part selected with known presence of visible gold in order to assist in assessing the sampling protocols for nugget gold. The Pitard sample was composed principally of XR3 but with some XR2 and mixed vein material so as to be more representative of the overall deposit.

At the request of Paul Morgan of DRA ten additional samples each of XR2 and XR3 mineralization were selected and sent for unconfined compressive strength (“UCS”) testing at Rocklab in Pretoria, South Africa. Samples were selected with a range of vein intensity, composition and host lithology. The average length of each sample was approximately 20 centimetres.

Further details of the metallurgical sampling programmes are provided in Section 7.2.2 Test Samples and Appendix 7-7-20 Otijikoto Metallurgical Sampling Report in the December 2012 Otjikoto Gold Definitive Feasibility Study Report.

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13.3 TEST WORK RESULTS

13.3.1 ORE CHEMICAL ANALYSIS AND MINERALOGY

Chemical analysis and mineralogical studies were performed at SGS Lakefield on the XR1, XR2 and XR3 Metallurgical Recovery Composite samples.

The following chemical analyses were performed:

  • Whole rock analysis in the mineralogical study.
  • Chemical analysis of Metallurgical Recovery Composites, which included gold assay by the screened metallics protocol.

The mineralogical study conducted had the following objectives:

  1.

Determine the bulk mineralogy of each composite sample.

  2.

Determine the gold and silver deportment of each composite, including mineral speciation, grain size, liberation and association.

The detailed results of the chemical analysis and mineralogical examinations complete with photomicrographs are presented in the Mineralogy Test Work Report in Appendix 7-7-3 and the SGS Lakefield Final Test Work Report in Appendix 7-7 in the December 2012 Otjikoto Gold Definitive Feasibility Study.

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13.3.1.1 Chemical Analysis

Whole Rock Analysis

The whole rock assay results are presented in Table 13.3 below.

Table 13.3:        Whole Rock Analysis and Head Assays

Sample ID XR1 XR2 XR3
  (%) (%) (%)
SiO2 50.7 50.0 51.4
Al2 O3 10.8 10.5 10.4
Fe2 O3 10.7 14.2 14.5
MgO 0.721 2.350 2.390
CaO 8.9 4.4 5.2
K2 O 0.115 0.085 0.058
TiO2 0.647 0.629 0.628
MnO 0.137 0.140 0.154
Cr2 O3 0.018 0.026 0.027
V2 O5 0.029 0.040 0.035
Na2 O 4.6 5.7 6.0
P2 O5 0.130 0.140 0.140
As <0.001 <0.001 <0.001
Fe 7.9 10.3 10.2
S= <0.05 2.74 3.08

All 3 Metallurgical Recovery Composites had similar whole rock compositions with the most significant difference occurring in sulfide content.

Chemical Analysis of Metallurgical Recovery Composites

The head analysis of the metallurgical recovery composites included gold assay by screened metallics protocol, Inductively Coupled Plasma (“ICP”)-Scan, Au and Ag by direct fire assay and Pulp Fusion Atomic Absorption Spectroscopy for Fe, Cu, Co, As, Sb, Cl, Si, Mg, Bi, Cd, Hg, F, Pb, Zn and Ni, sulphur speciation and carbon speciation by various methods as shown in Table 13.4 and Table 13.5 below. Table 13.4 shows that the repeatability of the gold assay determination by direct fire assay is poor for XR1 and XR2. This variation in head grade is an indication of presence of coarse nugget gold in the Otjikoto deposit. As indicated in Tables 13.4 and 13.5, the base metal assays for XR1, XR2 and XR3 are low, the copper assays range from 0.012% – 0.016% (Table 13.4) . Organic carbon, Arsenic, Antimony and Bismuth levels are very low concentrations and at these level would not present any metallurgical processing issues. The mercury levels were <0.3 g/t for all ore types.

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Table 13.4:        Head Analysis of XR1, XR2 and XR3 Metallurgical Composites

Head Analysis Method XR1
Composite
XR2
Composite
XR3
Composite
Au g/t Cut A Fire Assay and FAAS 2.77 0.84 1.11
Au g/t Cut B Fire Assay and FAAS 0.58 0.47 1.08
Ag g/t Cut A FAAS < 0.5 < 0.5 < 0.5
Ag g/t Cut B FAAS < 0.5 < 0.5 < 0.5
Fe % XRF 8.14 10.3 10.1
Cu % XRF 0.016 0.015 0.012
Co % XRF < 0.01 < 0.01 < 0.01
As % XRF < 0.001 < 0.001 < 0.001
As % re-assay2 XRF < 0.001 < 0.001 < 0.001
As g/t re-assay3 Hydride generation AA 4 1 --
As g/t re-assay4 Hydride generation AA 4 < 1 --
Sb % XRF < 0.002 < 0.002 < 0.002
Cl g/t XRF 40 60 90
Si % XRF 23.5 23.8 23.4
Mg % XRF 0.48 1.53 1.52
Al % XRF 5.89 5.94 5.83
Bi % FAAS < 0.002 < 0.002 < 0.002
Cd % FAAS < 0.0005 < 0.0005 < 0.0005
Hg g/t Cold Vapor AAS < 0.3 < 0.3 < 0.3
Hg g/t re-assay2 Cold Vapor AAS < 0.3 < 0.3 --
Hg g/t re-assay3 Cold Vapor AAS < 0.3 < 0.3 --
F % Ion Selecticve Electrode 0.059 0.045 0.035
Pb % FAAS < 0.002 < 0.002 < 0.002
Zn % FAAS 0.002 0.002 0.001
Ni % FAAS 0.003 0.004 0.004
C(t) % Leco 2.35 2.41 2.08
C(g) % Leco 0.04 0.05 0.03
TOC leco % Leco < 0.05 0.22 0.12
CO3 % Leco 10.4 10.8 9.34
S % Leco 0.04 2.98 4.05
S % re-assay Leco 0.04 2.88 --
S= % Leco < 0.05 2.77 3.86
SO4 % Leco < 0.1 < 0.1 < 0.1
S° % Leco < 0.05 < 0.05 < 0.05

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Table 13.5:        Elements ICP-Scan XR1, XR2 and XR3 Metallurgical Composites

Head Analysis Method XR1
Composite
XR2
Composite
XR3
Composite
Ag g/t ICP < 2 < 2 < 2
Al g/t ICP 7040 53400 59500
As g/t ICP < 30 < 30 < 30
Ba g/t ICP 20.3 13.2 8.7
Be g/t ICP 0.14 0.4 0.42
Bi g/t ICP < 20 < 20 < 20
Ca g/t ICP 17300 30500 43200
Cd g/t ICP < 2 < 2 < 2
Co g/t ICP 31 24 30
Cr g/t ICP 53 111 87
Cu g/t ICP 128 41.2 114
Fe g/t ICP 266000 102000 112000
K g/t ICP 862 808 627
Li g/t ICP < 9 < 9 < 5
Mg g/t ICP 14600 13700 15200
Mn g/t ICP 7860 1060 1260
Mo g/t ICP < 5 < 5 < 5
Na g/t ICP 320 33600 47700
Ni g/t ICP 37 41 45
P g/t ICP 203 586 648
Pb g/t ICP < 20 < 20 < 20
Sb g/t ICP < 10 < 10 < 10
Se g/t ICP < 30 < 30 < 30
Sn g/t ICP < 20 < 20 < 20
Sr g/t ICP 18.3 45.6 58.4
Ti g/t ICP 291 3480 3940
Tl g/t ICP < 30 < 30 < 30
U g/t ICP < 30 < 30 < 20
V g/t ICP 15 207 216
Y g/t ICP 2.3 16.7 21.3
Zn g/t ICP 216 < 40 21
Te g/t ICP < 50 < 50 < 50
As g/t Hydride
generation
ICP
< 30 < 30 < 30

Three samples from each composite XR1, XR2 and XR3 were rotary split out and sent for screen metallics assays. The results are shown in Table 13.6. This table shows that the repeatability for the 1 kg screen metallics assays was poor. The variation in head grade is an indication of the presence of coarse, nugget gold in the samples. The comparison of the screen metallics gold assays (Table 13.6) and direct fire assay (Table 13.4) showed significant differences in the results. This difficulty in measuring the gold assays was a major reason for performing the Pitard sampling test to establish the sampling and assay protocols for the mine, mill and assay laboratory operations.

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Table 13.6:        Screen Metallics Assay Results for the Metallurgical Recovery Composites

Sample ID Head
Grade
Au, g/t
+150 Mesh -150 Mesh % Au Distribution
%
Mass
Au, g/t
%
Mass
Au, g/t
+150
Mesh
-150
Mesh
a b
XR1 Comopsite - sample A 1.84 2.74 28.46 97.3 0.75 1.43 42.4 57.6
XR1 Comopsite - sample B 1.04 2.87 13.54 97.1 0.65 0.70 37.2 62.8
XR1 Comopsite - sample C 0.95 2.93 6.04 97.1 0.63 0.97 18.6 81.4
XR2 Comopsite - sample A 1.02 1.98 11.36 98.0 0.95 0.68 21.9 78.1
XR2 Comopsite - sample B 3.43 2.74 94.70 97.3 0.50 1.22 75.6 24.4
XR2 Comopsite - sample C 1.14 2.95 12.14 97.1 0.74 0.87 31.4 68.6
XR3 Comopsite - sample A 5.85 2.65 172.00 97.4 0.83 1.84 77.8 22.2
XR3 Comopsite - sample B 2.08 2.56 53.51 97.4 0.87 0.59 65.8 34.2
XR3 Comopsite - sample C 4.63 2.57 140.00 97.4 0.72 1.39 77.8 22.2

13.3.1.2 Bulk Mineralogy

The XR1, XR2 and XR3 metallurgical recovery composites were subjected to the following analysis to determine bulk mineralogy:

  • QEMSCANTM analysis
  • Analysis by X-Ray Diffraction (“XRD”)

13.3.1.3 QEMSCANTM analysis

XR1 was composed of major amounts of plagioclase (41.9%), moderate amounts of quartz (19.9%), calcite (13.3%) and Fe-oxides(13.2%) and a minor amount of chlorites (4.9%) and other silicates (2.6%) .

XR2 was composed of major amounts of plagioclase (45.3%), moderate amounts of quartz (15.1%), ankerite (11.0%) and Fe-oxides (10.3%) and a minor amount of pyrite (5.6%), other silicates (3.7%), chlorites (3.5%) and calcite (2.1%) .

XR3 was composed of major amounts of plagioclase (49.2%), moderate amounts of ankerite (14.4%) quartz (14.2%) and pyrrhotite (10.8%) and a minor amount of calcite (2.5%) and chlorites (2.5%) .

The mineral compositions as by QEMSCAN were confirmed by the XRD analyses,

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A.        Gold and Silver Deportment Studies

The XR1, XR2 and XR3 Metallurgical Recovery Composites were evaluated in gold deportment studies following heavy liquid separation (“HLS”) and further concentration by super-panning.

13.3.1.4 Gold Distribution

The overall gold distribution for sample XR 1 was the following:

(1)        The liberated, exposed and locked gold minerals account for 84.3%, 0.4% and 15.4% of total gold, respectively;

(2)        Heavy liquid products:

 

84.7% of Au was recovered in the HLS 2.8 Sink fraction, including 84.3% liberated and 0.4% exposed.

 

15.4% of Au remained in the HLS 2.8 Float fraction all of which was locked.

The overall gold and silver distribution for sample XR2 was the following:

(1)        The liberated and locked gold and silver minerals account for 84.3% and 16.6% of the total gold and silver, respectively;

(2)        Heavy liquid products:

 

97.2% of Au and Ag was recovered in the HLS 3.1 Sink fraction, including 80.5% liberated and 16.6% locked.

 

One grain, representing 2.9% of the total Au, remained liberated in the HLS Float fraction.

The overall gold distribution for sample XR3 was the following:

(1)        The liberated, exposed and locked gold minerals account for 85.8%, 3.7% and 10.5% of total gold, respectively;

(2)        Heavy liquid products:

5.2% of Au was recovered in the HLS 3.1 Sink fraction, including 85.8% liberated, 3.7% exposed and 3.1% locked.
  4.8% of Au remained in locked the HLS Float fraction.

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13.3.2 MATERIALS HANDLING TEST WORK

Materials handling test work was conducted by Jenike & Johanson Limited (“JJL”), who were requested to perform flow property tests on samples of crushed XR1, XR2 and XR3 composite gold ore samples. Based on the results obtained Jenike and Johanson provided functional design recommendations for the crushed ore stockpile and reclaim system.

The test program consisted of the following tests:

  • Particle density - done using a 100 ml graduated cylinder.
  • Compressibility - determines the bulk density vs. consolidating pressure relationship.
  • Loose and Compacted Bulk Density - done using a 20-litre pail.
  • Flow Function - for calculating the critical outlet dimensions.
  • Wall Friction - for calculating mass-flow hopper angles.
  • Critical Chute Angle - for calculating critical chute angle.

The detailed material handling test work results and design parameters are presented in Appendix 7-4 Materials Handling Test Work and Design Criteria in the December 2012 Otjikoto Gold Definitive Feasibility Study.

Based on the test results, the following crushed ore stockpile reclaim system hopper/feeder design was proposed:

  • An upper pyramidal hopper, 5000 x 2500 mm at the top and 4284 x 1488 mm at the bottom with 14° wall slopes.
  • A 350 mm vertical spool piece for a spile bar gate
  • A 1050 high feeder interface
  • An 1800 mm wide apron feeder.

13.3.3 COMMINUTION TEST WORK AND GRINDING MODELLING RESULTS

Grinding Test Work

A grindability study was performed by SGS Lakefield on drill core samples from the Otjikoto deposit. The grindability program was performed on 3 composite and 17 variability samples. A summary of the test results are detailed in Table 13.7 below and the detailed grinding circuit design report is in Appendix 7-7-5 Grinding Circuit Test Work & Modelling Report of the December 2012 Otjikoto Gold Definitive Feasibility Study. The grinding test work results were used by SGS Lakefield for modelling with JKSimMet in order to size the grinding mills.

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Table 13.7:        Grindability Test Summary and Overall Statistics

DWT test results are bold

Grinding Circuit Power Requirements and Mill Sizing

The bench-scale grindability test results as presented in section 7.2.4.3 of the December 2012 Otjikoto Definitive Feasibility Study were used to design a grinding circuit capable of treating 2.4 Mtpa (304 t/h at 90% availability) from a primary crusher product P80 of 150 mm down to a product size P80 of 74 microns during the first years of operation. (The plant availability was later revised by B2Gold to 94% which increased the annual mill throughput to 2.5 Mtpa.) The grinding circuit design allows for a possible expansion by adding a pebble crusher to achieve 3.0 Mtpa (381 t/h at 90% availability).

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The grinding circuit design was based on the bench-scale grindability results for the XR3 ore type using the 85th percentile. The results for the XR3 85th percentile grindability data are presented in Table 13.8. XR3 was used as the design basis for the grinding circuit as it represented the hardest ore type of significant tonnage in the Otjikoto deposit.

Table13.8:        Grindability Data Set used for Mill Selection

The grinding circuit was sized based on the modelling and simulation results using two independent approaches namely;

  • JKSimMet modelling - performed by Mr F. Verret of SGS Lakefield
  • Particle breakage rate modelling – performed by Mr P. Morgan of DRA

A series of scoping simulations were performed to select the mill power and SAG and ball mill such to satisfy plant throughput requirements for both the SAB (SAG Mill/Ball Mill) and SABC (SAG Mill/Ball Mill/Pebble Crusher) circuit configurations. Based on the scoping simulations as well as consultation between DRA and B2Gold, the following mill selections and power requirements were determined:

  • SAG Mill: 24.0 ft Ø x 14.5 ft EGL and 4000 kW motor
  • Ball Mill: 17.0 ft Ø x 28.0 ft and 4000 kW motor

Upon further discussions with the mill supplier (Metso Minerals Industries, Inc. - “MMII”), the final ball mill selected was 16.5 ft ∅ x 28 ft EGL. Because of the wide variability in ore hardness, a variable speed drive will be installed on the SAG mill.

In order to provide an estimation of the expected plant throughput based on the selected mill sizes and installed motor power, variability simulations were run for both the SAB and SABC circuit based on the variability grindability results. The simulations indicated a substantial amount of variability in the ore hardness and mill throughput for all ore types. All the simulations were done using a fixed SAG mill ball charge of 12% and critical speed 75% and it was noted that for the hardest samples the target throughput rates of 304 tph and 381 tph were not achieved for the SAB and SABC circuit respectively. However, a series of simulations were undertaken for the hardest ore type ( A x b = 35.9) using a 15% SAG mill ball charge and a critical speed of 78%, which showed that the target throughput rates of 304 tph and 381 tph for the SAB and SABC circuit could be exceeded.

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UCS Test Work

The UCS test results on the XR2 and XR3 ore types along with the crushing work indices were provided to the gyratory crusher suppliers as part of the data for specifying the primary crusher design. The detailed UCS test work results can be found in Appendix 7-7-6 UCS Test Work Report of the December 2012 Otjikoto Gold Definitive Feasibility Study.

13.3.4 SGS LAKEFIELD GRAVITY CONCENTRATION AND LEACH TEST WORK ON COMPOSITE SAMPLES

Gravity concentration and leach test work was performed at SGS Lakefield on the XR1, XR2 and XR3 Metallurgical Recovery Composite samples. The detailed report can be found in Appendix 7-7 SGS Lakefield Final Test Work Report of the December 2012 Otjikoto Gold Definitive Feasibility Study.

The gravity concentration and leach test work carried out on the 3 composites was aimed at determining the optimum milling circuit grind. The grind optimization test programme was comprised of the following:

  • Grinding to a P80 of 100 mesh
  • Gravity concentration on each composite
  • Intensive cyanidation of gravity concentrates
  • Cyanidation of gravity tailings at 100 mesh, 150 mesh and 200 mesh

Once the optimum grind size was determined, bulk leach tests were performed in order to produce samples for downstream testing. The flowsheet for the entire test work programme on the XR1, XR2 and XR3 Metallurgical Recovery Composite samples is detailed in Figure 13.3 below.

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Figure 13.3:        Metallurgical Recovery Composites Test Work Flowsheet

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The gravity concentration testing conducted at SGS Lakefield was aimed at producing a gravity concentration tailings product for direct cyanidation at various grinds in order to determine the optimum mill grind. The gravity concentrates were then subjected to intensive leaching in order to obtain a preliminary indication of the expected gravity circuit gold recovery. It is important to note that the purpose of this test work was to determine the optimum mill grind and that all recoveries and reagent consumptions reported in this phase of test work were not used as a basis for design. The design basis for recovery and reagent consumption was based on the test work results for the metallurgical recovery variability test work that was conducted using the optimized conditions from the CANMET leach test work.

The results of the gravity concentration and intensive cyanidation of gravity concentrates are presented in Table 13.9 and Table 13.10 below.

Table 13.9:        Summary of the gravity concentration results on composite samples

Test
No.
Composite
Sample
Feed
wt.
(kg)
Tailing
P80
(micron)
Mass pull
wt.
(%)
Recovery
Au
(%)
**Adjusted Calc. Conc.
Au
(g/t)
Tailing
Au *
(g/t)
Head Grade
Au Calc.
(g/t)
Head Grade
Au Direct
(g/t)
G1 XR3 70 174 0.22 70.9 604 0.56 1.91 1.75
G2 XR2 60 155 0.20 82.5 885 0.38 2.15 1.70
G3 XR1 60 139 0.11 59.5 999 0.78 1.91 1.64

*

average of triplicate assays

**

adjusted concentrate grade Au (considering the back calc. values from intensive leach)

Table 13.10:        Results of the intensive cyanidation of gravity concentrates

Test
Number
Composite
sample
Weight of
Concentrate
(g)
P80 of
residue
(µm)
HEAD
(Calc.)
(g/t)
CN
Consumption
(kg/t)
Au in
PLS
(mg)
Au in
Residue
(mg)
Au
Extraction
(%)
CN1 XR3 141.2 174.0 619.3 88.2 57.5 0.453 99.9
CN2 XR2 107.9 155.0 823.1 8.9 89.9 2.260 99.7
CN3* XR1 51.5 139.0 1465 60.36 3.37 0.02 100

* presents the overal results of three intensive leaches conducted on the residue

For XR2 and XR3, the gravity concentration recoveries were 82.5% and 70.9% respectively. The gravity concentrates intensive leach tests indicated a gold recovery of 99.7% and 99.9% respectively after 48 hours of leaching.

For XR1 the gravity concentration recovery was 53.5% . The gravity concentrate intensive leach test initially produced low gold recovery after 48 hours of leaching and the residue was found to contain coarse gold grains. The residue was sent for SEM-EDS analysis, and the results indicated that the coarse particles were gold grains and not maldonite. The material was then leached again for 48hours, followed by 0.5 hours of cyanidation in a laboratory scale pebble mill and finally an 80 hour cyanide leach to achieve a final gold recovery of 100%.

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Gravity Tailings Cyanidation Grind Series Test Results

The gravity tailings were subjected to a cyanide leach at 500ppm NaCN for a period of 72 hours. The results of these cyanide leach tests are presented in Table 13.11 below:

Table 13.11:        Results of the Cyanidation tests on Gravity Concentration Tailings

Composite CN
Test
No.
Feed
Size
P80
% Au
Extraction
Head Au, (g/t) Residue
assay
Reagent
Consumption

(kg/t)
               
    (µm) 48 h 72 h Assay Calc. (g/t) NaCN CaO
XR3 CN4 152 58.7 73.5 0.56 0.63 0.17 0.91 0.59
CN5 100 74.3 87.8 0.56 0.78 0.10 1.01 0.59
CN6 70 84.6 96.7 0.56 0.60 0.02 1.45 0.66
XR2 CN7 149 87.7 88.7 0.38 0.38 0.04 0.14 0.45
CN8 105 92.8 91.9 0.38 0.41 0.03 0.16 0.43
CN9 83 95.5 95.7 0.38 0.47 0.02 0.16 0.47
XR1 CN10 142 79.8 88.4 0.78 0.54 0.06 0.09 1.08
CN11 97 81.0 86.7 0.78 0.66 0.09 0.08 0.95
CN12 70 72.3 91.6 0.78 0.59 0.05 0.09 1.03

Based on a high level economic review of the increased cost of grinding power to achieve the finer grind relative to the benefit of increased recovery, a mill grind of 80% passing 75μm (200 Mesh) was selected as the design basis. This grind was used in the CANMET leach optimization test work in order to determine the optimum reagent suite, and the required leach residence time for each ore type.

Preg-Robbing Tests

A Carbon–in-Leach (“CIL”) test was conducted on each composite at the target grind of 74 µm. The presence of preg-robbing can be indicated by a significant difference in the final residue gold assays when comparing conventional cyanidation and carbon-in-leach. The similarity of the direct cyanidation and CIL test results suggested that there was no preg-robbing in these composites.

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13.3.5 CANMET LEACH OPTIMIZATION TEST WORK

Leach optimization test work was performed at CANMET using gravity tailings samples generated during the gravity concentration and leach test work programme performed at SGS Lakefield. The detailed report is presented in Appendix 7-8 CANMET Leach Optimization Test Work Report of the December 2012 Otjikoto Gold Definitive Feasibility Study.

The CANMET leach optimization test work programme was aimed at determining the optimum leach conditions for each ore type; however XR3 was the primary focus because it is considered the most difficult ore type to treat due to its pyrrhotite content.

The following leach conditions were evaluated and optimized during this test programme:

  • Pre-treatment with air (dissolved oxygen (“DO”)>5 ppm) and pre-treatment with oxygen (DO>16 ppm)
  • Addition of lead nitrate
  • Leach Residence Time
  • Concentration of free cyanide in leach solutions

A summary of each test and the conditions evaluated is detailed in Table 13.12 below:

Table 13.12:        CANMET Leach Optimization Test Conditions

Test No.


Feed


Objective


Targetted Test Conditions
pH DO Temp. Pretreat     Cyanidation  
      Length Pb(NO3)2 Length   NaCN  Pb(NO3)2
  (ppm) (oC) (h) (g/t) (h) (ppm) (g/t)
1 XR3 4h PT, pH 10.5, 500ppm NaCN 10.5 >5 20 4 0 72 500 0
2 XR3 500 ppm NaCN 10.5 >5 20 0 0 72 500 0
3 XR3 4h PT, pH 10.5, 250g/t LN, 500ppm NaCN 10.5 >5 20 4 250 72 500 0
4 XR3 8h PT, pH 10.5, 250g/t LN, 500ppm NaCN 10.5 >5 20 8 250 72 500 0
5 XR3 0.5h PT, pH 10.5, 250g/t LN, 500ppm NaCN 10.5 >5 20 0.5 250 72 500 0
6 XR3 7h PT, pH 10.5, 100g/t LN, 500ppm NaCN 10.5 >5 20 7.0 100 72 500 0
7 XR3 7h PT, pH 10.5, 50g/t LN, 500ppm NaCN 10.5 >5 20 7.0 50 72 500 0
8 XR3 7h PT, pH 10.5, 100g/t LN, 400ppm NaCN 10.5 >5 20 7.0 100 72 400 0
9 XR3 7h PT, pH 10.5, 100g/t LN, 300ppm NaCN 10.5 >5 20 7.0 100 72 300 0
10 XR3 7h PT, pH 10.5, 100g/t LN, 300ppm NaCN, 16 ppm O2 10.5 >16 20 7.0 100 72 300 0
11 XR3 7h PT, pH 10.5, 100g/t LN, 300ppm NaCN 10.5 >5 20 7.0 100 72 300 0
12 XR3 7h PT, pH 10.5, 100g/t LN, 300ppm NaCN, 48 hours 10.5 >5 20 7.0 100 48 300 0
13 XR3 7h PT, pH 8, 100g/t LN, 300ppm NaCN, 48 hours 8 >5 20 7.0 100 48 300 0
14 XR3 7h PT, pH 10.5, 100g/t LN, 200ppm NaCN, 48 hours 10.5 >5 20 7.0 100 48 200 0
15 XR3 7h PT, pH 10.5, 100g/t LN, 300ppm NaCN, cement 10.5 >5 20 7.0 100 72 300 0
 
XR2-1 XR2 7h PT, pH 10.5, 100g/t LN, 500ppm NaCN, 48 hours 10.5 >5 20 7.0 100 48 500 0
XR2-2 XR2 7h PT, pH 10.5, 100g/t LN, 400ppm NaCN, 48 hours 10.5 >5 20 7.0 100 48 400 0
XR2-3 XR2 7h PT, pH 10.5, 100g/t LN, 300ppm NaCN, 48 hours 10.5 >5 20 7.0 100 48 300 0
XR2-4 XR2 7h PT, pH 10.5, 400ppm NaCN, 48 hours 10.5 >5 20 7.0 0 48 400 0
XR2-5 XR2 400ppm NaCN, 48 hours 10.5 >5 20 0 0 48 400 0
 
XR1-1 XR1 7h PT, pH 10.5, 100g/t LN, 500ppm NaCN, 48 hours 10.5 >5 20 7.0 100 48 500 0
XR1-2 XR1 7h PT, pH 10.5, 100g/t LN, 400ppm NaCN, 48 hours 10.5 >5 20 7.0 100 48 400 0
XR1-3 XR1 7h PT, pH 10.5, 100g/t LN, 300ppm NaCN, 48 hours 10.5 >5 20 7.0 100 48 300 0

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Based on the selected optimized leach conditions for XR3 of 7 hours of pre-treatment with air, 100 g/t lead nitrate addition, DO>5 ppm and 48 hours of leach retention time at 300 ppm NaCN from Test No. 12, a repeat of the test using the optimized leach conditions was conducted to determine the reproducibility of the results achieved for the optimum leaching conditions. The tests produced similar leach kinetics as shown below in Figure 13.4 and leach residues with the same gold content (0.03 g/t). The calculated gold leach extractions for Tests 12 and 17 were 91.5% and 93.6%, respectively which was due to the difference in calculated head grades of 0.35 g/t Au and 0.47 g/t Au. The cyanide consumptions were very similar (0.19 kg/t NaCN and 0.24 kg/t NaCN).

Figure 13.4:        Evaluation of reproducibility of leach kinetics and gold extraction for the selected optimum leach conditions for the XR3 ore sample

Leach Optimization for XR1 and XR2

From the optimum leach conditions selected for XR3, a series of leach optimization tests were also conducted for the XR1 and XR2 ore types. The focus of these tests was to optimize NaCN solution concentrations and determine the optimum lead nitrate addition requirement.

A summary of the optimized leach conditions for each ore type as determined by the CAMMET testing programme is presented in Table 13.13 below. The optimized leach conditions for XR3 were used as the design basis for the Otjikoto leach circuit. The leach circuit is comprised of one 3450 m3 pre-aeration tank to allow for 6.9 hours of pre-treatment followed by seven 3450 m3 leach tanks which allow for 48 hours of leach residence time at a feed solids concentration of 45% (w/w). The leach circuit design includes provision for the addition of compressed air, sodium cyanide and lead nitrate.

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Table 13.13:        Summary of Optimized Leach Conditions from CANMET Testing

  XR1 XR2 XR3
Pre-treatment None 7 Hours at pH 8 7 Hours at pH 8
Lead Nitrate Addition None 100g/t 100g/t
Free Cyanide Concentration 300ppm 400ppm 300ppm
Dissolved Oxygen Concentration >5ppm >5ppm >5ppm
Optimum Leach Residence Time 24 Hours 48 Hours 48 Hours

13.3.6 KNELSON GRAVITY CONCENTRATION AND INTENSIVE LEACH TEST WORK AND MODELLING/DESIGN

Gravity Concentration and Intensive Leach Test work

Knelson Research & Technology Centre conducted gravity concentration and intensive leach testing on splits of the XR1, XR2 and XR3 Metallurgical Recovery Composites from SGS Lakefield. The purpose of this test work was to determine the Gravity Recoverable Gold (“GRG”) value of each of the composites and the amenability of each ore sample to gravity concentration. The final test report for the Extended Gravity Recoverable Gold (”E-GRG”) and cyanidation test work can be found in Appendix 7-7-9 FLSmidth Knelson Gravity Concentration & Intensive Leach Test Work Report in the December 2012 Otjikoto Gold Definitive Feasibility Study.

The GRG test work programme was comprised of the E-GRG component, followed by intensive cyanidation of E-GRG concentrates. In this test work the gravity tailings were then subjected to direct cyanidation.

The results of the E-GRG test work were used to size and model the plant performance of the gravity circuit for Otjikoto. These simulations provided an estimation of the expected gravity circuit performance based on the samples tested for the XR1, XR2 and XR3 ore types.

The overall results for the E-GRG and cyanidation test work are presented in Table 13.14 below. The test results indicated high gold recoveries and both the E-GRG recovery and overall recoveries were in agreement with the SGS Lakefield test results for the Metallurgical Recovery Composite samples.

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Table 13.14:        Overall E-GRG and Cyanidation Test Work Recovery for XR1, XR2 and XR3 Composites

  XR1   XR2   XR3
           
Feed Grade (g/t) 1.55   1.80   1.51
Tails Grade (g/t) 0.082   0.030   0.057
           
  % Rec   % Rec   % Rec
Gravity Circuit          
                 Knelson Recovery 76.4   89.0   88.9
                 Intensive Leach 99.8   99.9   99.3
Total Gravity Circuit Recovery 76.2   89.0   88.2
           
Gravity Tails Leach 78.1   84.9   67.1
           
Combined Au Recovery 94.8   98.3   96.1

Overall gold recoveries for the XR1, XR2 and XR3 metallurgical recovery composites were very good, ranging from 94.8% to 98.3% . It should be noted that the gold head grades were quite low in the Knelson gravity tails leach tests for all 3 composites due to the gravity recovery testing methodology used. The low head grades led to a lower calculated gravity tails leach recoveries, but did not impact the overall gold recoveries.

Gravity Concentration Modelling and Circuit Design

The results of the E-GRG testing carried out on XR1, XR2 and XR3 metallurgical recovery composite samples by FLSmidth Knelson were used to model the gravity recovery of the full scale plant operations using KC MOD*Pro. This model has been used numerous times for both Greenfield simulations, as well as in operating plants, which allows a comparison of predicted results with actual.

For the Otjikoto project the modelling results were used to:

  • Determine the size and number of Knelson gravity concentrator units required for the full scale operation based on treating the entire ball mill discharge stream.
  • Determine the size of the Acacia leach reactor based on the simulated Knelson gravity concentration circuit performance and required leach time as determined by the test work.
  • Simulate the expected plant performance in order to provide an estimation of the expected gravity recovery that can be achieved for the full scale plant operations based on the samples tested in the E-GRG test work programme.

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The Knelson gravity circuit modelling report can be viewed in Appendix 7-7-10 FLSmidth Knelson Gravity Circuit Modelling Report in the December 2012 Otjikoto Gold Definitive Feasibility Study, a summary of these findings are presented below.

The following design data was used for the gravity circuit modelling:

  • Plant throughput – 304 mtph
  • Intensive leaching – Consep Acacia Reactor
  • Acacia leach time – Possible extended leach due to very coarse gold
  • Concentrating cycle time – 40 minutes
  • Knelson cone style – Generation 5, low mass yield, or possible mix of G5 and higher mass G6 cones
  • Circulating load – 300%
  • Sulfides in ore – Moderate – up to 5%
  • Gravity is installed on ball mill discharge and treats 100% of flow
  • Availability – Availability has not been taken into account, client to add a circuit availability factor.

The gravity test results for the Otjikoto samples showed unusually coarse GRG. Knelson indicated that for full scale plant operations, the coarseness of the GRG will be somewhat mitigated by the re-circulation effect in the cyclones, however, the GRG is coarse and abundant, and as such, a robust gravity circuit is required to recover a substantial portion of the gold. For this reason Knelson recommended treating the entire ball mill discharge through a gravity concentration circuit. In most cases, Knelson does not recommend treating the entire circulating load when leaching is the downstream process, however due to the very unusual nature of the GRG, it was recommended as the best treatment option for Otjikoto.

Based on the decision to treat the entire ball mill discharge it was determined that the circuit would require 4 KC-QS4 concentrators. It was recommended that due to the coarse nature of the GRG, provision should be made to allow for longer leach times of 48 hours. The longer leach times are to be achieved by selecting an oversized Acacia CS8000 (4,000-12,000 kg/batch) to allow for flexibility to run longer leach cycles. The modeling results are summarized in Table 13.15.

Table 13.15:        Gravity Circuit Modelling Results

Sample Gravity
Equipment
Gravity in BM Discharge
Gravity Recovery Concentrate Data
(% Au) (% GRG) (kg/day) (g/tonne)
XR1
XR2
XR3
4-QS48
4-QS48
4-QS48
64.8
80.1
71.9
90.4
91.6
89.3
6192
6192
6192
1169
2303
1500

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13.3.7 LAKEFIELD RECOVERY VARIABILITY TEST WORK

Gravity concentration and leach test work was performed at SGS Lakefield on drill core variability composite samples using the optimized CANMET leach conditions. The test work was aimed at determining the expected mill gold recoveries and reagent consumption figures for each ore type, as well as recovery variability. The detailed results from this test work can be viewed in Appendix 7-7-7 SGS Lakefield Final Test Work Report in the December 2012 Otjikoto Gold Definitive Feasibility Study.

The test work programme was performed on 46 variability composite samples as follows:

  • Eight (8) XR1 Composite samples
  • Fifteen (15) XR2 Composite Samples
  • Three (3) XR2-3 (XR4) Composite Samples
  • Twenty (20) XR3 Composite Samples

The test procedure used to simulate the gravity/whole ore leach process was comprised of the following:

  • Determination of direct head grade by screened metallics
  • Grinding to a P80 of 200 mesh
  • Gravity Concentration on each variability composite
  • Intensive Cyanidation of gravity concentrates (48 Hour leach)
  • Direct Cyanidation of gravity tailings using the CANMET optimized leach conditions identified for each ore type

The optimized leach tests indicated high gold recovery for all ore types. For XR1, the overall gold recovery for the eight variability composite samples ranged from 96.2% - 98.2% . The overall recovery was comprised of the gravity circuit recovery and the cyanide leach recovery. Gravity recovery ranged from 49.5% -77.6%, while the leach recovery ranged from 82.6% -95.2% . The average gravity tailings leach cyanide and lime additions were 0.55 kg/t and 1.02 kg/t respectively. Cyanide and lime test additions were used to estimate plant consumptions since there won’t be recovery of reagents in the Otjikoto process flowsheet.

For XR2, the overall gold recovery for the fifteen variability composite samples ranged from 96.4% -99.0% . Gravity recovery varied from 59.7% -87.1%, while the leach recovery varied from 83.1% -94.4% . The average gravity tailings leach cyanide and lime test additions were 0.69 kg/t and 0.86 kg/t, respectively.

For XR2-3 (XR4), the overall gold recovery for the three variability composite samples ranged from 94.6% -96.5% . Gravity recovery ranged from 54.7% -68.8%, while the leach recovery ranged from 87.2% -88.2% . The average gravity tailings leach cyanide and lime addition rates were 0.41 kg/t and 0.81 kg/t respectively.

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For XR3, the overall gold recovery for the twenty variability composite samples varied from 90.7% -99.4% . The sample with an overall gold recovery of 90.7% was considered an anomaly as the leach recovery was influenced by a single high residue assay of 1.28 g/t Au likely due to the presence of coarse gold that reported to the gravity tail. Gravity recovery ranged from 62.9% -86.1%, while the leach recovery ranged from 57.5% -93.4% . The average gravity tailings leach cyanide and lime addition rates were 0.59 kg/t and 0.53 kg/t, respectively.

It was noted that for all intensive leach tests the cyanide addition was as per the intensive leach test procedure and was in excess of the requirement. For this reason the cyanide consumption for the intensive leach process was not used for plant opex estimation, rather this consumption figure was provided by the equipment vendor as determined based on the Acacia reactor size and solution make-up requirements for full scale plant operations.

13.3.8 CARBON ADSORPTION TEST WORK AND MODELLING

The results of the CIP test work were for confirmation of the design parameters as presented in the process design criteria which can be viewed in Appendix 7-7-15 Process Design Criteria of the December 2012 Otjikoto Gold Definitive Feasibility Study. The final carbon adsorption test work and modelling results can be found in Appendix 7-7-7 SGS Lakefield Final Test Work Report of the December 2012 Otjikoto Gold Definitive Feasibility Study.

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13.3.9 SETTLING AND RHEOLOGY TEST WORK FOR THICKENER SIZING

FLSmidth Salt Lake City Inc. (“FLS”) carried out bench scale sedimentation tests on Gravity Tails (Pre-Leach Thickener) and Cyanide Destruction (“CND”) Product (Tailings Thickener) samples on behalf of B2Gold for the Otjikoto Gold Project. The detailed test report can be viewed in Appendix 7-7-11 FLSmidth Settling & Rheology Test Work Report of the December 2012 Otjikoto Gold Definitive Feasibility Study. The test work was conducted on the XR1, XR2 and XR3 Metallurgical Recovery Composites to determine the pre-leach thickener and tailings thickener performance for each ore type.

The following was evaluated during this test work programme:

  • Flocculant screening
  • Determination of feedwell design criteria related to effective feed conditioning for flocculation,
  • Settling tests
  • Measurement of thickened mud rheology

Flocculant screening showed that an anionic polyacrylamide flocculant with a high molecular weight and medium charge density produced the best settling rates and overflow clarity for all ore types and Ciba MF1011 was recommended as a suitable flocculant type for use on plant operations.

Pre-Leach and Tailings Thickener Design Parameters

Based on the test work findings the recommended pre-leach and tailings thickener design parameters are presented in Table 13.16 below.

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Table 13.16:        Thickener Design Parameters

  Gravity Tails (Grind) CND Product (Leach Tails)
    Thickener     Thickener  
Sample XR1-G XR2-G XR3-G XR1-L XR2-L XR3-L
Solids loading (Tpd) 7296 9120 9120 7296 9120 9120
Unit Area (m2/tpd) 0.043 0.034 0.034 0.043 0.034 0.034
Floc Dose (g/t) 30-40 10-20 10-20 30-40 10-20 10-20
Diameter (m) 20 20
Side Wall Height (m) 2.7 5
Floor Slope 9.5o 30o
Design Underflow Density (wt%) 55 55 55 58-60 75-76 75-76
Underflow Retention Time (hr) <1 <1 <1 1-2 1-2 1-2
Design Underflow Yield Stress (Pa) <30 <30 <30 40-80 30-80 <20
Design Overflow Solids (ppm) 200 100 100 200 100 100

The recommended design parameters for the pre-leach and tailings thickeners were based on the following:

  • The XR1 sample was significantly more difficult to settle primarily because a small fraction of the solids (likely a clay fraction) didn’t flocculate well.
  • The XR1 sample did show very high flux rates with well flocculated solids but cloudy supernatant.
  • The XR1 sample represents a small fraction of the overall ore body that will exist only during the initial phase of operation.
  • The Pre-Leach thickener will be operated to produce leach pulp at 45-55 wt% solids.
  • The Tailings thickener will target an underflow solids concentration of 55-60 wt% solids.

13.3.10        SO2/AIR CYANIDE DESTRUCTION TEST WORK AND DESIGN

SO2/Air cyanide destruction test work was performed at SGS Lakefield in Ontario, Canada under the guidance of Dr. Erik Devuyst. The test work was conducted on gravity tailings leach product generated from bulk leach tests on each of the XR1, XR2 and XR3 Metallurgical Recovery Composite samples. The results from this testing were sent to Erik Devuyst to provide design criteria for the cyanide destruction circuit as follows:

  • Cyanide destruction circuit residence time
  • Cyanide destruction tank size and mixer power requirement
  • Cyanide destruction air supply requirement
  • Recommended sizing of reagent make-up and storage tanks for the cyanide destruction circuit

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  • Preliminary estimate of reagent consumption based on metallurgical recovery composites
  • Updated reagent consumption estimates based on variability composites

The detailed process design and final report for the cyanide destruction circuit can be found in Appendix 7-7-12 Cyanide Destruction Circuit Design Parameters of the December 2012 Otjikoto Gold Definitive Feasibility Study.

SGS Lakefield Cyanide Destruction test work on metallurgical recovery composite samples

The detailed procedures and results of the cyanide destruction test work are presented in Appendix 7-7-7 SGS Lakefield Final Test Work Report of the December 2012 Otjikoto Gold Definitive Feasibility Study.. The objective of the testing was to establish the reagent consumptions and design parameters for the SO2/air CND process. The leach product from each of the metallurgical composites was subjected to cyanide destruction testing. The standard procedure for cyanide destruction using SO2/Air was started with a batch test to produce treated product with low residual cyanide for the continuous tests.

The results of the cyanide destruction tests are shown in Table 13.17. They indicate that for the metallurgical composites, the target CNWAD of <10 mg/L is achievable for all three composites.

Table 13.17:        CND Results of the Gravity Tailings Bulk Leach Product Samples

Feed Stream Dens. Ret. pH Assays
(mg/L)
SO2 Lime Cu
(%) (min) CNt CNwad Cu Fe SCN (g/gCNwad)
XR1
CND 1-1
CND 1-2
Feed 45   10.3 281 277 1.3 0.6 -      
Effl.
Effl.
45
45
59
58
8.5
8.6
2.82
2.81
0.7
0.63
0.07
0.09
0.89
0.98
<2
<2
5.36
4.27
2.54
3.27
0
0
XR2
CND 2-1
CND 2-2
Feed 45   10.4 212 177 8.2 30.2 -      
Effl.
Effl.
45
45
57
56
8.5
8.5
4.38
12
0.21
1.43
0.13
0.47
1.19
5.26
17
19
4.02
3.74
2.69
3.99
0.39
0.27
XR3
CND 3-1
CND 3-2
Feed 45   10.3 331 269 9.5 24.1 -      
Effl.
Effl.
45
45
58
56
8.5
8.5
45.7
77
1.68
37.9
7.39
8.88
13
26.1
300
320
4.04
3.88
1.3
1.14
0.19
0.11

The test results as presented in Table 13.17 were used to determine the typical and design conditions shown in Table 13.18. Based on these conditions a CND reactor volume of 1090m3 was recommended with an air supply requirement of 4000 scfm. The process design is based on a milling rate of 381 tph at 45 wt % solids with a P80 of 75 µm.

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Table 13.18: CND circuit design criteria based on the gravity tailings bulk leach composite samples

Design Milling Dens. Retn. SO2 Lime Cu 2+ Feed
CNwad
(t/h) (Wt%) (min) (g/gCNwad) (mg/L) (mg/L)
Typical 304 45 112 4.0 3.0  20.0 200
Design 380 40-50 80-100 5.0 4.0  50.0 325*

Cyanide destruction circuit design based on the XR1, XR2 and XR3 variability samples

Once the variability test work using the optimized leach conditions had been completed at SGS Lakefield, the leach solution assays were sent to Erik Devuyst for re-estimation of the expected CND reagent consumptions. The solution assays for the optimized leach tests conducted on the variability samples indicated a substantially different feed chemistry for the CND circuit as compared to the feed chemistry derived from the initial bulk leach tests conducted on the composite samples. This testing using the optimized leach conditions indicated that the feed to CND was expected to have lower levels of CNwad, Copper and Iron in solution, than those achieved for the initial 72 hour bulk leach tests at 500 ppm NaCN concentration.

Based on the solution chemistry for variability samples with optimized leach conditions the typical and design parameters for the CND circuit were re-evaluated, these updated parameters are presented in Table 13.19 below. The updated parameters showed that CND circuit residence time could possibly be reduced and that the air supply requirement could be reduced by as much as 50%. Both of these changes would reduce the capital cost for the circuit. It was thus recommended by Erik Devuyst that confirmation CND testwork should be performed on XR3 ore samples to confirm the possibility of reduced circuit residence time and air requirements. His recommendation was to only test XR3 gravity tailings leach product samples as XR1 and XR2 gravity tailings leach samples will work under the same or less severe operating conditions.

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Table 13.19:        CND circuit design criteria based on the gravity tailings variability samples

Design Milling Dens. Retn. SO2 Lime Cu2+ Feed
CNwad
(t/h) (Wt%) (min) (g/gCNwad) (mg/L) (mg/L)
Typical 304 45 60-75 4.0 2.5  20.0 100
Design 380 40-50 60 5.0 4.0  50.0 160*

It was decided that for the purpose of the feasibility study the CND circuit residence time and air supply requirements would be based on the SGS Lakefield CND test work performed on the Metallurgical Recovery Composite samples. The leach solution assay results for the variability test work using the optimized leach conditions were then used to provide estimations for the CND reagent consumptions for determination of plant operating cost.

13.3.11        SAMPLING TEST WORK PITARD TEST

The Pitard sampling test was performed at SGS Lakefield on a 1200 kg bulk sample of predominantly the XR3 ore type. Results of this test were provided to Dr. Francis Pitard who used this information to review the gold sampling characteristics for the Otjikoto Project and provide recommendations for mine, mill and assay laboratory sampling protocols.

The test procedure was comprised of the following process steps:

  • Sample Preparation and characterization
  • Gravity Separation
  • Intensive leaching of gravity concentrates
  • Bulk Cyanidation of Gravity tailings

The gold assay results for each process step were then used to calculate the head grade for the XR3 1200kg bulk ore sample. The results of this test were then sent to Dr. Pitard for evaluation.

Based on the results of this test as well as experience, Dr. Pitard provided detailed recommendations for the sampling protocol that would best be suited to the Otjikoto ore deposit and is presented in Appendix 7-7-13 Pitard Sampling System Design Reports of the December 2012 Otjikoto Gold Definitive Feasibility Study. Dr. Pitard also provided a detailed report with suggestions for the automated plant sampling systems.

13.3.12        TAILINGS TEST WORK

Tailing geochemical and geotechnical test work was conducted at SGS Lakefield and AMEC. The test work was aimed at providing a detailed characterization of the tailings for the design of the tailings storage facility (“TSF”).

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The geochemical testing was conducted at SGS Lakefield, with the following forming part of this scope:

  • Comprehensive solution and solids analyses
  • Modified acid base accounting
  • Net acid generation testing
  • Deionised water leach
  • Toxicity Characteristic Leaching Procedure (“TCLP”) Testing
  • Whole rock analyses

The geotechnical testing was conducted at SGS Lakefield and AMEC, with the following forming part of this scope:

  • Specific Gravity determination
  • Determination of particle size distribution
  • Settling Tests
  • Drained Settling Tests
  • Standard Proctor Tests
  • Determination of Atterberg Limits
  • Air Drying Tests
  • Rowe cell consolidation with hydraulic conductivity (Amec)
  • Triaxial Tests (Amec)

The detailed results of this test work can be viewed in in Appendix 7-7-14 Tailings Test Work Results and Reports of the December 2012 Otjikoto Gold Definitive Feasibility Study. The interpretation of these results and the impact on the tailings facility design are discussed in the tailings facility design report (Refer to Section 8 of the December 2012 Otjikoto Gold Definitive Feasibility Study).

13.3.13        PLANT RECOVERY

Based on the results of the variability composite testing with the optimized leach conditions as detailed Section 13.3.7, it is clear that high overall gold recovery was achieved as a result of the high GRG recovery as well as high leach gold leach extractions. Average life-of-mine gold recoveries are estimated to be 95.6% based on the empirical recovery formulas applied in the financial model by B2Gold. The contribution of the total GRG recovery (Gravity and Intensive dissolution) and leach recovery to the overall recovery is shown in Figure 13.5.

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Figure 13.5:        Overall Gold Recovery Contributions from Variability Test Work

Figure13.6 shows the relationships between sample head grade and GRG recovery. It is clear that a lot of scatter or variability exists, and that poor correlations are observed, as reflected in the R2 of the correlations. It therefore does not seem prudent to link gravity recovery estimates to feed grade, given that the average GRG contribution is 75%, and that a poor fit is observed between gravity recovery and feed grade.

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Figure 13.6:        Total GRG Recovery as a function of head grade from Variability Test Work

In order to provide an estimation of plant recovery based on the variability samples and testing that was performed at SGS Lakefield a statistical analysis of the test work results was conducted and is presented below.

Statistical Analysis of Variability Test Work Recovery Using Monte Carlo

The Monte Carlo probability distributions (derived from variability test results) for the overall recovery achieved in the variability tests were prepared. This analysis has only been performed for the XR1, XR2 and XR3 test results as the three (3) tests conducted on the XR2-3 (XR4) ore samples does not provide sufficient data to provide any meaningful results from a statistical analysis.

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Figure 13.7:        XR1 Monte Carlo Probability Distribution of Variability Test Work Recovery

Figure 13.8:        XR2 Monte Carlo Probability Distribution of Variability Test Work Recovery

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Figure 13.9:        XR3 Monte Carlo Probability Distribution of Variability Test Work Recovery

The probability distributions showed that high overall recoveries can be expected, although the distributions show that there is downside risk of achieving lower recoveries (albeit still good). The 90% confidence recovery range for all samples is shown on the graphs, and could be seen to provide a range of between 3 – 5% over this confidence range.

13.3.14        ESTIMATION OF OTJIKOTO PLANT RECOVERY

In order to provide an estimate of the expected recovery for full scale continuous plant operations, the bench scale laboratory recoveries were discounted by a factor of 1.5% . This discount factor is an estimated figure in order to account for process inefficiency and solution gold losses due to:

  • Plant operational disturbances
  • Soluble gold losses
  • Reagent addition inconsistencies
  • Carbon fines losses to tailings
  • Gold losses during plant start-up and shut down

Based on the discount factor of 1.5%, the estimated overall plant recovery for the XR1, XR2 and XR3 ore types are presented in Figure 13.. The estimated mean recovery for each ore type is shown relative the expected recovery range based on a 90% confidence interval.

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Figure 13.10:        Estimated Full Scale Plant Recovery Distribution for XR1, XR2 and XR3

At the request of Mr J. Rajala of B2 Gold, empirical models (based on mean correlations) were used for estimating plant recovery, mine cut-off grade and for the purpose of financial modelling. These recovery estimates were for full scale plant operations based on the mine production schedule and planned mill feed grade. The grade/recovery curves and empirical equations were used to estimate gold recovery as function of head grade particularly for gold grades near cutoff.

The empirical correlations presented in Figure 13.11 – Figure 13.13 below provide a logarithmic data fit to the laboratory recoveries as determined from the test work carried out on the variability composite samples using the optimized leach conditions. These model predicted recoveries, were then further discounted by 1.5% in order to obtain an estimate of the plant recovery based on the mill feed grade and mine production schedule.

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Figure 13.11:        Overall Gold Recovery as a function of Head Grade for the XR1 Variability Composites

Figure 13.12:        Overall Gold Recovery as a function of Head Grade for the XR2 Variability Composites

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Figure 13.13:        Overall Gold Recovery as a function of Head Grade for the XR3 Variability Composites

The discounted calculated recoveries from the empirical models presented in Figure 13.11 – Figure 13.13 are shown relative to the discounted expected recovery ranges for the 95% confidence intervals as detailed in Figure 13.10. It should be noted that the recovery estimate for the XR2-3 (XR4) ore type was determined based on the averaged results for XR2 and XR3 as there was not enough available information from test work to provide a more accurate recovery estimate for this ore type.

Figure 13.14. - Figure 13.17 show that based on mill feed grades as presented in the mine production schedule, the discounted calculated plant recovery estimates fall within the 90% recovery confidence interval as determined from the Monte Carlo analysis and presented in Figure 13.10 above.

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Figure 13.14:        Calculated Au Recovery based on Empirical Model Fit and Mine Production Schedule for XR1

Figure 13.15        Calculated Au Recovery based on Empirical Model Fit and Mine Production Schedule for XR2

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Figure 13.16:        Calculated Au Recovery based on Empirical Model Fit and Mine Production Schedule for XR2-3 (XR4)

Figure 13.17:        Calculated Au Recovery based on Empirical Model Fit and Mine Production Schedule for XR3

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14 MINERAL RESOURCE ESTIMATES

14.1 SUMMARY

The current resource model represents an update of previous models prepared by RSG Global (van der Merwe et al., 2005), SRK (Wanless et al., 2007; Wanless et. al., 2009; McDonald et. al., 2011) and BRME (Nicolls, 2011 and 2012).

In the current model, mineralized zone outlines were based on logged lithology, vein percentage and gold grades. The outlines define a nominal 0.4 g/t gold cut-off. Gold grades were estimated into a 3D block model using a mix of ID3 and OK (ordinary Kriging) estimation methods. A single indicator was used to better define high grade mineralization at a threshold of approximately 0.8 g/t gold. The total resource at the Otjikoto Project at a 0.4 g/t gold cut-off within a $1,350/oz. optimized pit is shown in Table 14.1.

Table 14.1: Otjikoto Mineral Resource Estimate, August 2012

  INDICATED INFERRED
Weathering kTonnes Au g/t kOunces kTonnes Au g/t kOunces
Ox/Trans 3,049 1.31 129 93 0.78 2
Sulphide 25,899 1.53 1,276 55 1.83 3
Total 28,949 1.51 1,405 149 1.17 6

Notes:
1) Mineral resources that are not mineral reserves do not have a demonstrated economic viability.
2) Due to the uncertainty that may be attached to inferred mineral resources, it cannot be assumed that all or any part of an inferred mineral resource will be upgraded to an indicated or measured mineral resource as a result of continued exploration.
3) Mineral resources are inclusive of mineral reserves.

14.2 DATABASE

The database used for this exercise was provided by Mike Glover (B2Gold database manager) and Liisa Kawali (Otjikoto database manager) in MS-Access. The primary database was issued on May 31, 2012 (Otjikoto DDH DB May 31 2012 LK V2.accdb). Certain key tables including assays, lithology and collars were updated with the June 6, 2012, version (Otjikoto DDH DB Jun 06 2012 LK.accdb).

The database was thoroughly validated and any errors were corrected.

RAB holes were excluded from this exercise due to concerns about the sample quality. In addition, certain holes lacking assays and lithology were not used. Holes with lithology but no assays were used to create oxidation surfaces and OTB marble wireframes.

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A summary of the holes drilled by year and type that were used in the resource is presented in Table 14.2.

Table 14.2:        Summary by year and type of holes used in resource database

  DDH RC TOTAL
YEAR #HOLES MS #HOLES MS #HOLES MS
1999 11 3,146 0 0 11 3,146
2001 9 1,397 9 621 18 2,018
2003 46 6,311 48 3,995 94 10,306
2004 17 1,316 0 0 17 1,316
2005 24 5,881 0 0 24 5,881
2006 66 15,081 76 4,485 142 19,566
2007 127 22,096 150 13,410 277 35,506
2008 51 9,083 39 2,730 90 11,813
2010 26 7,641 71 5,926 97 13,567
2011 199 50,164 44 5,368 243 55,532
2012 111 13,867 0 0 111 13,867
TOTAL 687 135,982 437 36,535 1124 172,517

*includes some condemnation holes and holes that were only used for geology (i.e. missing assays)

14.3 SOFTWARE

Exploratory data analysis was completed in MS-Excel and B2Gold’s proprietary EDA software. Variography was completed using Snowden Supervisor 7.

All interpretations, wire framing and block modeling were executed using Datamine Studio 3.

14.4 GEOLOGY

The geology of the Otjikoto deposit is summarized in Section 7 “Geological Setting and Mineralization” of this report. Of primary importance to the mineral resource estimation are the following:

  • Gold mineralization is concentrated in sheeted veins that are parallel to the northeast- striking S0/S1 foliation (locally 035°/25°) and form high grade shoots trending approximately 015° in the plane of mineralization.
  • The host rocks to mineralization are mainly hornfels and albitite, while waste rocks are dominantly biotite schists and marble.
  • Gold mineralization at Otjikoto is quite coarse and relatively high nugget values are observed.

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  • A well-developed 5 to 15 metres thick hardpan/calcrete unit overlies the deposit and oxidation extends to a depth of approximately 40 metres from surface.

Representative geological sections are presented in Figures 14.1 and 14.2.

Figure 14.1:        Summary geological section 6700N looking northeast

Figure 14.2:        Sample long section 4750E looking southeast

14.5 EXPLORATORY DATA ANALYSIS (EDA)

A comprehensive examination of logging parameters with regards to controls on grade was completed. The results of this study ultimately indicated that the only useful logging parameters for modeling were lithology and vein percent (Figure 14.3) . The majority of ore grade material occurs in the lithologies hornfels and albitite with greater than ~5% veining (Figures 14.4 and 14.5) . Logged pyrite (Py) and pyrrhotite (Po) abundance also show weak correlation with gold grades.

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Figure 14.3:        EDA Assays by lithology

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Figure 14.4:        Albitite with gold in vein bins

Figure 14.5:        Hornfels with gold in vein bins

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14.6 INTERPRETATION AND WIREFRAMING

Interpretations of mineralized zones were created using lithology, vein percent, sulphide abundance and gold grade at a nominal 0.4 g/t gold cut-off. Grades slightly below 0.4 g/t gold were included along the margins of zones or along strike/dip for the sake of continuity. The 0.4 g/t gold threshold was chosen for the following reason:

  • The proximity to a natural break that distinguishes mineralized from non-mineralized material at approximately 0.25-0.35 g/t gold;
  • 0.4 g/t gold is near but below the proposed cut-off grade (at the time of modelling the proposed cut-off was ~0.5 g/t gold); and
  • Historically 0.4 g/t gold grade shells have been used at Otjikoto.

Interpretations were created on 25 to 50 spaced vertical cross sections and 25 spaced vertical long sections.

Cross sectional interpretations were reconciled in 3D and joined to create wireframes. Interpretations were generally extended 50 metres beyond drilling or half way to the next drillhole that showed mineralization was absent.

An interpretation of the high grade core at 0.8 -0.9 g/t gold was also created, but due to time constraints, was not completely reconciled in 3D, and not directly used for the remainder of this study. A grade indicator was used in lieu of the high grade wireframe.

The bottom of transition and bottom of calcrete/hardpan surfaces were created by determining the thickness of hardpan and transition in each hole. This thickness was interpolated into a 2D grid and subtracted from the topography digital topography model (DTM) that was similarly gridded to produce the final surfaces. Areas that have very poor data coverage were given the mean thickness for each material type.

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14.7 DOMAINS

For the purpose of capping, variography, estimation and classification, mineralized zones were divided into seven different stratigraphic/structural/positional domains (Table 14.3) . Figure 14.6 shows a diagrammatic cross section of the position of each domain.

Table 14.3:        Resource domains

Stratigraphic
Domain
Description
SHALLOW Shallow domain just below oxide boundary at south of deposit
HW Mineralized zones in the hanging wall of the main zones
MAIN Largest most continuous zone in hanging wall to OTB marble
FW Mineralized zones in footwall of main zone but above OTB marble
MISC Miscellaneous zones in hanging wall, usually single hole intercepts
K2/WOLFSHAG Mineralized zones at north end of deposit below the footwall marble
WASTE Anything not in other domains

Figure 14.6:        Schematic showing general location of domains (looking north)

14.8 CAPPING AND COMPOSITING

The effect of statistical outliers was limited by capping assay values by each major structural/stratigraphic domain. Capping levels were determined from distribution plots

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(probability) and decile statistics. Assays were capped prior to compositing according to the scheme outlined in Table14.4. Unsampled intervals were treated as null.

Table 14.4:        Capping by Domain

Stratigraphic
Domain
Capping Level (g/t)
Number capped
Percentile
SHALLOW 4.5 5/337 98.5%
HANGING WALL 6.5 14/593 97.6%
MAIN 45 16/20139 99.9%
FOOTWALL 7.5 8/700 98.8%
MISC 4.5 8/110 92.7%
K2/WOLFSHAG 12 9/307 97.1%
WASTE 5.3 57/79,679 99.9%

The Main domain was further subdivided into high-grade and low-grade subdomains at a 0.8 -0.9 g/t gold cut-off for use in the indicator model.

2.5 metres downhole composites were created with breaks at mineralized zone contacts. This length was chosen based on grade continuity, desired coefficient of variation (“CV”) reduction, and Selective Mining Unit (“SMU”) size (i.e. half the proposed bench height). Composites lengths were permitted to vary a little to avoid the problem of residuals at the end of intervals.

Final capped composite statistics by domain are presented in Figure 14.7:

Figure 14.7:        Composite statistics by domain

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14.9 BLOCK MODEL PARAMETERS

Table 14.5:        Block model parameters

  X Y Z
Model origin 718900 7787000 1050
Block size 6.25 6.25 5
Number of blocks 448 560 94

The parent block dimensions were chosen based on drill density and grade variability but mostly on proposed mining SMU.

Sub-cells were created along the X and Y dimensions down to 3.125 metres. Blocks were filled in the Z dimension to fit the wireframes. Parent cells were sub-celled at weathering, mineralization and OTB contact.

Although the model would have benefited by being rotated, the difficulty caused by the rotation to users of the model outweighed the potential benefit so none was applied.

14.10 LITHOLOGY MODEL

A lithology model (as distinguished from mineralization) was created by interpolating an indicator using inverse distance squared (ID2) weighting for each of the following major rock types into the block model. The block was assigned the rock type of the dominant indicator.

Table 14.6:        Lithology Model codes

ROCK Description
1000 Albitite
2000 Albitite-Bio Schist
3000 Hornfels
4000 Marble (not OTB)
5000 OTB Marble
6000 Schist
7000 Undiff. rock
8000 Hardpan/Calcrete
9000 Mineralized rock

Oxidation/weathering was applied to the block model using the wireframes for each of these material types. The codes used for the weathering model are outlined in Table14.7.

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Table 14.7:        Lithology Model codes

WETH
WEATHERING
INTENSITY
800 Hardpan/Calcrete
200 Oxide/Trans
100 Fresh

14.11 DENSITY

Density was applied to the model by calculating the mean of SG samples by major rock type, weathering and mineralization state. A few obviously erroneous samples were removed from the dataset prior to calculations. Table 14.8 outlines the SG’s used in the model.

Table 14.8:        Density values applied to model

Oxidation Mineralization ALB ALBIO HORN MARB OTB SCHIST UNDIFF HARDPAN
                   
Ox/Trans Mineralized                
    2.64 2.59 2.58 2.57 2.57 2.40 2.56 2.49
  Unmineralized 2.61 2.46 2.56 2.49 2.49 2.40 2.50 2.49
Fresh Mineralized                
    2.84 2.75 2.78 2.76 2.78 2.77 2.78 2.49
  Unmineralized                
    2.76 2.74 2.74 2.74 2.73 2.77 2.75 2.49

*red-shaded cells contained insufficient samples. Values for these sub-domains were inferred from other domains or by regression.

14.12 METALLURGICAL DOMAINS

Metallurgical domains have been defined by oxidation state and dominant sulphide composition. Figure 14.8 shows the average pyrite (Py) to pyrrhotite (Po) ratio within the mineralized zones. The average was calculated across the entire mineralized zone so some of the internal variability has been smoothed out. That said, this method indicates clear zonation of sulphide species through the deposit. A polygon “cookie-cutter” was used to tag the block model with these domains.

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Figure 14.8:        Metallurgical domains

14.13 GRADE ESTIMATION

14.13.1 VARIOGRAPHY

As noted by previous authors, grades at the Otjikoto Project are extremely erratic, resulting in poor model variograms. For all but the Main domain, no valid variograms could be produced. The suite of experimental and fitted model variograms for the Main domain is presented in Appendix IV in the DFS. Model variogram parameters used for interpolation are presented in Table14.9.

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Table 14.9:        Variogram parameters

    Structure 1-Spherical Structure 2-Spherical
Vario
Nugget
Sill
Range-
principle
Range-
secondary
Range-
cross
Sill
Range-
principle
Range-
secondary
Range-
cross
Au 0.65 0.25 75 40 25 0.10 180 100 45
Indic 0.55 0.35 75 40 25 0.10 180 100 45
AU_LG 0.50 0.25 70 30 20 0.25 150 75 35
AU_HG 0.60 0.20 65 30 15 0.20 120 60 25

The orientation of the principle axis was oriented along the plunge of shoots within the overall plane of mineralization. The actual rotation angles used for the model were controlled by the local orientation of mineralized zones using Datamine’s dynamic anisotropy option (see search parameters).

Note that the nugget for each model variogram was manipulated downward from an average of approximately 0.7 -0.8 (normalized) to 0.5 -0.65 because early estimates were overly smoothed, producing ore grade material in marginal grade areas. The sills were proportionally modified to maintain an overall normalized sill of 1.

14.13.2 SEARCH CRITERIA

Search parameters were set up to ensure that blocks would be informed by on average 6-8 drillholes in the core of the deposit. Search ellipse parameters and search criteria are outlined in Table14.10.

Table 14.10:        Search criteria

Rotation of the ellipse was 40o clockwise looking down the positive Z axis, 28o counter clockwise looking down the positive Y axis, and 25o counter clockwise looking down the positive Z axis. Note, however, that the orientation was adjusted locally based on interpretations. For all interpolation runs, model variogram and search orientations were controlled by Datamine’s dynamic anisotropy. Using this method, orientations are modified (within specified limits) based on the local wireframe/interpretation on a block by block basis.

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14.13.3        ESTIMATION PLAN

Preliminary estimates showed that the kriged values were overly smoothed, spreading grade from the high-grade, north-trending shoots into areas of marginal grade, and severely decreasing the grade of blocks within the shoot. Ideally, wireframes would have been used to separate the very high-grade from the rest of the deposit, but time did not permit this.

Instead, a grade indicator was created at a nominal 0.8 -0.9 g/t gold cut-off. This threshold corresponds to a modest natural break in grade distribution. High grade indicator values (“1”) were assigned to intervals of contiguous zones above or near the chosen threshold. Spatially isolated assays above 0.8 g/t gold were not designated as high grade. Some consideration was given to adjacent holes while assigning the indicator, so marginal material may have been designated as “high-grade” for continuity sake even though it didn’t strictly meet the 0.8 g/t gold cut-off. Given the high grade variability and nuggety nature of the deposit, this vague definition seems reasonable.

The indicator was interpolated into the block model using ID3 interplantation. For better resolution of the indicator, blocks within the mineralized zones were re-blocked to a 2.5 metre bench prior to estimation.

High and low grade values were estimated using ordinary Kriging. The high and low grades were combined into a single block grade using a 50/50 rule. That is to say, if the indicator was greater than 0.5 g/t gold, the block was assigned the high-grade kriged value; if the indicator was lower than 0.5 g/t gold, the block was assigned the low-grade kriged value.

Grades for all domains other than Main were estimated using ID3 interpolation without an indicator since valid variograms were not available and high-grade shoots were not observed. Hard boundaries were applied between mineralized zones (as a whole) and waste data.

14.14 CATEGORIZATION

Mineral resource confidence classification has been defined taking into account the 2010 CIM Definition Standards for Mineral Resources as outlined below:

Inferred Mineral Resource

An “Inferred Mineral Resource” is that part of a Mineral Resource for which quantity and grade or quality can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified, geological and grade continuity. The estimate is based on limited information and sampling gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings and drillholes

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Indicated Mineral Resource

An “Indicated Mineral Resource” is that part of a Mineral Resource for which quantity, grade or quality, densities, shape and physical characteristics can be estimated with a level of confidence sufficient to allow the appropriate application of technical and economic parameters, to support mine planning and evaluation of the economic viability of the deposit. The estimate is based on detailed and reliable exploration and testing information gathered through appropriate techniques from locations such as outcrops, trenches, pits, workings, and drillholes that are spaced closely enough for geological and grade continuity to be reasonably assumed.

Confidence categorization was applied to the model based on the criteria outlined in Table14.11.

Table 14.11:        Resource categorization criteria

Stratigraphic
Domain
Resource Categorization
SHALLOW

Inferred: blocks within areas of up to 100X100 metre spacing, except some isolated discontinuous zones were set to “Other”
Other/Endowment: any estimated blocks not inferred

HW

Indicated: blocks within areas up to 25X50 metre drill spacing except some isolated discontinuous zones were set to “Other”
Inferred: blocks within areas up to100X100 metre drill spacing, except some isolated discontinuous zones were set to “Other”
Other/Endowment: any estimated block not indicated or inferred

MAIN

Indicated: blocks within areas up to 25X50 metre drill spacing except some isolated discontinuous zones were set to “Other”
Inferred: blocks within areas up to100X100 metre drill spacing except some isolated discontinuous zones were set to “Other”
Other/Endowment: any estimated block not indicated or inferred

FW

Indicated: blocks within areas up to 25X50 metre drill spacing except some isolated discontinuous zones were set to “Other”
Inferred: blocks within areas up to100X100 metre drill spacing, except some isolated discontinuous zones were set to “Other”
Other/Endowment: any estimated block not indicated or inferred

MISC

Other/Endowment: all estimated blocks

K2/WOLFSHAG

Other/Endowment: all estimated blocks

WASTE

Inferred: Blocks within 26X45 meters of a drillhole and estimated with at least 2 drillholes
Other/Endowment: all other estimated blocks

Any blocks categorized as “Other” were reset to a grade of 0.01 g/t gold prior to pit optimization and mine planning.

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Due to the nugget nature of the gold distribution in the deposit and lack of tightly spaced detailed sampling, no resources have been classified as Measured. Only Indicated resources were used for mining planning and economic analysis for conversion to the reserves outlined in Section 15 “Mineral Reserve Estimate” of this report. Due to the geological complexity of the deposit there is no guarantee that the Inferred resources can be converted to Indicated resources. “Other” resources are not NI 43-101 compliant and are not reported in this document.

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Figure 14.9:        Resource categorization

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14.15 BLOCK MODEL VALIDATION

The block model was validated using the following methods:

  1)

Visual comparison of composites to the block model on sections and long sections;

  2)

Comparison of composite and block model statistics; and

  3)

Comparison of interpolation methods by easting, northing and elevation on swath plots.

14.15.1 VISUAL COMPARISON

Composites were compared to the block model on every section and long section to ensure that the estimate appropriately reflected drillhole grades. The following observations were made:

  • The indicator adequately creates zones for high-grade/low-grade domaining and is representative of the volume of each material. Although some local over and under estimates were observed, they are minor in extent and common to this type of estimate.
  • Final grades are representative of the drillhole data.
  • Local grade variability is sometimes high making it difficult to compare blocks to composites.

14.15.2 COMPOSITE AND BLOCK MODEL STATISTICS

Final block grade (AUC2) statistics were compared to composites and the nearest neighbour (“NN”) model as a representation of de-clustered composites to determine if any global biases were introduced during the estimate. The direct (without indicator) ID3 and kriged models are also presented for comparison. (Table 14.12)

Final indicated block grades are within 2-3% of the NN model and within 4% overall. This is within reasonable limits.

The CV of the final model is about 60% of the NN model indicating that a significant level of smoothing is occurring. Further study should be completed to determine if this level of smoothing is appropriate for this deposit.

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Table 14.12:        Comparison of composite and block model statistics

14.15.3 COMPARISON OF ESTIMATION METHODS BY EASTING, NORTHING AND ELEVATION

Swath plot of average grade by coordinate axis were created to compare various estimation methods (Figure 14.10) . The final model (AUC2) tracks the NN model quite well but seems to be very slightly lower in all the swath plots.

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Figure 14.10:        Swath plot comparison of models

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14.16 MINERAL RESOURCE STATEMENT

The total undiluted mineral resources, unconstrained and constrained by optimized pits are outlined in Tables 14.13 and 14.14, respectively. The effective date of the mineral resource estimate is August 2012.

Table 14.13:        Undiluted Mineral Resources (unconstrained by optimized pit) (*base case highlighted in pink)

    INDICATED INFERRED
Weathering cut-off kTonnes   Au-g/t kOunces kTonnes   Au-g/t kOunces
Oxide 0.3 3,700 1.17 139 376 0.71 9
0.4 3,551 1.20 137 361 0.73 8
0.5 2,781 1.41 126 315 0.77 8
0.6 1,904 1.81 111 232 0.85 6
               
Sulphide 0.3 42,822 1.18 1,619 36,569 0.64 758
0.4 39,545 1.24 1,581 29,441 0.71 676
0.5 28,915 1.53 1,425 18,309 0.88 515
0.6 20,035 1.97 1,270 10,482 1.12 378
               
Total 0.3 46,522 1.18 1,758 36,945 0.65 767
0.4 43,096 1.24 1,718 29,802 0.71 684
0.5 31,696 1.52 1,551 18,624 0.87 523
0.6 21,939 1.96 1,381 10,714 1.12 384

Table 14.14:        Undiluted resource within a $US1,350/oz. pit (*base case highlighted in pink)

    INDICATED INFERRED
Weathering cut-off  kTonnes Au-g/t kOunces kTonnes   Au-g/t kOunces
Oxide    0.3 3,174 1.28 130 93 0.78 2
   0.4 3,049 1.31 129 93 0.78 2
   0.5 2,499 1.50 121 92 0.79 2
   0.6 1,848 1.84 109 80 0.82 2
               
Sulphide    0.3 27,015 1.48 1,290 63 1.66 3
   0.4 25,899 1.53 1,276 55 1.83 3
   0.5 21,015 1.78 1,205 50 1.96 3
   0.6 15,720 2.20 1,112 40 2.34 3
               
Total    0.3 30,189 1.46 1,420 156 1.13 6
   0.4 28,949 1.51 1,405 149 1.17 6
   0.5 23,514 1.75 1,325 143 1.20 6
   0.6 17,568 2.16 1,221 120 1.33 5

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Notes:
1) Mineral resources that are not mineral reserves do not have a demonstrated economic viability.
2) Due to the uncertainty that may be attached to inferred mineral resources, it cannot be assumed that all or any part of an inferred mineral resource will be upgraded to an indicated or measured mineral resource as a result of continued exploration.
3) Mineral resources are inclusive of mineral reserves.

The July 30, 2012 VBKOM whittle pit (Run1-Shell40-Milawa balanced version) was used for all tabulations and analysis within pits.

Key parameters used to create this pit are outlined in Table14.15.

Table 14.15:        Whittle pit key parameters for constrained in-pit resource

Parameters Value*
Gold Price $1350/oz.
Maximum pit slopes 50o in fresh, 40o in weathered
Mining Cost $1.91/tonne for waste, $2.09/tonne for ore
Mining Recovery 98%
Mining Dilution Average of ~10%
Processing costs $15/tonne
Process recovery gold 91% weathered, 94-96% Fresh
G&A $9,000,000/annum
Royalty 3%
Mill limit 2.4 Million tonnes per annum

*Note: Costs reported are as determined at time of resource estimation (July 30, 2012 VBKOM whittle pit).

Current cost and gold price assumptions suggest the cut-off grade for Otjikoto will be between 0.45 and 0.50 g/t gold. The resource base case is at a cut-off of 0.4 g/t gold. Grade tonnage curves are presented in Figure 14.11.

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Figure 14.11:        Grade tonnage curves

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14.17 COMMENTS ON SECTION 14

The Qualified Person responsible for the Mineral Resource estimate is Mr. Tom Garagan, P.Geo, Senior VP of Exploration for B2Gold Corp. Mineral Resources for the Otjikoto Project, have been estimated using drill data, performed to industry best practices (CIM, 2003), and conform to the requirements of CIM Definition Standards, 2010.

The Otjikoto Project is fully permitted; B2Gold Namibia has received a mining license and owns the surface rights to the area of the proposed infrastructure. The only remaining areas of uncertainty that could potentially materially impact the Mineral Resource estimate at the Otjikoto Project are the price of gold and long term exchange rate fluctuations.

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15 MINERAL RESERVE ESTIMATE

The mineral reserve for the Otjikoto Project was generated as part of the DFS completed on the

Otjikoto Project in December 2012 that is summarized herein in the Technical Report “Otjikoto Gold Project NI 43-101 Technical Report Feasibility Study” dated February 25, 2013. The mineral reserve has not changed and is based on a block model resource and mine plan that envisions open pit mining using conventional hard rock mining techniques. The Qualified Person for the mineral reserve estimate is Hermanus J. Kriel, CEO, Senior Mining Engineer, Pr. Eng, VBKom Namibia Consulting Engineers (Pty) Ltd.

Variations of the price of diesel and heavy fuel oil, and changes in the price of gold could materially affect the mineral resources and mineral reserves in either a positive or negative manner. At the time of this report there is no known permitting, metallurgic, environmental or social economic conditions that would have a material effect on the mineral reserve estimate of the Otjikoto Project.

15.1 MINERAL RESERVE CLASSIFICATIONS

The mineral reserves for the Otjikoto Project were developed by applying the relevant economic and design criteria to the resource model in order to define the economically extractable portions of the resource. The reserves were developed and are disclosed in this report in accordance with NI 43-101.

Mineral Reserve

Mineral reserves are subdivided in order of increasing confidence into probable mineral reserves and proven mineral reserves. A probable mineral reserve has a lower level of confidence than a proven mineral reserve. The reserves for the Otjikoto Project are all in the probable category.

15.2 PIT OPTIMIZATION METHODOLOGY

The open pit mine design is based on conventional floating cone techniques to establish guides to mineable shapes within the mineral resource block model. The block model was imported into Gemcom’s Whittle pit optimization software for final analysis. Whittle optimization software is an industry standard program used worldwide to assist in the development of open pit mine planning. Whittle uses a series of economic constraints as well as slope angle limitations and ore recoveries to establish the most economic cone possible. Whittle optimization is an iterative process using costs developed during previous studies, refined to be as accurate as possible. Although a detailed description of the optimization methodology is beyond the scope of this report, the following section provides a brief summary. The optimization process can be divided into two processes, as follows:

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  • Creation of a range of nested pit shells of increasing size achieved by varying the product price and generating a pit shell at each price point; and
  • Selection of the optimal pit shell by generating various production schedules for each pit shell and calculating the NPV for each schedule, the output of this process being a series of “pit versus value” curves.

Table 15.1:        Pit optimization input parameters

                                                   Parameter Unit Value
Base Currency   USD
Exchange Rate ND:USD 8.00
Discount Rate (%) 5.00%
Commodity Price    
Ounce Conversion to Grams constant 31.10348
Gold Price - Base Case (USD/oz) 1 350
Government Royalty (%) 3.0%
Net Gold Price (USD/oz)
(USD/g)
1 309.50
Gold Price - Spot Price (USD/oz) 1 600
Gold Price - Bear Case (USD/oz) 1 000
Gold Price - Bull Case (USD/oz) 1 700
Gold Price - Blue Sky Case (USD/oz) 2 200
Selling Cost (Transport) (USD/oz) 8.00
Diesel Cost (USD/L) 1.06
Power Cost (USD/kWhr) 0.20
Mining Block Model Dimensions    
Block Model Name   otj_jun20_pctmod3.csv
Origin X(m), Y(m), Z(m) 718 900, 7 787 000, 1 050
Extent X(m), Y(m), Z(m) 448, 560, 94
Block Size X(m) x Y(m) x Z(m) 6.25 x 6.25 x 5
Geotechnical Design Parameters    
Weathered    
Overall Slope Angle - Pit Design Sector 1 (deg) 40
Overall Slope Angle - Pit Design Sector 2 (deg) 40
Overall Slope Angle - Pit Design Sector 3 (deg) 40
Overall Slope Angle - Pit Design Sector 4 (deg) 40
Fresh    
Overall Slope Angle - Pit Design Sector 1 (deg) 50
Overall Slope Angle - Pit Design Sector 2 (deg) 50
Overall Slope Angle - Pit Design Sector 3 (deg) 50
Overall Slope Angle - Pit Design Sector 4 (deg) 50
Ramp Width    
Two Way (m) 21.5

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Parameter Unit Value
One Way
Ramp Gradient
(m)
(1:#)
16
10
Mining Factors    
Dilution
Diluent Grade
Ore Loss
Minimum Ore Parcel
(%)
(g/t)
(%)
(%)
variable per SMU block
0.00
2%
35%
Mining Operating Costs (excluding SG&A cost)    
Start-Up Mining Cost Factor Year 1
Year 2
110%
100%
Steady State Reference Mining Operating Cost
Depth Factor for Mining Operating Cost
Incremental Ore Mining Operating Cost
(USD/tonne)
(USD/tonne/10m bench)
(USD/tonne)
1.91
0.025
0.18
Plant Parameters    
Process Recovery Ramp-Up (% of steady state recovery) Year 1
Year 2
98.0%
100.0%
Processing Recovery - XR1 (oxides, includes transitional & hardpan)
Processing Recovery - XR2 (pyrite dominant mineralization)
Processing Recovery - XR3 (pyrrhotite dominant mineralization)
Processing Recovery - XR2/3 (Mix pyrrhotite ) and pyrite
%
%
%
%
91.0%
95.8%
93.6%
94.7%
     
Mining Limit    
Year 1
Year 2
Year 3
Steady State
Mtpa
Mtpa
Mtpa
Mtpa
10
17.5
25
25 +
Plant Limit    
Year 1
Year 2
Year 3
Steady State
Mtpa
Mtpa
Mtpa
Mtpa
1.8
2.4
2.4
2.4
Plant Operating Costs (excluding SG&A cost)    
Start-Up Processing Cost Factor Year 1
Year 2
110%
100%
Processing Cost - Oxide Material - Steady State
Processing Cost - Transition Material - Steady State
Processing Cost - Sulphide Material - Steady State
USD/t ROM ore
USD/t ROM ore
USD/t ROM ore
15.00
15.00
15.00
Fixed Cost - Supervision, General & Administrative (SG&A) Cost
Year 1 M USD/annum 9.00

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Parameter Unit Value
Year 2 M USD/annum 9.00
Steady State M USD/annum 9.00

15.3 GEOTECHNICAL PARAMETERS FOR PIT DESIGN

SRK Consulting completed a feasibility level geotechnical field investigation and analysis of the Otjikoto pit slopes and bench details. The resulting pit slope and bench parameter recommendations are summarized in Table 15.2 below.

Table 15.2:        Otjikoto Pit slope architecture

Parameter Hanging wall Foot wall
Weathered Rock Profile
Bench Heights
Bench Angles
Berm Width
Stack Height
Stack Angles
Catch Berm
10m
75o
9.15m
50m
45o
15m
10m
75o
9.15m
50m
45o
15m
Fresh Rock Profile
Bench Heights
Bench Angles
Berm Width
Stack Height
Stack Angles
Catch Berm
10m
75o
3.4m
50m
50°
15m
10m
75o
3.4m
50m
50°
n/a

15.4 ULTIMATE PIT DESIGN

The ultimate pit design is based on the optimum pit shell and geotechnical parameters described above. Additionally, the pit is designed to meet the operational constraints summarized in Table15.3 below. The resulting pit design is shown in Figure 15.1 below.

Table 15.3:        Otjikoto pit design operational constraints

Description Value
Haul Road Gradient
Haul road width for double-lane traffic
Haul road width for single-lane traffic
Minimum practical mining working width
Switchback operating radius
1:10
21.5 m
16 m
40 m
60 m

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Figure 15.1:        Plan view of the Otjikoto pit design

15.5 COMPLIANCE WITH THE WHITTLE SHELL

The optimal whittle shell represents the shell with the highest NPV but not practical to mine as there are no access ramps. When the optimal shell is converted into a practical pit, the NPV of the resultant pit is expected to be lower because of the extra waste that will have to be mined to make room for access ramps. It is however important that the difference in volumes and overall value is kept at a minimum. Table 15.4 below shows a comparison between the ultimate pit design content (Indicated material above a cut-off grade of 0.4 g/t) and the optimal whittle shell.

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Table 15.4:        Comparison between the ultimate pit design and the optimal Whittle shell

Unit Ultimate
Pit Design
Optimal
Whittle Shell
%
Variance
ROM Ore
Metal Content
Metal Content
Ore Grade
Total Waste
Stripping Ratio
tonnes
grams
ounces
g/t
tonnes
29 405 338
41 714 452
1 341 151
1.42
164 266 053
5.59
27 640 000
40 997 497
1 318 100
1.48
149 188 000
5.40
6%
2%
2%
-4%
10%
4%
Total Tonnes tonnes 193 671 391 176 828 000 10%

As indicated above, there is a 10% increase in the total waste and an increase of 6% in the run-of-mine ore tonnes at a lower average grade of 1.42 g/t as compared to 1.48 g/t. This results in a net increase in stripping ratio from 5.40 to 5.59. The results also indicate a marginal increase in the metal contained. It is important to note that the additional ore mined is of low grade and therefore does not add significant value to the project. Generally, there is a close correlation between the optimal whittle shell and ultimate pit design.

15.6 CUT-OFF GRADE CALCULATION

The method employed for classifying material mined as ore and waste should not be confused with the method for establishing the limits of mining. If a block of material falls inside the optimized mining limits then the question is not whether to mine the block but whether to process the material. This study is based on the assumption that a block of material should be processed if the income derived from the sale of product covers at least the cost of processing. The marginal cut-off grade is therefore the grade at which the income from the sale of product is equal to or more than the cost of processing. Cut-off grades are calculated on a break-even basis and the approach assumes the cost of mining material out of the pit to the waste dump is a sunk cost as it is intrinsic to the mining process, regardless of whether the material is ore or waste. The assessment of whether material is ore or waste occurs once it has been removed from the pit. Similarly, capital is a once-off cost that is not applicable to the instantaneous evaluation of a tonne of material to determine its classification.

The break-even cut-off grade determines whether a tonne of material is ore on the basis that the revenue generated has to be greater or equal to the additional cost of that tonne processed through the plant. The marginal cut-off grade is therefore the grade at which the income from the sale of product is equal to or more than the cost of processing. The marginal break-even cut-off grade for ore is calculated as follows:

Where:

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Total cost of processing includes the incremental cost of mining ore as opposed to mining waste; Net price includes a state royalty of 3% and a selling cost of USD8/oz; and Processing recovery of gold varies based on material type.

The cut-off grades calculated for the different material types are summarized in Table 15.5 along with the relevant input parameters.

Table 15.5:        Break-even cut-off grades

                               Item Unit XR-1 XR-2 XR-3 XR-4
Processing Cost
Incremental Ore Cost
Total Cost of Ore
Price
Price
Selling Cost
Selling Cost
Royalty
Net Price
Recovery
Cut-off Grade
USD/tonne
USD/tonne
USD/tonne
USD/oz
USD/g
USD/oz
USD/g
%
USD/g
%
g/t
15.00
0.18
15.18
1350
43.40
8
0.26
3
41.84
91.0%
0.40
15.00
0.18
15.18
1350
43.40
8
0.26
3
41.84
95.8%
0.38
15.00
0.18
15.18
1350
43.40
8
0.26
3
41.84
93.6%
0.39
15.00
0.18
15.18
1350
43.40
8
0.26
3
41.84
94.7%
0.38

15.7 OTJIKOTO PROBABLE MINERAL RESERVE

The probable mineral reserves for the Otjikoto Project are provided in Table 15.6 below. These reserves are based on the pit designs discussed above. The mineral reserves have been shown to be economic and are reasonable for the statement of probable reserves. The final probable mineral reserves for the Otjikoto Project are 29.4 million tonnes of ore at a diluted grade of 1.42 g/t resulting in 1.3 million ounces (39.3 million grams) of contained gold at a stripping ratio of 5.59:1.

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Table 15.6:        Probable Mineral Reserve, effective date December 2012

Parameter Unit Value
Total Ore
Total Waste
Stripping Ratio
Grade
Metal Content
tonnes
tonnes

g/t
ounces
29 405 338
164 266 053
5.59
1.42
1 341 151

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16 MINING METHODS

The objective of the pit design process was to transform the pit shells obtained from the optimization into a practical pit, with the inclusion of ramps, bench and berm configurations by taking all the required inputs into account. The practical pit design forms part of a critical input for the scheduling and conversion of resources into reserves. The WhittleTM pit optimization outputs, the design criteria and geotechnical constraints were used as input parameters in order to design the practical final pit. Pushbacks were based on the interim selected WhittleTM shells and designed using the recommended geotechnical parameters and pit design criteria derived from the equipment strategy as well as current world best practices. The designs were created using SurpacTM and MicromineTM mining software.

Two important considerations for the pit design were the pushback strategy and the positioning of the access ramps. The optimization exercise has indicated that improved value can be generated by this project through an optimum extraction sequence. The starting point of an optimum scheduling sequence is an informed decision regarding pushbacks. The selected interim and ultimate pit shells were used as basis for the practical pit and pushback designs.

The WhittleTM pit optimization exercise resulted in a selection of two lower revenue factor shells, the combination of which provide an optimum extraction sequence, which ensures that grade to the mill is maximized in the early years and waste stripping is deferred as far as possible into the future. The selected shells provided some guidance towards the location of interim stage designs.

16.1 PIT SLOPE AND BENCH DESIGN PARAMETERS FOR PIT DESIGN

SRK Consulting completed a feasibility level geotechnical field investigation and analysis of the Otjikoto pit slopes and bench details (The final SRK Geotechnical Report is included in the DFS dated December 2012). The field investigation included twelve oriented core drill holes in multiple pit interior, foot wall, and hanging wall locations. This was followed by a complete lab test regimen to determine the rock mass and joint strength parameters. The stability analysis included stereograph, limit equilibrium (SLIDE) and stress deformation (PHASE2) analyses in drained and un-drained scenarios. The resulting pit slope and bench design recommendations are summarized in Table 16.1 below.

Table 16.1:        Otjikoto Pit slope architecture

Parameter Hanging wall Foot wall
Weathered Rock Profile
Bench Heights
Bench Angles
Berm Width
10 metres
75o