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 Surpac^TM and Micromine^TM 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, 22^nd 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 10^th 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 7^th to 12^th, 2012
• February 3^rd to 5^th, 2012
• August 21^st to 24^th, 2012
• November 20^th to 22^nd, 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 40^o clockwise looking down the positive Z axis, 28^o counter clockwise looking down the positive Y axis, and 25^o 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 Whittle^TM 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 Whittle^TM 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 Surpac^TM and Micromine^TM 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 Whittle^TM 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 9.15 metres	10 metres 75o 9.15 metres

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

16.2 BENCH HEIGHT AND MINING DIRECTION

The decision regarding the bench height takes into account the geometry of the ore body, the required mining rate, and equipment type and size. Under these considerations the drilling and blasting in ore and selective waste will be carried out in 5 metre benches and mined selectively in a single 5 metre flitch while bulk waste benches will be blasted and loaded in 10 metre benches. To reduce dilution, mining of ore and selective waste will start at the hanging wall and proceed towards the foot wall. Furthermore, ore will generally be mined during daylight hours.

16.3 HAUL ROAD DESIGN

The haul road design parameters were established taking into consideration the type and size of material hauling equipment that will be used during the operation.

The dimensions of the haul road were based on a Hitachi EH1700 dump truck using global standards of good practice. Many of the guidelines specify that the vehicle operating width should be multiplied by a factor of 3 for two-lane traffic and 2 for single-lane traffic in order to determine the effective operating width of the haul road and to incorporate the road infrastructure, for example, the safety berm and drainage channel. The haul road gradient and width are discussed below.

16.3.1 HAUL ROAD GRADIENT

A reduction in road grade significantly increases a vehicle's attainable uphill speed. Thus, haulage cycle times, fuel consumption, and stress on mechanical components, which results in increased maintenance costs, can be minimized to some extent by limiting the severity in grades.

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A haul road gradient of 1:10 (10% or 5.71º) was selected for the Otjikoto Project. The selection of the haul road gradient was based on the world best practice for the type of trucks that will be utilized and local weather conditions.

16.3.2 HAUL ROAD WIDTH

The equipment study conducted concluded that Hitachi EH1700 dump trucks or similar will be used to haul broken rock out of the pit and therefore the road dimensions were based on this type of truck equipment, taking into consideration global standards of good practice. Designing for anything less than this dimension will create a safety hazard due to a lack of proper clearance. In addition, narrow lanes often create an uncomfortable and unsafe operating environment, resulting in slower traffic and therefore impeding on production.

Rules of thumb for determining haulage road lane dimensions vary considerably from one reference source to another. For the purpose of this report, the effective operating width of the haul road for a single-lane was calculated by multiplying the physical truck-operating width by a factor of 1.8 and a factor of 2.7 for a double-lane haul road.

The selected Hitachi EH1700 has a physical truck-operating width of 6.250 metres.

The haul road width for double-lane haul road was calculated as follows:

• Safety berm height = 1.35 metres
• Safety berm width = 3.85 metres (based on a Bridgestone 27.00R49 E4 radial tire with a diameter of 2.7 metres)
• Drainage channel = 0.5 metres
• Design width calculation = (2.7*6.250) metres + 3.85 metres + 0.5 metres = 21.225 metres
• Practical design width = 21.5 metres

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Figure 16-1:        Two-lane traffic haul road configuration

Because the last five levels of each pushback will be open and in use for a limited period of time, extra measures would be taken to reduce waste stripping. These last five levels will have a relatively low stripping ratio which implies less equipment maneuvering on these levels. The road width is decreased to allow for one-way traffic.

The haul road width for single-lane haul road was calculated as follows:

• Design width calculation (1.8*6.250) metres + 3.85 metres + 0.5 metres = 15.6 metres
• Practical design width = 16 metres

16.3.3 SURFACE HAUL ROADS

The surface haul road width is designed for a width of 25.5 metres, derived by adding a second safety berm (3.85 metres) to the 21.5 metres practical design width for double-lane traffic.

16.4 ULTIMATE PIT DESIGN

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 reduced at the lower levels of the pit to minimise waste stripping as much as possible. The design work was performed in Gemcom’s Surpac^TM and Micromine^TM

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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.

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 16-2 below and the internal phases are shown in Figure 16-3 below. The ultimate pit will have dual access established along the final limits as indicated in Figure 16-2 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.

The ultimate pit dimensions are as follows:

•	Length	1,970 metres
•	Width	585 metres
•	Depth	245 m
•	Area	816,800 m2 (81.68 ha)

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Figure 16-2:        Otjikoto ultimate pit design

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Figure 16-3:        Otjikoto ultimate pit and internal phase designs

In Table 16.2 below is a summary of the indicated material content with applied mining modifying factors of the ultimate pit above a cut-off grade of 0.4 g/t. The total indicated run-of mine ore tonnes contained in the ultimate pit design are 29.41 million tonnes at an average head grade of 1.42 g/t, and the contained ounces sum up to 1.34 million ounces. The stripping ratio for the final pit design is 5.59 resulting in 164.27 million tonnes of waste to be stripped in order to expose the 29.41 million tonnes of ore. The life of mine is 12.46 years assuming a plant throughput of 2.5 Mtpa, and a plant commissioning ramp-up of 1.96 Mtpa for the first year of production and 2.34 million tonnes per annum for the second year of production. All figures quoted are above a gold cut-off grade of 0.4 g/t and mining modifying factors applied.

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Table 16.2:        Otjikoto ultimate pit design ore and waste content - indicated material

	Met Type	Rock Type	Unit	Pushback_01	Pushback_02	Pushback_03	Total
ROM Ore Tonnes	XR1 XR2 XR3 XR2/3	Oxide/Trans Pyrite (fresh) Pyrrhotite (fresh) Pyrite & Pyrrhotite	tonnes tonnes tonnes tonnes	2 336 956 3 434 104 283 277 3 237 073	771 845 3 470 436 1 982 438 5 639 392	54 737 459 075 6 048 096 1 687 910	3 163 538 7 363 615 8 313 811 10 564 375
Total ROM Ore			tonnes	9 291 409	11 864 111	8 249 818	29 405 338
Metal Content	XR1 XR2 XR3 XR2/3	Oxide/Trans Pyrite (fresh) Pyrrhotite (fresh) Pyrite & Pyrrhotite	grams grams grams grams	3 078 574 4 323 240 467 597 6 292 590	778 908 3 298 968 3 487 932 7 397 951	40 320 316 643 10 659 874 1 571 856	3 897 801 7 938 851 14 615 403 15 262 397
Metal Content	Total		grams	14 162 001	14 963 758	12 588 693	41 714 452
AU Head Grade			g/t	1.52	1.26	1.53	1.42
Waste		Hardpan/Calcrete Otb Marble Min/Unmineralized	tonnes tonnes tonnes	6 486 481 5 410 19 505 626	7 036 041 209 155 43 253 581	8 384 386 572 787 78 812 587	21 906 908 787 352 141 571 793
Waste	Total		tonnes	25 997 516	50 498 776	87 769 760	164 266 053
Total Tonnes			tonnes	35 288 926	62 362 887	96 019 578	193 671 391
Stripping Ratio		Tonnes of waste per tonne of ore		2.80	4.26	10.64	5.59

As indicated in the table above, 164.27 million tonnes of waste will be mined to expose 29.41 million tonnes of ore, resulting in a stripping ratio of 5.59 over the life of mine. The total waste content is made up of un-mineralized and mineralized rock below the cut-off grade of 0.4 g/t, OTB marble, and hardpan/calcrete. The major lithological unit contained inside the pit design is the albitite / hornfels /schist package followed by the weathered bedrock and transition material.

16.4.1 GRADE-TONNAGE CURVE

Grade-tonnage curves are used to determine the volume and grade of material in face of variations of the cut-off grade. The grade-tonnage curve in Figure 16-4 below was calculated based on the diluted, indicated ore content of the pit design assuming 2% ore loss and including only blocks with greater than 35% ore. At a cut-off grade of 0.4 g/t approximately 29 million tonnes of ore will be mined at an average grade of 1.4 g/t.

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Figure 16.4:        Otjikoto Pit Grade-Tonnage Curve

16.5 ANNUAL PIT DESIGNS

As a part of the overall scheduling, annual pit progressions maps complete with haul road access were reviewed to ensure that all mining areas will have ramp access throughout the mine life. See Figures 16-4 and 16-5 below for examples of annual pit maps.

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Figure 16.5:        Otjikoto pit progression map, year 3

Figure 16.6:        Otjikoto pit progression map, year 6

16.6 MINE ROCK STORAGE FACILITY

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 perimeter 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 16-3 below shows the phased and total capacity and Figure 16-6 below shows the waste dump configuration and progression.

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Table 16.3:      Rock dump capacities

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

Figure 16.7:        Otjikoto Mine Rock Storage Facility

16.7 MINE EQUIPMENT SELECTION AND SCHEDULE

The surface mine has been planned as an open-pit mine using diesel down the hole (“DTH”) blast hole drills, diesel 100 t class haul trucks, and 180 tonne class hydraulic shovels to achieve the most efficient material movement possible. Blast hole drilling for both ore and waste material utilizes 325 kW class diesel drills. Primary mine ore and waste production is achieved using 180 tonne hydraulic excavators in both front and backhoe configuration, depending on the specific lithology encountered, and Cat 777 haul trucks. A Cat 992 loader is available for additional loading capacity, and a secondary fleet consisting of smaller excavators, bulldozers, graders, and wheel dozers will be in place to support the primary mining fleet.

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16.7.1 BLAST HOLE DRILLING EQUIPMENT

The blast hole drills for Otjikoto are scheduled to operate at all times. There will be several areas in operation at any one time as the ore and selective waste zones are drilled on a smaller 3.3 metres pattern using 115 mm holes and the bulk waste zones are spaced at 5.3 metres using 171 mm holes. Based on the required hole size and drill depth the project will use Atlas Copco D60 or similar drill rigs, equipped with DTH hammer drill bits.

Blasting will be performed under a comprehensive contract to include blast design, the emulsion and ANFO supply and storage, loading, blasting, and all other blasting associated activities.

See Table 16.4 below for the penetration rates in ore, selective waste, and bulk waste.

Table 16.4:        Blasthole Drilling Penetration Rates

Parameter	Unit	5m – Ore			5m - Waste			10m - Waste
Bench Height	m	5	5	5	5	5	5	10	10
Hole Diameter	mm	102	115	127	102	115	127	165	171
Burden	m	2.8	3.1	3.3	2.9	3.2	3.4	3.2	3.3
Spacing	m	3.1	3.4	3.6	3.2	3.5	3.7	1.2	1.3
Operational Activities
AVG. HOLE TO HOLE MOVE	min	2.60	2.60	2.60	2.60	2.60	2.60	2.60	2.60
TIME
AVG. LEVELING TIME	min	0.98	0.98	0.98	0.98	0.98	0.98	0.98	0.98
AVG. STEEL ADD TIME	min	0.91	0.91	0.91	0.91	0.91	0.91	0.91	0.91
AVG. STEEL RETRACT TIME	min	1.17	1.17	1.17	1.17	1.17	1.17	1.17	1.17
AVG. STEEL RACK TIME	min	0.72	0.72	0.72	0.72	0.72	0.72	0.72	0.72
AVG. JACK RETRACT TIME	min	0.65	0.65	0.65	0.65	0.65	0.65	0.78	0.78
Outputs
Total Drill Cycle Time	min	21.43	17.19	18.18	21.49	17.23	17.87	28.81	29.60
Total Hole Length	m	5.8	5.8	5.8	5.8	5.8	5.8	11.2	11.2
Pure Penetration Rate	m/hr	24	34	31	24	34	32	31	30
Drilling Time	min	14.41	10.17	11.16	14.47	10.21	10.85	21.66	22.45
Non Drilling Time	min	7.02	7.02	7.02	7.02	7.02	7.02	7.15	7.15
Total Time per Hole	min	21.43	17.19	18.18	21.49	17.23	17.87	28.81	29.60
Overall Drill Factor	%	67%	59%	61%	67%	59%	61%	75%	76%
Operator Efficiency	%	90%	90%	90%	90%	90%	90%	90%	90%
Overall Penetration Rate	m/hr	14.5	18.1	17.1	14.5	18.1	17.5	21.0	20.5

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16.7.2 ORE AND WASTE LOADING AND HAULING

The Otjikoto gold mine is planned to operate as a large open pit operation transporting at steady state 3.8 Mtpa of ore either to the plant or a stockpile or up to 17.1 Mtpa of waste rock of which the total does not exceed 19.2 Mtpa. A life of mine of 12 years with an additional two years of plant feed from the LG stockpile is expected for the Otjikoto Project. To achieve this various truck simulations were run for ore and waste hauling for deriving the most cost-effective and applicable ground handling solution.

Through the course of previous studies, various equipment combinations were assessed through trade-off studies with the focus on primary equipment. Various truck sizes in combination with different shovel sizes from a number of different OEM’s were analysed. The trucks ranged from 40 t articulated dump trucks to 100 t rigid frame dump trucks. The shovels traded-off were done with three different shovels, 120 t, 190 t and 250 t class shovels. For application purposes, the backhoe shovels would load the 5 metre ore and selective waste benches whilst the face shovels would be used for loading the 10 m bulk waste benches.

The suggested load and haul fleet for in-pit loading and hauling was selected by considering the results from the equipment trade-off study and the feasibility study. The costing was based on updated quotes from all OEM’s and the load and haul cycle study.

Although bulk waste will be mined on 10 m benches while ore and selective waste are mined on 5 metre benches, the Otjikoto mine will use similar equipment for both areas. This is to simplify the maintenance, scheduling, training, and other aspects of fleet operation. Loading equipment will consist of 180 tonne class hydraulic excavators (Cat 6018 or similar). One will be in the traditional backhoe configuration for mining steeply dipping areas in ore and selective waste. The other two excavators will be in a front shovel configuration for loading bulk waste and areas in selective waste and ore with wider and steeper ore and waste zones. Additionally, the mobility of a Cat 992 front end loader will alleviate any shortfalls in loading capacity.

100 t class rigid diesel haul trucks (Cat 777 or similar) will be the only haulage trucks on site. Based on the productivity, capital cost, and equipment availability (notably tires) the 100 t class trucks are the most effective haul trucks for the Otjikoto open pit mine.

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Table 16.5:        Global Equipment Parameters

Calendar Assumptions	Description	Unit	Values
Calendar Time	CTYEAR	days / year	365.0
		hours / year	8760.0
Lost Operating Time per year (i.e. Public Holidays, Planned Site Shutdown etc.)	LOTYEAR	days / year	3.0
		hours / year	72.0
Weather Delays per year	WDYEAR	days / year	5.0
		hours / year	120.0
Operating Standby per Year	OSYEAR	hours / year	192.0
Working Time per Year	WTYEAR	hours / year	8568.0
Shift Breakdown
Number of Shifts per day	Shifts		3
Scheduled Hours per Shift	Hours	hours / shift	8,0
In Shift Delays (minutes)
OS=Operating Standby (i.e. Equipment not running)
OD=Operating Delay (i.e. Equipment running but not productive)
Shift Change/Start-Up1	OD GENERAL	mins / shift	30.0
Lunch Break2	OSTRUCK	mins / shift	0.0
	OS SHOVEL	mins / shift	0.0
	OSSUPPORT	mins / shift	0.0
In-Cab Shift Breaks	OSGENERAL	mins / shift	9.0
Equipment Inspection	OSGENERAL	mins / shift	9.0
Fuelling/Service	OSGENERAL	mins / shift	15.0
Blasts per Week		#	2.0
Delay per Blast		Mins	120.0
Blast Delay per Shift	ODGENERAL	mins / shift	5.4
Safety / briefing / Ad Hoc Meetings	OSGENERAL	mins / month	268.0
		mins / shift	3.0
Inter-bench Movement - Large Loading Unit3	ODLARGE LOADING UNIT	mins / shift	3.0
Inter-bench Movement - Small Loading Unit4	OD SMALLLOADING UNIT	mins / shift	3.0
Manoeuvre Delay - Drill Rig	ODDRILLING	mins / week	60.0
		mins / shift	3.0
Face Preparation - Large Loading Unit	ODLARGE LOADING UNIT	mins / shift	3.0
Shift Utilisation
Available Time per shift	ATSHIFT	mins / shift	480
Total Operating Standby per Shift	OSGENERAL	mins / shift	45
	OSTRUCK	mins / shift	36
	OSSHOVEL	mins / shift	45
	OSSUPPORT	Mins	45
Equipment Specific Operating Standby5	Trucks	Minutes	0.0
	Loading Unit	Minutes	0.0
	Drill	Minutes	0.0

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Calendar Assumptions	Description	Unit	Values
	Front End Loaders	Minutes	0.0
	Dozers	Minutes	0.0
	Graders	Minutes	60.0
	Water/Diesel Trucks	minutes	111.0
	Wheel Dozers	Minutes	30.0
Total Utilised Time per Shift (UT SHIFT)6	UTSHIFT = (ATSHIFT - ∑OSSHIFT )
	Trucks	Mins	444
	Loading Unit	Mins	435
	Drill	Mins	435
	Front End Loaders	Mins	435
	Track Dozers	Mins	435
	Graders	Mins	375
	Water Trucks	Mins	324
	Wheel Dozers	Mins	405
Use of Availability (U% SHIFT)7	U%SHIFT = UTSHIFT / ATSHIFT
	Trucks	%	93%
	Loading Unit	%	91%
	Drill	%	91%
	Front End Loaders	%	91%
	Track Dozers	%	91%
	Graders	%	78%
	Water Trucks	%	68%
	Wheel Dozers	%	84%
Shift Operating Efficiencies
Operational Efficiency		operating mins/hr	53.0
Performance Loss	PL	mins lost per shift	55.8
Operating Loss8	OL = ∑OD + PL
	Trucks	Mins	91.4
	Loading Unit	Mins	97.4
	Drill	Mins	133.0
	Front End Loaders	Mins	91.4
	Track Dozers	Mins	91.4
	Graders	Mins	91.4
	Water/Diesel Trucks	Mins	91.4
	Wheel Dozers	Mins	91.4
Net Operating Time9	NOT = UT – OL
	Trucks	Mins	345.6
	Loading Unit	Mins	330.6
	Drill	Mins	295.0
	Front End Loaders	Mins	336.6
	Track Dozers	Mins	336.6
	Graders	Mins	276.6

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Calendar Assumptions	Description	Unit	Values
	Water Trucks	Mins	225.6
	Wheel Dozers	Mins	306.6
Operating Efficency10	UOE = NOT / UT
	Trucks	%	77.8%
	Loading Unit	%	76%
	Drill	%	67.8%
	Front End Loaders	%	77.4%
	Dozers	%	77.4%
	Graders	%	73.8%
	Water Trucks	%	69.6%
	Wheel Dozers	%	75.7%

Table 16.6:         Typical Haul Cycle Detail

Field	Description
Pit	1
Pushback	1
Bench	1500
Material	Ore
Destination	Crusher
Cycle Time (min)	16.7
Cycle Distance (m)	3680
Fuel (l)	9.78
Haul Segment	Distance (m)	Time (min)	Fuel (l)
Queue time at loader 1	0	1.50	0.23
Spot time at loader 1	0	0.55	0.23
Loading time 1	0	4.05	0.61
Loading area	50	0.26	0.22
In pit	370	0.73	0.60
Ramp	100	0.42	0.74
Switchback	60	0.26	0.17
Flat	0	0.00	0.00
Flat	160	0.36	0.37
Traffic Circle	120	0.63	0.47
Flat	500	0.84	0.87
Ramp	170	0.71	1.24
Flat	260	0.52	0.54

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Figure 16.8:         Otjikoto Primary Fleet

16.7.3 SECONDARY AND SUPPORT EQUIPMENT

These are equipment units whose function falls outside that of the primary production equipment’s scope, but which the mine would not function without. Primary production cost is directly impacted by a number of aspects related to ancillary equipment. Support equipment is the lifeline of reliable and cost-effective mining production by supporting the primary production equipment with the following activities:

• Keeping loading, tipping and haul road areas clean thus prolonging tyre life and making the operation safe;
• Maintaining haul road conditions thus prolonging tire life and making the operation safe;
• Suppressing dust emissions from a health, safety, environmental and financial perspective;
• Supporting the full equipment maintenance and diesel requirements to remote track- propelled equipment and breakdowns;
• Bench preparation and leveling;
• Fuelling of track mounted equipment and dump trucks; and
• Rehabilitation – Track Dozers.

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The collective support fleet cost per tonne is relatively low within overall mining cost. The secondary equipment will be discussed in this sections after which the total operating costs and the secondary equipment fleet profile will be estimated. The Secondary equipment fleet was also selected from results of a previous trade-off study. The required secondary fleet is shown in Figure 16.7 below.

Figure 16.9:         Secondary Fleet

In addition to the secondary fleet a support fleet has been included in project evaluation. This includes service and lubrication trucks, tyre handler, compactors, rock breakers, light stands, cranes, skid steer loaders, forklifts, light delivery vehicles, and other necessary equipment.

16.7.4 SURFACE AND PIT HYDROLOGY

The Otjikoto pit and waste dump are in a dry area with no permanent or seasonal channels, but with short duration intense rainfalls during the rainy season. Therefore an adequately sized diversion channel will carry any surface water flows around the pit to a settling pond at the south-eastern limit of the pit.

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Excess water accumulating in the pit due to groundwater drainage and rain will be removed through in pit dewatering and will be pumped to a return water dam. The maximum pump capacity was estimated based on the inflow of 3,000 cubic metres expected during the unlikely event of a flood. There is always going to be a sump located at the lowest level to store inflow water temporarily before pumping.

16.8 PRODUCTION SCHEDULE

Due to the geological setting, the production schedule and stockpile have significant impacts on the Otjikoto Project NPV. Additionally, the total mining capacity directly affects the mine fleet capital costs. Therefore several production scenarios were reviewed to determine the most economic and operationally sound solution.

The primary goals of the selected production scenario are to limit mine fleet capital while maintaining annual gold production above 100,000 ounces. Total fleet capacity is limited to 19 million tonnes of total mining per year. The strategy to defer waste stripping whilst still utilizing the value adding strategy of the 0.6 g/t stockpile cut-off grade was also used to achieve these goals. This caused pushback life to be as low as possible while overlapping as little as possible due to reduced waste stripping. The low overlapping makes the schedule more practical. Whilst the low production target adds the benefit of a reduced fleet requirement, especially during ramp-up, the high cut-off grade adds to the project value. The production plan is shown in Table 16-7 below.

The pushback release for the deferred waste and 0.6 g/t stockpile cut-off grade schedule is illustrated in Figure 16.8 Below.

Figure 16.10:         Otjikoto pit pushback schedule

This schedule allows mining to take place in only 2 pushbacks at a time and as little as possible overlapping between the 3 pushbacks. Additionally, this schedule results in gold production above 100,000 ounces annually for the first five calendar years of production. The ROM and low grade stockpiles will be milled during years 6-10 and at the end of mine life.

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Table 16.7:         Mine Production Schedule

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16.9 MINE PERSONNEL SCHEDULE

Mine personnel estimates include both operating and salaried-staff personnel. Operating personnel are estimated as the number of people required to operate trucks, loading equipment, and support equipment to achieve the production schedule. Mine staffing has been based on the people required for supervision and support of mine production.

There are numerous forms of shift schedules in the surface mining industry, all with their associated merits and shortfalls. After analysing several shift schedules, it has been determined that a 3 panel full calendar shift schedule with 3 eight hour shifts is the most effective for the Otjikoto Project. This schedule maximizes annual available operating time, which cascades down to the operator and maintenance labour complement resulting in fewer workers per shift to keep up with the higher production rate. The management, technical services, and day maintenance personnel will work an extended shift to overlap with the three shifts. The labour requirements over the life of mine have been determined with detailed hour and shift calculations, resulting in the labour complement shown in Figure 16-9 below.

Figure 16.11:         Otjikoto mine labour over the life of mine

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For the purpose of estimating the labour costs, the Patterson grading scale, which is widely used in southern Africa surface operations, was applied. As shown in Table 16-8, each position was assigned a grade that corresponds to a salary band as reflected in the table.

These salary bands, include the basic rate, overtime and shift payments, payroll burdens, employee benefits including health schemes, annual and long service leave and pension fund contributions. Travel allowance is not included in the total cost to company.

Table 16.8:         Patterson job grades with associated “cost to company” bands

Level Description	Job Level	U$	N$
		Annual Cost to Company	Annual Cost to Company
Unskilled/Virtually	A1	8,554	68,429
Semi-Skilled Unskilled	A2	10,024	80,190
Operators	B1	23,415	187,324
Supervisors	B2	30,232	241,853
Tradesmen	C1	31,515	252,117
Junior Professionals	C2	51,121	408,969
Senior Professionals	D1	87,461	699,684
Management (Local)	D2	126,818	1,014,542
Senior Management (Local)	E1	163,561	1,308,487
Senior Management (Local Expat)	E2	294,827	2,358,612

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17 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 and a plant availability of 94%. A 25% design factor was included for sizing the primary crusher, conveyors, ball mill, thickeners, cyanide destruction circuit, reagent systems and mainstream pumps which would facilitate a future expansion. The SAG mill circuit would only require the addition of a pebble crusher for the expansion.

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 cyanidation
• Cyanide leaching of gravity tailings
• Carbon-in-Pulp (CIP)
• Cyanide destruction
• Tailings disposal
• Acid wash and Elution
• Electrowinning and Gold Room
• Carbon regeneration
• Reagents make-up and distribution
• Air services and Plant water service

17.1 PROCESS DESIGN CRITERIA

The metallurgical testing presented in Section 13 “Mineral Processing and Metallurgical Testing” of this report as well as the DFS provided the data to finalize the process design criteria and the Otjikoto mill flow sheet. The major process design criteria are presented below in Table 17.1. Please note the XR3 pyrrhotite dominant ore type is listed for the crushing, grinding, gravity recovery, leach and cyanide consumption criteria as this one type was used as the basis for most of the mill design because of its physical, chemical and mineralogical characteristics and it is also the predominant ore type in the Otjikoto deposit. The detailed process design criteria is presented in the Process Design Criteria in the DFS.

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Table 17.1:         Otjikoto Major Process Design Criteria

Table 17.1: Otjikoto Major Process Design Criteria
Criteria	Unit
Operating Year	d	365
Annual Mill Throughput	Mtpa	2.5
Feed Grade (Life-of-Mine)	g/t	1.42
Crusher Availability	%	65.0
Crushing Throughput	mtph	440
Crushing Work Index (XR3)	kWh/t	16.9
Mill Availability	%	94.0
SAG Mill Feed Size, 80% Passing	µm	150,000
Grinding Throughput	mtph	304
Ball Mill Circulating Load	%	300
Bond Ball Mill Work Index (XR3)	kWh/t	9.3
Bond Abrasion Index (XR3)	g	0.466
Grind Size, 80% Passing	µm	75
Gravity Gold Recovery (XR3)	%	70.8
Leach Residence Time (XR3)	hr	48
CIP Residence Time	hr	6
Cyanide Consumption (XR3)	kg/t	0.59
Overall Gravity+Leach/CIP Recovery (Life-of-Mine)	%	95.6
Average Annual Gold Production (years 1-5)	oz	141,000
Average Annual Gold Production (LOM)	oz	112,000

17.2 PROCESS DESCRIPTION

The simplified Otjikoto gravity/whole ore leach flow sheet is shown below in Figure 17.1. The mill flow sheet consists of the following unit operations:

Ore receiving and Crushing
The crushing circuit has a 42 inch x 65 - inch (1,067 mm x 1,651 mm) gyratory crusher, with a nominal throughput rate of 440 tph. Run-of-Mine ore from the open pit operations will be delivered by 100 tonne trucks to the crusher. The crushed product at a P80 of 150 mm, is conveyed to a 7,000 tonne (live capacity) mill feed stockpile.

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Grinding Circuit
The grinding circuit consists of a 24 ft diameter x 14.5 ft EGL (7.11 metres ∅ x 4.42 metres EGL) SAG mill operated in closed circuit with a trommel for pebble screening on the mill discharge and a vibrating screen to produce ball mill circuit feed. The SAG mill product is combined with tailings from the gravity concentration circuit before being pumped to a classification cyclone cluster. Cyclone underflow is fed to a 16.5 ft diameter x 28 ft EGL (5.03 metres ∅ x 8.53 metres EGL) ball mill and the entire ball mill discharge is fed to a gravity concentration circuit for recovery of coarse free gold. Both grinding mills are powered with 4,000 kW (5,364 HP) synchronous motors. The circuit is designed to produce an 80%

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Figure 17.1:         Otjikoto Mill Flowsheet

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passing product size (P₈₀) of 75 µm. The SAG mill circuit design includes a future allowance for the installation of a pebble crusher

Gravity Concentration and Intensive Leaching

The gravity concentration circuit is comprised of four (4) Knelson 48 inch (1,219 mm) concentrator units which will be operated in parallel. The gravity concentrate collected from the Knelsons will be stored and then processed by intensive cyanide leaching in a ConSep CS8000 Acacia Reactor on a batch basis. The design allows for processing of batches on either a 24 hour or a 48 hour basis.

Cyanide Leach Circuit

Cyclone overflow from the grinding circuit is thickened in a pre-leach thickener to achieve a solids content of 45% in the underflow product prior to cyanidation. The leach circuit is comprised of one pre-aeration tank with a residence time of 6.9 hours and seven leach tanks with a total residence time of 48 hours.

Carbon-in-Pulp Circuit

The CIP circuit is comprised of six CIP tanks in series with a design residence time of one hour per tank. The circuit is designed for carbon concentrations in the range 15 -25 g/L with loaded carbon batch transfers of 5 tonnes every 24 hour cycle.

Cyanide Destruction

The Otjikoto cyanide destruction circuit is based on the SO2-Air process and is comprised of a single tank with a design residence time of 90 minutes. The circuit is designed to reduce weak acid dissociable cyanide (CNwad) concentrations to less than 10 ppm.

Tailings Thickening and Disposal

Product slurry from the cyanide destruction circuit is thickened in a tailings thickener to achieve an underflow density of 55%-65% solids. Thickened tailings are pumped to a tailings impoundment and water reclamation facility situated approximately 3 km from the mill.

Acid Wash, Elution and Carbon Regeneration

Loaded carbon is processed in 5 tonne batches on a 24 hour cycle for recovery of gold. The loaded carbon is treated in an acid wash followed by elution and regeneration of barren carbon in a rotary kiln. The design of the elution circuit is based on the pressure split Anglo-American Research Laboratories process (Pressure split “AARL”). Gold is recovered in the eluate stream during the elution process.

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Electrowinning and Gold Room

The electrowinning circuit treats eluate from the carbon elution circuit and pregnant solution from the Acacia Reactor on a batch basis for the recovery of gold from solution, using atmospheric sludging electrowinning cells. Electorwinning sludge will be fluxed and smelted in an induction furnace to produce gold bars for shipment to a refinery.

Reagent Make-Up and Distribution

The reagent make-up and distribution facilities cater for the make-up, storage and distribution of the following reagents:

• Sodium Cyanide
• Caustic Soda
• Hydrated Lime
• Lead Nitrate
• Hydrochloric Acid
• Sodium Metabisulphite (“SMBS”)
• Copper sulphate
• Flocculant
• Elution De-Scalant
• Process Water Anti-Scalant

Table 17.2 presents a summary of the mill reagent consumptions as a function of the XR1, XR2 and XR3 ore types.

Table 17.2:         Mill Reagent Consumptions by Ore Type

Reagent	Ore Type
	XR1	XR2	XR3
Sodium Cyanide (kg/t)	0.55	0.69	0.59
Caustic Soda (kg/day)	30	30	30
Hydrated Lime (kg/t)	1.29	1.18	0.82
Lead Nitrate (g/t)	100	100	100
Hydrochloric Acid (t/d)	0.83	0.83	0.83
Sodium Metabisulfite (kg/t)	0.70	0.83	0.75
Copper Sulphate (g/t)	50	126	155
Flocculant (g/t)	40	20	20
Elution De-Scalant (g/t)	5	5	5
Process Water Antiscalant (g/t)	14.8	14.8	14.8

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Plant Services

The design includes facilities for the generation, storage and distribution of the following air services:

• Plant and instrument air at a design pressure of 7.5 Bar.
• Leach air supply at a design pressure of 3.0 Bar
• Cyanide destruction air supply of 3.0 Bar.

The design includes storage and distribution facilities for the following plant water services:

• Process Water
• Reclaim Water
• Fresh Water
• Fire Water
• Potable water
• Gland Service water

The mill fresh water consumption is projected at 1 Mm³/a at the mill design throughput of 2.5 Mtpa. Fresh water will be supplied from wells on the Otjikoto property for both potable and process needs. Average overall plant power consumption during steady state mill operation will be approximately 32 kWh/t of ore processed. Electrical power will be generated on site using heavy fuel oil generators.

17.3 LABORATORY

Jack Stanley, owner of jsAnalytical Laboratory Consultants Ltd., designed a laboratory to support the exploration, mining and milling functions of the Otjikoto Project operation. The scope of work included the following:

• Laboratory layout
• Staffing Requirements
• Equipment Selection and Costing
• Design of the Ventilation System

The lab design provides for the analysis of up to 250 samples per day, the primary element being gold and secondary elements being Fe, S, and Ag but not limited to these. Analysis consists of preparation equipment and methods that take into consideration the occurrence of free gold, gravimetric methods, wet chemistry procedures and instrument methods, using Atomic Absorption, MP-AES, and LECO.

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The major components of this laboratory are the sample preparation lab, fire assay lab, wet chemistry lab, metallurgical lab, low level lab (environmental) and the instrument room. Support areas include rooms for mechanical, electrical, gas bottle storage, washrooms and offices.

The detailed laboratory design report can be found in the Laboratory Design Report of the DFS.

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18 PROJECT INFRASTRUCTURE

This section will provide an overview of the general site infrastructure for the Otjikoto Project along with the G&A staffing. It will cover site preparation, layout, buildings and facilities, tailings, waste, and water management, health, safety, environment, security, and G&A capital and operating costs.

The Otjikoto Project will consist of an open pit surface mine feeding an ore processing facility with a capacity of 2.5 Mtpa. The thickened tailings will be pumped to a fully HDPE lined paddock type storage facility with processing return water and storm water collection ponds. See Figure 18.1 for a general facilities arrangement and Figure 18.2 for the construction schedule.

Figure 18.1:         General facilities arrangement

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Figure 18.2:         Construction Schedule

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18.1 SITE LAYOUT AND PREPARATION

The site has been arranged to minimize transportation distances while maintaining sufficient room for planned operations and potential expansions. The open pit is in the north-western portion of the mine area, immediately adjacent to the primary crusher, ROM stockpile, low grade stockpile, and mine waste facility. The processing plant has been situated to the south of the mine waste dump to allow for potential pit expansion to the south, increase access from the mine access road, and reduce dust from mine operations. The mine facilities are located closer to the open pit to reduce travel time between the mine and the mine facilities and to separate light vehicle traffic from heavy mine equipment.

Before construction of the process and mine facilities commences, the sites and buffer areas will be appropriately prepared. The first step is to remove the invasive thornbush and other woody vegetation. This will be followed with a standard clearing and grubbing, then topsoil will be stockpiled. All civil areas will be graded prior to any compaction or detailed earthwork for concrete foundations. In areas with dynamic or other heavy loads the nodular calcrete will be removed and replaced with an engineered foundation fill.

The overall industrial area will be fenced and entrance points will be manned with security guards. In addition, the administration building, processing facilities, and mine facilities will have appropriate higher security fencing and access control.

18.2 INTERNAL AND ACCESS ROADS

Mine access will be from the B1 national highway directly to the administration area via the dedicated, private mine access road. All roads will be constructed to purpose based on the expected traffic, and will be properly graded and drained with culverts as needed based on local drainage patterns.

As the Otjikoto mine design was developed it was recognised that the existing district road (the D2808) intersected both the northern area of the planned pit and the waste rock dump. In Namibia, district roads are secondary rural roads, most often made of in-situ material (dirt surface). In conjunction with the RA and in compliance with the Namibian Roads Ordinance 17 of 1972 a formal application for the deviation of a portion of the D2808 was applied for in early 2012. Following prescribed legislation a public consultation process was completed as well as public advertisement of the intended deviation. No objections were recorded by the RA, who independently managed the process. The RA Board approved the deviation in September 2012 and Ministerial approval was granted in early 2013. This ministerial decision was publicized on February 13, 2013. A portion of the D2808, approximately 14 km in length, will be relocated approximately 4 km north of its current location, and will thus circumvent the operational area. The mine access road will be a 7 metre wide gravel road, apart from a 150 metre length at the B1 highway which will be asphalt to match the B1 highway and reduce dust. The mine access road will be approximately 2 km long to the administration building and processing facilities and an additional 1 km from the processing facilities to the mine facilities. The mine access road will be provided with required road signs, notably at the rail level crossing and approaching the intersection with the B1 highway. The project also provides for widening the B1 highway for at least 240 metres either side of the mine access road centreline. The provision of these works is premised on the assumption that the road intersection permit application submitted by B2Gold Namibia will be approved by the relevant road and rail authorities.

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Site roads will be sized according to the expected traffic and use. All roads near and connecting the administration, processing and mine facilities will be sized appropriately for two way traffic consisting of heavy on-road delivery vehicles and standard light vehicles and constructed using locally available gravel. Remote roads such as pipeline and borehole access will be single lane with pull-outs and will be constructed on native ground with 150 mm of gravel when necessary.

18.3 BUILDINGS AND FACILITIES

The layout of the buildings is organized to separate the various functions and maintain secure areas for the independent functions while keeping all facilities within a manageable distance. The general facilities consist of:

• Administration complex, with facilities for overall site management, safety inductions, and general and administrative functions. See Figure 18.3.
• Processing facility complex, containing the grinding and leaching facilities, along with management and engineering offices, change house, workshop, warehouse, and assay laboratory facilities. See Figure 18.3.
• Power generation facility, within the processing facility limits but within an additional layer of security/safety fencing and access control. The fuel storage area (HFO and diesel) will be adjacent to the power generation facility. See Figure 18.3.

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Figure 18.3:         Processing and Administration facility

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Mine facilities will consist of management and engineering offices, change house, EMV workshop and light vehicle workshops, wash bay, and warehouse. See Figure 18.4. Other smaller facilities will include the explosives magazine, crusher, mine access gate house, return water pump house, and any other necessary structures.

18.4 TAILINGS MANAGEMENT

A scoping level trade-off study was completed in Q2, 2012 to determine the cost and water consumption impact of conventional, thickened, or filtered tailings. It was determined that thickened tailings would present the best balance of water consumption and cost. The thickened option results in a tailings slurry consisting of 55% solids.

Following this analysis, SRK Consulting was contracted to perform and manage the geotechnical investigation of the civil works, including the tailings facility area. Epoch was contracted to complete the feasibility level design of the tailings storage facility.

Geotechnical test pit logging has shown the presence of nodular calcrete in discrete areas throughout the site. This will require field adjustment to the earthworks and foundations of the tailings facility, which has been taken into account in the designs and costs of the project.

The tailings storage facility proposed for the Otjikoto mine is to be constructed using upstream construction techniques; in particular the paddock system is to be employed. The paddock system of tailings facility wall building is common throughout the gold extractive industry in southern Africa. In this system the tailings are allowed to desiccate to create a consolidated and strong material as the outer wall of the disposal facility is raised above the inner beach. A penstock system will be constructed in the center of the paddock to direct return water to the return water pond as quickly as possible. The return water dam will be constructed within the larger storm water dam, which has been designed with a runoff system to contain the 50 year storm event. See Figure 18.5.

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Figure 18.4:         Mine facilities

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The tailings storage facility has been designed to contain 36 million tonnes at a deposition rate of 3.0 million tonnes per year. The storage facility and ponds will be fully lined with HDPE liner. Monitoring wells and liner under drains will be installed and regularly sampled.

18.4.1 ENGINEERING ANALYSIS

The following engineering analyses were completed for the design of the tailings storage facility and related works:

• Slope stability in various drainage and phreatic conditions.
• Seepage analysis in various drainage and phreatic conditions.
• Hydrological runoff analysis for storm water control design.

The above engineering analyses were based on:

• Geotechnical samples collected by SRK.
• Tailings samples from the processing analysis.
• Hydrologic parameters from Otavi rainfall records (Peens and Associates), see Figure 18.6.

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Figure 18.5:         Tailings Storage facility

The results of the analyses indicate that the facility will be stable under normal and undrained cases during operations and closure.

18.4.2 REHABILITATION

The side slopes will be topsoiled and vegetated on an on-going basis as each slope bench is completed, using topsoil stockpiled during construction. At closure the final tailing surface will be graded to drain, topsoiled, and vegetated. The penstock intakes and outlets will be sealed and any exposed piping and valves will be removed.

18.5 HYDRAULIC DESIGNS

The Otjikoto Project is in an arid location with rainfall occurring only during the rainy season, most significantly December through March. Therefore the hydraulic designs and water management plans focus on isolating the storm water from the process water and minimizing water use. There are no permanent or seasonal waterways on the site, but sheet flows and small channel flows can occur during high intensity/short duration storms.

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The site topography slopes gradually to the west, towards the B1 highway. Therefore storm water and any other non-contacted water that flows outside of the disturbed areas will be diverted around the facilities with a system of storm water ponds, diversion, and culverts. This system will direct the flow to the existing channels and culverts associated with the B1 highway.

All water falling directly on the industrial areas (contacted water) or otherwise contacted (fissure water from the mine pit, return and storm water from the tailings facility) will be stored and used in the mining and processing facilities.

18.5.1 STORM WATER DAMS

There are plans for two storm water dams. The largest is designed with a capacity of 14,000 m³ to contain all water falling on the processing facility terrace during a 1:50 year rainfall event. The smaller storm water dam provides capacity of 4,000 m³ and will contain all runoff from the mining facilities. Both storm water dams will include a spillway and silt traps.

18.5.2 WATER STORAGE DAMS

There will be two water storage dams. The first will be the reclaim process water dam, which will receive water from the tailings facility and will contain 4,000 m³ in an HDPE lined pond. Water pumps will supply water to the processing plant. The second water storage dam is the pit dewatering dam. Pit water will be pumped from the in-pit sump and dewatering bore holes to the pond near the southeast pit crest. The dam will have a capacity of 3,500 m³ and will provide water for dust suppression and mineral processing.

18.5.3 GROUNDWATER HYDROLOGY AND DEWATERING

A numerical groundwater model was developed by SLR in 2011 and 2012 to estimate inflows to the mine pit and analyse impacts of the pit inflows and borehole pumping on the local and regional water resources. The data and conclusions of this model were applied to the design of the mine dewatering system, the processing facility, and were referred to for the various permits required for construction and operation of the Otjikoto Project.

The numerical model was based on various data from borehole records of the Namibia Department of Water Affairs and measurements taken by AGES and SLR during hydrocensus activities in 2008 and 2011. See Table 18.1.

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Table 18.1:         Monthly Rainfall

	Oct	Nov	Dec	Jan	Feb	Mar	Apr	May	Jun	Jul	Aug	Sep	Total
Mean Rainfall (mm)	20	50	79	123	138	92	36	6	1	0	1	3	688
Mean Wet Year Rainfall (mm)	29	60	98	145	183	113	49	6	1	0	1	3	870
Mean Dry Year rainfall (mm)	6	33	51	88	92	60	24	0	0	0	0	0	445

18.5.3.1 Regional and Site Hydrogeology

The aquifer system of the Otjikoto hydrogeology study area consists of three major units: the Karibib marble formation, the surrounding and underlying quartz biotite and graphitic schists or the Okonguarri formation, and the calcrete layer on top of the Okonguarri. The Karibib marble formation and the associated groundwater system are enclosed and bordered by relatively impervious rock units with decisively different geological and hydrogeological characteristics. Depth to groundwater measurements in numerous locations shows steep gradients near the boundaries of the formations while the flow gradient is flat within the Karibib marble aquifer. Flow within the Karibib marble aquifer is from the NE to the SW, parallel to the strike of the marble outcrop. In the biotite and graphitic schists the flow is directed to the NW with a steep gradient across the geological strikes towards the center of the Platveld basin. Due to the significant water level and flow direction difference between the marble and the schist it is assumed that flow from the marble in the east to the biotite schist in the west is limited.

Near the surface it is assumed that a discontinuous perched aquifer exists in the calcrete. These small discrete aquifers are not directly connected to the deeper regional aquifer due to the impermeable layers and are fed by recent rainfall.

18.5.3.2 Groundwater Model Setup

The numerical groundwater model was developed for the local area using the modelling package Feflow 5.4 (www.feflow.info). The model uses 203,652 elements and 137,116 nodes. The model domain covers 665 km² (Figure 18.6) . Rivers and faults are included explicitly, and other important modelling zones were delineated. The model context is based on existing 1:50,000 topographical data and 1:250000 geological data. Hydrogeological parameters are based on hydrocensus data and additional drilling of 25 boreholes that were subjected to aquifer test to determine the representative parameters for each of the local aquifer units. The model was calibrated to 93 observation boreholes by varying the hydrogeological parameters of the various units and boundaries.

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Figure 18.6:         Groundwater Model

18.5.3.3 Mining and Post-Mining Simulations

After calibration, the general mine development plan was added as input to the model to determine the pit inflows as the pit develops over approximately 11 years. The radius of influence extends between 4 km and 4.5 km to the west, east, and south. To the north the radius of influence extends to 7.5 km. The pit inflows increase gradually up to a maximum of roughly 5600 m³/day, but could be far less depending on site specific evaporation. The dewatering system has been sized to pump at least 6000 m³/day.

After mining is completed the pit will form a shallow pit lake with a surface water elevation between 1,390 and 1,430 metres, depending on evaporation. The lake will be primarily fed by groundwater. Due to the radius of influence any potential contaminants from the tailings facility, mine waste dump, or processing plant will flow towards the pit lake.

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18.5.3.4 Mine Dewatering

A detailed mine dewatering scenario has not been completed, but it is expected that dewatering will be effective starting from pit depth of 20 to 30 metres below ground surface. Groundwater inflows are expected below pit depths of 40 metres via small fissures. Vertical and/or horizontal wells would be effective to reduce the local phreatic levels in the pit walls to increase pit stability.

18.5.3.5 Site Water Balance

The groundwater model, tailings facility water balance, and processing facility all serve as inputs to the site water balance. Tables 18.2 and 18.3 show water balance scenarios with and without water from mine pit dewatering boreholes. Based on these results the mine will require between 2,461 m³/day and 3,315 m³/day for normal operations. This makeup water will be sourced from pit dewatering wells and water supply wells in the Karibib aquifer. According to the Otjikoto Groundwater Flow Model Update (SLR Feb 14 2013) the lower make up water limit (roughly 900,000 m³/year) is sustainable indefinitely and the maximum amount (roughly 1,300,000 m³/year) is sustainable for at least five years.

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Table 18.2:         Water balance with pit de-watering

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Table 18.3:         Water balance without pit de-watering

18.6 WASTE AND WASTEWATER MANAGEMENT

Self-contained, containerized wastewater processing units will be installed at the construction camp during construction and will be gradually relocated to the permanent facilities. Additional units will be installed as needed. The outflows will not require further processing or treatment and will serve as process water for the mineral processing facility.

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A policy will be established to minimize usage and maximize recycling of domestic wastes such as paper, aluminium, glass, plastics, etc. and receptacle will be provided throughout the facilities. Instructions will be given to all personnel (including cleaning staff) in the use of these receptacles.

Dangerous or hazardous waste will be collected and stored briefly before being transferred to a suitable permitted facility, either on-site or off-site depending on the specific materials and requirements.

18.7 POWER SUPPLY AND DISTRIBUTION

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

The generators will supply power to the plant at 6.6 kV rather than the more widely used 11 kV to permit a bypass facility around the SAG mill variable speed drive 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, plus the capacity to start both mills; this figure to exclude at least one generator and include a 10% margin. The generators will be fuelled with HFO, which will be transported to the Otjikoto Project site from Walvis Bay via the national highways. A storage facility with minimum ten days fuel supply has been allowed for to support the generators. The necessary fuel storage capacity is subject to revision once the fuel supply and logistics have been finalised.

18.7.1 MAIN SUBSTATION

The main substation will be a thermally insulated, fan pressurized brick structure housing a 19 panel 6.6 kV 25 kA switch board and will supply all plant, mining, and other loads. The switchboard will be supplied from the power station via two groups of feeder cables. Each group will be of adequate capacity to supply the entire site.

The 6.6 kV switchgear will be of metal clad vacuum interrupter type equipped with microprocessor protection relays and a 110 V dc tripping supply with arc detection. The tripping supply will be fed from the local 400 V minisub, direct from the main switchboard.

The plant loads consist of the mills plus 1600 kVa transformers fitted with the necessary protection and situated locally to the MCCs they are supplying. Mining loads will be supplied via two 6.6 kV overhead lines, one to the crusher area and the other to the mining facilities area. The return water pump station will also have a 6.6 kV overhead line.

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MV cables will be 6,35/11kV rated, sized for load current and for a 25 kA fault for one second. LV cables will be sized for load current and end of cable volt drop only. Cables will be fire retardant to SANS 1339 and SANS 1507 respectively. All cables will be run on heavy duty galvanised racking, generally in a vertical configuration.

18.8 DORÉ TRANSPORT

Doré will be stored on site in a locked, secure facility. The refinery company will collect the dore at the site and ship the dore to the refinery (most likely in S. Africa) via air transportation. The closest airport is Otjiwarongo but doré will most likely be shipped via Windhoek.

18.9 HEALTH, SAFETY, ENVIRONMENT, AND SECURITY

B2Gold has overall responsibility for the development of Health, Safety, Environmental (HSE) and Security policy and its implementation. B2Gold’s team will produce a project HSE plan detailing the requirements of HSE in compliance with the laws and regulations of Namibia as well as international best practices. All employees, visitors, contractors, subcontractors, suppliers, vendors, and consultants will comply with these requirements.

18.9.1 INDUSTRIAL HYGIENE

To safeguard the workers’ health hazards must be recognized and controlled. This can be accomplished by implementing engineering controls, work practice controls, and administrative controls. When effective work practices or engineering controls are not feasible or while such controls are being instituted, appropriate Personal Protective Equipment (PPE) must be used. To be effective, PPE must be individually selected, properly fitted, and periodically refitted, conscientiously and properly worn, regularly maintained, and replaced as necessary. Potential hazards can include air contaminants and chemical, biological, physical, and ergonomic hazards. To estimate the costs associated with industrial hygiene at the Otjikoto Project appropriate PPE has been included in the capital and operating costs based on the type and number of personnel on site.

18.9.2 SECURITY

Site wide security will be contracted to a reputable and capable contractor with experience in the mining industry in southern Africa. The project will have a single guarded entry point on the mine access road and will be fully fenced and monitored. Additional fencing and controlled access points will be established at each facility.

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18.9.3 FIRST AID

Due to the moderate distance (70 km) from the Otjikoto Project site to high quality emergency health care facilities in Otjiwarongo, the Otjikoto mine will have an emergency clinic and trauma room. Trained EMT personnel and ambulances will be on site at all times to respond rapidly to medical emergency situations. The emergency clinic will focus on short term care to prepare patients for transport to the hospital in Otjiwarongo.

18.9.4 TRAINING

The Otjikoto Project will use various types of training to meet both safety and performance needs. In addition to basic ethical and safety training, all employees will undergo site and task specific safety training prior to operating equipment. This training will be performed by experienced trainers or experienced operators. This training will be supplemented with continued development to enable a more skilled workforce.

18.9.5 ENVIRONMENTAL

The environmental programs are discussed in Section 20 “Environmental Studies, Permitting, and Social or Community Impacts” of this report.

18.10 GENERAL AND ADMINISTRATIVE

G&A expenses and systems are related to business operations not directly associated with production. For this study the G&A expenses have been included in each area (mine, plant, etc.) plus the overall management has been included in the design and costing of the project. This includes staff, buildings, transportation, supplies, outside services, and other costs associated with owning and operating the mine. Typical staffing is shown in Table 18.4 below.

Table 18.4:         General & Administrative Staffing

General Position	2013	2014	2015	2016	2017	2018	2019	2020	2021	2022	2023	2024	2025	2026
Administration	-	1	3	3	3	3	3	3	3	3	3	3	3	1
Human Resources	-	1	2	2	2	2	2	2	2	2	2	2	2	1
Information Technology	2	3	4	4	4	4	4	4	4	4	4	4	4	1
Accounting	3	6	8	8	8	8	8	8	8	8	8	8	6	2
Purchasing, Warehousing, Logistics	5	12	15	15	15	15	15	15	15	15	15	15	11	3
Environmental	1	2	2	2	2	2	2	2	2	2	2	2	2	1
Environmental Laboratory	-	0	1	1	1	1	1	1	1	1	1	1	1	0
Security	1	1	1	1	1	1	1	1	1	1	1	1	1	0
Safety	3	6	7	7	7	7	7	7	7	7	7	7	4	1
Medical Clinic	2	2	2	2	2	2	2	2	2	2	2	2	2	1
Employee Transportation	1	1	1	1	1	1	1	1	1	1	1	1	1	0
Cafeteria & Housing	1	1	-	-	-	-	-	-	-	-	-	-	-	-

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19 MARKET ANALYSIS

19.1 MARKET OVERVIEW

19.1.1 LATE 2000S FINANCIAL CRISIS

The late-2000s financial crisis (often called the Credit Crunch or the Global Financial Crisis) is considered by many economists to be the worst financial crisis since the Great Depression of the 1930s (Haidar, Jamal Ibrahim, 2012). It resulted in the threat of collapse of large financial institutions, the bailout of banks by national governments, and downturns in stock markets around the world. In many areas, the housing market had also suffered, resulting in numerous evictions, foreclosures and prolonged vacancies. It contributed to the failure of key businesses, a decline in consumer wealth estimated in the trillions of U.S. dollars, and a significant decline in economic activity, leading to a severe global economic recession in 2008 (Baily and Elliott, 2009 & Williams, 2012).

The financial crisis was triggered by a liquidity shortfall in the United States banking system in 2008. The collapse of the U.S. housing bubble, which peaked in 2006 (Tully, 2006), caused the values of securities tied to U.S. real estate pricing to plummet, damaging financial institutions globally (Public radio International, 2009 & Simkovic, 2011). Questions regarding bank solvency, decline in credit availability and damaged investor confidence had an impact on global stock markets, where securities suffered large losses during 2008 and early 2009. Economies worldwide slowed during this period, as credit tightened and international trade declined (International Monetary Fund, 2009).

Governments and central banks responded with unprecedented fiscal stimulus, monetary policy expansion and institutional bailouts. Although there have been aftershocks, the financial crisis itself ended sometime between late-2008 and mid-2009. While many causes for the financial crisis have been suggested, with varying weight assigned by experts, the United States Senate issuing the Levin–Coburn Report found that the crisis was not a natural disaster, but the result of “high risk, complex financial products; undisclosed conflicts of interest; and the failure of regulators, the credit rating agencies, and the market itself to rein in the excesses of Wall Street” (Levin and Coburn, 2011). Critics argued that credit rating agencies and investors failed to accurately price the risk involved with mortgage-related financial products, and that governments did not adjust their regulatory practices to address 21st-century financial markets (Declaration of G20 – Whitehouse Government, 2008). All of these economic and financial dynamics are linked to the value of the US dollar and the price of gold.

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19.1.2 HISTORIC GOLD PRICES

A graph showing the gold price over the past 15 years is given in Figure 19.1. The market correction of 2008/9 is clearly visible, but has had little impact on the general upward trend in the gold price. This bears out the market fundamentals of demand, supply and rising production cost as discussed above. A second market correction can be seen in 2011/12.

Figure 19.1:         15-year and 12-month historical gold prices (source: www.goldprice.org)

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In a North American Equity Research bulletin by J.P.Morgan of December 18, 2012, they refer to the market adjustments as periods of consolidation which historically have been followed by growth. The section relating to gold from that report is reproduced below as Section 19.2.

19.2 A FEW WORDS ON GOLD

We continue to expect gold to perform in both inflationary and deflationary times; hence, we feel the sector deserves attention. Inflationary times are the easiest ones to invest in gold and this supported the gold equities as gold moved from $250/oz to its $1,900 peak. Since September 2011, even though the Federal Reserves Bank’s balance sheet continues to increase, this has not been translating into new highs for gold.

Figure 19.2:         Gold Price Range Bound (Price in $/oz)

This could be another case of gold price consolidation prior to an upward breakout. Seasonally, gold often has been strong in January/February and September/October and it’s possible that gold could break higher in one of these periods in 2013. Any settlement of the fiscal cliff could be seen as offering financial stability and could put a cap on the gold price. However, even if a settlement is reached, delivery is likely to be difficult.

19.2.1 MARKET VOLATILITY

Figure 19.1 also shows a graph of the gold price over the last twelve months, illustrating the high volatility in the price of gold during the past year.

The following information in Section 19.3 has been taken from TD Securities Commodities Outlook 2013 which was published on December 10, 2012. This article discusses the global economics and their effect on the demand for gold and the gold price. TD Securities is predicting that the gold price will average above $1,800 per ounce in 2013 and 2014.

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The Qualified Person has reviewed these studies and believes that they support the assumptions in this report.

19.3 MACRO OUTLOOK

“Living in Interesting Times” into Early-2013, if US Goes over Fiscal Cliff  Mirroring the fear inspired by a European meltdown in late-2011, markets again look set to be obsessed with the risk of a major correction as we head into 2013. But this time, it’s the concern that the US will fall over its fiscal cliff that is responsible for investor angst. As such, TD Securities (TDS) expects fairly volatile market action into early-2013, with random fiscal cliff headlines triggering risk-off, risk-on trading, which will likely encompass the commodity spectrum.

While metal and energy markets have gone from cold, to hot, to cold, to finally finish strong in November, the knee-jerk nature of commodity trading based on headlines, technical and other non- fundamentals implies that there are no guarantees that we will finish the year on a high note. In fact, if the current political winds blow the US over the fiscal cliff and seeing no well-defined timeline for a solution at the time of writing, commodities could very well finish the year on a sour note, with this poor market condition prevailing into early-2013.

The concern is that the legislated tax increases and spending cuts, which amount to some $600 billion over the course of the year, if not reversed in the early part of the New Year could very well plunge the US in another recession. A loss of this much fiscal economic fuel for a prolonged period would not only put the US economy on to a path towards recession, but could derail the entire global economic outlook and commodity demand prospects into 2013.

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Figure 19.3:         Commodities Follow Global Growth – Moving up

More on faith rather than fact and despite the lack of any solid progress in the negotiations, markets are still pricing in some sort of dignified deal between the US President and the Republican House to the fiscal cliff problem in the near term. The view is that a solution must be found before the deadline and that failure not an option. Markets simply can’t believe that US politicians would be so cavalier as to risk a recession for pure partisan reasons. We too are in that camp.

This optimistic view, however, may be giving the US political class too much credit! If we don’t see progress in the negotiations soon, markets may very well react negatively.

Echoing a Chinese curse—“may you live in interesting times”— all this could very well be happening during very thinly traded holiday markets, which could exaggerate any potential correction.

MANY REASONS TO BUY THE DIPS AFTER US FISCAL CLIFF IS SETTLED FOR THE LONGER-TERM

An improvement in market risk appetite, along with tighter supply/demand conditions and rising investor interest, should help industrial precious metals, base metals and energy move higher in the period following the start of a new US Congressional session.

Even if the US slides over its fiscal cliff, a solution to the US budgetary problems early in 2013 should leave the economy relatively unscathed. A combination of Treasury Department administrative manoeuvres, along with the sheer fact that the fiscal drag from the Budget Control Act of 2011 would not have enough time to take a serious bite out of economic activity, implies that the US economy would resume its recovery. Of course, this will be helped by the Fed’s steady hand.

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Figure 19.4:         China Biggest Incremental Consumer

We project that the new Congress and the US President will work out a deal and avoid the worst of the impact on the real economy, which should allow US growth to trend higher. At the same time, the ECB along with Euro Zone (EZ) governments will likely take policy action, including balance sheet measures and some sort of debt load lightening, to stabilize EZ economies.

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Figure 19.5:         China Real Estate Turning a Corner

Furthermore, China should continue to get its groove back, which will help boost demand growth prospects, particularly for industrial commodities. China’s new leadership may very well add some fiscal pep to the broad domestic economy, while possible reductions of bank reserve requirement ratios (RRR) should help to boost the real estate segment. As such, demand is likely to pick up and we could see a restocking cycle commence as well, which should help metals like copper and PGMs. Oil and bulks should also get a lift.

PRECIOUS METALS OUTLOOK

Palladium and Silver to Outperform in 2013, Gold and Base Metals Look Promising Too

The powerful combination of fiscal cliff related risk-off trading along with concerns that fundamentals could deteriorate should the US suddenly stop growing are likely to see the metals complex retain its corrective bias into early-2013. Even gold—a traditional safe-haven asset—is expected to be under pressure if investors build cash positions in response to fiscal cliff tail risk.

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Figure 19.6:         Holding on Support for now

The Fed’s new bond buying program and any positive news out of China is likely to serve as support for gold and industrial metals, rather than a set of catalysts propelling prices higher, as long as, the fiscal cliff issue is not off the table. Of course, if the US central bank disappoints on December 12th, markets could get very interesting indeed.

However, once the US political community agrees on a fiscal deal, we should see the return of risk appetite and a much improved assessment on market fundamentals going forward. With western central bank stimulus, particularly the Fed, and stronger Chinese performance, the metals complex should move smartly higher, particularly precious metals.

Investors and Central Banks Working Together for a Better Gold Future

Ultra easy monetary policy, with balance sheet action in particular, has been driving gold markets ever since the Great Recession reared its ugly head. And, expectations of more monetary accommodation from the Fed, the ECB, the PBoC and other central banks should continue to drive this market to new highs in 2013.

But monetary policy is not the only way central banks around the world, especially those domiciled in nations that are up-and- coming power houses, are helping to support gold. The official sector physical buying has increasingly, with more on the way, been supporting the yellow metal. It has been a big physical buyer since 2011, with purchases of another 525 tonnes, if not more, expected next year.

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The combination of the very possible new build-out of the Federal Reserve’s balance sheet and extremely low yields should start to ignite inflation concerns when the economy moves up onto firmer ground. The Fed balance sheet is speculated to increase by $85 billion per month with MBS/Treasury purchases, which should also increase commercial bank reserves dramatically. As such, an even larger contingent of investors should look to the yellow metal as a hedge. If the US central bank stays its über accommodative course, even as growth is seen to be more solid, many in the market will likely worry that monetary policy may be behind the inflation curve.

Figure 19.7:         Inflation Expectations to Continue to Lift Gold Higher in 2013

Heightened concern that purchasing power and the real value of US dollar denominated debt obligations may erode in the future has traditionally driven investors of all stripes, ranging from retail to sophisticated institutional players, into gold. One only needs to look at the ETF market to confirm this—market segment holds some 84.9 million oz (2,630 tonnes), up some 8.5 million oz (265 tonnes) from the end of last year.

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Figure 19.8:         Broad USD Index Short term Issue for Gold – Big Boost longer-term

Figure 19:9:         Gold ETF Holdings

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Given the recent investor behavior and that the above mentioned conditions continue to be in play again in 2013, investors of all stripes are likely to be driven more into gold next year, and we expect to see a more robust use of the yellow metal as a hedge. This increase in demand is likely to continue to drive gold prices materially above its marginal cost of production—representing an “insurance” premium where markets are willing to pay for protection.

The recently introduced QE3 has already lifted inflation expectations and has contributed to a fairly positive price dynamic—a trend that looks set to continue. We project that gold will jump close to $2,000/oz in the latter part of 2013.

Figure 19.10:         More Fed Bond Buying Spells Higher Gold Price, eventually

Stronger economic activity should boost confidence, lifting the consumers’ willingness to borrow and the banks’ appetite to lend. With more lending and borrowing, the velocity of money should also jump, with an increasing amount of excess reserves held by US banks being converted into loans. Given that banks have excess reserve holdings measured in the trillions, there is a lot of potential for new loans out there should the appetite from both the supply and demand side return. The Fed’s QE3 program should continue to help the slowly recovering housing sector—the key conduit for monetary policy transmission.

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Figure 19.11:         Lending/Borrowing Appetite

Figure 19.12:         Gold Reserves in Emerging Markets

This latest version of QE is an open-ended program designed to help lift economic growth to levels which will generate robust employment growth. If growth actually does what the Federal Reserve desires, inflation risk will no doubt worry markets, especially since the US central bank has said that it is willing to keep the system flooded with liquidity “for a considerable time after the economic recovery strengthens”.

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We will get a situation where the central bank keeps the policy rate at very low levels while also artificially suppressing yields along the long end of the curve with their market activism in the Treasury and MBS markets, even as price pressures build. This implies that the US central bank may very well be willing to err on the side of inflation, draining liquidity only after the economy reaches potential.

While we don’t project a big breakout of inflation as an imminent problem, we do see it as a growing risk should the Fed miscalculate and wait too long to drain liquidity out of the system. In addition, the Fed is not targeting headline figures, nor is it trying to hit a spot target, but rather core inflation over a time. Higher market inflation expectations and the negative real interest rates resulting from the Fed’s various QE programs may raise investor alarms that the real value of public and private debt will drop considerably. This may very well prompt them to try to hedge against that risk using gold.

After more than a 20 year period of aggressive sales, central banks have started buying gold since 2010. By the end of 2012, central banks will have likely bought some 1,030 tonnes of bullion to diversify their FX reserves. Emerging nations were particularly active as they hold some 95% of FX reserves in dollars, yen, euros and sterling and now seek to diversify their multi trillion dollar holdings.

China, which holds more than $3.5 trillion in its FX reserves (60% in US dollars), is particularly concerned that the purchasing power of its claims against Europe, Japan, the UK and the US will erode, as these nations may become tempted to monetize their massive public debt loads. All these nations are already using some form of balance sheet accommodative action.

TDS expects to see the official sector continue to buy more gold—another 525 tonnes of gold bullion to be acquired next year alone, which should tighten the market and likelyattract other investors. At some 15%, central banks have less reserves in gold than historic trends would suggest. To increase gold’s share by only 5% to 20% would require a purchase of about 9,700 tonnes of the yellow metal—this is price positive and makes us comfortable in saying that another 525 tonnes of official sector purchases is likely in 2013.

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Table 19.1:         Commodity Price Forecasts

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19.4 IMPACT OF OTJIKOTO PROJECT ON GLOBAL GOLD SUPPLY

Global gold production has exceeded 2,000 tonnes per year for more than 20 years, reaching a peak in 2011 of approximately 2,700 tonnes. The Otjikoto Project will produce on average 141,000 ounces per annum over the first five years of production. This is not a significant amount of gold compared to global production and will not materially impact on the supply-demand dynamics in the market.

19.4.1 SELECTION OF GOLD PRICE AND EXCHANGE RATE

19.4.1.1 Historical Gold Price

The 24-month trailing average gold price is shown in Figure 19.13. For the purpose of calculating Mineral Resources, a gold price of $1,350 per ounce has been used, and a price of $1,550 per ounce has been used in the financial model of the DFS. As can be seen in Figure 19.2, the gold price has been above $1,350 per ounce for almost the entire two year period and above $1,550 per ounce for all but a few days during the past 18 months.

Figure 19.13: 24-month historical gold price (source: www.goldprice.org)

19.4.1.2 Historical Exchange Rate

The 24-month historical trailing average exchange rate of the South African Rand (“ZAR) versus the USD for the period February 1, 2011 through January 31, 2013 is shown in Figure 19.3. The 24-month values range from a low of 6.6279 to a high of 8.9305 with an average of 7.7931 ZAR per USD. The most current 12 month average (February 1, 2012 – January 31, 2013) has a range from a low of 7.5072 to a high of 8.9305 with an average of 8.2611 ZAR per USD. The Namibian Dollar (“NAD”) has historically been and, continues to be, on parity with the ZAR at a ratio of 1 to 1.

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Figure 19.14:         24-month historical ZAR: USD exchange rate (source: www.oanda.com)

19.4.1.3 Contracts

At the time of this report B2Gold is in final negotiations with an international engineering and construction firm to provide the detailed engineering and procurement for the processing plant and associated surface infrastructure. In addition, B2Gold is in advanced negotiations for an Engineering, Procurement and Construction contract for a power generating plant for the Otjikoto Mine and Mill facilities. Anticipated prices for both of these contracts fall within industry norms. All other activities required to develop the site fall within the scope and capabilities of the B2Gold.

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The product being produced for sale from the Otjikoto project is doré gold bars. No contracts are currently in place for the transportation, refining and sale of gold from Otjikoto. B2Gold has contracts in place for sale of gold and silver from three producing mines that it owns in Nicaragua and the Philippines. Contracts are typically put in place with well recognized international refineries and sales are made based on spot gold and silver prices. These contracts typically include fees for transportation of the product from the site, insurance, assaying, refining and an allowance for metal losses during refining. There are no gold hedges in place, and at this time none are contemplated. Being able to get contracts in place for the sale of doré gold bars prior to the start of production in the 4th quarter of 2014 is not seen as a risk to the Otjikoto Project.

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20 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACTS

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;
• the impact of the project on the environment, proposes monitoring programs for all phases of project development; and,
• mitigation strategies to produce potential project impacts.

The EIA updates an earlier impact assessment (completed in September 2011) to included design criteria for the mine, milling circuit, tailings design and infrastructure. The updated EIA was completed in June 2012 and included extensive public consultation. The EIA received Environmental Clearance from the MET in August 2012 and was incorporated into the Mining License Amplification Application approved by the MME in December 2012.

20.1 INTRODUCTION

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.

Along with many other developing countries, Namibian economic growth was affected negatively by the global financial crisis of 2008, with sharp cuts in demand for most of the natural resource products it produces. This crisis affected the industrialised countries in particular, and their economies have been slow in regaining momentum, weighed down by the excessive borrowing needed to finance stimulus programmes. Essentially, the world economy was prevented from entering into a prolonged depression, by strong growth prevalent in emerging economies, notably China. This growth has been the driver of a general global economic recovery, which began in earnest in 2010.

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The Otjikoto Project is situated in the large, central Otjozondjupa Region. The region is occupied by semi-arid savannah vegetation types, with commercial farmland on the west, and communal land on the east. The dominant land uses make use of the natural rangelands for extensive grazing with livestock and wildlife. Only in limited areas is crop production possible. Mining is important in a few localities, and essentially dominates the economy of Otjozondjupa Region. The Otjozondjupa Region hosts the Okorusu Fluorspar mine, the Ohorongo cement mine, and the Otjozondjupa and Purity manganese mines. The small urban centers of Otjiwarongo, Otavi, Grootfontein and others, provide services for these different land use activities.

20.2 PERMITTING

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 (such as the proposed change from power from the National grid to on-site power generation from HFO to the project) have to be assessed by the 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 EMP.

Permit type	Regulator	Date received/required
Environmental Clearance for EIA and EMP	MET	August 2012
Mining licence	MME	December 2012
Consumer installation certificate (Fuel farm)	MME	Prior to commissioning
Water abstraction	Ministry of Agriculture, Water and Forestry (MAWF)	Construction phase: 100,000 m3/annum was issued by MAWF in February 2013 Operations: ±1.2 Mm3 /annum required prior to abstracting
Water Discharge		Prior to commissioning the relevant facilities
Disposal of domestic and industrial effluent	MET/MAWF	Prior to commissioning of relevant facilities
Licence for explosives magazine	MME	Prior to commencing with blasting activities and construction of the
Storing and handling/use of	MME

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Permit type	Regulator	Date received/required
explosives and explosives burning permit		explosives magazine
Removal of protected plants/trees	MET	Prior to removed/cutting down any protected plant/tree species.

Other than as outlined above, B2Gold is not aware of any significant permits or environmental factors that may affect the right or ability to commence construction activities on site. Additionally, B2Gold is not aware of any reasons why additional operational permits will not be granted. B2Gold confirms that there are currently no requirements to post any reclamation bonds.

20.3 ENVIRONMENTAL IMPACT ASSESSMENT PROCESS

Development of the EIA was completed in accordance with the Environmental Management Act, 2007 (Government Gazette No. 4878) that was promulgated on 6 February 2012. In accordance with this legal framework the EIA approach included the following:

• The scoping process was conducted to identify the environmental issues associated with the proposed projects and to define the terms of reference for the required specialist studies and the EIA;
• Environmental specialist studies were commissioned in accordance with the relevant terms of reference;
• The specialists were selected on the basis of their expertise and knowledge of the project area;
• The EIA Report was compiled on the basis of the findings of the specialist studies;
• The Otjikoto Project EMP was prepared to elaborate on the mitigation objectives and include additional actions that were described in the EIA Report; and
• A project specific public participation process was conducted. As part of this process the regulatory authorities and interested and affected parties (IAPs) were given the opportunity to attend information sharing meetings, submit questions and comments to the project team, and review the background information document and scoping report. All questions and comments that were raised by the authorities and IAPs have been included and addressed in the EIA Report. Similarly, IAPs had the opportunity to review the EIA Report and provide additional comments before it was finalised. Public open days were arranged to present the findings of the EIA to IAPs.
• All comments and issues raised during the EIA review period were included in the Issues and Reponses Report provided to MET and MME.

The initial baseline specialist studies for the EIA were conducted between April 2007 and February 2008. In August 2008 the EIA scoping report was finalized and submitted to the MET. After Teal was acquired by Auryx Gold work resumed in July 2010. After the completion of the ‘Preliminary Environmental and Social Impact Assessment of the proposed Otjikoto Gold Mine Project’ (ASEC, September 2010), a number of outstanding specialist studies identified during the preliminary EIA Report were conducted between December 2010 and December 2011.

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20.4 SUMMARY OF ENVIRONMENTAL BASELINE CONDITIONS

20.4.1 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. Otjiwarongo (the closest government weather station) falls within the higher rainfall band of Namibia, with annual average rainfall ranging between 350 and 450 mm. Rainfall is higher at Otjiwarongo decreasing towards Outjo, which lies west of the project area. In general, Namibia has high evaporation rates. Evaporation rates at Otjiwarongo are between 1,960 to 2,100 mm per year increasing towards Outjo where it ranges between 2,100 to 2,240 mm.

Maximum temperatures within the Otjiwarongo area range between 34-36 °C and occur mainly between December and February. The lowest temperatures are between 4-6 °C during the months of June and July. Monthly maximum, minimum and average temperatures are provided in Figure 20.1.

Frost is associated with the region but is only likely to occur for 1-5 days per year. The wind field of the region represents a combination of the synoptic-scale circulation and the local land-sea breeze circulation. The prevailing wind field is from the east and east-south-east with a frequency of occurrence of over 22%. Less frequent winds occur from the westerly sector. Wind speeds are primarily below 5.0 m/s with strong winds (>6 m/s) occurring less than 0.1% of the time.

Figure 20.1:         Maximum, minimum and average monthly temperatures.

	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

20.4.2 SURFACE WATER

The Otjikoto Project area is located in the most eastern parts of the Ugab catchment, an ephemeral river catchment draining westwards into the Atlantic Sea. The regional surface water divide to the neighbouring Etosha and Omatako catchments is located towards the east and north of the Otjikoto farm. The main surface water flow in the vicinity of the mine site is in a north-western direction, following the gradient towards the Platveld basin. There are no clearly defined channels found on aerial photographs or satellite images. Surface water flow in this region occurs as overland flow accumulating in pans and depressions from where it infiltrates or evaporates.

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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. No baseline surface water quality data is available for the site.

20.4.3 GROUNDWATER

AGES (S. Africa) was appointed to conduct a baseline hydrogeological assessment of the Otjikoto Project in Namibia. The purpose of this investigation was to qualify the requirements for mine water supply, mine dewatering and environmental impacts that could be associated with the proposed mine development. The focus area was on the Otjikoto farm and the surrounding environment which might be affected by the development. Additional work, such as water quality analysis, a detailed geohydrological model, water quantity and recharge models, was conducted by SLR Namibia.

The geological formations form a number of hydraulic zones:

• Calcrete forms the upper layer, which would enhance groundwater recharge. The calcrete is too shallow to form an aquifer but is expected to be permeable and would allow seepage from mineralised stockpiles.

• The regional marble formations form a major aquifer that strikes south-east to north- west. The marble forms a fractured porous aquifer with a high permeability and storativity. Underground solution channels occur in the marble aquifers.

• The schist formations in which the mine is located form a minor aquifer with permeable zones formed by faulting.

• Regional fault zones link the marble and schist aquifers and form conduits for groundwater flow.

During October 2007 water samples were taken from 10 of the boreholes intersected in the September 2007 hydrocensus. These samples were taken to determine macro and micro chemical analysis, and compared to South African water quality guidelines. According to these guidelines the water is classified as water of marginal water quality (Class 2) for domestic water quality. This means that the water is only conditionally acceptable for potable requirements and disinfection and softening would be required to make the water suitable for potable and mine water supply.

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20.4.4 AIR QUALITY

Eight single dust fallout buckets and a PM10 minivol sampler were installed in November 2010. These served to measure background dust deposition and PM10 ambient concentrations, respectively. More than one year of data is included in this air quality impact assessment study (part of the EIA) to provide an indication of the background dust fallout rates and PM10 concentrations prior to the commencement of mining operations. In addition, all existing sources of air pollution in the region were identified and qualitatively described based on the associated pollutants and potential to contribute to the background ambient concentrations and dust fallout levels at the project area.

The main activity around the proposed Otjikoto Project is feedstock farming. The B1 national road is approximately 2 km west of the proposed mine with a few unpaved farm roads surrounding the site. These activities are likely to result in particulate emissions at a localised scale whereas SO2 and NOx are likely to result from biomass burning (veld fires). Traffic on the national road will add to both the particulate and gaseous emissions whereas windblown dust from exposed tilled land could be a significant source of particulate emissions.

Data from the Otjikoto Project dust fallout monitoring network indicate dust fallout rates generally below the SANS residential limit of 600 mg/m²/day. The highest dust fallout collected was during September 2011, exceeding the SANS residential limit. It is suspected that this was due to localised dust generating activities near the unit. On average, the dust monitoring unit (dust bucket) located next to the entrance to the mine property collected the highest dust fallout and this is due to dust generation on the unpaved access roads.

Measured PM10 daily concentrations are moderate in general, not exceeding the WHO_IT3 guideline. The more stringent WHO guideline was however exceeded on six occasions, mostly within the winter month of June. It should be noted however that the data record is very limited and the general background PM10 may be higher.

20.4.5 BIODIVERSITY

The area falls within the semi-arid Thornbush Shrubland vegetation type of the Acacia Tree-and-shrub Savanna Biome. The area receives a mean annual rainfall of about 450 mm, with a coefficient of variation of about 30%, meaning that rainfall is fairly unpredictable. The mean annual evaporation rate is about 2,000 mm. The landscape is characterized by dense tree and bush savannah, dominated by Acacia species and annual and perennial grasses. Bush density and grass composition are largely determined by micro-habitat conditions and rangeland management practices. Throughout the focal areas the vegetation is dense (difficult to walk through) and highly bush-encroached.

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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. It is characterised predominantly by Acacia spp., including A. reficiens, A. hebeclada subsp. hebeclada, A. mellifera subsp. detinens, A. fleckii, A. erubescens and several others. Other common tree species include Philenoptera nelsii and Peltophorum africanum in sandy areas, Combretum apiculatum subsp. apiculatum on limestone and rocky outcrops, and Ziziphus mucronata in various habitats. There are no plant species of sufficient conservation concern in any of the habitats affected by the proposed mine and, due to the relatively low sensitivity of the vegetation present, no special mitigation measures are necessary.

The vegetation of the proposed mining site is relatively homogenous. Except for the pan to the east of the site, there are no geographic features, e.g. mountains, which could provide for a diversity of habitats. The area is not pristine with quite a high level of bush encroachment. Sensitive or extreme habitat specific species are thus already expected to be locally extinct.

An annotated checklist of amphibians, reptiles and mammals occurring within the Otjikoto Project area is included in the EIA. Fourteen (14) species of amphibians, seventy-eight (78) species of reptiles and seventy-eight (78) species of mammals are known or expected to occur within the Otjikoto Project area. To a greater or lesser extent, most of these species can be expected to interface with the proposed mine. Overall, reptiles are not expected to be unduly influenced by the development of this project, as significant populations occur outside the project area. No mammal species are expected to be affected to the extent that their regional or national conservation status will be degraded. For amphibians, the only species of conservation concern is the Spotted Rubber Frog (Phrynomatis affinis). Although this species undoubtedly occurs within the area owned by B2Gold Namibia (although most likely outside of the mining project area in the pan area), the population in this area would be relatively small in comparison to the national range. The potential loss of this species in the footprint as well as in the extended mine area (worst case scenario) would not have a significant impact on this species' national conservation status.

Many bird species found in the Otjikoto Project area (as in other parts of Namibia) are highly nomadic, moving from place to place in search of appropriate conditions. (These differ from migratory species that undertake fixed annual movements, usually from north of the equator to the southern hemisphere and back). Nomadic species by contrast generally remain within Namibia or adjacent parts of the southern African sub-region, but move around widely and in no fixed pattern, to exploit patchy and unpredictable food, water and other ecosystem resources. Thus a number of nomadic species that occur in the area may not be present at certain times of the year or in some years, and their abundance may differ dramatically over time.

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The Otjikoto Project area does not fall within Namibia’s endemic zone and there are no endemic birds to Namibia in the project area. While a number of species recorded are endemic to the southern African sub-region (typically the “south-west arid” zoogeographic zone), these species all have large ranges in southern Africa and none are considered to be threatened in any way.

Six species recorded for the project area are listed in Namibia’s Red Data Book: Whitebacked Vulture listed as Near Threatened; Lappetfaced Vulture listed as Vulnerable; and Tawny, Booted, Martial and Bateleur Eagles all listed as Endangered. For all these species the main impact in the focal area would have been breeding disturbance and loss of breeding sites. A careful search of the most likely breeding sites (copses of taller trees) revealed no sign of nests. Based on a site survey, it was concluded that the loss of habitat resulting from the proposed mining and processing developments in the project area poses no significant conservation risk to the avifauna in the area.

20.4.6 SOCIO-ECONOMIC

The Otjozondjupa Region is the fifth largest region in Namibia, 105,460 km², with a total population of 142,400 people, giving a population density of 1.4 persons/km². The Otjozondjupa Region is divided into seven constituencies and the proposed mine is situated on the main road between the Otavi and Otjiwarongo Constituencies, approximately equidistant between them. The 2011 census shows a decrease in Otavi constituency’s population from 12,387 in 2001 to 11,800 in 2011, of which 5,200 live in the municipality of Otavi. By contrast, the more populated and economically active Otjiwarongo Constituency had grown from 23,412 people in 2001 to 30,400 in 2011, of which 28,000 inhabitants lived in the town of Otjiwarongo.

The mine is immediately surrounded by commercial farm land, which is used primarily for ranching cattle with some game hunting and tourism use in varying combinations. The closest direct neighbouring farms encircling the mine farms clockwise are: Erhadtshof, Bergershof, Okaruiputa, Erpfsfarm, Houmoad, Okaputa, Okaputa West, Stark, Fisher, Lardner, Hester, Embla and Tirol.

Generally, it can be seen that land use in the northern savannah is not highly attractive for the investor, with low financial profitability and return on investment. Tourism may be slightly more financially attractive, and this might explain the tendency for wildlife use to expand at the expense of livestock in this setting. In terms of economic contribution (contribution to the national income), tourism is distinctly more efficient than livestock production. Several forces affect the relative values for land uses in this setting. Bush encroachment has reduced the suitability of rangeland in the Otjikoto mine area for both livestock and tourism. Use of invader bush for commercial production of fuel can provide a financially and economically viable way of ameliorating this problem.

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20.4.7 ARCHAEOLOGY

An archaeological survey was completed as part of the baseline surveys. Two sites of archaeological/historical interest were found on the Otjikoto farm outside the area of development during a first site visit by the archaeological specialist. Mitigation measures include fencing off these areas and not disturbing the sites. In addition, there are four Christian burial sites on farms Gerhardshausen and Okaputa Nerd (also named Felsenquelle) which are also protected by the Burial Ordinance.

20.4.8 VISUAL

The Otjikoto Project area is covered by Tree and Shrub Savannah type vegetation. This is the dominant vegetation type in central Namibia. An ephemeral pan, situated on farms Otjikoto and Gerhardshausen, is surrounded by woody species, whereas the rest of the project area is composed of thorn bush thicket.

The following landscapes were defined as falling within the zone of visual influence of the proposed project. The landscape character of each physiographic rating units is:

• Game farms;
• Agricultural;
• Road/railway corridor;
• Ephemeral pan;
• Existing modified landscapes (e.g. from prospecting, power lines); and
• Ridgeline.

As a result of the moderate levels of landscape character, only the core receptors were identified as Key Observation Points (KOPs). The following KOPs were selected to be assessed in the Contrast Rating stage in order to determine visual impacts:

• B1 northbound;
• B1 southbound; and
• Cattle farm/D2808 eastbound.

20.4.9 NOISE

Ambient (background) noise measurements were taken during the day and early morning on November 1^st and 2^nd, 2010, respectively. From the data obtained, it can be seen that the ambient daytime sound levels ranges between 22.4 (_LA,90) and 18.6 (L_A,min) dBA. The average ambient night-time sound levels (_LA,90) ranged between 20.5 and 15.1 dBA (L_A,min). All samples illustrate the rural character of the area during periods with light winds and mainly natural sounds defining the acoustic character, the area is considered rural. According to South African

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National Standard (SANS) 10103:2008, the maximum average ambient noise levels for rural residential areas are 45 dBA (day) and 35 dBA (night).

Based on these observations, it is suggested that ambient noise levels of 22.4 -18.6 dBA and 20.5 – 15.1 dBA for the day and night, respectively, be used as a reference for calculating the impact caused by the proposed Otjikoto mining operation.

20.5 SUMMARY OF KEY ENVIRONMENTAL IMPACTS

The impact assessment looked at all the different potential impacts based on the Otjikoto Project baseline and proposed project design. Key findings include:

20.5.1 TOPOGRAPHY

The Otjikoto Project area is approximately 2-3 km east of the B1 National Road and it is not impossible that third parties may access the site by mistake or out of curiosity. In the unmitigated scenario, it is possible that the surface excavations and infrastructure present a risk to unaccompanied third parties during all phases. The proposed mitigation measures will focus on infrastructure safety and on limiting access to third parties and animals which reduces the probability of the impact occurring.

20.5.2 SURFACE WATER

In the unmitigated scenario, surface water may collect contaminants (hydrocarbons, salts, chemicals, metals, etc.) from numerous sources. At elevated pollution concentrations these contaminants can be harmful to humans if ingested directly or indirectly through contaminated vegetation, vertebrates and invertebrates. Therefore the precautionary approach has been applied with the result that contaminated surface water from the mine will be contained, re-used and/or treated.

Despite this precaution one can assume that due to the site specific climatic and hydrological conditions, if a surface run-off occurs, a large fraction of the surface water seeps into the ground water alluvium. Therefore mitigation measures are based on the principle of containing dirty water and diverting clean water as far as possible, so the drainage scheme will ensure that clean surface water flow is directed around structures such as the waste rock dumps and tailings disposal facility.

20.5.3 GROUNDWATER

Dewatering of the pit and borehole abstraction has the potential to reduce local groundwater levels and there are a range of sources/activities that have the potential to contaminate groundwater in all phases of the project. The proposed mitigation therefore focuses on containing pollution at source and/or preventing the off-site migration of pollution and preventing unacceptable groundwater pollution related to seepage and groundwater flow disruption impacts. In addition, it is proposed to manage the local water supply and to monitor and mitigate impacts if required.

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According to the groundwater model and hydrogeological investigations, no major impact is expected. B2Gold Namibia will ensure that if community supply boreholes are dewatered, the community will be provided with an alternative water source. Groundwater levels will be monitored at all pumping wells and monitoring boreholes as well as the nearest community supply boreholes, as per the monitoring programme included in the EMP. The groundwater flow model will be updated within one year after approval of the Otjikoto Project and every two years subsequently.

20.5.4 AIR

The air quality assessment focused on airborne particulates (PM10 emissions and total suspended particles). Gaseous pollutants (such as sulphur dioxide, oxides of nitrogen, carbon monoxide etc.) deriving from mine vehicles and equipment and from the gold smelting operations could not be assessed but were regarded by the specialist as potentially negligible in comparison to particulate emissions.

The main contributing sources to both the unmitigated and mitigated PM10 predicted impacts are firstly the unpaved roads, specifically between the open-pit and the waste rock dump, and secondly the materials handling operations and over the short term windblown dust from the TSF. In the unmitigated scenarios, the PM10 ground level concentrations exceed the WHO-IT3 limit at all sensitive receptors. With mitigation these exceedances can be reduced and the impact area becomes much smaller. Some exceedances are still predicted, regarding the WHO-IT3 daily limit. The annual WHO-IT3 limits for PM10 ground level concentration are not exceeded at any potential receptor areas in the mitigated scenario.

A number of uncertainties are associated with the calculation and simulation of fugitive dust from unpaved roads and windblown dust resulting in expected conservative ground level concentrations (GLCs). This is in addition to the inherent model uncertainty of -50% to 200%. Therefore, with effective mitigation the impacts reduce significantly and can be limited to on-site only as indicated by the annual average GLCs. The proposed mitigation measures therefore focus on the use of water sprays, vegetation cover at relevant areas and suppressants to achieve certain control efficiencies and monitoring the effectiveness of these measures.

20.5.5 NOISE

Excluding one receptor that belongs to B2Gold Namibia, the severity of the noise impact is considered to be low for all potential receptors. Considering a worse-case scenario the severity of the potential noise impact could be a medium. It should be noted that the unmitigated medium severity is based on the fact that there is the potential that the operation of the mine will increase the ambient sound levels at night by more than 7 dBA at this one property. As such, it is possible that the receptor will detect the change in ambient sound level, yet it is unlikely that the receptor may complain, as the total cumulative night-time noise level is still predicted to be less than 35 dBA (a level determined to be the level where rural communities may start to complain about the increased noise levels).

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20.5.6 BIODIVERSITY

In the unmitigated scenario, biodiversity will be disturbed in the following ways:

• Firewood collection can lead to loss of woody plants and their key role as a nutrient source;

• Illegal poaching of game, game birds, reptiles and other animals;

• Illegal collection and loss of seeds and plants;

• Deposition of dust on plant communities and impact on plant vitality;

• Introduction of invasive plant species;

• Interference with movement of large game;

• Habitat fragmentation isolates organisms from wider populations;

• The presence of vehicles in the area can cause road kills especially if drivers speed;

• White light attracts large numbers of invertebrates, which become easy prey for predators this can upset the invertebrate population balances;

• Noise and vibration pollution may scare off vertebrates and invertebrates and influence migration paths;

• The presence of open pits and water dams may lead to the death/injury of animals; and

• Pollution emissions, residues from the mining activities and general litter may impact on the survival of individual plants, vertebrates and invertebrates.

In addition, periodic surface water run-off and the existence of near-surface water resources are understood to be key ecological drivers for vegetation, vertebrates and invertebrates within and downstream of the proposed project site because it recharges the moisture content of the aquifer. This in turn supports a range of vegetation, in particular larger trees, which in turn support a range of invertebrates and vertebrates. In addition, water points on farms, on which wild mammals and birds also depend, may become more difficult to replenish. However, the Otjikoto Project is located in an area of low biodiversity sensitivity and there are extensive areas of similar habitat in the surrounding area. Therefore, the potential significance is considered to be medium which would reduce to between medium and low with the successful implementation of the mitigation measures.

20.5.7 VISUAL

The severity of visual impacts is determined by assessing the change to the visual landscape as a result of mine related infrastructure and activities. The most visible aspect of the proposed mining project will be the proposed waste rock dump which, at 40 metres high, is considered to have a high visual impact without mitigation. Other elements of the mine including the processing plant, open pit and TSF are all lower in height and benefit from the screening provided by surrounding vegetation.

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Key mitigation measures therefore include:

• Special attention needs to be given to the entrance to the B1 national road, as well as to the eastern section where the new road diverges from the old. At these locations, a 2.5 metre natural looking berm needs to be raised to act as a screening wall. The berms need to be planted with local endemic plants to ensure that the views down the old road are screened;

• All painted surfaces should blend into the natural surroundings.

• Rehabilitation of all the faces of the waste rock dump to grass / scrub bushes and some trees;

• Reduce the angle of the waste rock dump slope if not suitable for rehabilitation;

• An agreement needs to be reached with the local authority to ensure that the trees between the railway line and the road are not removed; and

• For the closure phase:

o B2Gold Namibia will establish a mechanism to ensure that the rehabilitation of the mine is properly funded to ensure that sufficient funds are available to implement the rehabilitation and mitigations required for closure.

o All components of the infrastructure used during operation, apart from the waste rock dump and tailings disposal facility, must be removed. The site must be visually ‘cleaned up’ so as to portray an uncluttered landscape.

o The ground where processing plants were located must be decontaminated and then covered by the earth used for the berm and landscaped into a natural form in alignment with the natural hydrological patterns.

20.5.8 ARCHAEOLOGY

The nature of the likely impacts on the archaeological sites include the direct impacts of earthmoving activities and the indirect impacts of accelerated soil erosion caused by the clearance of vegetation, and potential damage from blasting activities. Given the sensitive nature of the archaeological sites and when taking into account existing impacts from on-site exploration activities, the significance of these impacts without mitigation is high, but this could be reduced to low/medium with mitigation, as follows:

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• Any sites identified by the specialist will be marked and protected, and the area within cleared of all encroaching bush.

• A formal “No-Go Area” policy, as set out in the Otjikoto Project EMP, will be developed and implemented.

• All workers (temporary and permanent) will be educated about the archaeological sites that may be encountered and appropriate training will be provided.

• In the event that mining activities will unavoidably encroach on an archeological site, B2Gold Namibia is legally obliged to take all necessary steps either to protect the sites, or to re- locate the burials according to official directives.

If there are any chance finds of archaeological sites that have not been identified and described in the specialist report, B2Gold Namibia will follow its chance find procedure. The key component of which is to ensure that the site remains undisturbed until a specialist has assessed the site, assessed the potential damage, advised on the necessary management steps and advised on the requirements for stakeholder consultation and permitting.

20.5.9 TRAFFIC

The mine will have access off the main B1 road which is designed to carry traffic of this nature. Mine related traffic will travel from both the Otavi and Otjiwarongo directions on a daily basis with all traffic using the B1 from which the re-routed D2808 will provide the site access. Traffic on the B1 would be a combination of B2Gold Namibia, national and international long distance goods haulage, and local business, tourism, other mining and private traffic. The highest levels of additional B2Gold Namibia traffic would be likely to be experienced during the construction phase which is relatively short term. Levels over the operational life of the mine would be lower and traffic in the decommissioning phase will reduce to the extent that production related traffic ceases. In the closure phase traffic will be limited to just occasional trips for monitoring and aftercare activities.

20.5.10         ECONOMIC

The life of mine (LoM) is characterized by an 18 month intensive construction phase involving the capital investment of approximated US$242 million, followed by operations lasting some ten years, merging into a decommissioning phase, followed by mine closure after 14 years. Clearly investment costs, operational costs and income will vary over the 14 year LoM, but the average direct contribution to Gross Namibian Income (GNI) of the mine during LoM will be very significant in local terms and will impact the total national GNI of Namibia. Further, due to the value added or income multiplier effect, this direct impact will induce some US$103 million in additional GNI within the broader economy.

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The whole mining and processing operation will create an average of 454 jobs, which will peak at 540 jobs. Studies have calculated that for every job created by a mine, a further additional 1.5 job opportunities are created by suppliers and contractors, which will result in a further 681 jobs created by suppliers, contractors and service providers. Through employment and skills development, the proposed mine will contribute to the three national development goals of – reducing income inequality, increasing job creation and economic growth. It will also contribute to the Otjozondjupa Regional Council’s strategic objective “to improve regional economic development and employment”.

The economic spin-offs from the mine’s construction and operations will create empowerment opportunities in a range of skills and activities. Employment provides income to the employees, their immediate household members and to others living elsewhere in Namibia who depend on cash remittances. Thus, the assessment of this impact can be summarised as having a high beneficial effect. The work experience and skills gained through the opportunities that the mine brings will have lasting benefits for all employees nationally.

The negative land and natural resource use costs which are estimated to happen over the LoM have been measured at being just over 3% of the positive contribution of the mine. So the economic benefits clearly outweigh the negative economic impacts that may occur and the impact has a positive high severity both with and without mitigation. The other potential negative economic impact is concerned with mine closure. This potential impact is common to all mines and must be planned for and mitigation provided because the major economic benefits that the increase in mining will bring to the country cannot be realised without the eventual impact of closure.

20.5.11         SOCIAL

The population of the town of Otavi has grown from 3,800 in 2001 to 5,200 in 2011. On a comparatively larger scale, Otjiwarongo municipality has grown from 23,000 to 31,000 over the same period. Otjiwarongo’s capacity to meet the needs of the Otjikoto Project, in terms of available goods and services, is good. Therefore it would be sensible for the two towns to anticipate and plan for substantial in-ward migration of around 7,000 people, shared between the two towns.

In-ward migration can have positive and negative impacts such as:

• Stimulating the local economy and community organization. Mineworkers will require housing and this will stimulate the construction of housing in both Otavi and Otjiwarongo. Wages and salaries will be injected into the local economy where they live.

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• Developing informal settlements which increase demand/need for basic infrastructure – housing, clean water supplies, sanitation, and electricity and waste management systems.
• Increasing pressure on government services such as health and education facilities, their staffing and running cost requirements.
• Increasing incidence of social ills including alcoholism, drug abuse, prostitution, gambling & criminality. Alcohol abuse is part of the accepted social norm in Namibia and is often stimulated by cash earnings which increase the likelihood of domestic violence (usually against women and children), unprotected sex and the spread of HIV. The influx of job seekers may increase over-crowding, which increase the spread of TB.

Unmitigated, these social ills can have severe consequences which last beyond the life of mine but a number of policies and actions can be taken which can mitigate the impact and reduce their severity. However these efforts must be carefully co-ordinated with those of local government and the community, through planned engagement and public-private partnerships.

The potential negative environmental impacts associated with the proposed project are expected to be of mainly low to medium significance and provided that the relevant mitigation measures are successfully implemented there is no environmental reason why the proposed Otjikoto Project should not proceed. In addition, the Otjikoto Project will have significant positive economic impacts of benefit to the local, regional and national economies of Namibia.

20.6 ENVIRONMENTAL MANAGEMENT PLAN

The EMP documents a series of individual management programs (MPs) that are designed to meet legal requirements and minimise the impacts associated with the implementation of the Otjikoto Project. A draft EMP was prepared as part of the scoping phase of the EIA process for the proposed Otjikoto mine in 2010 and has been updated to take into consideration the findings of the latest impact assessment (2012). It is the intention that the EMP should be seen as a “living document” which will be amended during the operation of the Otjikoto Project, as the activities might change or new ones be introduced. This is in accordance with Section 50 (g) of the Namibian Minerals (Mining and Prospecting) Act, 33 of 1992, which states that the holder of a mining license shall undertake the periodic review of the EMP(s) should circumstances change.

MPs were created for the following:

• Stakeholder consultation/communications;
• Safety and security;
• Surface water/storm water;
• Groundwater;
• Air quality;

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• Noise and vibration;
• Biodiversity;
• Visual impacts;
• Archaeology;
• Traffic control;
• Socio-economics;
• Resources;
• Soils; and
• Waste management.

20.6.1 STAKEHOLDER CONSULTATION/COMMUNICATIONS

Experience has demonstrated that projects that establish early and ongoing stakeholder engagement processes that continue throughout the life of the project have a greater rate of success, as well as reduced business risk. Additionally, routine monitoring and reporting are integral to ensuring continual assessment and improvement.

A stakeholder database has been created for the site and is periodically updated. The database includes contact information, stakeholder representatives, preferred contact information and additional information specific to each stakeholder.

The Stakeholder Engagement Plan (“SEP”) has been developed to: identify project stakeholders; describe a process to ensure timely and effective communication between the project site and affected stakeholders throughout the life of the project; document all meaningful consultation and disclosure; establish a formal complaint process, and include a monitoring and reporting component.

The SEP is a living document and will be updated from time to time (no less than annually) based on project developments and feedback provided by stakeholders.

An Annual Monitoring Report (“AMR”) will be prepared and publicly disclosed as part of on-going stakeholder engagement. The report for each site will be prepared prior to the end of the 1^st quarter of the calendar year.

Management will commit to coordinating activities and assigning overall responsibility; hiring and training sufficient personnel, creating clear reporting guidelines between management and a community liaison leader, managing contractor risks, and tracking stakeholder engagement.

20.6.2 SAFETY AND SECURITY

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The objective of the Safety and Security MP is to prevent physical harm to third parties and animals from potentially hazardous excavations and infrastructure. The MP is made up of the following components:

• General (third party) safety and security;
• Construction workers (after hours) - Contractors camp; and
• Occupational Health & Safety.

20.6.3 SURFACE WATER MANAGEMENT PROGRAM

There are a number of sources in all project phases that have the potential to pollute surface water, particularly in the unmanaged scenario. In the construction and decommissioning phases these potential pollution sources are temporary in nature, usually existing for a few weeks to a few months. Although these sources may be temporary, the potential pollution may be long term. The operational phase will present more long term potential sources and the closure phase will present final land forms that may have the potential to contaminate surface water through long term seepage and/or run-off.

The Surface Water MP is made up of the following components:

• Altering drainage pattern;
• Pollution of surface water – general;
• Industrial effluent;
• Domestic effluent; and
• Spills.

This MP includes a commitment to minimize fresh water use and maximize recycle from the tailings pond as well as segregating dirty and clean water streams.

20.6.4 GROUNDWATER

Potential groundwater quality and quantity impacts can be an issue during the construction and operation of the various mining activities and infrastructure unless measures are undertaken to prevent and mitigate such impacts. The purpose of the Groundwater MP is to provide for methods to be followed to achieve such mitigation.

This MP is made up of the following components:

• Water Supply and Dewatering of Mine Pit; and
• Contamination of Groundwater.

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20.6.5 AIR QUALITY

There are a number of sources in all phases of the project that have the potential to pollute the air. In the construction and decommissioning phases these potential pollution sources are temporary in nature, usually existing for a few weeks to a few months. The operational phase will present more long term potential sources and the closure phase will present final land forms that may have the potential to pollute the air through long term wind erosion. The EIA showed that the more significant emissions are - inhalable particulate matter less than 10 microns in size (PM10). The commitments derived from the EIA Report with regards to air quality form the basis of the MP and include dust monitoring and dust mitigation strategies such as road watering and chemical suppressants.

20.6.6 NOISE AND VIBRATION

There is a range of construction, operation and decommissioning activities that have the potential to generate noise and cause related pollution. Noise pollution will have different impacts on different receptors because some are very sensitive to noise and others are not. For example, mine workers in general do not expect an environment free of mine related noise and so they will not be as sensitive to environmental noise pollution at work. In contrast, local communities are likely to be more sensitive to unnatural noises and so any change to ambient noise levels because of mine related noise will have a negative impact on them, although studies have shown that only one property is likely to be affected. In addition the mine will introduce blasting vibration and the potential for flyrock into an area where there is currently none. The commitments derived from the EIA with regards to noise and vibrations forms the basis of this MP and include: monitoring, blast design and stakeholder engagement to ensure that receptors are not being impacted.

20.6.7 BIODIVERSITY MANAGEMENT

To understand biodiversity, one must appreciate all of its components. It is not just about the species of plants and animals and the different habitats in which they live (biodiversity patterns) but the way that factors such as wind, water, steepness of slope and presence of pollinators, affect the habitats and the species living in them (ecosystem processes). The commitments derived from the EIA with regards to biodiversity form the basis of this MP and include:

• managing the physical destruction of biodiversity;
• reduction of water resource as an ecological driver; and
• managing general disturbance.

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20.6.8 VISUAL MANAGEMENT PROGRAM

Visual impacts of the open pit, processing plant and TSF will generally be limited. Views of the waste rock dump present the greatest visual exposure. The commitments derived from the EIA Report with regards to visual impacts form the basis of this MP. The objective of the management measures is to limit visual impacts.

This MP is made up of the following component:

• Visual disturbance.

20.6.9 ARCHAEOLOGY MANAGEMENT PROGRAM

There are a number of activities/infrastructure in all phases that have the potential to damage archaeological resources. The commitments derived from the EIA with regards to archaeology, form the basis of the MP and include:

• measures required to protect existing sites;
• chance find procedures; and
• Demarcation of “no go” area (not currently anticipated to impact the proposed footprint).

20.6.10         TRAFFIC CONTROL

The activities associated with the mine have traffic impacts, particularly in the construction and operational phases. This MP aims to provide measures to limits the negatives impacts. The commitments derived from the EIA Report with regards to traffic issues form the basis of this MP. This MP is made up of the following component:

• Infrastructure – road use.

The objective of the management measures is to reduce the potential for safety and vehicle related impacts on road users.

20.6.11         SOCIO-ECONOMICS MANAGEMENT PROGRAM

The activities associated with the mine have socioeconomic impacts in all mine phases – mostly positive. These impacts relate to amongst others employment/job creation, inward migration, improvement of local- and regional economies, land use issues and community safety and security. The socioeconomic MP tries to enhance the positives and minimize negative impacts through:

• development of employment opportunities;
• best practices procurement policies and procedures;

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• capacity building and educational opportunities;
• development of small to medium enterprises;
• investment in sustainable businesses;
• implementation of a corporate responsibility fund that is transparent and represents the needs of the stakeholders; and,
• stakeholder engagement

20.6.12         RESOURCES

This MP provides management actions regarding resources like water and provides means of reducing the consumption of resources. It is made up of the following components:

• Consumption of energy;
• Consumption of water;
• Use of manufactured materials; and
• Consumption of fuel.

20.6.13         SOILS

Management of soils is important as mining is a temporary land use where-after rehabilitation is the key to re-establishing post closure land capability that will support conservation, agricultural and tourism type land uses. Soil is a key part of rehabilitation. The commitments derived from the EIA Report with regards to soil form the basis of this MP.

The objective of the management measures is to ensure that all topsoil stripping, stockpiling and replacement operations will be undertaken in a manner that limits impacts on the soil functionality and to ensure it can be used for rehabilitation as and when required.

20.6.14         WASTE MANAGEMENT

Waste is generated during all phases of the mine. This MP deals with solid waste management.

The MP is made up of the following components:

• Non-hazardous solid waste (non-mineralised);
• Hazardous solid waste (non-mineralised); and
• Medical waste.

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Table 20.1:         Waste Inventory list:

Waste type	Waste specifics (example of waste types)	Source
Non-hazardous solid waste (non-mineralised)	Metal cut-offs, rubber, wood, cardboard/paper, used PPE, etc.	Across site
Hazardous solid waste (non- mineralised)	Printer cartridges, sewerage, batteries, hydrocarbons (oils, grease), fluorescent bulbs, etc.	Admin building, workshops, plant
Medical waste	Syringes, material with blood stains, bandages, etc.	First Aid Centre

The objective of the management measures is to ensure proper storage, removal, transportation and disposal of waste.

20.7 RECLAMATION

The mine reclamation and closure costs have been estimated based on the Mine Closure Framework prepared by SLR. The economic model includes approximately US$12.2 million for reclamation and closure. Most of the reclamation costs are incurred during operations as part of on-going reclamation. This includes interim reclamation of the waste dump and tailings slopes to minimize dust. Additional costs are related to removal of structures and retrenchment payouts to employees. A mechanism for accrual of reclamation costs will be agreed upon with the MET prior to commencement of operations.

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21 CAPITAL AND OPERATING COSTS

21.1 CAPITAL COST ESTIMATE

The capital cost estimate for the mine, process plant, power supply, surface facilities and all related infrastructure for the Otjikoto Project has been prepared in accordance with the standard industry practices for this level of study and to a level of definition and intended accuracy of +/- 10%. This section covers the capital costs for the 2.5 Mtpa processing plant and surface mine.

The capital cost estimate for the Otjikoto Project’s processing plant and infrastructure was compiled by the project management team supported by the primary feasibility consultant, DRA, and other feasibility study engineering contractors that contributed significantly to the capital cost estimate form and basis. DRA provided the plant process and infrastructure capital costs. VBKOM provided the form and basis for the surface mining capital costs inclusive of the mining equipment and mine development costs. Epoch provided the designs and quantities for the TFS cost estimates. The project schedule was completed by DRA.

In deriving construction costs, DRA consulted with contractors, vendors and service suppliers to establish costs in today’s market as much as possible. DRA has addressed construction personnel, material and equipment availability, work methods and areas of risk.

The capital cost estimate includes direct and indirect costs covering all of the traditional items typical of any project, and incorporated DRA’s and the Otjikoto Project management team’s experience based on similar projects that have been constructed over the last few years.

21.1.1 SUMMARY

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 Namibia.

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 Minerals Project (Pty) Ltd and the B2Gold’s project’s management team.

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The capital cost estimate for the Otjikoto Project is approximately US$244.2 million as of December 2012, subject to qualifications, assumptions and exclusions, all of which are detailed following.

The pre-production capital cost summary and distribution are shown in Table 21.1.

Table 21.1:         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

21.1.2 BASIS OF ESTIMATE

The direct costs have been based on the following information:

• Process flow diagrams, process design criteria, piping and instrumentation diagrams, mechanical equipment list, site layout and general arrangement drawings, electrical single line diagrams.
• Budget submissions for the design/supply of new primary and secondary equipment provided by vendors in accordance with specifications and/or data sheets developed by the engineering groups involved.
• Prices for materials based on supplier/vendor quotations and in-house data and current market conditions.
• Quantity take-offs for materials provided by the engineer.

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• Contractor unit rates provided by local and regional construction contractors with whom DRA has solicited and/or worked previously on similar projects in Africa.
• Owner’s labor rates based on a regional standardized labor index “Patterson job classification and pay scale” was based on the most recent available information of 2011 and then inflated 8% for the 2012 year and 8% for the 2013 year.
• Productivities for construction activities of installing equipment and materials provided by DRA based on their familiarity with the project location and local and regional conditions. DRA has reviewed the productivities that can be expected for this location and compared them to actual results from other similar projects as well as with in-house data.
• VBKOM provided the form and basis for the surface mining capital costs inclusive of the mining equipment and the mine development costs.
• Epoch provided the design and quantities for the tailings pond cost estimate.
• Supply and installation prices from experienced vendors of pre-engineered and modular buildings.
• Freight cost allowances for mechanical equipment were obtained from vendor quotations. Transportation costs for steelwork and platework were quoted by the fabricator/erector. Other transport costs were based on a tonnage or “per load” basis. When freight costs were provided by the vendor/supplier in a quote, then DRA utilized these vendor supplied freight costs. The freight costs for civil and concrete materials are included in the unit rates. The mining equipment cost is inclusive of the freight cost to site.
• Geotechnical information / recommendations provided by SRK (Pty) Ltd.
• An engineering, procurement and construction management (EPCM) project execution strategy and project schedule was prepared with costs by DRA.

21.1.3 DIRECT COSTS

Direct costs include costs for all new equipment, new materials and installation of all permanent facilities associated with:

• Crushing, material handling and processing facilities
• Process buildings
• Infrastructure roads and site preparation
• Power generation and distribution
• Tailings storage facility
• Warehousing facilities
• Administration facilities
• Mine maintenance facilities
• Mine buildings
• Yard services and other utilities
• Communication systems

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• Mine equipment
• Plant mobile equipment
• Site general mobile equipment
• Construction equipment
• Fuel storage facility to be supplied by vendor, no cost allocation
• Explosives facility to be supplied by vendor, no cost allocation

21.1.4 INDIRECT COST

Indirect costs include the following:

• Mechanical and Electrical spares
• First fills at processing plant
• Engineering, procurement and construction management services (including travel expenses)
• Surveying
• Owner’s costs
• Owner’s construction management
• Start up and commissioning allowance
• Clearing and grubbing
• Security and game fencing
• Construction personnel transport
• Owners mobile equipment costs
• Vehicle purchase, operation and maintenance
• Construction camp facility purchase and operation and maintenance

21.1.5 WORKING CAPITAL

An equivalent of 25% of the first month’s production operating cost and 25% of the first month’s net production value is included as the working capital in the projected pre-production cash flow statement. This should prove to be acceptable given that it is typical for a gold mine located within a developed infrastructure to be able to ship the dore to the refinery weekly or as desired. Typically, the refinery will pay a provisional payment of up to 98% of the estimated gold value for early settlement, payable within less than 5 days of receipt.

21.1.6 CONTINGENCY

The overall contingency for the project development has been estimated by B2Gold as follows:

The mine equipment, construction equipment, owner’s costs and the owner’s construction management did not receive a contingency allowance.

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The contingency applicable to the process plant (milling) estimate was by estimated by DRA by assessing the risks applicable to estimating inaccuracies. Which has been assessed (as shown below) by reviewing the engineering and estimating confidence pertaining to the overall supply and erection costs within each cost discipline.

The estimating inaccuracy allowance has then been assessed as a percentage for the individual cost components, from which an overall estimating inaccuracy has been derived.

The estimating inaccuracy is assessed as a measure of the confidence in the design and estimating processes/outputs. The confidence in each design and engineering discipline was assessed against the following scale:

HH	(High High confidence)	5%
H	(High Confidence)	7.5%
HM	(High Medium Confidence)	10%
M	(Medium Confidence)	12.5%
ML	(Medium Low Confidence)	15%
L	(Low Confidence)	20%

The assessment produces an average assessed confidence level for each discipline. The estimating inaccuracy allowance addresses the inaccuracies in the quantity, rates and prices used in the estimate and does not address any risks associated with the unknown or scope changes.

The overall contingency and accuracy level of the process plant estimate was determined by DRA to be 9.58% . The contingency for the plant was later reduced by B2Gold.

21.1.7 QUALIFICATIONS AND ASSUMPTIONS

The following assumptions were made in preparing the estimate:

• Diesel price of $1.056 per liter.
• Namibian Dollar to US Dollar exchange rate of 8.5.
• Namibian Dollar to Euro exchange rate of 11.1909.
• Construction work pricing based on unit price rates.
• Vendors supplied budgetary quotes for major equipment and materials.
• The construction man camp was secured in December 2012 with acceptable erection date.
• The primary gyratory crusher, SAG mill and Ball mill were ordered in December 2012 with acceptable delivery dates to site.
• A concrete batch plant will be established at site and adequate aggregate will be available locally.

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• Soil conditions will be adequate for foundation bearing pressures.
• Construction activities will to be carried out in a continuous program.
• Bulk materials such as rebar, structural steel and plate, electric cable and piping are all readily available in the scheduled timeframe.
• Capital equipment is available in the timeframes scheduled since availability has been verified with suppliers.
• Namibian work visas will be granted timely to key construction expatriates.

21.1.8 PRICING

None of the pricing for commodities or the design/supply of equipment is based on binding quotations, other than the construction man camp, primary gyratory crusher, SAG Mill and Ball Mill, which have been purchased with binding purchase orders. Budgetary quotations have been obtained from vendors and contractors for major equipment and unit rates. “Budgetary quotation” generally means that indicative pricing has been provided for specified equipment, materials and productivity based on the available information and engineering but no commitment has been made to provide the equipment or materials at this price at a future date.

21.1.9 TAXES

All taxes are excluded in the capital cost estimate. B2Gold Namibia expects to receive approval from the Namibian Ministry of Finance for an exemption on VAT tax for construction equipment and materials and approval from the Namibian Ministry of Trade and Industry on the Export Processing Zone License being exempted from all related import customs duties in connection with imported goods related to construction.

21.1.10 PROJECT CURRENCY, ESTIMATE BASE DATE AND FOREIGN EXCHANGE

All project capital costs were modelled in the anticipated payment currency, or the currency the equipment or materials were quoted in. Exchange rates were applied in the model and all project capital costs were then expressed in United States dollars (US$, USD) with the following provisions:

• Costs are based on December 2012 and prior market conditions with no provision for escalation beyond this date.

• Costs submitted in other currencies have been converted to US$ at the following exchange rates:

o 8.5 NAD : 1 USD (Namibian Dollar to US Dollar)

o 11.1909 NAD : 1 EUR (Namibian Dollar to Euro)

o 0.1079 NAD : 1 JPY (Namibian Dollar to Japanese Yen)

o 8.5 NAD : 1 CAD (Namibian Dollar to Canadian Dollar)

o 8.95 NAD : 1 AUD (Namibian Dollar to Australian Dollar)

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o 1 NAD : 1 ZAR (Namibian Dollar to South African Rand

• No provision has been made for any taxes or fees applicable to currency charges.

21.1.11 ACCURACY

The capital cost estimate, including contingency, for the mine, process plant, tailings storage facility and infrastructure has been prepared to a level of +/- 10%.

21.1.12 PROJECT IMPLEMENTATION

The capital cost estimate is based on the assumption that B2Gold Namibia will follow the project implementation plan as described. Any deviation from this plan may have a material impact on both the schedule and costs.

21.1.13 EXCLUSIONS

The capital cost estimate does not include allowances for:

• Escalation during construction

• Scope changes

• Interest and or financing costs during construction

• All surcharges and any other statutory taxation, levies or government duties; - Not receiving approval on the VAT tax and Customs duties exemptions

• Schedule delays and associated costs such as those caused by:

o Scope changes

o Unidentified geotechnical conditions

o Extraordinary climatic events

o Labor disputes

o Permit application delays

o Work permit delays

o Receipt of information beyond the control of the Owner or contractors

o Schedule recovery or acceleration

o Owners sunk costs

o Research and exploration drilling

o Sustaining capital

o Permitting costs

o Working capital

o Closure costs

o Salvage values

o Delays in primary equipment delivery

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o Surface mine development of pre-production waste stripping and ore stockpiling during the construction estimated at $33.3 million

o Intermediate rock excavation for foundations

o Hard rock excavation, including blasting for foundations

o Haulage of material greater than 2 km for disposal of excavated material and / or supplying suitable material for construction purposes

21.2 PROJECT DIRECT AND INDIRECT COSTS

21.2.1 QUANTITIES AND UNIT PRICING

Engineering material take-offs have been based on quantities derived by the engineering groups from the Otjikoto Project general arrangement drawings, sketches and similar projects.

21.2.2           EARTHWORKS

Quantities have been based on: topographic drawings at 2 metre contour intervals, findings from SRK’s geotechnical investigations at site and DRA’s past experience.

The earthworks unit rates were calculated based on data obtained from regional contractors and DRA’s historic experience for similar projects. The contractor’s unit rates are all inclusive for supplying, operating and maintaining the equipment. An additional cost of 25% was included for Preliminary and General (“P&G”) based on DRA’s past experience for similar projects.

It has been assumed that structural backfill, granular base, road base and sub-base will be supplied from local borrow pits. The unit costs associated with these materials include borrow pit development, crushing and screening and transport and placement costs.

21.2.3           CONCRETE, FORMWORK AND REINFORCING STEEL

Civil quantities were quantified from the drawings and layouts by Holley & Associates (Structural and Civil Consulting Engineers). For minor items of construction, quantities were estimated or factorised by DRA.

A preliminary bill of quantities was issued to major civil construction companies to obtain rates & P&G’s.

A commercial adjudication was carried out and one of the quotations was chosen. The estimate was then populated with the final quantities from Holley & Associates and the rates from the chosen civil contractor.

The civil preliminary and general (P&G) have been taken as 30% based on the figure received from the civil contractor.

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21.2.4           STRUCTURAL STEEL

Steelwork quantities for major structures were estimated by Holley & Associates from General Arrangement and layout drawings produced by DRA. Quantities for minor structures were estimated or factorised by DRA

A bill of quantities was produced by DRA to cover all steelwork and ancillary equipment (sheeting, grating, stairtreads, handrailing, etc.) and this was issued on enquiry to prominent steelwork fabricators/erectors. A commercial adjudication was carried out and one of the quotations was selected. The estimate was populated with the final steelwork quantities from Holley & Associates and the rates from the chosen contractor.

Structural steel P&G’s have been taken as 40% of the total supply and installation cost, in line with the figure quoted by the selected contractor.

21.2.5           MECHANICAL EQUIPMENT

From the Equipment List and PFD’s, an enquiry register was compiled and enquiries were issued to multiple vendors for each piece of major and minor mechanical equipment. Technical adjudications of the quotations were carried out and the pricing information for each item was incorporated into the estimate. The erection cost for the mechanical equipment was based on rates received from the steelwork & platework fabrication/erection contractor. For minor and/or ancillary items the erection costs were taken from previous quotations and/or the DRA historical data base.

The mechanical P&G’s were taken at 40% of the erection component only, as quoted by the steelwork & platework contractor.

21.2.6           PLATE WORK, LININGS AND TANKS

The Plate Works, Linings and Tanks were itemized and priced in accordance with the mechanical equipment list, the process flow sheets, process design criteria and the piping and instrumentation diagrams. Plate work weights were calculated with allowances made for any necessary stiffeners, weirs, launders, etc. The unit prices include regionally available plate purchase, detailing, fabrication and installation.

The platework P&G’s were taken at 40% of the total supply and installation cost, in line with the contractor’s quotation.

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21.2.7           PIPING AND VALVES

A bill of quantities was produced by DRA to cover all piping and valves by way of a detailed Piping material take off generated from preliminary routings of all lines shown on the P&ID’s and using the plant area GA’s to quantify lengths and quantity of fittings. This was issued on enquiry to prominent piping fabricators/erectors. A commercial adjudication was carried out and one of the quotations was selected. The estimate was populated with the final piping quantities from DRA and the rates from the chosen contractor.

The manual valves were costed by going out on enquiry with a Valve Schedule taken off the P&ID’s. All Actuated Vales were costed as part of the Control & Instrumentation Bill of Quantities.

Overland and non-process piping was quantified and priced in the infrastructure section of the estimate.

Piping P&G’s were assigned at 72.17% of the total supply and installation cost

21.2.8           ELECTRICAL, CONTROL AND INSTRUMENTATION

From the Mechanical Equipment List and Process Flow Diagrams, the DRA Electrical Department produced motor lists and instrument lists and then quantified and specified all electrical, control and instrumentation equipment. This was issued for enquiry to prominent electrical vendors.

From the General Arrangement drawings and Block Plan, the DRA Electrical Department determined the lengths and quantities of cabling, racking etc. This information was sent out on enquiry to prominent vendors.

Contractors for the supply of all equipment were chosen, as was a contractor for the installation portion of the work. Rates from the chosen contractors were used to populate the estimate.

Electrical & Instrumentation P&G’s were assigned at 14.52% of the total supply and installation cost.

21.2.9           BUILDINGS

Buildings were sized to accommodate the requirements of the process plant, mine and administration areas. The buildings were then priced with suppliers of pre-fabricated building manufacturers. The selected pricing for the building and erection was then included in the estimate.

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The building P&G has been taken as 25% based on the quotation received from the contractor.

21.2.10           BELT CONVEYORS

A preliminary design of all belt conveyors was carried out by the DRA engineering department in accordance with the belt profiles as depicted on the general arrangement drawings, the duty to meet the process requirements and the design to meet the general engineering design criteria. Calculations for each conveyor were carried out and the mechanical equipment components and steelwork content were quantified. Vendor quotations for the mechanical equipment were used and the steelwork fabrication rates were utilised from the steelwork vendor as described in section 4.4. The erection cost for the mechanical equipment and steelwork was based on rates received from the steelwork & platework fabrication/erection contractor. For minor and/or ancillary items the erection costs were taken from previous quotations and/or the DRA historical data base.

The mechanical equipment portion of the P&G’s was taken at 40% of the erection component, whilst the steelwork portion of the P&G’s was taken as 40% of the supply and erection cost as quoted by the steelwork & platework contractor.

21.2.11           FREIGHT

Freight costs for mechanical equipment were obtained from vendor quotations and were included in the mechanical equipment costs. Freight costs for steelwork and plate work were quoted by the fabricator/erector and included in these respective costs. Other freight costs were based on a tonnage or “per load” basis as estimated by DRA based on previous experience at similar sites.

21.2.12           ENGINEERING, PROCUREMENT AND CONSTRUCTION MANAGEMENT

The Engineering, Procurement and Construction Management (“EPCM”) costs cover the project management, detailed engineering, and procurement and construction management costs directly associated with the implementation of the project.

Project Management and EPCM costs relating to the process plant and associated infrastructure have been based on EPCM man hours calculated by DRA. The rates used are DRA 2012 rates. The man hours were calculated using the implementation program and project team organization and are shown in the EPCM man hour schedule, which also includes reimbursable and sundry costs. These DRA EPCM costs were then decreased by approximately 25% to recognize that B2Gold Namibia would utilize key expatriate supervisors to perform a combination of the EPCM combined with multiple prime engineering contracts and some self-performed owner construction. These key expatriate supervisors have worked on previous successful projects associated with B2Gold and Bema Gold Corporation. An additional $2.4 million was added to the estimate for the owners’ construction management costs.

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21.2.13           FIRST FILLS AND SPARES

Allowances have been made for the First Fills of process grinding media and reagents and for consumables based on DRA’s estimate of the requirements.

Spare parts costs have been included in the estimate based on a factorised estimate of the equipment estimates. Consideration was also given to vendors recommendations of the requirements to cover strategic and operational spares.

The provision for spares was changed by B2Gold following a change in the implementation strategy.

21.2.14           PRELIMINARY AND GENERAL

P&G costs include all contractors’ overheads such as contractual requirements (safety, sureties, insurance, etc.), the site establishment and the removal thereof, and company and head office overheads. They also cover supervision, travel to and from the site, contractor supplied temporary facilities, offices and lay-down areas, tools and contractors’ equipment.

P&G costs have been allowed for at a percentage of the value of the works to be executed and are shown separately against each estimate sub-heading at the bottom of the estimate summary sheet. The P&G cost was determined by DRA based on previous historical experience at similar sites.

21.3 SURFACE MINE

The surface mine equipment capital cost is estimated based on the quantity of equipment required to achieve mine production. The surface mine equipment requirements were prepared by VBKOM and the project management team as based on the mine development plan. The suppliers / vendors provided budgetary quotations on the equipment, including delivery and erection to site. The purchase price is from the actual vendor quotes and then B2Gold Namibia anticipated they would then enter into a lease to own equipment fleet agreement based on the general terms of 5% interest per year over a 5 year amortization schedule.

Table 21.3 shows the estimated mine mobile equipment capital purchasing schedule. The initial surface mining capital is estimated to be US$18.0 million (this is the total of the lease payments and miscellaneous capital purchases during the pre-production period). This does not include the surface mine development of pre-production waste stripping and ore stockpiling during the construction period estimated at $33.3 million.

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Table 21.3:           Surface Mine Mobile Equipment Capital Purchase Schedule

Table 21.3 Surface Mine Mobile Equipment Capital Purchase Schedule
SUMMARY			Total Cost Pre- Production (USD$)
Surface Mine - Lease			15 971 064
Surface Mine - Non-Lease			2 119 779
Total (USD$)			USD 18 090 844
SURFACE MINE - LEASE	Pre-Production Quantity 2013 - 2014	Unit Cost (USD$)	Total Cost Pre-Production (USD$)
Astra HD9 44.42 Lube Truck	2	215 323	430 646
Astra HD9 64.38 Service Truck	2	298 806	597 611
Astra HD9 64.54 Tractor Truck	1	190 617	190 617
CAT 16 M - Grader	2	901 289	1 802 577
CAT 730 DB - Diesel Tanker 25000L	1	467 760	467 760
CAT 730 WB - Water Tanker 27000L	2	467 760	935 519
CAT 777 D - Mine Truck 110t	13	1 318 645	17 142 391
CAT 834 H - Wheel Dozer	1	1 027 016	1 027 016
CAT 988 H - Loader 6.4m3	1	779 957	779 957
CAT 992 K - Loader 10.7m3	2	1 954 812	3 909 624
CAT CS76XT Vibratory Compactor 17T	1	175 496	175 496
CAT D10 T - Track Dozer	1	1 117 242	1 117 242
CAT D9 T - Track Dozer	1	790 820	790 820
CAT MD6290 Surface Mine Drill	6	1 374 670	8 248 021
Feeler FD160 Forklift 16T	1	215 882	215 882
Feeler FD50 Forklift 5T- All Terrian	1	48 824	48 824
Hitachi EX 1900 Backhoe 12.0m3	1	3 319 881	3 319 881
Hitachi EX 1900 Shovel 12.0m3	2	3 355 750	6 711 499
Hitachi ZW310 Loader w/ MC3200 Tire Handler	1	597 412	597 412
Hitachi ZX670LCR-5G Excavator w/ Sandvik Rammer BR7013 Rockbreaker, Quick Coupler	1	1 064 440	1 064 440
Ingersoll Rand Compressor	1	17 647	17 647
Martin RGN1207244WB Lowboy Trailer 120T	1	321 882	321 882
New Holland B110B Backhoe, Loader	1	74 899	74 899
New Holland LM 1745 High Lift Platform	1	136 809	136 809
New Holland SSL223 Skid Steer Loader	2	45 294	90 588
Pilot Crushtec VT1A 9m Mobile Tower Light w/ generator	12	32 941	395 294
Terex AC100/4 Rough Terrian Crane 100T	1	1 384 008	1 384 008
Total (USD$)			USD 48 083 728
LEASE PAYMENT (USD$)			USD 15 971 064
Note Partial payment during pre-production for 1 each CAT MD6290 Drill, 1 each CAT 992
Loader and 1 each Hitachi EX1900 Shovel with balance of payment paid during production.
SURFACE MINE - NON-LEASE	Pre-Production Quantity 2013 - 2014	Unit Cost (USD$)	Total Cost (USD$)
Toyota Hilux 3.0D 4x4 - Pickup Double Cab	21	41 518	871 880
Toyota Quantam Personnel Bus - 14 passenger	4	44 883	179 531
MTC A 150/5 Multi-stage Diesel Pump w/ Discharge Pipeline	2	89 412	178 824
Miscellaneous Office Equipment	1	889 544	889 544
Total (USD$)			USD 2 119 779

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21.4 PROCESS FACILITIES

The process facilities equipment capital cost is estimated based on the major process equipment, ancillary equipment, and infrastructure components necessary to process a nominal 2.5 million tonnes of ore per year from the surface mine deposit. The purchase prices for the major equipment are from actual vendor quotes as requested by DRA from the vendors and received during the DFS period. Table 21.4 lists the estimated process initial capital purchases by area. The initial pre-production process plant capital is estimated to be US$106.3 million.

Table 21.4:           Process Plant Initial Capital Purchase

SUMMARY	Total Cost Pre-Production (USD$)
Process Plant	106,279,096
Total (USD$)
PROCESS PLANT	Total Cost Pre-Production (USD$)
Civils	8,553,124
Structural	8,388,857
Platework Supplies	10,055,367
Mechanical Supplies	35,895,716
Piping Supplies & Erection	4,682,147
Electrical & Instrumentation Supply & Installation	16,590,829
Freight	3,988,786
Project Services	343,580
P&G - Preliminary & General	17,780,689
Total (USD$)	USD 106,279,096

21.5 ELECTRICAL & POWER GENERATION

The site construction power and the construction camp power will be taken from Cenored line power and distributed.

The permanent electrical and power generation capital cost estimate is based on the process flow diagrams, process design criteria, mechanical equipment list, site layout and general arrangement drawings, and the electrical single line diagrams. The electrical power generation will be performed by a HFO power generation plant capable of providing 15 MW of installed generating capacity as derived from the motor list by DRA.

Budget quotations were obtained from vendors for the HFO power generation plant based on these specifications. Vendor representatives will be engaged to oversee the installation of the larger equipment. Where minor equipment was not quoted by a vendor, DRA has estimated the cost from similar projects.

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The purchase prices for the major equipment are from actual vendor quotes as requested by DRA from the vendors and received during the DFS period. Table 21.5 lists the estimated electrical and power generation initial capital by area. The initial pre-production electrical and the leased power generation capital is estimated to be US$8.4 million.

Table 21.5:           Electrical and Power Generation Initial Capital

SUMMARY	Total Cost Pre-Production (USD$)
Electrical and Power Generation - Lease	7,121,299
Electrical and Power Generation - Non-Lease	1,248,495
Total (USD$)	USD 8,369,793
ELECTRICAL & POWER GENERATION - LEASE	Total Cost Pre-Production (USD$)
Generator Plant (Wartsila Equipment)	21,361,453
Generator Plant Civil, Structural, Building (70mX21.45m)	7,389,112
Total (USD$)	USD 28,750,565
LEASE PAYMENT (USD$)	USD 7,121,299
ELECTRICAL & POWER GENERATION - NON-LEASE	Total Cost Pre-Production (USD$)
Electrical Supplies, MCC	809,281
P&G - Preliminary & General	439,214
Total (USD$)	USD 1,248,495

21.6 TAILINGS STORAGE FACILITY

Epoch performed the feasibility level engineering design and cost estimate. The quantities were based on the site topographic drawing to 2 metre contour intervals and the recommendations from the SRK geotechnical investigation. The unit rates used in the cost estimate are based on the rates from DRA in the earthwork and civil (concrete) unit rates.

An additional cost for unmeasured items was included at 15% of the construction costs.

The P&G cost of 15% is only on the supply and installation of the liner in the tailings storage facility since B2Gold Namibia has included costs for procuring the mobile equipment necessary for the owner to self-perform these earthworks. Table 21.6 outlines the initial capital costs for the TSF.

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Table 21.6: Tailings Storage Facility Initial Capital

SUMMARY	Total Cost Pre-Production (USD$)
Tailings Storage Facility	25,311,262
Total (USD$)	USD 25,311,262
TAILINGS STORAGE FACILITY	Total Cost Pre-Production (USD$)
TSF Site Clearance:	996,176
TSF Earthworks:	3,929,100
TSF Drainage:	10,404,271
TSF Concrete Structures:	1,111,550
TSF Pipework:	1,467,607
TSF Miscellaneous:	245,590
RWD-Return Water Dam & SWD-Storm Water Dam Site Clearance:	15,494
RWD & SWD Earthworks:	481,895
TSF Drainage:	1,207,593
TSF Concrete Structures:	670,799
TSF Miscellaneous:	18,673
Unmeasured Items at 15%:	3,082,312
P&G Costs at 16%	1,115,496
Detailed Engineering Design	282,353
Construction Supervision, QC/QA	282,353
Total (USD$)	USD 25,311,262

21.7 CONSTRUCTION EQUIPMENT

The purpose of the mobile construction equipment is for B2Gold Namibia to be able to perform miscellaneous site construction and maintenance projects. The suppliers / vendors provided budgetary quotations on the equipment, including delivery to and erection on site. The purchase price is derived from the actual vendor quotes combined with B2Gold Namibia’s expectation to enter into a lease to own equipment fleet agreement based on the general terms of 5% interest per year over a 5 year amortization schedule.

Table 21.7 shows the estimated mobile construction equipment capital purchasing schedule. The initial surface mining capital is estimated to be US$2.7 million (this is the total of the lease payments and miscellaneous capital purchases during the pre-production period).

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Table 21.7:           Construction Equipment Capita Purchase Schedule

Table 21.7 Construction Equipment Capital Purchase Schedule
			Total Cost Pre-Production
SUMMARY			(USD$)
Construction Equipment - Lease			2 735 817
Total (USD$)			USD 2 735 817
	Pre-Production Quantity	Unit Cost	Total Cost Pre-Production
CONSTRUCTION EQUIPMENT - LEASE	2013 - 2014	(USD$)	(USD$)
CAT 988 H - Loader 6.4m3	1	779 957	779 957
CAT D9 T - Track Dozer	1	790 820	790 820
CAT CS76XT Vibratory Compactor 17T	1	175 496	175 496
Hitachi ZX670LCR-5G Excavator w/ Sandvik Rammer BR7013 Rockbreaker, Quick Coupler	1	625 882	625 882
Sandvik Crushing & Screening Plant	1	1 917 576	1 917 576
CAT 740 DT Articulated Dump Truck	3	732 872	2 198 616
Scania RR410LA 6x4 MSZ Tractor	3	165 059	495 176
Tridem 34m3 Twin Bin 3 axle side tipper semi-trailer	6	49 221	295 324
Total (USD$)			USD 7 278 847
LEASE PAYMENT (USD$)			USD 2 735 817

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21.8 EARTHWORKS

Quantities based on: topographic drawings at 2 metre contour intervals, findings from SRK’s geotechnical investigations at site and DRA’s past experience.

The earthworks unit rates were calculated based on data obtained from regional contractors and

DRA’s historic experience for similar projects. The contractor’s unit rates are all inclusive for supplying, operating and maintaining the equipment. An additional cost was included for P&G of 25% was derived based on DRA’s past experience for similar projects.

It has been assumed that structural backfill, granular base, road base and sub-base will be supplied from local borrow pits. The unit costs associated with these materials include borrow pit development, crushing and screening and transport and placement costs.

Table 21.8:           Earthworks Capital

SUMMARY	Total Cost Pre-Production (USD$)
Earthworks	14,162,386
Total (USD$)	USD 14,162,386
EARTHWORKS	Total Cost Pre-Production (USD$)
Roads	3,972,653
Bulk Earthworks - Terraces & Platforms	3,562,416
Stormwater Control & Management	220,491
Process Water Control	859,342
Raw Water Control	1,712,234
Potable Water Reticulation & Treatment	3,835,251
Total (USD$)	USD 14,162,386

21.9 MINE INFRASTRUCTURE, MINE BUILDINGS

The mine infrastructure was designed with the input from VBKOM, DRA and B2Gold Namibia’s project team. Costs for the infrastructure was based on estimated concrete quantities by DRA and on unit rates as described above in Section 21.2.3. The costs for the buildings, including erection were priced by DRA with regional suppliers of pre-fabricated building manufacturers.

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Table 21.9:           Mine Infrastructure and Mine Buildings Capital

SUMMARY	Total Cost Pre-Production (USD$)
Mine Infrastructure and Mine Buildings	7,059,627
Total (USD$)	USD 7,059,627
MINE INFRASTRUCTURE, MINE BUILDINGS	Total Cost Pre-Production (USD$)
Offices	356,648
Change House	479,112
EMV Workshop	2,422,685
EMV Workshop Washbay	72,400
Tyre Changing Workshop	319,749
LV Workshop	488,010
Diesel & Lube Storage Facility	29,400
EMV Washbay	123,080
Explosive Magazine	10,675
Non Explosives Magazine	3,663
Gate House	14,793
Ablution Block	1,469
Mining Warehouse	1,128,183
P&G - Preliminary & General	1,609,760
Total (USD$)	USD 7,059,627

21.10 BUILDINGS AND ANCILLARY FACILITIES

The quantities for the buildings and ancillary facilities were based on the site general arrangement drawings and building requirements as determined by DRA and B2Gold Namibia’s project team. The costs for these facilities were based on estimates by DRA as based on actual unit rates from contractors, historical experience from similar projects and pre-fabricated building vendors

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Table 21.10:           Buildings and Ancillary Facilities Capital

SUMMARY	Total Cost Pre-Production (USD$)
Buildings and Ancillary Facilities	7,647,843
Total (USD$)	USD 7,647,843
BUILDINGS AND ANCILLARY FACILITIES	Total Cost Pre-Production (USD$)
Fire Fighting Reticulation	570,272
Fire Prevention & Control	750,663
Sewer Treatment & Reticulation	526,595
General Buildings	1,118,357
Plant Buildings	4,681,956
Total (USD$)	USD 7,647,843

21.11 MECHANICAL & ELECTRICAL SPARES

DRA calculated the allowance for the mechanical and electrical spares was as follows: strategic spares are on 2.5% of the mechanical cost of the process plant equipment, operating spares are on 4% of the mechanical cost of the process plant equipment, pipes and valves are on 2% of the supply cost of the piping and valves and 7.5% of the electrical supply cost.

Table 21.11:           Mechanical and Electrical Spares Capital

SUMMARY	Total Cost Pre-Production (USD$)
Mechanical and Electrical Spares	3,279,148
Total (USD$)	USD 3,279,148
MECHANICAL AND ELECTRICAL SPARES	Total Cost Pre-Production (USD$)
Piping & Valves Critical Spares	78,619
Electrical Spares Critical Spares	1,204,410
Strategic Spares	1,996,119
Total (USD$)	USD 3,279,148

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21.12 OWNER’S COSTS

Owner’s costs were derived by B2Golds project team. These costs include the following:

• Staffing of key positions in the process plant and training of the operators and maintenance personnel
• Staffing of key positions and costs for normal consumables for the general site areas of administration, human resources, information technology, accounting, purchasing and warehousing, environment, medical and safety
• Contractor costs for providing site security, employee transportation to / from site and at site, catering costs and operating costs at the site construction camp
• Costs for the first fill of the lubricants at the process plant, process plant commissioning and SHE costs
• The costs for the general mobile equipment were from vendor quotations and previous estimates based on historical pricing

The costs were derived from the expected labor rates inclusive of full payroll overhead, estimates for normal office consumables based on B2Gold’s historical experience, contractor quotations for providing services and from DRA’s costs for the first fill lubricants and providing SHE management at site.

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Table 21.12:           Owners Costs

SUMMARY	Total Cost Pre-Production (USD$)
Owners Costs - Non-Leased	19,113,572
Owners Costs - Leased	755,854
Total (USD$)	USD 19,869,426
OWNERS COSTS - NON-LEASED	Total Cost Pre-Production (USD$)
Process Plant - Pre-Production	1,579,156
Administration	1,246,731
Human Resources	7,060
Information Technology	89,157
Accounting	448,412
Purchasing, Warehousing, Logistics	569,989
Environmental	125,006
Security	957,252
Safety	327,093
Medical Clinic	106,818
Employee Transportation	621,870
Cafeteria & Housing	5,287,173
Process Plant (First Fill Consumables, Training, Vehicles, Miscellaneous)	1,030,109
Site General (Office Equipment, Communications, Man Camp, Vehicles, Miscellaneous)	5,002,114
Process Plant (First Fill Lubricants)	167,700
General Non-Lease (Vehicles, Commissioning, SHE, Miscellaneous)	1,547,932
Total (USD$)	USD 19,113,572
OWNERS COSTS - LEASED	Total Cost Pre-Production (USD$)
New Holland SSL223 Skid Steer Loader	90,588
New Holland B110B Backhoe, Loader	74,899
Feeler FD50 Forklift 5T- All Terrian	48,824
Feeler FD160 Forklift 16T	215,882
Terex AC100/4 Rough Terrian Crane 100T	1,384,008
All wheel drive 8 ton platform truck with 5 ton crane	101,044
Total (USD$)	USD 1,915,245
LEASE PAYMENT (USD$)	USD 755,854

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21.13 OWNERS CONSTRUCTION MANAGEMENT

B2Gold Namibia’s construction costs were estimated by B2Gold project team to cover the labor and travel expenses for their construction management supervision personnel at site.

Table 21.13:           Owners Construction Management

SUMMARY	Total Cost Pre-Production (USD$)
Owners Construction Management	2,400,000
Total (USD$)	USD 2,400,000
OWNERS CONSTRUCTION MANAGEMENT	Total Cost Pre-Production (USD$)
Owners Construction Management	2,400,000
Total (USD$)	USD 2,400,000

21.14 ENGINEERING, PROCUREMENT AND CONSTRUCTION MANAGEMENT

The EPCM costs were derived by DRA based on historical experience at similar projects for an EPCM contractor to provide the necessary services. DRA costs have been calculated based on the development of an organization chart, estimating the number of personnel required. The labor costs are all inclusive and are based on DRA’s detailed EPCM labor schedule. These costs also include all travel, accommodations, site office expenses and other reasonable sundry expenses. The charge out rates are consistent with other projects with similar turnaround schedules.

The calculated EPCM cost was reduced by B2Gold as described above.

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Table 21.14:           Engineering, Procurement and Construction Management (EPCM)

SUMMARY	Total Cost Pre-Production (USD$)
Engineering, Procurement and Construction Management	13,595,085
Total (USD$)
EPCM	Total Cost Pre-Production (USD$)
Quantity Surveyors and Expediting	1,583,219
EPCM-Engineering	3,574,548
EPCM-Design	3,606,679
EPCM-Project Services	994,377
EPCM-Construction Management	2,528,578
EPCM-Disbursement	887,338
Plant Fire Protection System	89,216
Surveying Services	161,576
Geotech Investigation	70,588
Other Consultants	98,965
Total (USD$)	USD 13,595,085

21.15 CONTINGENCY

The overall contingency for the project development has been estimated by B2Gold as follows:

The mine equipment, construction equipment, owner’s costs and the owner’s construction management did not receive a contingency allowance.

The contingency applicable to the process plant (milling) estimate was by estimated by DRA by assessing the risks applicable to estimating inaccuracies. Which has been assessed (as shown below) by reviewing the engineering and estimating confidence pertaining to the overall supply and erection costs within each cost discipline.

The estimating inaccuracy allowance has then been assessed as a percentage for the individual cost components, from which an overall estimating inaccuracy has been derived.

The estimating inaccuracy is assessed as a measure of the confidence in the design and estimating processes/outputs. The confidence in each design and engineering discipline was assessed against the following scale:

HH	(High confidence)	5%
H	(High Confidence)	7.5%
HM	(High Medium Confidence)	10%

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M	(Medium Confidence)	12.5%
ML	(Medium Low Confidence)	15%
L	(Low Confidence)	20%

The assessment produces an average assessed confidence level for each discipline. The estimating inaccuracy allowance addresses the inaccuracies in the quantity, rates and prices used in the estimate and does not address any risks associated with the unknown or scope changes.

The overall contingency and accuracy level of the process plant and infrastructure estimate was determined by DRA to be 9.58% .

The contingency for the plant was later reduced by B2Gold.

Table 21.15:           Contingency

SUMMARY	Total Cost Pre-Production (USD$)
Contingency	15,501,396
Total (USD$)
CONTINGENCY	Total Cost Pre-Production (USD$)
Process Plant	8,383,108
Electrical & Power Generation	1,475,989
Tailings Storage Facility	1,931,582
Earthworks	1,424,180
Mine Infrastructure, Mine Buildings	648,534
Buildings, Ancillary Facilities	764,676
Mechanical and Electrical Spares	409,894
Process Plant (First Fill Lubricants)	20,963
General Non-Lease (Vehicles, Commissioning, SHE, Miscellaneous)	125,080
General - Leased	95,762
EPCM	221,629
Total (USD$)	USD 15,501,396

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Table 21.16:           Life of Mine Sustaining Capital Costs

SUMMARY	Total Cost (USD$)
Sustaining Capital	109 819 237
Total (USD$)	USD 109 819 237
Life of Mine Sustaining Capital	Total Cost (USD$)
Surface Mine - Lease	61 125 810
Surface Mine - Non-Lease	4 352 427
Construction Equipment - Lease	5 471 635
Process Plant	5 176 471
Electrical & Power Generation - Lease	25 428 991
General- Lease	1 511 708
General - Non-Lease	953 665
Site General	5 798 529
Total (USD$)	USD 109 819 237

Table 21.17:           Surface Mine Life of Mine Sustaining Capital Costs

SUMMARY	Total Cost (USD$)
Surface Mine Sustaining Capital	70 949 872
Total (USD$)	USD 70 949 872
Life of Mine Sustaining Capital	Total Cost (USD$)
Surface Mine - Lease	61 125 810
Surface Mine - Non-Lease	4 352 427
Construction Equipment - Lease	5 471 635
Total (USD$)	USD 70 949 872

Table 21.18:           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

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Table 21.19:           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

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Table 21.20:           Surface Mine Annual Operating Costs

Table 21.21:           Surface Mine Annual Production

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Table 21.22:           Life of Mine Unit Operating Costs by Activity

LIFE OF MINE SURFACE MINING COSTS (inclusive Pre-Stripping)	Total	% of Total
Drilling	$0.33	13%
Blasting	$0.52	21%
Loading & Hauling	$0.85	34%
Dewatering & Pumping	$0.07	3%
Dumps & Haul Roads	$0.11	4%
Mine General	$0.13	5%
Mine Technical Services	$0.10	4%
Mine Maintenance	$0.32	13%
Production Laboratory (Distributed)	$0.04	2%
Total (USD$) / Rock Tonne Mined (t-r)	$2.48	100%

Table 21.23 a:           Life of Mine Unit Operating Costs by Consumable Category

LIFE OF MINE SURFACE MINING COSTS (inclusive Pre-Stripping)	Total	% of Total
O&M Labor Cost	$0.30	12%
Equipment O&M Cost (Maintenance) (excludes Diesel Fuel)	$0.48	19%
Production Laboratory (Distributed)	$0.04	2%
Drilling Supplies: Bits, Rods, Hammers, Subs	$0.07	3%
Explosives Supplies	$0.46	19%
Contractor Services	$0.32	13%
Fuel: Diesel & Gasoline	$0.65	26%
Road Dust Stabilizer - Dust-A-Side	$0.03	1%
Other	$0.13	5%
Total (USD$) / Rock Tonne Mined (t-r)	$2.48	100%

Table 21.23 b:           Life of Mine Site General Unit Operating Costs by Activity

	Total Cost
SITE GENERAL	(USD$)
Administration	1.36
Human Resources	0.04
Information Technology	0.04
Accounting	0.14
Purchasing, Warehousing, Logistics	0.21
Environmental	0.04
Environmental Laboratory	0.02
Security	0.18
Safety	0.13
Medical Clinic	0.03
Employee Transportation	0.22
Cafeteria & Housing	0.06
Total (USD$) Per Ore Tonne Processed	USD 2.47

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21.16 PROCESS PLANT OPERATING COSTS

The annual process plant operating costs for the 2.5 million ore tonne per year process plant are depicted in Table 21.24. This table provides the overall estimated annual operating cost summary for the process plant and is based on 2.5 million ore tonnes per year with a mill availability of 93.7%, gold recovery of 95.5%, ore rehandle of 35% and 365 scheduled operating days per year. The Table 21.25 shows the annual process plant production. Tables 21.26 and 21.27 show the average operating costs by activity and by primary consumable categories.

The life of mine process plant operating costs is estimated to be US$14.02 per tonne of ore processed at the processing rate of 2.5 million tonnes per year.

The Sodium Cyanide price of $1770 per tonne delivered to site was obtained from OCD chemicals by B2Gold, this price was significantly less than the other quoted prices received. The reagent pricing received for Sodium Cyanide varied from $1770 per ton to $4692 per ton with an average price of $3221 per ton. Due to the large variation in quoted prices for Sodium Cyanide, the operating cost variation relative to Cyanide pricing was included in the process plant operating cost risk analysis as presented in Section 7 of the Otjikoto Definitive Feasibility Study – Decemeber 2012 and below.

21.16.1           PROCESS PLANT OPERATING COST RISK ANALYSIS

A risk analysis was performed to establish the operating cost variability, based on test work variability results obtained for the grinding circuit power consumptions, plant reagent consumptions and sensitivity to the Sodium Cyanide reagent price. The results of this risk analysis are present in Figure 21.1 below:

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Figure 21.1:           Estimated Plant Opex based on mill power and reagent consumptions from variability test work

Based on the fact that there was less data for the XR1 and XR2-3 (XR4) ore types on which to base the operating cost variability analysis, these operating cost estimates would be considered to be less accurate than for XR2 and XR3. For XR1 the 90% confidence interval for operating cost is 12.91$ /t – 14.97$ /t. For XR2 the 90% confidence interval for operating cost is 14.15$ /t-17.16$ /t. For XR3 the 90% confidence interval for operating cost is 13.72$ /t – 16.34$ /t. For XR2-3(XR4) the 90% confidence interval for operating cost is 14.13$ /t – 16.52$ /t. Based on this variability analysis an average life of mine operating cost range of 13.84$ /t – 16.44$ /t is expected based on the process plant production schedule as provided by B2Gold. The variation in operating cost for all ore types was driven by the variation in sodium cyanide cost, leach cyanide consumption, SMBS consumption and milling circuit power consumption. These variances are summarized in Figure 21.2 -Figure 21.8

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Figure 21.2: Estimated Plant Opex Variability for XR1

Figure 21.2: Factors Affecting Plant Opex Variability for XR1

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Figure 21.3: Estimated Plant Opex Variability for XR2

Figure 21.4: Factors Affecting Plant Opex Variability for XR2

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Figure 21.5: Estimated Plant Opex Variability for XR3

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Figure 21.6: Factors Affecting Plant Opex Variability for XR3

Figure 21.7: Estimated Plant Opex Variability for XR2-3 (XR4)

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Table 21.24:           Processing Annual Operating Costs

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Table 21.25:           Processing Annual Production

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Table 21.26:           Life of Mine Unit Processing Operating Costs by Activity

LIFE OF MINE PROCESSING COSTS - OPERATING	Total	% of Total
Feed Crusher / Rehandle	$0.01	0%
Crushing	$0.46	4%
Grinding	$4.12	38%
Gravity Concentrate, Acacia Leach	$0.52	5%
Leaching	$2.09	19%
CIP Carbon In Pulp	$0.25	2%
Carbon Elution & Regeneration	$0.23	2%
Cyanide Destruction	$0.92	9%
Electrowinning & Refinery	$0.12	1%
Tailings Storage Facility	$0.46	4%
Water System: Fresh-Process-Reclaim	$0.36	3%
Reagent Systems	$0.52	5%
Metallurgical Laboratory	$0.07	1%
Process General	$0.39	4%
Production Laboratory (Distributed)	$0.24	2%
Total (USD$) / Ore Tonne Processed	$10.75	100%
LIFE OF MINE PROCESSING COSTS - MAINTENANCE	Total	% of Total
Feed Crusher / Rehandle	$0.21	7%
Crushing	$0.14	4%
Grinding	$0.40	12%
Gravity Concentrate, Acacia Leach	$0.10	3%
Leaching	$0.06	2%
CIP Carbon In Pulp	$0.16	5%
Carbon Elution & Regeneration	$1.52	46%
Cyanide Destruction	$0.10	3%
Electrowinning & Refinery	$0.10	3%
Tailings Storage Facility	$0.19	6%
Water System: Fresh-Process-Reclaim	$0.04	1%
Reagent Systems	$0.11	3%
Metallurgical Laboratory	$0.02	1%
Process General	$0.11	3%
Total (USD$) / Ore Tonne Processed	$3.26	100%
Total (USD$) / Ore Tonne Processed	$14.02

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Table 21.27:           Life of Mine Processing Unit Operating Costs by Consumable Category

LIFE OF MINE PROCESSING COSTS (inclusive operating and maintenance)	Total	% of Total
O&M Labor Cost	$0.86	6%
Equipment O&M Cost (Maintenance)	$2.91	21%
Production Laboratory (Distributed)	$0.24	2%
Electric Power & Generation	$5.57	40%
Reagents	$2.75	20%
Grinding Balls	$0.86	6%
Mill & Crusher Liners, Wear Steel	$0.63	4%
Other	$0.20	1%
Total (USD$) / Ore Tonne Processed	$14.02	100%

The annual process plant cost estimates above includes the following:

• Staffing and maintenance manpower complements and base salaries based on the Peterson pay scale, including a payroll burden of 36%

• Power consumption based on the estimated power draw by the equipment as estimated by DRA with the cost of the power based on the vendor quoted heavy fuel oil and estimated lube cost, the heavy fuel consumption and lube consumption based on estimates from vendor and historical performance from similar plants

• Reagent and grinding media consumption values based on actual metallurgical test work results for each ore type and the process design criteria. The reagent and grinding media prices based on vendor provided quotes.

• Maintenance costs based on an estimate of the capital costs and historical experience from similar plants.

• The labor costs have been estimated on the basis of having 4 crews, with 3 crews on and one crew off. The work schedule is 8 hour shifts per day with 244 days worked or 1952 hours worked per year. The rotation time is such that have the days off will be approximately half of the days worked.

21.16.2 SITE GENERAL OPERATING COSTS

The site general operating costs are based on B2Gold’s historical experience from similar projects and from vendor quotes for supplying services such as security, employee transportation. Staffing manpower complements and base salaries based on the Peterson pay scale, including a payroll burden of 36%. The site general operating costs and manpower are shown in Table 21.28 and 21.29 and 21.30.

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Table 21.28:           Site General Annual Operating Costs

Table 21.29:           Site General Annual Manpower

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Table 21.30:           Life of Mine Site General Unit Operating Costs by Activity

SITE GENERAL	Total Cost (USD$)
Administration	1.36
Human Resources	0.04
Information Technology	0.04
Accounting	0.14
Purchasing, Warehousing, Logistics	0.21
Environmental	0.04
Environmental Laboratory	0.02
Security	0.18
Safety	0.13
Medical Clinic	0.03
Employee Transportation	0.22
Cafeteria & Housing	0.06
Total (USD$) Per Ore Tonne Processed	USD 2.47

22 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 and 12,000 ounces of silver over the 12 years of full production from the process plant. Details of the reserve calculations and production schedules are shown in previous sections.

A Base Case pro forma cash flow statement was prepared. The pro forma statement projects potential revenues, refinery charges, transport costs and facility operating and capital costs to yield annual net cash flows which are then discounted to determine a project NPV. The pro forma cash flow statement is from 2013 through the life of the mine and does not include any corporate administration costs nor any costs prior to 2013.

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 NPV at 5% discount rate, 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.

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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.

The highest sensitivity for cumulative cash flow is the gold price followed by the currency exchange rate. A detailed analysis of these values and other metrics are contained further in this section.

Sensitivity analysis was assessed on three scenarios for Low, Base and High rates for these individual parameters: Gold price, Diesel fuel price, HFO price, and exchange rate for NAD to USD and to the local Namibian labor rate. The three sensitivity scenarios are shown in Table 22.1.

Table 22.1:           Sensitivity Analysis Parameters

Parameter	Low	Base	High
Gold Price (USD$ per ounce)	USD 1,350 /oz	USD 1,550 /oz	USD 1,750 /oz
Diesel Fuel Price (USD$ per liter)	USD 0.7925 /L	USD 1.0565 /L	USD 1.3206 /L
HFO Heavy Fuel Oil Price	USD 0.5903 /L	USD 0.7881 /L	USD 0.9838 /L
Exchange Rate: NAD:USD	7.6500NAD : 1USD	8.5000NAD : 1USD	9.3500NAD : 1USD
National Labor	-25%	0% - Unchanged from 2013 rates	25%

22.1 CASH FLOW BASIS

22.1.1 MINERAL RESERVES

Mineral reserves were derived for the Otjikoto Project deposit. These reserves are presented in Section 15.0 “Mineral Reserve Estimate” of this report.

The mineral reserves were developed by applying cut-off grades and developing surface mine plans using appropriate mining and processing methods and estimated costs. The cut-off grades were approximately 0.40 g/t for the surface mine. Gold values were assessed at a price of US$1,350 per ounce gold and an estimated overall gold recovery of approximately 94.18% was used to estimate the cut-off grade. Silver content was not considered in calculating potential mineable resources.

Total life of mine ore mined and processed from the surface mine will be 29.4 million tonnes at an average grade of 1.42 grams per tonne. The total life of mine contained gold delivered from the surface mine is approximately 1,341,000 ounces. The total life of mine gold produced from the process plant is approximately 1,263,000 ounces.

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22.1.2 MINE PERMITTING AND DEVELOPMENT SCHEDULE

Pre-production development of the Otjikoto surface mine deposit is scheduled to begin in the second quarter of 2013. Site preparation for the construction of the process plant and ancillary facilities is scheduled to begin in the second quarter of 2013. The grinding mills and primary crusher and construction camp were secured and ordered in December 2012. The process plant is expected to be ready to receive ore in the fourth quarter of 2014. All permits that are required to construct the mine; process plant and ancillary facilities are in place.

22.1.3 MINE PLANS AND SCHEDULES

The surface mine commences pre-stripping in the second quarter of 2013 with approximately 8.1 million waste tonnes and approximately 1.0 million ore tonnes mined prior to the fourth quarter of 2014. The surface mine will maintain a high grade ore stockpile, greater than 0.6 g/t gold, and a low grade ore stockpile, greater than or equal to 0.4 g/t gold but less than 0.6 g/t gold, during the life of the surface mine. The mine development and production schedule was developed to optimize the gold grade of the ore delivered to the mill during the early production years. In addition to the initial pre-production development, the surface mine is scheduled to have two pushbacks during its mine life. The annual life of mine average rock tonnes mined per day from the surface mine is approximately 46,300 tonnes per day, with the minimum being approximately 39,100 tonnes per day and the maximum being 52,500 tonnes per day. The life of the surface mine is scheduled to be approximately 12 years. The surface mine is scheduled to produce approximately 29.4 million ore tonnes at 1.42 grams per tonne gold and 164.3 million waste tonnes with the strip ratio of 5.59 during the life of the mine.

The surface mine is scheduled to be mined out in mid-year 2025 while the process plant continues to process ore from the ore stockpile through mid-year 2026.

22.1.4 METALS PRODUCTION

Projected gold and silver production in doré is summarized in Table 22.2. The life of mine total gold recovery is estimated to be 95.5% and the silver recovery is estimated to be 60%. The gold production in doré is estimated to be approximately 1,281,000 ounces.

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Table 22.2:           Life of Mine Gold and Silver Production

	Years 1 - 5 Annual Average (ounces per year)	Life of Mine Annual Average (ounces per year)	Life of Mine Total (ounces)
YEAR
PRODUCTION
Gold Produced (troy ounces)	141,331	104,322	1,281,013
Silver Produced (troy ounces)	12,056	11,612	141,828

22.1.5 TRANSPORT, TREATMENT AND REFINING COSTS

The doré bars will be transported from the Otjikoto Project site to the yet to be designated refiner by a combination of highly secured transport vehicles and / or fixed rotor or fixed wing aircraft.

The costs associated with transporting, insuring and refining doré are listed in Table 22.3. These cost factors are typical of the industry and are based on B2Gold’s experience with other refineries.

Table 22.3:           Typical Transportation and Refining Terms

TYPICAL TRANSPORTATION AND REFINING TERMS (all costs in USD Dollars)
REFINING, TRANSPORTATION COSTS
Refinery Payable Metal Recoveries
Gold - Refinery Payable Recovery	99.90%	Estimate based on Dore Gold content over 50%.
Silver - Refinery Payable Recovery	99.00%	Estimate based on Dore Silver content.
Non-Payable Base Metal, Slag in Dore	3.00%	Estimate based on expected content.
Dore Transportation, Security, Insurance
Basic Charge	$3,050.00	Dedicated Service Basic Charge per shipment or lot.
Weight Surcharge	$10.00	Charge per kg.
Fuel Surcharge	$0.95	Charge per kg.
Screening Surcharge	$0.07	Charge per kg.
Refinery Charges
Refinery Treatment Charge	$0.50	Charge per net weight received troy ounces of Dore.
Assay Charge	$75.00	Charge per Lot.
	2	Estimate # of lots per month.
Early Settlement Fee - Finance Charge	0.04%	Estimate based on 2% per annum, for 7 days at value of gold.
Other Costs of Sale
Gold Transfer Charge	$0.20	Transfer Charge per gold ounce.
Silver Transfer Charge	$0.02	Transfer Charge per silver ounce.

22.1.6 ROYALTIES AND TAXES

The cash flows calculated on the pro forma include the estimated costs for the royalty and other taxes. The Minerals Act Royalty is 3% of the NPV (net sales) of gold and silver. The VAT tax is 15% VAT on domestic goods and services and 16.5% on imported goods and services. An exemption for imported project capital goods and services is being applied with the Ministry of Finance for approval of zero rated VAT exemption. This approval is expected to be received, thus the VAT tax on imported goods and services for the project capital would be reduced to 0%. A refund on the 15% VAT on domestic goods and services is expected to be approved and the expected refund cycle period is estimated to be two months. The Customs Duties and Fees has various rates for the various categories of items imported into Namibia. An exemption from the customs duties and fees for imported project capital goods and services has been applied for and approval is expected. The Normal Income Tax rate for corporations is 37.5% of profit less exemptions. The exemptions include all equipment and project capital costs, which are expensed in the year incurred or in the first year of production since there is no depreciation schedule. Unused exemptions may be carried forward and used in subsequent year(s) to reduce tax liability.

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22.1.7 OPERATING COSTS

Operating costs for the Otjikoto Project are summarized in Table 22.4. Details regarding the development of the operating costs are included in Section 21 “Capital and Operating Costs” of this report.

Table 22.4:           Life of Mine Cost of Production Unit Operating Cost Summary

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

22.1.8 SUMMARY OF PARAMETERS

Values of key parameters used during preparation of the pro forma cash flow statement are presented in Table 22.5. Included are such values as ore tonnage and grade, waste tonnage, metal recoveries, total metal production, capital and operating expenses and metal prices (Base Case shown).

This information is also found in Table 22.6; however, the summary in Table 22.5 is provided for ease of use and reference. Sensitivity analyses on metal price, diesel fuel price, heavy fuel oil price, exchange rate and national labor rate are presented in Section 22.4.

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Table 22.5a:           Base Case Parameters Used for the Cash Flow Analysis

	Life of Mine Total Cost USD$	Life of Mine Unit Cost per Tonne Processed USD$
Exchange Rates: NAD : ???		USD - United States dollar
NAD : USD	8.5	NAD - Namibian dollar
NAD : EUR	11.1909	EUR - Euro
NAD : JPY	0.1079	JPY - Japanese Yen
NAD : SEK	1.2984	SEK - Swedish Krona
NAD : CAD	8.5	CAD - Canadian
NAD : AUD	8.95	AUD - Australian
NAD : ZAR	1	ZAR - South African rand
Metal Prices
Gold Price (USD$ / troy ounce)	$1,550.00
Silver Price (USD$ / troy ounce)	$30.00
Unit Fuel Cost: (delivered cost to site)
Fuel: Diesel / AGO (NAD$ / Liter)	$8.98
Fuel: 180cst-HFO-Heavy Fuel Oil (NAD$ / Liter)	$6.69
PRODUCTION STATISTICS
Processing
Gold Produced (troy ounces)	1,281,013
Silver Produced (troy ounces)	141,828
Ore Tonnes Processed	29,408,899
Gold Grade Processed (g/t)	1.42
Recovery - Gold (%)	95.51%
Ore Tonnes Processed per Day	5,884
Surface Mine - Otjikoto
High Grade - Ore Tonnes Mined	18,204,242
Gold Grade Mined [>=0.60 g/t] (g/t)	1.99
Low Grade - Ore Tonnes Mined	11,204,657
Gold Grade Mined [0.40 >= g/t<0.60] (g/t)	0.49
Total Ore Tonnes Mined >=0.40 g/t	29,408,899
Total Gold Grade Mined (g/t)	1.42
Waste Tonnes Mined - Operations	127,140,279
Waste Tonnes Mined - Prestripping	37,125,766
Waste Tonnes Mined - Total	164,266,045

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Table 22.5b:           Base Case Parameters Used for the Cash Flow Analysis (continued)

	Life of Mine Total Cost USD$	Life of Mine Unit Cost per Tonne Processed USD$
REVENUE
Gross Production Value	$1,989,824,696	$67.66
Less: Refining, Transportation Costs	$4,484,144	$0.15
Net Production Value	$1,985,340,552	$67.51
COST OF PRODUCTION
Surface Mining - Otjikoto	$480,488,452	$16.34
Surface Mining - Otjikoto Prestripping	($94,046,483)	($3.20)
Processing	$413,787,284	$14.07
Site General	$82,334,943	$2.80
Total Cost of Production (less Prestripping)	$882,564,195	$30.01
OTHER OPERATING COSTS
Royalties	$59,560,217	$2.03
Taxes	$185,991,038	$6.32
Other (Income) Expense	($971,946)	($0.03)
Total Other Operating Costs	$244,579,308	$8.32
NON-OPERATING COSTS
Reclamation & Closure	$12,284,356	$0.42
Project Financing Costs	$0	$0.00
Total Other Non-Operating Costs	$12,284,356	$0.42
OTHER COSTS
CapX: EPCM-Process Plant	$114,662,205	$3.90
CapX: EPCM-Indirects	$21,299,616	$0.72
CapX: Owners Infrastructure	$93,822,832	$3.19
CapX: Owners Pre-Production & Sustaining	$108,783,755	$3.70
Surface Mining - Otjikoto Prestripping	$94,046,483	$3.20
Total Other Costs	$432,614,891	$14.71

22.2 PROJECT CASH FLOW

Table 22.6 shows the pro forma cash flow statement. Otjikoto Project development is shown to commence in 2013 with ore processing operations starting in late 2014. The surface mine ramps up to full production in 2014 and a significant ore stockpile is created and maintained during the life of mine. The surface mine is mined out in mid-year 2025. The process plant attains full production in the first quarter of 2015 and the last ore is processed in mid-year 2026. Note there are several years of reclamation and closure work required and some of it is expected to be completed concurrent with operations and the remaining work will be completed after operations and beyond 2028, the costs for which were consolidated in the final year.

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The cash flow statement in Table 22.6 “Pro forma Income Statement” section shows the revenue, cost of production, other operating costs, non-operating costs and other costs.

The cash flow statement in Table 22.7 “Production Summary” summarizes the processing production. Also note in this section the annual average metal grades, estimated processing recoveries and amount of gold produced.

Table 22.8, the “Net Production Value” section summarizes estimates of gross production value and the refinery schedule pay factors used to estimate refinery charges and fees.

Table 22.9, the “Summary of Surface Mining Costs” shows the costs for: drilling, blasting, loading and hauling, dewatering and pumping, dumps and haul roads, mine general, mine technical services, mine maintenance and production laboratory. The surface mine production and ore stockpile tonnes and grade are shown. There is approximately 1,045,000 tonnes of ore mined and stockpiled prior to the commissioning of the process plant.

Table 22.10, the “Summary of Processing Cost” shows the costs for operation and maintenance in: feed crusher / rehandle, crushing, grinding, gravity concentrate / acadia leach, leaching, carbon in pulp, carbon elution / regeneration, cyanide destruction, electrowinning / refinery, tailings storage facility, water system: fresh-process-reclaim, reagent systems, metallurgical laboratory, process general, and production laboratory.

Table 22.11, the “Summary of Site General Cost” shows the costs for: administration, human resources, information technology, accounting, purchasing / warehousing / logistics, environmental, environmental laboratory, security, safety, medical clinic, employee transportation, and cafeteria & housing.

Table 22.12, the “Summary of Distributed Account Costs” shows the costs for: electric power and generation, production laboratory, vehicle O&M (operation & maintenance).

Table 22.13, the “Change in Ore Stockpile and Mill Work In Process Inventory” shows these costs.

Table 22.14, the “Other Costs” shows the change in supplies inventory costs.

Table 22.15, the “Reclamation and Closure Costs” shows these costs.

Table 22.16, the “Other Operating Costs” shows the royalty and tax costs and other income / expense including the salvage value.

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The NPV were calculated from the resulting cash flow at discount rates of 0%, 5%, and 8%. At a 5% discount rate, the project NPV for the Base Case is US$402.4 million pre-tax and US$243.4 million after tax with a Return on Investment (“ROI”) of 30.4% and 23.6% respectively. A simple payback occurs within the first two and a half years of operation.

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Table 22.6:           Pro forma Income Statement

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Table 22.7:           Production Summary

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Table 22.8:           Net Production Value

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Table 22.9a:           Summary of Surface Mining Costs

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Table 22.9b:           Summary of Surface Mining Costs (continued)

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Table 22.10a:           Summary of Processing Costs

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Table 22.10b:           Summary of Processing Costs (continued)

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Table 22.11:           Summary of Site General Costs

Table 22.12:           Summary of Distributed Account Costs

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Table 22.13           Change in Ore Stockpile and Mill Work In Process Inventory

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Table 22.14:           Other Costs

Table 22.15:           Summary of Reclamation and Closure Costs

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Table 22.16:           Other Operating Costs

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22.3 CASH FLOW ANALYSIS

The most significant parameter affecting projected revenues is future gold price. For this reason, three gold price scenarios were considered, the results from which are presented in the next section for comparison. These future gold prices were considered to be fixed during the life of mine period for the purpose of this analysis.

Table 22.17:           Cash Flow Analysis Cases

Parameter	Low Case	Base Case	High Case
Gold Price (USD$ per ounce)	USD 1,350 /oz	USD 1,550 /oz	USD 1,750 /oz

22.3.1 ECONOMIC RESULT SUMMARY

Table 22.18 presents the economic results for the three cases as derived from the pro forma cash flow statement.

Table 22.18:           Economic Results for Gold Price Cases

	Base Case	Low Case	High Case
Gold Price (USD$ per ounce)	USD 1,550 /oz	USD 1,350 /oz	USD 1,750 /oz
Pre-Tax NPV @ 0% Discount Rate	USD 658.8 million	USD 403.0 million	USD 914.7 million
Pre-Tax NPV @ 5% Discount Rate	USD 402.4 million	USD 222.3 million	USD 582.5 million
Pre-Tax NPV @ 8% Discount Rate	USD 299.7 million	USD 150.8 million	USD 448.6 million
Pre-Tax ROI	30.4%	20.1%	39.8%
Simple Payback - (Years from first production)	2.5 years	3.2 years	2.1 years

Note that in all cases, the life of mine total operating cash cost per gold ounce produced is US$689 per ounce. The life of mine total cash cost per gold ounce produced ranges from US$736, US$729 and US$742 per ounce respectively per Table 22.18, due to the effect of the royalty payment.

22.4 SENSITIVITY ANALYSIS

In addition to investigating differences in NPV and ROI due to forecasted gold prices, sensitivities were calculated for negative 25%, negative 10% and positive 10% and 25% change in revenue and operating cost and pre-production capital expenditures. Sensitivities were also conducted on the gold price, diesel fuel, heavy fuel oil, exchange rate and national labor cost. These sensitivity analyses are in the following tables and figures.

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Table 22.19:           NPV at 5% Discount - Sensitivity to Net Production Value, Operating Cost and PreProduction Capital

Net Production Value (Revenue)	Net Present Value (NPV) 5% Discount rate	Operating Cost	Net Present Value (NPV) 5% Discount rate	Pre-Production Capital	Net Present Value (NPV) 5% Discount rate
-25%	USD 52,982,241	-25%	USD 574,608,677	-25%	USD 439,824,319
-10%	USD 262,643,258	-10%	USD 471,293,833	-10%	USD 417,380,090
Base Case - 0%	USD 402,417,270	Base Case - 0%	USD 402,417,270	Base Case - 0%	USD 402,417,270
10%	USD 542,191,282	10%	USD 333,540,708	10%	USD 387,454,451
25%	USD 751,852,300	25%	USD 230,225,864	25%	USD 365,010,221

Table 22.20:           Sensitivity of Cash Flow to Gold Price, Diesel Fuel Price, Heavy Fuel Oil Price, Exchange Rate, National Labor Cost

Parameter	Approximate % Parameter Change from Base Case	Case	Pre-Tax: Life of Mine Cumulative Cash Flow (in millions)
Gold (USD$1350/ounce)	-13%	Low	USD 403 M
	0%	Base Case	USD 659 M
Gold (USD$1750/ounce)	13%	High	USD 915 M
Diesel (USD$0.79/Liter)	-25%	Low	USD 695 M
	0%	Base Case	USD 659 M
Diesel (USD$1.32/Liter)	25%	High	USD 623 M
HFO (USD$0.59/Liter)	-25%	Low	USD 707 M
	0%	Base Case	USD 659 M
HFO (USD$0.98/Liter)	25%	High	USD 611 M
Exch-Rate: 7.65NAD:1USD	-25%	Low	USD 539 M
	0%	Base Case	USD 659 M
Exch-Rate: 9.35NAD:1USD	25%	High	USD 757 M
National Labor (-25%)	-25%	Low	USD 685 M
	0%	Base Case	USD 659 M
National Labor (+25%)	25%	High	USD 632 M

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Figure 22.1:           Sensitivity of Cash Flow to Gold Price, Diesel Fuel Price, Heavy Fuel Oil Price, Exchange Rate, National Labor Cost

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23 ADJACENT PROPERTIES

23.1 MINERAL OCCURRENCES

Figure 23.1 illustrates the location of the Otjikoto Project in relation to known mineralization in the region. Several minor gold occurrences are present to the south of the Kombat mine; however, there is no known gold mineralization of economic interest within the immediate vicinity of the Otjikoto Project.

Several small-scale amethyst quarries are present on the property but are not of economic interest. Two mines of interest are located in the region, Okorusu Fluorspar and the Kombat Copper-Lead mines. A third mine, the famous Tsumeb (Ongopolo) Lead-Copper-Zinc-Silver Mine, is located 105 km northeast of the Otjikoto Project and while not located directly in the vicinity of Otjikoto, it is described here for reference to the “Tsumeb Type” regional base metal metallogeny.

The information provided in this section is from the sources indicated. The Qualified Person has been unable to verify any of the adjacent property information contained in this Section and any information contained herein is not necessarily indicative of mineralization on the Otjikoto Project.

Figure 23.1:           Mineral Occurrence map of region (after data from MME)

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23.2 OKORUSU FLUORSPAR MINE

The following description is modified after http://www.mindat.org/loc-5842.html and McDonald (2011).

The Okorusu open pit fluorite mine is located 35 km east of the Otjikoto Project and is part of a syenite-carbonatite ring dike complex where fluorite has replaced pegmatitic carbonatite within the contact metamorphic aureole around the complex. Mining at the locale has been conducted since the 1920’s with the current mine operated since 1988 by Okorusu Fluorspar (Pty) Ltd.

The Early-Cretaceous alkaline rocks/sövite complex intrudes Neoproterozoic Damara series quartzites, marbles and biotite schists. The metasedimentary rocks have undergone alkalic metasomatism (fenetization) in the vicinity of the intrusion. Fluorite occurs in veins and as replacement bodies in the marbles and biotite schists. Rare Earth Element (REE) bearing minerals occur in beforsitic carbonatite dikes and carbonate-fluorite-bearing metasomatites within the complex.

23.3 KOMBAT COPPER-LEAD MINE

Small scale mining of the Kombat copper-lead-silver mine started in the early 1920’s but the main commercial production took place between 1962 and 2008. Production from 1962 to 1981 is reported as 5.53 million tonnes at 2.32% copper, 1.94% lead and 18 g/t silver (Innes and Chaplin, 1986). The mine workings were flooded in 2008. In 2012 a Canadian junior mining company, Kombat Copper Inc., acquired the rights to the mine and is actively exploring the property.

The Kombat mine is situated on the northern flank of the Otavi Valley Syncline, with phyllites of the Kombat formation underlain by dolostone of the Hüttenberg formation. The ore bodies are structurally controlled zones comprised of epigenetic, hydrothermal and metasomatic replacement as well as fracture-fill Cu-Pb-(Ag) deposits falling under the general Tsumeb-type mineralization classification. Bornite, chalcopyrite and galena are the principal ore minerals in the hypogene zone with lesser amounts of sphalerite and tennanite. Chalcocite, digenite and malachite dominate in the supergene zone. Vanadium rich minerals are a common accessory along with the gangue minerals calcite, quartz and minor barite. The ore mineralization is believed to be associated with the second Damara tectono-thermal event and related to basin dewatering and migration of metalliferous brines up high angles faults and shear zones (Innes and Chaplin, 1986).

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23.4 TSUMEB (ONGOPOLO) LEAD-COPPER-ZINC-SILVER MINE

The Tsumeb mine operated between 1905 and 1996 with approximately 30 million tonnes mined at an average grade of 10% lead, 4.3% copper, 3.5% zinc, 100 g/t silver and 50 ppm germanium (Melcher, 2003). The ore deposit is a complex, cross cutting, pipe shaped body with a vertical extent in excess of 1,500 m. There are four principal types of ore present in the deposit: massive peripheral, manto, disseminated and stringer ores and oxides. Most of the ore occurs as replacement style mineralization within feldspathic sandstones. The sandstones are believed to have in-filled a karst collapse breccia pipe. Later mineralizing brines exploited the permeable sandstones, reacting in part with carbonaceous material, resulting in the deposition of the ore minerals. The principal ore minerals are galena, sphalerite, tennantite, enargite, bornite, chalcocite and djurleite (Cu₃₁S₁₆) with lesser amounts of chalcopyrite and germanite (Cu₂₆Fe₄Ge₄S₃₂). An additional 250 or more mineral species are present within the deposit indicating a complex fluid assemblage(s) and depositional environment (Lombard et al. 1986).

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24 OTHER RELEVANT DATA AND INFORMATION

B2Gold completed this feasibility study in February 2013 on the Otjikoto Project in Namibia entitled “Otjikoto Project NI 43-101 Technical Report Feasibility Study”. This NI 43-101 Technical Report summarizes all relevant data and sections from the DFS completed in December 2012 for B2Gold by DRA and VBKom and other Contributors. The DFS describes the development of the mineral resources, mining reserves, mining methods, processing selection and economics of the Otjikoto Project based on an estimate of costs and commodity prices.

There is no other relevant data and other information on the Otjikoto Project that has not been discussed in the other sections of this report.

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25 INTERPRETATION AND CONCLUSIONS

Results of the December 2012 DFS have confirmed the economic viability of the Otjikoto Project. Probable mineral reserves of 29.4 million tonnes of ore at a gold grade of 1.42 g/t containing 1.34 million ounces of gold have been established.

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. Since the completion of the DFS, B2Gold has advanced the project in the following areas:

• key members of the construction team have been hired,
• key Windhoek office personnel have been hired and satellite offices have been established in Otavi and Otjiwarongo,
• long lead time equipment orders have been placed for process plant equipment,
• negotiations for the mine fleet and power plant are at final decision points,
• a contract is in place and detailed design engineering is underway, and
• initial site work has started which included brush clearing, camp construction and fencing.

The Otjikoto Project will be developed as an open pit mine, where run-of-mine ore will be trucked to the plant, crushed, and then treated in a grinding circuit utilizing conventional SAG and ball mills, and a CIL recovery process.

The DFS estimates the annual production for the first five years is projected to be approximately 141,000 ounces of gold per year at an average operating cash cost of $525 per ounce, and for the LoM approximately 112,000 ounces of gold per year at an average operating cash cost of $689 per ounce. The total pre-production capital costs are estimated to be $244.2 million. Refer to Section 21 “Capital and Operating Costs” of this report for additional details.

The plant facility and support infrastructure will be built to support a plant expansion from the initial processing capability of 2.5 million tonnes per annum to 3.0 million tonnes with minimal additional capital expenditure.

The financial modeling for the Otjikoto Project indicates robust economics. At a gold price of $1,550 per ounce, the project is projected to yield a positive after-tax NPV of $243.4 million at a discount rate of 5%. The IRR after-tax is 23.6% . The project has a 32 month pay-back period after first gold production.

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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 increase annual average gold production.

The most significant parameter affecting project revenues is future gold price. Therefore, a sensitivity analysis was conducting using three different gold prices. In all cases, the simple payment ranged between 2.1 and 3.2 years. In addition to the sensitivity analysis on gold price, sensitivities were completed on diesel fuel price, heavy fuel oil price, exchange rate, and national labour costs. While these sensitivity analyses indicate robust economics, if these parameters vary outside of the sensitivity analyses, there is the potential to impact economics of the project.

Additionally, B2Gold is not aware of any pending legislation or regulatory changes that may change project development timelines or economics. If changes occur to the regulatory framework, it has the potential to negatively impact project development timelines and costs.

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26 RECOMMENDATIONS

Based on the interpretations and conclusions, 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. All of the government permits and licenses required for construction have been received.

The total pre-production capital costs are estimated to be $244.2 million, which includes working capital for supplies, taxes and owners costs.

Pre-Production Capital Costs	$ millions
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	2.7
Mine Infrastructure, Mine Buildings	7.1
Buildings, Ancillary Facilities	7.6
Owners Costs – Other	19.8
Mechanical & Electrical Spares	3.3
Owners Construction Management	2.4
EPCM	13.6
Contingency	15.5
Total	244.2

The current Otjikoto Project execution plan encompasses project management by B2Gold and utilizes a delivery method comprised of a combination of engineering, procurement, and construction combined with multiple prime engineering contracts and some self-performed owner construction.

Engineering and procurement will be managed from the B2Gold corporate offices in Vancouver, British Columbia, Canada with construction management occurring from the Otjikoto Project site. The majority of equipment and supplies will be sourced from South Africa, North America or European suppliers and will be ocean shipped to the seaport at Walvis Bay, Namibia, with subsequent overland delivery utilizing contracted commercial trucking firms.

The site development will take place year round, utilizing a work force of experienced Namibian nationals, trained and supervised by Namibian and expatriate supervisors. Many of the expatriate supervisors have worked on previous successful projects associated with B2Gold and Bema Gold Corporation.

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Logistics will be supported from the existing B2Gold support structure in place and operating in Namibia. Construction activities have commenced. The mills and primary crusher were ordered in December 2012. The grinding mills, crushing equipment and the construction man camp have been secured and partial payment has been made. Site construction work has started for the camp. It is anticipated that the HFO generators and the mining fleet will be procured or contracted for by the end of February 2013. The process plant is expected to be ready to receive ore in the fourth quarter of 2014.

B2Gold Namibia should continue to work closely with communities around the Otjikoto Project area to develop a corporate social responsibility (“CSR”) program that benefits all stakeholders. To date, B2Gold Namibia is supporting community driven projects in public health, education, development of small to medium enterprises and environmental conservation. Projects included scholarships to historically disadvantaged students, support to the Namibian Institute of Mining and Technology, a partnership with the Cheetah Conservation Foundation to monitor and protect endangered species, identification of potential small to medium enterprises within the communities where our workers live, support of educational needs with the Ministry of Education, support of programs that help at risk teens, and providing equipment and facilities to support local medical needs, and this work should remain a high priority.

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27 REFERENCES

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Wilton, J., Lombard, P. J. A. and Philpot, H. G. (2002). Discovery of the Otjikoto gold deposit in Namibia. Proceedings of the 11^th Quadrennial IAGOD Symposium and Geocongress 2002, 22-26 July 2002 Windhoek, Namibia.

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