Filed by Avantafile.com - Nevsun Resources Ltd.

EX-99.1 2 exhibit99-1.htm TECHNICAL REPORT Filed by Avantafile.com - Nevsun Resources Ltd. - Exhibit 99.1

Bisha Mine

Eritrea, Africa

NI 43-101 Technical Report

Prepared for:
Nevsun Resources Ltd.

Prepared by:
Paul Gribble, C. Eng., FIMMM,
   Bisha Mining Share Company J
Jay Melnyk, P.Eng.,
   Mining Consultants, Inc.
Peter Munro, BAppSc.,
   Mineralurgy Pty. Ltd.

Effective Date:
31 December 2013

IMPORTANT NOTICE

This report was prepared as a National Instrument 43-101 Technical Report for Nevsun Resources Ltd. (Nevsun) by AGP Mining Consultants Inc. (AGP) and Mineralurgy Pty. Ltd. (Mineralurgy). The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in AGP’s and Mineralurgy’s services, based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by Nevsun subject to the terms and conditions of its contracts with AGP and Mineralurgy. This contract permits Nevsun to file this report as a Technical Report with Canadian Securities Regulatory Authorities pursuant to National Instrument 43-101, Standards of Disclosure for Mineral Projects. Except for the purposes legislated under provincial securities law, any other uses of this report by any third party is at that party’s sole risk.

NEVSUN RESOURCES LTD.
BISHA MINE, ERITREA, AFRICA
NI 43-101 TECHNICAL REPORT

Table of Contents

1 Summary.	1
1.1. Location and Access.	1
1.2. Mineral Tenure, Surface Rights, and Royalties.	1
1.3. Permits.	2
1.4. Environment.	2
1.5. Geological Setting and Mineralization.	2
1.6. Drilling.	4
1.7. Sample Preparation, Analyses, and Security.	4
1.8. Data Verification.	5
1.9. Mineral Resource Estimate.	5
1.10 Mineral Processing and Metallurgical Testing.	9
1.11 Mineral Reserves Estimates.	11
1.12 Mine Plan.	13
1.13 Process.	14
1.14 Markets.	15
1.15 Capital and Operating Costs.	16
1.16 Conclusions.	16
1.17 Recommendations.	17
1.17.1 Mineral Resources.	17
1.17.2 Mineral Reserves.	18
1.17.3 Metallurgical	20
1.17.4 Exploration.	20
2 Introduction and Terms of Reference.	21
2.1. Qualified Persons, Principal Contributors, and Areas of Responsibility.	21
2.2. Site Visits and Scopes of Personal Inspections.	22
2.3. Effective Dates.	23
2.4. Information Sources and References.	23
2.5. Previous Technical Reports.	23
3 Reliance on Other Experts.	25
3.1. Mineral Tenure.	25
3.2. Surface Rights and permitting.	25
3.3. Environmental Liabilities.	25
3.4. Social and Community Impacts.	25
4 Property Description and Location.	26
4.1. Property and Title in Eritrea.	26

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4.2. Property Ownership.	28
4.3. Mineral Tenure.	29
4.4. Surface Rights.	30
4.5. Royalties and Encumbrances.	30
4.6. Property Agreements.	31
4.7. Permits.	31
4.8. Environmental Liabilities.	32
4.9. Social Licence.	32
5 Accessibility, Climate, Local Resources, Infrastructure and Physiography.	33
5.1. Accessibility.	33
5.2. Climate.	33
5.3. Physiography.	35
5.4. Local Resources and Infrastructure.	35
5.4.1 Local Resources.	35
5.4.2 Infrastructure.	36
5.4.3 Power.	36
5.4.4 Water.	36
5.4.5 Communications.	37
6 History.	38
6.1. Bisha and Harena.	38
6.2. Northwest.	38
6.3. Mogoraib River Exploration Licence.	38
6.3.1 Summary Prospect Descriptions.	42
7 Geological Setting and Mineralization.	46
7.1. Regional Geology.	46
7.2. Property Geology.	48
7.2.1 Stratigraphy.	49
7.2.2 Structure.	50
7.2.3 Metamorphism.	51
7.2.4 Alteration.	51
7.2.5 Weathering.	51
7.3. Deposit Geology and Mineralization.	51
7.3.1 Bisha Main Zone.	51
7.3.2 Harena Deposit.	67
7.3.3 Northwest Deposit.	68
7.3.4 Hambok Deposit.	82
8 Deposit Types.	88
8.1. Deposit Classifications.	88

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8.2. Previous Interpretation.	90
8.2.1 Bimodal Siliciclastic VMS Deposit Model	90
8.3. Recent Advances.	91
8.3.1 Bisha and Northwest Deposits.	91
8.3.2 Hambok Deposit.	92
8.3.3 Harena Deposit.	92
8.3.4 Seafloor Accumulation vs. Sub-seafloor Replacement.	92
9 Exploration.	95
9.1. Grids and Surveys.	95
9.2. Geological Mapping.	102
9.3. Geochemical Sampling.	102
9.3.1 Stream Sediment Sampling.	102
9.3.2 Rock Chip Sampling.	102
9.3.3 Soil Geochemical Sampling.	102
9.3.4 Termite Mound Sampling.	104
9.3.5 Soil pH Geochemical Sampling.	104
9.3.6 Auger Geochemical Sampling.	104
9.4. Remote Sensing and Satellite Imagery.	105
9.5. Geophysics.	105
9.5.1 Ground Geophysics.	105
9.5.2 Aerial Geophysics.	106
9.6. Pits and Trenches.	108
9.7. Petrology, Mineralogy, and Research Studies.	112
9.7.1 History.	112
9.7.2 Summary of Recent Research on the Bisha Deposit.	113
9.7.3 Summary of the Recent Research on the Northwest Deposit.	114
9.7.4 Summary of the Recent Research on the Hambok Deposit.	114
9.7.5 Summary of the Recent Research on the Harena Deposit.	114
9.8. Geotechnical and Hydrological Studies.	115
9.9. Exploration Potential	115
10 Drilling.	117
10.1 Bisha Deposit.	117
10.2 Harena Deposit.	119
10.3 Northwest Deposit.	120
10.4 Hambok Deposit.	121
10.5 Drill Methods.	122
10.5.1 Core Drilling.	122
10.5.2 Reverse Circulation Drilling.	123
10.6 Geological Logging.	124
10.6.1 Core Logging.	124

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10.6.2 Chip Logging.	125
10.7 Sample Recovery.	125
10.7.1 Bisha.	125
10.7.2 Northwest.	125
10.7.3 Hambok.	126
10.8 Collar Surveys.	126
10.9 Down Hole Surveys.	126
10.9.1 Bisha.	126
10.9.2 Northwest.	126
10.9.3 Harena.	127
10.9.4 Hambok.	127
10.10 Hydrological Drilling.	127
10.10.1 Bisha Deposit.	127
10.10.2 North West Bisha.	127
10.10.3 Harena.	127
10.10.4 Hambok.	127
10.11 Geotechnical Drilling.	128
10.12 Metallurgical Drilling.	128
11 Sample Preparation, Analyses, and Security.	130
11.1 Sample Collection and Preparation.	130
11.1.1 Trench Sampling.	130
11.1.2 Diamond Core Sampling.	130
11.1.3 Grade Control RC Sampling.	132
11.1.4 Bulk Density Sampling.	133
11.2 Sample Analysis.	133
11.2.1 Analytical Laboratories.	133
11.2.2 Analytical methods.	134
11.3 Quality Assurance and Quality Control	136
11.3.1 Nevsun/BMSC QA/QC Protocols for Drill Programs, 2003-2005.	136
11.3.2 BMSC QA/QC Protocols for Drill Programs, 2006-2010.	137
11.3.3 BMSC QA/QC Protocols for Drill Programs, 2011.	138
11.3.4 BMSC QA/QC Protocols for Drill Programs, 2012-2013 Diamond Core.	139
11.3.5 BMSC QA/QC Protocols for Drill Programs, 2013 Grade Control RC.	148
11.3.6 Sanu Resources QA/QC Protocols for Drill Programs, 2006-2010.	149
11.3.7 Sanu Resources QA/QC Protocols for Drill Programs, 2011-2012.	150
11.3.8 BMSC QA/QC Protocols for Drill Programs, 2013 Diamond Core, Hambok.	151
11.4 Databases.	154
11.5 Sample Security.	154
11.5.1 Chain-of-Custody.	154

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11.5.2 Sample Storage.	154
11.6 Author's Statement.	155
12 Data Verification.	156
12.1 Independent Sampling and QA/QC Audits by Cube.	156
12.2 Data Verification by BMSC.	157
13 Mineral Processing and Metallurgical Testing.	158
13.1 Metallurgical Testwork.	158
13.1.1 Metallurgical Samples.	159
13.1.2 Composite Samples.	162
13.2 Grinding Testwork.	163
13.2.1 2005 SGS Lakefield Test Program.	163
13.2.2 2010 Test Program.	164
13.2.3 2011-2012 Test Program.	165
13.2.4 2013 Northwest Program.	165
13.2.5 2013 Hambok Program.	165
13.2.6 2013-2014 Bisha Main Primary Program.	165
13.3 Mineralogy.	165
13.3.1 2005 SGS Lakefield Test Program.	165
13.3.2 Supergene Mineralization.	167
13.3.3 Transition Supergene and Pyrite Sand.	167
13.3.4 Northwest Satellite Deposit.	167
13.3.5 Hambok Deposit.	168
13.3.6 Primary Mineralization.	168
13.4 Cyanidation Testwork.	168
13.4.1 2005 SGS Lakefield Test Program.	168
13.4.2 2005 SGS South Africa.	169
13.5 Flotation Testwork.	169
13.5.1 2005 SGS Lakefield Test Program Oxide Mineralization.	169
13.5.2 2010 Mintek Test Program.	171
13.5.3 2010 and 2011-2012 Maelgwyn Test Programs.	171
13.5.4 Transitional Supergene and Pyrite Sand at ALS Metallurgy Kamloops.	172
13.5.5 2013 Northwest Satellite Deposit at ALS Metallurgy Kamloops.	172
13.5.6 2013 Supergene Mineralization at ALS Metallurgy Burnie.	173
13.5.7 2013 Hambok Deposit at SGS Mineral Services UK.	173
13.5.8 2013-2014 Primary Mineralization at ALS Metallurgy Kamloops.	173
14 Mineral Resource Estimates.	174
14.1 Bisha Main Estimate.	174
14.1.1 Project Sample Database.	174
14.1.2 Geological Model	176

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14.1.3 Alteration and Weathering.	181
14.1.4 Mineralization Domains.	182
14.1.5 Database Coding.	182
14.1.6 Compositing.	183
14.1.7 Bulk Density.	183
14.1.8 Sample Type.	184
14.1.9 Recovery.	185
14.1.10 Exploratory Data Analysis.	186
14.1.11 Variography.	191
14.1.12 Search Neighbourhood Analysis.	198
14.1.13 Block Model Parameters.	198
14.1.14 Grade Estimation.	199
14.1.15 Block Model Validation.	199
14.1.16 Classification of Mineral Resources.	200
14.1.17 Reasonable Prospects of Economic Extraction.	200
14.2 Harena Mineral Resource Estimate.	202
14.2.1 Summary.	202
14.3 Northwest Estimate.	204
14.3.1 Project Sample Database.	204
14.3.2 Local Grid Transformation.	204
14.3.3 Geological Model	205
14.3.4 Alteration and Weathering.	210
14.3.5 Mineralization Domains.	210
14.3.6 Database Coding.	212
14.3.7 Compositing.	212
14.3.8 Bulk Density.	213
14.3.9 Recovery.	213
14.3.10 Exploratory Data Analysis.	215
14.3.11 Variography.	217
14.3.12 Search Neighbourhood Analysis.	223
14.3.13 Block Model Parameters.	223
14.3.14 Grade Estimation.	224
14.3.15 Block Model Validation.	224
14.3.16 Classification of Mineral Resources.	225
14.3.17 Reasonable Prospects of Economic Extraction.	225
14.4 Hambok Estimate.	227
14.4.1 Project Sample Database.	227
14.4.2 Geological Model	228
14.4.3 Alteration and Weathering.	230
14.4.4 Mineralization Domains.	230
14.4.5 Database Coding.	231
14.4.6 Compositing.	231

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14.4.7 Bulk Density.	232
14.4.8 Recovery.	232
14.4.9 Exploratory Data Analysis.	232
14.4.10 Variography.	233
14.4.11 Search Neighbourhood Analysis.	233
14.4.12 Block Model Parameters.	233
14.4.13 Grade Estimation.	234
14.4.14 Block Model Validation.	234
14.4.15 Classification of Mineral Resources.	235
14.4.16 Reasonable Prospects of Economic Extraction.	236
14.5 Mineral Resource Statement.	238
15 Mineral Reserve Estimates.	242
15.1 Key Assumptions/Basis of Estimate.	242
15.1.1 Pit Slopes.	242
15.1.2 Net Smelter Return Calculations.	242
15.1.3 Operating Costs.	245
15.1.4 Pit Optimization and Pit Phase Design.	245
15.2 Dilution Adjustments.	246
15.3 Conversion Factors from Mineral Resources to Mineral Reserves.	246
15.4 Mineral Reserves Statement.	246
15.5 Factors that May Affect the Mineral Reserve Estimates.	248
16 Mining Methods.	249
16.1 Geotechnical	249
16.1.1 Overview.	249
16.1.2 Main Zone.	249
16.1.3 Harena Zone.	252
16.2 Pit Design.	254
16.3 Cutoff Grades.	256
16.4 Mine Plan.	258
16.5 Waste Rock Storage.	260
16.6 Blasting and Explosives.	262
16.7 Grade Control	263
16.8 Reconciliation.	264
16.9 Hydrogeology.	264
16.9.1 Pit Dewatering.	264
16.9.2 Runoff Water.	264
16.9.3 Fereketetet River Interception and Diversion.	265
16.10 Mining Equipment.	265
17 Recovery Methods.	267
17.1 Supergene Process Plant.	267

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18 Project Infrastructure.	271
19 Market Studies and Contracts.	273
19.1 Concentrate Quality.	273
19.2 Commodity Price Projections.	275
19.3 Contracts.	275
20 Environmental Studies, Permitting and Social or Community Impact.	276
20.1 Environmental Regulatory Framework.	276
20.2 Baseline Studies.	276
20.3 Environmental Issues.	277
20.4 Closure Plan.	277
20.5 Permitting.	279
20.6 Considerations of Social and Community Impacts.	280
20.7 Discussion on Risks to Mineral Resources and Mineral Reserves.	280
21 Capital and Operating Costs.	281
21.1 Capital Cost Estimates.	281
21.1.1 Introduction.	281
21.1.2 Process Capital Costs.	281
21.1.3 Tailings Management Facility Expansion.	281
21.1.4 Mine Capital Costs.	281
21.1.5 Sustaining Mill, Camp, and Port Capital	282
21.2 Operating Cost Estimates.	284
21.2.1 Basis of Estimate.	284
21.2.2 Mine Operating Costs.	284
21.2.3 Process Operating Costs.	284
21.2.4 G&A Operating Costs.	284
21.2.5 Downstream Costs.	284
21.3 Operating Cost Summary.	284
22 Economic Analysis.	286
23 Adjacent Properties.	287
24 Other Relevant Data and Information.	289
25 Interpretation and Conclusions.	290
26 Recommendations.	293
26.1 Mineral Resources.	293
26.1.1 Bisha.	293
26.1.2 Harena.	293
26.1.3 Northwest.	293
26.1.4 Hambok.	294

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26.2 Mineral Reserves.	294
26.2.1 Bisha.	294
26.2.2 Harena.	295
26.2.3 Northwest.	295
26.2.4 Hambok.	295
26.3 Metallurgical	296
26.4 Exploration.	296
27 References.	297
28 Certificate of Authors.	305

Tables

Table 1-1: Summary of Drilling for Each of the Deposits.	4
Table 1-2: Combined Mineral Resource Estimate - Bisha, Harena, Northwest, and Hambok Deposits (Effective Date 31 December 2013)	6
Table 1-3: Bisha Resource Estimate (Effective Date 31 December 2013)	6
Table 1-4: Harena Mineral Resource Estimate (Effective Date 31 December 2013)	7
Table 1-5: Northwest Mineral Resource Estimate (Effective Date 31 December 2013)	7
Table 1-6: Hambok Mineral Resource Estimate (Effective Date 31 December 2013)	8
Table 1-7: Bisha and Harena Reserves Estimate (Effective Date: 31 December 2013)	12
Table 2-1: Report Areas of Responsibility.	22
Table 2-2: Dates of Site Visits.	22
Table 5-1: Distances by Road to the Property.	33
Table 6-1: Sanu Resources Exploration Program 2003 - 2012.	38
Table 6-2: Sanu Reported Expenditure during Tenure of the Mogoraib EL.	45
Table 8-1: Diagnostic Criteria for Sub-Seafloor Replacement and Seafloor Accumulation.	93
Table 8-2: Bisha Deposits vs. Sub-seafloor Replacement Criteria Comparison*.	93
Table 9-1: Summary of Work Completed.	96
Table 9-2: Northwest Trenching 2012, Summary of Results.	111
Table 10-1: Summary of Drilling for the Bisha Main Deposit	119
Table 10-2: Summary of Drilling for the Harena Deposit	120
Table 10-3: Summary of Drilling for the Northwest Deposit	121
Table 10-4: Hambok Deposit Mogoraib EL Drill Hole Summary Table.	122
Table 10-5: Bisha Deposit Metallurgical Drill Hole Summary.	128
Table 10-6: NW Bisha Deposit Metallurgical Drill Hole Summary.	128
Table 10-7: Harena Deposit Metallurgical Drill Hole Summary.	129
Table 11-1: Analytical Laboratories used for drill samples.	133
Table 11-2: Assay Laboratory Analytical Techniques.	135
Table 11-3: Copper - CRM Performance Summary, ALS Laboratory Vancouver, 2011.	139
Table 11-4: Zinc - CRM Performance Summary, ALS Laboratory Vancouver, 2011.	139
Table 11-5: Blank Performance Summary.	140
Table 11-6: Copper - CRM Performance Summary, ALS Laboratory Vancouver.	141
Table 11-7: Zinc - CRM Performance Summary, ALS Laboratory Vancouver.	142
Table 11-8: Gold - CRM Performance Summary, ALS Laboratory Vancouver.	142
Table 11-9: Duplicate Sample Performance Summary, ALS laboratory Vancouver, Diamond Core, 2011.	144

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Table 11-10: Duplicate Sample Performance Summary, ALS laboratory Vancouver, Diamond Core, 2012 - 2013	144
Table 11-11: Duplicate Sample Performance Summary, SGS On-Site laboratory, Bisha Mine RC Grade Control, May to September 2013.	145
Table 11-12: Duplicate Sample Comparison by Laboratory.	148
Table 11-13: Blank Performance Summary.	148
Table 11-14: Copper - CRM Performance Summary, SGS On-Site Laboratory, Bisha Mine.	149
Table 11-15: Copper - CRM Performance Summary, Genalysis Laboratory Perth, 2011 - 2012.	150
Table 11-16: Zinc - CRM Performance Summary, Genalysis Laboratory Perth, 2011 - 2012.	150
Table 11-17: Copper - CRM Performance Summary, Genalysis Laboratory Perth, 2013.	151
Table 11-18: Zinc - CRM Performance Summary, Genalysis Laboratory Perth, 2013.	152
Table 11-19: Duplicate Sample Performance Summary, Hambok 2011 - 2012.	153
Table 11-20: Duplicate Sample Performance Summary, Hambok 2013.	153
Table 12-1: Independent Verification Samples - Mineralized Interval Comparison.	156
Table 13-1: Metallurgical Sample Drill Hole Locations.	160
Table 14-1: Bisha Main Drill Hole Summary Table.	175
Table 14-2: Bulk Density Values Used for this Estimate vs. 2012 Resource Estimate.	184
Table 14-3: Mineralization Domains and High-Grade Capping.	190
Table 14-4: Variogram Parameters, Arsenic in Supergene.	192
Table 14-5: Variogram Parameters, Copper in Supergene.	193
Table 14-6: Variogram Parameters, Gold and Silver in Supergene.	194
Table 14-7: Variogram Parameters, Arsenic in Primary.	194
Table 14-8: Variogram Parameters, Copper in Primary.	195
Table 14-9: Variogram Parameters, Zinc in Primary.	196
Table 14-10: Variogram Parameters, Gold in Primary.	197
Table 14-11: Variogram Parameters, Silver in Primary.	197
Table 14-12: Block Model Definition.	199
Table 14-13: Mineral Resource Commodity Prices.	201
Table 14-14: Northwest Drill Holes and Sample Metres at 18 October 2013.	204
Table 14-15: UTM to Local Grid Transformation - Common Points.	205
Table 14-16: Sulphide Mineralization Interpretation Criteria.	207
Table 14-17: Bulk Density Sample Data Statistics by Estimation Domain.	214
Table 14-18: Assigned Bulk Density Values.	215
Table 14-19: Copper High-Grade Assay Cuts by Estimation Domain.	215
Table 14-20: Zinc High Grade Assay Cuts by Estimation Domain.	216
Table 14-21: Lead High Grade Assay Cuts by Estimation Domain.	216
Table 14-22: Gold High Grade Assay Cuts by Estimation Domain.	216
Table 14-23: Silver High Grade Assay Cuts by Estimation Domain.	217
Table 14-24: Arsenic High Grade Assay Cuts by Estimation Domain.	217
Table 14-25: Iron High Grade Assay Cuts by Estimation Domain.	217
Table 14-26: Domain Groupings.	218
Table 14-27: Copper Variogram Parameters.	219
Table 14-28: Zinc Variogram Parameters.	220
Table 14-29: Gold Variogram Parameters.	221
Table 14-30: Silver Variogram Parameters.	222
Table 14-31: Arsenic Variogram Parameters.	223
Table 14-32: Block Model Definitions.	224
Table 14-33: Mineral Resource Commodity Prices.	226
Table 14-34 Hambok Deposit Drill Hole Summary Table.	228

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Table 14-35: Variogram Parameters for Hambok Estimation.	233
Table 14-36: Block Model Extents, Massive Sulphide-Hosted Mineralization.	234
Table 14-37: Mineral Resource Commodity Prices.	236
Table 14-38: Combined Mineral Resource Estimate - Bisha, Harena, Northwest and Hambok Deposits (Effective Date 31 December 2013)	238
Table 14-39: Bisha Resource Estimate (Effective Date 31 December 2013)	238
Table 14-40: Harena Mineral Resource Estimate (Effective Date 31 December 2013)	239
Table 14-41: Northwest Mineral Resource Estimate (Effective Date 31 December 2013)	239
Table 14-42: Hambok Mineral Resource Estimate (Effective Date 31 December 2013)	240
Table 15-1: Reserves Metal Prices.	243
Table 15-2: Doré Parameters.	243
Table 15-3: Copper Concentrate Recoveries.	243
Table 15-4: Copper Concentrate Shipping and Smelting Terms.	244
Table 15-5: Zinc Concentrate Recoveries.	244
Table 15-6: Zinc Concentrate Shipping and Smelting Terms.	245
Table 15-7: Bisha and Harena Reserves Estimate (Effective Date: 31 December 2013)	247
Table 16-1: Bisha Main Zone Slope Designs.	251
Table 16-2: Harena Slope Designs - Primary Sulphide Mine Phase.	253
Table 16-3: Pit Phase Volumetrics.	257
Table 16-4: Summarized Mine Plan.	259
Table 16-5: Current Mine Production Fleet	265
Table 16-6: Purchases Required for Additions and Replacements.	266
Table 17-1: LOM Process Plant Throughput	269
Table 19-1: ICP Scan on Flotation Concentrates.	273
Table 21-1: Mining Replacement and Additions Cost ($ '000s)	282
Table 21-2: Capital Cost Summary ($ '000s)	283
Table 21-3: Annual Operating Costs ($ '000S)	285

Figures

Figure 4-1: Property Location Map.	26
Figure 4-2: Nevsun Property Ownership Diagram.	28
Figure 4-3: Location of the Bisha Mining Licences and the Mining Agreement Area.	29
Figure 5-1: Property Access and Site Layout	34
Figure 6-1: Sanu 1:50,000 Geological Map of Mogoraib River Exploration Licence with VTEM Anomalies and Prospects	42
Figure 7-1: Arabian-Nubian Shield.	46
Figure 7-2: Neoproterozoic Terranes of Eritrea.	47
Figure 7-3: Geological Map of Eritrea.	48
Figure 7-4: Bisha Property Stratigraphic Section (Barrie, 2004)	49
Figure 7-5: Schematic Cross-Section of Mineralization and Alteration of the Bisha Deposit, and Characterization of the Main Weathering Domains.	53
Figure 7-6: Simple Weathering Profile through Bisha Sulphides.	54
Figure 7-7: Weathering Sequence, Northern Bisha Pit, Beneath NE Hill Cutback.	56
Figure 7-8: Oxide-Soap-Acid-Pyrite Sand Sequence, Northern Bisha Open Pit	57
Figure 7-9: Copper Supergene, Bisha Open Pit (left) and Fault Controlled Copper Supergene, Northern Bisha Open Pit (right)	59
Figure 7-10: Bisha Deposit Alteration Patterns.	61

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Figure 7-11: 000° North-South Shear Zone Truncating Bisha Deposit	62
Figure 7-12: 020° Shear, Bisha Deposit	63
Figure 7-13: 045° Sub-Vertical Shears, Bisha Deposit	64
Figure 7-14: Supergene Copper Distribution Relative to Structures and Underlying Primary Mineralization.	65
Figure 7-15: Oxide Gold Mineralization Distribution Relative to Faulting and Shearing.	66
Figure 7-16: Bisha Deposition Model after McPhie (2013)	67
Figure 7-17: Northwest Deposit Project Area, Bisha Mine, Eritrea.	69
Figure 7-18: Northwest Mineralization and Structures; Plan View, Local Grid N at 045o	70
Figure 7-19: Mineralization Styles, Northwest Deposit	72
Figure 7-20: Rhyolite and Mudstone Distribution, Northwest Deposit, Plan View.	75
Figure 7-21: Schematic Section: Interpreted Volcano-Sedimentary Facies - Architecture for the Stratigraphy at Northwest	76
Figure 7-22: Interpreted Structural and Massive Sulphide Domains, Northwest Deposit - Plan View.	78
Figure 7-23: Northwest Deposit and Associated Alteration - Plan View.	80
Figure 7-24: Northwest Deposit - Massive Sulphide, Geology, and Weathering Surfaces. Cross-Section 20550 mN	81
Figure 7-25: Hambok: Looking South towards Deposit from Gossan Hill	82
Figure 7-26: Hambok Deposit, Drill Coverage.	83
Figure 7-27: Chlorite-Epidote-Altered Pillow Basalt	84
Figure 7-28: Simplified Relationship between Facies (Schematic Section)	86
Figure 7-29: Typical Section of Mineralization Zones.	87
Figure 8-1: VMS Deposit Classifications.	88
Figure 8-2: Cross-Section of Idealized Mineralization (top) and Alteration (bottom) Zonation Patterns in a Footwall Pipe beneath a Typical VHMS Deposit	89
Figure 8-3: Bisha Bimodal Siliciclastic VMS Model Schematic.	91
Figure 9-1: 2010 Gravity Survey Results.	107
Figure 9-2: Map Showing Interpreted Location of Airborne Magnetic Anomalies.	109
Figure 9-3: Northwest Zone 1:2,000 Scale Geological Map and Trenches.	110
Figure 9-4: Trenches at the Aderat Prospect, Mogoraib EL (1:2,000 Scale)	112
Figure 10-1: Drill Hole Location Map.	118
Figure 13-1: Metallurgical Drill Hole Locations with Supergene Wireframe within Bisha Main Pit	161
Figure 13-2: Metallurgical Drill Hole Locations for Bisha Primary Testwork 2013-2014, Longitudinal Section.	162
Figure 14-1: Major Mineralized Zones at the Bisha Main Deposit, Looking Northwest	177
Figure 14-2: Vertical Cross-Section at 1715500N through the Bisha Main Mineralization, Looking North.	179
Figure 14-3: Plan View of Massive Sulphide Zones and Controlling Structures.	180
Figure 14-4: Schematic Cross-Section of Mineralization and Alteration.	181
Figure 14-5: Bisha Massive Sulphide Zones - Diamond Core Recovery vs. Copper (%)	185
Figure 14-6: Bisha Remnant Oxide Zone - Diamond Core Recovery vs. Silver (g/t)	186
Figure 14-7: Copper Distribution for the Primary Sulphide, Main Zone - Histogram.	187
Figure 14-8: Copper Distribution for the Primary Sulphide, Main Zone - Log Probability.	188
Figure 14-9: Parrish (decile) Analysis for the Primary Sulphide, Main Zone.	189
Figure 14-10: Southern Massive Sulphide Zone - Directions of Spatial Continuity, Looking Northwest	192

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Figure 14-11: Constraining Mineral Resource Pit Shell with Massive Sulphide Mineralization and Underground Mineral Resource, Perspective View Looking ENE.	202
Figure 14-12: 3D Cutaway View of Constraining Mineral Resource Pit Shell, with Mineralization and As-Mined Pit; Looking Northeast	203
Figure 14-13: Modelled Sulphide Domains - Sliced at 480 mRL, approx. 80 m Vertical Depth, Plan View.	206
Figure 14-14: Interpreted Dykes - Mafic (blue), Intermediate (green), QFP (magenta), Plan View.	208
Figure 14-15: Interpreted Structural Features and Massive Sulphide Domains, Plan View.	209
Figure 14-16: Mineralization Domains for Copper (left) and Zinc (right), Plan View.	211
Figure 14-17: Mineralization Domains: Oxide and Supergene Gold (left), High-Grade Primary Gold (right), Plan View	212
Figure 14-18: Northwest Massive Sulphide Zones - Diamond Core Recovery vs. Copper Grade.	214
Figure 14-19: Northwest Resource-Constraining Pit Shell with Copper and Zinc Mineralized Zones, Perspective View Looking Northeast (Local Grid)	227
Figure 14-20: Hambok Massive Sulphide Body and Oxide Zone, Perspective View Looking Northeast	229
Figure 14-21: Vertical Section of Hambok Massive Sulphides showing Internal Copper-Zinc Zonation; Section at approx. 1705100N.	231
Figure 14-22: Hambok Massive Sulphide Zone - Diamond Core Recovery vs. Zinc Grade.	232
Figure 14-23: Hambok Massive Sulphide: Slope of Regression Distribution, Value >0.5 (inset, all blocks)	235
Figure 14-24: Perspective View of Hambok Mineralization with Constraining Pit Shell, Looking North.	237
Figure 16-1: Main Zone Domains and Geotechnical Units.	251
Figure 16-2: Harena Domains and Geotechnical Units.	253
Figure 16-3: Bisha Ultimate Pit Design.	255
Figure 16-4: Harena Ultimate Pit Design.	256
Figure 16-5: Bisha Site Layout	261
Figure 16-6: Harena Mine Site Layout	262
Figure 17-1: Supergene Process Flowsheet (crushing, grainding, and tailings circuits common with oxide)	268
Figure 17-2: Oxide Process Flowsheet (crushing, grainding, and tailings circuits common with supergene)	270
Figure 23-1: Chalice Gold Mines Licence Holding and Bisha Properties.	288

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Abbreviations, Acronyms, and Units of Measure

Acid Mine Drainage.................................................................................................................... AMD

Acid Rock Drainage.................................................................................................................... ARD

AGP Mining Consultants Inc...................................................................................................... AGP

Airborne versatile time domain electro-magnetic............................................................... VTEM

ALS Global.................................................................................................................................... ALS

Annum (year).............................................................................................................................. a

Average Coefficient of Variation.............................................................................................. ACV

Base of Complete Oxidation..................................................................................................... BOCO

BGC Engineering Inc................................................................................................................... BGC

Bisha Mining Share Company................................................................................................... BMSC

Bisha Property............................................................................................................................. the Property

Boart Longyear Eritrea Limited................................................................................................ Boart Eritrea

Bulk Leach Extractable Gold..................................................................................................... BLEG

Carbon-in-Leach......................................................................................................................... CIL

Centimetre................................................................................................................................... cm

Certified Reference Materials.................................................................................................. CRMs

Check Samples............................................................................................................................ CS

Coefficient of Variation.............................................................................................................. CV

Conceptual Closure and Reclamation Plan............................................................................. CCRP

Copper.......................................................................................................................................... Cu

Cube Consulting.......................................................................................................................... Cube

Cubic centimetre........................................................................................................................ cm³

Cubic feet per minute................................................................................................................ cfm

Cubic metre................................................................................................................................. m³

Day................................................................................................................................................ d

Days per week............................................................................................................................. d/wk

Days per year (annum).............................................................................................................. d/a

Degree.......................................................................................................................................... °

Diameter...................................................................................................................................... ø

Diamond Drill Hole..................................................................................................................... DDH

Dry metric tonne........................................................................................................................ dmt

Electromagnetic.......................................................................................................................... EM

Elevation (metres) ..................................................................................................................... El

Environmental Impact Assessment.......................................................................................... EIA

Environmental Impact Declaration.......................................................................................... EID

Eritrean National Mining Corporation..................................................................................... ENAMCO

Exploration License ................................................................................................................... EL

Gemcom Whittle V. 5.5.4.......................................................................................................... Whittle

General Exploration Drilling Pty Ltd........................................................................................ GED

Global positioning system.......................................................................................................... GPS

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Global Positioning System......................................................................................................... GPS

Gold............................................................................................................................................... Au

Grade Control Planning............................................................................................................. GCX

Gram............................................................................................................................................. g

Grams per tonne......................................................................................................................... g/t

Greater than................................................................................................................................ >

Hanging Wall Copper Zone....................................................................................................... HWCu

Hectare (10,000 m²)................................................................................................................... ha

Hour.............................................................................................................................................. h

Hours per day.............................................................................................................................. h/d

Hours per week.......................................................................................................................... h/wk

Hours per year............................................................................................................................ h/a

Induced polarization.................................................................................................................. IP

Inductively-coupled plasma (Chemical Analysis Instrument).............................................. ICP

Inverse Distance Squared......................................................................................................... IDW²

Kilogram....................................................................................................................................... kg

Kilogram per year....................................................................................................................... kg/a

Kilograms per cubic metre........................................................................................................ kg/m³

Kilograms per hour..................................................................................................................... kg/h

Kilograms per square metre..................................................................................................... kg/m²

Kilometre..................................................................................................................................... km

Knight Piésold (South Africa).................................................................................................... KP SA

Lead.............................................................................................................................................. Pb

Lerchs–Grossmann..................................................................................................................... LG

Less than...................................................................................................................................... <

Litre............................................................................................................................................... L

Litres per minute........................................................................................................................ L/m

Major Pontil Pty Ltd.................................................................................................................... Major Pontil

Mass spectrometer (Analysis Instrument).............................................................................. MS

Mass submerged in water......................................................................................................... Mw

Measure of acidity or alkalinity of a solution......................................................................... pH

Metre............................................................................................................................................ m

Metres above sea level ............................................................................................................ masl

Metres per minute..................................................................................................................... m/min

Metres per second..................................................................................................................... m/s

Micrometre (micron)................................................................................................................. µm

Milliamperes................................................................................................................................ mA

Milligram...................................................................................................................................... mg

Milligrams per litre..................................................................................................................... mg/L

Millilitre........................................................................................................................................ mL

Millimetre.................................................................................................................................... mm

Million........................................................................................................................................... M

Million ounces............................................................................................................................. Moz

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Million tonnes............................................................................................................................. Mt

Mineralurgy Pty. Ltd................................................................................................................... Mineralurgy

Ministry of Environment............................................................................................................ MoE

Ministry of Mines........................................................................................................................ MoM

Minute (plane angle).................................................................................................................. '

Minute (time).............................................................................................................................. min

Month........................................................................................................................................... mo

MWH Geo-Surveys Inc............................................................................................................... MWH

Net Smelter Return.................................................................................................................... NSR

Nevsun Resources Ltd............................................................................................................... Nevsun or the Company

NI 43-101 Technical Report...................................................................................................... the Report

Non-Acid Forming....................................................................................................................... NAF

Optical Emission Spectroscopy (Analysis Instrument).......................................................... OES

Ordinary Kriging.......................................................................................................................... OK

Ounce........................................................................................................................................... oz

Overall Bias.................................................................................................................................. OABias

Parts per billion........................................................................................................................... ppb

Parts per million.......................................................................................................................... ppm

Percent......................................................................................................................................... %

Potentially Acid Forming........................................................................................................... PAF

Qualified Person......................................................................................................................... QP

Quality Assurance and Control................................................................................................. QA/QC

Quantitative Kriging Neighbourhood Analysis....................................................................... QKNA

Relative Difference Plots........................................................................................................... RDP

Reverse Circulation.................................................................................................................... RC

Reverse Circulation Grade Control.......................................................................................... RCGC

Reverse Circulation Hole........................................................................................................... RCH

Rock Mass Rating........................................................................................................................ RMR

Rock Quality Designator............................................................................................................ RQD

Run-of-Mine................................................................................................................................ ROM

Sanu Resources Ltd.................................................................................................................... Sanu

Second (plane angle)................................................................................................................. "

Second (time).............................................................................................................................. s

Seismic Magnitude..................................................................................................................... Ms

Silver............................................................................................................................................. Ag

Slope of Regression.................................................................................................................... Sr

Socioeconomic and Environmental Impact Assessment...................................................... SEIA

Specific Gravity........................................................................................................................... SG

Square Centimetre..................................................................................................................... cm²

Square kilometre........................................................................................................................ km²

Square metre.............................................................................................................................. m²

Standard deviation..................................................................................................................... SD

Sulphur dioxide........................................................................................................................... SO₂

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Tailings Management Facility.................................................................................................... TMF

Thousand tonnes........................................................................................................................ kt

Three-dimensional..................................................................................................................... 3D

Tonne (1,000 kg)......................................................................................................................... t

Top of Fresh Rock....................................................................................................................... TOFR

Underground............................................................................................................................... UG

Universal Transverse Mercator................................................................................................ UTM

US dollar....................................................................................................................................... US$

Volcanic Massive Sulphide........................................................................................................ VMS

Waste Rock Storage Facilities................................................................................................... WRF

X-Ray diffraction......................................................................................................................... XRD

Year (annum)............................................................................................................................... a

Zinc................................................................................................................................................ Zn

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1 Summary

Nevsun Resources Ltd. (Nevsun) retained AGP Mining Consultants Inc. (AGP) and Mineralurgy Pty. Ltd. (Mineralurgy) to work with staff at Bisha Mining Share Company (BMSC) to update the Mineral Resources and Mineral Reserves for the Bisha Mine (including the Bisha Main, Harena, Northwest, and Hambok deposits), and to prepare an updated Technical Report (the Report) for Bisha and its satellite deposits (the Property) located in Eritrea, Africa.

Nevsun holds a 60% interest in the Property, through a 60% interest in BMSC. The Eritrean National Mining Corporation (ENAMCO) holds the remaining 40% interest. BMSC is the operator for the Bisha and Harena mining licences, the mining agreement area, and the Mogoraib River Exploration Licence (Mogoraib EL).

This Report was prepared in compliance with National Instrument 43–101, Standards of Disclosure for Mineral Projects (NI 43–101), and documents the results of ongoing exploration work and mine production on the Property in support of the Nevsun press release dated 18 February 2014, entitled “Nevsun Increases Indicated Resource Tonnes by 29% and Contained Copper by 22%.”

1.1 Location and Access

The operation is located 150 km west of Asmara, 43 km southwest of the regional town of Akurdat, and 50 km north of Barentu, the regional, or Zone Administration Centre, of the Gash-Barka District, in Eritrea, East Africa.

Access to the Property is by paved road from Asmara to Akurdat, a distance by road of 181 km and then 52 km from Akurdat via an all-weather unpaved road, which is currently being upgraded. The drive from Asmara to the Bisha camp (also referred to as Bisha Village) takes approximately 4 hours.

1.2 Mineral Tenure, Surface Rights, and Royalties

The Property comprises two mining licences covering an area of 24.0 km², (16.5 km² for Bisha Main and the Northwest Zones and 7.5 km² for Harena), a Mining Agreement Area covering an area of 39.0 km²for Bisha, and the 73.1 km² Mogaraib River Exploration Licence (Mogaraib EL). BMSC is the operator for all of the licences.

Under the terms of the mining agreement, BMSC has the exclusive right of land use in the Mining Licence Area that is granted within the mining agreement area. This right is subject to the acquisition and settlement of any third-party land-use rights by payment of compensation and/or relocation at the expense of BMSC.

Royalties payable include an Eritrean Government gross royalty of 5.0% on precious metals and 3.5% on base metals.

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1.3 Permits

BMSC holds all the necessary permits to support a mining operation.

1.4 Environment

A Socioeconomic and Environmental Impact Assessment (SEIA) is conducted as part of the mine licencing process. An SEIA report was completed in December 2006 and submitted to the Ministry. A Mining Licence for the Operation was issued on 26 May 2008. Issuance of the licence is accepted by BMSC as SEIA approval by the Impact Review Committee, as required under the Eritrean Mining Regulations. The Harena SEIA Addendum was submitted December 2011 with a conditional mining licence granted in July 2012. All Ministry of Mines (MoM) and Ministry of Environment (MoE) SEIA requirements have been satisfied with respect to the Harena Mining Licence.

1.5 Geological Setting and Mineralization

The BMSC deposits are located within the Arabian-Nubian Shield. The Precambrian rocks are exposed on either side of the Red Sea in western Arabia and northeastern Africa. The rocks are a collage of volcanic arcs, granitoid intrusions, volcano-sedimentary basins, and shear zones. The Arabian-Nubian Shield has a wide range of deposit types and settings, including volcanic massive sulphide (VMS) deposits. Eritrea is divided into several north- or northeast-trending Proterozoic terranes, which are separated by major crustal sutures. The western Nakfa Terrane comprises low-grade metamorphosed calc-alkaline volcanics and sediments, and hosts base metal mineralization in the Gash-Barka district, including the Bisha polymetallic deposits.

The Bisha deposit is a series of massive sulphide lenses that trend north-south and dip steeply to the west. The strike extent is 1.2 km and the thickness of the lenses varies from 0 m to 70 m. Mineralization extends to a depth of 500 m below surface and has not been closed off. The deposit is hosted by an intensely foliated, bimodal sequence of predominantly tuffaceous metavolcanic rocks. Felsic lithologies appear to directly host the mineralization, predominate overall, and form the hanging wall stratigraphy. A significant component of mafic metavolcanic rocks occurs in the more obviously bimodal footwall, which is exposed mainly to the east of known mineralized zones.

Deep weathering has affected the Bisha Main Zone lenses by removing most of the sulphide and producing high-grade supergene blankets enriched in gold and copper. The depth of oxidation is approximately 30 m to 35 m, with supergene sulphides developed to approximately 65 m in depth. The oxidation of massive sulphides generated strong acid solutions that have progressively destroyed the sulphides and host rock, forming a horizon of extremely acid-leached material between the oxide and supergene / primary domains. To the west of the massive sulphide lenses is the Hanging Wall Copper Zone, which is restricted to the supergene horizon and has a zinc massive sulphide primary component. The Hanging Wall Copper Zone has a north–northeast strike, and converges towards the Bisha Main Zone towards the north. The primary massive sulphides extending to depth are zinc dominated with a series of internal high-grade lenses. A zone of stringer sulphides with low-grade copper is present in the footwall of the primary massive sulphides. The oxide zone is largely mined-out.

The Harena deposit is a northwest-dipping, tabular, massive sulphide body with a strike length of approximately 400 m. The deposit remains open to the southeast. Host rocks to the Harena deposit are a bimodal, hydrothermally-altered suite of basalts and rhyolite-dacite volcanics. There are three distinct zones of mineralization, namely oxide, supergene and primary, varying in thickness from 5 m to 15 m. The supergene is poorly developed as compared with the Bisha deposit, and the deposit lacks stringer mineralization. The primary sulphides are a mixture of zinc and copper mineralization. The oxide zone, developed to approximately 45 m in depth, is largely mined-out.

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The Northwest deposit is made up of an oxide and supergene zone and a series of primary massive sulphide lenses hosted in felsic volcanics, with rhyolites developed in the hangingwall. The deposit strikes at approximately 045°, dipping steeply to the northwest, with overall strike and dip lengths of some 800 m and 230 m, respectively. The deposit is open at depth and along strike. The gold bearing oxide zone is developed over the strike length of the primary massive sulphides in two or more sub-horizontal lenses some 10 m to 15 m thick. The oxide and supergene zones are similar in nature to the Bisha deposit with oxidation developed to approximately 40 m in depth. The supergene zone is best developed over the southern massive sulphide lenses. The primary massive sulphide zones are disrupted, and displaced by north-south trending faults into three zones, south, central, and north. Most of the Resource is captured in the copper-dominated Central Zone where copper forms a carapace around pyrite in the thicker zones. The thickness of the massive sulphide varies from some 5 m to 75 m. The massive sulphides have associated sulphide stringer mineralization in both footwall and hangingwall, which contain lenses of zinc mineralization. A narrow zone of massive sulphides lying to the east runs parallel and has similar strike length, but lacks associated stringer sulphide development. Zones of gold mineralization are recognized in the primary zone that feed into the oxide/supergene gold blanket.

The Hambok deposit comprises a primary copper/zinc sulphide zone, representing the majority of the deposit, and a minor oxide gold component. The host rocks are rhyolitic and basaltic units. The primary massive sulphide mineralization is a single body, with a faulted displacement interpreted at depth in the northeast of the deposit. The massive sulphide zone strikes at approximately 015°, dipping steeply to the east, with overall strike and dip lengths of some 975 m and 400 m, respectively. The thickness of the massive sulphide varies from about 5 m to 75 m. Semi-massive and stringer sulphides are not well developed in this deposit. The mixed copper and zinc mineralization is best developed in the upper parts of the massive sulphide, with the down dip parts dominated by pyrite. The oxide unit, as currently understood, is narrow and poorly developed, with minor gold enhancement. Oxidation is developed down to approximately 50 m.

Significant advances have been made in the understanding of the geology and mineralization of the deposits since the last Report, particularly that the massive sulphide units are interpreted to have been formed as sub-seafloor replacement rather than seafloor deposition.

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1.6 Drilling

Drilling on the four deposits has been undertaken over a number of diamond-core and reverse-circulation campaigns from 2002 to 2013. Drill programs have been completed, primarily by contract drill crews, largely supervised by BMSC geological staff for Bisha, Harena, and Northwest, and by Sanu Resources Ltd. (Sanu) and BMSC staff in the case of Hambok. Drilling undertaken is summarized in Table 1‑1.

Table 1-1: Summary of Drilling for Each of the Deposits

Deposit	No. of DDH	Length of DDH (m)	No. of RCH	Length of RCH (m)
Bisha	681	100,188	514	8,643
Harena	98	11,891	7	646
Northwest	240	46,200	47	873
Hambok	102	26,790	42	2,676

All hole collars are surveyed; Hambok collars were re-surveyed by BMSC. Down-the-hole survey was mostly completed using Eastman single‑shot or Reflex EZ-Shot camera methods; gyroscopic methods were employed in 2013.

BMSC or contract geologists logged core and chip samples. Core is also geotechnically logged and photographed. Core samples vary from approximately 1 m to 3 m in length; RC samples are between 1 m and 2 m in length. Since 2012, all holes are sampled in their entire length.

Core drilling for metallurgical testwork samples has taken place at Bisha, Harena, and Northwest, with drilling of 40, 2, and 2 holes, respectively. Core drilling for geotechnical evaluation has been completed at Bisha, Harena, and Northwest, being drilling of 32, 4, and 8 holes, respectively. Open holes for hydrogeological assessment or water management has been completed at Bisha, Harena, and Northwest, being 29, 7, and 4 holes, respectively.

All analytical, lithological and associated data are stored in an industry standard acQuire database system maintained by BMSC.

1.7 Sample Preparation, Analyses, and Security

Sample analysis has been completed by independent ISO-registered and accredited laboratories for each of the deposits. Laboratories include ALS Vancouver (ALS), Genalysis Laboratory Services PTY Ltd. (Intertek Genalysis), and the African Assay Laboratories (Tanzania) Ltd. (SGS) operated run-of-mine (ROM) laboratory.

The sample collection and preparation, analytical techniques, security and quality assurance / quality control (QA/QC) protocols implemented for the Bisha, Harena, Northwest, and Hambok deposits are consistent with standard industry practice, and are suitable for the purpose of Mineral Resource estimation and the reporting of exploration results. The sampling procedures are adequate for and consistent with the style of base metal and gold mineralization found at these deposits.

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1.8 Data Verification

A series of data verification exercises have been undertaken during the project history for each deposit. Additional work has been completed since the last report, including complete validation of the Bisha and Hambok databases, and an independent sampling exercise for Northwest. The data used in Mineral Resource estimation is considered adequate and reliable.

1.9 Mineral Resource Estimate

A major update of geological interpretation has been made since the last Report for the Bisha deposit, reviewing new and old data. This resulted in more detail in the supergene and primary zones, introduction of structural control in the distribution of mineralization, and more cognizance given to mineral processing characteristics. The Mineral Resource estimates for the Northwest and Hambok deposits are presented for the first time.

Mineral resources were estimated, classified, and reported in accordance with the current CIM Definition Standards for Mineral Resources and Mineral Reserves. Reasonable prospects of eventual economic extraction for each deposit were assessed using a Lerchs–Grossmann (LG) optimized pit shell, using parameters defined for estimation of Mineral Reserves, but with commodity prices some 15% higher. Inferred Resources below the Bisha optimized pit shell that represent underground potential were also estimated.

Mineral Resources reported here for Bisha Main and Harena are inclusive of Mineral Reserves. BMSC cautions that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability. All Mineral Resources have an effective date of 31 December 2013, and are shown in Table 1‑2 to Table 1‑6.

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Table 1-2: Combined Mineral Resource Estimate – Bisha, Harena, Northwest, and Hambok Deposits (Effective Date 31 December 2013)

Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Contained Metal
Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000 lb)	Zn ('000 lb)	Au ('000 oz)	Ag ('000 oz)
*Indicated*
Oxide Phase	Variable	480	-	-	6.6	20	-	-	100	300
Supergene Phase	Variable	8,480	3.41	-	0.6	25	638,650	-	160	6,920
Primary Phase	Variable	32,260	1.05	4.59	0.6	36	743,920	3,223,350	570	36,430
Total		41,220					1,382,570	3,223,350	830	43,650
*Inferred*
Oxide Phase	Variable	570	-	-	3.9	19	-	-	61	350
Supergene Phase	Variable	110	1.37	-	3.4	18	4,200	-	10	70
Primary Phase	Variable	1,752	0.80	4.19	0.7	33	31,230	163,210	40	1,910
Total	-	2,432					35,430	163,210	111	2,330

Table 1-3: Bisha Resource Estimate (Effective Date 31 December 2013)

Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Contained Metal
Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000 lb)	Zn ('000 lb)	Au ('000 oz)	Ag ('000 oz)
*Indicated*
Oxide Phase	40.55	410	-	-	6.8	21	-	-	90	270
Supergene Phase	39.55	7,460	3.68	-	0.6	27	605,500	-	150	6,590
Primary Phase	39.55	21,070	1.05	5.87	0.7	47	487,770	2,726,870	480	31,770
Total	-	28,940					1,093,270	2,726,870	720	38,630
*Inferred*
Oxide Phase	40.55	30	-	-	7.3	39	-	-	10	30
Supergene Phase	39.55	10	7.23	-	0.1	10	2,200	-	0	0
Primary Phase	39.55	1,300	0.80	4.50	0.5	36	23,100	129,600	20	1,500
Total	-	1,340	-	-	-	-	25,300	129,600	30	1,530

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Table 1-4: Harena Mineral Resource Estimate (Effective Date 31 December 2013)

Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Contained Metal
Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000 lb)	Zn ('000 lb)	Au ('000 oz)	Ag ('000 oz)
*Indicated*
Oxide Phase	42.41	70	-	-	5.5	14	-	-	10	30
Primary Phase	41.41	1,800	0.65	3.91	0.6	23	25,760	154,990	30	1,350
Total	-	1,870					25,760	154,990	40	1,380
*Inferred*
Oxide Phase	40.55	20	-	-	5.9	8	-	-	0	10
Primary Phase	39.55	350	0.75	4.10	0.8	32	5,700	31,200	10	350
Total	-	370					5,700	31,200	10	360

Table 1-5: Northwest Mineral Resource Estimate (Effective Date 31 December 2013)

Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Contained Metal
Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000 lb)	Zn ('000 lb)	Au ('000 oz)	Ag ('000 oz)
*Indicated*
Oxide Phase	40.70
Supergene Phase	39.70	1,020	1.47	-	0.2	10	33,150	-	10	330
Primary Phase	39.70	2,530	1.04	1.08	0.3	13	58,020	60,250	20	1,050
Total	-	3,550					91,170	60,250	30	1,380
*Inferred*
Oxide Phase	40.70	500	-	-	3.7	18	-	-	50	300
Supergene Phase	39.70	100	0.80	-	3.7	19	2,000	-	10	70
Primary Phase	39.70	100	0.90	0.90	2.9	15	2,400	2,400	10	60
Total	-	700					4,400	2,400	70	430

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Table 1-6: Hambok Mineral Resource Estimate (Effective Date 31 December 2013)

Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Contained Metal
Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000 lb)	Zn ('000 lb)	Au ('000 oz)	Ag ('000 oz)
*Indicated*
Oxide Phase	44.45	-	-	-	-	-	-	-	-	-
Primary Phase	43.45	6,860	1.14	1.86	0.2	10	172,370	281,240	40	2,260
Total	-	6,860					172,370	281,240	40	2,260
*Inferred*
Oxide Phase	44.45	20	-	-	1.5	17	-	-	1	10
Primary Phase	43.45	2	0.90	0.20	0.2	8	30	10	0	0
Total	-	22					30	10	1	10

The following notes should be read in conjunction with Table 1-2 to Table 1-6, above:

Notes:

(2) NSR values were calculated for each block using both Indicated and Inferred categories, metal prices, recoveries, appropriate smelter terms, and downstream costs. Metallurgical recoveries, supported by metallurgical testwork, were applied as follows:

a) Bisha Oxide Zone: recoveries of 88% and 22% were applied for gold and silver, respectively.

b) Harena Oxide Zone: a recovery of 75% was applied for gold.

c) Bisha Supergene Zone: recoveries of 88%, 46%, and 50% were applied for copper, gold, and silver, respectively.

d) Bisha Hanging Wall Zone: recoveries of 85%, 46%, and 50% were applied for copper, gold, and silver, respectively.

e) Bisha Transition Zone (mixed zinc and secondary copper zone below the supergene): recoveries as per supergene zone were applied.

f) Bisha Primary Zone: recoveries to copper concentrate of 85%, 36%, and 29% were applied for copper, gold, and silver, respectively. Recoveries to zinc concentrate of 83.5%, 9%, and 20% were applied for zinc, gold, and silver, respectively.

g) Harena Primary Zone: recoveries to copper concentrate of 85%, 36%, and 29% were applied for copper, gold, and silver, respectively. A zinc recovery to zinc concentrate of 72% was applied.

h) Northwest Oxide Zone: recoveries of 88% and 22% were applied to gold and silver, respectively.

i) Northwest supergene zone: recoveries of 87%, 46%, and 50% were applied for copper, gold, and silver, respectively. Zinc was not assigned a recovery, as the values are isolated on the fringes of the deposit.

j) Northwest Primary Zone: recoveries to copper concentrate of 87%, 36%, and 29% were applied for copper, gold, and silver, respectively. Recoveries to zinc concentrate of 81%, 9%, and 20% were applied for zinc, gold, and silver, respectively.

k) Hambok Oxide Zone: recoveries of 88% and 22% were applied to gold and silver, respectively.

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l) Hambok: recoveries to copper concentrate of 88%, 87%, 36%, and 29% were applied for copper, zinc, gold, and silver, respectively. Preliminary metallurgical characterization studies, but not full testing, have been completed for Hambok.

(3) Mineral Resources are reported within the pit shell generated using the specified commodity prices, using NSR block grade cutoffs derived as above. Tonnages were rounded to the nearest 10,000 tonnes, and grades were rounded to two decimal places for copper and zinc, one decimal place for gold, and no decimal places for silver. Tonnages and grades for the Inferred category are further rounded, reflecting the uncertainty that attaches to this category.

(4) Rounding as required by reporting guidelines may result in apparent summation differences between tonnes, grades, and contained metal contents.

(5) Tonnage and grade measurements are in metric units. Contained gold and silver ounces are reported as troy ounces, contained copper and zinc pounds as imperial pounds.

(6) Stockpile tonnages are included as Indicated Resources in the totals given in the tables for Bisha and Harena.

(7) The Bisha Primary Inferred Resource includes an underground resource. This was derived by defining a shape around contiguous blocks where an NSR of $100 was achieved. The value of NSR $100 represents the processing cost plus approximately $60/t mining cost.

(8) Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

1.10 Mineral Processing and Metallurgical Testing

In 2005, metallurgical testwork was done at SGS Lakefield in Canada for feasibility studies for the oxide, supergene, and primary mineralization material types. The results of this work were used for the design of the current 2 Mt/a design capacity process plant at Bisha to treating oxide ore by cyanide leaching producing doré bullion. This plant achieved commercial production in February 2011 and operated until Q3 2013. SGS Lakefield’s work on the Bisha supergene copper and primary copper-zinc mineralization included quantitative mineralogical examination, together with bench- scale testwork on comminution and flotation. A key focus area in processing the Bisha supergene copper mineralization due to the high pyrite:copper sulphide ratio in the feed, is ensuring acceptable liberation of sulphide minerals. Flotation tests showed that the copper concentrate with a grade around 30% Cu could be made at a recovery of 88%. Mineral liberation of the primary copper-zinc mineralization was very high for a volcanogenic massive sulphide deposit at a modest grind of P₈₀ of 53 µm with the expectation of making relatively clean copper and zinc concentrates at high metals recoveries; this was supported by the flotation testwork. Based upon the samples tested, this work demonstrated that conventional grinding and flotation technologies could be used to treat the supergene copper and primary copper-zinc mineralization to produce copper concentrate at 25% Cu, copper and zinc concentrate at 55% Zn at high metals recoveries.

The prime aim of work done from 2010 to mid-2012 by three metallurgical testing laboratories in South Africa—Maelgwyn Mineral Services Africa, Mintek, and SGS South Africa—has been both to confirm the results of the 2005 SGS Lakefield work on Bisha supergene copper mineralization and produce process design data. The latter were required by the engineering company SENET of South Africa, the EPCM contractor for the flotation section and ancillary facilities at Bisha for the treatment of supergene copper mineralization. Copper concentrate production commenced in mid-2013. Unforeseen issues have not arisen from this work, with the main “outcome driver” of metallurgical performance being the separation of copper sulphide minerals from the pyrite. Additional quantitative mineralogy work has been done to support the metallurgical testwork.

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Additional comminution, cyanidation, and flotation work has been done in 2012 on samples from the Harena deposit, by SGS South Africa and Maelgwyn Mineral Services Africa. Maelgwyn Mineral Services Africa demonstrated that Harena primary copper-zinc mineralization could be treated by flotation to produce saleable copper and zinc concentrates though with somewhat lower zinc concentrate, grade and zinc recovery compared with Bisha primary copper-zinc mineralization.

A limited number of observations during Maelgwyn’s testwork on Bisha supergene mineralization indicated the presence of minor amounts of enargite/tennantite and arsenopyrite, which could result in higher arsenic levels in the concentrate than previously observed. While arsenopyrite can be expected to follow pyrite and not report to the supergene copper concentrate, under normal conditions for the recovery of copper sulphide minerals, enargite/tennantite will report to the copper concentrate. Since insufficient information was available to quantify the spatial limits of the enargite/tennantite at the time of report writing, the conservative approach of assumed homogeneity throughout the supergene zone was taken. Assuming 75% of the arsenic is contained in enargite/tennantite with the remainder in arsenopyrite, an arsenic recovery to copper concentrate of 67.5% was estimated. This trend has subsequently been confirmed by examination of monthly performance data. Further testwork is ongoing to better quantify and delineate the enargite/tennantite and arsenopyrite within the supergene zone. Blending should allow the supergene copper concentrate to be kept below the 0.5% As limit generally required for copper concentrate sold to most custom smelters.

From late 2012 through to mid-2013, ALS Metallurgy Kamloops has done testwork both on “transitional supergene” and pyrite sand from Bisha, and samples from the Northwest satellite deposit. The former indicated beneficiation by flotation was possible, and the Northwest material showed similar mineralization and metallurgical performance to analogous Bisha Main mineralization.

A program of metallurgical testwork is currently underway on Bisha primary mineralization at ALS Metallurgy Kamloops to both confirm the results of the SGS Lakefield investigations on a wider suite of samples, and to generate data for the design of the plant to treat this material.

It is the qualified persons (QPs)’ opinion that the metallurgical testwork completed to date on the Bisha Main operation has been appropriate to establish reasonable processing routes for the different mineralization styles in the various deposits, to a level sufficient to support Mineral Reserves declaration:

Sample testwork has been based on mineralization that is typical of the various mineralization types currently interpreted in the deposits.

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Recoveries used in the current mineral resource and mineral reserve estimations are consistent with metallurgical testwork on the various mineralization types. Realistic values were used to estimate metal recoveries where there are limited testwork data.
The production performance achieved to date on the plant modifications currently under construction for the treatment of Bisha Main supergene mineralization and the associated metals recovery factors are considered appropriate to support mineral resource and mineral reserve estimation.
Locked cycle flotation testwork has been done to support the design of the primary mineralization supergene flotation circuit and estimation of associated metals recovery factors.
Additional batch and locked cycle flotation testwork is recommended to support the future design of the primary ore flotation circuit and support associated metals recovery factors.
It is recommended that further investigation be done into the effects of oxidation on supergene and primary flotation performance recognizing that while this issue is common to other volcanogenic massive sulphide deposits, the very high pyrite content of Bisha material may make it more susceptible to this phenomenon.

1.11 Mineral Reserves Estimates

The Proven and Probable Mineral Reserves at the operation have been classified in accordance with the 2010 CIM Definition Standards for Mineral Resources and Mineral Reserves. Mineral Reserves are defined within a mine plan, with open pit phase designs guided by LG optimized pit shells, generated using metal prices for copper, zinc, gold, and silver of $2.90/lb, $0.92/lb, $1,175/oz, and $20/oz, respectively. The NSR cutoffs ($/t) are supergene phase $39.55 for Bisha; and primary phase $39.55 for Bisha and $42.41 for Harena; remaining oxide $40.55 for Bisha and $42.41 for Harena. The summary of the Mineral Reserves are shown in Table 1‑7.

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Table 1-7: Bisha and Harena Reserves Estimate (Effective Date: 31 December 2013)

Zone	Tonnes (‘000s)					Contained Metal
Zone	Tonnes (‘000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000s lb)	Zn ('000s lb)	Au ('000s oz)	Ag ('000s oz)
*Bisha Probable Mineral Reserve Estimate*
Oxide Phase	430	-	-	6.50	20	-	-	90	280
Supergene Phase	7,400	3.57	-	0.61	27	582,420	-	145	6,420
Primary Phase	18,390	1.02	5.66	0.68	46	413,540	2,294,730	402	27,200
Total	26,220					995,960	2,294,730	637	33,900
*Harena Probable Mineral Reserve Estimate*
Oxide Phase	80	-	-	4.93	16	-	-	13	-
Primary Phase	1,160	0.64	3.57	0.52	22	16,370	91,300	19	820
Total	1,240					16,370	91,300	32	820
*Combined Bisha and Harena Probable Mineral Reserve Estimate*
Oxide Phase	510	-	-	6.25	19	-	-	103	280
Supergene Phase	7,400	3.57	-	0.61	27	582,420	-	145	6,420
Primary Phase	19,550	1.00	5.54	0.67	45	429,910	2,386,030	421	28,020
Total	27,460					1,012,330	2,386,030	669	34,720

Notes:

(1) NSR Cutoff ($/t): Oxide Phase $40.55 for Bisha and $42.41 for Harena: Supergene Phase $39.55 for Bisha and Primary Phase $39.55 for Bisha and $42.41 for Harena. Mineral Reserves are defined within a mine plan, with phase designs guided by LG pit shells, generated using metal prices for copper, zinc, gold and silver of $2.90/lb, $0.92/lb, $1,175/oz, and $20/oz respectively. The mining cost applied was the 2014 budget mining cost with appropriate haulage cost adjustments. The total ore based costs (process, G&A, and stockpile re-handle) are $40.55/t for oxide, and $39.55/t for supergene and primary ores. Harena ore-based costs include an additional $2.63/t overland ore haulage cost. Overall pit slopes varied from 34.5° to 44° for Bisha and from 29° to 35.5° for Harena.

(2) NSR values were calculated using diluted Indicated grades, metal prices, recoveries and appropriate smelter terms, and downstream costs. Metallurgical recoveries, supported by metallurgical testwork, were applied as follows:

a) Bisha Oxide Zone: recoveries of 88% and 22% were applied for gold and silver, respectively.

b) Harena Oxide Zone: a recovery of 75% was applied for gold; 80 kt of oxide remain in the Harena pit.

c) Bisha Supergene Zone; recoveries of 88%, 46%, and 50% were applied for copper, gold, and silver respectively.

d) Bisha Hanging Wall Zone: recoveries of 85%, 46% and 50% were applied for copper, gold and silver, respectively.

e) Bisha Transition Zone: (mixed zinc and secondary copper zone below the supergene): the same metallurgical parameters as for the Bisha Supergene Zone were applied.

f) Bisha Primary Zone: recoveries to copper concentrate of 85%, 36%, and 29% were applied for copper, gold, and silver, respectively. A recovery to zinc concentrate of 83.5% was applied for zinc. Gold and silver reporting to the zinc concentrate are not expected to be payable.

g) Harena Primary Zone: recoveries to copper concentrate of 85%, 36% and 29% were applied for copper, gold and silver respectively. A zinc recovery to zinc concentrate of 72% was applied. Gold and Silver reporting to zinc concentrate are not expected to be payable.

(3) Mineral Reserves are reported within Bisha and Harena ultimate pit designs, using NSR block grade, where the marginal cutoff is the total ore-based cost stated above. Tonnages are rounded to the nearest 10kt and grades are rounded to two decimal places with the exception of silver, which was rounded to zero decimal places.

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(4) Rounding as required by reporting guidelines may result in apparent summation differences between tonnes, grade, and contained metal content.

(5) Tonnage and grade measurements are in metric units. Contained gold and silver ounces are reported as troy ounces, contained copper and zinc pounds as imperial pounds.

(6) The life-of-mine (LOM) stripping ratios in tonnes for Bisha and Harena are 5.6: 1 and 5:7: 1, respectively.

(7) The Bisha Probable Oxide Mineral Reserve includes the pyrite sand and uncrushed DSO stockpiled material, being 320 kt at 7.8 g/t Au in the stockpile as of 31 December 2013.

(8) The Bisha Probable Supergene Mineral Reserves includes 105 kt at 5.25% Cu in the stockpile as of 31 December 2013.

(9) For Bisha and Harena, conversions from Mineral Resources to Mineral Reserves included a mining dilution of 2 m applied around contiguous blocks that exceeded the appropriate NSR cutoff value.

1.12 Mine Plan

The Bisha Main and Harena deposits are being mined by conventional open pit mining methods. The Bisha Main pit consists of nine individual pit phases, where the first three phases target oxide production, the second three target Supergene production and the final three phases will target Primary production. The oxide pit phases have now been exhausted, and Phases 4, 5, and 6 are currently providing supergene ore. Stripping for the primary mineralization has started in Phase 8.

The Harena pit features two pit phases, one targeting oxide production (which was completed in 2013), and the final phase targeting primary production. Some remaining oxide mineralization will be mined within the Phase 2 pit and will be processed at the end of the mine life. The Harena pit is currently inactive.

Although the initial oxide production phase of the operation is complete, small quantities of oxide mineralization remain within some of the pit phases and in stockpiles. These materials will be processed at the end of the primary phase of the operation.

Drilling and blasting is performed on 5 m benches, with loading carried out on 2.5 m flitches to minimize dilution and mining loss. Primary loading is performed with RH40s, loading Cat 775 trucks. The mine is scheduled to work 360 d/a, with five days allowed for delays due to weather disruptions. The plant is scheduled to operate 365 d/a.

Over the remaining LOM, it is expected that 150 Mt of waste rock will be produced from the Bisha Main pit and placed in two wasterock storage facilities (WRFs) to the east and southeast of the open pit, plus a backfill dump in the north end of the ultimate pit. It is also expected that 8 Mt of waste rock will be produced from the Harena Pit and placed in two WRFs to the east and southeast of the open pit.

The mine grade control process currently features vertical and angled RC holes piercing multiple benches on a 10 m x 10 m pattern. Two owner-operated Schramm T450 drill rigs are used to drill up to 2 benches (10 m) ahead of mining. Each drill rig has a side-mounted sampling system consisting of a cyclone and underlying sample splitter. Grade control collects a primary 2 kg to 3 kg sample, and bags an up to 20 kg residual sample on a metre-drilled basis. Grade Control samples are currently assayed by the SGS on-site laboratory for Ag, Cu, Zn, As, and Pb using a 3-acid digestion with atomic absorption finish, with Au determined by fire assay. Assay results are input into the geological database, then ordinary kriged. The blocks are used to design ore polygons, which are mapped out by surveyors on the shot muck.

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The processing of oxide ore ended in mid-2013. Commissioning of the supergene plant started in Q3 2013, and commercial production announced 2 December 2013. Supergene material will be processed alone until mid-2016, when the zinc flotation plant comes online to begin processing primary phase material. Due to the sub-horizontal and undulating contact between the supergene and primary mineralized materials, there will be a multi-year period where both supergene and primary materials are mined. During this overlap period, both supergene and primary ores will be treated in campaigns of appropriate durations, to avoid mixing the ore types. However, no interruption to production is anticipated as campaign processing of different ore types is an established practice in the treatment of base metals sulphide ores.

The mine delivers ore to the ROM pad, where it is dumped into several different short-term stockpiles. This ore is then reclaimed by a front-end loader to the dump pocket, following a blending plan that is provided daily and modified as required based on plant performance. Longer-term stockpiling of non-oxide material has been minimized to limit oxidization of material.

The current production fleet will require additions as the stripping increases to expose Primary ores. At peak production, which occurs in 2018, the equipment requirements are four excavators, one front-end loader, four drills, and twenty-five haul trucks.

1.13 Process

The oxide processing facility achieved commercial production in February 2011, and has operated at an average throughput rate of 1.8 Mt/a from start-up until July 2013.

The oxide plant facilities include a primary crusher, SAG and ball grinding mills, cyanide leach/carbon-in-leach (CIL) circuit, cyanide destruction circuit, refinery to produce doré bullion, tailings thickener, tailings discharge system, and the necessary reagent, water, and air systems.

Bisha has three different types of mineralization: oxide, supergene, and primary; each requiring a specific process flowsheet. Downstream of the existing crushing and grinding circuit, the additional process equipment to treat the supergene mineralization is now installed and was commissioned in mid-2013. This copper plant expansion consists of flotation cells for copper roughing and cleaning duties, regrind mills for rougher concentrate, copper concentrate thickener, and pressure filters, copper concentrate load-out building, copper flotation reagent systems, flotation air blowers, and pressure-filter air compressors.

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Some transitional material remains that is a mixture of oxide and supergene mineralization mined as the oxide was depleted. An initial 100,000 tonnes of this material was treated in the new copper flotation circuit with the flotation tails leached in the CIL circuit, with the intent of maximizing copper recovery and minimizing gold losses. Additional transitional material will be mined and stockpiled for future processing or direct sale. The cyanide leach circuit was not operated after this amount of transitional ore was treated.

The plant is currently ramping up to the design performance of making a 30% Cu concentrate grade at 88% Cu recovery.

There have been the expected commissioning problems in changing over from a hydrometallurgical process of CIL treatment of oxide gold ore, to physical beneficiation of base metal sulphides by flotation making a concentrate, with equipment issues, familiarizing the workforce with new unit operations, and other concerns, exacerbated by logistical issues.

In December 2013 a total of 135,083 tonnes of supergene ore was processed at head grade of 6.3% Cu, 1.21 g/t Au, and 31.4 g/t Ag, producing 21,880 tonnes of copper concentrate assaying 32.8% Cu, 4.19 g/t Au, and 131 g/t Ag at 84.4% Cu recovery, 56.1% Cu recovery, and 67.8% Ag recovery.

Similarly, before the supergene mineralization is exhausted, the additional equipment required to process the primary mineralization will be installed and commissioned, to permit a smooth transition to processing primary mineralization with minimum interruption and shutdowns.

The crushing and grinding “front-end” of the current plant, which processed Bisha oxide ore for the cyanide leaching, is currently being used for the supergene ore and will be used for the primary ore. Additional equipment currently being installed for treating the supergene mineralization includes flotation cells for copper roughing and cleaning duties, regrind mills for rougher concentrate, copper concentrate thickener and pressure filters, copper concentrate load-out building, copper flotation reagent systems, flotation air blowers, and pressure-filter air compressors.

For the treatment of primary mineralization, additional equipment will include zinc roughing and cleaning flotation circuits, zinc concentrate regrind mill, zinc concentrate thickener and pressure filters, zinc concentrate load-out building, zinc flotation reagent systems, additional zinc flotation air blower and zinc pressure filter air compressor.

1.14 Markets

BMSC has negotiated contracts with two smelters for the sale of approximately 60% of its future copper concentrate. Normal commercial terms are included in the concentrate contracts. Negotiations are underway for the sale of the remaining copper concentrates.

Discussion for the sale of future zinc concentrates to be produced from the treatment of primary material will commence in 2015. Terms contained within the concentrate sales contracts are likely to be typical of, and consistent with, standard industry norms, and be similar to such contracts elsewhere in the world.

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1.15 Capital and Operating Costs

For the Bisha mine, the majority of the capital cost has already been spent, with the oxide phase complete and the copper phase reaching commercial production in December 2013, and currently ramping up to planned throughput.

SENET of South Africa built the copper phase project under EPCM contract. SENET was the same contractor who successfully built the original Bisha plant. The total estimated cost of this phase of the project was $110 million, which was completed on time and under budget.

The zinc phase project is currently estimated at $89 million. Mining capital requirements are $39 million during the mine life, for mobile fleet replacements and additions. The remaining TMF expansions have been estimated at $40 million. Sustaining capital related to process replacements/additions, camp, port, and general/administrative replacements/additions have been estimated at $31 million.

The operating costs are based on BMSC’s 2014 budget. The mining cost (inclusive of operations, maintenance, grade control, and geology and laboratory services) was $2.73/t, and an appropriate incremental haulage cost per bench of $0.01/t/5 m. The LOM average cost with incremental haul additions is $2.87/t mined.

The process cost is based on BMSC’s 2014 budget process cost estimate of $28.74/t andincludes maintenance, power, labour, and consumables.

G&A costs are based on the 2014 Budget estimate $10.04/t milled.

1.16 Conclusions

The Mineral Resources and Mineral Reserves have been successfully updated for the property. The QPs believe that there are no issues with regard to the technical information that would materially impact on mineral resource and mineral reserve estimates, that the resource and reserve estimates have been properly prepared using acceptable methods, and that they may be relied upon for project economic analysis. The project shows robust economics and the initial capital payback has already occurred.

The Bisha Mine is a well-established open pit operation, with pre-production having commenced in 2011. The mine has completed oxide phase and has transitioned to the supergene phase in Q2 2013 with a planned throughput of 2.4 Mt/a. The Primary flotation plant is projected to come online in 2016 and the mine life has been projected to 2025.

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1.17 Recommendations

The QPs recommendations for the Bisha operation, satellite deposits, and regional work are described below.

1.17.1 Mineral Resources

The recommendations for each deposit are as described below.

Bisha

Continue in-house grade control drilling to provide additional detail in delineation of supergene mineralized domains and identification of the zinc depletion surface that indicates the start of the primary zone.
Drilling of the down dip extension of the main massive sulphide zone to the west and below the current optimal pit shape limiting the Mineral Resource in order to gain better understanding of the potential for underground mining. The envisaged program of eight holes for 4,800 m would cost in the order of $1.75 million; continuation of such a program would be based on results.
Continue in-house in-situ density testing of the supergene mineralized zone.
Complete an in-house desktop study of underground potential with a view to increasing Resource inventory and understanding of future direction for mining. See also recommendations for Reserves below.

Harena

Drilling for potential strike and dip extension in the centre and south of the deposit. A program of eight holes for 2,400 m is envisaged to determine potential for relatively shallow mineralization costing an estimated $800,000.

Northwest

Drilling of the Eastern zone to better define and upgrade Resource category; identify potential for shallow mineralization to be added to the current Resource. A program of 16 holes for 1,500 m at an estimated cost of $0.6 million is envisaged.
RC drilling of oxide and supergene gold cap to improve sample recovery and better delineate these zones with accompanying upgrade to Indicated category. A program of 120 short holes for 7,200 m at a cost of approximately $250,000 is envisaged.
Down dip drilling of the primary massive sulphides to identify potential as follow-up to the successful deeper drilling completed in 2013. A program of 21 holes for 6,000 m at an approximate cost of $2.1 million is envisaged.

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Hambok

Complete Resource infill program within the current Resource area to extend on strike and better delineate near surface primary mineralization. A program of eight holes for 2,850 m is envisaged at a cost of approximately $1.0 million.

RC drilling of the oxide zone to examine potential for shallow mineralization with strike extension, particularly to the south, and for better delineation. An initial program of 30 holes for 2,250 m is recommended at an approximate cost of $800,000.

1.17.2 Mineral Reserves

The recommendations for each deposit are described below.

Bisha

Complete an in-house trade-off study examining possible underground mining versus ultimate pit expansion. Approximate cost (limited input from external consultants) is $50,000.
Geotechnical:

- BMSC needs to examine blasting practices to develop a method of incorporating geotechnical information into the blast designs, since rock quality varies considerably within domains.

- Continue development of Slope Management Plan

Hydrogeological:

- BMSC should review the pit dewatering plans for each zone at Bisha Main with consideration of the updated pit designs. The ability of the planned dewatering system to achieve the pit slope depressurization required by the slope designs presented in the current work should be confirmed. The number of wells, locations of wells, and requirements for horizontal drains may require revisions from those presented for previous open pit designs.

- The Knight Piésold hydrogeological report (2014) identified three aquifers that will influence the seepage of groundwater into the pit, of which the middle Fractured Bedrock aquifer is the most significant as it will be encountered throughout the life of mine. This aquifer, however, needs to be correlated with the major structures and the geotechnical domains, which are expected to have different hydraulic conductivities and will change as mining progresses. The fresh rock unit (FRK) has very good RQDs and should have a lower hydraulic conductivity than the overlying WRK (weathered rock unit) in which most mining is currently taking place.

- Minimizing seepage into the pit will require the drilling and equipping of at least four in-pit dewatering boreholes that will intersect the major structures that are thought to be recharging the permeable fractured rock around the orebody. Sumps and interception drains are required for daily management of seepage in the floor of the mining area.

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- An ongoing program of targeted horizontal depressurization holes into the pit slopes needs to be implemented as mining progresses to relieve the pore pressures that build up behind the pit walls.

Reconciliation:

- Confirm process plant production data vs. concentrate shipment outturn data once a consistent pattern of concentrate movements has been established.

- Confirm metallurgical response of transition material identified between supergene and primary zones

- Additional testwork to examine processing characteristics for the transition zone of mixed copper minerals, with zinc, at the top of the primary zone; approximate cost of $200,000.

Harena

Additional geotechnical drilling targeting the primary sulphide Harena open pit should be completed to provide data for rock mass characterization, hydrogeological characterization, and structural geology assessments. Geotechnical core logging, packer testing, and core orientation should be completed in 3 to 5 holes.
Metallurgical testwork on supergene material, as for primary transition mineralization above, is dependent on potential quantity, which will be assessed.

Northwest

Complete a prefeasibility study leading to Mineral Reserve estimate from existing data for improved LOM planning options; approximate cost is $200,000.

Hambok

In line with Eritrean government’s requirements:

Complete an in-house study to better understand project economics; approximate cost (limited input from external consultants) is $25,000.
Preparatory to a prefeasibility study leading to a Mineral Reserve estimate:

- complete a program of six geotechnical holes, testing, analysis and reporting, approximate cost is $1.0 million

- complete a program of three metallurgical test holes with associated testwork, analysis and reporting, approximate cost is $500,000.

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- complete a program of four hydrogeological test holes, with associated pump testing, analysis and reporting, approximate cost is $700,000

- complete a program of waste rock characterization, approximate cost is $200,000

Complete a prefeasibility study leading to Mineral Reserve estimate for improved LOM planning options; approximate cost is $20,000.

1.17.3 Metallurgical

Additional metallurgical studies for Bisha Main, Harena and the Northwest zone include:

Additional batch and locked cycle flotation testwork is recommended to support the future design of the primary ore flotation circuit and support associated metals recovery factors.
Investigate the effects of oxidation on supergene and primary flotation performance recognizing that this issue is common to other volcanogenic massive sulphide deposits.
Investigate the occurrence and deportment of gold and silver in Bisha primary mineralization to improve payability.

These metallurgical recommendations have an approximate cost of $1.5 million.

1.17.4 Exploration

Additional exploration work is envisaged both regionally and around the Bisha/Harena/Northwest deposits and programs of diamond drilling, geophysics, geochemistry, and associated works are planned.

Within the Mogoraib EL, drilling will be along the Hambok mineralized trend and related targets. The approximate cost of exploration activities for 2014 is $2.85 million.

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2 Introduction and Terms of Reference

Unless specified, all measurements in this report use the metric system. The report currency is expressed in United States dollars and uses Canadian English. The currency used in Eritrea is the Nakfa. The exchange rate for $1.00 is equal to 15 Nakfa.

2.1 Qualified Persons, Principal Contributors, and Areas of Responsibility

The QPs, as defined in NI 43–101, responsible for the preparation of the report are:

Jay Melnyk, P.Eng., Principal Mining Engineer (AGP)
Paul Gribble, C.Eng., FIMMM, Chief Resource Geologist (BMSC)
Peter Munro, BAppSc, Senior Principal Consulting Engineer (Mineralurgy).

Many have also contributed their time and expertise in preparation of the report:

Serge Smolonogov, MAIG, Technical Services Manager (BMSC)
Andri Wiratama, B.Sc., Senior Life-of-Mine Planner (BMSC)
Derek Kinakin, M.Sc., P.Geo., P.G., Senior Engineering Geologist, BGC Engineering Inc. (BGC), Canada
Jeremy Christensen, M.Sc., Pr.Sci.Nat., SAIEG, Principal Engineering Geologist, Geologica, South Africa
Matthew Bampton, MAusIMM, Senior Consultant Geologist, Cube Consulting (Cube) Australia
Adrian Shepherd, MAusIMM CP (Geo.), Senior Consultant Geologist (Cube).

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Areas of responsibility for the report sections are given in Table 2‑1.

Table 2-1: Report Areas of Responsibility

Report Section	Responsible Author(s)	Report Section	Responsible Author(s)
1	All	15	JM
2	PG	16	JM
3	JM	17	PM
4	PG	18	JM
5	JM	19	PM
6	PG	20	JM
7	PG	21	JM
8	PG	22	JM
9	PG	23	PG
10	PG	24	All
11	PG	25	All
12	PG	26	All
13	PM	27	All
14	PG	28	All

Note: Jay Melnyk (JM), Paul Gribble (PG), Peter Munro (PM), All (JM/PG/PM).

2.2 Site Visits and Scopes of Personal Inspections

The QPs and principal contributors have conducted site visits to, or are based at, the operation as shown in Table 2‑2.

Table 2-2: Dates of Site Visits

Name	Site Visit Dates
Jay Melnyk	Frequent: most recent, 9 to 15 November 2013
Paul Gribble	On site, 17 December 2012 to date
Peter Munro	None
Serge Smolonogov	On site, January 2012 to date
Andre Wiratama	On site, February 2013 to date
Derek Kinakin	Frequent: most recent, 2 to 5 October 2013
Jeremy Christensen	Frequent, 2012 and 2013
Matthew Bampton	Frequent, 2012 and 2013
Adrian Shepherd	December 2012

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2.3 Effective Dates

The Report has a number of data cutoff dates:

Drill data and ore control assay cutoff date of 30 October 2013
Density determination cutoff date of 31 August 2013
Metallurgical test results cutoff date of 13 December 2013
Surveyed month-end pit surface dated 31 December 2013.

Based on these data cutoff dates, the effective date for the Mineral Resources and Mineral Reserves were taken to be 31 December 2013. There were no material changes to the scientific and technical information between the effective date and the signature date of the Report, other than ongoing grade control sampling and production reporting as expected of an operating mine. Therefore, the effective date of this technical report is considered to be 31 December 2013.

2.4 Information Sources and References

Section 27 includes a list of information sourced from reports and other documents, which are cited in this report. BMSC provided technical data for the preparation of Mineral Resource and Mineral Reserve estimation. AGP, Mineralurgy, and BMSC have relied on other experts in the fields of mineral tenure, surface rights, and permitting, as outlined in Section 3.

2.5 Previous Technical Reports

Nevsun filed a number of previous Technical Reports on the Project:

Melnyk, J., Waldegger, M., Kinakin, D., Munro, P., Thomas, D. (2012). Bisha Polymetallic Operation, Eritrea, Africa, NI 43-101 Technical Report for Nevsun Resources Ltd., effective date 31 August 2012.
Thomas, D., Melnyk, J., Kozak, A., Khera, V. (2011). Nevsun Resources Limited, Bisha Polymetallic Operation Eritrea, Africa, NI 43-101 Technical Report to Nevsun Resources Ltd., effective date 1 January 2011, and revised 29 March 2011.
Waller, S., Reddy, D., Melnyk, L. (2006). Nevsun Resources (Eritrea) Ltd, NI 43-101 Technical Report on the Feasibility Assessment, Bisha Property, Gash-Barka District, Eritrea: unpublished technical report to Nevsun Resources Ltd., effective date 5 October 2006.
Yu, F., Reddy, D., Brisebois, K., and Melnyk, L., (2005). Nevsun Resources (Eritrea) Ltd. Bisha Property, Gash-Barka District, Eritrea, NI 43-101 Technical Report and Preliminary Assessment, 30 December 2005: unpublished technical report to Nevsun Resources Ltd., effective date 30 December 2005.

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Reddy, D., and Brisebois, K. (2004). Technical Report on the Bisha Property and Resource Estimate of the Bisha Deposit, Gash-Barka District, Eritrea, 1 October 2004: unpublished technical report to Nevsun Resources Ltd., effective date 18 November 2004.
Barrie, C.T., and Giroux, G.H. (2009). Hambok Deposit, Mogoraib Exploration License, Gash-Barka District, Western Eritrea 43-101 Technical Report and Preliminary Resource Assessment Prepared for Sanu Resources Ltd.

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3 Reliance on Other Experts

The authors have relied upon information derived from the following expert report, pertaining to mineral tenure, surface rights, permitting, environmental liabilities, and social issues:

Bourchier, F., 2014, Re: Bisha Operation, letter from Chief Operating Officer of Nevsun to Jay Melnyk, AGP Mining Consultants Inc., dated 14 March 2014.

3.1 Mineral Tenure

The QPs have not reviewed the mineral tenure, nor independently verified the legal status or ownership of the Project area or underlying property agreements. The QPs have relied upon information obtained from the above-mentioned document. This information is used in Sections 4.2, 4.3, and 4.4 of this report.

3.2 Surface Rights and permitting

The QPs have relied on information regarding the status of the current Surface Rights, Road Access, and Permits through opinions and data supplied by BMSC experts through the above-mentioned document. This information is used in Sections 4.5 and 4.7 of this report.

3.3 Environmental Liabilities

The QPs relied on information regarding Environmental Liabilities through opinions and data supplied by BMSC experts through the above-mentioned document. This information is used in Section 20.3 of this report.

3.4 Social and Community Impacts

The QPs have relied on information regarding the status of Social and Community Impacts through opinions and data supplied by BMSC experts through the above-mentioned document. This information is used in Section 20.6 of this report.

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4 Property Description and Location

The Property is located 150 km west of Asmara (233 km by road), 43 km southwest of the regional town of Akurdat, and 50 km north of Barentu, the regional, or Zone Administration Centre of the Gash-Barka District (Figure 4‑1), in Eritrea, East Africa.

The Property is centred at 1,711,000 N and 334,500 E (UTM Zone 37), or 15°28’ N and 37°27’ E.

Figure 41: Property Location Map

4.1 Property and Title in Eritrea

Within the State of Eritrea, property title is granted in Agreements with the State of Eritrea under the provisions of Proclamation No. 68/1995 “A Proclamation to Promote the Development of Mineral Resources.”

Licences are granted and identified according to the level of exploration work completed on a property. Properties are granted under the following licence types: prospecting licences, exploration licences, or mining licences. Properties can be obtained under one type of licence and can be converted to the subsequent type if all obligations are met and the titleholder is not in breach of any provisions of the Proclamation and the appropriate application (with fees) are submitted.

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A Mining Licence entitles the licensee a 90% interest and the State of Eritrea holds the remaining 10% interest, without cost. The State may acquire up to an additional 30% (total not exceeding 40%) by agreement with the licensee and by funding their share of the development and operating costs.

Under the Regulation of Mining Operations (Legal Notice 19/1995), the holder of a Mining Licence shall pay the Eritrean government:

Royalty for all minerals produced (see below)
Income tax in accordance with the Proclamation No. 69/1995
Licence renewal fee
Annual rental fees for licence areas (as described above).

Additionally, the holder of a licence and his contractors shall pay a 0.5% customs duty on all imports into Eritrea of equipment, machinery, vehicles, and spare parts (excluding sedan style cars and their spare parts) necessary for mining operations.

The royalty to be paid by a licensee pursuant to Article 34 (1) of the proclamation shall be as follows:

For precious minerals the royalty is 5.0%
For metallic and non-metallic minerals including construction minerals, the royalty is 3.5%
For geothermal deposits and mineral water, the royalty is 2.0%.

Notwithstanding this law, a lesser rate of royalty may be provided by agreement with the licensing authority, when it becomes necessary to encourage mining activities.

Taxation rates are described in the Proclamation No. 69/1995 “Proclamation to Provide for Payment of Tax on Income from Mining Operations.” A holder of a mining licence shall pay income tax on the taxable income at a rate of 38%. Taxable income is to be computed on a historical accrual accounting basis by subtracting from gross income for the accounting year by taking into consideration all allowable revenue, expenditure, depreciation, which, for tax purposes, is deducted straight-line over four years, reinvestment deduction and permitted losses.

If any licensee transfers or assigns, wholly or partially, any interest in the licence, the proceeds shall be taxable income to the extent that such consideration exceeds the amount of his un-recovered expenditure.

Withholding taxes and personal income taxes of non-residents of Eritrea are identified within the proclamation. If the licensee contracts a company or person, who is not resident in Eritrea for services in Eritrea, the licensee will pay taxes on behalf of such a person. Taxes will be paid at the rate of 10% on the amount paid. For the purposes of this article in the proclamation, a person is temporarily present in Eritrea if he performs work in the country for more than 183 days in any accounting year. The compensation received by an expatriate employee of the licensee or his contractor shall be subject to an income tax at a flat rate of 20%.

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The holder of a Mining Licence producing exportable minerals can open and operate a foreign currency account in Eritrea and retain abroad a portion of his earnings to be able to pay for importation of machinery, pay for services, for reimbursement of loans and for compensation of employees and other activities that may contribute to enhancement of the mining operations.

4.2 Property Ownership

An Eritrean registered corporation, BMSC, holds the Bisha Property. The shareholder structure of BMSC is 60% Nevsun and 40% ENAMCO.

ENAMCO agreed in October 2007 to purchase the 30% paid participating interest. The purchase price of $253.5 million was agreed to between ENAMCO and Nevsun in August 2011 with the amount being paid over time based on free cash flow generated by the Operation.

The Nevsun interest in the Property and Nevsun inter-corporate holdings are summarized in Figure 4‑2.

Figure 42: Nevsun Property Ownership Diagram

Source: Nevsun, 2012

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4.3 Mineral Tenure

The Property comprises two mining licences covering an area of 24.0 km², (16.5 km² over the Bisha and the Northwest Zone and 7.5 km² over Harena) and a mining agreement area covering an area of 39.0 km². BMSC is the operator for all of the licences.

In October 2012, BMSC purchased the Mogoraib EL. In the first year of tenure, this licence covered some 97.4 km², which was subsequently reduced by 25% in accordance with Eritrean policy to some 73.1 km². At the time of writing BMSC has tenure over the licence until July 2014; renewal of the licence will be the result of the normal reapplication process. The Mogoraib EL is centred at approximately 1,705,000 N 325,000 E.

The licence areas that form the Property are shown in Figure 4‑3.

Figure 4‑3: Location of the Bisha Mining Licences and the Mining Agreement Area

Note: In this report, “Bisha Village” is more frequently referred to as “Bisha Camp.”

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BMSC have the exclusive right to apply for and be granted multiple Mining Licences within the Mining Agreement Area. The Mining Agreement entitles BMSC to apply for a mining licence valid for a period of up to 20 years, with renewal periods of up to 10 years providing:

sufficient ore has been defined to demonstrate continued economic viability of Mining Operations
BMSC has fulfilled the obligations specified in the Mining Licence and the Mining Agreement
BMSC is not in breach of any provision of the Mining Proclamation and which would constitute grounds for suspension or revocation of the Mining Licence.

The Eritrean Ministry of Energy and Mines (the Ministry) granted the Bisha Mining Licence on 26 May 2008 and is valid for a period of 20 years. The Harena Mining Licence was granted on 06 July 2012 and is valid for a period of up to 10 years, with renewal periods of up to 10 years.

An Exploration Licence may be converted to a Mining Licence upon the acceptance by the State of Eritrea of an appropriate feasibility study and Environmental Impact Assessment (EIA) report.

The annual rental fee for an Exploration Licence is 53,200 Nakfa, and the annual licence renewal fee is 6,000 Nakfa (about $3,500 and $400), respectively. BMSC’s 53 km² exploration licence officially expired in May 2012. BMSC and the Ministry officials are working together to re-establish an exploration licence or licences that cover a larger area so that BMSC may carry out a more significant regional exploration program. BMSC has advised the State that it wishes to expand its exploration efforts and the State has welcomed this approach.

BMSC has surveyed the boundaries of the Mining Licence Area in accordance with the Mining Proclamation law. BMSC is not required to survey the Mining Agreement Area or to place Mining Agreement Area boundary markers. Exploration licences also do not require survey.

4.4 Surface Rights

Under the terms of the Mining Agreement, BMSC has the exclusive right of land use in the Mining Licence Area that is granted within the Mining Agreement Area. This right is subject to the acquisition and settlement of any third-party land-use rights by payment of compensation and/or relocation at the expense of BMSC, in accordance with Eritrean Government Proclamation No. 68/1995, “Proclamation to Promote the Development of Mineral Resources and the Mining Agreement.”

4.5 Royalties and Encumbrances

Royalties payable include an Eritrean Government gross royalty of 5% on precious metals and 3.5% on base metals.

There are no encumbrances on the property.

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4.6 Property Agreements

In December 2007, BMSC concluded a confidential Mining Agreement with the Government of the State of Eritrea containing all the normal provisions governing the future development and operations for the Bisha Property.

AGP reviewed the confidential document and is satisfied that the terms of the agreement are consistent with the assumptions used in the Mineral Reserve estimation and financial analysis.

4.7 Permits

For the mining operations, grant of the mining lease provided permission to construct and operate the Bisha Mine. A permit has been granted for use of water from the Mogoraib River and construction of necessary water diversion structures.

These permits are sufficient to ensure that mining activities at Bisha are conducted in accordance with the appropriate National laws. BMSC has conditional approval for the Mining Licence for Harena subject to completion of detailed geotechnical design, hydrogeological studies, and waste rock characterization testing by the end of December 2012. BMSC completed these studies and results were duly submitted to the relevant authorities before the end of 2012.

BMSC commenced the process for the mine development portion of the Project in 2004, undertaking environmental and socioeconomic baseline studies and an environmental assessment.

The Ministry approved the Terms of Reference for the project environmental and SEIA in March 2006. A SEIA report was completed in December 2006 and submitted to the Ministry.

During 2007, a review of the report was conducted by the Ministry of Land, Water and Environment, by an appointed “Impact Review Committee.” Comments and queries raised by the latter were addressed by BMSC in 2007–2008. A Mining Licence for the Project was issued on 26 May 2008. Issuance of the licence is accepted by BMSC as SEIA approval by the Impact Review Committee, as required under the Eritrean Mining Regulations. The Harena SEIA Addendum was submitted December 2011 with the conditional mining licence granted in July 2012. The SEIA Addendum is still under review.

As the transportation route has already been constructed as part of the national transportation system, an assessment of the environmental effects that were associated only with the transport of hazardous materials (i.e., cyanide and fuels) and increased traffic was addressed in the SEIA.

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4.8 Environmental Liabilities

The key environmental issues assessed by the SEIA studies and addressed in Project associated risk assessments and the environmental management plan is as follows:

Direct footprint disturbance of 442 ha (Bisha) and 200 ha (Harena) with associated potential for loss of land use, habitat, soils loss, and drainage disturbance.
Groundwater impacts from both extraction of Project supply water from new wells and excavation of an open pit.
Water quality impacts arising from potential for acid rock drainage (ARD), including the need to ensure that there is adequate post-closure monitoring to ensure no post-closure problems with water quality.
Soil and water quality impacts arising from the storage and use on site of hazardous chemicals, including cyanide.
Changes to local surface drainage patterns due to construction of a site surface water management system, including flood control and diversion works.
Air quality impacts, most significantly from surface haulage on unsealed roads.

BMSC has provided a remediation bond with the State of Eritrea in the amount of $15 million. BMSC has also accrued in its financial records a provision for mine rehabilitation of $22.8 million for the estimated costs of reclamation, remediation, and post-closure monitoring.

4.9 Social Licence

Since exploration and environmental baseline data collection began, considerable effort was spent developing support for the Project by fostering local relationships, developing a strong local workforce, educating stakeholders about the Project and mining in general, and providing stakeholders with regular Project updates and, where appropriate, site visits. The key socioeconomic issues assessed by the SEIA studies and addressed in the proposed social management and related plans are as follows:

Direct footprint disturbance of 442 ha (Bisha) and 200 ha (Harena) with associated potential for displacement of people and their customary use of the land (although it is noted that the affected area is sparsely populated and only lightly used).
Influx of people seeking employment with associated potential issues, including pressure on existing social infrastructure.
Inward investment and creation of direct and indirect employment opportunity to 31 December 2013, the total land disturbed was 600 ha, and total land rehabilitated was 24.6 ha.

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5 Accessibility, Climate, Local Resources, Infrastructure and Physiography

5.1 Accessibility

Asmara is the capital city of Eritrea and is serviced by regular international flights.

Access to the Property is by paved road from Asmara to Akurdat, and then from Akurdat via an all-weather unpaved road, which is currently being upgraded. The route is shown in Figure 4‑1. The drive from Asmara to the Bisha camp takes approximately 4 hours. The main distances by road to the Bisha Property are summarized in Table 5‑1. The principal port for importation of heavy equipment is Massawa on the Red Sea coast, which is about 350 km from the Property by road, via Asmara to the east.

Table 51: Distances by Road to the Property

From	To	Distance (km)	Condition
Asmara	Akurdat	181	Paved, all-weather road
Akurdat	Adi Ibrahim	28	Unpaved, all-weather road
Adi Ibrahim	Hashakito	19	Unpaved, all-weather road, being upgraded, approx. half now paved
Hashakito	Bisha Camp	5	Unpaved, all-weather road
Asmara	Bisha Camp	233	4 hour drive
Bisha Camp	Bisha Main Deposit	4	Unpaved, all-weather road
Bisha Plant	Harena Deposit	10	Unpaved, all-weather road

5.2 Climate

The climate in the area is semi-arid, with elevated temperatures year-round. During the hot season in April and May, the average temperature is +42°C, although temperatures may rise to +50°C for short periods. The main rainy season is between June and September, and periodic flooding of the Mogoraib and Barka rivers can result in spectacular flash floods. Occasional rain may also fall during April and May. Total rainfall is sparse, with an annual precipitation of between 250 mm and 300 mm.

The rainy season causes periodic, short-lived difficulty in travel off the main highways, although exploration work is possible year round. During the period of exploration work by BMSC, the precipitation has only occasionally been sufficient to flood the local rivers. All mining activities are planned on a year-round basis.

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Figure 51: Property Access and Site Layout

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5.3 Physiography

The Property is located on a flat to rolling, desert-like plain along the western foot of the Central Highlands. The plain is at 560 masl and contains scattered vegetation and few trees. Steep hills and ridges rise above the plain; the Bisha, Wade, and Neve peaks reach elevations of up to 1,226 masl above the alluvial plain at the southern boundary of the Property (Figure 5-1). A smaller seasonal tributary, the Fereketetet River, flows north-northwest into the Mogoraib River. The Fereketetet River crosses the Property and passes immediately west of the Bisha Gossan Zone.

5.4 Local Resources and Infrastructure

The following subsections detail the local resources, and the existing and projected infrastructure associated with the Project.

5.4.1 Local Resources

There are few local resources in the Bisha area.

In 2004, Klohn Crippen Berger Ltd. (KCB) conducted a preliminary land-use survey near the proposed mine site. It was determined that approximately 96% of the area was used by local herders as pasture for livestock, and used seasonally for activities including agriculture (dryland crops), domestic livestock migration, and accessing wells and burial sites. Currently the land is overgrazed, which is related to ongoing drought conditions and pressures from livestock foraging. This study was conducted in consultation with people from local communities.

An additional survey was conducted in February 2006 for the six communities within the Bisha area; these included Tekeret, Adi-Ibrihim, Hashakito, Jimmel, Adarat-Harenay and Takawda. Most of the people in the region are located in permanent settlements, but utilize the Bisha area as one of the many used for grazing livestock, planting crops, and accessing watering areas, which in some cases involves migrating distances up to 200 km, as herders move through the region in search of suitable grazing lands.

The village of Mogoraib is the local administration centre for the Dighe Sub-zone within the Gash-Barka District. The village has a small refugee resettlement site, and subsidiary military and commercial interests. The village contains a well-equipped, eight-person health centre with a nursing staff capable of taking care of small medical problems in preparation for referring patients to larger, better-equipped hospitals in Akurdat and Keren. Camp Mogoraib is a military training site located just outside the village boundaries. With the presence of the mine development and exploration project at Bisha, this camp has been re-activated as a security post from its previous care/maintenance basis.

Few basic goods are commercially available in the region, either in Mogoraib or in Akurdat. The main centre for support of exploration and project development is the capital city, Asmara.

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The local population has no exploration or mining culture, and training of local staff is ongoing.

5.4.2 Infrastructure

Current onsite operation infrastructure includes:

open pit
process plant
tailings and waste-rock storage facilities
offices
maintenance and laboratory facilities
fuel storage areas
on-site power plant
airstrip (not shown as located immediately north of Figure 5-1)

The key areas are indicated in Figure 5-1.

In addition, there is off-site infrastructure that includes container port and ship-loading facilities at the port of Massawa. Concentrate is trucked in specially built reusable intermodal containers from Bisha Mine to the port of Massawa, and stacked at the existing container facility. The concentrate in the containers is discharged into the bulk carriers using a mobile crane along with Rotainer’s Lid-Lift equipment, which rotates the container a full 360° after lifting the lid.

5.4.3 Power

Electric power for the mine and processing plant site is supplied from a diesel-fuelled power station located adjacent to the process facilities. The container port near Massawa receives power from the local utility.

5.4.4 Water

Process water is sourced from recycling within the plant, and additional needs are supplemented from freshwater sources. The process was designed to maximize the recycle of process water, and included installing a tailings slurry thickener to recover process water prior to pumping to the tailings containment system. This approach served to minimize the evaporation losses that result with the typically large water surface area in tailings containment systems. Even though evaporation rates in this region are very high, a tailings-management facility, supernatant water reclaim pumping-system is installed to reclaim seasonal decant water from this source. In addition, water from the pit is pumped to the process plant, which further reduces dependence on raw water.

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Freshwater is supplied to the property from groundwater. Two well farms have been established by BMSC, the first approximately 1 km south of the open pit on the western bank of the non-perennial Fereketetet River (which also serves to dewater the Bisha pit), and the second 5 km to the west, adjacent to the Mogoraib River (Figure 5-1). Potable water sourced from the well fields is pumped to a potable water plant utilizing chlorination filtration and ultraviolet radiation treatment.

5.4.5 Communications

Current site communication is via radio, cellular service, and a satellite communications system.

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6 History

6.1 Bisha and Harena

BMSC has no record of any exploration or development work on the Bisha or Harena areas prior to 1996.

In late 1996, Ophir Ventures, a private Canadian company, conducted prospecting in the Bisha area that resulted in the discovery of the surface exposure of the Bisha deposit.

Nevsun was granted a prospecting licence for Bisha in May 1998, and the Bisha Main Zone was recognized through drilling by Nevsun in 2002. In 2006, BMSC was created with ownership as 60% Nevsun and 40% ENAMCO, as described in Section 4.2.

6.2 Northwest

During 2005, Sanu Resources (Sanu) drilled 11 diamond core holes immediately north of the area of what is described in this report as the Northwest deposit. The results were inconclusive, with only narrow zones of massive sulphide intersected in four holes. Minor zinc but no copper mineralization was found. Sancu relinquished this licence area before 2009.

6.3 Mogoraib River Exploration Licence

The Mogoraib EL was granted to Sanu in October 2003. Sanu reported that it had no record of any previous exploration in the licence area. Exploration works started that year and continued to 2012, when BMSC purchased the licence. A summary of activities undertaken is given in Table 6‑1, and the prospects described in this table are shown in Figure 6‑1. All information and figures in this section are sourced from Sanu reports and data available to BMSC.

Table 6‑1: Sanu Resources Exploration Program 2003 – 2012

Year	Work Type	Work Completed
2003–2004	Airborne geophysical survey	Mai-Melih area (electromagnetic (EM) and magnetic surveys)
	Bulk Leach Extractable Gold (BLEG) sampling	217 samples from 290 km² area
	Detailed geological mapping	Mai Melih gossan, Ashelli, Nageib (1:1,000 scale) Tekewuda, Ankerite village prospects (1:2,000 scale)
	Soil/rock chip sampling in several areas	Mai Melih: 207 soil, 60 grab Ashelli: 257 soil, 46 grab Tekewuda: 458 soil, 20 grab Nageib: 305 soil, 14 grab Ankerite: 117 soil, 10 grab

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Year	Work Type	Work Completed
	Ground EM surveys	Mai Melih: 9.9 km Ashelli: 7.2 km Sileceous breccia: 7.2 km Tekewuda: 5.4 km Total: 29.7 line-km
	Trenching	579 m of 9 trenches in Mai Melih gossan, and 278 trench samples
2004–2005	Gravity surveys	Mai Melih: 19 lines, 1.2 km Siliceous breccia: 13 lines, 1.0 km Ashelli: 15 lineslines, 0.9 km Tekewuda: 10 Lines. 1.0 km Shabait: 13 lines, 1.4 km Total: 70 lines over a 5.5 km² area
	Soil sampling	Total: 2,412 soil samples from all prospects
	Ground EM surveys	Mai Melih: 37.8 line-km
	Gravity surveys	Mai Melih, Hambok, Ashelli, Tekewuda, Shabayet Total: 2,031 gravity points
March 2005–2006	Soil sampling	220 soil samples from Hambok prospect.
	Rock chip sampling	64 rock chip samples from 290 km² area
	Ground EM surveys	Hambok (south): 6.3 km Hambok (North) Ankerite: 18.0 km Mogoraib River: 9.0 km Total: 33.3 line-km
	Gravity survey	10,452 gravity points on extended grids of 2004–2005 surveys in the same areas
	Drilling	50 DDH holes. Total of 8,411 m drilled at Mai Melih, Ashelli, and Hambok: Mai Melih 510 m Ashelli, 619 m Hambok 7,282 m
2007	Detailed gravity survey	Regional: 300 m x 300 m grid, and Gossan hill: detailed 100 x 50 m grid survey, 1,948 gravity points; Shabait West: detailed 100 x 50 m grid, 796 gravity points
	Ground EM	Horizontal max-min and deep EM in Ankeritre, Shabait, and Ashelli prospects
	Ground magnetic survey	Total 500 line-km completed in Hambok, Ankerage, Shabait, Watey Ridge, Gossan Hill, Ashelli and Tekewuda prospects
	Drilling (extension)	Hambok: 2 holes, 717 m Hambok North: 2 holes, 207 m

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Year	Work Type	Work Completed
		Ankerite: 3 holes, 690 m Taebeit: 2 holes, 400 m Shabait: 4 holes, 1,058 m Ashelli: 4 holes, 520 m Total: 17 holes, 3,592 m
	Regional geochemistry stream sediment sampling	Regional: 3,075 samples, 400 x 100 m grid Bisha North: 815 samples, 200 x 50 m grid Bisha North: 137 samples, 400 x 50 m grid Gossan Hill: 906 samples, 100 x 50 m grid Wetey North: 405 samples, 400 x 50 m grid Ashelli-Akub: 695 samples, 100 x 50 m grid Ashelli-Adal: 235 samples, 100 x 50 m grid Mogoraib central: 600 samples, 400 x 50 m grid Total: 6,868 samples
	Trenching	Shabait West: 3 trenches Shabait South: 1 trench Shabait North: 4 trenches Ashelli North and Hambok North: 5 trenches
	Rock chip/grab sampling	Total 421 Chips and 96 Grab samples
	Geological mapping	Ashelli South and Shabait (1:2,000 scale) prospect mapping; In area between Hambok, Shabait, Tekewuda, and Southern boundary of the licence, 1:10,000 scale
2008–June 2009	Rock chip/grab sampling	Total 329 chip/grab samples
	Trenching/channel and pit sampling	Aderat prospect: total 198 trench and channel chip samples; Aderat prospect: 27 pits dug to expose bedrock for mapping objectives
	Drilling	Aderat-Ankerite prospect: 6 DDH 1,209 m drilled; Shabayet prospect: 1 DDH
	Geological mapping	Regional mapping (1:10,000 scale)
	Geological mapping	Hambok North, Aderat, Ankeri: detailed mapping of backhoe trenches
July 2009–July 2010	Geological mapping	Regional (1:100,000 scale) mapping updates. Prospect mapping: Akub and Basur. Detailed mapping of Hambok-Ankerite
	Trenching	Two pits extended for trenching. Forty-five samples collected in trenches in Mogoraib EL
	Drilling	Hambok: 3 DDH infill, 820 m Hambok South IP anomaly: 1 DDH 137 m deep Akub: 1 DDH, 74 m
	Ground magnetic survey	Aderat and Hambok: 16 km² IP survey at 200 m spacing and 50 m station grids

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Year	Work Type	Work Completed
	Ground magnetic survey	Aderat and Hambok: 16 km² IP survey at 200 m spacing and 50 m station grids
	Geophysical IP survey	Total 130 km² at 400 m line spacing covering Mogoraib EL area
Aug 2010–2011	Airborne versatile time domain electro-magnetic (VTEM)	Hambok: Block 1 reveals east-dipping sulphide body Block 2: moderately strong conductor, 60° SE-dipping Block 7–8: two conductors 4 km SW of Ashelli, shallowly E-dipping
	Ground magnetic survey	Akub prospect: 19 lines at 200 m spacing, 3.6 km² covered
	Drilling	Hambok: 4 DDH infill holes; 987 m were drilled
	Geological mapping	Regional (1:100,000 scale) mapping updates
July 2011–June 2012	Drilling	Hambok: DDH-DT combined: 27 infill holes, 8,962 m drilled targeting primary massive sulphides. RC: 42 holes, 2,675 m for Au oxide drilling program
	Follow-up prospection VTEM	Ground follow-up of VTEM and Hambok South VTEM (Blocks 1-11) interpretation for drill targeting
	Prospect re-mapping	Follow-up of Mogoraib EL 1:100,000 scale regional mapping

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Figure 6‑1: Sanu 1:50,000 Geological Map of Mogoraib River Exploration Licence with VTEM Anomalies and Prospects

6.3.1 Summary Prospect Descriptions

Exploration activities completed by Sanu between 2004 and 2012 for the different prospects in the Mogoraib EL area are summarized in this section.

Mai Melih Prospect

The Mai Melih prospect lies outside the northern boundary of the licence area. The gossan outcrops of the Mai Melih area range from 2 m to 30 m thick, extending over 400 m along strike. These gossans were the main focus of exploration by Sanu.

The ground EM and airborne geophysical survey results confirmed strong conductors that correlated well with the gossan and a prominent gravity trough; the gossan were thought to indicate oxidation of massive sulphides.

Three holes were drilled in the Mai Melih prospect. The results of the drilling were not positive, as only thin, low-grade mineralized horizons were encountered, and a drill hole directly underneath the main gossan outcrop did not intersect any mineralization. No further exploration work was done by Sanu in this area.

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Siliceous Breccia (Hambok)

Extending several hundred metres along strike on the north side of Hambok, geochemical sampling in this area returned very promising results during the early stages of the work completed by Sanu. Geophysical surveys also indicated that the siliceous breccia gossans correlated well with the EM signatures. In 2012, 4 DDH were planned, but no further drilling was completed in this area.

Ashelli Gossan Prospect

The gossan is close to the western edge of the licence. Rock chip samples indicated anomalous values of base metals and gold.

A gravity signature was identified, together with an EM conductor in close proximity (100 m to 300 m), that was thought to represent a potential massive sulphide body. Four diamond holes (520 m) were drilled in this area. Results of the drilling showed Au and Ag mineralization in two holes, one of which had a significant intercept of a 20.8 g/t Au over 1 m, within an intersection of 22 m at 1.58 g/t Au. Sanu planned to carry on exploration work in the area but did not complete any further work.

Tekewuda

Tekewuda lies in the southeastern extremity of the Mogoraib EL. Sanu carried out a gravity survey, 1 km in length at 100 m line spacing, plus prospect mapping on a 1:2,000 scale in the area. A ground EM survey, comprising 5.4 line-km, and rock chip sampling were also completed.

The EM survey did not show any significant results or any correlation with the gossans. The results of the gravity survey provided some positive signatures at about 100 m west of the gossan. Sanu did not indicate any intention of future exploration work in the Tekewuda area.

Nageib

The area lies between the Ashelli gossan prospect and Melih village. Detailed (1:1,000 Scale) geological mapping and soil geochemistry sampling was done in the area. The results of sampling were disappointing, and no further work was anticipated.

Aderat-Ankerite Prospect

The Ankerite prospect is situated north-northeast of the Hambok deposit along the same strike/mineralized trend. Prospect geological mapping was followed by the collection of 117 soil geochemistry and 10 grab samples from the Ankerite area during 2005 and 2006. As a result, Sanu drilled 3 DDH in 2007.

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Exploration activities continued at the prospect after this period, and in 2009, 6 DDH were drilled in which massive sulphide-style mineralization was intersected. Four out of the six holes intersected disseminated and stockwork mineralization, out of which two holes intersected Zn and Au-Cu-Zn mineralization. Drill hole ANK-006 intersected two mineralized zones, a shallow zone of 1.2% Zn over 47 m (drilled thickness), and another, deeper Au-Cu-Zn mineralization zone of 0.83 g/t Au, 0.71% Cu, and 3.3% Zn over 29 m.

Further geophysical IP and ground magnetic survey was carried out in the area. Sanu planned to continue exploration in this area using geophysical tools.

Hambok

Hambok was the main focus of exploration activities for Sanu. Programs of geochemical sampling, EM surveys, gravity data acquisition, and drilling identified Hambok as a resource development project.

In 2006, exploration drilling at Hambok determined the mineralization style and geometry of the massive sulphides. Sanu’s interpretation was that the massive sulphide mineralized zone was bounded by structures defined as hanging wall and footwall faults.

Ground magnetic surveys, resource development drilling in Hambok and Hambok North, and prospect mapping on 1:2,000 scale continued from 2006–2009.

Sanu completed preliminary drilling of the Hambok deposit during 2009, and produced a Technical Report. During 2010–2012, a VTEM geophysical survey was conducted, with infill diamond core drilling of the massive sulphide zone, and shallow RC drilling targeting the oxide mineralization.

Several prospects evolved within and around the Hambok deposit, namely Hambok South, Hambok North, Hambok West, Hambok SE prospect, and Gossan Hill Prospect, which is 5 km southwest of the Hambok deposit. Large amounts of geophysical data acquisition were completed by Sanu in Hambok and the associated prospects during the 2010 exploration phases and subsequently.

Exploration Expenditure

The expenditure reported by Sanu for its exploration work on the Mogoraib EL is incomplete, but is shown in Table 6‑2.

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Table 6‑2: Sanu Reported Expenditure during Tenure of the Mogoraib EL

Period	Reported Approximate Expenditure ($)
November 2003 to April 2005	270,000
May 2005 to June 2006	610,000
July 2006 to December 2007	not reported
January 2008 to June 2009	1,000,000
July 2009 to July 2010	466,000
August 2010 to June 2011	Not reported
July 2011 to June 2012	3,000,000

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7 Geological Setting and Mineralization

7.1 Regional Geology

The BMSC projects are located within the 0.85 to 0.55 Ga Arabian-Nubian Shield (Barrie et al., 2007). Precambrian rocks are exposed on either side of the Red Sea in western Arabia and northeastern Africa (Figure 7‑1). The rocks are a collage of volcanic arcs, granitoid intrusions, volcano-sedimentary basins, and shear zones (Johnson, 2013).

The Arabian-Nubian Shield comprises a wide range of deposit types and settings, including volcanic massive sulphide (VMS) deposits. Ages of VMS deposits within the Nubian Shield to range between 850 Ma to 680 Ma (Johnson, 2013). Deposit settings for Cu-Zn-Pb-Au VMS systems are considered as island arc (Johnson, 2013) or back-arc (Barrie et al., 2007). Within the Nubian Shield, there are at least 60 VMS deposits.

Figure 7‑1: Arabian-Nubian Shield

Note: Location of Bisha and Eritrea (modified from Johnson et al., 2003 in Wasylik, 2011).

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The Neoproterozoic basement that covers Eritrea is subdivided into four terranes:

Barka to the west

Hagar in the north

Nakfa in the centre

Arig in the east.

The Nakfa terrane is further subdivided into eastern and western zones (Figure 7‑2). The Eritrean terranes strike north to northeast and are separated by major crustal sutures. The Nakfa terrane comprises mafic and ultramafic cumulates, diorites and granodiorites that are structurally overlain by greenschist facies metavolcanic and volcaniclastic metasedimentary rocks with calc-alkalic island-arc affinities (Johnson et al., 2003).

The Nakfa terrane is not highly deformed, except along narrow shear zones (De Souza Filho et al., 1998). The volcanic-sedimentary rocks generally strike north-south more or less 25° (Massawe, 2012). Younger geological events during the Cretaceous and Tertiary relate to the opening of the Red Sea. These events are displayed as minor faults and dykes in the BMSC licence area (Massawe, 2012).

Figure 7‑2: Neoproterozoic Terranes of Eritrea

Source: Barrie et al., 2007

Note:

Locations of the VHMS deposits: Bisha, NW = Northwest,
H = Harena, HM = Hambok, E = Emba Derho, A = Adi Nefas,
D = Debarwa, and AR = Adi Rossi.

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7.2 Property Geology

On the western margin of the Nakfa terrain the Property is underlain by low-grade metamorphosed (upper greenschist to lower amphibolite facies) volcanics and sedimentary units (Figure 7‑3).

The precious metals-enriched volcanogenic massive sulphide (VMS) deposits on the property are hosted by an intensely foliated, bimodal sequence of generally weakly stratified, predominantly tuffaceous metavolcanic rocks (Greig, 2004). Felsic lithologies appear to directly host the mineralization, predominate overall, and form the hanging wall stratigraphy. The felsic lithologies are mainly exposed to the west and southwest of the mineralized zones, and grade upward into a sequence of generally fine-grained volcaniclastic rocks. A significant component of mafic metavolcanic rocks occurs in the more obviously bimodal footwall, which is exposed mainly to the east of known mineralized zones.

Figure 7‑3: Geological Map of Eritrea

Source: Nevsun, 2011

To the east and south, the metavolcanic rocks are intruded by felsic to mafic intrusive rocks, now foliated, including those of the aerially extensive Bisha Gabbroic Complex. Sedimentary rocks overlie the felsic component and have been mapped to the west of, and parallel to, the stratigraphic units that host mineralization.

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7.2.1 Stratigraphy

Regionally the sedimentary rocks consist primarily of greywacke, siltstone, shale, marble, and feldspathic arenites with less common conglomerate, magnetic ironstone, quartzite, and massive sulphide lenses. The regional volcanic sequence includes fine-grained pyroclastic rocks of mafic to intermediate composition and pillowed mafic flows, felsic ash and lapilli tuffs (Figure 7‑4).

Figure 7‑4: Bisha Property Stratigraphic Section (Barrie, 2004)

In general, stratified rocks underlying the property can be divided into two parts: an upper, predominantly felsic volcanic that is capped by sedimentary rocks; and a lower volcanic part that is clearly bimodal, at least in the south and east. This lower, bimodal volcanic part appears to be capped by the stratiform mineralized horizons at Bisha and Northwest deposits.

The stratigraphic section near the Bisha deposit comprises, from the base (at the Bisha Gabbroic Complex contact) to the top: carbonates and fine-grained siliciclastic rocks, including siliceous iron formation; felsic lapilli and ash crystal lapilli tuffs with intercalated minor mafic flows and hyaloclastite; and fine-grained volcaniclastic/siliciclastic rocks. Volcanic rocks comprise approximately 50% of the stratigraphic section ±2.5 km from the deposit horizon.

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Rhyolites are the predominant volcanic rock type. These are mostly tuffs, with minor blocky flows and agglomerates present immediately west of the Bisha Main and Northwest deposits. Dacites comprise only approximately 5% of the volcanic strata. Other volcanic rocks include tholeiitic basalts. The strata are cut by Neoproterozoic granite-syenite intrusions and minor mafic dykes/sills, and by Cenozoic felsic and mafic dykes. One suite of quartz and feldspar phyric rhyolite/granite dykes is texturally and chemically distinct from the other felsic strata. They occur as rhyolite porphyry, or as granitic rocks.

Carbonates, quartzites, and siliceous iron formation are present in the lower stratigraphic section to the east of Bisha. The presence of carbonates indicates a relatively shallow depositional environment.

The Bisha Gabbroic Complex is a large (225 km²), partly layered, gabbro–gabbro–norite intrusion that forms high hills in the central and southern part of the Property. The complex extends in a north–northeast to south–southwest orientation for 25 km, and has a maximum width of approximately 12 km (immediately south of the Property). The complex appears to cut strata, but has undergone penetrative deformation, and is presumable coeval, or nearly coeval with the strata. The Bisha Gabbroic Complex is tholeiitic, and compositionally similar to basalts.

The Hambok deposit appears to occupy the eastern flank of a broad anticlinorium cored by basaltic rocks and ultramafic and granitic intrusions, with the Bisha Gabbroic Complex lying to the east. A subordinate, unnamed gabbro–diorite–pyroxenite intrusive complex cores an anticline between Hambok and Ashelli, with the Ashelli and Mai Melih massive sulphide prospects on the western flank of this anticlinorium. The top of the bimodal volcanic sequence appears to be marked by graphitic schists, chert and carbonate horizons, and pelitic sedimentary rocks. The Hambok deposit is present within a sequence of chloritic, volcaniclastic rocks with lesser, massive basaltic to andesitic(?) lavas and felsic tuff.

7.2.2 Structure

In the Property area, rock units generally trend north–northeast, with moderate to steep dips to the east and west. The Bisha Gabbroic Complex broadly forms a north-plunging antiform that appears overturned, with dips generally steep to the east. It appears that volcanic and sedimentary strata were thrust against this buttress from the west-southwest, forming a nape-like structure, with internal antiforms and synforms on a scale of hundreds of metres.

The stratigraphy and principal tectonic fabrics at the Property have been disrupted, at least locally, by late-stage brittle faults. Because of the relatively poor exposure in the area, these are expressed in the main as well-developed topographic lineaments.

Once considered to be significant, regional-scale folding on deposit-scale mineralization distribution is now secondary to the recognition that all of the Bisha VMS deposits defined to date are structurally controlled. Regional and secondary faulting form boundaries to mineralization and have been noted to offset, as well as redistribute, base metals mineralization.

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Evidence of structural deformation having a greater influence than folding on stratigraphy and mineralization is seen from new field mapping, open pit mapping, core logging, and physical mining of ore.

7.2.3 Metamorphism

Nakfa Terrain greenstone belt rocks such as the volcanic and sedimentary units at the Property exhibit upper greenschist to lower amphibolite facies metamorphism. The presence of chlorite, fine-grained amphibole, and local garnet in the mafic rocks supports greenschist to lower amphibolite facies metamorphism (Greig, 2004).

7.2.4 Alteration

Footwall alteration is typically pervasive quartz + chlorite alteration of tuffs, which may extend for tens of metres below massive sulphide units. Immediately below the massive sulphides, there is a thin, but variable, (<3 m thick) zone of silicification and K-feldspar replacement (Chisholm et al., 2003). This zone is more variable in intensity and thickness than the chlorite alteration, and in some cases is entirely absent.

Hanging wall alteration is typically pervasive quartz + muscovite alteration of tuffs, which may extend for tens of metres above massive sulphide units.

Other, lesser-observed alteration mineral assemblages include carbonate, epidote, and albite. These alteration styles range from weak and patchy to strong/intense and pervasive.

7.2.5 Weathering

The weathering profile atop the various VMS deposits varies in depth from 50 m to 100 m below surface. The profile is at times complex as a result of groundwater oxidization and leaching. The profile is affected by low rainfall that falls in two months of the year. Low rainfall and resultant low groundwater flow or circulation provides an environment for the groundwater to become highly acidic, depending on interaction with the local or surrounding geology.

7.3 Deposit Geology and Mineralization

7.3.1 Bisha Main Zone

General

The Bisha deposit is a series of massive sulphide lenses that occur over a 1.2 km north-south trending strike extent. The thickness of the lenses is variable from 0 m to 70 m. Mineralization extends to a depth of 500 m below surface, and as yet has not been closed off. Significant advances have been made in the understanding of the Bisha orebody geology from 2012 to 2014, culminating in a series of reports (Bampton, 2013; Bampton, 2013; Ashley, 2013; McPhie, 2013).

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Creating a coherent geological model for Bisha in 2013 allowed the mineralization at Bisha to be more accurately constrained within a framework of fault domains, and a more accurate distribution and subdivision of weathering surfaces. Previous Mineral Resource Estimates for Bisha were based on mineralization models. The 2013 Mineral Resource and Reserve estimation process used the Bisha Geological Model to guide and constrain creation of a new Bisha Mineralization Model.

Mineralized Stratigraphy

In July 2013, mining of the Bisha deposit had reached the supergene mineralized zone. The mining process had cut a profile through a greater than 35 m deep weathering profile.

Deep weathering has affected Bisha Main Zone lenses by removing most of the sulphide, and producing high-grade supergene blankets enriched in gold, copper, and lead, in particular. The gossan zone can vary in composition from highly siliceous and somewhat ferruginous, to a massive goethite–hematite–jarosite gossan. Oxidation is approximately 30 m to 35 m deep in outcrop areas, but is variable in sand-covered areas. Supergene sulphides occur at depths between 35 m and 65 m, with accompanying carbonate, sulphate, phosphate, silicate, halide, and native base‑metal minerals.

Oxidation of the massive sulphides generated strong acid solutions that have progressively destroyed the sulphides and host rock. A horizon of extremely acid-leached material, or “soap,” has developed between the oxide and supergene/primary domains.

To the west of the massive sulphide lenses of the Bisha Main Zone, there is a zone of copper mineralization referred to as the hanging wall copper zone. The mineralization is located in the stratigraphic hangingwall to the main massive sulphide lenses. The hangingwall copper zone is restricted to the supergene horizon, and has a primary massive sulphide component. The hangingwall copper zone has a north–northeast strike, and converges towards the Bisha Main Zone towards the north. Drill hole intercepts have down-hole thicknesses varying from 2.7 m to 63.5 m.

Within the weathering zone, mineralization has been geochemically partitioned into distinct zones, relative to the:

base of complete oxidation (BOCO)
top of acidification (TOA)
top of fresh rock (TOFR).

Relationships are complicated by the deposit being originally mineralogically zoned and altered parallel to a sea floor, and then rotated into a sub-vertical position. The end product of secondary alteration processes is strongly influenced by the original metals and alteration distribution. The complexity of the impacts of weathering on mineralization and waste are depicted in Figure 7‑5.

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Figure 7‑5: Schematic Cross-Section of Mineralization and Alteration of the Bisha Deposit, and Characterization of the Main Weathering Domains

Source: After Cube Consulting 2013

From Surface to the top of fresh rock, the following sequence can be expected:

residual and transported soils and alluvials
near-surface mixed zone of gossan, ferruginous oxide, saprolite, and silica-clay sand
horizontal zone of acidification producing a white silica-clay rich band (no sulphides), locally known as soap
horizontal zone of acidification producing a black silica-lead rich band (sulphides appear)
horizontal zone of unconsolidated pyrite sand sitting atop Cu rich supergene
copper rich supergene
podiform transitional zone of Zn + Cu at the Zn-rich primary contact.

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The complex weathering profile has been established through the process of mining, with subsequent follow-up diamond and RC drilling, on-and-off site metallurgical testwork. Figure 7‑6 provides a much simpler, but more visual weathering profile than the all-encompassing schematic presented in Figure 7‑5.

Figure 7‑6: Simple Weathering Profile through Bisha Sulphides

Note: Average grades shown here are illustrative only and not representative of 2013 Resource and Reserve estimated values.

The various weathered facies of the Bisha VMS system are described in the following sub-sections.

Gossan

The Ferruginous to massive goethite–hematite–jarosite gossan was the remnant of surface oxidation of the massive sulphides. The gossan consisted of a large mound of red–brown oxide material ranging from fine sand to dense cobbles and boulders, distributed randomly or as groups or possible remnants of stratigraphic ‘horizons.’ The boulders and cobbles were usually extremely siliceous. The depth of oxidation was variable, being on the order of 30 m to 35 m in outcrop areas. The unit had a high gold content; the relatively low base metal values (copper, zinc) are due to leaching during oxidation. Banded, white, opaque, quartz veins ranging up to 0.5 m thick and several metres long occurred in some of the gossans.

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During mining, most of the mineralized stratigraphy was not outcropping. Strong saprolite and saprock occurred above the mineralization, representing weathered hangingwall volcanic facies. The interface between saprolite/saprock where mineralization was closest to surface was generally a gossan underlain soap, facies of the weathered Bisha VMS system.

Breccia

Flanking the gossan was a breccia unit, which appears to have been a product of oxidation, lateritic weathering, and desegregation of the original rock, as opposed to it being a structural feature. The unit was mostly quartz breccia or silicified fragments within oxidized material. The breccia occurred as an extremely high-gold grade unit within the Bisha open pit.

Oxide

The majority of the free dig within the Bisha pit was soft, brown, iron-rich oxide (limonite, goethite, and hematite). The oxide varied in thickness from 5 m to 30 m depending on location in the open pit, and carried grades in excess of 6 g/t Au with high associated silver. Being above the BOCO, there were no sulphides present within the saprolite, and no relict textures of its original precursor rock type. The oxide is distinct from the heavily altered saprolite and saprock that occurs above the VMS system in the hangingwall of the deposit.

Soap

The soap (acid leached) horizon is the result of the extremely acidic nature of the massive sulphide oxidation that caused development of a highly leached “front,” producing very friable remnants consisting of mostly clay and silica. The thickness of the soap horizon was variable, ranging from 0.5 m to 15 m, and averaging 3 m in thickness. The unit had variable gold and silver values, and was devoid of base metal mineralization.

Soap contacts are sharp, even when inter-fingering with underlying acid, or overlying oxide facies. Minor horizontal lenses of acid can occur within the soap, as can discolouration through presence of minor iron oxides. The overlying saprock, acid-leach zone (soap), acid zone, and pyrite sand relationship is shown in Figure 7‑7 and Figure 7‑8.

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Figure 7‑7: Weathering Sequence, Northern Bisha Pit, Beneath NE Hill Cutback

Notes:

From top down—ferruginous saprock (VMS hanging-wall waste);
soap (Fe mottled, white silica, minor clay, variable gold);
acid (silica–galena–pyrite–silver–gold); pyrite sand (crystalline, unconsolidated +90% pyrite).
White soap band (Soap 2) unusually splits acid into two units).

Acid Zone

The “acid” mineralization mined at Bisha Mine was exposed in March 2012, and represents a weathering horizon poorly defined during the Feasibility stages of the Bisha Project. The poor definition is attributable to the acid mineralization that corresponds with significant zones of poor drill-core recovery. The acid zone is considered an extension of the soap zone as the majority of the acid zone comprises silica and clay. The defining black colouration of the acid zone results from concentration of lead (galena) within this unit, starkly contrasting the acid from the white, overlying soap zone. The contact between acid and soap was sharp. Examples are shown in Figure 7‑7 and Figure 7‑8. The acid zone had high concentrations of lead, gold, and silver. The very high gold grades averaged 25 g/t, with associated silver values over 1,000 g/t. Acid ore had to be managed very carefully through the Bisha processing plant. The percentage of pyrite within the acid zone increased from top to bottom, becoming noticeably visible within the ore towards the base of the zone. Presence of high amounts of pyrite with associated high gold and silver grades was problematic with respect to processing through a CIL circuit. A large proportion of the acid ore was sold as Direct Shipping Ore (DSO).

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Figure 7‑8: Oxide-Soap-Acid-Pyrite Sand Sequence, Northern Bisha Open Pit

Note:

Base of oxide is marked by the hematitic band (red) overlying a thin limonite stained soap horizon (yellow-white) carrying stringers of acid (black). The acid facies is a 0.5 m thick continuous horizon with a very sharp contact with underlying pyrite sand.

Pyrite Sand

The pyrite sand horizon is lies directly atop the copper mineralized supergene zone. Pyrite sand was generally 0.5 m to 5 m thick, with mining in 2014 still encountering pyrite sands in areas where it is now over 10 m in thickness. The generally deeper sections of pyrite sand are in areas of deeper weathering such as along ore/waste boundaries and along major fault structures. The pyrite sand consists of fine, crystalline sand sized particles of unconsolidated pyrite. The contact with the acid zone is knife sharp as it is a compositional change from a silica dominated unit to a sulphide dominated unit. Gold grades within the pyrite sand were highly variable and it is considered that this relates to concentrations of ‘Acid’ stringers and laminae through the pyrite sand. On average, the pyrite sand averages 10 g/t au with low silver values. Towards the base of the pyrite sand, copper values increase in proximity to the supergene. The pyrite sand/supergene contact appears gradational rather than sharp. Examples of this material and contacts are shown in Figure 7‑8.

Supergene

Supergene mineralization is copper-enriched, and occurs between 35 m and 65 m below surface. As in the supergene enrichment of porphyry deposits, oxidation of massive sulphides caused the descending waters to become acidic, and leach copper and other metals. The metals were deposited generally as covellite, and some chalcocite, at the base of the acid and oxide domains. Sooty secondary sulphides coat and replace primary sulphides. An exposure of supergene material is illustrated in Figure 7‑9.

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Extensive work has been undertaken between 2012 and 2014 to increase the understanding of the main supergene zone above the Bisha primary sulphides. Work included additional metallurgy, density determination testwork, ore characterization, petrological analysis, and copper-arsenic distribution modelling. Through this work, there is an understanding that arsenic is distributed within pyrite, arsenopyrite, enargite, tennantite, and tetrahedrite (Johnston, H. and C. Chan, 2014). Ashley (2014) petrology report has also determined that a significant proportion of what was previously considered chalcocite is actually digenite, a much more reactive form of chalcocite.

Better understanding of arsenic distribution has identified that copper-arsenic minerals (such as enargite, tennantite, and tetrahedrite) need to have their concentrations and distributions carefully defined within the supergene zone. The presence of these copper-arsenic species has an impact on process recoveries of copper, and concentration of arsenic in concentrate. Understanding the geometallurgy of the deposit will allow for selective mining and blending of various ore types to produce optimal process recovery for copper and reduced arsenic in concentrate.

The presence of digenite as a copper mineral explains the high oxidation rates of supergene mineralization within the Bisha open pit. Digenite is a catalyst for rapid oxidation and resultant heat generation.

In the hangingwall copper zone, the mineralization is predominantly supergene copper mineralization consisting of chalcocite and covellite stringers in chloritic-altered rocks: chalcocite, covellite, and hematite in the soap rocks. Mineralized zones proximal to the massive sulphides appear to be quite erratically distributed. Supergene copper minerals tend to occur as a horizon gently dipping away from the massive sulphide lenses. The geometry of hangingwall copper mineralization is now considered strongly influenced by the presence of multiple 045° trending faults. The faults visibly offset mineralization in the northern extent of the Bisha open pit, and have potentially remobilized and acted as fluid conduits channeling groundwaters responsible for forming and distribution supergene copper. This relationship is shown in Figure 7‑9.

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Figure 7‑9: Copper Supergene, Bisha Open Pit (left) and Fault Controlled Copper Supergene, Northern Bisha Open Pit (right)


Note: Blocky massive pyrite, chalcocite-digenite, covellite rich supergene ore. Ore oxidation appears as light green malachite coatings on surfaces.	Note: Massive supergene having sharp fault contact with adjacent pale white bleached unmineralized saprock. Fault contact mapped as being one of a number of mineralization bounding 045° faults responsible for high grades at the northern end of Bisha pit and distribution of supergene into the hangingwall copper zone.

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Primary Sulphide

Primary sulphide mineralization occurs typically below a vertical depth of 60 m to 70 m. Sulphide minerals are predominantly pyrite, sphalerite, and lesser chalcopyrite. Sphalerite is the dominant economic mineral in this zone, and is now exposed at the southern end of the Bisha Main Zone deposit.

Ashley (2013), noted that pyrite in the massive sulphide could in occur in concentrations of up to 97%, and sphalerite in concentrations of up to 55%. Minor amounts of galena, arsenopyrite, pyrrhotite, tetrahedrite, and marcasite were also observed. In terms of paragenesis, Ashley (2013) considered pyrite to have been deposited prior to base metals and pyrrhotite, based on textural evidence.

Sulphide textures include semi-massive, massive, banded/laminated, minor folds, clasts, and disseminated sulphides within chloritized volcanics.

Reflected light microscopy on polished sections collected from massive sulphide were found to be composed of either pyrite-rich fragments embedded in a pyrite-sphalerite-trace galena, chalcopyrite, minor gangue (quartz, feldspar, sericite, and carbonate) matrix, or coarse-grained pyrite with interstitial chalcopyrite, trace sphalerite, and minor gangue (quartz, feldspar, sericite, and carbonate).

Bampton (2013) defined a transition zone between supergene copper and primary zinc sulphide mineralization. The transition is marked by the first appearance of sphalerite (ZnS) coexisting with secondary copper minerals covellite, chalcocite, bornite, enargite, tennantite and arsenopyrite. The transition zone is estimated to extend 10 m to 15 m in irregular pockets below the top of first sphalerite occurrence (termed the zinc depletion surface in Section 14), after which chalcopyrite is the main copper mineral in sphalerite and pyrite dominated massive sulphide.

Alteration

Chlorite, sericite, and silica alteration occur in the Bisha deposit hangingwall and footwall wallrocks. Hangingwall alteration is more intense than footwall alteration. This was modelled as illustrated in Figure 7‑10.

Structure

Folding has previously been interpreted as a key ore redistribution process post mineralization deposition (Barrie, 2004; Barrie, 2007). Pit mapping, grade control drilling and re-logging of historic core have shown previously interpreted “fold-limb” mineralization to be mineralization redistribution along and between structures. Faults and shears break the Bisha mineralization into separate domains, and are responsible for creating mineralization “spurs” striking to the southwest and northeast off the Bisha Main Zone.

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Figure 7‑10: Bisha Deposit Alteration Patterns

Note:

Bisha massive sulphides (red) and alteration patterns for A) chlorite, B) sericite and C) silica. Lighter shades indicate moderate, and darker shades indicate strong/intense alteration. Modelling is based on extracting alteration codes taken from drill logs, and using Leapfrog software to create the alteration distribution patterns.

The previously interpreted “West Limb” of Bisha supergene ore is associated with 020° and 045° faulting to give the apparent shape of an anticlinal fold at the northern limit of the Bisha orebody. The Bisha Main Zone is truncated by a clear, sharp 045° fault at its most northern extent. The current structural interpretation has come together as part of the Bisha 2013 Geological Model Project (Bampton, 2013). Synthesis of data to create the structural framework for the Bisha deposit outlined below was by Bampton (2013).

000^°North-South Shears

Two and potentially three north-south shear corridors (000°) are currently interpreted to truncate the Bisha deposit (Figure 7‑11). The shears occur as semi-continuous zones throughout the deposit. Contacts with massive sulphide are sometimes sharp, but more frequently there is some degree of sulphide brecciation, which defines the shear zone, often as a weak, clast-supported “crackle breccia.” The north-south shears represent the earliest of three different set of faults that have been defined. Displacement along these structures has not yet been determined.

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Figure 7‑11: 000° North-South Shear Zone Truncating Bisha Deposit

020° Shears

The footwall side of the primary and supergene massive mineralization in the north of the pit is often characterised by brecciation. The orientation of brecciation is consistent with a shear zone striking approximately 020°, with a dip of -65° to -75° to the west (Figure 7‑12). This broadly corresponds to the location and orientation of gold mineralization in the oxide phase away from the main structures and a hematite-altered shear zone visible in the northeast wall of the Bisha open pit. In the northern part of the Bisha deposit, 020° shear is interpreted to offset the 000° shear corridor, and thus may represent a later event. Displacement due to these structures has not yet been determined.

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Figure 7‑12: 020° Shear, Bisha Deposit

Note:

The 020° shear is interpreted as having offset the 000° shear corridor.

045° Sub-Vertical Faults

The northern margin of the orebody appears to be truncated by a 5 m to 7 m wide fault zone striking approximately 045° and dipping -85° to the southeast. Gold mineralization in the oxide phase follows this margin. Presence of a small pod of MSUL below the oxide gold mineralization may indicate this is a fault sliver of the main Bisha massive sulphide orebody, displaced approximately 200 m to the south-west (strike component). Dip component of displacement due to these structures has not yet been determined. These shears appear to truncate or offset the 000° structures, and are thus part of a later event, illustrated in Figure 7‑13.

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Figure 7‑13: 045° Sub-Vertical Shears, Bisha Deposit

Note:

The 045° faults are interpreted as having offset the 000° shear corridor.
Isolated massive sulphide lens in the north is interpreted to be a raft of primary mineralization offset
from the main deposit by 045° faulting.

The currently recognized major structural elements are interpreted as having displaced and/or truncated primary, massive sulphide mineralization. The same structural elements are also correlated with the occurrence, distribution, and enrichment of supergene and oxide mineralization (Figure 7‑14 and Figure 7‑15). The 000°, 020° and 045° structures have been preferred pathways for groundwater and potentially late stage hydrothermal fluids that have mobilized copper, gold, and silver out of the massive sulphide Bisha VMS deposit, concentrating these metals within the weathering profile of the deposit.

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Figure 7‑14: Supergene Copper Distribution Relative to Structures and Underlying Primary Mineralization

Note:

The supergene copper is best developed directly above richest zones of zinc massive sulphide within primary mineralization.
There is no supergene copper development where 020° shear cuts massive sulphides and offsets 000° shear zone breaking the Bisha deposit into two structurally separate domains.

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Figure 7‑15: Oxide Gold Mineralization Distribution Relative to Faulting and Shearing

Note:

Gold oxide mineralization (yellow) is best developed and has the highest grade proximal to major structures.
At the northern extent of the Bisha open pit, the 045° fault truncates the Bisha deposit and has both oxide (yellow) and Hangingwall supergene copper (green) developing along the strike of the fault structures.

Stratigraphy

The Bisha deposit is hosted within rhyolite, rhyolitic breccias, mudstone, and mafic volcanic rocks. McPhie (2013) proposes that Bisha massive sulphide deposition was sub-seafloor into coherent rhyolite. Fragments of rhyolite can be found within massive sulphide.

Rhyolite/mineralization contacts can be sharp to gradational. The rhyolite is moderately to strongly foliated, with occasional zones of quartz and feldspar phenocrysts, perlite, and apparent clastic textures. A coarse clast-supported rhyolitic breccia overlies and flanks coherent rhyolite. Breccia clasts are monomict, and comprise coarse perlitic rhyolite clasts.

A mudstone occurs amidst coherent and brecciated rhyolite facies. The mudstone is thinly bedded and planar. Underlying the mudstone is a polymict breccia-conglomerate that is internally massive, composed of pale and dark aphanitic clasts, and moderately to strongly foliated.

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Logging reports that mafic volcanics occur in the hangingwall and footwall of the deposit. Massawe (2012) reported mafic volcanics in the footwall of Bisha.

Although McPhie (2013) proposes that the Bisha deposit was emplaced sub-seafloor, petrographic work by Ashley (2013) suggests that there are mineralogies indicative of exhalative deposition, and textures that could be interpreted as representing seafloor “black smokers,” and thus a polygenetic origin can be envisaged. McPhie’s schematic model is shown in Figure 7‑16.

Figure 7‑16: Bisha Deposition Model after McPhie (2013)

Note:

Simplified cartoon. Massive Bisha sulphides are enclosed within altered rhyolites and coherent rhyolite on the outer. Breccia facies flank and separate stacked coherent rhyolite facies. McPhie (2013) advocates a sub-seafloor emplacement model for the Bisha deposit

7.3.2 Harena Deposit

The Harena deposit has been traced over a strike length of 400 m, and is interpreted to be a northwest-dipping, tabular, massive sulphide body, closed off by drilling to the northeast, but open to the southwest.

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The Harena host rocks deposit are a bimodal, hydrothermally altered suite of basalts and rhyolite dacite volcanics. The stratigraphic succession has appreciable siliciclastic rocks up-section to the west, and is dominated by rhyolite and dacite tuffs proximal to the deposit, with minor intercalated basaltic rocks. The deposit has a distinct footwall that contains kyanite and andalusite. Both minerals are often noticeably chloritized. The kyanite and andalusite are interpreted as having been formed following metamorphism of aluminium-bearing seafloor sediments. There is a graphitic component to some of the massive sulphide intersections. A number of late dykes have cut the near surface mineralization at Harena, making determinations of the actual widths of the zone difficult.

Surficial weathering processes have produced three distinct zones of mineralization. These include a surface oxide/gossan overlaying a very thin secondary supergene horizon, which grades into a primary massive sulphide horizon at depth.

The primary massive sulphides are predominantly made-up of fine- to medium-grained subhedral to anhedral pyrite, with interstitial and/or enriched layers of sphalerite and chalcopyrite.

The gossanous horizon contains frequent anomalous levels of gold and silver. The depth of oxidation appears to be approximately 45 m to 50 m. Both the oxide and sulphide mineralized zones are approximately 400 m long, and vary in thickness between 5 m and 15 m.

In the opinion of the QP, the deposit settings, lithologies, and structural and alteration controls on mineralization at Harena are well understood, and the geological understanding at Harena is sufficient to support Mineral Resource and Mineral Reserve estimation.

7.3.3 Northwest Deposit

General

The Northwest deposit comprises a series of poly-metallic massive sulphide lodes that have been defined over a strike length of 800 m (Figure 7‑17), striking northeast, and dipping from 70° northwest to subvertical. The deposit is thickest at the centre, tapering to widths of less than 8 m at its strike limits. The central portion of the deposit is over 85 m wide. In cross-section, the deposit is a wedge that narrows down-dip. Resource drilling has effectively defined the deposit to a maximum of 250 m below surface. Exploration drilling indicates that a mineralized stringer vein system still exists at depths of 350 m below surface.

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Figure 7‑17: Northwest Deposit Project Area, Bisha Mine, Eritrea

The Northwest deposit can be subdivided into three domains. The northern Main Lode is the largest domain, being a copper-dominated massive- and semi-massive sulphide lode (also referred to in this report as the Central Zone) that increases in zinc content northwards. The Southern Lode is a zinc-dominated, discontinuous massive- and semi-massive sulphide body. As with the Bisha deposit, pyrite is the main sulphide component of the VMS deposit.

A separate narrow VMS lode has been defined on the footwall side of the Northwest deposit (Eastern Lode). The Eastern Lode is poorly drill defined, but for its width carries appreciable copper and zinc base metal grades. The Eastern Lode is at the earliest stages of being defined and understood.

As with the Bisha deposit, Northwest has a gold oxide cap and supergene copper zone. These parts of the stratigraphy are not as well endowed as Bisha in gold, silver, and copper, reflecting the overall lower grade of the underlying Northwest primary mineralization. The oxide profile is still problematic due to core recovery problems during drilling, a problem also encountered across the Bisha deposit. At this stage, there is still significant potential for resource upgrade in the oxide and supergene portions of the Northwest deposit through infill drilling of areas of poorest core recovery.

The majority of information presented is as a result of extensive resource development drilling and testwork from 2011 to mid-2013. Completed works include geotechnical, hydrogeological, metallurgical, and waste-rock characterization drilling and testwork.

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Figure 7‑18: Northwest Mineralization and Structures; Plan View, Local Grid N at 045^o

Source: Shepherd, 2013

Note:

Mineralization is strongly influenced by shearing and faulting, which has split the deposit into Northern and Southern Lodes. A separate footwall Eastern Lode has been identified near surface, but is poorly defined.

Sulphide Mineralization Occurrence

General

Three mineralization styles were identified from logging of lithology and visual percentage of pyrite—massive sulphide, semi-massive sulphide, and stringer mineralization (Figure 7‑19). The mineralization styles have specific element, element-ratio, and physical-property distributions that could be validated, i.e., arsenic, barium, zinc, iron, density, and applying the Ishikawa alteration index (Shepherd, 2013). Application of secondary geochemical methods allowed areas of logging ambiguity and inconsistently to be quickly resolved, and robust mineralization domains to be modelled.

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Massive Sulphides

Massive sulphide zones are here defined as areas where the proportion of logged visual sulphides is greater than 50% (majority pyrite). The Main, Southern and Eastern Lodes each have massive sulphides at their core, with all other mineralization styles peripheral to the massive sulphides.

The Main Lode can be sub-domained across a major shear zone into a copper-dominated southern zone and a northern zinc-dominated zone. The Main Lode contains the largest body of Northwest sulphides varying in thickness from 8 m at its northern extent to 85 m in the central, thickest portion of the Lode. The Main Lode has a steep westerly dip of approximately 75°.

The Southern Lode is physically separated from the Northern Lode by an interpreted shear zone. The Southern Lode is narrow (5 m to 25 m wide), steeply dipping (-85° westward), and discontinuous over its strike extent. Zinc sulphides dominate the Southern Lode.

The Eastern Lode lays parallel to and approximately 140 m to grid east, of the Main Lode. The Eastern Lode has a strike extent of 600 m and dips steeply to the west at approximately 85°. The lode is only 3 m to 6 m wide and currently defined by six shallow drill holes within 100 m of surface.

Massive sulphide contacts are generally sharp, but can grade over several metres into more diffuse semi-massive and stringer mineralization.

Semi-Massive Sulphide

Semi-massive sulphide zones are defined as areas where the percentage of logged visual sulphides is 25% to 50% (majority pyrite). Semi-massive sulphides are a gradational extension to massive sulphides, with less intense pyrite development. Semi-massive sulphides are generally indicative of a more distal edge to the main sulphide body, displaying interlayering and banding with the host rock, and less intense footwall stringer sulphides.

A large semi-massive sulphide zone occurs about the Southern Lode, as massive sulphides grade into sulphide stringer alteration at the southern distal strike and dip extent of the Northwest deposit.

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Figure 7‑19: Mineralization Styles, Northwest Deposit

Note:

Massive sulphides (pink), semi-massive sulphides (green) and stringer zone (yellow) defining the Northwest VMS deposit. Plan sliced at 480 mRL, approximately 80 m below surface.

Stringer Sulphides

Stringer sulphide zones represent a more diffuse zone of logged sulphides (5% to 25% visible sulphides) of which pyrite is the main component. Stringer sulphides define the footwall zones of massive and semi-massive sulphide mineralization. The stringers are typically within a chlorite sericite alteration zone, and have a distinctive, finely banded, and anastomosing nature. The stringer sulphides are highly variable in geometry, discontinuous, and represent the footwall feeder and alteration zone to overlying more massive sulphides.

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The Northwest stringer sulphide zone has been interpreted as a single large wedge shaped zone that totally encapsulates the massive and semi-massive sulphides, and represents the limits of the mineralising and alteration system. The north-south alteration zone extends for a strike length of 1 km, and currently, has not been closed off by drilling to the north. The zone generally has a sub-vertical, sharp hangingwall contact and a more diffuse and moderately west-dipping (approx. 85°) footwall contact.

Oxide Mineralization Occurrence

Oxide Zone

The Northwest deposit has an oxide zone above the BOCO surface. Within this zone, original rock types have been weathered to the extent that all sulphides are oxidized with no primary or secondary sulphides present, and the original colour of the fresh rock is unrecognizable.

Low-grade gold has been defined within the oxide zone, reporting grades of 0.5 to 1.5 g/t Au. The gold zone occurs above the supergene gold zone, and is developed as a sub-horizontal near surface band (5 m to 10 m thick) within 20 m of the surface. This zone is typically highly weathered and ferruginous, and lies above the BOCO surface.

Soap Zone

Soap is a local term that refers to a zone of intense acid leaching, predominantly above and adjacent to massive sulphide bodies. Leaching destroys the primary texture of the original rock, leaving a clay silica residue which when wet has a greasy to ‘soapy’ nature. Soap is defined as all material below the BOCO surface but above the base of the soap surface. Decomposed sulphides can be present toward the base of this unit, as disseminated chalcocite and covellite.

A supergene gold domain has developed within the soap zone directly above the massive sulphide, mainly within and straddling the acid leach interface (soap surface). The gold mineralization appears to be unrelated to faulting. Higher gold grades are developed over the thickest portions of the central massive sulphide within soap material.

The average core recovery within the soap zone is 35%, with discontinuous and restricted zones of high-grade gold. As a result of the low sample recoveries, the confidence level when estimating gold within the soap zone is low.

Supergene Zone

The Northwest copper supergene zone includes all material lying above the TOFR and below the soap zone surface. Within the supergene zone the colour and texture of the fresh rock is recognizable, but partial weathering and discolouration of the rock material has occurred. The supergene zone is the transition between oxide and sulphide minerals, and contains both primary (chalcopyrite) and secondary copper minerals (chalcocite/covellite).

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Sulphide Mineralogy

The Northwest sulphide mineralogy is dominated by pyrite (up to 99%) with minor amounts of sphalerite, chalcopyrite, galena, and pyrrhotite. Trace arsenopyrite and possible bismuth telluride were also recorded during petrographic analysis (Ashley, 2014b).

Mineralogical assessments by QEMSCAN/BMA (Johnston, and Sloan, 2013), defined covellite, chalcocite, bornite, enargite, and tennantite within the supergene zone. Chalcocite was the main copper contributor to the supergene.

Chalcopyrite was the main copper contributor to primary sulphides, whereas sphalerite made-up over 95% of the primary zinc mineralogy in parts of the orebody dominated by zinc (Southern Lode and northern extent of Main Lode).

Lithology

The Northwest deposit is within a sequence of rhyolite, rhyolitic and polymictic breccias, mudstone, and metasandstone (Figure 7-20). The host sequence to the mineralization is an altered glassy rhyolite. The rhyolite is fine grained, foliated, and has zones with feldspar and quartz phenocrysts. The rhyolite also displays relict perlite and clastic textures.

Monomict rhyolitic breccias within the volcano-sedimentary sequence have sharp clast boundaries. The rhyolitic breccia is inferred to be breccia that has formed above or to the side of rhyolitic flows.

A number of thinly bedded mudstone horizons with occasional clasts have been logged on the hangingwall side of mineralization. Also distal to mineralization on the hangingwall side of the Northwest deposit are weakly bedded to massive, fine- to medium-grained metasediments (siltstone/ sandstone).

McPhie (2013) considers the Northwest deposit host stratigraphy to be analogous to the volcano-sedimentary sequence hosting the Bisha deposit. As with Bisha, the association of coherent and monomictic rhyolite breccia at the Northwest deposit is typical of submarine felsic lavas and/or domes generated by effusive eruptions. In this situation, sulphides occur entirely within rhyolitic units, with hydrothermal alteration affecting hangingwall and footwall host volcanics, proximal breccias, and metasediments. The mechanism of emplacement is proposed to be sub-seafloor replacement of glassy, fractured rhyolitic host facies (Figure 7‑21).

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Figure 7‑20: Rhyolite and Mudstone Distribution, Northwest Deposit, Plan View

Note: Association of major logged and interpreted lithological units with massive sulphide mineralization.

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Figure 7‑21: Schematic Section: Interpreted Volcano-Sedimentary Facies– Architecture for the Stratigraphy at Northwest

Souce: McPhie, 2013. Not to scale.

Note: Faults, dykes and minor facies are not shown

Structure

A number of structural zones were interpreted from the trends established by surface mapping and trenching, with targeted core re-logging and core photo examination used to confirm the structural interpretation. These zones are highlighted by offsets, rapid changes in orientation and terminations of the massive sulphide body, particularly to the south where the South and Main Lodes are separated by an area of discontinuous and irregular sulphide lenses. The identified faulting has preferred orientations of northwest-southeast to north-northwest-south-southwest, with a dominant dextral strike slip movement inferred.

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Two main fault zones have been interpreted that significantly disrupt and displace the massive sulphide body (Shepherd, 2013):

Zbei Fault Zone – a 20 m to 50 m wide north-northwest-south-southwest trending zone of multiple fault planes that separates the southern northwest-trending massive sulphides from the north-south trending Main Lode. Erratic pods of massive sulphide with variable orientations lie in this zone, representing “sliced and diced” portions of sulphide mineralization.
Hbei Fault Zone – a northwest-southeast trending zone of multiple faults that have caused a rapid thickness and orientation (dip) of the central massive sulphide to the north. A vertical displacement is evidenced by rapid change in the depth extent of the central and northern massive sulphide across the structure.

Other faults (Figure 7‑22) defined by Shepherd (2013) include the:

Secondary Fault Zone – striking 80° toward 105°
Footwall Fault – dipping at between 80° and 85° toward 315°
Arab Fault – striking 043°, running parallel to mineralization
020° Faults – with strike slip sense of movement, and perhaps a dip slip component.

The foliation across volcanics and metasediments at Northwest is steeply dipping to either the northwest or southeast, and at times tightly crenulated. The foliation is also pervasive within mineralized rock, but is less apparent in the felsic host rock (Kinakin, 2013).

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Figure 7‑22: Interpreted Structural and Massive Sulphide Domains, Northwest Deposit – Plan View

Source: Shepherd, 2013

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Intrusive Dykes

Three main dyke types have been logged and interpreted:

	1)	North-south trending mafic to intermediate dyke, with thicknesses ranging from 1 m to 5 m
	2)	Moderately west-dipping intermediate dykes within the footwall alteration zone, with average widths of 1 m to 5 m
	3)	Cross-cutting, sub-vertical felsic quartz-feldspar dykes trending west-northwest up to 10 m in true thickness, and offsetting mineralized zones.

These dykes indicate zones of faulting and structural discontinuity. The dominant mafic hanging-wall dyke swarm marks the stratigraphic top of the mineralising system, and has developed within a shear zone. This sub-vertical, 5 m to 10 m wide structural zone is continuous along the western edge of the massive sulphide body for most of the strike length of the Northwest mineralization.

Alteration

A weak to strong alteration halo of chlorite, sericite, and silica occurs in the hangingwall and the footwall of the Northwest deposit (Figure 7‑23).

Chlorite appears to have a rough zonation, with the strongest alteration close to mineralization. Sericite alteration is proximal to mineralization. Silica alteration is more distal to mineralization than chlorite and sericite, and more broadly dispersed through footwall and hangingwall lithologies.

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Figure 7‑23: Northwest Deposit and Associated Alteration – Plan View

Note:

Massive sulphides (beige) with alteration patterns for A) chlorite, B) sericite, and C) silica.
Lighter shades indicate moderate alteration, and darker shading strong-to-intense alteration.
Data taken from drill hole logs and visualised using Leapfrog software.

Overburden

Transported sheetwash and alluvial material blankets the Northwest deposit in areas of lower topography between weathered outcrops. The overburden consists of poorly consolidated brown clays, quartz, ferruginous gravel, and pebbles. The overburden may extend up to 2 m to 3 m below surface. Trenching across the deposit was fortunate to reach to the base of overburden at the maximum reach of the excavator in use (3 m).

Weathering

The weathering profile at Northwest extends to between 50 m and 80 m below surface (Figure 7‑24). The weathering profile transitions from a 2 m to 3 m transported cover to weakly developed gossan directly over mineralization, ferruginous saprolite or saprock, depending on proximity of mineralization to surface. Mineralization is not equidistant to surface along the length of the Northwest deposit. The weathering profile is much deeper towards the south of the deposit, with no development at or near surface gossan.

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Figure 7‑24: Northwest Deposit – Massive Sulphide, Geology, and Weathering Surfaces. Cross-Section 20550 mN

As with the Bisha deposit, the weathering profile above mineralization progresses through an oxide zone into the soap zone (inclusive of acid horizon), pyrite sand, then into a weakly developed copper supergene zone. Not all units are continuous and may be absent depending on location relative to mineralization.

On the hangingwall of mineralization, the weathering profile through waste is typically saprolite transitioning to saprock and then fresh rock. Distal to mineralization the weathering profile may be restricted to a very limited zone of saprock development.

Metamorphism

The stratigraphy and mineralization in the region of the Northwest deposit is considered to have undergone greenschist facies metamorphism. Regional metamorphism is characterized by chlorite‑muscovite alteration, with rare garnets occurring in sheared or strongly foliated zones.

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7.3.4 Hambok Deposit

General

The Hambok deposit lies within the Western Nakfa terrane, and is part of a sequence of Late Proterozoic mafic to felsic volcanic rocks, and sedimentary rocks, including pelites, chert, and carbonate units. The deposit comprises a primary copper/zinc sulphide zone, representing the majority of the deposit, and a minor oxide gold component. The primary massive sulphide mineralization is a single body, with a faulted displacement interpreted at depth in the northeast of the deposit. The massive sulphide zones strike at approximately 015°, dipping steeply to the east, with overall strike and dip lengths of some 975 m and 400 m, respectively. The thickness of the massive sulphide varies from about 5 m to 75 m. A general view of the deposit is given in Figure 7‑25 and overview of the deposit, showing drilling, is shown in Figure 7‑26. Additional illustrations for Hambok can be found in Section 14.

Figure 7‑25: Hambok: Looking South towards Deposit from Gossan Hill

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Figure 7‑26: Hambok Deposit, Drill Coverage

Setting

The hangingwall rocks of the Hambok deposit consist of the following principal facies (McPhie, 2013):

chlorite-epidote-altered massive basalt
chlorite-epidote-altered pillow basalt (Figure 7‑27)
chlorite-magnetite-altered fluidal-clast basalt breccia, polymictic basalt breccias
strongly altered coherent rhyolite, and strongly foliated felsic metavolcanics.

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Figure 7‑27: Chlorite-Epidote-Altered Pillow Basalt

These facies have been interpreted as follows:

The presence of amygdales and fine grain size indicate that the chlorite-epidote-altered massive basalt comprises lavas dominated by coherent facies.
The local presence of pillow basalt constrains the setting to being subaqueous. Given the association with massive sulphide deposits, the subaqueous setting is likely to have been submarine.
The chlorite-magnetite-altered fluidal-clast basalt breccia is a submarine fountain deposit, generated by injection of a jet of molten lava into the water column. The molten basalt broke into large (>5 cm) droplets that developed quenched rims and vesicular cores, while the smaller droplets quench-fragmented to produce angular basalt clasts in the matrix. This facies forms at the seafloor and is commonly associated with both massive basalt lavas and with pillow basalt, as observed in core. Fountaining occurs in response to surges in discharge rate that may be very brief or sustained. Fountain height is typically only tens of metres above the seafloor, so this facies is usually taken as a good indicator of proximity to basalt vents (i.e., within tens to hundreds of metres). The basalt succession is very thick (approx. 370 m) and no other facies are interleaved, implying proximity to source vents and relatively rapid accumulation.
The green, quartz-phyric rhyolite intervals are interpreted to be post-sulphide dykes because they have finer grained margins at contacts with other facies, and although altered, they do not contain sulphides.
A dyke‐ or sill‐like granitoid body, 300 m to 600 m thick and over 30 km long, cuts the volcanic package only a few tens of metres east of Hambok. It trends north-northeast, parallel to the regional fold axes. Compositionally, this dyke or sill consists of equigranular biotite granite, but local biotite‐hornblende syenite and granodiorite phases are present. Although the outcropping margins of the dyke or sill commonly have tectonic overprinting and cataclastic textures, drill intersections at Hambok and the Aderat prospect indicate that the granite is intrusive. It is marked by a strong, negative residual gravity anomaly with steep gradient against the mafic volcanic section.

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Alteration

At Hambok, chloritization of both hanging wall and footwall rocks is evident based on observation and a limited number of whole rock analyses. The presence of the both hangingwall and footwall chloritic rocks that bound the massive sulphide mineralization preclude the use of chlorite alteration as a tool for determining stratigraphic facing directions.

Magnetite commonly accompanies chlorite in the alteration envelope around the massive sulphide deposit, and is also present and locally abundant in the stratigraphic hanging wall. It occurs as fine disseminations, as heavy disseminations, and as wispy layers on a millimetre to centimetre scale, intercalated with the volcanics and commonly found in the massive sulphide itself. The thicker accumulations have 10% to 30% magnetite over 10 cm to 30 cm intervals.

Sericite occurs as very pale, greenish‐cream white, fine‐grained aggregates interstitial to pyrite in the Hambok massive sulphide, and is a common constituent of the massive sulphide mineralization. The colour is close to that of sphalerite. The hanging wall mafic and felsic volcanic rocks are variably sericite‐altered and this is evidenced by pale, greenish‐cream white replacement of feldspar phenocrysts in felsic volcanic rocks and the groundmass in mafic rocks.

Mineralization

The deposit is mostly covered by alluvium, with a small gossan at the northern end as the only surface expression. Without the extensive exposure from mining of the Bisha deposit, the oxide zone is much less well defined, mostly intersected in RC drill holes. The current interpretation is of a narrow zone less than 5 m wide with low values of precious metals, with oxide weathering present to some 50 m below surface. Although there is minor evidence for the development of pyrite sand from the RC drilling no coherent units have been interpreted. A supergene zone, such as developed at Bisha, does not appear to be present.

The primary massive sulphide mineralization is effectively a single body, with a faulted displacement interpreted at depth in the northeast of the deposit. The massive sulphide zones strike at approximately 015°, dipping steeply to the east, with overall strike and dip lengths of about 975 m and 400 m, respectively. The thickness of the massive sulphide varies from about 5 m to 75 m and the contacts are frequently seen to be sheared or faulted.

The massive sulphides can be separated into a number of intervals ranging from approx. 6 m to ~17 m in thickness. These sulphide intervals are separated by small intervals of chlorite-epidote altered massive basalt and/or green, quartz-phyric rhyolite(?). Contacts between the sulphide and the chlorite-epidote-altered massive basalt are typically gradational over a few centimetres although sharp contacts have been observed. The chlorite-epidote-altered massive basalt intervals also locally include irregular blobs and bands of sulphides. In contrast, contacts between the green, quartz-phyric rhyolite(?) and the sulphides are consistently sharp. In addition, contacts between the green, quartz-phyric rhyolite(?) and the chlorite-epidote-altered massive basalt are also consistently sharp. These separating units have not yet been successfully correlated between drill sections. Development of semi-massive and stringer sulphides are much more limited than those seen at the Bisha and Northwest deposits.

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Observation suggests that the chlorite-epidote-altered massive basalt has a similar style and intensity of alteration both above and below the sulphide-mineralization; epidote and chlorite are both conspicuous above and below, but in between, epidote is either minor or absent. Sulphides occur within the amygdales and also disseminated in the basalt. These relationships suggest that the massive sulphide has formed by replacement of the chlorite-epidote-altered massive basalt at multiple levels as illustrated in Figure 7‑28. The chlorite-epidote altered massive basalt does not show any textural changes close to contacts with the massive sulphide intervals, and the thicknesses of the basalt intervals between the sulphide intervals vary widely (<1 m to approx. 5 m). No other facies (e.g., mudstone) is interleaved with the basalt. Therefore, the sulphide intervals do not appear to be occurring at boundaries between basalt lava units. It was concluded therefore, that the full thickness of basalt succession (>100 m) was in place when replacement by sulphides began. It is also clear that the green, quartz-phyric rhyolite(?) occurs as dykes that post-date both the basalt host succession and the sulphides.

Figure 7‑28: Simplified Relationship between Facies (Schematic Section)

The massive sulphide mineralization shows a number of different textures and forms (as described above), but these zones cannot be readily correlated between drill sections or drill holes. It is, suggested however, that the observed grade distribution might be related to the different zones. Pyrite predominates and can be massive, fine or coarse grained. Economic mineralization comprises disseminated chalcopyrite and sphalerite; in some parts of the massive sulphides, these minerals can clearly be seen as separate centimetric bands. This has been validated by hand-held XRF work that shows the zones to be copper-rich-zinc poor and vice versa.

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A metallurgical characterization study completed in 2013 demonstrated that good recoveries and liberation could be achieved, and the banding of mineralization might be a factor. Recoveries of copper were 88% for a 25% concentrate, for zinc, 87% for a 57% concentrate. The study also showed that the sulphide mineralization is low in deleterious minerals such as arsenic. A small percentage of barite was present in the samples submitted, but very little galena.

The economic minerals are largely distributed in the upper part of the massive sulphide unit, i.e., that nearest surface, with three zones identified: hangingwall, footwall, and central. The hangingwall and footwall zones generally continue down dip further than the central zone, but with increasing dip, the mineralization is dominated by pyrite. This zonation is illustrated in Figure 7‑29.

Figure 7‑29: Typical Section of Mineralization Zones

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8 Deposit Types

The Bisha Main Zone is a large precious- and base metal-rich VMS deposit. The Northwest and Harena deposits are small to medium size VMS deposits. The Hambok deposit located on the Mogoraib EL is the second largest base-metal VMS deposit in the BMSC licence areas.

Previous work broadly determining the VMS deposit type and depositional setting for the Bisha, Northwest, Harena, and Hambok deposits (Barrie et. al., 2007), has been expanded by recent work by McPhie (2013) and Shepherd (2013).

8.1 Deposit Classifications

VMS deposits have a range of emplacement styles. Figure 8‑1 show four styles of VMS deposits, massive sulphides are shown for seafloor deposits (A) and deposits dominated by infiltration and replacement (B, C, and D). Graphic modified from Doyle and Allen (2003).

Figure 8‑1: VMS Deposit Classifications


A. Seafloor mound		B. Autoclastic breccias

C. Replacement below lava sill		D. Replacement within lava / intrusion

Irrespective of emplacement style, VMS deposits share the following common features, as outlined and illustrated by Gifkins et al. (2005). They are:

Hosted by submarine volcanic or volcano-sedimentary facies
The same age as the host sequence

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Host rocks vary from coherent to clastic volcanic or sedimentary facies, and range from basalt, andesite, dacite, through to rhyolite in composition
Composed of massive sulphide, mainly pyrite, with lesser amounts of sphalerite, chalcopyrite, and galena
Commonly aligned to strata, but not always
Stringer sulphides commonly underlie the massive sulphide and may contain economic copper. Note: Bisha deposits are exceptions, having stringer sulphides extending above and below massive sulphide
Ore metals are commonly zoned from Cu at the stratigraphic base to Zn, Pb, Ag, Au, and Ba, in general order to the top. Note, there are many variations to this (Figure 8‑2)
Intense hydrothermal alteration of the footwall volcanic rocks stratigraphically below the massive sulphide to chlorite, sericite, and quartz is common. By comparison, hanging wall rocks are weakly altered or can be unaltered. Note: the Bisha deposits are exceptions, having strong alteration above and below.

Figure 8‑2: Cross-Section of Idealized Mineralization (top) and Alteration (bottom) Zonation Patterns in a Footwall Pipe beneath a Typical VHMS Deposit

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Note: Modified from Gifkins et al., 2005

8.2 Previous Interpretation

8.2.1 Bimodal Siliciclastic VMS Deposit Model

Tucker Barrie visited the Bisha Project and concluded that a bimodal siliciclastic model (Barrie, C.T. et al., 1999) was most appropriate for the Bisha deposit (Barrie, 2004). Bimodal siliciclastic deposits form in lithological sequences composed of roughly equal proportions of volcanic and siliciclastic rocks. Based on field mapping Greig (2004) indicated that the host rock at Bisha was principally felsic volcanic rock (variably altered felsic lapilli, lapilli ash tuffs, crystal tuffs, and minor felsic dykes). Felsic volcanic rocks at Bisha were identified as more abundant than mafic rocks and calc-alkalic in composition, while mafic rocks are of tholeiitic composition.

Barrie (2004) considered the Bisha Main Zone deposit to be similar to those of the Iberian Pyrite Belt, and developed a VMS model for the Bisha Main Zone deposit, as shown in Figure 8‑3. The model incorporated local features such as the Bisha Gabbroic Complex and dominantly felsic and siliciclastic host rocks.

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Figure 8‑3: Bisha Bimodal Siliciclastic VMS Model Schematic

Source: Barrie, 2004

8.3 Recent Advances

Professor Jocelyn McPhie of the University of Tasmania, Australia, was contracted in September 2013 to complete a review of volcanic facies hosting the Bisha, Northwest, and Hambok Deposits. The objective was to define principal facies, stratigraphy, possible correlations, facies architecture, and relationships to sulphide mineralization for each of the deposits. McPhie (2013) submitted a detailed internal report summarizing these outcomes.

A significant outcome of the volcanic facies review by McPhie was the realization that the Bisha VMS deposits represent sub-seafloor replacive and impregnative styles of bimodal-siliciclastic VMS deposits, closely associated with rhyolitic and basaltic lavas.

8.3.1 Bisha and Northwest Deposits

Where previous workers were reliant on geochemistry and logging from heavily altered diamond core and field exposures, McPhie (2013) established from fresh open pit and drill core exposures that the host successions for the Bisha and Northwest Deposits mainly comprised coherent rhyolite and rhyolite breccia units, interleaved with mudstone and polymictic breccia. The rhyolite association is typical of thick submarine lavas or domes.

The work by McPhie (2013) strongly suggests that the mineralization was formed by sub-seafloor replacement and impregnation rather than sulphide deposition at the seafloor. This is supported by evidence of: the footwall being strongly to intensely altered, with the hanging wall also being strongly altered; massive sulphide intervals having gradational contacts with stringer sulphides both above and below, and; massive sulphide appearing to occur within thick rhyolite units that do not correspond to seafloor positions.

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8.3.2 Hambok Deposit

The Hambok massive sulphides are hosted by different lithologies than those at the Bisha and Northwest deposits. The host succession to Hambok massive sulphides comprises submarine, amygdaloidal basaltic lavas, including pillow lava and fluidal clast breccia that collectively suggests a relatively near volcanic vent setting (McPhie, 2013). The massive sulphide at Hambok is interpreted to have formed by sub-seafloor replacement of the host basalt succession.

Chlorite-epidote-altered basalts at Hambok display:

no textural changes close to contacts with the massive sulphide
widely varying thicknesses of basalt intervals between massive sulphide lenses, from less than 1 m to approximately 5 m
no other lithological facies within the basalt, e.g., mudstone
no obvious association of massive sulphide lenses with basalt lava flow contacts; massive sulphide lenses occur away from basalt lava flow contacts.

8.3.3 Harena Deposit

The Harena deposit was not reviewed by McPhie (2013), but is also considered by BMSC Geologists to be replacive from interpretation of data collected from drill hole logging. Harena has strong alteration and stringer mineralization on hangingwall and footwall sides of the massive sulphides.

8.3.4 Seafloor Accumulation vs. Sub-seafloor Replacement

Using criteria defined by Doyle and Allen (2004), physical features evident in a VMS deposit can be used as diagnostic indicators as to whether a massive sulphide deposit has formed in a seafloor or sub-seafloor environment (Table 8‑1).

Table 8‑2 applies the criteria of Doyle and Allen (2004) to the Bisha, Hambok, Harena, and Northwest deposits using features noted by McPhie (2013), Ashley (2013), and BMSC Geologists from open pit mapping and drill core logging.

Table 8‑2 highlights that Bisha, Hambok, and Northwest all have sub-seafloor diagnostic criteria 1, 2, 3, and 5. With respect to discordance between ore and host rocks (criterion 4), this has been difficult to establish due to strong alteration, and requires further work to accurately determine relationships.

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Table 8‑1: Diagnostic Criteria for Sub-Seafloor Replacement and Seafloor Accumulation

Criteria	Sub-Seafloor Replacement in Massive Sulphides
1*	Relics of the host rock within the sulphide deposit
2*	Facies displaying very rapid emplacement of the host lithofacies
3*	Replacement fronts between the sulphide deposit and host deposit
4	Discordance with the enclosing host lithofacies
5	Strong hangingwall alteration, similar in style and intensity to footwall alteration
Criteria	Seafloor VMS Accumulation
1**	Sedimentary clastic sulphide textures
2**	Sulphide chimney textures
3**	Exhalite at the ore horizon
4**	Fossil tube worms and bivalves
5	Facies indicating slow accumulation rate of the host rocks and/or occurrence of VMS ores between units of rapid emplacement rate
6	Asymmetric alteration pattern – strong footwall and weaker hangingwall

Notes:

Text modified from Doyle and Allen, 2004
* Diagnostic with 4-5 may suggest replacement
** Diagnostic and 5-6 suggest seafloor deposition

Table 8‑2: Bisha Deposits vs. Sub-seafloor Replacement Criteria Comparison*

Deposit	Style	Immediate Host Lithofacies	(1): Host Facies within Mineral Deposit	(2): Rapidly Emplaced Host Facies	(3): Replacement Fronts	(4): Discordance between Ore and Host Rocks	(5): Hangingwall Alteration	Interpreted Environment
Bisha	lenses	rhyolite	Y	Y	Y	+/-	strong	sub-seafloor with some contrary textures⁽¹⁾
Hambok	lenses	basalt	Y	Y	Y	+/-	no difference/strong	sub-seafloor
Harena	lens	rhyolite	+/-	Y	Y	+/-	strong	sub-seafloor (?)
Northwest	lenses	rhyolite	Y	Y	Y	+/-	strong	Sub-seafloor

Notes:

*After Doyle and Allen, 2003
Bisha Information sourced from drill hole logs, McPhie (2013) and Ashley (2013)

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The Harena deposit meets criteria 2, 3, and 5. Petrography is required to resolve relationships of mudstones and altered felsic volcanics within massive sulphides (criterion 1). Similarly, as with the other deposits, criteria 4 also needs resolution. As sedimentary and volcanic lithologies have been logged within the Harena massive sulphides, and criteria 2, 3, and 5 are met, Harena is considered a sub-seafloor replacement style deposit.

Ashley (2013) provides information from Bisha that also suggests that, in terms of deposit models, Bisha may also be a hybrid, with there being microscopic textures and mineralogies that could be related to chimney structures and exhalite occurring at the ore horizon. This could indicate that a portion of the Bisha deposit was also accumulated on the sea floor. This is not unlikely, as Barrie et al. (2007) indicate that the Bisha deposit was likely the product of a very long and continuous process of hydrothermal deposition.

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9 Exploration

Exploration activities at the various deposits and regional locales have included geological mapping, geochemical sampling, geophysical surveys, and drilling, and are summarized in Table 9‑1. Exploration activities and results are discussed in the following subsections; drilling results are discussed in Section 10.

Activities were conducted by BMSC personnel, or by consultants and contractors appointed by Nevsun or BMSC.

Exploration work completed by Sanu Resources (Sanu) for the Mogoraib River exploration licence, which includes the Hambok deposit, can be found in Section 6.

9.1 Grids and Surveys

The coordinate system used for all data collection and surveying is the Universal Transverse Mercator (UTM) system, Zone 37 and geographic coordinates in WGS84 (World Geodetic System, 1984).

During the 1999 exploration program, Nevsun established a local grid (not based on UTM coordinates) over the gossan area with a baseline 5.9 km in length oriented at an azimuth of 010° from magnetic north. Individual lines were usually spaced 200 m apart, and were of variable lengths.

At Bisha and Harena, the local grid constructed at the beginning of the 2003 program conforms to the UTM coordinate system, with a baseline oriented at 0° and section-lines oriented at 090°. Section-lines were usually spaced 100 m apart, except over the Bisha Main Zone, where they were spaced 25 m apart for drilling.

At Northwest, a local grid coordinate system replaced the older UTM-based grid that was used for all drilling campaigns prior to November 2012. Since the mineralization in Northwest strikes approximately UTM 045°, the local grid system was used to facilitate interpretation and modelling work. The majority of the drilling to date has been drilled perpendicular to the local grid.

Hambok uses the UTM coordinate system.

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Table 9‑1: Summary of Work Completed

Year	Deposit	Company	Type of Work	Description
1996	Regional	Ophir Ventures	Regional grassroots exploration	Prospecting, mapping, and sampling
1998	Regional	Nevsun	Property evaluation	Property examination and acquisition
1998	Regional	Nevsun	Property grassroots exploration	Reconnaissance scale geological mapping (1:50,000), geochemical stream sediment sampling
1999	Bisha	Nevsun	Geophysical surveys	Geophysical surveys – MaxMin horizontal loop EM and magnetometer
			Geological mapping	Property scale (1:5,000)
			Geochemical sampling	Soil sampling on three grid lines
2002	Bisha	Nevsun	Drilling	6 core drill holes (B-01 to B-06) totalling 759.0 m
2002	Bisha	Nevsun	Geological mapping	Discovery outcrop area (1:1,000)
2003	Bisha	Nevsun	Drilling	47 diamond drill holes (B-07 to B-53, B-02a) totalling 6,722.6 m
			Trenching	36 trenches sampled and mapped
			Geophysical surveys	Airborne EM and magnetometer (325 km²), pulse EM and ground magnetometer (73.5 line km), gravimetric survey (40 km)
			Geological mapping	Deposit scale (1:1,000), property scale (1:2,500), and regional scale (1:10,000) geological mapping
			Geochemical sampling	Stream sediment (165 samples), soil (39 samples), termite mound (115 samples), and auger and pit (33 samples)
			Petrographic study	11 thin sections by Vancouver Petrographics
			Metallurgical testing	2 oxide samples, 2 copper supergene mineralization and 2 primary mineralization samples
			Bulk density	260 samples determined on site, 44 samples sent to ALS Chemex for determination
2003	Bisha	Nevsun	Drilling	93 core drill holes (B-54 to B-146, and deepening B-40) totalling 11,750.8 m
2003	Bisha	Nevsun	Drilling	2 air blast holes for water wells completed by Eritrean Drilling

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Year	Deposit	Company	Type of Work	Description
			Geophysical surveys	Pulse EM, horizontal loop EM (151 line km), gravimetric survey (107.6 km)
			Geochemical sampling	pH soil survey, soil sampling (40.3 line km), whole rock (REE), regional prospecting
			Metallurgical testing	2 oxide samples and 2 copper supergene mineralization samples at PRA in Vancouver, some minor work at Kappes Cassidy in Nevada
			Petrographic study	13 thin sections by Vancouver Petrographics
			Bulk density	611 samples determined on site, 68 samples sent to ALS Chemex for determination
2004	Bisha	Nevsun	Drilling	163 core drill holes (B-147 to B-309) totalling 28,879.50 m
			Drilling	42 reverse circulation drill holes (BRC-01 to BRC-42) totalling 2,097.3 m
			Drilling	BRCD-26, BRCD-27, BRCD-32 to BRCD-34, BRCD-37, BRCD-38, BRCD-41, and BRCD-42) totalling 308.70 m
			Drilling	15 reverse circulation holes for water wells totalling 768 m
			Geophysical surveys	Gravimetric survey, 65.2 line km
			Geological mapping	Deposit scale (1:1,000) mapping and regional prospecting
			Geochemical sampling	Soil sampling (111.6 line km), whole rock (REE), prospecting
			Petrographic study	16 thin sections, 2 polished sections
			Bulk density	311 samples determined on site, 697 samples sent to ALS Chemex for determination
			Geotechnical work	All drill core oriented
			Environmental	Base line study implemented
			Metallurgical testing	2 primary sulphide samples tested at PRA in Vancouver
			Hydrological	Studies commenced
			Archaeological	Studies commenced
			Physical properties tests	On selected core samples of massive sulphide by JVX Geophysics
				Table continues...

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Year	Deposit	Company	Type of Work	Description
2005	Bisha and Harena	Nevsun	Drilling	112 diamond drill holes totalling 16,074.3 m
			Petrographic study	10 thin sections by Vancouver Petrographics
			Geochemistry	Whole rock analyses, petrographic studies, soil sampling
			Feasibility studies	Metallurgical sampling, testwork, geotechnical studies, other studies
			Geological mapping	Deposit scale (1:1,000) Harena and proposed tailings containment area
			Geophysical surveys	Gravity, HLEM, ground magnetometer
			Hydrological	Ongoing studies
			Archaeological	Ongoing studies
			Trenching	5 trenches at Harena (166.5 line m)
			Geotechnical work	All drill core oriented, point load and packer testing
			Metallurgical testing	1 bulk sample of primary sulphide tested at SGS Lakefield in Canada
			Bulk density	Harena and Northwest Zone prospects
			Geotechnical pit excavation	59 pits in areas of the proposed processing plant, accommodation and tailings dam
2006	Bisha, Northwest, Regional	Nevsun	Drilling	8 core drill holes totalling 1,680 m, 4 holes in Bisha Main and 4 other in Bisha Northwest
			Feasibility studies	Metallurgical sampling, testwork, geotechnical studies, other studies
			Geochemistry	Soil geochemistry on the Northwest Barite, HW copper zone, target 4 to target 9 geophysical anomalies
			Geophysical surveys	IP/Resistivity survey, Harena, Northwest Zone, Northwest Barite Hill and south of Bisha Main deposit
			Trenching and pitting	9 trenches on gold in soil anomalies and HW copper zone, 14 pits excavated in HW copper zone
			Geological mapping and prospecting	Regional scale
			Exploration of a source of aggregate	42 test pits in a basaltic dyke
			GPS Surveys	Proposed mine site, roads to Massawa and port area

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Year	Deposit	Company	Type of Work	Description
2007	Bisha	BMSC/ Nevsun	Geophysical surveys	Gravimetric survey, 13.5 line km. Target 9 area
2008	Bisha, Northwest, Regional	BMSC/ Nevsun	Prospecting	Targets 4 and 9
			Pitting	Targets 4 and 9, Bisha South and Northwest Barite Hill. 10 pits
			Trenching	Targets 4 and 9, Bisha South and Northwest Barite Hill. 6 trenches for 466 line m
			Geological mapping	Target 9 area 1:2000 scale
2009	Bisha, Harena	BMSC	Geophysical surveys	Gravimetric Survey, 32 line km
			Drilling	17 diamond drill holes totalling 2,163.5 m in Harena
			Geological mapping	Northwest of T9 area 1:5000 scale
			Geotechnical work	9 oriented drill core holes and point load testing
			Metallurgical testing	Bulk sampling of supergene material tested at Mintek in South Africa
2010	Northwest	BMSC	Drilling	47 core drill holes totalling 4,366 m
			Geophysical surveys	Gravimetric survey, 70 line km. South Tabakin (33 line km) and Northwest Barite Hill (37 line km)
			Petrographic studies	6 thin sections by Vancouver Petrographics and 11 polished thin sections by Carleton University
2011	Bisha, Northwest, Harena, Hambok	BMSC	Drilling	126 DDH, 2 metallurgical test and 6 geotechnical holes were drilled in Bisha. 23 DDH holes in Northwest Bisha, 5 DDH holes in Harena, 6 DDH holes in Hambok comprising 168 core drill holes totalling 35,780.2 m (including geotechnical and metallurgical holes below).
			Hydrogeology	2 water bore holes drilled at Harena and 16 RC water bores were drilled in Hambok
			Metallurgy	4 core drill holes (2 at Bisha and 2 at Harena) for oxide and sulphide metallurgical tests
			Geotechnical	10 oriented drill core holes (6 at Bisha and 4 at Harena) for geotechnical testwork
		Arcadia University	Geochemistry	686 whole rock lithogeochemistry samples from drill core

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Year	Deposit	Company	Type of Work	Description
2012	Bisha, Northwest, Harena, Hambok	BMSC	Drilling	31 DDH holes totalling 4,278.7 m drilled in Bisha Main (including geotechnical and metallurgical holes below). 77 DDH holes drilled in Northwest Bisha (including 8 geotechnical and 2 metallurgical testing holes; 4 DDH holes, and 2 RC water bores) were drilled in Harena. In Hambok, 25 DDH and 26 RC exploration holes were drilled
			Geotechnical	12 oriented drill core holes totalling 3,046.5 m and UCS, DST and Brazilian tests in Bisha Main. 8 Geotechnical oriented (EZ-Mark) holes were drilled in Northwest Bisha
			Metallurgical	12 core drill holes totalling 819.2 m for HW copper and supergene metallurgical tests in Bisha Main and 2 metallurgical test holes were drilled in Northwest Bisha. 5 Metallurgical samples were processed for Harena at ALS metallurgy Services. 4 metallurgical test samples from Bisha Main.
			Trenches	16 Trenches comprising 1,118.0 m in Bisha Northwest deposit (included as drill holes in the resource model)
			Geological mapping	Deposit Scale (I:2,000) Mapping in Bisha Northwest
2013	Bisha, Northwest, Hambok	BMSC	Drilling	33 DDH holes in Bisha (Bisha Deep /Extension) total 7,228.17 m. It also includes Metallurgy test holes. Other 474 RC Grade control holes, 6,243 m. 91DDH holes in Northwest Bisha, total 50,475.69 m. 4 other aquifer testing RC water bores, total 772 m were also drilled. 8 DDH holes in Hambok, Mogoraib EL, total, 2,713 m.
2013	Bisha, Northwest, Hambok	BMSC	Metallurgical	In Bisha Main Deposit. Metallurgical testwork mentioned above were total 10 DDH holes drilled 1,006.17 m. This was for the Zn phase III. In Hambok, Metallurgical sampling for flotation test and other metallurgical testing including mineralogical testing is done. Total 13 samples were processed at SGS. In Northwest Bisha 20 Composite samples, taken from Northwest 2 MET holes. In Bisha Main, 37 samples composited from drill core samples of 08 of the Metallurgy test holes, i.e., for mineralogical assessment. In Hambok, 13 samples were processed by SGS lab.

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Year	Deposit	Company	Type of Work	Description
2013	Northwest	BMSC	Hydrogeology	4 water bores in Northwest Bisha, total 772 m.
2013	Mogoraib EL	BMSC	Trenches	2 Exploratory Trenches were prepared, in Aderat, Mogoraib EL. Logging Assaying and Mapping done.
2013	Bisha, Northwest	BMSC	Petrographic studies	40 polished core samples Bisha Main Deposit and 33 samples from Northwest Bisha were studied by Paul Ashley Petrographic Services for Optical Microscopy analysis.
2013	Hambok	BMSC	Geotechnical	A preliminary assessment of the Hambok deposit was made using available geotechnical data.

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9.2 Geological Mapping

Geological mapping was completed to provide information on outcrop and gossan extents, geological units, and structure, at scales ranging from regional to property, including 1:50,000, 1:10,000, 1:5,000, 1:2,500, and 1:1,000.

Deposit Scale (1:2,000) detailed geology mapping was completed in the Northwest area from July to November 2012. A team of Consultant geologists from Taiga Consultants Ltd. (Calgary, Canada) was actively involved in this work.

9.3 Geochemical Sampling

9.3.1 Stream Sediment Sampling

Nevsun carried out stream sediment sampling in 1998, covering an area of 100 km². In 2003, 165 stream sediment samples were collected at an approximate density of one sample per square kilometre.

The stream sediment surveys were considered an effective method of delineating areas with potential for base and precious metal mineralization. The main anomalous areas for copper, lead, zinc, and gold, based on the combined 1998 and 2003 results, were the Okreb area (outside of the then current Bisha Exploration Licence), the Bisha Main Zone southwards towards the Harena area, and the NW Barite Hill area.

9.3.2 Rock Chip Sampling

During mapping and prospecting of the area, 461 rock chip samples were collected. The samples were used to help focus on prospective areas for VMS mineralization.

9.3.3 Soil Geochemical Sampling

In 1999, soil samples were collected over the Bisha gossan outcrop (Nevsun, 2003) that showed a distinct base metal anomaly. The samples were not analyzed for gold.

Between 2003 and 2006, 14,069 soil samples were collected and used to investigate geophysical anomalies, often in areas with minimal or no outcrop. This sampling also defined geochemical anomalies over selected geophysical targets and zones of known mineralization.

Initial soil sampling was performed in 2003, comprising orientation soil, termite mound, auger, and pit sampling. During the second exploration campaign in 2003, 40.3 line-km of soil sampling was completed over the NW Barite Hill, Bisha Main, and Harena areas. Based on the results of the orientation soil sampling, soil samples were collected at a shallow depth of no more than 10 cm, at intervals spaced 25 m apart, along grid lines spaced 50 m to 100 m apart. The grid lines were surveyed using a differential GPS unit.

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Soil sampling in 2004 comprised 111.6 line-km completed over the Bisha Main Zone, Harena, and NW Barite Hill areas. The soil sampling resulted in the definition of a coincident multi-element gold, silver, and lead anomaly over the Bisha Main Zone deposit.

At Harena, anomalous copper, lead, zinc, silver, and gold values were identified in the south-central portion of the gridded area. At NW Barite Hill, the analytical results of the soil geochemical survey defined a widespread coincident copper/zinc soil anomaly with subdued, weakly coincident gold, silver, and lead.

As part of the 2005 exploration program, additional soil sampling (5,005 samples) was completed southwest of the Harena prospect, east of the Bisha Main Zone, and over the proposed tailings containment area, in order to complete area coverage, or close-off previously-defined anomalous geochemical results.

Over the Northwest Zone grid, 843 soil samples were collected. For many of the contoured elemental plots, the highest values were present immediately west of the prospect; this represents an area where the footwall rhyolites rise to the surface, as the massive sulphide component of the deposit is at 30 m to 50 m depth, trending north-northeasterly, and plunging shallowly to the north.

A grid for soil and geophysical surveys covered the Harena area. In total, 2,724 soil samples were collected over this target. Three targets were identified on the Harena soil sample grid; these three soil anomaly areas may indicate the presence of VMS mineralization at depth:

At 1707000N/335450E, or approximately 750 m in the stratigraphic hanging wall from the Harena deposit, highlighted by all of the metals analyzed, except Cu and Fe
At 1705300N/333850E, or 2.7 km southwest along strike from the Harena deposit, highlighted by Cu, Zn, Hg, S, and U
At 1710000N/336750E, highlighted weakly by Cu, Fe, and Mo.

During 2006, soil samples were taken on the Target 4, 6, 7, 8, and 9 geophysical anomalies (Figure 9‑2). Targets 4 and 9 returned anomalous geochemical signatures; Targets 6, 7, and 8 did not.

Additional sampling was performed over the Northwest Zone, NW Barite Hill, and the Hanging Wall Copper Zone.

At the Bisha Main Zone deposit, additional soil samples were collected on a 100 m x 100 m grid spacing to complete the geochemical coverage of the interpreted location of the Hanging Wall Copper Zone. The limited additional sampling at the Bisha Main Zone did not enhance the soil geochemical interpretation for the area.

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Additional soil sampling was completed over the Northwest Zone, extending the soil coverage westward. Grid lines were placed using hand-held GPS units in combination with chaining and back sight methods. Soil samples were collected at a shallow depth of no more than 10 cm along the grid lines, at 25 m station intervals. Results from the samples collected show a continuation of the previously-defined mineralization trends. The gold-in-soil geochemical trend appeared to continue to the southwest and remains open in this direction. The copper-in-soil geochemical anomaly appears to trend towards the copper-in-soil anomaly defined over the Bisha Main Zone deposit. Additional, infill sampling remains to be completed to provide complete coverage between the Bisha Main Zone and Northwest Zone soil grids.

Soil sample coverage over the NW Barite Hill was resurveyed northward from UTM 1716800N to the northern exploration licence boundary, to provide coverage on 100 m x 100 m grid lines. The area was resampled because the previous analytical method used had a higher detection limit than the remainder of the survey data for the Project, making levelling of the analytical data problematic. The coverage area was also extended to conform to the soil geochemical coverage to the south. There were no significant soil geochemical anomalies detected.

9.3.4 Termite Mound Sampling

Over an approximate area of 55 km², 107 termite mound samples (including four duplicates) and 8 auger samples of the mounds were collected. The termite sampling provided some additional geochemical information, but was limited to areas with mounds.

9.3.5 Soil pH Geochemical Sampling

Soil geochemical surveys that collected pH signatures were conducted over the known Bisha Gossan to test the theory that a pH measurement can identify the change in pH related to the presence of massive sulphides (and related generation of acid conditions related to oxidation of the sulphides), even below alluvial cover. Nevsun found that the pH technique works very well in the delineation of known sulphide mineralization in alluvial-covered areas, and considered that it may be used to define new targets. Unfortunately, the survey reacts to a wide variety of types of underlying chemical differences, and thus produces a large number of anomalies that need to be prioritized.

9.3.6 Auger Geochemical Sampling

During 2003, 39 samples (23 soil samples and 16 auger samples) were collected to test the hand auger as a geochemical sampling tool. Auger sampling was not determined to be advantageous, and therefore was not continued.

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9.4 Remote Sensing and Satellite Imagery

Two remote sensing studies have been performed.

Nevsun prepared Landsat images of the area in 1998. Using the Landsat image with a translucent (30%) topographic overlay, a preliminary interpretation was made of the immediate Bisha occurrence area. The interpretation focused on structural features that were subsequently plotted on the geology map as well as the preliminary gravity maps.

During 2003, Earth Resource Surveys Inc. (ERSI), based in Vancouver, Canada, completed a remote sensing investigation for the Bisha Project and western Eritrea. The survey mapped alteration types and interpreted major structural features using different Landsat bands, and highlighted alteration and structural trends (Chisholm et al., 2003).

9.5 Geophysics

9.5.1 Ground Geophysics

Horizontal loop and pulse EM surveys were completed in 1999, 2003, and 2005. Most of the EM surveys were successful in delineating mineralization and other features.

Magnetometer surveys were completed on portions of the Project in 1999, 2003, and 2005. The magnetometer surveys showed lithological contrasts and anomalies over the Bisha Main Zone, but features were generally less distinct than in the EM surveys. Chisholm et al. (2003) observed north-northeast-trending features that were interpreted to show a fault zone that transects the Bisha Main Zone deposit.

Gravity surveys were carried out during 2003, 2004, 2005, 2007, and 2009. The filtered residual gravity surveys provided good definition of the Bisha Main Zone, Northwest Zone, and Harena massive sulphide mineralization.

In December 2005, Nevsun conducted a reinterpretation of previous airborne geophysical surveys and ground geophysical surveys. Induced polarization (IP) and resistivity surveys were carried out over the Harena, Northwest Zone, and NW Barite Hill prospects, and over the area to the south of the Bisha Main Zone. The results of the IP/resistivity surveys showed a distinct low-resistivity anomaly coincident with the massive sulphide body at the Northwest Zone. At Harena, there is a low amplitude resistivity anomaly associated with the sulphide mineralization. The IP and resistivity responses at the NW Barite Hill area are characterized by weak to moderate chargeability highs, with associated resistivity lows.

In April 2007, BMSC initiated a 13.5 line-km gravity survey over a newly interpreted target area on the Project referred to as Target 9. MWH Geo-Surveys Inc. (MWH) conducted the gravity survey. The survey lines were spaced 200 m apart, with stations spaced at 25 m intervals along the survey lines. The gravity survey identified a weak residual gravity anomaly with coincident EM conductors. This, combined with previously-identified soil geochemical anomalies, outlined an area of potential VMS mineralization that required follow-up investigation.

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In 2009, the gravity survey was extended to the southwest corner of the exploration licence, with an additional 32 line-km of data collected (Figure 9‑1). The survey provided further definition of gravity highs to the southwest, and along strike of the Harena deposit. Ongoing analysis of the data should help further identify exploration targets in the area that may have potential for VMS mineralization.

In 2010, gravity surveys were initiated in two areas: a 37 line-km survey over the NW Barite occurrence, and a 33 line-km survey in an area directly south of the Tabakin hills, south of Bisha. The survey lines were spaced 200 m apart, with stations spaced at 25 m intervals along the survey lines. Results of the survey south of the Tabakin hills identified two gravity anomalies associated with mafic intrusions. The survey at the NW Barite occurrence identified two strong gravity anomalies, which were drilled in June 2010, intersecting mafic tuffs with no significant mineralization. Survey results are shown in Figure 9‑1.

MWH performed the gravimetric surveys in 2009 and 2010.

9.5.2 Aerial Geophysics

In March 2003, a combined airborne EM and magnetometer fixed-wing survey was conducted over an area of approximately 325 km². The survey identified a number of horizons that were considered prospective for VMS mineralization.

In December 2005, Nevsun reinterpreted the 2003 airborne EM survey, resulting in several priority areas being demarcated for further work. Several priority VMS targets were identified (marked as second-tier anomalies on the map), as shown in Figure 9‑2.

The Target 4 area, located immediately to the west of the Northwest Zone, exhibits a coincident airborne EM and residual gravity anomaly. The position of the anomaly is marked on the ground by an abundance of gossanous boulders, suggesting the presence of massive sulphide mineralization.

Target 9 is located approximately 5 km to the southwest of the Bisha Main Zone deposit. Soil sampling over the area indicated by the airborne EM responses returned anomalous results from multiple elements. The area is considered prospective for VMS mineralization.

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Figure 9‑1: 2010 Gravity Survey Results

Source: Nevsun

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9.6 Pits and Trenches

In 2003, a series of 36 trenches were excavated over various parts of the Bisha Main and Northwest Zones. Rock samples returned elevated base metal and precious metal results.

In 2005, five trenches (166.5 m) were excavated at the Harena prospect, nine trenches were excavated in the Hanging Wall Copper Zone and the Northwest Zone, and 14 pits were excavated in the Hanging Wall Copper Zone. The trenches in the Hanging Wall Copper Zone intersected significant intervals of copper mineralization, with rock samples returning grades ranging from 0.27% Cu to 2.68% Cu, and 0.3% Zn to 1.33% Zn.

In 2008, Targets 4 and 9, Bisha South, and NW Barite Hill were subject to trenching and pitting, with ten pits and six trenches completed. The trenches at Targets 4 and 9 intersected mudstones and shales with varying amounts of graphite. The Bisha South trenches did not intersect any significant mineralization. At NW Barite Hill, the trenches intersected chlorite-sericite-altered schist with anomalous lead values up to a maximum of 0.23% Pb.

The trench data described above were not used in geological modelling or resource estimation.

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Figure 9‑2: Map Showing Interpreted Location of Airborne Magnetic Anomalies

Source: Nevsun

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At the Northwest Zone, 16 trenches totalling 1,118 m were completed in 2012 (Figure 9‑3). Detailed logging suggested the presence of gossans, highly weathered zones, and leached stringer zones, confirming the geological mapping work done earlier in the year. The trench assays show elevated values for gold in several trenches in the southern, central, and eastern parts of the Northwest deposit, as shown in Table 9‑2. This work highlighted the mineralized zones, and helped in locating the shallow oxide/supergene gold distribution by exposing gossans and oxidized stringer zones projecting to the surface. Some of these results are incorporated into the geological modelling and Resource Estimate for the Northwest Zone.

Figure 9‑3: Northwest Zone 1:2,000 Scale Geological Map and Trenches

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Table 9‑2: Northwest Trenching 2012, Summary of Results

Trench N^o	East	North	RL	Trench Length (m)	From	To	Width (m)	Au g/t
NWTR001	338426.1	1717459.8	555.3	14	0	5.5	5.5	3.77
					7.5	14	6.5	0.41
NWTR002	338417.4	1717555.5	553.5	120	51	55	4	0.52
NWTR003	338462.7	1717579.0	553.9	74	26	28	2	0.2
NWTR004	338561.9	1717663.8	553.8	126	56	60	4	1.39
					68	72	4	0.44
NWTR005	338609.1	1717685.8	554.7	10
NWTR006	338543.6	1717560.1	553.2	18	0	12	12	1.69
					14	18	4	0.37
NWTR007	338525.1	1717697.7	553.3	20	11	12	1	0.38
					16	18	2	0.21
NWTR008	338553.1	1717740.7	553.3	22	18	22	4	0.37
NWTR009	338821.9	1717900.9	556.5	30	12	24	12	0.75
NWTR010	338844.6	1717946.8	555.7	96	80	94	14	0.62
NWTR011	338865.4	1717996.1	553.8	140	98	100	2	0.87
NWTR016	338555.6	1717585.5	553.2	36	6	16	10	0.27

In mid-2013s, two trenches were completed at Aderat, a prospect in Mogoraib EL following up old anomalies and gossans in the area. The objective was to expose a possible northern extension of the Hambok mineralized trend following existing anomalies. Additional trenches are planned for this prospect. The trenching is shown in Figure 9‑4.

Strong hematite-limonite leaching in a 43 m long section of the weathered zone was exposed in one of the trenches with pockets of gossan. One trench cross-cuts an old Sanu Resources trench. Assays showed no significant base metal results; one zone of 7 m has copper values of approximately 0.25% Cu in one of the trenches.

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Figure 9‑4: Trenches at the Aderat Prospect, Mogoraib EL (1:2,000 Scale)

9.7 Petrology, Mineralogy, and Research Studies

9.7.1 History

Petrographic studies were undertaken on mineralization samples, including thin and polished sections, during 2003–2004. Thin section samples were collected from different types of lithology that Nevsun personnel had problems identifying during drill core logging. Two polished section samples were prepared from the primary massive sulphide lens (Bisha South), representing the zinc-rich zone and the copper-rich zone. The results of the study generally confirmed the lithological types logged by Nevsun personnel.

Over 500 whole-rock analyses were completed during 2003 and 2004 (Daoud, 2004). Winchester-and-Floyd trace-element diagrams show that the mineralization host rocks have a mafic to intermediate affinity (basaltic to dacitic chemistry). Alkali-silica diagrams (Lebas et al., 1986) show that the Bisha volcanic rocks vary from picrites (highly enriched in MgO) to rhyolites; however, the diagrams are affected by alteration, and therefore rocks plotting in the rhyolite field are probably silica-altered (enriched) intermediate rocks.

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In 2005, 18 whole-rock analyses were completed on samples from the Harena prospect. The results of the study show there are two populations of basaltic (mafic) and rhyolitic (felsic) rocks.

In April 2010, Brett Atkinson, a student at Carleton University, Canada, completed a thesis entitled “A Petrographic and Microprobe Study of Oxide Gold-Silver Mineralization and Gangue at Bisha, Eritrea.” The results of the study show that the Bisha oxide zone contains very pure native gold with some silver minerals (naumanite, acanthite, and chloroargyrite). These ore minerals are hosted by a matrix of colloform goethite and hematite, quartz, traces of other silicates, barite, trace anglesite, and jarosite (mainly plumbojarosite with trace natrojarosite).

9.7.2 Summary of Recent Research on the Bisha Deposit

In January 2013, 37 samples from the Bisha supergene zone of mineralization were analyzed as part of the metallurgical testwork program. These samples were composited from 230 individual sub-samples of drill core from 8 drill holes (MET-024, MET-025, MET-026, MET-027, MET-028, MET-029, MET-030, and MET-031). The objective was to determine the mineral composition of the samples. A combination of Bulk Mineral Analyses (BMA) using QEMSCAN and XRD analyses were used to achieve this objective.

The mineralogical analysis showed that pyrite was the dominant mineral in the samples. Only four samples contained less than 90% sulphide mineral content. The copper sulphide mineral assemblage was relatively complex. On average, 40% of the copper was contained in chalcopyrite, with a further 38% in covellite. Small amounts of chalcocite, bornite, and enargite/tennantite were also recorded. Significant zinc was observed in the form of sphalerite in a small number of samples. Arsenic was observed in copper sulphide minerals enargite, tennantite, and tetrahedrite. The assumption made in earlier studies was that arsenic was present exclusively as arsenopyrite, and thus amenable to flotation with pyrite.

In late 2013 and early 2014, Paul Ashley of Petrographic and Geological Services, Australia (Ashley), completed further petrological studies. Forty polished core samples from the Bisha Main deposit were tested. This study was carried out to further characterize the Bisha lithologies to better understand the geological setting. Although not specifically a study of mineralization, the work examined the complexity of the arsenic mineral assemblage of enargite with covellite and digenite.

In August 2013, Prof. Jocelyn McPhie (University of Tasmania), a world-renowned volcanic facies specialist, spent two weeks reviewing Bisha, Northwest, and Hambok stratigraphies as visible in core and outcrop across Bisha tenements. Detailed outcomes of the review are presented in “Volcanic Facies of the Bisha and Hambok Licence Areas, Eritrea,” (McPhie, 2013). Key new information includes evidence that the Bisha Deposit formed through sub-seafloor replacement and impregnation rather than sulphide deposition at the seafloor.

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Work by Cube Consulting (Bampton, 2013) in conjunction with BMSC Technical Services staff, established that the main controls on Bisha mineralization physical and grade distribution, post initial deposit emplacement, were structural. Outcomes of this work are presented in BMSC internal reports “Bisha Main Geological Model” and “Bisha Project Arsenic Model.” Previous literature defined Bisha as being a folded VMS deposit.

9.7.3 Summary of the Recent Research on the Northwest Deposit

Thirty-three DDH core samples, as polished thin sections, were processed and examined by Ashley in late 2013. The objective was to assist in the characterization of Northwest lithologies and to review ore mineral deportment. A summary of the results of this study are given below.

All samples (with one exception) show metamorphism, hydrothermal alteration, and penetrative deformation. Of the observed sulphide mineralization, pyrite is the predominant mineral.

Several samples in the suite are interpreted to show effects of supergene alteration, ranging from incipient to strong replacement. The effect of moderate supergene alteration is evident in trace covellite development from chalcopyrite, and the development of fine-grained secondary pyrite from pyrrhotite.

Stronger supergene alteration occurs in samples with likely replacement of prior sulphides and formation of goethite, and minor jarosite formation of veins of chalcedonic silica and goethite/jarosite.

The majority of sulphides are considered pre-metamorphic, with evidence of hydrothermal alteration. Where sulphides are more concentrated, ore textures could reflect direct precipitation from submarine exhalative fluids.

Prof. J. McPhie (2013) concluded, as with the Bisha Deposit, that Northwest had direct stratigraphic similarities to the Bisha deposit, indicative of mineralization having formed through sub-seafloor replacement and impregnation, rather than sulphide deposition at the seafloor. The association of rhyolitic lithologies with mineralization was observed, and found to be important to mineralization localization at the Bisha and Northwest deposits.

9.7.4 Summary of the Recent Research on the Hambok Deposit

Prof. J. McPhie (2013), as part of a lithostratigraphic review of VMS deposits across BMSC’s tenements, concluded that the Hambok deposit was formed through sub-seafloor replacement of a basalt host succession. The mechanism of formation is similar to that for the Bisha and Northwest deposits, but within a different stratigraphic succession (basaltic versus rhyolitic).

9.7.5 Summary of the Recent Research on the Harena Deposit

A Master of Science thesis titled “Petrology and Lithogeochemistry of Host Rocks to the Bisha and Harena Cu-Zn-Au Volcanic Hosted Massive Sulphide Deposits, Eritrea” was completed in September 2012 by R.J. Massawe (Acadia University). The main outcome of this work was recognition that the Bisha and Harena deposits are not hosted at the same stratigraphic level within the regional volcanic stratigraphy.

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9.8 Geotechnical and Hydrological Studies

Geotechnical and hydrological programs conducted were sufficient to support mining studies, and are ongoing (refer to Section 16).

Geotechnical studies carried out since the previous Technical Report are summarized in Table 9-1. In addition, BGC Engineering (Vancouver) made a site visit in late 2013, reviewing current pit conditions in relation to the current update of Mineral Reserves.

Geotechnical mapping of the Bisha open pit, and modelling and mapping of the pit wall geology and structure, has been continuous from late 2012 to date. With the help of consultant firms Geologica, Cube, and Gnomic Exploration Services, methods have been developed and established, with routine data now adding to a GIS database. Additional 3D modelling of this data is now planned to enhance the value of this work to the operating mine.

Eight geotechnical holes were drilled and logged at Northwest by the end of 2012. The data and results from associated testwork was modelled and analyzed by BGC, giving initial geotechnical domains for this deposit. The domains divided the deposit into three vertical zones, with varying slope angle and berm characteristics. These domains were employed as a part of the Mineral Resource estimate described in this report.

As a part of the Hambok Mineral Resource estimate, a preliminary analysis of geotechnical data derived from standard geotechnical core logging was made. This initial study established an overall pit slope based on available deposit data, rather than an assumed default, and was used as part of the Mineral Resource estimate described in this report.

Hydrogeological drilling is discussed in Section 10.

9.9 Exploration Potential

Since the May 2012 Technical Report, BMSC purchased the Mogoraib EL, which includes the Hambok deposit, and has tenure until July 2014, after which the normal reapplication process will take place. The Mogoraib EL has good exploration potential along mineralized trends parallel to the Bisha trend. The current exploration program for this licence is summarized below:

regional diamond drilling testing targets along mineralized trend in Hambok North, Hambok South and at Aderat; Hambok West
generative ground geophysics at Mogoraib EL and within Bisha mine areas
downhole borehole geophysics
geochemistry – assessment and regional program

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geological mapping, rock sampling, and trenching.

Additional potential exists at the current deposits, as summarized below, and more fully detailed in Section 26:

Bisha – down-dip/strike potential at the south of the Main Zone
Northwest – strike and down-dip potential
Harena – strike and down-dip potential
Hambok – strike potential.

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10 Drilling

Drilling on the four deposits described in this report has been undertaken in a number of diamond drill (DD) and reverse circulation (RC) campaigns from 2002 to 2013. Three of the deposits, Bisha Main, Northwest, and Harena, lie close to each other, with Hambok lying some 18 km to the southwest in the Mogoraib EL. The drilling areas are shown in Figure 10‑1.

Drill programs have been completed, primarily by contract drill crews, largely supervised by BMSC geological staff for Bisha, Harena, and Northwest, and by Sanu and BMSC staff in the case of Hambok.

10.1 Bisha Deposit

Six hundred and eighteen DDH were drilled between 2002 and 2013, including 9 RC pre-collared DDH, 14 geology-related blast holes, 33 geotechnical drill holes, and 40 metallurgical testwork drill holes. The drilling is summarized in Table 1‑1. The RC pre-collared DDH, some of the metallurgical holes, and the geotechnical holes were not included in the resource estimate described in this report. A series of other drilling campaigns, namely the BEXT and BNDD holes, were also not included in the resource estimate.

Five hundred and fourteen RC holes, including 474 grade control holes, were also drilled in 2013 in grade control drilling campaigns. A total of 498 DD and 474 RC grade control holes were used in the Mineral Resource Estimate.

Diamond drilling was mostly completed by Boart Longyear Eritrea Limited (Boart Eritrea) using LF90 track-mounted drill rigs; modern RC grade control drilling was carried out in-house, with BMSC employing two Schramm T450 Rigs. Major Pontil Pty Ltd. (Major), an Australian subsidiary of Major Drilling Inc., based in Queensland, Australia, performed the 2004 RC drilling.

Much of the massive sulphide mineralization in the Bisha Main deposit has been well defined with drilling spaced at 25 m x 25 m or closer in some areas. Drilling density decreases with depth with the deposit remaining open at depth in the south.

The drilling completed at the deposit is of sufficient density and quality for meaningful geological interpretation and Mineral Resource and Mineral Reserve estimation, as described in detail in Section 14 and Section 15 of this report.

With the oxide zone now mined out, the sample recovery problems experienced in this zone have only minor influence on the Mineral Resource and Mineral Reserves.

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Figure 10‑1: Drill Hole Location Map

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Table 10‑1: Summary of Drilling for the Bisha Main Deposit

Year	Phase	Range of Hole No.	No. of DDH	Length of DDH (m)	No. of RC Holes	Length of RC Holes (m)	Total No. of Holes	Total Length (m)
2002	-	B-001 to B-006	6	810.9	-	-	6	810.9
2003	I	B-002a, B-007 to B-0 53	48	6,724.8	-	-	48	6,724.8
2003	II	B-058 to B-146	79	9,861.7	-	-	79	9,861.7
2004	-	B-147 to B-309	163	28,954.1	-	-	163	28,954.1
2004	-	BRC-001 to BRC-041*	-	-	40	2,117.40	40	2,117.4
2004	-	BRCD-026 to BRCD-42*	9	308.8	-	282.90	9	591.7
2005	I	B-310 to B-367	58	7,520.0	-	-	58	7,520.0
2005	I	GT-01 to GT-05	6	937.0	-	-	6	937.0
2005	II	MET-05-01 to MET-05-08	8	1,212.5	-	-	8	1,212.5
2005	II	BH-01 to BH-12, BH-14, and BH-15	14	410.0	-	-	14	410.0
2006	-	B-368 to B-371	4	1,014.0	-	-	4	1,014.0
2009	-	GT-006 to GT-014	9	1,201.5	-	-	9	1,201.5
2009	-	MET-009 to MET-011	3	224.0	-	-	3	224.0
2010	I	B-372 to B-378	7	1,062.0	-	-	7	1062.0
2010	I	MET-012 to MET-017	6	478.0	-	-	6	478.0
2011	I	B-379 to B-399	21	2,592.4	-	-	21	2,592.4
2011	II	B-400 to B-504	105	24,494.1	-	-	105	24,494.1
2011	II	MET-018 to MET-019	2	180.0	-	-	2	180.0
2011	II	GT-015 to GT-020	6	695.5	-	-	6	695.5
2012	I	B-505 to B-511	7	409.5	-	-	7	409.5
2012	-	MET-020 to MET-031	12	820.7	-	-	12	820.7
2012	-	GT-021 to GT-032	12	3,048.5	-	-	12	3,048.5
2013	-	B-512, BEXT001 to BEXT014, BMT032 to BMT041, BNDD001 to BNDD008, BRC (RC grade control holes)	33	7,228.2	474	6,243	507	13,471.2
Total			618	100,188.1	514	8,643.3	1,132	108,831.4

10.2 Harena Deposit

A total of 105 drill holes have been completed in Harena, comprising 98 DDH and 7 RC holes drilled between 2005 and 2012. The DD holes include four geotechnical and two metallurgical test holes. A summary of the drilling is given in Table 10‑2.

Only DDH were used for resource estimation in the primary zone; for the oxide/supergene gold zones additional information was derived from the RC holes. Not all holes were included in the estimation work.

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Diamond drilling was completed by Boart Eritrea; RC drilling was carried out in house, with BMSC employing an Atlas Copco 850 rig.

For the oxide and supergene mineralization, drill hole spacing was 25 m x 25 m. For the primary mineralization a pattern of 50 m x 50 m spaced holes was maintained.

The drilling completed across the deposit is of sufficient density and quality for meaningful geological interpretation and Mineral Resource and Mineral Reserve estimation, as described in detail in Section 14 and Section 15 of this report.

Table 10‑2: Summary of Drilling for the Harena Deposit

Year	Phase	Range of Hole No.	No. of DDH	Length of DDH (m)	No. of RC Holes	Length of RC Holes (m)	Total No. of Holes	Total Length (m)
2005	I	H-001 to H-020	20	2,909.7	-	-	20	2,909.7
2005	II	H-021 to H-027	7	1,162.0	-	-	7	1,162.0
2009	I	H-028 to H-044	17	2,164.0	-	-	17	2,164.0
2010	I	H-045 to H-047	3	450.0	-	-	3	450.0
2010	II	H-048 to H-081	34	2,466.7	-	-	34	2,466.7
2011		H-082 to H-086	5	851.5	-	-	5	851.5
2011	II	HGT-001 to HGT-004, HMET001 to HMET002, HWB2011-001, and HWB2011-002	6	764.0	2	110	8	874.0
2012	I	H-087 to H-092, and HWB12-001 to HWB12-005	6	1,123.1	5	536	11	1,659.1
Total			98	11,891.0	7	646	105	12,537.0

10.3 Northwest Deposit

A total of 287 drill holes have been drilled at the Northwest deposit in a series of campaigns since 2003, comprising 240 DD and 47 shallow RC grade control exploration-related holes. The DDH include eight geotechnical and two metallurgical test holes. A summary of the drilling is given in Table 10‑3.

Only DDH were used for resource estimation in the primary zone; for the oxide/supergene gold zones, additional information was derived from the RC holes and trenches. Not all holes were included in the estimation work.

Diamond drilling was completed by Boart Eritrea; RC drilling was carried out in-house, with BMSC employing an Atlas Copco 850 rig.

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Much of the massive sulphide mineralization in the Northwest deposit has been well defined, with drilling spaced at 25 m x 25 m or closer in some areas. Drilling density decreases with depth, with the deposit remaining open at depth and along strike.

The drilling completed across the deposit is of sufficient density and quality for meaningful geological interpretation and Mineral Resource estimation, as described in detail in Section 14 of this report. The recovery problems in the Oxide and supergene Zones were countered by classification of this material as Inferred Resources. Additional drilling will be needed to upgrade the resource category.

Table 10‑3 Summary of Drilling for the Northwest Deposit

Year	Phase	Range of Hole No.	No. of DDH	Length of DDH (m)	No. of RC Holes	Length of RC Holes (m)	Total No. of Holes	Total Length (m)
2003	I	B-054 to B-083	14	2,033.5	-	-	14	2,033.5
2005		BND-001 to BND-011 (Sanu)	11	2,392.77	-	-	11	2,392.77
2005	I	NW-001 to NW-022	22	3,901.85	-	-	22	3,901.85
2006	I	NW-023 to NW-026	4	666	-	-	4	666
2011	II	NW-027 to NW-049	23	4,325.7	-	-	23	4,325.7
2012	II	NW-050 to NW-099	50	9,293.86	-	-	50	9,293.86
2012		NWGT001 to NWGT008	8	1,568.1	-	-	8	1,568.1
2012		NWMT001 to NWMT002	2	471.0	-	-	2	471.0
2012-2013	III	NWDD100 to NWDD206	106	21,546.8	-	-	106	21,546.8
?	-	NWAA001 to NWAI001	-	-	47	872.5	47	872.5
Total			240	46,199.58	47	872.5	287	47,072.08

10.4 Hambok Deposit

A total of 145 holes have been drilled across the Hambok deposit in a series of campaigns, comprising 102 DD and 42 RC holes. A summary of the drilling is given in Table 10‑4.

Only DDH were used for resource estimation in the primary zone; for the oxide gold zones information was mostly derived from the RC holes. Not all holes were included in the estimation work.

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Table 10‑4 Hambok Deposit Mogoraib EL Drill Hole Summary Table

Year	Phase	Range of Hole No.	No. of DDH	Length of DDH (m)	No. of RC Holes	Length of RC Holes (m)	Total No. of Holes	Total Length (m)
2006	I	HAM-001 to HAM-045	44	7,282.5	-	-	44	7,282.5
2006	II	HAM-045 to HAM-055 and HAM-058	12	5,105	-	-	12	5,105
2007		HAM-056, HAM-060, and HAM-061	3	766	-	-	3	766
2010		HAM-10-062 to HAM-10-065	4	974	-	-	4	974
2011		HAM-11-066 to HAM-11-071, HAM-RC-11-01 to HAM-RC-11-16	6	1,877	16	978	22	2,855
2012	I	HAM-12-072 to HAM-12-096, HAM-RC-12-17 to HAM-RC-12-42	25	8,072	26	1,698	52	9,770
2013	II	HAM-13-097 to HAM-13-104	8	2,713	-	-	8	2,713
Total			102	26,789.5	42	2,676	145	29,465.5

Diamond drilling was completed by Boart Eritrea in 2006 for Sanu and in 2013 for BMSC, with the remaining diamond cored holes completed by General Exploration Drilling Pty Ltd, (GED) an Australian-based company. GED typically employed an RC pre-collar of up to 100 m.

Much of the massive sulphide mineralization in the Hambok deposit has been well defined, with drilling spaced at 50 m x 50 m or closer in some areas. Drilling density decreases with depth, with the deposit remaining open at depth and along strike.

Additional drilling is needed in the oxide zone to better define this zone.

10.5 Drill Methods

10.5.1 Core Drilling

Diamond core drilling was undertaken to provide geological, mineralogical, metallurgical, hydrological, and geotechnical information on the various deposits.

Kluane International Drilling (Kluane), a contractor based in Vancouver, Canada, completed diamond drilling at Bisha between 2002 and 2003 using a “man-portable” drill rig. The unit uses a 1.5 m long NTW-core barrel (55.1 mm diameter core), which reduces in bad ground to a BTW-sized (41.0 mm diameter core) 1.5 m or 3.0 m long core barrel.

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Boart Longyear, a contractor based in North Bay, Ontario, Canada used Longyear 44 skid-mounted wire-line rigs. Holes were collared with HQ core (63.5 mm diameter) until ground conditions necessitated a reduction to NQ-sized core (47.6 mm diameter). Not all holes were reduced if the ground conditions permitted reasonable penetration or if ground conditions were not favourable for reduction to a smaller core diameter (for example, if the ground was badly fractured). At various times holes were collared with PQ (85.0 mm), and then reduced to HQ core in order to improve recovery, particularly in the oxide zone at Bisha, Harena, and Northwest. Success with this method was variable; improvements were also made by reducing penetration rates and using shorter core runs.

GED followed a similar drilling methodology, and made some use of RC pre-collars of up to 100 m.

Core barrels were retrieved by wire line. Upon retrieval, the split tube was opened by the driller’s helper, who transferred the core to a galvanized steel core box. The core was marked where it was manually broken to fit into the box. Drill depths were marked with wooden or plastic blocks. During 2013, the depth markers were augmented with the EzyMark™ orientated core block marker system.

On completion of most holes PVC casing was inserted and cemented in place with the hole number marked.

10.5.2 Reverse Circulation Drilling

Major performed the 2004 RC drilling at Bisha. They used a universal drill rig (UDR-650-P35 combination drill) with a centre-sample return, triple-wall system to drill holes with a diameter of 136 mm.

Sample discharge and sample splitting equipment consisted of cyclone collectors mounted above Jones splitters for both wet and dry drilling. Representative samples were stored in plastic chip trays for geological logging. Each chip tray represented about 2 m of drilling.

Records kept by Sanu Resources do not indicate the contractor used for their RC drilling programs; it is assumed to be GED, which had previously drilled pre-collars for diamond holes across the Hambok deposit.

Current RC grade control drilling at Bisha, the results of which feed into the Resource estimation, is being carried out in-house by BMSC. Two RC Schramm T450 rigs operate in the Bisha Pit and environs. The shallow 10 m to 30 m deep holes are typically drilled on a 10 m x 10 m staggered pattern. Patterns are varied according to the nature of the grade control program. Sample discharge and sample splitting equipment consists of cyclone collectors mounted above Jones 2-tier riffle splitters for both wet and dry drilling. Representative samples are stored in plastic chip trays for geological logging. A subsample is also retained for additional analytical work.

The sample database validation process completed for Bisha (described in Section 14) resulted in the exclusion of data that was previously used in the Resource Estimates for the now-mined oxide zone. Historical Bisha blast hole and rip-line samples in the sample database are now considered redundant.

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10.6 Geological Logging

10.6.1 Core Logging

BMSC Process

Core is transferred to the BMSC core management facility for logging, sampling, and storage. The core logging and storage facility includes a large covered area for logging, handling, splitting, and storage of core within the mine perimeter. The core was logged by a geologist for geological and geotechnical elements, including lithology, alteration, mineralization, structure and geotechnical information, and sample recovery. Logging methods and procedures have evolved with the input of consultant geologists from various groups during 2012 and 2013 (Taiga and Gnomic Exploration Services).

The geologist determines the sampling intervals, adhering to lithological intervals. The sample intervals are identified by waterproof Tyvek tags indicating a sample interval, and crayon marks on the core for the beginning and end of each sample. The geologist also marks the cut line for the core cutters to follow if there is potential for any apparent bias in mineralization.

In the 2013 drill programs across the Northwest and Hambok deposits, core orientation was achieved using the EzyMark™ system. The system uses a down hole tool that captures the imprint of the core face at the start of a run of core using a ring of movable nail heads. A coloured chinagraph pencil amidst the nails assists in matching the nails when the core break is flat. Three ceramic balls at the back of the orientation tool lock into position on contact of the nail head with core. The position of the ceramic balls is equivalent to being the bottom of core at the end of a drilling run. Matching the nail-head imprint and pencil mark to core once it is in the core tray allows the core to be oriented relative to the bottom of the hole.

The core is photographed using a digital camera, and then transferred for sample collection and/or to the core storage area.

Sanu Process

The Sanu core sample process, as recorded in their reports, was as follows:

The Sanu core logging and storage facility included a large covered area for logging, handling, splitting, and storage of core within the Mogoraib field camp. After the core arrived in camp, it was washed, and meterage blocks were checked to ensure that no errors were present in the runs recorded during drilling. The logging system included codes for key aspects of the geology and the style of mineralization. The mineralization was logged on hard copy log forms. The logged information was later entered into an Excel spreadsheet by the same geologist that logged the hole. The key data types captured during core logging were geotechnical, lithologic, alteration, structure, and mineralization.

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A geologist determined the sampling intervals, with 1 m being the target interval. If natural lithological breaks occurred within a designated sample interval, that sample interval was then adjusted to coincide with the natural break. No mineralization-based distinctions were made in the massive sulphide units. The sample intervals were identified by felt pen marks on the core for the beginning and end of each sample. The geologist also marked the cut line for the core cutters to follow if there was potential for an apparent bias in mineralization. The cut lines were marked along the axis of the core, perpendicular to the mineralized interval.

Generally, more than one geologist was involved in logging a single hole, often working in pairs. This was done to reduce observational drift and to maintain uniformity in observation between individuals doing the core logging. All core was photographed. No core orientation work was carried out.

10.6.2 Chip Logging

Mine geologists logged chips. The chip trays are labelled, and stored in a locked storage container located at the BMSC core management facility. Logging was performed using standardized geological logging codes, with data recorded on hard copy logging forms that were later transferred to electronic format.

Digital backup copies of the geological data are stored at BMSC’s site office, BMSC’s Asmara office, Nevsun’s head office in Vancouver, and at Nevsun’s off-site data storage in Vancouver. All hard copy logs are archived in files, labelled, and stored at BMSC’s site offices.

10.7 Sample Recovery

10.7.1 Bisha

Sample recovery in diamond core in the supergene and primary zone is good, as described in Section 14. Average recovery in the mineralized zones is approximately 97%. Sample recovery in RC grade control drilling is approximately 80%.

10.7.2 Northwest

Sample recovery in diamond core in the primary zone is good, as described in Section 14. Average recovery in both the copper- and zinc-mineralized zones is approximately 97%. Sample recovery in the oxide and supergene is poor, averaging some 35%. This is inadequate for estimation of Mineral Resources above the category of Inferred. Additional drilling is planned to rectify the situation.

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10.7.3 Hambok

Sample recovery in diamond core in the primary zone is good, as described in Section 14. Average recovery in both the copper- and zinc-mineralized zones is approximately 98%. Sample recovery for the RC programs was estimated by Sanu to be approximately 60% overall. Additional drilling is planned to further assess this situation. All oxide material was classified as Inferred in this estimate.

10.8 Collar Surveys

The drill collars were surveyed using a Trimble Pro-Mark 2 Global Positioning System (GPS) instrument consisting of a base station and rover unit with a radio link. The GPS unit is capable of sub-metre accuracy.

All Bisha, Harena, Northwest, and Hambok (2013 campaign), hole collars are routinely surveyed using this method. BMSC completed a resurvey of the holes drilled by Sanu at Hambok where the original survey was by hand-held GPS only. More than 90% of the collar locations were found. Original easting and northing were found to be reasonably accurate, with only discrepancies found in elevation. Where holes were not located, their elevation value was obtained by interpolation onto a digital terrain surface created from the resurveyed collars.

10.9 Down Hole Surveys

10.9.1 Bisha

Down hole survey methods for Bisha holes in early campaigns used acid tests, Eastman single-shot cameras, or Reflex instrumentation. Typically, measurements were taken at an initial 20 m depth down the drill holes, and subsequently every 50 m thereafter, unless drill hole conditions dictated otherwise. Single-shot surveys were dominant, with an Eastman single shot or Reflex EZ-Shot camera used by the drill contractor until the end of the 2012 drilling campaign. During 2013, the survey method changed to a gyroscopic down hole survey method. The holes surveyed by gyro were B-512 and the BEXT and BNDD series.

10.9.2 Northwest

Surveys were completed using an Eastman single shot or Reflex EZ-Shot camera operated by the contractor until March 2013. Typically, measurements were taken at an initial 20 m depth down the drill holes, and subsequently every 50 m thereafter, and at the end of the hole. From hole NWDD162, gyroscopic methods were applied. An HSHA (high-speed high accuracy) north-seeking gyro probe manufactured by Gyro Technologies Ltd of Tewkesbury, England, was supplied and used by Well Force International contractors. Using this method, survey measurements were taken every 10 m down hole, and then 20 m up hole, providing QA/QC for the survey.

In the creation of a local grid at Northwest, all surveys were converted by the addition of 45° to local azimuths in the BMSC acQuire database management system.

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10.9.3 Harena

The first 18 drill holes completed at the Project have no down hole surveys. Later drill holes were surveyed using acid tests, Eastman single shot, or Reflex EZ-Shot instrumentation. Typically, measurements were taken at an initial 20 m depth down the drill holes, and subsequently every 50 m thereafter, unless drill hole conditions dictated otherwise.

10.9.4 Hambok

Sanu is reported to have used a combination of Eastman single shot or Reflex EZ-Shot instrumentation during their drilling campaigns, operated by the drilling contractors. Typically, measurements were taken at 50 m down hole, and a collar value provided. All holes drilled by BMSC used the gyroscopic method, measuring at 10 m intervals down hole.

10.10 Hydrological Drilling

Holes have been completed for water wells and for groundwater level and flow monitoring at Bisha, Harena, and Northwest. No work has been completed at Hambok.

Drilling was completed by Eritrean Core Drilling Company (ECDC), a local Asmara contractor, in 2003, 2011, and 2012. Major also undertook water bore drilling during 2004. In 2013, further drilling was undertaken by the Colonnade Mining Group (Eritrea) Ltd (Colonnade).

10.10.1 Bisha Deposit

Some 23 holes were completed during the development stages of Bisha. During 2013, an additional hydrogeological drilling program of six 200 m-deep holes was completed by Colonnade. The work was monitored by Knight Piésold (South Africa) (KP SA), who also carried out subsequent pump testing and reporting.

10.10.2 North West Bisha

During 2013, four dewatering/testing holes were planned and drilled around the deposit to make an initial assessment of the area. Drilling of the 200 m-deep holes was completed by Colonnade. The work was monitored by KP SA, who also carried out subsequent pump testing and reporting.

10.10.3 Harena

Two dewatering/testing holes were completed in 2011, with a further five 140 m-deep holes drilled in 2012. Drilling was undertaken by ECDC. The work was monitored by KP SA, who also carried out subsequent pump testing and reporting.

10.10.4 Hambok

No hydrological drilling work has been completed at Hambok to date.

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10.11 Geotechnical Drilling

Thirty-two diamond drill holes were completed for geotechnical evaluations on the Bisha deposit from 2005 to 2012, four diamond drill holes for the Harena deposit in 2012, and eight for the Northwest deposit in 2012. No geotechnical drilling has been undertaken at Hambok.

All drilling was completed by Boart Eritrea. Core orientation was undertaken using the Spear system in 2005, the Reflex ACT I system in 2009 and 2011, and the Reflex ACT II system in 2012. BMSC geologists trained by geotechnical consultants from BGC Engineering Inc. (Vancouver) (BGC) and Geologica (South Africa) performed geotechnical core logging. Data from the logging includes total core recovery, rock quality designation (RQD), number of natural discontinuities, intact rock strength, and weathering grade. Detailed observations regarding the orientation, roughness, and infilling of any measured discontinuities were also collected. The data from the drilling programs has been organized and compiled into the geotechnical tables of the acQuire database.

BGC completed additional off-site testwork and analysis. The results of this work are discussed in Section 16.

10.12 Metallurgical Drilling

Forty drill holes have been completed for metallurgical testwork on the Bisha deposit, and two drill holes each on the Northwest and Harena deposits. The drilling is summarized in Table 10‑5 to Table 10‑7. Results of the metallurgical drilling and testwork are discussed in Section 13.

Table 10‑5: Bisha Deposit Metallurgical Drill Hole Summary

Year	Phase	Range of Hole No.	No. of Holes	Drilled Length (m)
2005	II	MET-05-01 to MET-05-08	8	1,212.5
2009	-	MET-009 to MET-011	3	224.0
2010	I	MET-012 to MET-017	6	478.0
2011	II	MET-018 to MET-019	2	180.0
2012	-	MET-020 to MET-031	12	820.7
2013		BMT-032 to BMT-041	9	962.0
Total			40	3,877.2

Table 10‑6: NW Bisha Deposit Metallurgical Drill Hole Summary

Year	Phase	Range of Hole No.	No. of Holes	Drilled Length (m)
2012		NWMT001 to NWMT002	2	471.0
Total				471.0

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Table 10‑7: Harena Deposit Metallurgical Drill Hole Summary

Year	Phase	Range of Hole No.	No. of Holes	Drilled Length (m)
2011	II	HMET001 to HMET002	2	1,200.9
Total				1,200.9

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11 Sample Preparation, Analyses, and Security

BMSC staff has been involved with, or are responsible for, the following:

sample collection
core splitting
sample preparation of geochemical, pit, trench, RC, core, and grade control samples
delivery of samples to the analytical laboratory
density determinations
sample storage
sample security.

In this report sample preparation, analysis, quality control and security are described in detail only for sampling since the last Technical Report, with summaries of or reference to older data, as appropriate. Detail regarding historical samples can be found in the AGP, 2012; Bisha Polymetallic Operation, Eritrea, Africa NI 43-101 Technical Report for Nevsun Resources. Where BMSC has relied on samples taken and processed by Sanu Resources (Sanu), relevant extracts from Sanu reports are presented by way of information.

11.1 Sample Collection and Preparation

11.1.1 Trench Sampling

Trench sampling was used in the estimate of Inferred Mineral Resources for the Northwest deposit. The trench locations and the results obtained are given in Section 9.

The trenches were excavated to a depth of 0.5 m to 2.0 m depending on the difficulty of excavation or breaking the rock. The trenches were mapped and then sampled, with samples taken at 2 m intervals breaking at lithological contacts. 480 samples were taken in total. Trench samples were analysed by ALS and preparation, despatch and analysis are as described below for drill core.

11.1.2 Diamond Core Sampling

BMSC

For holes drilled in all the Northwest campaigns and BMSC 2013 drilling at Bisha and Hambok, holes were sampled for their full length at a target length of 1.0 m per sample. Sample intervals vary based upon mineralogical and lithological contacts. The logging geologist sets out the sample regime.

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Standard diamond cutting blades flushed with fresh water are used to halve the core. Highly broken core pieces are cut along the axis if possible or the core is split using a trowel down the middle of the tray row and handpicked or scooped to ensure representative samples are obtained. Cutting lines may be drawn on the core. The remaining half core is returned to the box and boxes stacked in numerical order by hole.

The technicians or samplers under the supervision of technicians place half of the core in individual trays laid out in numerical order. Samples are then placed in a drying oven for 12 to 18 hours at between 80°C and 100°C. The samples then pass through a jaw crusher to >75% of sample passing 10 mesh (2 mm) screen. The sample is then split using a Jones-type riffle splitter to achieve a sub-sample weight of approximately 200 g to 300 g. The sample reject is then bagged, labelled with the original sample ID and put into storage at Bisha site. The sub-samples are packed and then placed in large plastic shipping barrels. When samples are ready to be shipped the samples lists are combined with a sample submission form and enclosed with the samples in plastic drums. Samples are despatched to Asmara in BMSC vehicles with BMSC staff and are held at the BMSC Head Office before onward despatch.

ALS in Vancouver carried out analysis of Northwest samples. In this example, the sample information with required analytical procedures is emailed to ALS so that the sample shipment can be tracked and the laboratory is made aware of the pending arrival of samples. The sample barrels are submitted to the Eritrean Ministry of Mines for inspection and submission to customs. A seal is placed on the barrels and the sample barrels are shipped via air transport directly to ALS in Vancouver. At ALS the samples are pulverised to better than 85% passing 75 µm.

Genalysis carried out analysis for Hambok samples for consistency of analysis with the Sanu campaigns. In this example, the plastic barrels are taken by road to their Horn of Africa preparation facility near Asmara. Here, samples are pulverised and the pulps despatched by courier to the Genalysis Laboratory in Perth, West Australia, for analysis. Pulverization is to better than 85% passing 75 µm.

For holes drilled at Bisha and Harena that are used in the current Mineral Resource, sampling methods and preparation were the same as those described above, excepting that for the most part holes were sampled only through the mineralized zone.

Sanu Resources

According to available reports, Sanu’s sampling procedures up to the end of 2009 were as follows:

“All aspects of sample preparation up to the time of submission to the Horn of Africa Sample Preparation Lab in Asmara, including core cutting, sampling, bagging, numbering, and insertion of duplicate sample tags, was conducted by Sanu employees or consultants hired and managed directly by Sanu.”

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“Sampling was started at the geologic contact between the hanging wall and the sulfides (unless there was significant disseminated sulfides in the wall rocks), and continued in 1 m intervals; the first interval being either slightly longer or shorter that 1 m. The core was cut lengthwise in‐two (split) by a diamond saw. One‐half of the cut core was again cut lengthwise (quartered) for the sample to be analyzed. One‐quarter of the core was packed in plastic bags and sent to the Genalysis preparation facility in Asmara, leaving ¾ core in the core box.”

“Crushing and splitting of the sampled quarter‐core was conducted by the Horn of Africa Preparation Laboratory in Asmara. The Horn of Africa Preparation Laboratory was installed with the assistance of Genalysis Laboratories in Perth, Australia. Drill core samples were processed in the following steps:

The samples, each 1 m section of quarter core weighing approximately 1.8 kg (assuming a specific gravity of 4.5, 47.6 mm core diameter, and (2 mm x 47.6 mm x 2) loss due to core cutting), were sorted and ordered numerically after receipt
The samples were placed in a drying oven for 12 to 18 hours at 80°‐100°C
The samples were crushed to >75% of the sample passing 10 mesh or 2 mm screen
For drill holes HAM‐001 through HAM‐019, the entire sample was pulverized in a single ring and puck style bowl in a LM2 pulverizing machine with C1000 (1 kg) or C2000 (2 kg) bowl to >85% passing less than 75 μm
As outlined below, fine preparation duplicate samples were separated at this stage for drill holes HAM‐001 through HAM‐019
For drill holes HAM‐020 through HAM‐061, the samples were crushed only
As outlined below, coarse preparation duplicates were separated at this stage for drill holes HAM‐020 through HAM—061
The pulverized or crushed samples were packed in plastic bags and shipped to Genalysis in Australia.”

From 2010 onwards, the procedures were as above but with half core samples and half core remaining in the core boxes. BMSC can attest to this.

11.1.3 Grade Control RC Sampling

RC grade control samples from drilling in the Bisha pit are sampled at 1 m intervals. Following riffle splitting the sample is sent to the SGS on-site laboratory. The samples are placed in a drying oven for 12 to 18 hours at between 80°C and 100°C. The samples then pass through a jaw crusher to >75% of sample passing 10 mesh (2 mm) screen. The sample is then split using a Jones-type riffle splitter to achieve a sub-sample weight of approximately 200 g to 300 g. The samples are then pulverised to better than 85% passing 75 µm. A subsample is also retained for additional analytical work.

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11.1.4 Bulk Density Sampling

To determine bulk density, BMSC took and used two types of sample, namely drill core and rock samples. The latter are taken from within the open pit (pit floor and mining faces) and the ROM stockpiles. The sample locations are recorded by the survey department. A total of 315 samples were taken in the latest program to determine bulk densities for the supergene mineralized zone.

Bulk density for all samples is determined using Archimedes principle. Rock samples are dried and then wrapped in foil prior to weighing in air and water. BMSC recognise that the sample collection method has potential bias but all efforts are made to avoid this. The work completed to date has successfully addressed reconciliation of ROM tonnages, and densities determined by this method have been used in the resource estimation process.

11.2 Sample Analysis

11.2.1 Analytical Laboratories

Core and RC Samples - BMSC

A number of different laboratories have been used by Nevsun/BMSC over time as shown in Table 11‑1. Sanu submitted all their samples to Genalysis as described in Section 11.1.2.

Table 11‑1: Analytical Laboratories used for drill samples

Period	Crushing	Pulverization	Analysis
2002 to 2003 Ph I	ALS	ALS	ALS
2003 Ph II to 2006	On site	ALS	ALS
2009 to 2010 Ph I	Horn of Africa (Genalysis)	ALS Romania	ALS (Romania for Au)
2010 Ph II to present	On site	ALS SGS on-site	ALS – drill core SGS – RC Grade Control
2013 (Hambok)	On site	Horn of Africa (Genalysis)	Genalysis (Perth)

ALS established an on-site sample preparation facility in July 2003, which was a containerized facility purchased from ALS who also reviewed laboratory start-up and provided training. From September 2003 to 2006 splits from the core and RC drilling were crushed to -2 mm and sent to ALS in Vancouver for subsequent pulverization and analysis.

From 2009 and for Phase I of 2010 work, BMSC sent drill core to the Genalysis managed Horn of Africa preparation facility near Asmara. Samples were couriered to ALS Romania for gold analysis. The Romanian laboratory then sent the samples to ALS in Vancouver for multi-element analysis including base metals.

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In October 2010, SGS re-furbished the on-site prep facility and on-site crushing continued prior to samples being sent to ALS in Vancouver for subsequent pulverization and analysis. SGS managed the on-site laboratory using BMSC staff employing SGS standard analytical procedures. In September 2013, SGS took full control of management and staffing of the on-site laboratory, effectively becoming one of the SGS group laboratories.

BMSC discontinued using ALS in Romania for gold analysis after determining there was no benefit to cost or schedule.

ACME Laboratory (ACME), in Vancouver, Canada, performed check analyses on pulp duplicate materials.

Core and RC Samples – Sanu Resources

According to available reports, “the majority of analyses from Hambok were conducted at Genalysis Laboratory Services Pty Ltd in Perth, Australia. “The National Association of Testing Authorities Australia has accredited Genalysis Laboratory Services Pty Ltd, following demonstration of its technical competence, to operate in accordance with ISO/IEC 17025, which includes the management requirements of ISO 9001: 2000” (quote from Genalysis website, October 2008).

ALS in Vancouver analyzed a portion of duplicate samples from Hambok. “The Vancouver facility operates under ALS Laboratory Group's global Quality Management System and is in compliance with ISO 9001:2000 for the provision of assay and geochemical services according to QMI Management Systems Registration. The laboratory has also been accredited to ISO 17025 standards for specific laboratory procedures by the Standards Council of Canada (SCC)” (quote from ALS website, October 2008).”

Laboratory Certification

ALS, Genalysis, SGS, and ACME are ISO-registered and are internationally recognized analytical facilities that are independent of Nevsun and BMSC.

11.2.2 Analytical methods

The analytical methods used by each laboratory for core and RC samples are summarized in Table 11‑2.

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Table 11‑2: Assay Laboratory Analytical Techniques

Laboratory	Analyte	Method Code	Method	Description	Detection Limit	Comments
ALS Vancouver (Primary Laboratory)	Au	Au-AA23	Routine Gold Fire Assay	30 g, fire assay, AAS finish	0.005 – 100 ppm Au
	Au	Au-GRA21	Over-limit gold >100 ppm	30 g fire assay, Gravimetric finish	0.05 – 1,000 ppm Au
	Cu, Pb, Zn, Ag, As, Fe (41 elements)	ME-ICP41a	Multi-element Aqua Regia Partial Digestion	1 g, two acid (HN0₃, HCl) digest, ICP-AES analysis	5 – 50,000 ppm Cu 10– 50,000 ppm Pb, Zn	Multi-element scanning
	Cu, Pb, Zn	OG46	Over-limit Cu, Pb, Zn	1 g, two acid (HN0₃, HCl) digest, ICP-AES analysis	0.001 – 20% Cu, Pb 0.001 – 30% Zn	Analytical method modified for higher precision
SGS Bisha (Onsite Laboratory)	Au	FAA303	Routine Gold by Fire Assay	30 g, fire assay, AAS finish	0.01 - 100
	Au	FAA505	Over-limit gold >100ppm	30 g fire assay, Gravimetric finish	0.05 – 1,000 ppm Au
	Cu, Zn	AAS12B	Aqua Regia Partial Digestion	0.25 g*, two acid (HN0₃, HCl) digest, ICP-AES analysis	0.5 – 10,000 ppm Cu 1 – 10,000 ppm Zn 2 – 10,000 ppm Pb	Multi-element scanning
	Cu, Zn	AAS23C	Over-limit Cu, Zn	0.25 g*, two acid (HN0₃, HCl) digest, ICP-AES analysis	unknown	Analytical method modified for higher precision
Genalysis Perth (Umpire laboratory; Primary laboratory for Hambok, BMSC)	Cu, Pb, Zn	AR01_OE01	Multi-element Aqua Regia Partial Digestion	1 g, two acid (HN0₃, HCl) digest, ICP-OES analysis	1 ppm – 2% Cu, Zn 1 – 5,000 ppm Pb
	Cu, Pb, Zn	4AH_OE01	Over-limit Cu, Pb, Zn	1 g, four acid (HN0₃, HCl, HClO_4,HF) digest, ICP-OES analysis	unknown	Analytical method modified for higher precision
Genalysis Perth (Primary laboratory Sanu)	Au	FA50/AAX	Routine Gold by Fire Assay	50 g, fire assay, AAS finish	Not reported
Genalysis Perth (Primary laboratory Sanu)	Cu, Pb, Zn	AX/OES	Multi-element multi acid digestion	Multi acid digest (not defined), ICP-OES analysis	Not reported

Note: *Assay aliquot size changed in 2013 from 0.25 g to 1 g

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11.3 Quality Assurance and Quality Control

11.3.1 Nevsun/BMSC QA/QC Protocols for Drill Programs, 2003–2005

Since drill programs before 2003 did not encounter significant mineralization, QA/QC programs were not well established.

From 2003 onwards, all of the core and RC drilling programs included certified reference materials (CRMs) and also included blanks, twin sample duplicates, and coarse preparation duplicates. Each drill program report documented the protocols and results of the QA/QC program. Nevsun purchased the CRMs from Geostats Pty Ltd (Geostat), a well-known and certified reference material manufacturer, located in Australia.

The QA/QC sample insertion protocol employed by Nevsun for all core and RC drill sampling subsequent to the 2002 program was as follows:

six certified CRM control samples per 100 samples; three gold (B, D, and F) and three base metal (A, C, and E)
one coarse blank sample of barren material per 100 samples plus the random insertion of barren material in mineralized zones
one quartered core “twin” duplicate sample per 100 samples
two coarse preparation duplicates per 100 samples.

Summaries of Program Results

2003/2004 (304 DDH)

QA/QC samples comprised 11% of the sample stream. After some consideration, AMEC considered that any bias in the CRM was acceptable. Some problems were encountered with the ‘blank’ material as described below in the summary for 2005. For these reasons, sensible conclusions regarding cross-contamination were not possible. Twin duplicate analysis showed an acceptable level of precision when inherent nugget effect and values close to the detection limit were taken into account and AMEC concluded that sample variance was within an acceptable range. Coarse duplicate analysis was also regarded as within acceptable ranges.

2005 (121 DDH)

QA/QC samples comprised almost 11% of the sample stream. By virtue of analysis of control charts and accuracy plots (average vs. best values) AMEC considered that any bias values shown by the CRM were acceptable. Twin duplicate analysis showed an acceptable level of precision and AMEC concluded that sampling variance for the 2005 campaign was satisfactory. Coarse duplicate analysis was also regarded as within acceptable ranges.

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The coarse blank material was sourced from near the Project. This material usually consists of limestone and/or dolomite, considered to consist of barren rock without any appreciable precious metal or base metal content. After review of the logs, sample batches, and data, AMEC noted that the blank material is not barren, and thus the true values of the blank for each metal are not known. AMEC recommended that Nevsun purchase a commercial blank for use. If the use of a coarse blank material was continued, then AMEC suggested it should be of a clean, barren material with no obvious oxidation surfaces or patches iron oxides such as limonite or hematite.

11.3.2 BMSC QA/QC Protocols for Drill Programs, 2006–2010

All of the drilling programs included certified reference materials (CRMs) and also included blanks, twin sample duplicates, and coarse preparation duplicates. A limited number of holes were drilled in this period.

BMSC used the same CRMs purchased from Geostat as those used in the 2002-2005 drilling campaigns. The reference material includes a range of low-grade, mid-grade, and high-grade precious and base metal standards with certified values showing statistically-acceptable confidence limits.

Summaries of Program Results

2006 (10 DDH)

QA/QC samples comprised almost 11% of the sample stream. AMEC’s review of CRM results found the analytical precision to be satisfactory. No pulp or external check samples were submitted.

2009 (17 DDH)

QA/QC samples comprised 15% of the sample stream; all holes were from Harena. AMEC’s review of CRM results found the analytical precision to be satisfactory. Sub-sampling precision from coarse duplicates was not acceptable, but the number of samples included was very low.

2010 (44 DDH)

QA/QC samples comprised 15% of the sample stream; most holes were from Harena. AMEC’s review of CRM results found the analytical precision to be satisfactory. Sub-sampling precision from coarse duplicates was again not acceptable. AMEC recommended investigation of the sample preparation work and a program of umpire analysis. The author has found no record of this recommended work being completed.

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11.3.3 BMSC QA/QC Protocols for Drill Programs, 2011

2011 (105 DDH)

BMSC conducted a QA/QC program that included the insertion of certified reference materials (CRMs), blanks, twin sample duplicates, and coarse preparation duplicates. AGP reviewed the compiled QA/QC dataset. In total, the QA/QC samples comprise 16% of the total sample analyses.

For consistency, BMSC used the same CRMs purchased from Geostat as those used for previous drilling campaigns. The reference material includes a range of low-grade, mid-grade, and high-grade precious and base metal standards with certified values showing statistically-acceptable confidence limits.

In their review, performance of the CRMs was considered acceptable by AGP. In total 619 CRMs were submitted into the sample stream at a rate of 1 insertion per 15 samples. Twelve gold standards were used; however, 90% of the insertions came from six standards with grades representative of range of grades found in the deposit. AGP considered that in general the standards performed well.

In that the current Resource estimate is dominated by copper and zinc, BMSC did not make any additional cross-checks for gold analysis carried out in 2011. During 2012 and 2013 the AcQuire database at the Bisha site was formed, with the accompanying validation and rigour, as described in Section 11.4. As a result of this work BMSC was able to undertake an additional review of copper and zinc data that is given below.

Blanks

The performance of blanks for copper and zinc showed a failure rate of 9% in both cases. This can be attributed to the persistent problem of the nature of the material used for ‘blank’ analysis.

Certified Reference Material

Various CRMs (assay standards) as noted above were used to monitor the accuracy and precision of the assay laboratory and detect any possible bias in the reported analytical values. The method of assessment used here is as described in Section 11.3.4.

The results are summarized in Table 11‑3 and

Table 11‑4. The performance of the CRMs for copper and zinc assays is acceptable with only minor negative bias.

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Table 11‑3: Copper – CRM Performance Summary, ALS Laboratory Vancouver, 2011

CRM	Certified Value (ppm)	Std. Dev.	Lab. Method	No. Assays	Accuracy Test	Precision Test	% Passing 2SD	% Bias	Period in Use
GBM398-4	3891	195	ME-ICP41a	112	PASS	PASS	99%	1%	2011
GBM309-15	34390	1232	ME-ICP41a	51	PASS	PASS	100%	-2%	2011
GBM900-10	151370	7949	OG46	72	PASS	PASS	100%	-5%	2011
GBM996-7	23483	1456	ME-ICP41a	28	PASS	PASS	100%	-1%	2011

Table 11‑4: Zinc – CRM Performance Summary, ALS Laboratory Vancouver, 2011

CRM	Certified Value (ppm)	Std. Dev.	Lab. Method	No. Assays	Accuracy Test	Precision Test	% Passing 2SD	% Bias	Period in Use
GBM398-4	5117	229	ME-ICP41a	112	PASS	PASS	93%	1%	2011
GBM996-7	110344	6018	OG46	28	PASS	PASS	100%	-5%	2011

Duplicates

The results for pairs of duplicate samples (original and duplicate) were analysed as described in Section 11.3.4 and result shown in Table 11‑9. The filtered field duplicates for Cu and Zn analyses have an estimated average coefficient of variation (ACV) of less than 15%, which is within the accepted tolerance levels for field duplicates for both these elements. No bias was apparent in the field duplicate analyses.

In summary, BMSC concludes that the overall accuracy and precision of the assay data for copper and zinc relating to the CRMs is within tolerance limits, and no obvious trend or major bias is apparent within the primary assay data. The minor bias does require further investigation but is not considered material for the Mineral Resource estimate.

11.3.4 BMSC QA/QC Protocols for Drill Programs, 2012-2013 Diamond Core

202 DD holes were drilled in 2013-2013, with the majority at the Northwest deposit. BMSC conducted a QA/QC program that included the insertion of certified reference materials (CRMs), blanks, twin sample duplicates, and coarse preparation duplicates. For consistency, BMSC used the same CRMs purchased from Geostat as those used for previous drilling campaigns. A detailed review was made independently by Cube and is, reproduced in part in this section.

Blanks

The source of coarse blank material is as described above. Assays for the blank material were assessed by graphing the actual assayed value and the maximum accepted value, which was assigned as 50 ppm for Cu, Zn and 0.05 ppm for Au. A maximum accepted value of 10 times the lower analytical detection limit has been used for each of the elements, and removes the potential for bias and precision issues which increase close to the assay method detection limit. Blanks should return a value less than the accepted value at least 95% of the time.

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The results are summarized in Table 11‑5 and indicate that the assay blanks data are within acceptable limits. No obvious contamination issue is apparent within the primary assay data. Failures are, as described above, generally attributable to the nature of the blank material.

Table 11‑5: Blank Performance Summary

Laboratory	Analyte	No. of Samples	No. of Samples Fail	% Passing
ALS Vancouver	Cu	919	7	99.2
	Zn	980	11	98.9
	Au	967	0	100

A total of 982 assay blank samples were inserted into the sample stream which comprises approximately 3% of the total assay samples submitted.

Certified Reference Material

Various CRMs (assay standards) were used to monitor the accuracy and precision of the assay laboratory and detect any possible bias in the reported analytical values. The performance of the CRM sample data was assessed by plotting the replicate assay values of the CRMs against time on the control charts. Good quality analysis of the CRMs will be characterised by a random distribution of data points around the certified mean value, with 95% of the data points lying within two standard deviations of the mean (Abzalov, 2008). If more than 5% of the CRM’s submitted are outside three standard deviations of the certified mean value, then corrective action should be taken. In addition, no trends or significant bias should be observed in the control charts.

Any obvious assay “outliers” that are likely to be the result of sample mishandling or transcription errors were removed from the dataset prior to analysis, to avoid any skewing of the dataset.

In addition, the CRM data set was assessed using two statistical tests to demonstrate that the analytical accuracy and precision of the assays were comparable to the certified value and standard deviation of the CRM, and considered acceptable within the 95% confidence limit (Abzalov, 2008).

A total of 1,696 CRM samples were inserted into the sample stream which comprises 5% of the total samples. All CRMs with less than five replicate assays were considered as having insufficient data to assess the performance of the CRM and were excluded from the analysis.

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The results are summarized in Table 11‑6, Table 11‑7 and Table 11‑8. In general, a bias of greater than 5% is considered not acceptable.

The performance of the CRMs for copper assays is generally acceptable, with one CRM GBM906-16 showing successive replicate assays outside of two standard deviations. The majority of the other CRM replicate assays show a small but consistent negative bias of 1% to 5%. The poor performance of GBM906-16 is not replicated by other CRM in the same batches. All other copper CRMs performed within acceptable limits for all grade ranges over extended time periods and demonstrate that there is no systemic analytical accuracy issue for copper at the ALS laboratory.

The performance of the CRMs for zinc assays are generally acceptable but show the same consistent negative bias of 1% to 5% as demonstrated by the copper CRM replicate assays (Table 11‑7).

All of the gold CRMs returned replicate assay results within the acceptable limits for analytical accuracy for the ALS Vancouver laboratory (Table 11‑8).

Both the copper and zinc CRM replicate assays from ALS show a consistent negative bias in the range of 1% to 5% in this period. Although the bias is noticeable, it is not considered significant or material to the mineral resource estimate.

Table 11‑6: Copper - CRM Performance Summary, ALS Laboratory Vancouver

CRM	Certified Value (ppm)	Std. Dev.	Lab Method	No. Assays	Accuracy Test	Precision Test	% Passing 2SD	% Bias	Period in Use
GBM301-9	2,881	196	ICP41a	106	PASS	PASS	100%	-4%	2012-2013
GBM304-11	104,011	3,070	OG46	93	PASS	PASS	99%	-1%	2013
GBM309-14	28,025	1,176	ICP41a	70	PASS	PASS	93%	-1%	2012-2013
GBM309-15	34,390	1,232	ICP41a	122	PASS	PASS	97%	-3%	2012-2013
GBM310-15	237,854	8,084	OG46	3	NA	NA	100%	1%	2012
GBM398-4	3,891	195	ICP41a	35	PASS	PASS	97%	1%	2003-2013
GBM900-10	151,370	7,949	OG46	49	PASS	PASS	100%	-2%	2003-2012
GBM906-16	106,807	4,617	OG46	22	FAIL	FAIL	23%	16%	2012
GBM907-13	16,853	614	ICP41a	175	PASS	PASS	96	-1%	2012-2013
GBM908-11	177,033	5,807	ICP41a	2	NA	NA	100%	-3%	2012
GBM908-15	50,027	1,515	ICP41a	85	PASS	PASS	94%	-3%	2013
GBM909-14	21,898	842	ICP41a	191	PASS	PASS	98%	-2%	2012-2013
GBM995-8	264	23	ICP41a	48	PASS	PASS	100%	-3%	2012-2013
GBM996-7	23,483	1,456	ICP41a	38	PASS	PASS	100%	-1%	2003-2013
Total	14 CRMs			1,072					2003-2013

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Table 11‑7: Zinc – CRM Performance Summary, ALS Laboratory Vancouver

CRM	Certified Value (ppm)	Std. Dev.	Lab. Method	No. Assays	Accuracy Test	Precision Test	% Passing 2SD	% Bias	Period in Use
GBM301-9	7,208	433	ICP41a	105	PASS	PASS	98%	-5%	2013
GBM309-14	227,192	10,735	OG46	66	PASS	PASS	100%	-1%	2012-2013
GBM309-15	123,691	3,266	OG46	122	PASS	PASS	77%	-4%	2012-2013
GBM310-13	108,471	4,824	OG46	66	PASS	PASS	100%	-1%	2012-2013
GBM310-15	11,931	499	OG46	3	NA	NA	100%	-2%	2012
GBM398-4	5,117	229	ICP41a	35	PASS	PASS	94%	-2%	2003-2013
GBM900-10	25,632	1,607	ICP41a	48	PASS	PASS	94%	0%	2003-2012
GBM906-16	4,783	418	ICP41a	22	PASS	PASS	96%	3%	2012
GBM907-13	66,270	2,937	OG46	175	PASS	PASS	99	-2%	2012-2013
GBM908-11	23,604	1,104	ICP41a	2	NA	NA	100%	-3%	2012
GBM909-14	65,582	2,291	OG46	191	PASS	PASS	99.5%	-2%	2012-2013
GBM995-8	124,308	8,386	OG46	25	PASS	PASS	72%	-8%	2012-2013
GBM996-7	110,344	6,018	OG46	33	PASS	PASS	100%	-2%	2003-2013
Total	13 CRMs			895					2003-2013

Table 11‑8: Gold – CRM Performance Summary, ALS Laboratory Vancouver

CRM	Certified Value (ppm)	Std. Dev.	No. Assays	Accuracy Test	Precision Test	% Passing 2SD	% Bias	Period in Use
G300-8	1.07	0.06	5	NA	NA	100%	1%	2013
G308-3	2.50	0.11	4	NA	NA	100%	0%	2013
G396-2	0.12	0.03	8	PASS	PASS	100%	-14%	2012
G396-3	4.82	0.29	6	PASS	PASS	100%	0%	2013
G396-5	7.36	0.29	84	PASS	PASS	98%	2%	2012-2013
G396-6	13.5	0.5	6	PASS	PASS	100%	0%	2003
G397-2	4.49	0.18	1	NA	NA	100%	0%	2013
G399-2	1.46	0.09	1	NA	NA	100%	3%	2004
G399-4	4.78	0.21	172	PASS	PASS	94%	0%	2005-2013
G399-6	2.52	0.14	131	PASS	PASS	99%	2%	2005-2013
G900-10	13.85	0.53	32	PASS	PASS	91%	-3%	2005-2013
G901-1	2.58	0.13	2	NA	NA	100%	0%	2006-2012
G909-4	7.52	0.3	26	PASS	PASS	85%	2%	2013
G995-3	2.66	0.14	10	PASS	PASS	100%	-3%	2003-2005
G999-7	2.52	0.12	5	NA	NA	100%	2%	2013
Total	15 CRMs		493					2003-2013

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Given the identified CRM replicate assay issues, the overall accuracy and precision of the assay data relating to the CRMs is within tolerance limits, and no obvious trend or major bias is apparent within the primary assay data.

Duplicates

The results for pairs of duplicate samples (original and duplicate) are plotted as X/Y scatter plots and Relative Difference Plots (RDP). Scatter plots allow for direct comparison of the data pairs and the assessment of general dispersion, data regression, as well as the presence of any outliers. RDP plots evaluate the coefficient of variation (CV) for each data pair (the difference between pairs relative to the pair mean) and allow the measurement of the relative precision error between duplicate pairs based on the Average Coefficient of Variation (ACV). The ACV can be used as a universal measure of relative precision (Abzalov, 2008).

Overall, the precision of the duplicate samples was within acceptable limits and no obvious trend or bias was identified. A total of 2,291 duplicate samples (for ALS) were collected which represents an insertion rate of 6.4%, and is within the acceptable range of an insertion rate for duplicate samples of 5-10%. The results for each type of duplicate samples are summarized in Table 11‑10. A discussion of the results follows the tables below.

Field check sampling duplicates (QC Code DUP) were collected from either a quarter cut diamond core or a crushed core sub-sample and inserted sequentially within the sample stream at a nominal rate of 1 in 100 samples. A total of 1,550 field duplicates were submitted to the primary ALS laboratory, which comprises 4.3% of the total sample analyses.

Analysis of the field duplicates quantifies the total sampling error from sampling collection and preparation to the assay process. The filtered field duplicates for Cu and Zn analyses have an estimated ACV range of between 13% and 15%, which is within the accepted tolerance levels for field duplicates (15% for Cu/Zn). Gold analyses have an ACV of 16%, which is below the acceptable precision limit of 30% for Au. No bias was apparent in the field duplicate analyses.

Coarse reject check sampling (preparation) Duplicates (QC Code Coarse Dup) are a 200 g to 300 g sub-sample riffle splits from the crush rejects and inserted sequentially into the sample stream (in batch).

A total of 584 coarse split duplicates were analysed by the ALS laboratory and comprise 1.6% of the total sample analyses. Comparison of the original assays against the replicate values allows assessment of the precision error associated with the first sub-sample split during sample preparation. The overall estimated precision of the copper and zinc analyses is 9% ACV, which is within the accepted tolerance level for field duplicates. Gold duplicates assays have an ACV of 16% which is below the acceptable precision limit of 30% for Au. No significant bias was apparent in the coarse reject duplicate analyses.

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Table 11‑9: Duplicate Sample Performance Summary, ALS laboratory Vancouver, Diamond Core, 2011

Description	QC Code	Laboratory	Analyte	No. of Samples	No. Filtered Samples >1,000 ppm Cu, Zn >0.1 ppm Au	Outliers Removed >100% Absolute Relative Difference	Absolute Relative Difference	For Filtered Data, Outliers Removed				Comments
								ACV%	Assays within 10%	Assays within 20%	Assays within 50%
Field Duplicate	Dup.	ALS - Vancouver	Cu	753	286	11	19%	13%	63%	82%	96%	Acceptable Precision, no bias
			Zn	706	235	8	17%	12%	65%	85%	96%	Acceptable Precision, no bias

Table 11‑10: Duplicate Sample Performance Summary, ALS laboratory Vancouver, Diamond Core, 2012 – 2013

Description	QC Code	Laboratory	Analyte	No. of Samples	No. Filtered Samples >1,000 ppm Cu, Zn >0.1 ppm Au	Outliers Removed >100% Absolute Relative Difference	Absolute Relative Difference	For Filtered Data, Outliers Removed				Comments
Description	QC Code	Laboratory	Analyte	No. of Samples	No. Filtered Samples >1,000 ppm Cu, Zn >0.1 ppm Au	Outliers Removed >100% Absolute Relative Difference	Absolute Relative Difference	ACV%	Assays within 10%	Assays within 20%	Assays within 50%	Comments
Field Duplicate	Dup.	ALS - Vancouver	Cu	1,549	269	14	21%	15%	48%	71%	90%	Acceptable Precision, no bias
			Zn	1,549	257	5	19%	13%	52%	71%	93%	Acceptable Precision, no bias
			Au	1,550	309	0	22%	16%	48%	68%	88%	Acceptable Precision, no bias
Field Duplicate	Dup.	SGS Bisha	Cu	1,077	245	13	20%	14%	61%	77%	94%	Acceptable Precision, no bias
			Zn	1,074	197	11	16%	11%	60%	73%	96%	Acceptable Precision, no bias
			Au	1,050	318	0	33%	23%	36%	48%	75%	Acceptable Precision, no bias
Coarse Split Duplicates	Coarse Dup.	ALS - Vancouver	Cu	583	103	1	13%	9%	75%	91%	99%	Acceptable Precision, no bias
			Zn	584	104	2	13%	9%	73%	92%	97%	Acceptable Precision, no bias
			Au	584	129	0	22%	16%	71%	84%	95%	Acceptable Precision, no bias
Coarse Split Duplicates	Coarse Dup.	SGS Bisha	Cu	506	111	0	19%	13%	69%	86%	94%	Acceptable Precision, no bias
			Zn	506	96	4	13%	9%	78%	91%	99%	Duplicates are 19% higher for assays >2.5% Zn
			Au	485	174	0	40%	28%	30%	47%	71%	Low Precision
												Table continues…

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Description	QC Code	Laboratory	Analyte	No. of Samples	No. Filtered Samples >1,000 ppm Cu, Zn >0.1 ppm Au	Outliers Removed >100% Absolute Relative Difference	Absolute Relative Difference	For Filtered Data, Outliers Removed				Comments
Description	QC Code	Laboratory	Analyte	No. of Samples	No. Filtered Samples >1,000 ppm Cu, Zn >0.1 ppm Au	Outliers Removed >100% Absolute Relative Difference	Absolute Relative Difference	ACV%	Assays within 10%	Assays within 20%	Assays within 50%	Comments
Secondary Coarse Split Duplicate	Sec. Dup.	ALS vs. SGS Bisha	Cu	20,404	3,454	67	24%	17%	46%	72%	94%	SGS are 7% lower for assays <1.5% Cu
			Zn	20,390	3,132	81	22%	16%	46%	74%	95%	SGS are 7% lower for assays <2.5% Zn
			Au	22,432	3,920	0	46%	33%	29%	51%	83%	Low Precision
Umpire Pulp Duplicate	Sec. Dup.	ALS vs. Genalysis	Cu	157	156	1	17%	12%	57%	92%	97%	Genalysis are 11% lower for assays <1.5% Cu
			Zn	148	73	3	17%	12%	56%	73%	99%	Genalysis are 13% lower for assays <2.5% Zn
		SGS Bisha vs. Genalysis	Cu	77	77	3	16%	11%	47%	74%	100%	Genalysis are 8% lower for assays <1.5% Cu
			Zn	77	20	6	29%	21%	21%	57%	86%	Genalysis are 16% lower for all Zn assays
		ALS vs. SGS Bisha	Cu	77	77	3	13%	9%	65%	88%	100%	SGS are 5% lower for assays <1.5% Cu

Table 11‑11: Duplicate Sample Performance Summary, SGS On-Site laboratory, Bisha Mine RC Grade Control, May to September 2013

Description

QC
Code

Laboratory

Analyte

No. of
Samples

No. Filtered Samples
>1,000 ppm Cu, Zn
>0.1 ppm Au

Outliers Removed
>100% Absolute
Relative Difference

Absolute
Relative
Difference

For Filtered Data, Outliers Removed

Comments

ACV%

Assays
within
10%

Assays
within
20%

Assays
within
50%

Field Duplicate

DUP

SGS on-site

30%

21%

29%

61%

90%

Low precision, few samples

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Secondary Coarse Split Check Sampling Duplicates (QC Code Seconds Dup) are routine coarse sub-sample (200 g to 300 g) splits from each crushed sample which was submitted to the SGS Bisha onsite laboratory as a preliminary fast tracked assay. Comparison of these assays against the ALS samples allows a more complete assessment of any bias in the duplicates at the secondary laboratory.

A total of 20,402 secondary coarse duplicate samples were collected and analysed by both laboratories. The duplicated dataset comprises 58% of the total analyses. The filtered field duplicates for Cu and Zn analyses have an estimated ACV range of 16% and 17%, respectively, which exceeds the accepted tolerance levels for field duplicates (15% for Cu/Zn). Gold analyses have an ACV of 33%, which is also above the acceptable precision limit of 30% for Au. A negative bias of 7% for Cu and Zn exists with SGS Bisha analyses consistently reporting lower than the ALS analyses in the grade ranges of Cu 0.1% to 1.5% and Zn 0.1% to 2.5%. No such bias is apparent for the gold analyses.

Umpire Pulp Check Assaying Duplicates (QC Code Seconds Dup) are assays of mineralized pulps between laboratories performed to monitor any assay bias at the primary laboratory. A total of 157 assay pulp duplicates were retrieved from ALS Vancouver and re-submitted to an independent umpire laboratory (Genalysis in Perth, Western Australia) for check assay.

Comparison of the umpire pulp assays from Genalysis with the primary ALS analyses show an ACV of 12% for Cu and Zn analyses, which is above the 10% tolerance level expected for pulp duplicates. There is a significant negative bias of the Cu and Zn analysis, where the Genalysis analyses are reporting 11% and 13% respectively lower than ALS laboratory for Cu values <1.5% and Zn values <3%.

Comparison of the SGS Bisha laboratory analyses to the primary ALS assay results indicates similar precision and accuracy errors. The Cu and Zn analyses have an estimated ACV of 9% and 25%, which is above the acceptable tolerance level for zinc. A consistent small negative bias exists where the SGS Bisha analyses are reporting 4% and 5% respectively lower than the SGS Bisha laboratory for Cu values < 1.5% and Zn values <3%.

BMSC QA/QC Protocols Summary

The following areas of protocol are highlighted in the review completed by Cube:

The CRMs for copper, zinc and gold have overall performed within acceptable limits for all grade ranges over extended time periods and demonstrate that there is no systemic analytical accuracy issue at the ALS laboratory. However, a consistent small negative bias of 1% to 5% was identified that may be partly related to the two acid partial digestion method used when compared to the near ‘total’ CRM certified value.
Umpire duplicates analyses indicate that the primary ALS laboratory consistently overestimates the Cu and Zn grades (Table 11‑12). Both the SGS Bisha and Genalysis secondary laboratories report assays between 7% and 13% lower than the ALS laboratory. Comparison of the SGS Bisha and Genalysis pulp analyses indicate the SGS Bisha laboratory is overstating the Cu and Zn grades by 11% to 16% when compared with Genalysis.

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The CRM data contradicts the identified duplicate analyses bias with replicate CRM assays indicating the ALS and SGS Bisha analyses are underestimating the true base metal content in the order of 1% to 5%.

Cube made the following recommendations:

Further umpire pulp duplicate analysis is required to determine the source of the apparent bias at the ALS laboratory. It is recommended that a batch of mineralized pulps (>200 samples) are again re-submitted to both ALS laboratory and Genalysis laboratories for analysis. As part of the batch, two CRMS representative of the grade range should be inserted in the batch at a rate of 1 in 5. The increased CRM insertion rate will result in a statistically meaningful dataset on which to assess the accuracy of the laboratories and resolve the bias issue.
Both two acid and four acid digestion methods should be used for the umpire pulp assaying batch to confirm that there is no significant difference or bias between the two methods by using a larger dataset.
In addition, the contradictory duplicate analyses in conjunction with acceptable CRMs replicate assays indicate precision errors may be the result of an inhomogeneous and unrepresentative sub-sample which is being split from the coarse rejects, or that a too small analytical aliquot is being used. Further work is required to determine if an increase in the coarse sub-sample size (from 250 g to 500 g) and/or analytical aliquot (from 1 g to 10 g) results in a reduction of the precision figure.
About 1% of the CRM results were identified as sample mishandling or transcription errors and were excluded from the analysis as outliers. It is recommended that these samples are located and corrected in the database for future analysis.
Ensure CRMs are inserted at a rate of at least 1 in 10 in all umpire pulp duplicate re-assaying to allow assessment of the accuracy of the umpire laboratory.

Given the identified precision issues, the overall accuracy and precision of the assay data relating to the CRMs is within tolerance limits, and no obvious trend or major bias is apparent within the primary assay data. The identified precision errors and bias is not considered material for the Mineral Resource estimate but does require further investigation with the laboratories to establish the source of the errors.

BMSC notes the above and will carry out additional analysis.

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Table 11‑12: Duplicate Sample Comparison by Laboratory

Control Sample Type	Laboratory	Compared to Primary ALS laboratory
Control Sample Type	Laboratory	Cu (0.1% - 1.5 %)	Zn (0.1% - 2.5 %)	Au (>0.5 ppm)
Umpire Pulp Duplicates	SGS Bisha	-5%	-4%	na
Umpire Pulp Duplicates	Genalysis	-11%	-13%	na
Coarse Reject Duplicates	SGS Bisha	-7%	-7%	-2%
Independent Field Duplicates	SGS Bisha	-9%	na	-12%
Independent Field Duplicates	Genalysis	-8%	-3%	-8%

11.3.5 BMSC QA/QC Protocols for Drill Programs, 2013 Grade Control RC

Samples from the grade control campaign between May and September 2013 are included in the Resource estimate; QA/QC protocols for that period are described here. BMSC conducted a QA/QC program that included the insertion of certified reference materials (CRMs), blanks and twin sample duplicates. Some of the CRM overlap the analysis described for drill core samples given above. A total of 8,272 samples were taken with the insertion rate of quality control samples being at least 5%. All analysis is carried out by the SGS on-site laboratory. The number of samples available for quality control is relatively small at this early stage of grade control drilling in the supergene mineralization; BMSC plans to increase the proportion of quality control samples and better assessment can be made as the program progresses. No “round robin” assessment of laboratory performance had been carried out at the time of writing.

Blanks

The source of the blank material and the method of assessing results are as described above in Section 11.3.4.

The results are summarized in Table 11‑13 and indicate that the assay blanks data are within acceptable limits. No obvious contamination issue is apparent within the primary assay data. Failures are, as described above, generally attributable to the nature of the blank material.

Table 11‑13: Blank Performance Summary

Laboratory	Analyte	No. of Samples	No. of Samples Fail	% Passing
SGS on-site	Cu	119	2	99.9%

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Certified Reference Material (CRM)

Various CRMs (assay standards) were used to monitor the accuracy and precision of the assay laboratory and detect any possible bias in the reported analytical values. The method of assessment is as described in Section 11.3.4.

The results are summarized in Table 11‑14. In general, a bias of greater than 5% is not considered acceptable. The performance of the CRMs for copper assays is acceptable, with the same small negative bias observed for diamond core present, albeit for the SGS on-site laboratory.

Table 11‑14: Copper – CRM Performance Summary, SGS On-Site Laboratory, Bisha Mine

CRM	Certified Value (%)	Std. Dev.	Lab. Method	No. Assays	Accuracy Test	Precision Test	% Passing 2SD	% Bias	Period in Use
GBM301-9	0.29	0.0196	AAS23C	19	PASS	PASS	95%	-3%	4/13 – 9/13
GBM304-11	10.40	0.31	AAS23C	77	PASS	PASS	97%	0.3%	4/13 – 9/13
GBM908-15	28,025	1,176	AAS23C	75	PASS	PASS	96%	-1%	4/13 – 9/13
GBM309-15	3.44	0.123	AAS23C	62	PASS	PASS	97%	-2%	4/13 – 9/13
Total	4 CRMs			233					4/13 – 9/13

Duplicates

The results for pairs of duplicate samples (original and duplicate) were analysed as described in Section 11.3.4 and result shown in Table 11‑11. The ACV is above the acceptable precision limit of 15% for copper, but BMSC whilst noting this result also note that a very limited number of samples are available for this period. For future analysis, the population will increase and analysis that is more meaningful can be completed.

In summary, BMSC concludes that the overall accuracy and precision of the assay data relating to the CRMs is within tolerance limits, and no obvious trend or major bias is apparent within the primary assay data. The identified precision errors and bias does require further investigation but is not considered material for the Mineral Resource estimate.

11.3.6 Sanu Resources QA/QC Protocols for Drill Programs, 2006-2010

BMSC were not able to observe any of the Sanu sampling or quality control measures first hand. A review of the Sanu data shows that some 68 CRM and 145 preparation duplicates were inserted in the sample stream and submitted to Genalysis. This represents an insertion rate of approx. 9%. No blanks or field duplicates are recorded. In addition, the SRM used during this period were designed for gold and had only low values for base metals, thus being of no value for copper and zinc, the target elements.

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Overall BMSC consider that while Sanu followed acceptable industry practice in their approach to sampling, preparation and security, their approach to QA/QC measures was inadequate. The samples in this program represent just less than 50% of the samples used in the resource estimate.

11.3.7 Sanu Resources QA/QC Protocols for Drill Programs, 2011-2012

During this phase of work records show that Sanu inserted quality control samples at a rate of some 17% overall, being 12% duplicates, 3% CRM and 2% blanks. The proportion of duplicates is acceptable, the proportion of blanks a little low. No record of analysis of these control samples by Sanu was found; BMSC analysis is given below.

Blanks

The blank failure rate exceeded 10%, which is not acceptable, but the proportion of blank samples is very low.

Certified Reference Material

Various CRMs as noted above were used to monitor the accuracy and precision of the assay laboratory and detect any possible bias in the reported analytical values. The method of assessment is as described in Section 11.3.4.

The results are summarized in Table 11‑15 and Table 11‑16. The performance of the CRMs for copper and zinc assays is generally acceptable, noting the low number of analyses, with a small negative bias present.

Table 11‑15: Copper – CRM Performance Summary, Genalysis Laboratory Perth, 2011 – 2012

CRM	Certified Value (ppm)	Std. Dev.	Lab. Method	No. Assays	Accuracy Test	Precision Test	% Passing 2SD	% Bias	Period in Use
CDN-GCS-14	10130	430	4AH_OE01	21	PASS	PASS	95%	-4%	2011-2012
CDN-GCS-15	4510	200	4AH_OE01	45	PASS	PASS	95%	-2%	2011-2012
CDN-CM-2	10130	430	4AH_OE01	33	PASS	PASS	94%	-3%	2011-2012
CDN-HC-2	46300	2,600	4AH_OE01	21	PASS	PASS	100%	-3%	2011-2012

Table 11‑16: Zinc – CRM Performance Summary, Genalysis Laboratory Perth, 2011 – 2012

CRM	Certified Value (ppm)	Std. Dev.	Lab. Method	No. Assays	Accuracy Test	Precision Test	% Passing 2SD	% Bias	Period in Use
CDN-HC-2	2590	140	4AH_OE01	21	PASS	PASS	90%	5%	2011-2012

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Duplicates

The results for pairs of duplicate samples (original and duplicate) were analysed as described in Section 11.3.4 and result shown in Table 11‑19. The ACV for both copper and zinc is within the acceptable precision limit.

In summary, the identified bias is not considered material for the Mineral Resource estimate. Given the generally acceptable results of the Sanu 2011-2012 quality control program, and the continuity of sample process, laboratory and staff, BMSC also accepts the sampling carried out in earlier programs as being suitable for Mineral Resource estimation.

11.3.8 BMSC QA/QC Protocols for Drill Programs, 2013 Diamond Core, Hambok

Samples from the BMSC diamond-drilling program of eight holes in 2013 are included in the Resource estimate; QA/QC protocols for that period are described here. BMSC conducted a QA/QC program that included the insertion of CRMs, blanks and twin sample duplicates. Analysis at Genalysis Perth was continued for consistency with the Sanu work, with samples crushed, dried, and split at the Bisha Mine facility and pulverization completed at Genalysis Horn of Africa laboratory near Asmara. Sample pulps were despatched to Perth, as described in Section 11.1.2.

Blanks

The blank failure rate was less than 5%, which is acceptable. The rate of insertion of blanks was 3%.

Certified Reference Material

The results are summarized in Table 11‑17 and Table 11‑18. The performance of the CRMs for copper and zinc assays is generally acceptable with a small negative bias persisting.

Table 11‑17: Copper – CRM Performance Summary, Genalysis Laboratory Perth, 2013

CRM	Certified Value (ppm)	Std. Dev.	Lab Method	No. Assays	Accuracy Test	Precision Test	% Passing 2SD	% Bias	Period in Use
GBM301-9	2,881	196	4AH_OE01	20	PASS	PASS	100%	-1%	2013
GBM304-11	104,011	3070	4AH_OE01	26	PASS	PASS	84%	-3%	2013
GBM309-15	34,390	1232	4AH_OE01	37	PASS	PASS	97%	-3%	2013
GBM907-13	15,432	614	4AH_OE01	17	PASS	PASS	82%	-3%	2013
GBM908-15	50,027	1515	4AH_OE01	24	PASS	PASS	100%	-1%	2013
GBM909-14	21,898	842	4AH_OE01	12	PASS	PASS	75%	-5%	2013

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Table 11‑18: Zinc – CRM Performance Summary, Genalysis Laboratory Perth, 2013

CRM	Certified Value (ppm)	Std Dev	Lab Method	No. Assays	Accuracy Test	Precision Test	% Passing 2SD	% Bias	Period in Use
GBM301-9	7,208	433	4AH_OE01	20	PASS	PASS	100%	-3%	2013
GBM304-11	109	23	4AH_OE01	26	PASS	PASS	100%	10%	2013
GBM309-15	123,691	3266	4AH_OE01	33	PASS	PASS	91%	-2%	2013
GBM907-13	66,270	2937	4AH_OE01	17	PASS	PASS	100%	-	2013
GBM909-14	65,582	2291	4AH_OE01	12	PASS	PASS	100%	-2%	2013

Duplicates

The results for pairs of duplicate samples (original and duplicate) were analysed as described in Section 11.3.4 and result shown in Table 11‑20. The rate of insertion was 5%. No meaningful analysis is possible as too few analyses exceed the limit of detection filter.

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Table 11‑19: Duplicate Sample Performance Summary, Hambok 2011 - 2012

Description

QC
Code

Laboratory

Analyte

No. of
Samples

No. Filtered
Samples
>1,000 ppm Cu, Zn
>0.1 ppm Au

Outliers
Removed >100%
Absolute Relative
Difference

Absolute
Relative
Difference

For Filtered Data, Outliers Removed

Comments

ACV
%

Assays
within
10%

Assays
within
20%

Assays
within
50%

Field Duplicate

DUP

Genalysis Perth

21%

15%

50%

75%

95%

Acceptable Precision, no bias

14%

10%

52%

86%

98%

Acceptable Precision, no bias

Table 11‑20: Duplicate Sample Performance Summary, Hambok 2013

Description

QC
Code

Laboratory

Analyte

No. of
Samples

No. Filtered
Samples
>1,000 ppm Cu, Zn
>0.1 ppm Au

Outliers
Removed >100%
Absolute Relative
Difference

Absolute
Relative
Difference

For Filtered Data, Outliers Removed

Comments

ACV
%

Assays
within
10%

Assays
within
20%

Assays
within
50%

Field Duplicate

DUP

Genalysis Perth

122

Insufficient samples

122

Insufficient samples

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11.4 Databases

Since the previous Technical Report all historical data for all deposits has been loaded into the AcQuire™ database system established at the Bisha mine site. Work commenced in 2012 and extensive data validation was undertaken during this process, particularly as applies to the Bisha Main deposit data. All analytical and associated data are now routinely loaded into this database with appropriate QA/QC controls. Analytical data are examined on a batch-by-batch basis prior to loading to the database. The database is also the repository for handheld XRF (Niton) data, geotechnical mapping data and metallurgical QEMScan data.

11.5 Sample Security

11.5.1 Chain-of-Custody

The chain-of-custody for core samples collected and being shipped from site is as follows:

core is transported to the Bisha camp by the drill contractors or BMSC employees and placed in the core logging area
logging and sample preparation area is in a fenced compound
core samples are crushed and sub-sampled
crushed samples are bagged and placed in sealed barrels
each barrel has a list of samples written on the outside of the container
a sample submission form accompanies each barrel
barrels are transported to Asmara in mine-owned vehicles arranged by BMSC.

The sample barrels are submitted to the Eritrean Ministry of Mines for inspection and submission to customs, a customs seal is placed on the barrels and the sample barrels are shipped to Vancouver, where ALS staff receive and clears samples through Canadian customs. Where samples are sent to Genalysis, the chain of custody continues with the pulps couriered from the Horn of Africa preparation laboratory to Perth where the samples are received by Genalysis staff.

The security and chain-of-custody procedures are considered to be reasonable and acceptable.

11.5.2 Sample Storage

Retained pulp, pulp duplicate samples are stored onsite at the Bisha sample preparation facilities. The crushed residues and pulps from grade control sampling are stored at the Bisha site facility for possible additional analysis.

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11.6 Author’s Statement

The sample collection and preparation, analytical techniques, security and QA/QC protocols implemented for the Bisha, Harena, Northwest and Hambok deposits are consistent with standard industry practice and are suitable for the purpose of mineral resource estimation and the reporting of exploration results. The sampling procedures are adequate for and consistent with the style of base metal and gold mineralization found at these deposits.

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12 Data Verification

Data verification pre-mine production was undertaken by AMEC between 2004 and 2011 and post-mine production by AGP in 2011 and 2013. Additional verification work has been carried out in 2012 and 2013; one of the authors is a full time member of BMSC staff. Details of earlier verification work can be found in AGP, 2012; Bisha Polymetallic Operation, Eritrea, Africa NI 43-101 Technical Report for Nevsun Resources; effective date 31 May 2012 and filed on SEDAR. Since that report, the following data verification has taken place.

12.1 Independent Sampling and QA/QC Audits by Cube

During December 2012, Mr. Adrian Shepherd of Cube visited the property and selected two contiguous mineralized intervals from 2 DDH that were considered to be representative of the main mineralized zones at Northwest. Hole NW-023 was drilled from the east to the west and intersected the southern zinc – copper massive sulphide zone, while hole NW-047 was drilled from the west to the east and intersected the main central copper massive sulphide zone. Quarter core samples were taken utilising the original sample intervals. The samples were crushed and a 250 g to 350 g sub-sample riffle split derived as per routine BMSC sample preparation. The resulting 30 samples were sent to Genalysis in Perth for analysis. The results are given in Table 12‑1.

Cube concluded that “comparison of the mineralized intervals between the primary and independent verification duplicate sampling intervals in Table 12‑1 do not indicate any consistent bias between the duplicate (Genalysis) and the original assays (ALS). Cube concludes that the mineralized intercepts returned from the verification sampling confirm the tenor of mineralisation defined by the original SGS and ALS assays. Variations in grades can be attributed to the different sample sizes (1/4 core vs. 1/2 core) and the small 200 g crushed sub-sample which can introduce sample repeatability issues.

Table 12‑1: Independent Verification Samples – Mineralized Interval Comparison

Drill Hole	Domain	Material Type	From (m)	To (m)	Interval (m)	No. of Samples	Original (ALS)			Duplicate (Genalysis)
Drill Hole	Domain	Material Type	From (m)	To (m)	Interval (m)	No. of Samples	Cu %	Zn %	Au ppm	Cu %	Zn %	Au ppm
NW-023	South Zone 201	Primary Zinc	112.65	128.9	16.3	12	0.52	6.73	0.46	0.45	6.40	0.48
NW-047	North Zone 105	Primary Copper	119	137	18	18	0.47	0.03	0.23	0.46	0.04	0.26

In addition to the independent verification work described above, Mr. Shepherd also completed a detailed review of QA/QC protocols as described in Section 11.

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During 2013 Mr. Matt Bampton of Cube visited the property on a number of occasions. A part of his work was independent validation and verification of the sample database for the Bisha and Harena deposits. During this period, Mr. Bampton worked closely with BMSC data administration staff to ensure that as far as possible, all data to be used for future resource estimation work had been validated. Mr. Bampton made use of original survey, geology, and assay records in this work. BMSC subsequently clearly identified all data held within the AcQuire™ database system as being validated or now redundant.

12.2 Data Verification by BMSC

During the construction of the acQuire™ database at the Bisha site, the data administrators under the supervision of the Technical Services Manager completed a wide range of data validation tasks in order to ensure that, as far as possible, all data loaded was valid.

Under the supervision of one of the authors, Mr. Paul Gribble, BMSC carried out a number of validation exercises for data acquired from Sanu Resources for the Mogoraib EL. This included checks of hard copy logs against provided digital records, inspection and re-logging of diamond drill core and re-survey of drill hole collars for the Hambok deposit.

In summary, the author is satisfied that data verification is of a good standard, that no material problems relating to historical data remain, and that the data are suitable for the current Mineral Resource and Reserve estimation.

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13 Mineral Processing and Metallurgical Testing

13.1 Metallurgical Testwork

In 2005, metallurgical testwork was done at SGS Mineral Services in Lakefield, Canada (SGS Lakefield) to support Feasibility Studies for the oxide, supergene, and primary mineralization material types. Two separate drilling programs provided samples for the SGS Lakefield testwork: the first set to support scoping-level studies (Phase I), and the second set for the 2006 Feasibility Study (Phase II).

The core samples of Bisha supergene and primary mineralization were handled and prepared to minimize exposure to air. The oxide mineralization was essentially sulphide-free and, as the cyanide leach testwork was not unduly affected by sulphide oxidation, the oxide mineralization drill core did not require the same care in handling as the sulphide mineralization cores. There was a limit on the number of holes that could be drilled because of the cost of separately collecting these samples.

In 2010, metallurgical testwork was conducted at Mintek in South Africa, directed by the engineering company SENET, to provide plant design and predicted performance data for treatment of the Bisha supergene copper mineralization. The test program was conducted in two phases. The Phase I objectives were to replicate test conditions set by SGS Lakefield in 2005, and to make a sample copper concentrate for marketing purposes. As the results of those tests indicated that the ore sample delivered to Mintek did not behave as the samples previously tested at SGS Lakefield did, it was decided to conduct scouting investigations to determine the reasons for the variable response. For Phase II, the approach was to use a simplified reagent scheme to enhance copper flotation kinetics, and this met with limited success. Poor reproducibility resulted in the concentrate generation program being suspended to investigate the possible reasons for the different results.

Following the testwork at Mintek, Maelgwyn Mineral Services Africa (Pty) Limited (Maelgwyn) in South Africa was contracted to duplicate the test program attempted at Mintek. The objective of the work at Maelgwyn was to advance the testwork initiated at Mintek and demonstrate that the results could be reproducible at the given conditions. SENET, in conjunction with Eurus Mineral Consultants (Eurus), evaluated and optimized the proposed supergene flotation circuit design and expected performance. Eurus employed a proprietary simulation modelling technique using flotation rate kinetic data to construct a flotation circuit simulation model.

In 2011–2012, SGS South Africa (Pty) Ltd (SGS SA) did metallurgical testwork on oxide mineralization from the Harena portion of the deposit to confirm that the metallurgical performance of this material was similar to that for the analogous section of Bisha Main.

Similarly, in 2012 Maelgwyn worked on the supergene and primary mineralization at Harena to confirm that the metallurgical performances of these materials were similar to those for the corresponding mineralization types at Bisha Main.

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Material described as “transitional supergene” and pyrite sand were subjected to quantitative mineralogical examination and metallurgical testwork using both cyanidation and flotation at ALS Metallurgy Kamloops (ALS Kamloops) from November 2012 to May 2013.

In February 2013, quantitative mineralogical examination and metallurgical testwork commenced at ALS Kamloops on material from the Northwest satellite deposit.

In mid-2013, flotation tests were done at ALS Metallurgy Burnie (ALS Burnie) on Bisha Main supergene mineralization.

In mid-2013, mineralogy, flotation and Bond Work Index characterization tests were done at SGS Mineral Services UK Ltd (SGS UK) on Hambok massive sulphide mineralization.

Additional samples of Bisha Main primary mineralization from new drilling underwent a further program of metallurgical testwork from late 2013 at ALS Kamloops to confirm the results of the 2005 work at SGS Lakefield and to generate additional data for the design of the flotation circuit for processing the primary copper-zinc mineralization. This work was still underway at the time of writing this report, but enough data are available to draw initial conclusions.

13.1.1 Metallurgical Samples

Table 13‑1 summarizes the drill hole locations providing the samples for the 2005 and 2010 metallurgical test programs. Figure 13‑1 shows the spatial representation of the metallurgical drill holes within the pit outline, demonstrating a reasonable distribution of samples throughout the pit.

Drill holes Met 05-01 to Met 05-04 were drilled in March 2005 and comprised the Phase I portion of the program, producing five tonnes of samples. The testing program conducted on these samples provided the metallurgical results for the scoping phase of the Bisha Project.

Drill holes Met 05-06 to Met 05-08 were drilled in October 2005, and produced four tonnes of samples. This material was used to complete the Phase II metallurgical testing program to a level suitable for the 2006 Feasibility Study.

Drill holes Met 09 to Met 11 were drilled in October 2009, and produced 1.7 tonnes of samples. This material supplied samples for the metallurgical testwork at Mintek.

Drill holes Met 12 to Met 17 were drilled in July 2010, and were used for bench scale flotation tests at Maelgwyn.

A number of intervals logged as “transition” in Met 12, Met 16, and Met 17 represent material found within the gradational contact zone between the upper layer of supergene and the underlying primary mineralization. No testwork was done on this material.

Harena supergene and primary material came from drill holes H Met 01 and H Met 02.

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Table 13‑1: Metallurgical Sample Drill Hole Locations

Drill Hole Tag #	Drill Hole Coordinates	Azimuth	Dip	Ore Type	Depth From (m)	Depth To (m)
Met 05-01	1715400N, 339485E	270	-80	Oxide	10.5	36.0
				Supergene	39.0	67.5
				Primary	67.5	202.5
Met 05-02	1715500N, 339325E	90	-65	Primary	100.0	250.0
Met 05-03	1716050N, 339395E	vertical		Oxide	3.0	36.2
				Supergene	36.4	82.5
Met 05-04	1716050N, 339295E	290	-70	Oxide	-	36.0
				Supergene	36.0	84.0
Met 05-05	1715250N, 339520E	270	-80	Supergene	34.5	50.5
				Primary	50.5	115.0
Met 05-06	1715575N, 339425E	90	-90	Oxide	22.0	34.5
				Supergene	34.5	62.5
				Primary	62.5	152.5
Met 05-07	1716225N, 339425E	270	-80	Oxide	31.5	42.0
				Supergene	42.0	68.5
				Primary	68.5	124.3
Met 05-08	1715850N, 339400E	90	-70	Supergene	40.5	58.8
Met-09	1716128N, 339405E	270	-80	Supergene	36.15	84.00
Met-10	1715976N, 339381E	90	-75	Supergene	42.10	67.00
Met-11	1715698N, 339387E	90	-80	Supergene	38.80	68.45
Met-12	1715402N, 339497E	270	-80	Supergene	41.50	43.00
				Supergene	50.50	52.00
				Supergene	53.50	55.00
				Transition	61.00	62.50
Met-13	1715503N, 339294E	90	-80	Supergene	43.00	44.50
				Supergene	49.00	50.50
Met-14	1716052N, 339369E	90	-80	Supergene	38.50	40.00
				Supergene	52.00	53.50
				Supergene	61.00	62.50
Met-15	1716053N, 339286E	270	-70	Supergene	34.00	35.50
				Supergene	40.00	41.50
				Supergene	49.00	50.50
				Supergene	67.00	68.50
Met-16	1716051N, 339347E	90	-70	Transition	86.50	91.00
Met-17	1716050N, 339318E	270	-55	Transition	86.00	90.50

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Figure 13‑1: Metallurgical Drill Hole Locations with Supergene Wireframe within Bisha Main Pit

Source: Nevsun, 2010

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The transitional supergene and pyrite sand material used for testwork were taken in October 2012 from stockpiles generated in the course of mining the oxide ore.

Material for metallurgical testing of the Northwest satellite deposit came from two drill holes labelled Northwest Met Hole 001 and Northwest Met Hole 002; the former was described as “copper only,” with mineralization said to be analogous to Bisha supergene, and the latter described as “copper-zinc” mineralization being analogous to Bisha primary.

Hambok material came from 13 mineralized intervals taken from 7 drill holes, representing the range of massive sulphide material.

The flotation testwork done at ALS Burnie used half-core material from eight drill holes labelled Met 024 to Met 031. Figure 13‑2 shows the location of drill holes used to generate the samples for the 2013–2014 testwork programs on Bisha primary material.

Figure 13‑2: Metallurgical Drill Hole Locations for Bisha Primary Testwork 2013–2014, Longitudinal Section

Source: BMSC

13.1.2 Composite Samples

One master composite for the 2005 SGS Lakefield Phase I program was made up from the variability composite samples for each of the following three main mineralization types for Bisha Main:

oxide mineralization
supergene mineralization
primary mineralization.

For the 2005 Phase II program, two primary master composites typical of primary mineralization were produced, a zinc-rich master composite, and a low-zinc master composite. To the extent possible, the master composite samples were made to match the grades in the Mineral Resource as then estimated for each of the mineralization types.

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Seven supergene variability composite samples were produced from the eight metallurgical sample drill holes, ranging from a low of 1.82% Cu to a high of 15.9% Cu. Two master composite samples were produced from these variability composites, one for the 2005 SGS Lakefield Phase I testwork with a head grade of 1.93% Cu, and the other for the 2005 SGS Lakefield Phase II testwork assayed at 4.2% Cu.

Fifteen primary mineralization variability composite samples were generated from the eight metallurgical drill holes: seven for the Phase I testwork, and eight for the Phase II testwork.

A single composite was generated from the bulk of the drill samples sourced from supergene mineralization for the 2010 Mintek program.

The composite material from the Mintek program was used for the 2010 Maelgwyn program.

In addition, rougher flotation tests were performed on each of the metallurgical drill core samples from Met 12 through Met 17, with additional rougher kinetic tests.

Harena oxide material was made into four composite samples for comminution and cyanidation testwork, while a single composite each of supergene and primary material was used for the comminution and flotation testwork.

Two batches each were generated for the transitional supergene and pyrite sand materials.

The Northwest satellite deposit material was made into 15 composites for quantitative mineralogical examination, and five metallurgical composites for comminution and flotation testing.

The 13 mineralized intervals from the Hambok deposit were combined into a single composite for a short program of metallurgical characterisation.

Bisha primary material for the 2013–2014 testwork program was made into four master composites on the basis of assay and quantitative mineralogy data, namely mean Zn:Cu ratio (meant to represent the mean resource grade of the deposit), low Zn:Cu ratio (also a surrogate for cutoff-grade material), high Zn:Cu ratio, and low pyrite (representing stringer mineralization). Variability testing will be done on 47 separate drill intervals to give a larger number of metallurgical data points throughout the mineralized volume.

13.2 Grinding Testwork

13.2.1 2005 SGS Lakefield Test Program

Splits were taken from each of the three Bisha Main Phase I master composite samples for the standard grindability determinations, such as JK drop-weight tests, MacPherson autogenous mill tests, Bond rod mill and ball mill work indices, and abrasion indices. Since the oxide ore was indicated to potentially be the hardest of the three ore types, the grindability work on the oxide ore was expanded to include rod mill and ball mill Bond work index testing on the oxide ore variability samples, and a MacPherson mill test on the oxide ore master composite sample.

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In the context of the JK Tech mineralization database, the supergene material is in the very soft range, and the oxide and primary ores are in the soft to very soft range of resistance to impact breakage.

As part of the JK Tech sample property assessment, the relative densities of 30 randomly selected particles for each mineralization type in the 26.5 mm to 31.5 mm size range were determined by weighing each particle in water and in air. The majority of the supergene and primary ore samples had relative density values within the range of 4.6 to 5.0; however, the oxide ore samples displayed a bimodal histogram, with one set of values bracketing a relative density of value 2.5, and the other set bracketing a relative density value of 3.9. It was concluded that the ferruginous gossans had higher densities, while the acid oxide, breccias, and saprolite materials had lower densities.

Autogenous work index (AWi) values were determined in accordance with the procedures used for the standard MacPherson grindability test. The gross AWi resulting from the test was compared against the MacPherson database of operating plants to provide a correlated AWi value. AMEC’s experience with this conversion led to the conclusion that it was too low; therefore, AMEC added an additional factor between 1.4 and 1.6 to the correlated AWi. For the oxide ore, this resulted in a semi-autogenous grind (SAG) mill work index range of 11.8 kWh/t to 13.4 kWh/t, which bracketed the 12.7 kWh/t value back-calculated from the JK SimMet SAG mill model. Consequently, AMEC considered that there was good agreement between the two methods in sizing the SAG mill for the oxide ore. The supergene and primary ores are much softer than the oxide ore; therefore, neither of these ores had a significant input into sizing the grinding mills. Similar comparative calculations were therefore not performed on those composite samples.

Standard Bond rod mill and ball mill work index determinations were conducted on the master composite samples, and Bond ball mill work indices were determined for the harder oxide, supergene, and primary ore variability composite samples. The oxide ore was the hardest of the three ore types, and was used to size the grinding mills for the proposed plant. The lower rod mill work indices for all three mineralization types indicated these mineralization types were amenable to first-stage SAG mill grinding.

13.2.2 2010 Test Program

No grinding testwork was conducted as part of the 2010 Mintek and Maelgwyn test programs.

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13.2.3 2011–2012 Test Program

Bond ball mill work index were measured for each of the Harena oxide, supergene, and primary composites. The four Harena oxide samples had values ranging from 4.9 kWh/t to 20.1 kWh/t, whereas the supergene and primary mineralization types returned values around 10 kWh/t, which is in accordance with expectations for this type of deposit.

13.2.4 2013 Northwest Program

Standard Bond rod mill, ball mill, and abrasion tests were done on the five metallurgical composites. The results for Bond rod mill work index and abrasion index were at the lower end of the range of values determined for supergene and primary mineralization, while the Bond ball mill work index was at the upper end of the range for oxide mineralization.

13.2.5 2013 Hambok Program

A single Bond ball mill work index test gave a value within the range measured for primary mineralization.

13.2.6 2013–2014 Bisha Main Primary Program

Results of the Bond rod mill, ball mill, and abrasion tests on the four master composites were at the lower end of the range of values determined in the 2005 testwork for oxide, supergene, and primary mineralization that were used to size the grinding section.

13.3 Mineralogy

13.3.1 2005 SGS Lakefield Test Program

Oxide Mineralization

SGS Lakefield examined gold occurrence in a portion of Phase I oxide master composite sample using optical microscope and X-ray diffraction methods to evaluate potential metallurgical performance.

A portion of the oxide sample was concentrated on a superpanner and separated into three fractions. Results showed that gold in the oxide zone would not be amenable to gravity recovery, with good leach extractions requiring a comparatively fine grind.

Supergene Mineralization

Four of the Phase II supergene variability composite samples, the Phase II supergene master composite sample, and four individually selected supergene core samples were submitted for mineralogical analysis using the method of bulk mineralogical analysis (BMA). The samples were analysed by the QEMSCAN automated mineralogy system. The major minerals in the supergene mineralization were pyrite (72% to 96%), the secondary copper sulphide minerals covellite (1% to 9%), chalcocite (0.2% to 3%), and enargite, the primary copper sulphide chalcopyrite and bornite, and non-sulphide gangue (0.5% to 7%). Of the total copper mineralization, 40% to 80% occurred as secondary minerals, with 20% to 60% as primary minerals. There was significant sphalerite in two of the samples, and small quantities of molybdenite and galena in several samples. Non-sulphide gangue minerals were a very minor constituent of the samples examined, with a maximum of 7% in one. At a sizing of 80% -100 µm, nearly all of the copper was present as either a liberated primary or secondary sulphide mineral particle or an association between the two; only 4% to 5% of the copper (mainly secondary sulphides) was present as binary particles with pyrite, and 3% to 4% as “complex” (i.e., ternary or quaternary) particles.

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Liberated mineral release curves constructed for the supergene mineralization demonstrated that very high concentrate grades would be possible if the liberated copper minerals could be separated cleanly from the sulphide and non-sulphide gangue. The dominant issue in flotation processing of the supergene material is the very high pyrite:copper sulphide ratio.

Primary Mineralization

Seven primary mineralization variability composite samples collected during the Phase II sampling program and ground to a sizing of 80% -75 µm were submitted for QEMSCAN mineralogical study. Chalcopyrite was the most abundant copper mineral in the primary mineralization, with grades ranging between 1.5% and 6.0%. Zinc was present as sphalerite with a wide grade range, varying from 2.4% to 26.6%. Pyrite content was high, comprising 65% to 95% of the samples. Copper mineral liberation exceeded 80% in all samples, creating the expectation that a good copper concentrate grade should be achievable given the appropriate flotation conditions. As expected, pyrite was the major association of the chalcopyrite, being attached to 8% to 12% of the copper mineral particles contained in the sample. Approximately 5% of the copper was in complex particles.

Limiting grade/recovery curves indicated that the maximum possible copper concentrate grade would be in the low 30% Cu range at 90% recovery. In practice, actual separation by flotation is likely to be negatively affected by copper ions from the secondary copper sulphide minerals activating the pyrite and sphalerite.

Sphalerite in the composite samples was also very well liberated at 80% to 94%, with only about 1% in complex particles. The sphalerite/pyrite association ranged between 5% and 15%, and the sphalerite/copper association was only 1% to 2%. The zinc mineral associations suggest that the zinc should be readily recovered to high-grade zinc concentrates, and that zinc reporting to the copper concentrate will be due to deficient separation rather than inadequate liberation. Microprobe analysis of the sphalerite showed it to be a single compositional population with an iron content of 3% to 4%. This should allow the production of a high-grade zinc concentrate, assuming efficient flotation separation.

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An objective in obtaining good zinc flotation results will be to minimize the loss of sphalerite to the copper concentrate, which is expected to be a chemistry issue rather than one of liberation.

13.3.2 Supergene Mineralization

2010 Mintek Test Program

Quantitative mineralogical analysis was conducted on the flotation feed sample to compare the 2010 program flotation feed composite to the composites tested in 2005. Although the data output of the two separate testing facilities was not directly comparable, it was determined that the 2010 composite sample was lower in bornite and chalcopyrite, with higher contents of covellite and chalcocite, compared with the samples tested in 2005. It was not confirmed by Mintek that the differences in the two sample responses in flotation could be solely a result of mineralogy.

2010 and 2010–2012 Maelgwyn Test Programs

Quantitative mineralogical analysis was done in 2011 and 2012 on feed and test products, mostly for Bisha supergene material, by ALS Laboratory Group MLA Division – ALS Chemex in South Africa, SGS South Africa, and G&T Metallurgical Services Ltd in Canada (G&T was the predecessor entity to ALS Kamloops). The results were in accordance with the findings of the 2005 SGS Lakefield work. Enargite, tennantite, and arsenopyrite were identified in the supergene mineralization, with the first two minor mineral occurrences reporting to the copper concentrate, but arsenopyrite is expected to behave as pyrite and be rejected to the tailings.

G&T’s examination of a sample of Harena primary mineralization showed it to have a similar mineral composition to that of Bisha Main primary material.

13.3.3 Transition Supergene and Pyrite Sand

QEMSCAN analysis at ALS Kamloops determined that, for all samples, pyrite was the predominant sulphide mineral, with minor amounts of both primary and supergene copper sulphide minerals. Silver occurred as native silver, acanthite, and/or argentite, with some naumannite, a silver-selenium alloy. Non-sulphide minerals present were quartz, barite, iron oxides, and lead oxides.

13.3.4 Northwest Satellite Deposit

Mineral occurrences and liberation characteristics were similar to those for the Bisha Main deposit, with the “copper only” material having the same range of copper minerals as Bisha supergene, and the “copper-zinc” mineralization being similar to Bisha primary, albeit both with lower contents of the valuable minerals.

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13.3.5 Hambok Deposit

Quantitative mineralogical examination of the combined composite material by QEMSCAN at SGS UK showed a high pyrite mineral assemblage with liberation characteristics similar to that of Bisha primary material.

13.3.6 Primary Mineralization

Results from QEMSCAN analysis at ALS Kamloops generally agreed with the previous work done by SGS Lakefield in 2005.

The presence of secondary copper sulphide minerals at the tops of some of the drill holes indicated that supergene geological processes had taken place, and that such material should not be processed with the primary copper-zinc mineralization, due to the likelihood of copper ion activation of sphalerite and pyrite resulting in poor selectivity in copper flotation.

A detailed review of mineral associations indicated that while overall requirements for the liberation of chalcopyrite and sphalerite had not changed from the 2005 study, it might be possible to have a coarser flotation feed sizing than previously thought, provided there was sufficient regrinding.

13.4 Cyanidation Testwork

13.4.1 2005 SGS Lakefield Test Program

The initial set of cyanidation tests examined order-of-magnitude grind size and cyanide solution strength versus leach extraction.

Reduced cyanide solution strengths indicated reduced leach extractions, the differences diminishing the longer the leach time was extended. Leach extraction times varied with different test series, resulting in the variables of grind, CN concentration, and leach slurry density being examined in more detail. During Phase II testing, two oxide mineralization samples were subjected to a series of 24-hour leach tests with variations in grind P₈₀ of 60 μm and 75 μm, variations in leach slurry density of 45% and 50% solids, and variations in CN concentrations of 0.25 g/L and 0.5 g/L NaCN. The coarser grinds for both composite samples gave higher leach extractions. In line with expectations, the higher cyanide concentration resulted in higher leach extraction, and higher slurry densities resulted in lower leach extractions, although only by a small margin. To rationalize these grind versus extraction results, it was assumed that within the P₈₀ range of 60 μm to 85 μm, there was no difference in leach extraction. The gold deportment study and grind versus recovery tests in the P₈₀ range of 60 μm to 85 μm suggested that there was some porosity in the oxide mineralization matrix that allowed the cyanide to penetrate to the locked gold particles.

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13.4.2 2005 SGS South Africa

Cyanide leaching tests on the four composites of Harena showed medium to high gold dissolution, with a mean well above 80%; a good outcome considering the much lower gold head grade compared to Bisha Main oxide material. Higher CIL gold extractions compared to straight cyanide dissolutions showed two samples had significant preg-robbing characteristics. However, there should not be significant effect on gold recovery when treating Harena oxide material through the Bisha plant using standard practices, such as only adding cyanide when carbon is present, or reducing the activity of the preg robber in the ore by adding a hydrocarbon before cyanidation.

13.5 Flotation Testwork

13.5.1 2005 SGS Lakefield Test Program Oxide Mineralization

A single flotation test on the Phase I oxide mineralization master composite showed poor gold flotation selectivity, with rougher concentrate grades ranging from 55 g/t Au to 12 g/t Au for gold recoveries between 44% and 77%. The conclusion from the test result was that oxide mineralization was not a good candidate for upgrading by flotation.

Supergene Mineralization

Mineralogical studies confirmed that the supergene mineralization had high pyrite content, and that copper mineralization occurred primarily as covellite with lesser chalcocite, chalcopyrite, and bornite. There is a strong possibility that copper ions from covellite and chalcocite have activated pyrite surfaces. This phenomenon exacerbated the difficulty of the flotation separation between copper sulphide minerals and pyrite. Pyrite depression was the biggest challenge in evaluating the supergene mineralization flotation processing strategy.

During the Phase I program, over 30 batch flotation and two locked-cycle tests were conducted on the master composite sample. Grade/recovery curves for the supergene tests were essentially straight lines, with the slopes of the lines being mildly dependent on the degree of rougher concentrate regrinding. The conclusions drawn were:

A flotation feed P₈₀ of 75 μm will provide in excess of 90% Cu rougher recovery
Little difference was observed in effectiveness between reagents SIPX and PAX collector, therefore PAX was eventually selected, as it was noted as the better reagent for primary ore flotation
High lime additions were required in grinding, rougher flotation, and cleaner flotation to obtain maximum recovery of copper by depressing pyrite
The use of sodium sulphite in the primary grind and in regrinding limited the drop in redox potential and pH during the grinding stages, therefore less lime was required to maintain the pH through the course of the test.

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The Phase II supergene master composite had a grade of 4.2% Cu, which was more comparable to the 4.4% Cu grade then estimated for the mineral resource. Higher copper recoveries were achieved at this higher head grade than with the Phase I sample.

An additional 27 batch tests and four locked-cycle tests were conducted on the

Finer grinds provided an incremental increase in copper recovery; the flotation feed sizing was therefore changed from P₈₀ of 75 μm to P₈₀ of 55 μm.
Regrinding all of the supergene rougher concentrate provided only marginal improvement in the position of the grade/recovery curve for the subsequent three stages of cleaning as compared to the roughing-only grade/recovery curve.
Roughing kinetics tests were conducted on several supergene variability composite samples. Only a single test was run for each composite, therefore the test conditions were not optimized.

Primary Mineralization

These initial tests confirmed that the issue in flotation processing of the primary mineralization would be making a copper concentrate with minimum misplacement of sphalerite and pyrite. Zinc flotation was relatively easy, with recoveries to zinc rougher concentrates of 95% or more of the zinc remaining after copper flotation, which then upgraded to 55% Zn grade concentrates in two cleaning stages. The majority of the work on the primary mineralization was therefore focused on copper flotation.

The initial Phase I flotation tests on the primary mineralization indicated copper rougher recoveries up to 90% were achievable with a grind P₈₀ of 75 μm. Similar to the supergene mineralization, high lime additions were required. Because of sample deterioration, the optimum copper flotation test conditions were not considered to be established during the Phase I program. This sample deterioration is ascribed to excess moisture observed in the samples as received, allowing galvanic oxidation reactions of sulphide minerals to proceed. A major objective of the Phase II test program was to determine the conditions that would produce optimum metallurgical performance for copper and zinc from the primary mineralization.

Only a few zinc concentrate regrinding and cleaning tests were conducted during the Phase I program, but these generally produced good results.

The copper in the Phase II zinc-rich master composite sample had a higher degree of dissemination than the copper in the Phase I sample, and a grind P₈₀of 55 μm was necessary to achieve similar copper grade/recovery curves to those obtained in the Phase I program. The grind target was therefore changed from a P₈₀ of 75 μm to a P₈₀ of 55 μm both for the remaining testwork and for the plant design.

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Copper roughing kinetic tests showed similar flotation characteristics to the master composite samples, with copper recoveries ranging between 87% Cu and 96% Cu. Zinc roughing kinetic tests were also conducted on the variability samples. The majority of these tests recovered 96% or more of the remaining zinc to the zinc rougher concentrate.

As only single tests were conducted on the variability composite samples, these conditions were not optimized, and were not used to formulate the final grades and recoveries for the primary mineralization.

Initial flotation tests on both the zinc-rich and low-zinc Phase II master composite samples used a depressant combination of zinc sulphate and sodium cyanide in both the primary grind and in the copper section regrind. Later tests on the Phase II primary mineralization master composite focussed on substituting sodium sulphite for zinc sulphate/sodium cyanide depressant in both the primary grind and the regrind. Sodium sulphite gave improved selectivity against pyrite, but inferior selectivity against sphalerite in copper roughing and cleaning. Combinations of both depressants were also tested, which showed some promise.

13.5.2 2010 Mintek Test Program

The objective of the Mintek test program was to replicate the supergene flotation testwork completed in 2005. This program evaluated various circuit configuration and design issues, including primary and secondary rougher flotation rates, primary and secondary cleaner flotation rates, and cleaning tests with and without regrinding. These tests showed that the ore sample delivered to Mintek did not behave the same as the sample previously tested at SGS Lakefield did, as the flotation response was slower.

The supergene composite used in Mintek testwork had a head grade of 4.6% Cu and high iron content of 44%, which agrees with the known mineralogy. The final cleaner concentrate produced using the SGS Lakefield flowsheet and conditions gave a copper concentrate grade of 31.6% Cu at 87.2% recovery, compared with a concentrate grade of 30% Cu at 93% recovery assumed for the 2005 SGS Lakefield program.

13.5.3 2010 and 2011–2012 Maelgwyn Test Programs

The Maelgwyn test program on Bisha Main materials was originally designed to optimize the flotation parameters for Feasibility design of the supergene and primary mineralization. The flotation variables tested included feed sizing, flotation time, varying dosages in the reagent suite, mass pull, pH, effect of temperature, regrind, and extent of oxidation. Repeats were conducted on selected tests to confirm the reproducibility of results.

Following optimization of some conditions, additional batch flotation testwork was conducted on supergene sample composite material to generate kinetic data to allow for calculation of flotation rates in staged flotation, including roughers, cleaners, and cleaner scavengers. Rougher kinetic tests only were completed for primary mineralization composite samples. The kinetic data was used by Eurus to build simulation models of the circuit for further analysis and optimization. Locked-cycle flotation tests were done to support the expected metallurgical performance. A total of 35 batch and three locked-cycle tests were done on the supergene sample.

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The supergene test composite prepared for the Maelgwyn testwork contained 4.4% Cu, 0.50% Zn, 43.2% Fe, and 46.3% S. Locked-cycle flotation testing using optimized conditions produced a copper concentrate of 34% Cu at 87.3% recovery.

Maelgwyn initially did 15 batch flotation tests of two samples of Harena supergene primary mineralization, using basically the same flotation conditions established in its current program on Bisha supergene and derived from the previous 2006 SGS Lakefield work on Bisha Main primary mineralization. While rougher flotation recoveries of over 95% were achieved for both copper and zinc, efficient separation between the copper minerals and sphalerite could not be achieved with the limited number of scouting tests. Accordingly, a further 20 batch flotation tests were done, with the optimized performance for the primary mineralization being a copper concentrate grade of 28% Cu containing only 1.3% combined Pb and Zn at 94% recovery, and a zinc concentrate grade of 44% Zn at 87% recovery. When predicting the production plant performance for the Harena primary mineralization, the normal assumption used was that the zinc concentrate grade could be increased at the expense of decreasing the zinc recovery.

13.5.4 Transitional Supergene and Pyrite Sand at ALS Metallurgy Kamloops

Testing was limited to scouting rougher flotation, but results were encouraging, with indications that high-grade precious metals concentrates could be made at economic recoveries of gold and silver. Such a concentrate could be added to a copper concentrate, achieving good payment terms for the gold and silver.

13.5.5 2013 Northwest Satellite Deposit at ALS Metallurgy Kamloops

Flotation performance from batch tests with roughing and three stages of cleaning on “copper only” mineralization was very good, considering relatively low copper head grades of 0.4% Cu to 1.37% Cu with concentrate grades of ≈30% Cu at 80% to 90% Cu recoveries.

Metallurgical performance from “copper-zinc” mineralization was variable, with good results for both copper and zinc from one composite, while the other was poorer due to persistent sphalerite flotation into the copper concentrate. This was ascribed to activation of the sphalerite by copper ions; elevated levels of secondary copper sulphide minerals in this composite indicated that supergene geological processes had taken place.

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13.5.6 2013 Supergene Mineralization at ALS Metallurgy Burnie

Forty batches and five locked-cycle tests were done to investigate optimization of flotation processing of supergene mineralization. This used a slightly different flowsheet to that proposed for the initial plant commissioning on supergene material, in that all of the rougher concentrate was ground to P₈₀of 25 µm before three stages of cleaner flotation.

The locked-cycle tests showed that a copper concentrate assaying around 40% Cu at 85% recovery could be produced from supergene material.

13.5.7 2013 Hambok Deposit at SGS Mineral Services UK

Scouting flotation tests demonstrated that the composite of Hambok material had a similar metallurgical performance to that shown in laboratory tests on Bisha primary mineralization.

13.5.8 2013–2014 Primary Mineralization at ALS Metallurgy Kamloops

While the results to date using a similar flowsheet and reagent regime have generally been in accordance with the SGS Lakefield work done in 2005, the following differences will be incorporated into the process design for the treatment of primary mineralization:

Coarsening the flotation feed sizing from P₈₀ of 53 µm to P₈₀ of 75 µm while retaining the same approximate regrind sizing requirement of P₈₀ of 15 µm for copper rougher concentrate and P₈₀ of 20 µm for zinc rougher concentrate
Increasing the number of zinc cleaner flotation stages from two to three to ensure that a high-grade zinc concentrate is produced.

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14 Mineral Resource Estimates

Mineral Resources were estimated for the Bisha Main, Harena, Northwest, and Hambok deposits. The estimates were completed or supervised by Paul Gribble, C.Eng., FIMMM, Chief Resource Geologist for BMSC, and the Qualified Person for the Mineral Resource estimates. The effective date is 31 December 2013. Each estimate is separately described and reported below. Gemcom Surpac version 6.5 (Surpac) was used for all geological interpretation and grade estimation work; Isatis and Supervisor 8.1 were used for statistical and spatial analysis.

14.1 Bisha Main Estimate

14.1.1 Project Sample Database

Since the May 2012 estimate, the sample database has undergone a thorough QA/QC process, and any historical data considered unreliable was excluded.

The current interpretation and estimation was completed using Diamond Drill (DD) cored holes and recent Reverse Circulation (RC) grade control holes. Additional DD completed since the May 2012 technical report are metallurgical holes or exploration holes that do not affect the Resource. For the most part, the metallurgical holes represent partial intersections and are mostly excluded. This estimate, therefore, includes some 500 DD and 475 RC holes for approximately 30,700 and 5,840 samples respectively. The majority of the latter were drilled in 2012/2013. A summary of drilling at the deposit is given in Table 14‑1Table 1‑1; not all holes given here are included in the Estimate. A cutoff date of 30 September 2013 was applied to the RC drilling data. More detail on all drilling campaigns can be found in Section 10, Drilling.

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Table 14‑1: Bisha Main Drill Hole Summary Table

Year	Phase	Range of Hole No.	No. of DDH	Length of DDH (m)	No. of RC Holes	Length of RC (m)	Total No. of Holes	Total Length (m)
2002	-	B-001 – B-006	6	810.90	-	-	6	810.90
2003	I	B-002a, B-007 –B-0 53	48	6,724.76	-	-	48	6,724.76
2003	II	B-058 to B-146	79	9,861.7	-	-	79	9,861.7
2004	-	B-147 – 309	163	28,954.05	-	-	163	28,954.05
2004	-	BRC-001 – BRC-041 *	-	-	40	2,117.40	40	2,117.4
2004	-	BRCD-026 –BRCD-42 *	9	308.80	-	282.90	9	591.70
2005	I	B-310 to 367	58	7,520	-	-	58	7,520
2005	I	GT-01 to 05	06	937	-	-	06	937
2005	II	MET-05-01 to MET-05-08	08	1,212.5	-	-	08	1,212.5
2005	II	BH-01 to BH- 12, BH-14 and BH-15	14	410	-	-	14	410
2006	-	B-368 to B-371	04	1,014	-	-	04	1,014
2009	-	GT-006 to GT-014	09	1,201.5	-	-	09	1,201.5
2009	-	MET-009 to MET-011	03	224	-	-	03	224
2010	I	B-372 to B-378	07	1062	-	-	07	1062
2010	I	MET-012 to MET-017	06	478	-	-	06	478
2011	I	B-379 to B-399	21	2,592.4	-	-	21	2,592.4
2011	II	B-400 to B-504	105	2,4494.1	-	-	105	2,4494.1
2011	II	MET-018 to MET-019	02	180	-	-	02	180
2011	II	GT-015 to GT-020	06	695.5	-	-	06	695.5
2012	I	B-505 to B-511	07	409.5	-	-	07	409.5
2012	-	MET-020 to MET-031	12	820.7	-	-	12	820.7
2012	-	GT-021 to GT-032	12	3,048.5	-	-	12	3,048.5
2013	-	B-512, BEXT001 to BEXT014, BMT032 to 041, BNDD001 to BNDD008, BRC series (RC Grade control holes)	33	7,228.17	474	6,243	507	13,471.17
Total			618	1,00188.08	514	8,643	1,132	108,831.38

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14.1.2 Geological Model

Introduction

The mineralized components of the Bisha deposit are divided into three major zones, namely the oxide, supergene, and primary zones.

The large majority of the oxide component is mined out; a small zone remains in the northeast of the deposit.

The massive sulphide bodies of the supergene and primary zones are divided into a southern zone (the main zone) and a northern zone, whose dip extent is limited. The southern zone strikes at some 345° and dips steeply to the west, with strike and dip lengths of some 650 m and 600 m, respectively. The northern zone strikes at around 0° and dips steeply to the west, with strike and dip lengths of some 500 m and 100 m, respectively. In addition, there are massive sulphide pods in the hanging wall of the main massive sulphide units and development of supergene-mineralized material, also in the hanging wall. These last tend to be of lower-grade material. There are also additional zones of low-grade disseminated copper mineralization in the footwall and hanging wall of the massive sulphide bodies. The main mineralized zones are shown in Figure 14‑1.

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Figure 14‑1: Major Mineralized Zones at the Bisha Main Deposit, Looking Northwest

Note: Key: Red shapes = massive sulphides; Green shapes = Hanging wall copper; drill holes shown in dark green.

Methodology

A major reinterpretation of the geological model for Bisha has been undertaken since the May 2012 technical report. This work included review of drill core, logging data, grade control mapping, new RC data, and detailed pit wall mapping. The data was all combined in plan and sectional interpretations, with resulting three-dimensional (3D) models. This work, discussed in more detail in Section 7 and Section 8, resulted in reinterpretation of the supergene and primary zones for this estimate. The focus of reinterpretation was:

More detail in the supergene zone
Introduction of mineral processing characteristics as they relate to the supergene and primary zone mineralization

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Introduction of structural controls to the distribution of the mineralization. From the studies completed at the deposit, the mineralization is now seen as structurally controlled, with faulting playing a major role in an overall brittle regime, as opposed to the original folding model.

From an estimation and Mineral Resource perspective, this resulted in the following:

Interpretation of a high-grade copper core in the supergene zones immediately up-dip of the high-grade primary zone copper/zinc mineralization. The high-grade core is flanked by lower-grade material, largely confined to the massive sulphide mineralization. Some copper-mineralized sulphide stringer mineralization is present, dominantly in the footwall.
Extension of the Hanging Wall Copper Zone (HWCu) and reinterpretation of these mineralized zones, consistent with the steeply-dipping controlling structures striking at 045°.
Truncation of massive sulphides by steeply-dipping fault/shear structures trending at 000° and 020°.
Recognition of a “zinc-depletion surface” within the massive sulphide mineralization marking the change from supergene to primary zone. This surface is defined as the first recognition of sphalerite mineralization, which forms a consistent surface throughout the deposit.
Reinterpretation of the zinc-rich zones within the primary zone that led to tighter control of the high grade zones.
Reinterpretation of the copper-rich zones within the primary zone: a number of such zones or pods are present immediately below the zinc-depletion surface that also continue into the supergene high-grade zones to form separate domains.
Recognition of a transition zone below the zinc-depletion surface where secondary copper minerals and primary chalcopyrite are mixed with zinc mineralization.

A typical vertical cross-section is given in Figure 14‑2, controlling structures in Figure 14‑3, and a schematic section of the key mineralization and alteration features of the deposit in Figure 14‑4.

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Figure 14‑2: Vertical Cross-Section at 1715500N through the Bisha Main Mineralization, Looking North

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Figure 14‑3: Plan View of Massive Sulphide Zones and Controlling Structures

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Figure 14‑4: Schematic Cross-Section of Mineralization and Alteration

Source: after Cube Consulting 2013

14.1.3 Alteration and Weathering

Weathering surfaces, based upon lithology logs, core photos, and mineral assemblages, were constructed on 25 m spaced cross-sections to define the main weathering and oxidation domains. This work resulted in the modelling of gossan, base of complete oxidation, base of acid leach zone (locally known as “soap”), and the top of fresh rock surfaces. These models interacted with the models of mineralization.

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14.1.4 Mineralization Domains

Mineralization models were created on east-west cross-sections at 25 m intervals, with shapes interpreted in cross-section snapped to the drill hole traces to honour 3D locations. Cross-sectional models were joined together to form 3D-triangulated models. Most of the domains are defined by lithological features, such as massive sulphide, which generally have very sharp, well-defined contacts. Some zones are less well defined, and are constrained based on mineralized grade intercepts. These are:

Hanging Wall Copper, broadly a 0.5% Cu boundary
Internal zones within the primary massive sulphide, broadly a 0.5% Cu and 1% Zn boundary
Footwall Copper (pyrite stringer zone), broadly a 0.5% Cu boundary
Minor remnant oxide gold, broadly a grade contrast model.

Parrish analysis (refer to Section 14.1.10) identified the potential need for high-grade capping. This was largely obviated by the interpretation of sub-domains within the primary massive sulphide mineralization, thus:

High-grade zinc zones within the main (southern) massive sulphide body below the zinc-depletion surface, broadly an 8% Zn to 10% Zn boundary, often based on grade contrast rather than raw grade
High-grade copper zones within the massive sulphide body above and below the zinc-depletion surface, broadly a 10% Cu boundary, often based on grade contrast rather than raw grade.

The initial domain models were subsequently divided above and below important sub-horizontal surfaces, such as the zinc-depletion surface and weathering/alteration surfaces, as appropriate.

Sub-domaining in terms of search ellipsoid orientation took place as part of the estimation strategy.

The domains defined are summarized in Table 14‑3.

14.1.5 Database Coding

Drilling intervals within the mineralized domains described above in Section 14.1.4 were flagged with a unique database code by domain.

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14.1.6 Compositing

Copper, zinc, gold, silver, and arsenic samples were composited at 2.0 m lengths downhole, extracted from the uniquely coded intercepts stored from the various zone code database tables. The downhole compositing process utilized a “best fit” approach, which resulted in composites of variable but equal length within each contiguous coded interval, ensuring that the composite length was as near as possible to the nominated length. Use of 2.0 m downhole composites significantly increases the composite data support relative to model block size.

14.1.7 Bulk Density

Bulk density values were predominantly assigned on the basis of rock type and oxidation state, as defined by the interpreted geological wireframes. The values are based on a combination of:

Bulk densities from the previous resource estimate
Some generic values for typical weathering profiles in felsic sediments at [upper] greenschist facies
Bulk density values from in situ measurements in use at the mine, derived over the last year.

This approach was taken based on examination of historical bulk density values in use at Bisha. Problems with the veracity of the density information previously used in the resource work, especially in oxide material and the very top of the supergene zones, were noted. Melnyk et al. (2012) identified a probable sample bias issue, with the material being sampled from competent drill core relating to denser pieces of rock, and probably not representative of the oxidation domain. Recent determinations from in situ measurements have reduced this problem, but some potential in friable material still remains. The oxidized domains are now mostly mined out.

A comparison of the assigned bulk density values for the main rock types, with those (median values) used for the 2012 resource model (Melnyk et al., 2012), is shown in Table 14‑2.

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Table 14‑2: Bulk Density Values Used for this Estimate vs. 2012 Resource Estimate

	Bulk Density (g/cm³)	2012 Resource Bulk Density (g/cm³)
*Weathering Domain*
Oxide Waste	2.0	2.01
Acid Waste	1.8	1.8
Trans/Supergene Waste	2.3	Not defined
Fresh Waste	2.7	2.62
*Geology/Mineralization Domain*
Alluvials	2.0	Not defined
Gossan	2.25	Not defined
Supergene Cu	4.0	4.13/4.66
Primary Zn	4.5	4.47/4.49
Pyrite Sand	3.95	Not defined
Hanging Wall Cu (kaolinitic)	2.0	1.98

14.1.8 Sample Type

Overall statistics for copper and arsenic for the supergene massive sulphide domains were examined where there was an overlap of DD and RC drilling to assess the validity of combining the different sample types (different sample support) in the estimation. The analysis showed that the populations are very similar (in terms of mean, median, CV, and skewness), and that it was appropriate to combine these datasets for the estimation.

The mean assay data for the RCGC samples was around 10% higher for arsenic and around 7% lower for copper compared to the DD samples in the same area. This was considered to be borderline in terms of being a significant difference. A number of factors may be influencing this difference, including:

Quality of DD data at the top of the supergene (core recovery/friable)
Information effect (a greater density of information from RCGC data compared to DD)
Spatial bias (RC data does not fully intersect the supergene, or cover the full area)
Laboratory bias (different laboratories using different assay methods).

These factors will be investigated as drill coverage increases and as the mine develops.

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14.1.9 Recovery

Analysis of diamond core recovery for the massive sulphide zones shows that there is no correlation between sample grade and core recovery. This is illustrated for copper in Figure 14‑5; the distribution for zinc is similar.

Figure 14‑5: Bisha Massive Sulphide Zones – Diamond Core Recovery vs. Copper (%)

Analysis of diamond core recovery for the small remnant oxide gold zones does show correlation between sample grade and core recovery. This is illustrated for silver in Figure 14‑6. As a result of this distribution, samples with core recovery below 60% were excluded from the estimation process and this peripheral zone classified in the Inferred category, as further described in Section 14.1.16. Additional drilling is planned in this area during 2014 to address this situation.

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Figure 14‑6: Bisha Remnant Oxide Zone – Diamond Core Recovery vs. Silver (g/t)

14.1.10 Exploratory Data Analysis

Sample distribution and population characteristics for the supergene and primary zones were examined for each of the major domains via histograms, probability plots, and Parrish analysis, to validate the overall domain controls on mineralization, and to determine whether further domaining was required for treatment of obvious statistical outliers. These methods identified multiple populations as well as the need to consider high-grade capping. An example of this can be found in the primary massive sulphide main zone, illustrated in Figure 14‑7 to Figure 14‑9. Parrish (decile) analysis indicates that high-grade capping should be considered and this is borne out in both grade distribution charts. The solution was mostly the formation of a separate high-grade domain within the massive sulphide, often removing the need for high-grade capping and constraining grade estimation within discrete pods. Where discrete zones of high-grade mineralization could not be interpreted, a high-grade cut was applied. This particularly applies to the gold and silver component within both supergene and primary zones. Table 14‑3 shows the treatment by domain.

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Figure 14‑7: Copper Distribution for the Primary Sulphide, Main Zone – Histogram

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Figure 14‑8: Copper Distribution for the Primary Sulphide, Main Zone– Log Probability

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Figure 14‑9: Parrish (decile) Analysis for the Primary Sulphide, Main Zone

A small tonnage of oxide material remains in the northeast of the deposit. Data are sparse in this peripheral area, and examination of the sample population showed a number of analytical problems and a relationship between core recovery and grade. For these reasons only a subset of the available samples were used in the estimation process with an associated low confidence in any estimate.

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Table 14‑3: Mineralization Domains and High-Grade Capping

Zone	Unit	Domain	High-Grade Cap
Oxide	NE remnant		Samples selected @ >60% recovery*
Oxide	Pyrite sand	North	10 g/t Au, 200 g/t Ag
Oxide	Pyrite sand	South	10 g/t Au, 200 g/t Ag
Supergene	Main Massive Sulphide	North	10 g/t Au, 200 g/t Ag
Supergene	Main Massive Sulphide	South – High Grade	10 g/t Au, 200 g/t Ag
Supergene	Main Massive Sulphide	South – Low Grade	10 g/t Au, 200 g/t Ag
Supergene	Hanging Wall Copper	Massive Sulphide component, north	10 g/t Au, 200 g/t Ag
Supergene	Hanging Wall Copper	Disseminated component	10 g/t Au, 200 g/t Ag
Supergene	Hanging Wall Copper	Massive Sulphide component, south	10 g/t Au, 200 g/t Ag
Supergene	Footwall copper	Footwall stringers – north	10 g/t Au, 200 g/t Ag
Supergene	Footwall copper	Footwall stringers – south	10 g/t Au, 200 g/t Ag
Primary	Main Massive Sulphide	Sub-grade north – copper	2% Cu; 3 g/t Au, 110 g/t Ag
Primary	Main Massive Sulphide	Low-grade north – copper	3 g/t Au, 110 g/t Ag
Primary	Main Massive Sulphide	High-grade north – copper	None
Primary	Main Massive Sulphide	Sub-grade south – copper	2% Cu
Primary	Main Massive Sulphide	Low-grade south – copper	3 g/t Au, 110 g/t Ag
Primary	Main Massive Sulphide	High-grade south – copper	None
Primary	Main Massive Sulphide	Sub-grade north – zinc	2% Zn
Primary	Main Massive Sulphide	Low-grade north – zinc	3 g/t Au, 110 g/t Ag
Primary	Main Massive Sulphide	High-grade north – zinc	None
Primary	Main Massive Sulphide	Sub-grade south – zinc	2% Zn
Primary	Main Massive Sulphide	Low-grade south – zinc	3 g/t Au, 110 g/t Ag
Primary	Main Massive Sulphide	High-grade south – zinc	None
Primary	Hanging Wall Copper	Massive Sulphide component, north	2% Cu, 2% Zn, 2 g/t Au, 110 g/t Ag
Primary	Hanging Wall Copper	Disseminated component	2% Cu, 2% Zn, 2 g/t Au, 110 g/t Ag
Primary	Hanging Wall Copper	Massive Sulphide component, south	2% Cu, 2% Zn, 2 g/t Au, 110 g/t Ag
Primary	Footwall copper	Footwall stringers – north	3 g/t Au, 110 g/t Ag
Primary	Footwall copper	Footwall stringers – south	3 g/t Au, 110 g/t Ag

Note: *Cap effectively removed observed analytical problems and outliers

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14.1.11 Variography

Variogram models all made use of the 2 m composites derived from the mineralized domains described in Section 14.1.4. In domains where data are sparse, datasets were combined (e.g., footwall stringer zones and sub-grade zones within the massive sulphide). Where data proved to be insufficient for variography, parameters for grade estimation were adopted from representative mineralization.

Oxide Zone

As described above, available data are sparse in this peripheral zone. Reasonable continuity was obtained by using the 045° direction of controlling shears as a guide.

Supergene Zone

Variography identified a direction of maximum spatial continuity approximately following the strike of the main southern massive sulphide zone (≈345°), the main northern zone (~0°), and the massive sulphide component of the hanging wall copper (≈045° parallel to the interpreted bounding faults). Within these planes, there was no strong preferred dip or plunge exhibited in the supergene domains. As work was confined to the relatively thin supergene zones above the zinc depletion, surface anisotropy in the semi-major (down dip) direction was hard to quantify. This resulted in isotropy in the semi-major and minor directions being applied in some zones.

In general, the major orientations observed in the supergene, especially the high-grade copper zones, reflect the location of the underlying high-grade zinc zones, and indicate a steep control on mineralization. This differs from the previous interpretation, which modelled the supergene as a relatively flat zone of limited vertical extent.

Primary Zone

As observed in the supergene zone, directions of maximum spatial continuity approximately follow the strike and dip of the massive sulphide zones. Given the dip extent of this mineralization, it was also possible to observe plunge directions within the mineralization with the principal plunge typically dipping some 25° to the north. An example is given in Figure 14‑10, with the magenta line indicating the plunge orientation.

The variogram parameters used for grade interpolation in the various mineralized zones are detailed in Table 14‑4 to Table 14‑31, inclusive.

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Figure 14‑10: Southern Massive Sulphide Zone – Directions of Spatial Continuity, Looking Northwest

Table 14‑4: Variogram Parameters, Arsenic in Supergene

Domain	Relative Nugget (%)	Structure	Relative Sill (%)	Range (m)	Isatis Rotation (Surpac Convention)	Major/ Semi-Major	Major/ Minor
Main Massive Sulphide (N)	16	Structure 1	34	20	0°/-75°/0°	1	1
Main Massive Sulphide (N)	16	Structure 2	50	150	0°/-75°/0°	1	5
Hanging Wall Copper Massive Sulphide (N)	11	Structure 1	89	30	40°/0°/0°	1	1
Hanging Wall Copper (S)	19	Structure 1	53	25	0°/0°/0°	1	1
Hanging Wall Copper (S)	19	Structure 2	28	75	0°/0°/0°	1	3
Pyrite Sand (N)	14	Structure 1	86	50	0°/0°/0°	1	5
Pyrite Sand (S)	14	Structure 1	86	50	0°/0°/0°	1	5
Hanging Wall Copper (N)	15	Structure 1	85	55	40°/0°/0°	1	1
Main Massive Sulphide (S, high grade)	12	Structure 1	30	25	345°/-75°/0°	1	1
Main Massive Sulphide (S, high grade)	12	Structure 2	58	250	345°/-75°/0°	1	1
Main Massive Sulphide (S, low grade)	12	Structure 1	30	25	345°/-75°/0°	1	1
Main Massive Sulphide (S, low grade)	12	Structure 2	58	250	345°/-75°/0°	1	1

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Table 14‑5: Variogram Parameters, Copper in Supergene

Domain	Relative Nugget (%)	Structure	Relative Sill (%)	Range (m)	Isatis Rotation (Surpac Convention)	Major/ Semi-Major	Major/ Minor
Main Massive Sulphide (N)	8	Structure 1	45	20	0°/-75°/0°	1	1
Main Massive Sulphide (N)	8	Structure 2	47	140	0°/-75°/0°	1	5
Hanging Wall Copper Massive Sulphide (N)	5	Structure 1	95	30	40°/0°/0°	1	1
Hanging Wall Copper (S)	3	Structure 1	22	15	0°/0°/0°	1	1
Hanging Wall Copper (S)	3	Structure 2	74	90	0°/0°/0°	1	6
Pyrite Sand (N)	10	Structure 1	90	50	0°/0°/0°	1	5
Pyrite Sand (S)	10	Structure 1	90	50	0°/0°/0°	1	5
Hanging Wall Copper (N)	16	Structure 1	84	50	0°/0°/0°	1	1
Main Massive Sulphide (S, high grade)	9	Structure 1	47	30	345°/-75°/0°	1	1
Main Massive Sulphide (S, high grade)	9	Structure 2	44	300	345°/-75°/0°	1	1
Main Massive Sulphide (S, low grade)	9	Structure 1	47	30	345°/-75°/0°	1	1
Main Massive Sulphide (S, low grade)	9	Structure 2	44	300	345°/-75°/0°	1	1

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Table 14‑6: Variogram Parameters, Gold and Silver in Supergene

Domain	Relative Nugget (%)	Structure	Relative Sill (%)	Range (m)	Isatis Rotation (Surpac Convention)	Major/ Semi- Major	Major/ Minor
Main Massive Sulphide (N)	35	Structure 1	35	20	0°/-75°/0°	1	1
Main Massive Sulphide (N)	35	Structure 2	30	110	0°/-75°/0°	1	5
Hanging Wall Copper Massive Sulphide (N)	33	Structure 1	67	30	40°/0°/0°	1	1
Hanging Wall Copper (S)	31	Structure 1	29	25	0°/0°/0°	1	1
Hanging Wall Copper (S)	31	Structure 2	40	75	0°/0°/0°	1	3
Pyrite Sand (N)	30	Structure 1	70	50	0°/0°/0°	1	5
Pyrite Sand (S) Hanging Wall Copper (N)	20	Structure 1	35	15	0°/0°/0°	1	1
Pyrite Sand (S) Hanging Wall Copper (N)	20	Structure 2	45	150	0°/0°/0°	1	1
Main Massive Sulphide (S, high grade)	13	Structure 1	25	10	10°/-50°/0°	1	4
Main Massive Sulphide (S, high grade)	13	Structure 2	61	50	10°/-50°/0°	1	4
Main Massive Sulphide (S, low grade)	28	Structure 1	47	10	345°/-75°/0°	1	4
Main Massive Sulphide (S, low grade)	28	Structure 2	25	40	345°/-75°/0°	1	4
Main Massive Sulphide (N)	28	Structure 1	54	10	345°/-75°/0°	1	1
Main Massive Sulphide (N)	28	Structure 2	18	50	345°/-75°/0°	1	1
Hanging Wall Copper Massive Sulphide (N)	35	Structure 1	56	10	345°/-75°/0°	1	1
Hanging Wall Copper Massive Sulphide (N)	35	Structure 2	9	30	345°/-75°/0°	1	1

Table 14‑7: Variogram Parameters, Arsenic in Primary

Domain	Relative Nugget (%)	Structure	Relative Sill (%)	Range (m)	Isatis Rotation (Surpac Convention)	Major/ Semi- Major	Major/ Minor
Hanging Wall Copper Massive Sulphide (S, low grade)	3	Structure 1	65	40	345°/-75°/0°	1	1
Hanging Wall Copper Massive Sulphide (S, low grade)	3	Structure 2	32	200	345°/-75°/0°	1	1
Main Massive Sulphide (N, low grade)	3	Structure 1	65	40	0°/-75°/0°	1	1
Main Massive Sulphide (N, low grade)	3	Structure 2	32	200	0°/-75°/0°	1	1
Hanging Wall Copper Massive Sulphide (N)	6	Structure 1	5	9	40°/0°/0°	1	1
Hanging Wall Copper Massive Sulphide (N)	6	Structure 2	90	55	40°/0°/0°	1	1
Hanging Wall Copper Massive Sulphide (S)	1	Structure 1	99	70	0°/0°/0°	1	1
Hanging Wall Copper Massive Sulphide (S, high grade)	3	Structure 1	65	40	345°/-75°/0°	1	1
Hanging Wall Copper Massive Sulphide (S, high grade)	3	Structure 2	32	200	345°/-75°/0°	1	1
Footwall Copper (N)	13	Structure 1	87	30	10°/-50°/0°	1	4
Footwall Copper (S)	27	Structure 1	73	30	345°/-75°/0°	1	2

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Table 14‑8: Variogram Parameters, Copper in Primary

Domain	Nugget (%)	Structure	Sill (%)	Range (m)	Supervisor Rotation (Surpac Convention)	Major/ Semi-Major	Major/ Minor
Massive Sulphide Sub-Grade (S)	0.01	Structure 1	0.01	50	165°/-14°/-65°	1	3.3
Massive Sulphide Sub-Grade (S)	0.01	Structure 2	0.02	130	165°/-14°/-65°	1	3.3
Massive Sulphide Sub-Grade (N)	0.01	Structure 1	0.01	50	215°/-40°/-40°	1	3
Massive Sulphide Sub-Grade (N)	0.01	Structure 2	0.02	80	215°/-40°/-40°	1	3
Hanging Wall Cu Low Grade (N)	0.01	Structure 1	0.01	50	215°/-40°/-40°	1	3
Hanging Wall Cu Low Grade (N)	0.01	Structure 2	0.02	80	215°/-40°/-40°	1	3
Hanging Wall Cu Low Grade (S)	0.01	Structure 1	0.01	85	345°/5°/50°	3	5
Hanging Wall Cu Low Grade (S)	0.01	Structure 2	0.06	160	345°/5°/50°	3	5
Massive Sulphide High Grade	5	Structure 1	2.8	55	340°/0°/0°	2	5
Massive Sulphide High Grade	5	Structure 2	50	95	340°/0°/0°	2	5
Massive Sulphide Low Grade (S)	0.05	Structure 1	0.28	35	150°/20°/-60°	1	2.5
Massive Sulphide Low Grade (S)	0.05	Structure 2	0.26	80	150°/20°/-60°	1	2.5
Massive Sulphide Low Grade (N)	0.05	Structure 1	0.28	35	150°/20°/-60°	1	2.5
Massive Sulphide Low Grade (N)	0.05	Structure 2	0.26	80	150°/20°/-60°	1	2.5
Footwall Copper (N)	0.05	Structure 1	0.27	40	165°/10°/-70°	1	1
Footwall Copper (N)	0.05	Structure 2	0.37	95	165°/10°/-70°	1	1
Footwall Copper (S)	0.05	Structure 1	0.27	40	165°/10°/-70°	1	1
Footwall Copper (S)	0.05	Structure 2	0.34	85	165°/10°/-70°	1	1

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Table 14‑9: Variogram Parameters, Zinc in Primary

Domain	Nugget (%)	Structure	Sill (%)	Range (m)	Supervisor Rotation (Surpac Convention)	Major/ Semi-Major	Major/ Minor
Massive Sulphide Sub-Grade (S)	0.005	Structure 1	0.005	40	170°/0°/-70°	1	1
Massive Sulphide Sub-Grade (S)	0.005	Structure 2	0.04	95	170°/0°/-70°	1	1
Massive Sulphide Sub-Grade (N)	0.005	Structure 1	0.02	30	185°/0°/-45°	1	1
Massive Sulphide Sub-Grade (N)	0.005	Structure 2	0.03	90	185°/0°/-45°	1	1
Hanging Wall Cu Zone Sub-Grade (N)	0.005	Structure 1	0.02	30	185°/0°/-45°	1	1
Hanging Wall Cu Zone Sub-Grade (N)	0.005	Structure 2	0.03	90	185°/0°/-45°	1	1
Hanging Wall Cu Zone Sub-grade (S)	0.005	Structure 1	0.02	30	185°/0°/-45°	1	1
Hanging Wall Cu Zone Sub-grade (S)	0.005	Structure 2	0.03	90	185°/0°/-45°	1	1
Massive Sulphide High Grade Zones	5	Structure 1	20	45	170°/5°/-70° *	1	4
Massive Sulphide High Grade Zones	5	Structure 2	13	130	170°/5°/-70° *	1	4
Massive Sulphide Low Grade (S)	0.5	Structure 1	2.7	55	170°/5°/-70°	1.3	3.8
Massive Sulphide Low Grade (S)	0.5	Structure 2	1.3	150	170°/5°/-70°	1.3	3.8
Massive Sulphide Low Grade (N)	0.5	Structure 1	2.7	55	170°/5°/-70°	1.3	3.8
Massive Sulphide Low Grade (N)	0.5	Structure 2	1.3	150	170°/5°/-70°	1.3	3.8
Hanging Wall Cu Zone Low Grade (N)	0.5	Structure 1	2.7	55	170°/5°/-70°	1.3	3.8
Hanging Wall Cu Zone Low Grade (N)	0.5	Structure 2	1.3	150	170°/5°/-70°	1.3	3.8
Hanging Wall Cu Zone Low Grade (S)	0.5	Structure 1	2.7	55	170°/5°/-70°	1.3	3.8
Hanging Wall Cu Zone Low Grade (S)	0.5	Structure 2	1.3	150	170°/5°/-70°	1.3	3.8
Footwall Zinc (S)	0.5	Structure 1	2.7	55	170°/5°/-70°	1.3	3.8
Footwall Zinc (S)	0.5	Structure 2	1.3	150	170°/5°/-70°	1.3	3.8

Note: * Different orientations were used for each high-grade sub-zone

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Table 14‑10: Variogram Parameters, Gold in Primary

Domain	Nugget (%)	Structure	Sill (%)	Range (m)	Supervisor Rotation (Surpac Convention)	Major/ Semi-Major	Major/ Minor
Massive Sulphide (S)	0.02	Structure 1	0.07	35	160°/25°/-70°	1	2
Massive Sulphide (S)	0.02	Structure 2	0.10	110	160°/25°/-70°	1	2
Massive Sulphide (N)	0.02	Structure 1	0.06	30	180°/0°/90°	1	2
Massive Sulphide (N)	0.02	Structure 2	0.11	130	180°/0°/90°	1	2
Hanging Wall Cu Zone (N)	0.02	Structure 1	0.06	30	180°/0°/90°	1	1
Hanging Wall Cu Zone (N)	0.02	Structure 2	0.11	130	180°/0°/90°	1	1
Hanging Wall Cu Zone (S)	0.02	Structure 1	0.12	40	175°/5°/-65°	1	1
Hanging Wall Cu Zone (S)	0.02	Structure 2	0.05	80	175°/5°/-65°	1	1
Massive Sulphide High Grade (S)	0.03	Structure 1	0.02	35	20°/55°/55°	1	3
Massive Sulphide High Grade (S)	0.03	Structure 2	0.09	100	20°/55°/55°	1	3
Footwall Zone (N)	0.01	Structure 1	0.04	10	10°/0°/-50° *	1	4
Footwall Zone (N)	0.01	Structure 2	0.01	50	10°/0°/-50° *	1	4
Footwall Zone (S)	0.06	Structure 1	0.12	10	345°/0°/-75° *	1	4
Footwall Zone (S)	0.06	Structure 2	0.06	40	345°/0°/-75° *	1	4

Table 14‑11: Variogram Parameters, Silver in Primary

Domain	Nugget (%)	Structure	Sill (%)	Range (m)	Supervisor Rotation (Surpac Convention)	Major/ Semi-Major	Major/ Minor
Massive Sulphide (S)	40	Structure 1	215	50	80°/0°/-70°	1	5
Massive Sulphide (S)	40	Structure 2	360	95	80°/0°/-70°	1	5
Massive Sulphide (N)	40	Structure 1	5	30	180°/0°/90°	1	1
Massive Sulphide (N)	40	Structure 2	215	90	180°/0°/90°	1	1
Hanging Wall Cu Zone (N)	40	Structure 1	5	30	180°/0°/90°	1	1
Hanging Wall Cu Zone (N)	40	Structure 2	215	90	180°/0°/90°	1	1
Hanging Wall Cu Zone (S)	4	Structure 1	15	30	185°/5°/-65°	1	1
Hanging Wall Cu Zone (S)	4	Structure 2	40	90	185°/5°/-65°	1	1
Massive Sulphide High Grade (S)	40	Structure 1	325	60	350°/0°/70°	1	2
Massive Sulphide High Grade (S)	40	Structure 2	390	90	350°/0°/70°	1	2
Footwall Zone (N)	1.8	Structure 1	1.3	10	170°/5°/-70° *	1	4
Footwall Zone (N)	1.8	Structure 2	4	130	170°/5°/-70° *	1	4
Footwall Zone (S)	60	Structure 1	195	10	170°/5°/-70° *	1	4
Footwall Zone (S)	60	Structure 2	135	130	170°/5°/-70° *	1	4

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14.1.12 Search Neighbourhood Analysis

The modelling process looks to characterize the spatial relationship of the data using variography, and then seeks to implement search strategies aimed at producing a robust block estimate whilst minimizing estimation error and conditional bias. The search neighbourhood was optimized using Quantitative Kriging Neighbourhood Analysis (QKNA), a process that analyzes Slope of Regression and Kriging weights for various search neighbourhoods within a mineralized domain.

QKNA was undertaken for several test blocks within each of the main supergene and primary mineralized domains, looking at and testing key criteria, as follows:

Block coordinates and dimensions
Estimated grade
Kriging variance
Block Dispersion variance
Slope of Regression of estimated blocks z*(v) and theoretical true blocks z(v) (Vann et al., 2003)
Block discretization
Minimum and maximum number of informing samples.

A number of factors were taken into consideration when deciding on an appropriate search strategy, including data spacing, variogram models, estimation quality, and resource classification.

Generally, in moderately to well-informed areas of the model, the slope of regression approached 1.0, indicating that the potential for conditional bias was minimal using the chosen search strategy. By the same token, minimizing the number of negative weights for a given number of informing composites was also desirable. During this process, search ellipse orientations for interpolation derived from variography reflected the direction of maximum continuity within the mineralized domains.

Based on the selected optimal search neighbourhoods, the numbers of composites required for interpolation were determined. For the supergene, global figures of a minimum of 6 composites and a maximum of 15 composites were applied. For the primary, global figures of a minimum of 6 composites and a maximum of between 14 and 16 composites were applied. There were no restrictions in the interpolation for the maximum number of composites in a particular drill hole.

14.1.13 Block Model Parameters

The block model set-up took into account various factors including the geometry of the mineralized zones and the spacing of the informing data. For Bisha an un-rotated model was created given the general strike of the mineralization. As a part of the QKNA process described in Section 14.1.12, optimal block sizes were also selected. The block model is defined as given in Table 14‑12.

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Table 14‑12: Block Model Definition

	Minimum	Maximum	Model Extent
Easting	338800	339950	1,150
Northing	1714750	1716550	1,800
RL	0	700	700
Parent Cell X m	10	Min Sub-Cell X m	2.5
Parent Cell Y m	10	Min Sub-Cell Y m	2.5
Parent Cell Z m	5	Min Sub-Cell Z m	2.5

14.1.14 Grade Estimation

Ordinary Kriging (OK) was used for estimation of all elements in all mineralized domains. Each domain was separately estimated using the unique set of composite samples associated with that domain.

The majority of the blocks within the domains were estimated in the first pass based on the criteria given in the descriptions above. A second pass of estimation was required for a number of domains to estimate grade into peripheral areas. This pass used more relaxed search criteria, typically an increase in the search range and a reduction in the minimum number of informing samples.

In a few areas, the second pass search failed to estimate a small number of blocks, and these were given an assigned grade based on the mean grade of the composites for the particular domain.

14.1.15 Block Model Validation

Visual and statistical validation of the arsenic, copper, zinc, gold, and silver grade estimates for Bisha demonstrate reasonably robust model outcomes.

In summary, the validation methods used were:

Swath plots comparing the estimated grade parameters against the mean declustered and raw composite grades
Comparison of estimated grade and mean composite grade by domain
Examination of estimation variables such as Slope of Regression and Kriging efficiency.

All comparisons were satisfactory. Some smoothing of grade was apparent, but this is as expected, and desirable from the use of ordinary kriging as the interpolation method.

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14.1.16 Classification of Mineral Resources

The Mineral Resource for Bisha is a global estimate representing a reliable estimate of the total contained metal, but the current block estimates are likely to vary as compared with the actual grade/tonnage distribution that will be achieved during selective mining and over short production periods.

The Bisha mineralization has been sufficiently drilled and sampled to allow classification as an Indicated and Inferred Mineral Resource in accordance with the current CIM Definition Standards for Mineral Resources and Mineral Reserves. As with any non-rigidly defined classification there will always be some blocks within categories that depart from defined criteria. The classification employed reflects a practical combination of both geological knowledge and estimation quality parameters that may be more numerical in nature. This approach to classification aims to avoid creating a complex, numerically-based “mosaic” distribution of classified blocks.

Classification considered a number of factors:

Continuity of geology and mineralization: a mineralized volume is firstly based on geological interpretation, which is subsequently modified by application of a cutoff grade. This ensures that the mineralized volume is not extrapolated unreasonable distances beyond data limits. The models created satisfy these criteria
Data spacing and sample data quality: data spacing must be sufficiently dense to ensure continuity as defined by such studies; the veracity of the data being used for the estimate must be established. The models created satisfy these criteria
Estimation techniques: sufficient statistical and associated studies have been undertaken to ensure that the methodology used for the Bisha estimate is appropriate.

The small remnant oxide zone has been classified as Inferred, reflecting the low confidence in geological and grade continuity that results from poor quality sample data at the extremities of the available drilling data.

The supergene and primary zones were largely classified in the Indicated category. The factors described above were combined into simplified shapes by resource categories that were then applied to the block model, avoiding a ‘patchy’ distribution of blocks of differing categories. For the primary zone, where data diminishes down dip, estimation parameters such as slope of regression guided the limits of material in the Indicated category.

14.1.17 Reasonable Prospects of Economic Extraction

Reasonable prospects for economic extraction were made by applying a net smelter return (NSR)-based cutoff to blocks within a constraining optimized pit shell. LG optimization was achieved using Gemcom Whittle v. 5.5.4 (Whittle).

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The assumed long-term metal prices used for the optimization work as applied to Mineral Resources are shown in Table 14‑13. These metal prices are approximately 15% higher than those used in the estimation of Mineral Reserves, as described in Section 15.

Table 14‑13: Mineral Resource Commodity Prices

Metal Prices	Price
Copper ($/lb)	3.35
Zinc ($/lb)	1.05
Gold ($/oz)	1,350
Silver ($/oz)	23

The NSR calculation and pit optimization process considers many of the parameters used in Mineral Reserve estimation, as these parameters are well established within the working mine. These parameters include commodity price, 2014 budget costs for production and processing, process recoveries, concentrate grade, selling costs, and other ore-based costs. The optimization process also uses the current geotechnical model for the pit design.

An optimized pit shell results from Whittle using the parameters described above and the commodity prices given in Table 14‑13. No pit design has been completed. Blocks within the pit shell and below the mined pit floor, as at 31 December 2013, are reported as Mineral Resources above NSR cutoffs of $40.55/t for oxide material and $39.55/t for supergene and primary material.

In addition, mineralization below the pit shell described here was studied for potential for underground mining. The high-level study concluded that an appropriate underground mining cost per tonne would be $60/t. The mineralization was then analyzed for contiguous blocks with NSR values exceeding $100/t. A shape of sufficient tonnage and value was identified, and this tonnage is included in the Inferred Resource category given in the Mineral Resource statement.

The constraining pit shell, mineralization, and underground Inferred Resource are shown in Figure 14‑11.

The Mineral Resource statement for Bisha Main is given in Table 14‑39 in Section 14.5.

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Figure 14‑11: Constraining Mineral Resource Pit Shell with Massive Sulphide Mineralization and Underground Mineral Resource, Perspective View Looking ENE

14.2 Harena Mineral Resource Estimate

14.2.1 Summary

No additional drilling has been undertaken for the Harena deposit since the last technical report. The only changes since the Mineral Resource estimate reported in May 2012 are:

Changes to the commodity prices used to create the optimized pit shell
Changes to ore-based and other costs relating to NSR calculation as described in Section 14.1.17
Mining depletion of most of the oxide zone.

Blocks within the pit shell and below the mined pit floor, as at 30 June 2013, are reported as Mineral Resources above NSR cutoffs of $42.41/t for oxide material, and $41.41/t for supergene and primary material. No change was made to the Resource classification within the model. No mining has taken place since 30 June 2013.

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The constraining pit shell (in green), mineralized zones (in blue and yellow), and as-mined pit (in brown) are shown in a cutaway view in Figure 14‑12, demonstrating mining of most of the oxide material.

The Mineral Resource statement for Harena is given in Table 14‑40.

For additional detail, please refer to AGP Mining Consultants May 2012 report, “Bisha Polymetallic Operation, Eritrea, Africa: NI 43-101 Technical Report for Nevsun Resources.”

Figure 14‑12: 3D Cutaway View of Constraining Mineral Resource Pit Shell, with Mineralization and As-Mined Pit; Looking Northeast

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14.3 Northwest Estimate

14.3.1 Project Sample Database

The drilling database forms the basis of the estimate containing all the available and reliable assay data as of the 18 October 2013.

A total of 43,807 m of diamond drilling has been completed in a series of programs completed by BMSC since 2003, and the majority of this data are available for the Northwest estimate as detailed below in Table 14‑14.

Table 14‑14: Northwest Drill Holes and Sample Metres at 18 October 2013

Data Type	Hole Prefix	No. of Holes	Total Drill (m)	Total Sampled (m)	Average Sample Length (m)	Date Drilled
Data Type	Hole Prefix	No. of Holes	Total Drill (m)	Total Sampled (m)	Average Sample Length (m)	From	To
Diamond (DD)	B	14	2,033.5	1,393.5	1.6	Oct-03	Oct-03
	NW	26	4,567.9	3,636	1.5	Feb-05	Mar-06
	NW	73	13,619.6	13,177.9	1.1	Nov-11	Oct-12
	NWDD	106	21,546.8	20,479	1.0	Nov-12	Aug-13
Geotechnical	NWGT	8	1,568.1	310.4	1.3	Nov-12	Dec-12
Metallurgical	NWMT	2	471.0	463.5	1.2	Oct-12	Nov-12
		229	43,807	39,460.2	1.1
Trench*	NWTR	16	1,118.0	976.2	2.0	Nov-12	Jan-13
RC*	NWA	47	872.5	827.5	1.9	?	?

Note:

*The assay results and mapping from the trenches excavated in 2012–2013 were used to help define structural and mineralization trends at the surface. The trench and shallow RC assay data were used for the estimation of the oxide gold mineralization only.

14.3.2 Local Grid Transformation

Geology and mineralization strike at approximately 45° to the UTM true grid. The majority of the drilling to date has been drilled perpendicular to the mineralization on UTM grid azimuths of 135° and 315°. A local grid was established parallel to the drilling sections to facilitate the sectional drill interpretation and simplify resource estimation and block modelling.

Two common points were established to enable rotation of all the drilling data into the local grid as detailed below in Table 14‑15.

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Table 14‑15: UTM to Local Grid Transformation – Common Points

UTM Northing	UTM Easting	Local Northing	Local Easting	Comment
1717700	338100	20000	10000	Origin Point
1717900	338900	20707.107	10424.264	Point3

14.3.3 Geological Model

Introduction

The mineralized components of the Northwest deposit are divided into three major zones, namely oxide, supergene, and primary zones, with the latter dominant. A separate, narrower, massive sulphide zone lies to the east, running parallel to the main sulphide bodies.

The massive sulphide mineralized bodies of the primary zones are divided into a southern, a central (the Main Zone), and a northern zone. These zones are disrupted by fault or shear zones whose displacement is not defined, but the central zone is thought to represent the downthrown graben zone. The massive sulphides have associated semi-massive and stringer sulphide zones, with the latter best developed in the footwall. The massive sulphide zones strike at approximately 045°, dipping steeply to the northwest, with overall strike and dip lengths of some 800 m and 230 m, respectively. The eastern zone has similar dimensions, but lacks associated stringer sulphide development. This zone is not as well defined as the main zones. The mineralized zones are shown in Figure 14‑13.

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Figure 14‑13: Modelled Sulphide Domains – Sliced at 480 mRL, approx. 80 m Vertical Depth, Plan View

Methodology

The key geological components of the VHMS mineralization at Northwest have been progressively modelled since the infill and extensional diamond drilling program commenced in November 2012. The interpretation has been continually updated as more drilling data has become available.

Initial interpretation was on 25 m spaced hard copy E-W sections, and was accompanied by systematic re-logging of diamond core to resolve areas of geological uncertainty and complexity. The sections were digitized and 3D wireframes were created in Surpac. The resultant solid 3D models represent the main geological components controlling the mineralization, and are the basis for establishing grade continuity and the estimation of the Mineral Resource.

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Models

Sulphide mineralization as defined by logged visual pyrite percentage is the fundamental control on the base metal distribution and forms the basis for the geological framework. Additional geochemical indicators (such as arsenic, barium, zinc, iron, the Ishikawa alteration index, and density) were also used as secondary criteria to help define the orientations of best continuity for robust and geometrically simple geology solids.

The geochemical indicators were used to support and increase the confidence of the geological interpretation, particularly across mineralization and alteration boundaries. Each of the geochemical indicators has a distinct range of values that characterizes the three sulphide mineralization domains, and is particularly useful where the contacts are diffuse or the logging is inconclusive. The criteria developed and used for geological modelling are summarized in Table 14‑16.

Table 14‑16: Sulphide Mineralization Interpretation Criteria

Geological Domain	Litho Code	Logged Py	Fe %	As ppm	Ba ppm	Density (g/cm³)	Ishikawa Alteration Index
Massive Sulphide	MSUL	>50%	>35%	>200 ppm	NA	>3.7	NA
Semi-Massive Sulphide	SMSX	25%–50%	20%–35%	50–200 ppm	NA	2.9–3.7	NA
Stringer	STSX	5%–25%	10%–20%	<50 ppm	<500	2.5–2.9	>90

Three main types of dykes have been logged and interpreted:

	1.	North-south-trending mafic to intermediate dykes, with thicknesses ranging from 1 m to 5 m
	2.	Moderately west-dipping intermediate dykes within the footwall alteration zone, with average widths of 1 m to 5 m
	3.	Cross-cutting, sub-vertical, felsic quartz-feldspar dykes (QFP) trending W-NW, up to 10 m in true thickness, and offsetting mineralized zones.

The dykes indicate zones of faulting and structural discontinuity. The dominant mafic hanging wall dyke swarm marks the stratigraphic top of the mineralizing system and has developed within a shear zone. This sub-vertical, 5 m to 10 m wide structural zone is continuous along the western edge of the massive sulphide body for most of the strike length of the Northwest mineralization. The dykes are shown in Figure 14‑14.

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Figure 14‑14: Interpreted Dykes – Mafic (blue), Intermediate (green), QFP (magenta), Plan View

A number of structural zones were interpreted from the trends established from surface mapping and trenching, with targeted core re-logging and core photo examination used to confirm the structural interpretation. These zones are highlighted by offsets, rapid changes in orientation, and terminations of the massive sulphide body, particularly to the south where the South and Main Zones are separated by an area of discontinuous and irregular sulphide pods.

The identified faulting has preferred orientations of NW-SE to NNW-SSW, with a dominant dextral strike slip movement inferred.

Two main fault zones have been interpreted that significantly disrupt and displace the massive sulphide body, being the Zbei and Hbei Fault Zones. The latter is also interpreted as continuing southwards joining with a large boundary shear lying to the west of the Bisha deposit.

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Other faults have also been interpreted, and all structures are shown in Figure 14‑15.

Where possible, lithologies have also been modelled. Although this was hampered by strong alteration and inconsistencies in logging, two rhyolite plugs have been interpreted in the hanging wall to the north and the south. A laminated mudstone has also been identified in the hanging wall adjacent to the Main massive sulphide domain.

Figure 14‑15: Interpreted Structural Features and Massive Sulphide Domains, Plan View

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14.3.4 Alteration and Weathering

Weathering surfaces, based upon lithology logs, core photos, and mineral assemblages, were constructed on 25 m spaced cross-sections to define the main weathering and oxidation domains:

Oxide: above the Base of Complete Oxidation (BOCO) surface the rock has been weathered to the extent that all sulphides are oxidized with no primary or secondary sulphides present, and the original colour of the fresh rock is unrecognizable.
Acid Leach Zone (Soap): a zone of intense acid leaching, predominantly above and adjacent to the massive sulphide bodies, which destroys the primary texture of the original rock leaving a clay-silica residue which, when wet, has a greasy to ‘soapy’ nature. The zone is below the BOCO surface and decomposed sulphides can be present towards the base of this unit, as disseminated chalcocite and covellite.
Transitional (Supergene): all material lying above the Top of Fresh Rock (TOFR) surface and below the base of the acid leach zone. Within this zone the colour and texture of the fresh rock is recognizable, but partial weathering and discolouration of the rock substance has occurred. The zone is the transition between oxide and sulphide minerals, and contains both primary (chalcopyrite) and secondary (chalcocite/covellite) copper minerals.
Fresh: all material below the TOFR, where the rock is unaffected by weathering, with no evidence of discolouration, and all sulphide minerals are primary (i.e., chalcopyrite, arsenopyrite, etc.).

14.3.5 Mineralization Domains

Copper and Zinc

The estimation domains for copper and zinc were interpreted independently using the underlying geology models to establish the mineralization orientation and continuity from section to section. As the copper and zinc mineralization is known to be gradational in places across the massive sulphide boundaries, the resultant estimation domains can comprise a mixture of sulphide mineralization types. For the purposes of grade estimation and ensuring that the grade population distribution is preserved, this was not considered to affect the integrity of estimation.

Copper domains were modelled at a nominal 0.3% cutoff; domains for the stringer sulphides were also created in which values were estimated separately from the massive sulphide domains.

In a similar manner, zinc domains were modelled at a nominal 0.5% cutoff, again with a separate halo of surrounding stringer mineralization.

The copper and zinc domains are shown in Figure 14‑16.

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Arsenic, Iron, and Lead

The domains for arsenic and iron were interpreted to be the same as the massive sulphide models created in the geological modelling process. Lead was estimated using the zinc domains.

Figure 14‑16: Mineralization Domains for Copper (left) and Zinc (right), Plan View

Gold and Silver

Domains for gold mineralization were interpreted separately above and below the ‘soap’ interface, as the distribution of and control on gold mineralization are distinctly different between the oxide/supergene and primary material. Silver mineralization has a strong spatial correlation with gold, and has used the same domains for compositing.

The Supergene gold domain lies directly above the massive sulphide, mainly within and straddling the acid leach interface (‘soap’ surface) and appears to be unrelated to faulting. The higher gold grades are developed over the thickest portions of the central massive sulphide within ’soap’ material.

An upper low-grade (0.5 g/t Au to 1.5 g/t Au) saprolite (oxide) gold zone occurs above the supergene gold zone, and is developed as a sub-horizontal, near-surface band (5 m to 10 m thick) within 20 m of the surface. This zone is typically highly weathered and ferruginous and lies above the BOCO surface.

Gold wireframes were interpreted using a nominal 0.2 g/t cutoff within the oxide and soap material. Some mineralization could not be included due to lack of continuity.

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The semi-massive sulphide domains were utilized for the estimation of the broad low-grade gold and silver mineralization below the soap weathering domain. Several distinct and continuous higher-grade gold-silver zones were interpreted sub-parallel to the massive sulphide hangingwall and footwall contacts, and were modelled and estimated separately.

The gold and silver domains are shown in Figure 14‑17.

Figure 14‑17: Mineralization Domains: Oxide and Supergene Gold (left), High-Grade Primary Gold (right), Plan View

14.3.6 Database Coding

Drilling intervals within the mineralized domains described above in Section 14.3.5 were flagged with a unique database code.

Only diamond holes were used for the estimate, excluding those holes that were abandoned. The exception was the oxide/supergene gold estimate, where the shallow RC grade control holes and the trench assay data were also used.

14.3.7 Compositing

Copper, zinc, gold, silver, arsenic, lead, and iron samples were composited at 2.5 m lengths downhole, extracted from the uniquely coded intercepts stored from the various zone code database tables. The downhole compositing process utilized a best fit approach, which resulted in composites of variable but equal length within each contiguous drill hole-coded interval, ensuring that the composite length was as near as possible to the nominated length. Composites of less than 1 m were rejected. Use of 2.5 m downhole composites significantly increases the composite data support relative to model block size. All domains were further split by the soap and other weathering surfaces, with population distributions demonstrably distinct across the soap surface.

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14.3.8 Bulk Density

A total of 11,674 bulk density measurements were completed by BMSC using the water immersion method (Archimedes principle). The density samples were 10 cm to 20 cm segments of full core samples selected from the core trays at 4 m intervals down the drill hole. Density measurements taken represent more than 95% of the total metres drilled and sampled at Northwest, giving very good coverage.

The density sample data extracted from the supplied database was analyzed statistically and graphically with respect to lithology, weathering, and mineralization.

Estimating the bulk density below the soap weathering domain into the various sulphide domains was considered to be the most appropriate strategy. This estimation allowed for local variations in bulk density including less dense material in the transitional weathering domains which show a diffuse and irregular contact with the fresh material. The gradational nature of the TOFR surface does not warrant a hard boundary domain approach. Where unestimated, bulk density values were assigned. The density sample data statistics and assigned bulk densities are shown in Table 14‑17 and Table 14-18, respectively.

14.3.9 Recovery

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Figure 14‑18: Northwest Massive Sulphide Zones – Diamond Core Recovery vs. Copper Grade

Within the oxide and supergene zones core recovery has been unsatisfactory. The average core recovery within the supergene zone is 35%, with discontinuous and restricted zones of high-grade gold. There appears to be a bias where the highest gold-grade zones are related to the low core recovery areas, indicating that there is an upgrading of the gold grade caused by washing out of the fine sand/silt, and the retention of the more competent (silicified?) cored zones with a higher tenor of gold mineralization. For this reason, the confidence level when estimating these gold domains is low, and is reflected in the classification of these zones in the Inferred category. An RC drill program is planned to improve sample recovery and establish better spatial continuity, with the prospect of upgrading the Resource category once this work is completed.

Table 14‑17: Bulk Density Sample Data Statistics by Estimation Domain

Estimation Domain	No. of Samples	Min.	Max.	Mean	Median	CV
Massive Sulphide	1,155	1.27	5.11	4.05	4.25	0.15
Semi-Massive Sulphide	767	1.44	4.70	2.91	2.85	0.15
Stringer Sulphide	3,624	1.26	5.48	2.66	2.62	0.12
Outside Stringer (fresh)	3,010	1.41	5.32	2.49	2.50	0.08
Outside Stringer (transitional)	736	1.22	3.38	2.27	2.29	0.11

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Table 14‑18: Assigned Bulk Density Values

Bulk Density Domain	Assigned Value
Massive Sulphide (unestimated sub-domain 5200)	4.0
Stringer Sulphide (unestimated)	2.66
Outside Stringer (fresh)	2.5
Outside Stringer (transitional)	2.3
Soap	1.9
Oxide	1.6
Pyritic Sands (below Soap & inside mz_nw_au_supg20130409.dtm)	2.8
Air (above topography)	0
Dykes (transitional)	2.2
Dykes (fresh)	2.6

14.3.10 Exploratory Data Analysis

Sample distribution and population characteristics for the oxide, supergene, and primary zones were examined for the major domains via histograms, probability plots, and Parrish analysis, to validate the overall domain controls on mineralization, and to determine whether further domaining was required for treatment of obvious statistical outliers. These methods identified multiple populations as well as the need to consider high-grade capping. The high-grade capping strategy is summarized in Table 14‑19 to

Table 14‑25. The domain numbers refer to the numerical system applied to each of the mineralized domains, by element.

Table 14‑19: Copper High-Grade Assay Cuts by Estimation Domain

Domain	No. of Composites	Uncut Mean Cu (%)	Applied High Cut Cu (%)	Number Cut	Percentile Cut	Cut Mean Cu (%)	Cut CV
101	184	1.29	10	3	≈98^th	1.10	1.4
105	1,453	0.82	5	2	99^th+	0.82	0.8
190	585	0.06	2	4	99^th+	0.06	2.3
999 Soap	1,277	0.05	1	8	99^th+	0.04	2.8
999 Oxide	2,990	0.02	0.2	5	99^th+	0.02	1.0

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Table 14‑20: Zinc High Grade Assay Cuts by Estimation Domain

Domain	No. of Composites	Uncut Mean Zn (%)	Applied High Cut Zn (%)	Number Cut	Percentile Cut	Cut Mean Zn (%)	Cut CV
201	268	2.87	12	5	≈98^th	2.81	1.0
202	28	1.39	3	1	98^th	1.32	0.5
220	3	4.2	3	1	90^th	1.94	0.1
290	6,903	0.07	3	3	99^th+	0.07	2.1
999 Soap	1,277	0.03	0.5	7	99^th+	0.03	2.1
999 Oxide	2,990	0.02	0.3	6	99^th+	0.02	1.5

Table 14‑21: Lead High Grade Assay Cuts by Estimation Domain

Domain	No. of Composites	Uncut Mean Pb (%)	Applied High Cut Pb (%)	Number Cut	Percentile Cut	Cut Mean Pb (%)	Cut CV
211	16	0.23	0.5	1	98^th	0.21	0.6
290	6,852	0.01	0.3	9	99^th+	0.01	3.5
999 Soap	802	0.02	0.2	11	99^th+	0.02	1.4
999 Oxide	2,596	0.02	0.4	4	99^th+	0.02	1.7

Table 14‑22: Gold High Grade Assay Cuts by Estimation Domain

Domain	No. of Composites	Uncut Mean Au (g/t)	Applied High Cut Au (g/t)	Number Cut	Percentile Cut	Cut Mean Au (g/t)	Cut CV
900	80	0.71	3	3	97^th	0.63	1.2
910	191	0.70	7	3	≈98^th	0.61	2.0
950	336	3.0	25	8	≈98^th	2.06	2.3
999	3,990	0.05	1	11	99^th+	0.05	1.9
9100	401	0.37	5	2	99^th+	0.33	1.7
9110	119	0.21	1	5	≈96^th	0.19	1.3
9400	62	0.27	1	6	90^th	0.25	1.2
9500	2,023	0.31	10	6	99^th+	0.22	0.1
9600	4,623	0.09	2	23	99^th+	0.08	0.1
9904	30	5.14	25	2	96^th	3.97	1.5
9932	27	8.16	25	3	96^th	6.54	1.2
9936	27	3.54	25	1	99^th+	2.11	2.0

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Table 14‑23: Silver High Grade Assay Cuts by Estimation Domain

Domain	No. of Composites	Uncut Mean Ag (g/t)	Applied High Cut Ag (g/t)	Number Cut	Percentile Cut	Cut Mean Ag (g/t)	Cut CV
900	80	2.6	10	5	≈95^th	2.2	1.2
910	161	57.3	20	4	98^th	2.9	1.2
920	16	8.5	20	2	90^th	5.5	1.3
950	187	22.1	200	5	99^th	18.5	1.4
999	3,711	2.3	100	4	99^th+	2.0	2.6
9500	4,100	3.09	50	5	99^th+	2.9	1.9
9904	34	20.6	50	1	99^th+	17.3	0.7
9932	56	15.4	50	1	99^th+	12.4	1.5
9936	54	13.9	50	1	99^th+	8.5	1.8
9952	26	12.9	50	1	99^th+	11.5	1.8

Table 14‑24: Arsenic High Grade Assay Cuts by Estimation Domain

Domain	No. of Composites	Uncut Mean As (ppm)	Applied High Cut As (ppm)	Number Cut	Percentile Cut	Cut Mean As (ppm)	Cut CV
1000	205	557	3,000	1	99^th+	550	1.0
6000	5,785	51	400	4	99^th+	51	0.9
999 Soap	2,596	107	700	2	99^th+	106	1.0
999 Oxide	2,596	172	1,500	1	99^th+	172	1.0

Table 14‑25: Iron High Grade Assay Cuts by Estimation Domain

Domain	No. of Composites	Uncut Mean Fe (%)	Applied High Cut Fe (%)	Number Cut	Percentile Cut	Cut Mean Fe (%)	Cut CV
999 Soap	1,256	7.6	30	21	98^th	7.5	0.9

14.3.11 Variography

Variography was undertaken using Isatis software on the cut 2.5 m composites for copper, zinc, lead, gold, silver, arsenic, and iron. Variography for bulk density was undertaken on the sample data. The most informed domains were used for variography as shown in Table 14‑26.

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Table 14‑26: Domain Groupings

Group	Cu Domains	Zn-Pb Domains	Massive Sulphide Domain	Semi-Massive Sulphide Domain	Au-Ag Domains
North Zone	105, 107, 5200	201, 202–220 (all)	5000, 1200, 3000, 4000, 5100, 5200	9500, 9300, 9400, 5200	950, 900, 910, 920, 999 (Oxide/Supg.) All Primary Gold 9904, 9932, 9936, 9952,9953, 9955
Central Zone	105, 103,104, 106, 111		5000, 1200, 3000, 4000, 5100, 5200	9500, 9300, 9400, 5200
South Zone	101,102		1000, 1100, 2000	9100, 9110, 9200
Stringer Mineralization	190	290	6000	9600	9600

Note: bold indicates a domain on which variography was completed.

The best spatial continuity is generally orientated parallel to the strike of the mineralized zones, with the Central Zone domains showing a distinct plunge to the north. Minor modifications were made to the variogram orientations and anisotropy parameters used in the estimates to better suit the orientation of sub-regions within the mineralization domains. The resulting estimation parameters are shown in Table 14‑27 to Table 14‑31, inclusive.

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Table 14‑27: Copper Variogram Parameters

Zonecode	Nugget	Spherical 1				Spherical 2
Zonecode	Relative sill	Relative sill	Major Axis Range (m)	Major/ Semi-Major	Major/ Minor	Relative sill	Major Axis Range (m)	Major/ Semi-Major	Major/Minor	Plunge	Dip
101 fresh	0.11	0.16	45.0	1	4	0.73	70	1	4	0	85
101 trans	0.11	0.16	45.0	1	4	0.73	70	1	4	0	85
102	0.11	0.16	45.0	1	4	0.73	70	1	4	0	80
103	0.16	0.19	15.0	2	4	0.65	100	2	4	0	75
104	0.16	0.19	15.0	2	4	0.65	100	2	4	0	85
105 north fresh	0.17	0.32	15.0	1	4	0.51	65	1	4	0	70
105 north trans	0.17	0.32	15.0	1	4	0.51	65	1	4	0	70
105 central fresh	0.16	0.19	15.0	2	4	0.65	100	2	4	-25	80
105 central trans	0.16	0.19	15.0	2	4	0.65	100	2	4	-25	80
106	0.16	0.19	15.0	2	4	0.65	100	2	4	0	85
107 fresh	0.17	0.32	15.0	1	4	0.51	65	1	4	0	80
107 trans	0.17	0.32	15.0	1	4	0.51	65	1	4	0	80
111	0.16	0.19	15.0	2	4	0.65	100	2	4	0	90
5200 (EMS)	0.17	0.32	15.0	1	4	0.51	65	1	4	0	85
190 below soap fresh	0.26	0.26	6.0	1	4	0.48	75	1	4	0	75
190 below soap trans	0.26	0.26	6.0	1	4	0.48	75	1	4	0	75
999 soap	0.41	0.59	90.0	1	1	-	-	-	-	0	0
999 oxide	0.23	0.43	8.0	1	1	0.34	80	1	1	0	0

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Table 14‑28: Zinc Variogram Parameters

Zonecode	Nugget	Spherical 1				Spherical 2
Zonecode	Relative sill	Relative sill	Major Axis Range (m)	Major/ Semi- Major	Major/ Minor	Relative sill	Major Axis Range (m)	Major/ Semi-Major	Major/ Minor	Plunge	Dip
201	0.21	0.36	6.0	1	4	0.42	45	1	4	0	85
202	0.21	0.36	6.0	1	4	0.42	45	1	4	0	85
203	0.21	0.36	6.0	1	4	0.42	45	1	4	0	90
204	0.21	0.36	6.0	1	4	0.42	45	1	4	0	90
205	0.21	0.36	6.0	1	4	0.42	45	1	4	0	75
206	0.21	0.36	6.0	1	4	0.42	45	1	4	0	85
207	0.21	0.36	6.0	1	4	0.42	45	1	4	0	75
208	0.21	0.36	6.0	1	4	0.42	45	1	4	0	75
209	0.21	0.36	6.0	1	4	0.42	45	1	4	0	75
210	0.21	0.36	6.0	1	4	0.42	45	1	4	0	75
211 (EMS)	0.21	0.36	0.0	1	4	0.42	45	1	4	0	85
220	0.21	0.36	6.0	1	4	0.42	45	1	4	0	85
290 below soap	0.16	0.43	9.0	1	4	0.41	28	1	4	0	75
999 soap	0.21	0.25	20.0	1	1	0.54	60	1	1	0	0
999 oxide	0.21	0.25	20.0	1	1	0.54	60	1	1	0	0

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Table 14‑29: Gold Variogram Parameters

Zonecode	Nugget	Spherical 1				Spherical 2
Zonecode	Relative sill	Relative sill	Major Axis Range (m)	Major/ Semi- Major	Major/ Minor	Relative sill	Major Axis Range (m)	Major/ Semi- Major	Major/ Minor	Plunge	Dip
9904	0.2	0.55	4	1	2	0.25	25	1	2	0	70
9932	0.2	0.55	4	1	2	0.25	25	1	2	0	85
9936	0.2	0.55	4	1	2	0.25	25	1	2	0	75
9952	0.2	0.55	4	1	2	0.25	25	1	2	0	65
9953	0.2	0.55	4	1	2	0.25	25	1	2	0	70
9955	0.2	0.55	4	1	2	0.25	25	1	2	0	75
9100	0.15	0.52	8	1	2	0.33	50	1	2	0	80
9110	0.15	0.52	8	1	2	0.33	50	1	2	0	85
9200	0.15	0.52	8	1	2	0.33	50	1	2	0	85
9300	0.15	0.52	8	1	2	0.33	50	1	2	0	70
9400	0.15	0.52	8	1	2	0.33	50	1	2	0	65
9500	0.15	0.52	8	1	2	0.33	50	1	2	0	75
9600 (below soap)	0.15	0.52	8	1	2	0.33	50	1	2	0	75
5200	0.15	0.52	8	1	2	0.33	50	1	2	0	85
900 (oxide)	0.19	0.41	15	1	1	0.4	50	1	1	0	0
910 (oxide)	0.19	0.41	15	1	1	0.4	50	1	1	0	0
920 (oxide)	0.19	0.41	15	1	1	0.4	50	1	1	0	0
950 (supergene)	0.19	0.41	15	1	1	0.4	50	1	1	0	0
999 (above soap)	0.19	0.41	15	1	1	0.4	50	1	1	0	0

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Table 14‑30: Silver Variogram Parameters

Zonecode	Nugget	Spherical 1				Spherical 2
Zonecode	Relative sill	Relative sill	Major Axis Range (m)	Major/ Semi- Major	Major/ Minor	Relative sill	Major Axis Range (m)	Major/ Semi- Major	Major/ Minor	Plunge	Dip
9904	0.27	0.31	4	1	1	0.42	30	1	1	0	70
9932	0.27	0.31	4	1	1	0.42	30	1	1	0	85
9936	0.27	0.31	4	1	1	0.42	30	1	1	0	75
9952	0.27	0.31	4	1	1	0.42	30	1	1	0	65
9953	0.27	0.31	4	1	1	0.42	30	1	1	0	70
9955	0.27	0.31	4	1	1	0.42	30	1	1	0	75
9100	0.27	0.35	7	1	2	0.38	60	1	2	0	80
9110	0.27	0.35	7	1	2	0.38	60	1	2	0	85
9200	0.27	0.35	7	1	2	0.38	60	1	2	0	85
9300	0.29	0.25	10	1	2	0.46	130	1	2	0	70
9400	0.29	0.25	10	1	2	0.46	130	1	2	0	65
9500	0.29	0.25	10	1	2	0.46	130	1	2	0	80
9600 (below soap)	0.29	0.25	10	1	2	0.46	130	1	2	0	75
5200	0.29	0.25	10	1	2	0.46	130	1	2	0	80
900 (oxide)	0.19	0.41	15	1	1	0.4	50	1	1	0	0
910 (oxide)	0.19	0.41	15	1	1	0.4	50	1	1	0	0
920 (oxide)	0.19	0.41	15	1	1	0.4	50	1	1	0	0
950 (supergene)	0.19	0.41	15	1	1	0.4	50	1	1	0	0
999 (above soap)	0.19	0.41	15	1	1	0.4	50	1	1	0	0

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Table 14‑31: Arsenic Variogram Parameters

Zonecode	Nugget	Spherical 1				Spherical 2
Zonecode	Relative sill	Relative sill	Major axis range (m)	Major/ semi- major	Major/ minor	Relative sill	Major axis range (m)	Major/ semi- major	Major/ minor	Plunge	Dip
1000	0.33	0.37	10	1	4	0.3	50	1	4	0	85
1100	0.33	0.37	10	1	4	0.3	50	1	4	0	85
1200	0.09	0.35	35	2	4	0.56	130	2	4	0	90
2000	0.33	0.37	10	1	4	0.3	50	1	4	0	85
3000	0.09	0.35	35	2	4	0.56	130	2	4	0	70
4000	0.09	0.35	35	2	4	0.56	130	2	4	0	65
5000	0.09	0.35	35	2	4	0.56	130	2	4	-25	75
5100	0.09	0.35	35	1	1	0.56	130	1	1	0	85
5200	0.09	0.35	35	2	4	0.56	130	2	4	0	85
999 (oxide)	0.24	0.42	12	1	1	0.34	75	1	1	0	0
999 (soap)	0.24	0.42	12	1	1	0.34	75	1	1	0	0
6000 (below soap)	0.24	0.42	12	1	3	0.34	75	1	3	0	75

14.3.12 Search Neighbourhood Analysis

Search neighbourhood analysis was undertaken as generally described in Section 14.1.2.

Based on the selected optimal search neighbourhoods, the number of composites required for interpolation at Northwest was determined. Global figures of a minimum of 4 composites, and a maximum of 15 composites were applied. There were no restrictions in the interpolation for the maximum number of composites in a particular drill hole.

14.3.13 Block Model Parameters

The block model set-up took into account various factors including the geometry of the mineralized zones and the spacing of the informing data. For Northwest an un-rotated (local grid) model was created parallel to the general strike of the mineralization. The block model is defined as given in Table 14‑32.

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Table 14‑32: Block Model Definitions

	Minimum	Maximum	Model Extent
Easting	10,000	10800	800
Northing	19,870	21010	1,140
RL	200	660	460
Parent Cell X m	5	Min. Sub-Cell X m	1.25
Parent Cell Y m	10	Min. Sub-Cell Y m	2.5
Parent Cell Z m	10	Min. Sub-Cell Z m	2.5

14.3.14 Grade Estimation

OK was used for estimation of all elements in all mineralized domains. Each domain was separately estimated using the unique set of composite samples associated with that domain.

The majority of the blocks within the domains were estimated in a first pass, based on the criteria given in the descriptions above. A second pass of estimation was required for a number of domains to estimate grade in peripheral areas. This pass used more relaxed search criteria, typically an increase in the search range and a reduction in the minimum number of informing samples.

In a few areas, the second pass search failed to estimate a small number of blocks. Such blocks within the interpolated mineralized domain, and in the surrounding ‘waste’ domains outside of the mineralized domains, were assigned background values of 0.0001 ppm for gold and silver, and 0.0001% for copper, zinc, and lead. Unestimated blocks for arsenic and iron were assigned small background values of 10 ppm and 25 ppm, respectively. Background base metal and density values were assigned to all blocks within the modelled “barren” dyke models. Gold and silver were estimated as part of the primary high-grade gold interpolation across and within the dykes as a separate mineralization event.

14.3.15 Block Model Validation

Visual and statistical validation of the copper, zinc, and gold grade estimates for Northwest demonstrated robust model outcomes.

In summary, the validation methods used were:

Swath plots comparing the estimated grade parameters against the mean declustered and raw composite grades
Visual comparison of estimated against composite data values
Comparison of estimated grade and mean composite grade by domain.

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The visual comparison was good, with areas of high (and low) block model grades supported by the presence of high- (and low)-grade composites. All comparisons were satisfactory. Some smoothing of grade was apparent, but this is expected and desirable from the use of OK as the interpolation method.

14.3.16 Classification of Mineral Resources

The Mineral Resource for Northwest is a global estimate representing a reliable estimate of the total contained metal, but the current block estimates are likely to vary as compared with the actual grade/tonnage distribution that will be achieved during selective mining and over short production periods.

The Northwest mineralization has been sufficiently drilled and sampled to allow classification as an Indicated and Inferred Mineral Resource in accordance with the current CIM Definition Standards for Mineral Resources and Mineral Reserves. As with any non-rigidly defined classification there will always be some blocks within categories that depart from defined criteria. The classification employed reflects a practical combination of both geological knowledge and estimation quality parameters that may be more numerical in nature. This approach to classification aims to avoid creating a complex numerically-based mosaic distribution of classified blocks.

Classification considered a number of factors:

Continuity of geological and mineralization: a mineralized volume was based firstly on geological interpretation, subsequently modified by application of a cutoff grade. This ensured that the mineralized volume was not extrapolated unreasonable distances beyond data limits. The models created satisfy these criteria
Data spacing and sample data quality: data spacing must be sufficiently dense to ensure continuity as defined by such studies; the veracity of the data being used for the estimate must be established. The models created satisfy these criteria
Estimation techniques: sufficient statistical and associated studies have been undertaken to ensure that the methodology used for the Northwest estimate is appropriate.

Employing the foregoing criteria, the primary zones were largely classified in the Indicated category, representing the majority of the contained metal at this deposit. The oxide, supergene, and primary high-grade gold zones have all been classified as Inferred to reflect the low confidence in grade and geological continuity as a result of various identified factors, including poor core recovery, assay precision errors, and erratic, localized high-grade gold assays.

14.3.17 Reasonable Prospects of Economic Extraction

Prospects for economic extraction were made by applying an NSR-based cutoff to blocks within a constraining optimized pit shell. The LG optimization was achieved using Whittle.

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The assumed long-term metal prices used for the optimization work as applied to Mineral Resources are shown in Table 14‑33. These metal prices are approximately 15% higher than those used in the estimation of Mineral Reserves, as described in Section 15. .

Table 14‑33: Mineral Resource Commodity Prices

Metal Prices	Price ($)
Copper ($/lb)	3.35
Zinc ($/lb)	1.05
Gold ($/oz)	1,350
Silver ($/oz)	23

The NSR calculation and pit optimization process considered many of the parameters used in Mineral Reserve estimation, as these parameters are well established within the working mine. These parameters include commodity price, 2014 budget costs for production and processing, process recoveries, concentrate grade, selling costs, and other ore-based costs. The optimization process also uses the current geotechnical model for the pit design.

An optimized pit shell results from Whittle using the parameters described above and the commodity prices given in Table 14‑33. No pit design has been completed. Blocks within the pit shell are reported as Mineral Resources above NSR cutoffs of $40.70/t for oxide material, and $39.70/t for supergene and primary material.

Figure 14‑19 shows the pit shell constraining the Resource, with copper-mineralized zones in green, and zinc-mineralized zones in blue.

Mineralization below the constraining pit shell has not been analyzed for possible underground mining at this time. The Mineral Resource statement for Northwest is given in Table 14‑41 in Section 14.5.

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Figure 14‑19: Northwest Resource-Constraining Pit Shell with Copper and Zinc Mineralized Zones, Perspective View Looking Northeast (Local Grid)

14.4 Hambok Estimate

14.4.1 Project Sample Database

The drilling database forms the basis of the estimate, containing all of the available and reliable assay data as of 31 October 2013.

A total of 26,790 m of diamond drilling and 2,680 m of RC drilling was completed in a series of programs completed by Sanu and then BMSC since 2006. The majority of this data are available for the Hambok estimate, as detailed below in Table 10‑4. Some 4,400 samples represent the mineralized zones of the deposit.

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Table 14‑34 Hambok Deposit Drill Hole Summary Table

Year	Drilled by	Holes Drilled	No. of DDH	DDH (m)	No. of RC Holes*	RC (m)
2006	Sanu	HAM-001 to HAM-045	44	7,282.5	-	-
2006	Sanu	HAM-045 to HAM-055 And HAM-058	12	5,105	-	-
2007	Sanu	HAM-056, HAM-060,HAM-061	3	766	-	-
2010	Sanu	HAM-10-062 to HAM-10-065	4	974	-	-
2011	Sanu	HAM-11-066 to HAM-11-071, HAM-RC-11-01 to HAM-RC-11-16	6	1,877	16	978
2012	Sanu	HAM-12-072 to HAM-12-096, HAM-RC-12-17 to HAM-RC-12-42	25	8,072	26	1,698
2013	BMSC	HAM-13-097 to HAM-13-104	08	2,713	-	-
Total			102	26,789.50	42	2,676

Note: *Assay results and lithological data were only used for the oxide component of the Resource.

14.4.2 Geological Model

Introduction

The mineralized components of the Hambok deposit are divided into a primary copper/zinc sulphide zone, representing the majority of the deposit, and a minor oxide gold component.

The primary massive sulphide mineralization is a single body, with a faulted displacement interpreted at depth in the northeast of the deposit. The massive sulphide zones strike at approximately 015°, dipping steeply to the east, with overall strike and dip lengths of some 975 m and 400 m, respectively. Semi-massive and stringer sulphides are not well developed in this deposit, which is hosted in rhyolitic and basaltic units. The oxide unit, as currently understood, is narrow and poorly developed. The mineralized zones are shown in Figure 14‑20.

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Figure 14‑20: Hambok Massive Sulphide Body and Oxide Zone, Perspective View Looking Northeast

Methodology

Initial interpretation was on 25 m to 50 m spaced E-W sections, and was accompanied by re-logging of diamond core to understand the nature of the mineralization and to resolve areas of geological uncertainty and complexity. The sections were digitized and 3D wireframes were created in Surpac. The resultant solid 3D models represent the main geological components controlling the mineralization, and are the basis for establishing grade continuity and the estimation of the Mineral Resource.

The large majority of sulphide mineralization data derives from diamond core. The body is dominated by massive sulphides, frequently with a sharp and clear-cut boundary against the host rock. Developments of semi-massive and stringer sulphides are very limited and discontinuous, and models of these zones were not developed.

The small oxide zone is largely defined by RC drilling confined to a zone in the northern half of the deposit area. Detailed interpretation in this zone is limited due to data quality, and further work is required to better define this zone. Nevertheless, a coherent zone was developed based on records of gossan and associated oxide lithologies. At present, a supergene zone is not thought to be present.

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Displacement of the massive sulphide zone was clearly shown during the 2013 drilling program. Investigation of drill core and records showed that a flat-lying shear or fault was present causing this displacement.

14.4.3 Alteration and Weathering

A weathering surface, based on lithology logs (predominantly from RC drilling) was constructed on 50 m spaced cross-sections to define a simple oxidation domain. More modern drilling data would better define this and other surfaces. This surface was used to divide the oxide and sulphide parts of the deposit.

14.4.4 Mineralization Domains

Within the massive sulphide unit there is clear and persistent higher and lower grade zonation of copper and zinc mineralization. Observation shows a clear correlation of coincident copper and zinc values, and these zones can be readily interpreted on cross-section. Three zones are defined: on the hanging wall, footwall, and a central zone. The zones are best developed in the upper and thicker parts of the deposit, diminishing down-dip, and tapering toward the strike extents. The deepest massive sulphide intersections are often only represented by pyrite. The domains are illustrated in Figure 14‑21.

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Figure 14‑21: Vertical Section of Hambok Massive Sulphides showing Internal Copper-Zinc Zonation; Section at approx. 1705100N

14.4.5 Database Coding

Drilling intervals within mineralized domains (described above in Section 14.1.4) were flagged with a unique database code.

14.4.6 Compositing

Copper, zinc, gold, and silver samples were composited to 1.0 m lengths down hole, extracted from the uniquely coded intercepts stored from the various zone code database tables. The down hole compositing process utilized a best fit approach, which resulted in composites of variable but equal lengths within each contiguous drill hole-coded interval, ensuring that the composite length was as near as possible to the nominated length.

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14.4.7 Bulk Density

A total of 1,255 determinations were completed by Sanu and BMSC using the water immersion method. The density samples were 10 cm to 20 cm segments of full core samples selected from the core trays at approximately 10 m intervals, geology-dependent, during the Sanu programs, and at approximately 3 m intervals for the BMSC program.

Bulk density values were estimated within the massive sulphide using Inverse Distance Squared (IDW²), and model averages checked against raw data for acceptable correlation. A value of 3.2 g/cm³ was assigned to the oxide gold zone.

14.4.8 Recovery

Analysis of diamond core recovery for the massive sulphide shows that there is no correlation between sample grade and core recovery. This is illustrated for zinc in Figure 14‑22; the distribution for copper is similar.

Figure 14‑22: Hambok Massive Sulphide Zone – Diamond Core Recovery vs. Zinc Grade

14.4.9 Exploratory Data Analysis

Sample distribution and population characteristics for the oxide and primary zones were examined for the major domains via histograms, probability plots, and Parrish analysis, to validate the overall domain controls on mineralization, and to determine whether further domaining was required for treatment of obvious statistical outliers. These methods identified multiple populations as well as the need to consider high-grade capping. This is consistent with the creation of separate grade zone domains within the primary mineralization as described above. Further analysis showed that, by the defining of internal zones within the massive sulphide zone, the need for high-grade capping was removed.

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14.4.10 Variography

Variogram models all made use of the 1 m composites derived from the mineralized domains described in Section 14.4.4. The best spatial continuity for each of the sub-zones is generally orientated parallel to the strike and dip of the massive sulphide zone. Differing plunges were noted for each zone, being steeply south, shallowly south, and shallowly to the north for the footwall, central, and hanging wall zones, respectively. The resulting estimation parameters are shown in Table 14‑35.

Table 14‑35: Variogram Parameters for Hambok Estimation

Zone	Element	Nugget	Sill		Range (m)		Azimuth (°)	Plunge (°)	Dip (°)	Ratio
Zone	Element	Nugget	1	2	1	2	Azimuth (°)	Plunge (°)	Dip (°)	Semi- Major	Minor
Central Zone	Cu	0.36	0.2	0.44	114	267	184	-27	52	2.1	17.8
Central Zone	Zn	0.35	0.26	0.39	75	217	194	-9	63	3.0	31.0
Footwall Zone	Cu	0.25	0.24	0.51	164	295	180	-68	90	1.8	29.5
Footwall Zone	Zn	0.10	0.62	0.28	37	218	161	-76	90	3.6	36.3
Hanging Wall Zone	Cu	0.20	0.22	0.58	92	250	161	-52	27	2.0	19.2
Hanging Wall Zone	Zn	0.15	0.30	0.55	71	225	205	18	75	3.4	28.1
Oxide Zone	Au	0.17	0.83	-	38	-	200	24	60	2.4	4.8
Oxide Zone	Au	0.17	0.25	0.58	127	60	185	-21	62	1.9	14.1

14.4.11 Search Neighbourhood Analysis

Search neighbourhood analysis was undertaken, as generally described in Section 14.1.12.

Based on the selected optimal search neighbourhoods, the number of composites required for interpolation at Hambok was determined. Global figures of a minimum of 20 composites and a maximum of 55 composites were applied.

14.4.12 Block Model Parameters

The block model was orientated in the primary strike direction of 017°. The model parameters are given in Table 14‑36. Sub-blocking to 2.5 m was used in all directions. Sub-cells received the parent cell grade during estimation, and were not estimated separately.

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Table 14‑36: Block Model Extents, Massive Sulphide-Hosted Mineralization

Parameter	X	Y	Z
Minimum Coordinate	328520	1704640	0
Block size (m)	10	10	10
Extent (m)	800	1,400	1,000

14.4.13 Grade Estimation

Massive Sulphide-Hosted Mineralization

Four zones or domains were separately estimated within the massive sulphide mineralization, namely hangingwall, central, footwall, and remaining low-grade sulphide mineralization. Copper and zinc values were estimated for all Zones using OK, with ellipsoid searches determined by the analysis described in Section 14.4.10 onwards. Other elements (arsenic, iron, etc.) were estimated using the parameters derived for the copper estimation. Each domain was separately estimated using the unique set of composite samples associated with that domain. The number of informing samples per sampling location was controlled to prevent individual holes or samples having too large an influence on the estimate. A single pass strategy was used for estimation of each domain or sub-section thereof. A second pass with the same parameters was used to ensure all blocks were estimated.

Gold Mineralization

Gold values were estimated for the oxide zone above the oxidation surface using OK, with an ellipsoidal model determined by the variogram analysis and models described in Section 14.4.10 onwards. The zone was separately estimated using the unique set of composite samples associated with that zone. The interpolation pass strategy was as described for the massive sulphide zone. Gold values exist in the sulphide mineralization below the oxidation surface, but are generally low.

14.4.14 Block Model Validation

Validation of the model was by visual inspection of estimated block grades against informing samples, and by comparison of estimated values against informing samples. In addition, QKNA was carried out post-estimation to test the performance of the estimation and the suitability of selected parameters, such as block size, numbers of informing samples, and search ranges. Swath plots were also produced to test for bias between samples and block grades. BMSC concludes that the block model satisfactorily models the distribution and variability of the informing sample grades without undue bias or smoothing for the level of study.

Figure 14‑23 shows the distribution of the kriging slope of regression (Sr) parameter for the Hambok estimate for copper. The inset in the figure shows peripheral blocks, on the edges of the zones where values are lower, that mask the quality of the estimate in the core of the deposit. The main figure shows the high proportion of the model having slope of regression efficiencies in excess of 60%, demonstrating a robust estimate.

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Figure 14‑23: Hambok Massive Sulphide: Slope of Regression Distribution, Value >0.5 (inset, all blocks)

Main picture, blocks with Sr ≥0.5 within massive sulphides

14.4.15 Classification of Mineral Resources

The Mineral Resource for Hambok is a global estimate representing a reliable estimate of the total contained metal, but the current block estimates are likely to vary as compared with the actual grade/tonnage distribution that will be achieved during selective mining and over short production periods.

The Hambok mineralization has been sufficiently drilled and sampled to allow classification as an Indicated and Inferred Mineral Resource in accordance with the current CIM Definition Standards for Mineral Resources and Mineral Reserves. As with any non-rigidly defined classification there will always be some blocks within categories that depart from defined criteria. The classification employed reflects a practical combination of both geological knowledge and estimation quality parameters that may be more numerical in nature. This approach to classification aims to avoid creating a complex numerically-based mosaic distribution of classified blocks.

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Classification considered a number of factors:

Continuity of geology and mineralization: a mineralized volume is firstly based on geological interpretation, subsequently modified by application of a cutoff grade from which internal domaining was completed. This ensures that the mineralized volume is not extrapolated unreasonable distances beyond data limits. The model created satisfies these criteria
Data spacing and sample data quality: data spacing must be sufficiently dense to ensure continuity as defined by such studies; the veracity of the data being used for the estimate must be established. The models created satisfy these criteria; and
Estimation techniques: sufficient statistical and associated studies have been undertaken to ensure that the methodology used for the Hambok estimate is appropriate.

Employing the foregoing criteria, the primary zones were largely classified in the Indicated category. The factors described above were combined into simplified shapes by resource category that was then applied to the block model, avoiding a patchy distribution of blocks of differing category. For the primary zone where data diminishes down-dip, estimation parameters such as slope of regression and kriging efficiency guided the limit of material in the Indicated category. The oxide gold zone was largely classified as Inferred to reflect the lower confidence in grade and geological continuity in this zone.

14.4.16 Reasonable Prospects of Economic Extraction

Prospects for economic extraction were made by applying an NSR-based cutoff to blocks within a constraining optimized pit shell. The LG optimization was achieved using Whittle.

The assumed long-term metal prices used for the optimization work as applied to Mineral Resources are shown in Table 14‑37. These metal prices are approximately 15% higher than those used in the estimation of Mineral Reserves, as described in Section 15.

Table 14‑37: Mineral Resource Commodity Prices

Metal Prices	Price ($)
Copper ($/lb)	3.35
Zinc ($/lb)	1.05
Gold ($/oz)	1,350
Silver ($/oz)	23

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An optimized pit shell results from Whittle using the parameters described above and the commodity prices given in Table 14‑13. No pit design has been completed. Blocks within the pit shell are reported as Mineral Resources above NSR cutoffs of $44.45 for oxide material, and $43.45 for supergene and primary material.

Mineralization below the pit shell has not been analyzed for possible underground mining at this time. The constraining pit shell and mineralization are shown in Figure 14‑24.

The Mineral Resource statement for Hambok is given in Table 14‑42 in Section 14.5.

Figure 14‑24: Perspective View of Hambok Mineralization with Constraining Pit Shell, Looking North

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14.5 Mineral Resource Statement

Mineral resources are classified and reported in accordance with the current CIM Definition Standards for Mineral Resources and Mineral Reserves, and as described in Section 14.1.16, Section 14.3.16, and Section 14.4.15. Mineral Resources reported here for Bisha Main and Harena are inclusive of Mineral Reserves. BMSC cautions that Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

All of the Mineral Resource estimates described here have an effective date of 31 December 2013. The Mineral Resource estimates are summarized in Table 14‑38 to Table 14‑42.

Table 14‑38: Combined Mineral Resource Estimate – Bisha, Harena, Northwest and Hambok Deposits (Effective Date 31 December 2013)

Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Contained Metal
Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000 lb)	Zn ('000 lb)	Au ('000 oz)	Ag ('000 oz)
*Indicated*
Oxide Phase	Variable	480	-	-	6.6	20	-	-	100	300
Supergene Phase	Variable	8,480	3.41	-	0.6	25	638,650	-	160	6,920
Primary Phase	Variable	32,260	1.05	4.59	0.6	36	743,920	3,223,350	570	36,430
Total		41,220					1,382,570	3,223,350	830	43,650
*Inferred*
Oxide Phase	Variable	570	-	-	3.9	19	-	-	61	350
Supergene Phase	Variable	110	1.37	-	3.4	18	4,200	-	10	70
Primary Phase	Variable	1,752	0.80	4.19	0.7	33	31,230	163,210	40	1,910
Total	-	2,432					35,430	163,210	111	2,330

Table 14‑39: Bisha Resource Estimate (Effective Date 31 December 2013)

Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Contained Metal
Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000 lb)	Zn ('000 lb)	Au ('000 oz)	Ag ('000 oz)
*Indicated*
Oxide Phase	40.55	410	-	-	6.8	21	-	-	90	270
Supergene Phase	39.55	7,460	3.68	-	0.6	27	605,500	-	150	6,590
Primary Phase	39.55	21,070	1.05	5.87	0.7	47	487,770	2,726,870	480	31,770
Total	-	28,940					1,093,270	2,726,870	720	38,630
*Inferred*
Oxide Phase	40.55	30	-	-	7.3	39	-	-	10	30
Supergene Phase	39.55	10	7.23	-	0.1	10	2,200	-	0	0
Primary Phase	39.55	1,300	0.80	4.50	0.5	36	23,100	129,600	20	1,500
Total	-	1,340					25,300	129,600	30	1,530

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Table 14‑40: Harena Mineral Resource Estimate (Effective Date 31 December 2013)

Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Contained Metal
Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000 lb)	Zn ('000 lb)	Au ('000 oz)	Ag ('000 oz)
*Indicated*
Oxide Phase	42.41	70	-	-	5.5	14	-	-	10	30
Primary Phase	41.41	1,800	0.65	3.91	0.6	23	25,760	154,990	30	1,350
Total	-	1,870					25,760	154,990	40	1,380
*Inferred*
Oxide Phase	40.55	20	-	-	5.9	8	-	-	0	10
Primary Phase	39.55	350	0.75	4.10	0.8	32	5,700	31,200	10	350
Total	-	370					5,700	31,200	10	360

Table 14‑41: Northwest Mineral Resource Estimate (Effective Date 31 December 2013)

Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Contained Metal
Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000 lb)	Zn ('000 lb)	Au ('000 oz)	Ag ('000 oz)
*Indicated*
Oxide Phase	40.70
Supergene Phase	39.70	1,020	1.47	-	0.2	10	33,150	-	10	330
Primary Phase	39.70	2,530	1.04	1.08	0.3	13	58,020	60,250	20	1,050
Total	-	3,550					91,170	60,250	30	1,380
*Inferred*
Oxide Phase	40.70	500	-	-	3.7	18	-	-	50	300
Supergene Phase	39.70	100	0.80	-	3.7	19	2,000	-	10	70
Primary Phase	39.70	100	0.90	0.90	2.9	15	2,400	2,400	10	60
Total	-	700					4,400	2,400	70	430

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Table 14‑42: Hambok Mineral Resource Estimate (Effective Date 31 December 2013)

Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Contained Metal
Zone	NSR Cutoff ($/t)	Tonnes ('000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000 lb)	Zn ('000 lb)	Au ('000 oz)	Ag ('000 oz)
*Indicated*
Oxide Phase	44.45	-	-	-	-	-	-	-	-	-
Primary Phase	43.45	6,860	1.14	1.86	0.2	10	172,370	281,240	40	2,260
Total	-	6,860					172,370	281,240	40	2,260
*Inferred*
Oxide Phase	44.45	20	-	-	1.5	17	-	-	1	10
Primary Phase	43.45	2	0.90	0.20	0.2	8	30	10	0	0
Total	-	22					30	10	1	10

Notes:

(1) NSR Cutoff ($/t): variable as per tables above. Mineral Resources are defined within an optimal LG pit shell, generated using metal prices for copper, zinc, gold, and silver of $3.35/lb, $1.05/lb, $1,350/oz, and $23/oz, respectively, using blocks of all resource categories. The mining cost and total ore-based cost (process, G&A, and stockpile rehandle) applied was the 2014 budget mining cost with appropriate ore haulage costs for each satellite deposit. Overall pit slopes varied from 34.5° to 44° for Bisha, 29° to 35.5° for Harena, from 39° to 45° for Northwest, and 40° overall for Hambok (preliminary assessment)
(2) NSR values were calculated for each block using both Indicated and Inferred categories, metal prices, recoveries, appropriate smelter terms, and downstream costs. Metallurgical recoveries, supported by metallurgical testwork, were applied as follows:
a) Bisha oxide zone: recoveries of 88% and 22% were applied for gold and silver, respectively.
b) Harena oxide zone: a recovery of 75% was applied for gold.
c) Bisha Supergene zone; recoveries of 88%, 46%, and 50% were applied for copper, gold, and silver, respectively.
d) Bisha Hanging Wall zone; recoveries of 85%, 46%, and 50% were applied for copper, gold, and silver, respectively.
e) Bisha Transition Zone (mixed zinc and secondary copper zone below the supergene); recoveries as per supergene zone were applied.
f) Bisha Primary Zone; recoveries to copper concentrate of 85%, 36%, and 29% were applied for copper, gold, and silver, respectively. Recoveries to zinc concentrate of 83.5%, 9%, and 20% were applied for zinc, gold, and silver, respectively.
g) Harena Primary Zone; recoveries to copper concentrate of 85%, 36%, and 29% were applied for copper, gold, and silver, respectively. A zinc recovery to zinc concentrate of 72% was applied.
h) Northwest Oxide Zone; recoveries of 88% and 22% were applied to gold and silver, respectively.
i) Northwest Supergene Zone; recoveries of 87%, 46%, and 50% were applied for copper, gold, and silver, respectively. Zinc was not assigned a recovery, as the values are isolated on the fringes of the deposit.
j) Northwest Primary Zone; recoveries to copper concentrate of 87%, 36%, and 29% were applied for copper, gold, and silver, respectively. Recoveries to zinc concentrate of 81%, 9%, and 20% were applied for zinc, gold, and silver, respectively.
k) Hambok Oxide Zone; recoveries of 88% and 22% were applied to gold and silver, respectively.
l) Hambok; recoveries to copper concentrate of 88%, 87%, 36%, and 29% were applied for copper, zinc, gold, and silver, respectively. Preliminary metallurgical characterization studies, but not full testing, have been completed for Hambok.
(3) Mineral Resources are reported within the pit shell generated using the specified commodity prices, using NSR block grade cutoffs derived as above. Tonnages were rounded to the nearest 10,000 tonnes, and grades were rounded to two decimal places for copper and zinc, one decimal place for gold, and no decimal places for silver. Tonnages and grades for the Inferred category are further rounded, reflecting the uncertainty that attaches to this category.
(4) Rounding as required by reporting guidelines may result in apparent summation differences between tonnes, grades, and contained metal contents.

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(5) Tonnage and grade measurements are in metric units. Contained gold and silver ounces are reported as troy ounces, contained copper and zinc pounds as imperial pounds.
(6) Stockpile tonnages are included as Indicated Resources in the totals given in the tables for Bisha and Harena.
(7) The Bisha Primary Inferred Resource includes an underground resource. This was derived by defining a shape around contiguous blocks where an NSR of $100 was achieved. The value of NSR $100 represents the processing cost plus approximately $60/t mining cost.
(8) Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

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15 Mineral Reserve Estimates

15.1 Key Assumptions/Basis of Estimate

Mineral Reserves for Bisha and Harena are supported by a LOM plan, which was developed using the following key parameters.

15.1.1 Pit Slopes

BGC reviewed pit slope performance during a site visit in 2013 and made no changes to the pit slope design criteria they developed during 2012. These geotechnical design criteria were incorporated into updated Bisha and Harena pit slope designs. Overall pit slopes varied from 34.5° to 44° for Bisha, and from 29° to 35.5° for Harena. A detailed discussion of pit slope design parameters is provided in Section 16.1.

15.1.2 Net Smelter Return Calculations

Revenue will be generated from the sale of copper concentrates (which contain payable co-products of gold and silver) during the Supergene Phase, and both copper and zinc concentrates during the Primary Phase of the operation. Small quantities of oxides remaining within the pits and stockpiles will be processed at the end of the mine life through the currently inactive oxide circuit, producing doré with payable gold and silver. To capture the multi-rock type, variable recoveries by rocktype, and multi-element complexity, NSR values were calculated for block valuation.

The NSR grade determination considers the recoveries, concentrate grades, and penalties (where applicable) for each rock type, and applies the price and cost parameters shown in Table 15‑1 through Table 15‑6, resulting in a net value per tonne of ore, inclusive of all costs outside the mine gate. Only Measured and Indicated Mineral Resources were considered for processing. Inferred Mineral Resources were treated as waste.

As of 31 December 2013, the copper and gold prices used for Mineral Reserves ($2.90/lb and $1,175/oz, respectively) were below both the three-year trailing prices ($3.65/lb and $1,551/oz, respectively) and 31 December 2013 spot prices ($3.35/lb and $1,205/oz, respectively). The silver price used ($20/oz) is well below the three-year trailing average of $30/oz, and equal to the 31 December 2013 spot price of $20/oz. The zinc price used ($0.92/lb) is just below both the three-year trailing price of $0.93/lb, the 31 December 2013 spot price of $0.94/lb.

Applicable gross royalties are 3.5% for base metals and 5% for precious metals.

The Bisha oxide gold and silver recoveries are based on actual production at the Bisha mill. The Harena gold recovery is based on metallurgical testwork. Silver was not analyzed in the Harena oxide testwork.

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Table 15‑1: Reserves Metal Prices

Commodity	Unit	Cost
Cu	$/lb	2.90
Au	$/oz	1,175
Ag	$/oz	20.00
Zn	$/lb	0.92

Table 15‑2: Doré Parameters

Description	Unit	Bisha	Harena
*Recoveries to Doré*
Au	(%)	88.00	75.00
Ag	(%)	22.00	0.00
Metal Content in Doré	(%)	85.00	85.00
*Deductions and Payable Metal*
Au Payable	(%)	99.90	99.90
Ag Payable	(%)	99.50	99.50
*Refining Charge*
Refining, Treatment, Transport, and Insurance Charges	$/oz Au	5.12	5.12

Table 15‑3: Copper Concentrate Recoveries

Recoveries to Cu Concentrate	Bisha Supergene	Bisha Hanging Wall	Bisha Transition	Bisha Primary	Harena Primary
Cu (%)	88.00	85.00	88.00	85.00	85.00
Au (%)	46.25	46.25	46.25	36.00	36.00
Ag (%)	50.00	50.00	50.00	29.00	29.00
Cu Concentrate Grade (%)	30.00	22.00	30.00	25.50	25.50

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Table 15‑4: Copper Concentrate Shipping and Smelting Terms

	Unit	Value
*Payable Metal and Deductions*
Cu Payable	(%)	96.50
Cu Deduction	(Unit)	1.00
Au Pay Factor	(%)	94
Au Deduction	(g/dmt)	0
Ag Pay Factor	(%)	94
Ag Deduction	(g/dmt)	0
*Treatment, Shipping, and Refining*
Treatment, Shipping, and Refining	($/dmt)	289.00
Moisture	(%)	8.00
Transit Losses	(%)	0.125
Insurance Cost	(%)	0.20
*Refining Charges*
Cu	($/lb)	0.085
Au	$/oz	4.80
Ag	$/oz	0.40

The copper treatment, shipping, and refining charge shown in Table 15‑5 is the total of treatment, land freight, port, ocean freight, umpire, and marketing charges on a dry tonne basis.

The Bisha Supergene copper concentrate testwork identified the presence of enargite/tennantite. An arsenic recovery of 67.5% was estimated from the assumed ratio of enargite/tennantite to arsenopyrite. Arsenic in concentrate content was estimated at the block level, and was subject to a smelter penalty of $5/t per 0.1 increments above 0.2% in the NSR estimate for pit optimization and ore/waste delineation during mine planning.

Table 15‑5: Zinc Concentrate Recoveries

	Unit	Bisha Primary	Harena Primary
*Recoveries to Zn Concentrate*
Cu	(%)
Zn	(%)	83.50	72.00
Au	(%)	-	-
Ag	(%)	-	-
Zn Concentrate Grade	(%)	55.60	52.00

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Table 15‑6: Zinc Concentrate Shipping and Smelting Terms

	Unit	Value
*Payable Metal and Deductions*
Zn Payable	(%)	85.00
Zn Deduction	(Unit)	8.00
*Treatment and Shipping*
Treatment Shipping and Refining	($/t)	408.00
Moisture	(%)	8.00
Transit Losses	(%)	0.125
Insurance Cost	(%)	0.20

The above zinc treatment, shipping, and refining charge is the total of treatment, land freight, port, ocean freight, umpire, and marketing charges on a dry tonne basis.

15.1.3 Operating Costs

The waste and ore-based costs applied for pit optimization and mine planning were based on 2014 budget costs developed by BMSC. The mining cost (inclusive of ore control, geology, lab services, and a mining equipment-sustaining capital allowance) was $2.89/t, plus an appropriate incremental haulage cost per bench. The total ore-based costs (process, G&A, stockpile re-handle, and a TMF-sustaining capital allowance) are $40.55/t for oxide, and $39.55/t for supergene and primary ores. Harena ore-based costs include an additional $2.63/t overland ore haulage cost.

Because the ore-waste delineation was performed using an NSR block value, net of downstream costs, the total ore-based cost represents the marginal break even cut-off grade for pit optimization and mine planning purposes.

15.1.4 Pit Optimization and Pit Phase Design

LG pit optimization was performed to determine the economic limits of the deposits. The NSR grade item (refer to Section 15.1.2) was used to assign block values to Indicated blocks. Inferred material was treated as waste. The pit slope recommendations were applied for each spatial region. The waste and ore-based costs were applied as stated above. The December 2013 end-of-month as-built triangulated surface was used as the starting surface. The ultimate pit limit shells were based on the full NSR values calculated from the prices and parameters shown above. These shells were used to guide the design of the ultimate pits, which incorporate minimum mining widths and practical ramp access. Internal pit phase designs were guided by nested shells generated at revenue factors less than 1.0.

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15.2 Dilution Adjustments

The Mineral Resource estimates for Bisha and Harena are considered to be undiluted. A 2 m dilution skin was added at the time of ore and waste delineation for mine planning purposes. Skin blocks that did not have interpolated grades were given a low background values representative of non-mineralized material. The resulting dilution effect is 11% on tonnes for Supergene and 10% on Primary ore. No mining loss adjustments were made.

15.3 Conversion Factors from Mineral Resources to Mineral Reserves

Mineral Reserves have been modified from Mineral Resources by taking into account geological, mining, processing, and economic parameters and permitting requirements, and are therefore classified in accordance with the 2010 CIM Definition Standards for Mineral Resources and Mineral Reserves.

15.4 Mineral Reserves Statement

The QP for the Mineral Reserve estimate is Jay Melnyk, P.Eng., a principal of AGP Mining Consultants, Inc.

Mineral Reserves are reported at commodity prices for copper, zinc, gold, and silver of $2.90/lb, $0.92/lb, $1,175/oz, and $20/oz, respectively, and have an effective date of 31 December 2013.

Mineral Reserves for Bisha and Harena are summarized in Table 15‑7. The Mineral Resources stated in Section 14 are inclusive of the Mineral Reserves.

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Table 15‑7: Bisha and Harena Reserves Estimate (Effective Date: 31 December 2013)

Zone	Tonnes (‘000s)					Contained Metal
Zone	Tonnes (‘000s)	Cu %	Zn %	Au g/t	Ag g/t	Cu ('000s lb)	Zn ('000s lb)	Au ('000s oz)	Ag ('000s oz)
*Bisha Probable Mineral Reserve Estimate*
Oxide Phase	430	-	-	6.50	20	-	-	90	280
Supergene Phase	7,400	3.57	-	0.61	27	582,420	-	145	6,420
Primary Phase	18,390	1.02	5.66	0.68	46	413,540	2,294,730	402	27,200
Total	26,220					995,960	2,294,730	637	33,900
*Harena Probable Mineral Reserve Estimate*
Oxide Phase	80	-	-	4.93	16	-	-	13	-
Primary Phase	1,160	0.64	3.57	0.52	22	16,370	91,300	19	820
Total	1,240					16,370	91,300	32	820
*Combined Bisha and Harena Probable Mineral Reserve Estimate*
Oxide Phase	510	-	-	6.25	19	-	-	103	280
Supergene Phase	7,400	3.57	-	0.61	27	582,420	-	145	6,420
Primary Phase	19,550	1.00	5.54	0.67	45	429,910	2,386,030	421	28,020
Total	27,460					1,012,330	2,386,030	669	34,720

Notes:

(1) NSR Cutoff ($/t): Oxide Phase $40.55 for Bisha and $42.41 for Harena: Supergene Phase $39.55 for Bisha and Primary Phase $39.55 for Bisha and $42.41 for Harena. Mineral Reserves are defined within a mine plan, with phase designs guided by LG pit shells, generated using metal prices for copper, zinc, gold and silver of $2.90/lb, $0.92/lb, $1,175/oz, and $20/oz, respectively. The mining cost applied was the 2014 budget mining cost with appropriate haulage cost adjustments. The total ore based costs (process, G&A, and stockpile re-handle) are $40.55/t for oxide, and $39.55/t for supergene and primary ores. Harena ore-based costs include an additional $2.63/t overland ore haulage cost. Overall pit slopes varied from 34.5° to 44° for Bisha and from 29° to 35.5° for Harena.
(2) NSR values were calculated using diluted Indicated grades, metal prices, recoveries and appropriate smelter terms, and downstream costs. Metallurgical recoveries, supported by metallurgical testwork, were applied as follows:

a) Bisha Oxide Zone: recoveries of 88% and 22% were applied for gold and silver, respectively.

b) Harena Oxide Zone: a recovery of 75% was applied for gold; 80 kt of oxide remain in the Harena pit.

c) Bisha Supergene Zone; recoveries of 88%, 46%, and 50% were applied for copper, gold, and silver respectively.

d) Bisha Hanging Wall Zone; recoveries of 85%, 46% and 50% were applied for copper, gold and silver, respectively.

e) Bisha Transition Zone (mixed zinc and secondary copper zone below the supergene): the same metallurgical parameters as for the Bisha Supergene Zone were applied.

f) Bisha Primary Zone; recoveries to copper concentrate of 85%, 36%, and 29% were applied for copper, gold, and silver, respectively. A recovery to zinc concentrate of 83.5% was applied for zinc. Gold and silver reporting to the zinc concentrate are not expected to be payable.

g) Harena Primary Zone; recoveries to copper concentrate of 85%, 36% and 29% were applied for copper, gold and silver respectively. A zinc recovery to zinc concentrate of 72% was applied. Gold and Silver reporting to zinc concentrate are not expected to be payable.

(3) Mineral Reserves are reported within Bisha and Harena ultimate pit designs, using NSR block grade, where the marginal cutoff is the total ore-based cost stated above. Tonnages are rounded to the nearest 10kt and grades are rounded to two decimal places with the exception of silver, which was rounded to zero decimal places.

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(4) Rounding as required by reporting guidelines may result in apparent summation differences between tonnes, grade, and contained metal content.
(5) Tonnage and grade measurements are in metric units. Contained gold and silver ounces are reported as troy ounces, contained copper and zinc pounds as imperial pounds.
(6) The LOM stripping ratios in tonnes for Bisha and Harena are 5.6: 1 and 5:7: 1, respectively.
(7) The Bisha Probable Oxide Mineral Reserve includes the pyrite sand and uncrushed DSO stockpiled material, being 320 kt at 7.8 g/t Au in the stockpile as of 31 December 2013.
(8) The Bisha Probable Supergene Mineral Reserves includes 105 kt at 5.25% Cu in stockpile as of 31 December 2013.
(9) For Bisha and Harena, conversions from Mineral Resources to Mineral Reserves included a mining dilution of 2 m applied around contiguous blocks that exceeded the appropriate NSR cutoff value.

15.5 Factors that May Affect the Mineral Reserve Estimates

Factors that may affect the Mineral Reserve estimates include dilution; metal prices; smelter, refining, and shipping terms; metallurgical recoveries and geotechnical characteristics of the rock mass; capital and operating cost estimates; and effectiveness of surface and groundwater management.

The QPs are of the opinion that these potential modifying factors have been adequately accounted for using the assumptions in this report, and therefore the Mineral Resources within the mine plan may be converted to Mineral Reserves. Factors that may affect the assumptions in this report are:

Commodity price and exchange rate assumptions.
Ensuring marketability of concentrates in particular the copper concentrate during the Supergene Phase will require careful ore control and blending to minimize smelter penalties.
Mill throughput of the identified ore types may prove to be higher or lower than modelled. If certain rock types or delivered blends of rock types have lower throughputs than currently modelled, this would increase the processing cost, which would in turn increase the mill cutoff grade. If all other things held constant, this would tend to reduce the tonnage of the Mineral Reserve and the amount of contained metal. If throughput reductions are significant, this could reduce the size of the economic pit limits, further reducing the Mineral Reserve. Furthermore, a reduction in throughput would delay cash flow, resulting in a negative impact on Project economics.
Effective surface and groundwater management is important to the safety and productivity of the mining operation. If the currently planned water management methods prove to be ineffective, additional dewatering wells and/or sumps and pump systems may be required, which would add to the capital and operating costs, resulting in a negative impact on Project economics and a potential reduction in the Mineral Reserves.

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16 Mining Methods

16.1 Geotechnical

16.1.1 Overview

BGC Engineering Inc. (BGC) has undertaken Geotechnical assessments of the open pit slopes for the Main Zone and Harena Zone of the Bisha Mine. Slope design recommendations for the bench, inter-ramp, and overall slope scales have been provided for each zone, based on historical geotechnical drilling, recent (2011–2012) geotechnical drilling, and experience with the current pit walls of the Main Zone. AGP has used these open pit slope designs (Table 16‑1 and Table 16‑2) in the current Mineral Reserve estimate.

16.1.2 Main Zone

The Bisha Pit is planned to be approximately 1.5 km long and 1 km wide. The slope heights will range from 160 m to 290 m. BGC completed rock mass characterization, structural geology assessments, and slope stability assessments to develop the open pit slope designs for the Main Zone (BGC, 2012).

Most of the geotechnical data collected and used in the current work from 2004 to 2012 was collected by others (AMEC, Turner Mining and Geotechnical Pty Ltd, and site technical staff); these data were compiled by BGC. Geotechnical drilling results, oriented core data, discontinuity mapping, and geological models were provided by BMSC. Data from 33 geotechnical drill holes were used by BGC. Photogrammetric mapping and documentation of existing slopes in the Main Zone open pit was completed by BGC to provide additional structural geology data and to assess the performance of the mined slopes. Previously-completed geomechanical laboratory testing was supplemented by additional rock core samples selected by BGC. The laboratory-testing database consists of small-scale direct shear (10), uniaxial compressive strength (80), and Brazilian tensile strength (73) results.

The rock mass of the Main Zone has been divided into three geotechnical units, based on the intensity of weathering, properties of the intact rock, and geological units encountered. The geotechnical units used for the current study are “SRK,” or saprock, “WRK,” or weathered rock, and “FRK,” or fresh rock. Geotechnical core logging completed by BMSC and reviewed by BGC has been used to characterize the rock mass of the Main Zone.

The SRK geotechnical unit includes highly to completely weathered rock, including the saprolite, oxide ore, and strongly acidified or “soap” material. The rocks of the SRK unit are typically extremely weak to weak. The rock mass rating (RMR, 1976) for SRK is “very poor.” The WRK geotechnical unit includes slightly- to moderately-weathered rock, including the weathered and/or altered supergene ore. The rocks of the WRK unit are typically weak to strong. The rock mass rating for the WRK unit is “fair.” The FRK geotechnical unit includes all rocks below the WRK unit, including the sulphide ore. The rocks of the FRK unit are typically medium strong to strong. The rock mass rating for the FRK unit is “good.”

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The Main Zone has been divided into three structural domains; the domain boundaries are interpreted to be fold axes. The locations of the fold axes were interpreted by BGC based on a combination of available data for the orientation of the foliation from oriented core data provided by BMSC, as well as regional geological mapping. The structural domain boundaries include a syncline to the west side of the Main Zone dividing Domain I from Domain II, and an anticline to the east side of the Main Zone dividing Domain II from Domain III. Domain I and Domain III are characterized by steeply east-dipping foliation, and Domain II is characterized by steeply west-dipping foliation. Besides these project-scale fold axes, there are local variations in foliation orientation due to tight folding.

Open pit slope design recommendations have been provided for each design sector in each geotechnical domain (Table 16‑1). Design sectors are defined by the average azimuth of the anticipated wall orientations, based on geological structural controls on slope stability. Geotechnical domains result from the combination of structural domains and geotechnical units, resulting in nine distinct geotechnical domains (Figure 16‑1). The recommended inter-ramp angles vary from 31° to 46°, depending on the design sector and geotechnical domain. All of the slope designs assume that controlled blasting will be undertaken for the final walls of the pit.

Depressurization of the open pit slopes is required to achieve the open pit slope designs. The pre-development water table is approximately 15 m below ground surface (Knight Piésold, 2012). Some examples of seepage from the current slopes in the Main Zone have been identified. A combination of vertical wells and horizontal drains has been proposed to dewater the open pit and depressurize the slopes. Further evaluations of the pit dewatering system are required as part of future mine design studies.

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Figure 16‑1: Main Zone Domains and Geotechnical Units

Table 16‑1: Bisha Main Zone Slope Designs

Geotechnical Domain	Catch Bench Geometry			Inter-Ramp Geometry
Geotechnical Domain	Design Height Bh (m)	Face Angle Ba (°)	Width (varies with slope azimuth) Bw (m)	Maximum Height IRh (m)	Angle (varies with slope azimuth) IRa (°)
SRK-I	10	63	6.0 to 11.6	-	31 to 42
SRK-II	10	63	6.0 to 7.9	-	38 to 42
SRK-III	10	63	6.0 to 7.9	-	38 to 42
WRK-I	10	67	6.8 to 9.5	100	36 to 42
WRK-II	10	67	6.0 to 8.5	100	38 to 44
WRK-III	10	67	6.8 to 8.5	100	38 to 42
FRK-I	10	70	6.0 to 9.9	140	37 to 46
FRK-II	10	70	6.0 to 8.9	140	39 to 46
FRK-III	10	70	6.0 to 8.9	140	39 to 46

Note:

1. Summarized from Table 10, Bisha Mine - Main Zone Prefeasibility Level Open Pit Slope Designs, (BGC, 2012b)
2. Any walls developed in the SRK unit that are greater than 30 m high required detailed geotechnical review; slope design modifications may be required.
3. For rock mass control, SRK is assumed to be dry, WRK is assumed to be dry in the NE pit wall and 25% saturated elsewhere, and FRK is assumed to be 100% saturated.

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16.1.3 Harena Zone

The Harena Zone open pit is planned to be approximately 500 m long and 500 m wide. The slope heights of the ultimate open pit range from 115 m to 150 m. The development of the Harena deposit consists of two pit phases: mining of oxide ore, which is now complete, and mining of primary sulphide ores. BGC completed rock mass characterization, structural geology assessments, and slope stability assessments to develop open pit slope designs for the Harena Zone (BGC, 2012a).

The geotechnical database relied on for the current work includes geotechnical drilling, oriented core data, and laboratory testing results. Four geotechnical drill holes were completed by BMSC for the current work. A laboratory program of small-scale direct shear (31), uniaxial compressive strength (21), and Brazilian tensile strength (22) testing was completed. Data from previously completed exploration drill holes were also used in the current study. The data for the Harena Zone has been supplemented by observations of the existing pit slopes of the Main Zone.

The weathering profile is well developed in the foliated and folded felsic and mafic volcanic rocks of the zone. The rock mass near surface is completely- to moderately-weathered; fresh or unweathered rock is found approximately 50 m below the current ground surface. The rock mass of the Harena Zone was divided into four geotechnical units: “poor” quality weathered rock; oxide ore; “fair” to “good” quality fresh rock; and primary sulphide ore.

A structural geology model for the Harena Zone was developed from oriented core and lineament mapping data. The zone appears to be located in a single structural domain (Figure 16‑2). The geological structure of the zone is dominated by steeply northwest-dipping foliation. Major geological structures include foliation and three sets of faults inferred from the available lineament mapping data.

Open pit slope design recommendations have been provided for each design sector in each geotechnical domain (Table 16‑2). Design sectors are defined by the average azimuth of the anticipated wall orientations, based on geological structural controls on slope stability. The recommended inter-ramp angles vary from 34° to 43°, depending on the design sector of the pit. All of the slope designs assume that controlled blasting will be undertaken for the final walls of the pit.

Depressurization of the open pit slopes will be required to achieve the open pit slope designs. The pre-development water table is assumed to be similar in the Harena Zone as in the Main Zone. The design basis assumed a water table approximately 15 m below ground surface. Vertical wells may be adequate to dewater the open pit and depressurize the slopes. Further evaluations of the pit dewatering system are required as part of future mine design studies.

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Figure 16‑2: Harena Domains and Geotechnical Units

Table 16‑2: Harena Slope Designs – Primary Sulphide Mine Phase

Geotechnical Domain	Catch Bench Geometry			Inter-Ramp Geometry
Geotechnical Domain	Design Height Bh (m)	Face Angle (varies with slope azimuth) Ba (°)	Width (varies with slope azimuth) Bw (m)	Maximum Height IRh (m)	Angle (varies with slope azimuth) IRa (°)
WRK (Weathered Rock)	12	63	9.0 to 12.3	48 (Az 133)	33 to 38
FRK (Fresh Rock)	12	63 to 70	7.4 to 10.8	84 (Az 313) and 96 (Az 133)	35 to 46

Notes:

1. Summarized from Table 7, “Open Pit Slope Designs - Primary Sulphide Mine Phase,” (BGC, 2012a)
2. Interim slopes developed in the OXO (Oxide Ore) geotechnical unit should use the designs for WRK; interim slopes developed in the PSO (Sulphide Ore) geotechnical unit should use the designs for FRK.

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16.2 Pit Design

The Bisha and Harena deposits are being mined by conventional open pit mining methods. The Bisha pit consists of nine individual pit phases, where the first three phases targeted oxide production, the second three target supergene production, and the final three phases will target primary production. The oxide pit phases have now been exhausted, and Phases 4, 5, and 6 are currently providing supergene ore. Stripping for the primary mineralization has started in Phase 8.

The Harena pit features two pit phases, one targeting oxide production (which was completed in 2013), and the final phase targeting primary production. The Harena pit is currently inactive.

The pit designs are guided LG-optimized pit shells generated using the Whittle software package. The Bisha and Harena ultimate pit designs, along with the internal phases not yet being mined, were guided by LG shells generated using the optimization parameters discussed in Section 15.1 and using the pit slope guidance discussed in Section 16.1, with flattening added as required for ramp access.

Pit design parameters include: double lane ramp design width of 23 m, based on three times the width of the Cat 775 truck; ramp gradient of 10%; a nominal minimum mining width of 25 m; and smoothing of walls in areas where convex “noses” could cause geotechnical issues.

The ultimate pit designs for Bisha and Harena are shown in Figure 16‑3 and Figure 16‑4, respectively. Pit phase design volumetrics are shown in Table 16‑3.

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Figure 16‑3: Bisha Ultimate Pit Design

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Figure 16‑4: Harena Ultimate Pit Design

16.3 Cutoff Grades

To capture the multi-rock types and multi-element complexity, net smelter return (NSR) values were calculated for block valuation. The NSR is a net value per tonne of ore, inclusive of all recoveries and costs outside the mine gate, as listed in Section 15.1. Because the mineralization/waste delineation was performed using the NSR block value, the total ore-based cost represents the marginal break even cutoff grade for pit optimization and mine planning purposes. The total ore-based costs (process, G&A, and stockpile re-handle) are $39.55/t for supergene and primary ores. The remaining oxide to be processed at the end of the mine life has a cutoff of $40.55/t. Harena ore-based costs include an additional $2.63/t overland ore haulage cost.

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Table 16‑3: Pit Phase Volumetrics

Pit Phases	Ore Oxide (kt)	Au g/t	Ag g/t	Supergene Ore (kt)	Au g/t	Ag g/t	Cu %	As ppm	Primary Ore (kt)	Au g/t	Ag g/t	Cu %	Zn %	As ppm	Total Waste (kt)	Total Material (kt)	Strip Ratio (W/O)
Bisha NE Layback	43	3.82	86	40	0.86	12	1.38	582	-	-	-	-	-	-	2,823	2,906	34.1
Bisha Phase 4	5	1.21	46	1,803	0.68	27	4.25	1,035	63	0.79	24	0.61	1.80	965	2,704	4,575	1.4
Bisha Phase 5	51	2.12	84	2,141	0.74	42	4.53	1,346	126	0.66	31	1.29	2.80	995	1,833	4,151	0.8
Bisha Phase 6	9	1.21	31	2,539	0.50	17	2.76	738	327	0.75	33	1.27	2.49	961	8,339	11,213	2.9
Bisha Phase 7	-	-	-	570	0.62	22	2.02	636	4,437	0.72	39	0.94	4.53	761	6,352	11,359	1.3
Bisha Phase 8	-	-	-	81	0.05	4	1.32	102	5,867	0.68	48	1.03	6.26	633	46,615	52,563	7.8
Bisha Phase 9	-	-	-	124	0.04	3	1.42	49	7,569	0.66	49	1.06	6.06	702	75,046	82,740	9.8
Bisha Total	107	2.68	79	7,298	0.61	27	3.54	962	18,389	0.68	46	1.02	5.66	702	143,712	169,506	5.6
Harena Total	80	4.93	16	-	-	-	-	-	1,161	0.52	22	0.64	3.57	-	7,020	8,261	5.7

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16.4 Mine Plan

The mine plan presented in this report was carried out using the Surpac MineSched scheduling tool. Descent rates were limited to 50 vertical metres per year. Drilling and blasting will be performed on 5 m benches, with loading carried out on 2.5 m flitches to minimize dilution and mining loss. The mine is scheduled to work 360 days a year (d/a), with five days allowed for delays due to weather disruptions. The plant is scheduled to operate 365 d/a.

The plan starts from the December 2013 end-of-month surveyed surface. The processing of oxide ore ended in mid-2013. Commissioning of the supergene plant started in Q3 2013, and commercial production announced 2 December 2013. Supergene material will be processed alone until mid-2016, when the zinc flotation plant comes online to begin processing primary phase material. Due to the sub-horizontal and undulating contact between the supergene and primary mineralized materials, there will be an approximate two-year period where both supergene and primary materials are mined. During this overlap period, both supergene and primary ores will be treated in campaigns to avoid mixing of the two ore types. The mine plan presented in this report was developed using a throughput of 2.4 Mt/a for all materials.

Table 16‑4 shows the summarized mine plan.

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Table 16‑4: Summarized Mine Plan

	Unit	2014	2015	2016	2017	2018	2019	2020	2021	2022	2023	2024	2025	Total
Total Mined	t ('000)	20,428	19,837	21,119	21,160	20,465	19,273	17,656	14,201	12,180	7,681	3,500	-	177,500
Total Waste Mined	t ('000)	17,755	17,228	18,836	18,479	18,233	16,843	15,265	11,459	9,696	5,479	1,192	-	150,465
SR	W/O tonnes	6.6	6.6	8.3	6.9	8.2	6.9	6.4	4.2	3.9	2.5	0.5	-	5.6
Total Ore Milled	t ('000)	2,330	2,446	2,457	2,509	2,458	2,500	2,471	2,464	2,464	2,464	2,442	461	27,464
Cu Grade	%	4.68	3.81	2.24	1.14	0.92	0.97	1.08	1.07	1.14	1.02	0.87	-	-
Zn Grade	%	-	-	3.89	4.36	4.34	5.55	5.71	5.93	6.00	6.35	6.08	-	-
Ore Mined by Phase
Phase 4	t ('000)	1,333	566	55	-	-	-	-	-	-	-	-	-	1,954
Phase 5	t ('000)	1,109	1,209	-	-	-	-	-	-	-	-	-	-	2,317
Phase 6	t ('000)	231	798	1,847	-	-	-	-	-	-	-	-	-	2,875
Phase 7	t ('000)	-	-	72	2,298	2,044	592	-	-	-	-	-	-	5,007
Phase 8	t ('000)	-	37	253	322	171	1,773	2,092	1,300	-	-	-	-	5,948
Phase 9	t ('000)	-	-	56	61	17	65	299	1,442	2,484	1,704	1,566	-	7,693
Harena	t ('000)	-	-	-	-	-	-	-	-	0.3	498	742	-	1,235
Waste Mined by Phase														-
Phase 4	t ('000)	4,567	656	37	-	-	-	-	-	-	-	-	-	5,260
Phase 5	t ('000)	1,444	389	-	-	-	-	-	-	-	-	-	-	1,833
Phase 6	t ('000)	3,501	2,459	2,378	-	-	-	-	-	-	-	-	-	8,339
Phase 7	t ('000)	-	-	287	2,489	2,784	792	-	-	-	-	-	-	6,352
Phase 8	t ('000)	8,243	8,983	6,245	6,292	2,363	8,602	4,477	1,410	-	-	-	-	46,615
Phase 9	t ('000)	-	4,740	9,889	9,698	13,086	7,449	10,788	9,369	7,049	2,412	566	-	75,046
Harena	t ('000)	-	-	-	-	-	-	-	680	2,648	3,066	625	-	7,019

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16.5 Waste Rock Storage

Over the remaining LOM, it is expected that 144 Mt of waste rock will be produced from the Bisha Pit and placed in two waste rock storage facilities (WRFs) to the east and southeast of the open pit, as well as in a backfill dump located in the north end of the ultimate pit, as shown in Figure 16‑5. The east WRF nominally covers 80 ha, and the southeast WRF nominally covers 90 ha. An operational scheduling plan has been prepared for placement of the rock within the dumps that allows potential acid rock drainage (ARD) issues to be appropriately managed during operations and closure.

Waste rock geochemical characterization studies by Graeme Campbell & Associates (Graeme Campbell) in 2013 confirmed potential acid forming (PAF) waste being significantly greater in volume than non-acid forming (NAF) mine waste. Within the current pit design, PAF waste is all material up to and inclusive of material within the upper saprolite zone. This is a fundamental change to original Feasibility-level assumptions of PAF starting only at the top of fresh rock. Occurrence of PAF material in the oxide zone is a result of alunites and jarosites existing within the weathered portion of the stratigraphy. These minerals are sulphates as opposed to sulphides, which break down on weathering to produce ARD.

In terms of their effects on mine design, NAF versus PAF dump volumes have been adjusted to accommodate expected greater volumes of PAF and lesser volumes of NAF requiring stockpiled over the Bisha LOM.

Positive results from the Bisha waste rock characterization include a PAF management strategy that does not include having to use HDPE liners or complicated under-drainage for PAF dump construction. Use of a thick, friable NAF soil, gossan, or oxide as an underlying PAF dump base will be sufficient to stop potential ARD penetrating and migrating into local aquifers. The arid landscape does not produce enough rainfall to saturate potential waste dumps to the extent that fluids could migrate through the PAF dump. The NAF friable base is sufficient to absorb and wick away any localized ARD formed in the early phases of a PAF dump. As the PAF dump increases in height, compaction and lack of moisture penetration will isolate the core of the PAF dump from ARD formation. Armouring the PAF dump with an external skin of NAF material to a maximum seasonal rainfall wetting front depth will effectively isolate the underlying PAF from oxidation.

These principles have been reviewed and accepted by KP. Construction of the Southern PAF Dump in 2013–2014 has used the outlined arid environment PAF containment strategy. PAF material that was mined prior to completion of construction of the Southern PAF Dump has been placed atop the Northern NAF Dump for eventual burial below a NAF capping.

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It is also expected that 7 Mt of remaining waste rock will be produced from the Harena Pit and placed in two WRFs to the east and southeast of the open pit, as shown in Figure 16‑6. Graeme Campbell completed a Harena waste rock characterization study in 2012. The study and associated testwork identified that alunites and jarosites were not as prevalent at Harena as they were within the lower oxide profiles of the Bisha and Northwest deposits. As with Bisha, arid conditions allow for simple PAF material containment through use of a thick, friable NAF base (soil, gossan, upper oxide), with progressive armouring of PAF with NAF as the waste dump is advanced and ultimately encapsulated.

As cutback of the Harena Pit has not been commenced to access supergene or primary ore, there is currently no need for a Harena PAF Dump.

Figure 16‑5: Bisha Site Layout

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Figure 16‑6: Harena Mine Site Layout

16.6 Blasting and Explosives

Production blasting is performed using ANFO when dry conditions allow, and emulsion when wet conditions are present, necessitating an explosives mixture that is resistant to water. Site infrastructure currently includes an ANFO plant. Packaged emulsion is transported to site.

Reactive Ground was first discovered in North Phase 4 below the 525mRL in late 2013. Production was negatively affected through the reaction of supergene sulphides (mainly marcasite) with uninhibited explosives, which resulted in a number of high risk incidents including smoking and fuming and prematurely detonating holes. BMSC engaged the services of Peter Bellairs Consulting Pty Ltd (Bellairs), who performed tests on reactivity and sleep time tests samples on 103 samples collected from the pit. Bellairs recommended changing to a urea containing HEF inhibited emulsion. This inhibited product has successfully negated reactivity issues associated with Reactive Ground at Bisha. Ongoing monitoring and follow-up testwork is required.

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16.7 Grade Control

The mine grade control process currently features vertical and angled RC holes piercing multiple benches on a 10 m x 10 m pattern. Two owner-operated Schramm T450 drill rigs are used to drill up to 2 benches (10 m) ahead of mining. Plans are in place to increase this to 4 benches ahead of mining. Each drill rig has a side-mounted sampling system consisting of a cyclone and underlying sample splitter. Grade control collects a primary 2 kg to 3 kg sample, and bags an up to 20 kg residual sample on a metre-drilled basis. Each metre is chipped, with representative chips stored in plastic trays for future reference. All holes are logged by mine geologists, with hole collar and logged data captured in the acQuire geological database.

Grade Control samples are currently assayed by the SGS on-site laboratory for Ag, Cu, Zn, As, and Pb using a 3-acid digestion with atomic absorption finish, with gold determined by fire assay. The lab procedures include QA/QC checks, such as the use of multiple commercially-prepared standards, blanks, and lab duplicates checks. Lab performance on standards and blanks is monitored, and when variances exceed a pre-defined tolerance, re-assays are requested. Samples are also collected, and a process of routine QEMSCAN/BMA analysis begun, to define both grade distribution and deportment of elements by mineral species. QEMSCAN/BMA analysis is currently undertaken at ALS Kamloops, Canada.

Assay data returned to Technical Services’ Mine Geology section is input into the acQuire geological database. All incoming data are reviewed within acQuire for QA/QC issues before being accepted as validated and incorporated into the database. Mine geologists extract data from the acQuire database for use in Surpac mine planning software. Mine Geology has incorporated GCX Grade Control Planning (GCX) software as an add-on to Surpac. GCX allows for rapid design of drill patterns, analysis, and estimation of incoming assay data, creation of ore blocks from estimated data, and comparison of data against reserves. At present, all grade control assay data are ordinary kriged as part of the process of designing and estimating ore blocks for mining.

Designed ore blocks are marked out by the Mine Survey team and delineated by coloured tape according to set grade ranges. Mining of ore blocks is supervised by trained “ore spotters,” ensuring that material mined is to tapes, and, where possible, visible ore/waste boundaries. Within the supergene portion of the Bisha orebody, ore/waste contacts are highly visible and sharp, generally being defined by the presence of massive or majority sulphides against unmineralized altered wall rock.

At end of 2013, gold oxide and copper supergene ore was being mined concurrently as the mine unevenly transitions into full supergene. The uneven transition is a function of erratic orebody weathering, an undulating supergene alteration surface, and the presence of faulting that has locally extended oxides to greater depths than originally resolvable at the scale of reserve drilling. This affects reconciliations between reserves, grade control, and declared ore mined.

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Ore mined from the Bisha open pit is directed by ore spotters to the ROM pad. Ore on the ROM pad is stockpiled on the basis of material type (oxide, supergene) and grade range of material type. Routine grab sampling of stockpiles is used to verify material being sent to the stockpiles is within the grade range of the allocated stockpile.

Mine Geology is responsible for determining the blend of material being fed into the Primary Crusher, based on agreed-upon forecast processing requirements.

16.8 Reconciliation

Due to the very recent completion of the new Bisha resource model, and the current situation in the pit, where many areas are still in transition from oxide to supergene mineralization, it is considered too early to provide reconciliation data regarding supergene production.

16.9 Hydrogeology

16.9.1 Pit Dewatering

For the Bisha pit, depressurization of the open pit slopes will be required to achieve the open pit slope designs. The Bisha Pit is currently being dewatered using a series of in-pit sumps and pumping systems. Previously utilized perimeter wells have been lost due to mine progression. BMSC retained KP in 2011 to provide hydrogeological services for the Bisha mine. Based on the 2013 field program and assessment of collected data, KP recommends depressurizing / dewatering through installation of horizontal gravity-drained holes in the pit walls combined with sumps for water removal. The first series of 50 m-long horizontal drains are scheduled for installation during 2014. Further study of the Bisha main hydrogeology and optimization of dewatering systems is recommended.

For the Harena Pit, depressurization of the open pit slopes will be required to achieve the open pit slope designs. During the oxide phase, seven wells were drilled, and two equipped for dewatering, with further support by a pit sump. Since the completion of the oxide pit phase, dewatering has been stopped, resulting in the formation of a significant pit lake. Dewatering will need to be re-established before Harena mining restarts in 2021.

16.9.2 Runoff Water

Runoff from waste dumps will potentially be unsuitable for release. As a result, all waste dump contact runoff will be collected. The footprint of waste dumps around the pits will reach their maximum extent by the end of 2015. During a 200-year precipitation event, approximately 15,000 m³ of water could be stored in the runoff collection ponds.

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Diversions have been constructed near the mine and plant site. Runoff is collected in storm drains and used in the process plant if the water is not contaminated. Facilities with the potential to spill contaminants require curbs or berms to contain any spills.

16.9.3 Fereketetet River Interception and Diversion

The hydrology of the project site is such that high intensity, short duration rainfall events occur during the rainy months, resulting in flash flooding situations. The open pit is located in the ephemeral drainage of the Fereketetet River. Diversion works have been constructed to intercept flow in the Fereketetet River during surface runoff events to prevent water entering the pit during development and operations. The diversion works consist of a dike across the river upstream (southeast) of the pit, which ultimately forms part of the Project’s long-term southern waste dump. This dike intercepts surface flow and, in extreme events, diverts the flow over a dividing ridge and into the adjacent Shatera River to the east. Additional diversion channels divert an eastern tributary of the Fereketetet River to the Shatera River, and divert the western Fereketetet River tributary to the Mogoraib River.

16.10 Mining Equipment

The current production equipment fleet as of March 2014 is shown in Table 16‑5.

Table 16‑5: Current Mine Production Fleet

Equipment Unit	Current Fleet
Drills
ROC L8 Drill	3
Pantera 1500 Drill	1
Loading Equipment
Terex RH40 Shovel	1
Terex RH40 Excavator	3
CAT 990H Loader	2
Haul Trucks
CAT 775 Truck	15

To determine the number of additional equipment units required for each major fleet, productivities were calculated using first principles based on estimated annual operating hours and mechanical availabilities. An 80% availability and 85% utilization was applied to all equipment. At peak production, which will occur in 2017, the equipment requirements will be four excavators, two front-end loaders, four drills, and twenty-four trucks. Equipment replacements were based on the following projected equipment lives in thousands of operating hours: 40 for excavators, 45 for haul trucks, 40 for front-end loaders, 30 for dozers, and 25 for drills.

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Projected ancillary equipment requirements included three road graders, five tracked dozers, three water trucks, two wheel loaders, support trucks, a utility loader and tire manipulator, a lighting plant, a tractor-trailer, a crew bus, and pickup trucks.

Future equipment replacements and additions are shown in Table 16‑6. Ongoing mine planning work may result in modifications to the equipment requirements presented above. An expansion to the maintenance facilities from four to seven bays is currently being planned.

Table 16‑6: Purchases Required for Additions and Replacements

Equipment	2014	2015	2016	2017	2018	2019	2020	2021	2022	2023	2024
D&B Drills		2		1
CAT 775 Trucks	3		6	6	4	3	5
Dozers	1		3			1
Front-End Loader					1
Graders							2
Terex RH 40 Diggers	1		1	1	1		1
Dewatering Pumps	2		1	2			2			2

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17 Recovery Methods

17.1 Supergene Process Plant

Bisha has three different types of mineralization—oxide, supergene, and primary—each requiring a specific process flowsheet. The plan in the 2006 Feasibility Study to mine and process each zone in succession starting with the top, Oxide Zone, is now into the Supergene Phase, with the Oxide Zone being mostly depleted during 2013. Additional process equipment to treat the supergene mineralization was commissioned by mid-2013 and installed downstream of the crushing and grinding “front-end” part of the carbon-in-leach (CIL) plant to treat the oxide ore. This equipment consists of flotation cells for copper roughing and cleaning duties, regrind mills for size reduction of rougher concentrate, copper concentrate thickener and pressure filters, a copper concentrate load-out building, copper flotation reagent systems, flotation air blowers, and pressure filter air compressors. Figure 17‑1 shows a flowsheet of the process for supergene ore.

Before the supergene mineralization is exhausted, the additional equipment required to process the primary mineralization will be installed and commissioned. Due to the sub-horizontal and undulating contact between the supergene and primary mineralized materials, there will be a multi-year period where both supergene and primary materials are mined. During this overlap period, supergene and primary ores will be treated in campaigns to avoid mixing of the two ore types.

For the treatment of primary mineralization, additional equipment will include zinc roughing and cleaning flotation circuits, a zinc concentrate regrind mill, zinc concentrate thickener and pressure filters, a zinc concentrate load-out building, zinc flotation reagent systems, an additional zinc flotation air blower, and a zinc pressure filter air compressor.

Some stockpiling of supergene and primary mineralization types will occur. The effect of any possible sulphide mineral oxidation on flotation performance should be minimized by management practices currently used in the base metal sulphide sector, such as reduced wetting of broken ore.

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Figure 17‑1: Supergene Process Flowsheet (crushing, grainding, and tailings circuits common with oxide)

Source: BMSC, 2012

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Updated LOM process plant throughput projections are shown in Table 17‑1. Recovery assumptions are provided in Table 15‑2, Table 15‑3, and Table 15‑5.

Table 17‑1: LOM Process Plant Throughput

General	Unit	Oxide	Supergene	Primary
Yearly Throughput	(Mt)	1.6	2.4	2.4
Operating Days/Year	(d)	365	365	365
Overall Plant Availability	(%)	91.3	91.3	91.3
Throughput Rate	(t/operating hour)	200	300	300
Tonnes per operating day (average)		4,380	6,575	6,575

The 1.6 Mt/a to 2 Mt/a oxide CIL plant processing facility achieved commercial production in February 2011, and operated until Q3 2013.

The oxide plant facilities include a primary crusher, SAG and ball grinding mills, cyanide leach/CIL circuit, cyanide destruction circuit, refinery to produce doré bullion, tailings thickener, tailings discharge system, and the necessary reagent, water, and air systems. A simplified schematic of the process flowsheet is shown in Figure 17‑2.

Those oxide CIL plant facilities not required for processing of supergene ore have been de-commissioned, but will be brought back into service later to leach remaining oxide mineralization.

The process facilities consist of a cominution circuit, an oxide leaching circuit, a copper flotation circuit and downstream tailings management. The oxide leaching circuit is currently on care and maintenance.

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Figure 17‑2: Oxide Process Flowsheet (crushing, grainding, and tailings circuits common with supergene)

Source: BMSC, 2012

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18 Project Infrastructure

A Site layout plan covering site facilities is shown in Figure 5-1. Power and communication descriptions are included in Section 5 of this Report. Constructed on-site infrastructure includes:

A road network connecting the open pit to the main processing area, waste dumps and to the maintenance complex, tailings facility, and camp.
Administration building. Offices and cubicles are provided for the mine management and supervisory staff as well as for human resources, accounting, procurement, information technology, and safety staff.
Maintenance workshop, warehouse and laboratory complex. The complex is located at the south end of the plant area adjacent to mine access road.
The permanent camp complex is located approximately 5 km to the northwest of the plant site. The camp has a capacity to accommodate up to 900 people and includes accommodation, kitchen and dining facilities, recreation facilities, laundry, water treatment, sewage treatment, incinerators and emergency power facilities.
The explosive magazine and ANFO mixing plant are located 1,000 m southeast of the process plant site and 450 m east of the ultimate footprint of the waste dump.
Fuel is stored in a bermed tank farm with two 1,500 m³ storage tanks. Current average daily use can range from 75,000 L to 90,000 L. At this consumption rate, the storage capacity is approximately 27 to 35 days. Ongoing mine planning will determine whether additional capacity will be required in the future.
Power plant, consisting of twenty-four 0.8 MW generators, currently serving an average demand of 8.5 MW.
Process control system.
Communications system.
Water supply (potable and process).

The TMF is a HDPE lined facility with three downstream constructed rockfill embankments. The original facility was designed by AMEC, and BMSC has retained KP to design the first embankment raise, which at the time of writing, is nearing completion. AGP considers it a reasonable expectation that the current reserves, and possibile additional material from future mine life expansions, can be accommodated within the TMF with additional designed dam raises.

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Off-site infrastructure has been established at the port of Massawa to support the export of copper and zinc concentrates. Rather than using a conventional bulk storage approach, requiring bulk concentrate storage, reclaim systems, travelling hopper, transfer tower, and shiploader, BMSC’s concentrate transport and storage approach uses the Rotainer® system. Concentrate is trucked in specially built, reusable intermodal containers from the Bisha Mine site to the port of Massawa, and stacked at the existing container facility. Container trucking is contracted and no specialized prime mover (truck) or trailer equipment is required. The copper concentrate containers are discharged into the bulk carriers using a crane equipped with Rotainer’s Lid-Lift equipment, which rotates the container a full 360° after lifting the lid. This system minimizes material rehandle, allows blending on a container–by-container basis, and is reported to provide industry leading dust control.

Delivery to the mine operation of required operating consumables and fuel as well as delivery to the various smelters of mine produced concentrates is most often subject to third party contractors, be it land transport in country or sea freight to and from the ocean port. Largely there are many factors outside the control of Bisha, which can adversely affect the delivery of these key consumables or the export of these concentrates and cause delays or temporary stoppage in product movement. Recently Bisha has been unable to fully achieve transport logistical efficiencies in the transportation of copper concentrate from the mine site to port. Of particular concern has been the performance of the transport contractor responsible for land delivery of copper concentrates from mine gate to the port of Massawa. While there remains a risk that the contractor will not be able to transport the required volume of concentrate due to various factors such as an inadequate number of trucks, poor maintenance of those trucks, or an inadequate number of trained drivers to operate the trucks, the risks have been recently mitigated by the contractor order another forty five trucks to nearly double its fleet while training more drivers. It is understood that this should allow for sufficient transport of concentrate from minesite to the Port at the required volumes by mid-2014.

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19 Market Studies and Contracts

19.1 Concentrate Quality

A limited number of observations during testwork on Bisha supergene mineralization at Maelgwyn, indicated the presence of minor amounts of enargite/tennantite and arsenopyrite, which can result in higher arsenic levels in the concentrate than those shown in Table 19‑1. Subsequent investigations have found that arsenian pyrite is also present. While arsenopyrite and arsenian pyrite can be expected to follow pyrite and not report to the supergene copper concentrate, under normal conditions for the recovery of copper sulphide minerals, enargite/tennantite will report to the copper concentrate. As mentioned in Section 15.1.2 NSR Calculations, since insufficient information was available to quantify the spatial limits of the enargite/tennantite, the conservative approach of assumed homogeneity throughout the supergene zone was taken. Assuming 75% of the arsenic is contained in enargite/tennantite, with the remainder in arsenopyrite, an arsenic recovery to copper concentrate of 67.5% was estimated. This trend has subsequently been confirmed by examination of monthly performance data. Further testwork is ongoing to better quantify and delineate the enargite/tennantite and arsenopyrite within the Supergene Zone.

Table 19‑1 shows detailed assays for copper and zinc concentrates produced from supergene and primary mineralization during the 2005 SGS Lakefield testwork.

Table 19‑1: ICP Scan on Flotation Concentrates

Element	Low Zn Primary		High Zn Primary		Supergene
Element	Cu Conc.	Zn Conc.	Cu Conc.	Zn Conc.	Cu Conc.
Cu %	28.7	0.94	24.9	0.72	28.5
Fe %	29.4	7.57	28.1	8.15	27.5
Pb %	0.98	0.25	2.44	0.39	0.25
Mo %	<0.001	0.004	<0.001	<0.001	0.005
Zn %	4.3	56.4	6.11	55.4	1.89
As %	0.034	0.035	0.055	0.018	0.33
Sb %	0.012	0.012	0.017	0.006	0.006
U %	<0.002	<0.002	<0.002	<0.002	<0.002
Bi g/t	<50	<50	<200	<50	<200
Cd g/t	140	1900	200	1700	99
Co g/t	<30	<30	49	<30	96
In g/t	<200	<200	<200	<200	<200
Ni g/t	<20	<20	<20	<20	140
P g/t	<40	<40	<200	<40	<30
					Table continues…

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Element	Low Zn Primary		High Zn Primary		Supergene
Element	Cu Conc.	Zn Conc.	Cu Conc.	Zn Conc.	Cu Conc.
AI %	0.032	0.021	0.06	0.032	0.013
Ca %	0.13	0.09	0.11	0.12	0.1
Cr g/t	<40	<40	<40	53	250
Mg g/t	710	140	770	150	120
Mn g/t	77	1300	170	1800	70
Si %	0.18	0.16	0.22	<0.07	0.16
Ti g/t	<10	<10	<8	<10	<30
V g/t	<80	<80	<80	<80	<80
Na %	<0.002	0.045	0.006	0.002	<0.002
K %	0.003	0.004	0.003	0.009	0.007
Ga g/t	28	39	22	38	4
Ge g/t	31	<2	29	<2	9
Se g/t	110	71	110	51	100
Te g/t	85	29	140	36	320
TI g/t	12	6.4	21	8.4	9
C (t) %	0.41	0.13	0.79	0.14	0.16
S %	35.4	32.9	35.5	33.4	40.3
CI g/t	4	3	<10	5	14
F %	<0.01	<0.01	<0.01	<0.01	<0.01
Hg g/t	1.6	10.1	0.8	9.9	4.0
Au g/t	7.08	0.73	7.02	0.56	3.5
Ag g/t	375	119	431	88.3	144
Pt g/t	<0.02	<0.02	<0.02	<0.02	<0.02
Pd g/t	<0.02	<0.02	<0.02	<0.02	<0.02
INSOL %	0.56	0.20	0.6	0.20	0.28

Blending allows the supergene copper concentrate to be kept below the 0.5% As limit generally required for copper concentrates sold to custom smelters.

Instances of elevated selenium levels in the copper concentrate encountered early in the treatment of supergene material appear to have been related to a transition type of mineralization between the nominal zones of oxide and supergene. Tellurium levels in the concentrate from supergene ore treatment are elevated above those shown in Table 19‑1, but will only incur a minor penalty.

Metallurgical testwork underway on primary material at ALS Metallurgy Kamloops has confirmed the validity of the SGS Lakefield data shown in Table 19‑1; hence, at this stage, both the copper and zinc concentrates are expected to meet the quality requirements of custom smelters.

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19.2 Commodity Price Projections

Pit optimization, mine planning, and the base case financial analysis have used commodity prices for copper, zinc, gold, and silver of $2.90/lb, $0.92/lb, $1,175/oz, and $20/oz, respectively.

19.3 Contracts

BMSC has negotiated contracts with two smelters for the sale of 60% of its future copper concentrate, with the remainder to either be sold on the spot market or the option to enter further agreements.

Normal commercial terms are included in the concentrate contracts, and are similar to typical industry standards.

Negotiations will be underway no later than 2015 for the sale of the zinc concentrates to be produced from the treatment of primary material. Terms contained within the concentrate sales contracts are likely to be typical of, and consistent with, standard industry practices, and be similar to such contracts elsewhere in the world.

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20 Environmental Studies, Permitting and Social or Community Impact

20.1 Environmental Regulatory Framework

The Eritrean Government’s mining legislation outlines two key provisions for EIAs on projects. A “Proclamation to Promote the Development of Mineral Resources”, No. 68/1995, Article 43 and the Regulations on Mining Operations, Legal Notice No. 19/1995, Article 5, both state that an EIA must be completed and submitted before a mining licence is granted. The “National Environmental Assessment Procedures and Guidelines, March 1999” (NEAPG) outlines the procedure for undertaking environmental assessments and clearance of projects. Approvals are the responsibility of the Department of Environment (DoE) of the Ministry of Land, Water, and Environment.

The SEIA was conducted to comply with Eritrean requirements and with the International Finance Corporation Performance Standards on Social and Environmental Sustainability (IFC Performance Standards, April 2006) where the latter are more stringent or comprehensive than national requirements. As noted in Section 4.6, the SEIA report was submitted in December 2006 and following review by the appropriate Eritrean Government agencies, a Mining Licence was issued in May 2008, signifying that environmental approval had been granted. Subsequently, BMSC indicated that various environmental studies were proceeding in order to provide more information for the operational management of the Project, and an SEIA update was issued in early 2009. The update includes more detail on the implementation of the social and environmental management plans that will manage the impact of the project and ensure employment of the proposed mitigation/enhancement measures.

20.2 Baseline Studies

Environmental baseline studies were performed as part of the SEIA process during 2006. Baseline studies were conducted for both Bisha and Harena projects and included the following:

Atmospheric environment
Noise
Terrain, soils, geology, soil chemistry and seismicity
Vegetation
Wildlife
Hydrology
Geohydrology
Socioeconomic conditions
Land use

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Archeology and cultural resources.

20.3 Environmental Issues

The key environmental issues assessed by the SEIA studies and addressed in Project associated risk assessments and the environmental management plan is as follows:

Direct footprint disturbance of 442 ha (Bisha) and 200 ha (Harena) with associated potential for loss of land use, habitat, soils loss and drainage disturbance
Groundwater impacts from both extraction of Project supply water from new wells and excavation of an open pit
Water quality impacts arising from potential for ARD, including the need to ensure that there is no post-closure problems with water quality
Soil and water quality impacts arising from the storage and use on site of hazardous chemicals, including cyanide
Changes to local surface drainage patterns due to construction of a site surface water management system, including flood control and diversion works; and
Air quality impacts, most significantly from surface haulage on unsealed roads.

BMSC has provided a remediation bond with the State of Eritrea in the amount of $15 million. BMSC has also accrued in its financial records a provision for mine rehabilitation of $23 million for the estimated costs of reclamation, remediation, and post-closure monitoring.

20.4 Closure Plan

KP, was retained to update the conceptual closure and reclamation plan (CCRP) in 2012. Closure considered:

TMF: non-PAF cover to prevent erosion from wind and surface water run-off; tailings delivery pipelines, power lines, and associated infrastructure will be decommissioned and cleaned of potentially hazardous materials; roads required to access the TMF will be decommissioned unless required to provide access for post-closure inspection, maintenance, and monitoring.
WRFs: at closure the final slopes will be no steeper than 3:1; non-PAF cover to prevent erosion from wind and surface water run-off; for budgetary purposes revegetation has been included in the closure costs; drainage from both of the WRFs will be directed towards the open pit through lined channels.
ROM pad facility: at closure the final slopes will be no steeper than 3:1; non-PAF cover to prevent erosion from wind and surface water run-off; for budgetary purposes revegetation has been included in the closure costs; drainage from both of the WRFs will be directed towards the open pit through lined channels.

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Open pit: there will be a pit lake upon cessation of operations, and a preliminary estimate of water quality was developed to support the closure plan; the preferred closure option for the Bisha Pit is to berm the pit and monitor the water quality during filling to identify any potential risk to avian life during this time; man-made equipment and materials (e.g., mining equipment, dewatering apparatus, cables) with salvage value will be removed from the open pit; materials and equipment with no salvage value will be cleaned of any potentially hazardous substances and disposed of in the open pit or in the non-hazardous landfill; access into the open pit will be blocked by installation of a rock boulder barrier across the access ramp/s into the pit. At closure, a security fence or an earthwork berm or a combination of both will be constructed.
Process plant: demolition and removal of equipment and structures to leave the property in an environmentally sound condition that will sustain accepted post-closure land uses; as part of demolition of the process plant infrastructure, allowance is made for the handling and disposal of potentially hazardous materials (e.g., hydraulic oil, gasoline, etc.) encountered during the dismantling process. Non-hazardous debris will be disposed of within the open pit or in the non-hazardous landfill. Potentially hazardous materials will be properly containerized and shipped offsite for recycling or disposal at an approved facility.
Ancillary facilities: will be decommissioned, cleaned, and removed from site following similar procedures to those envisaged for the plant; equipment and building systems will be inspected and cleaned of potentially hazardous materials; if practicable, materials will be salvaged, otherwise the facilities will be demolished and the debris will be disposed of in the open pit or in the non-hazardous land fill.
Roads: The main access road to the mine site and any other onsite access roads will be left in place for a minimum period of five years post-closure, to allow access to the site for post-closure maintenance and monitoring activities; roads that are no longer required will be permanently decommissioned.

Closure and post-closure monitoring will document the progress of the closure and reclamation effort. The elements of these monitoring programs will include:

Inspection of the physical conditions (e.g., for evidence of erosion and landslides) at the end of the initial rainy season post-closure
Inspection of the plantings after the first year post-closure
After two years, evaluation of the effectiveness of the reclamation effort (e.g., number and type of plant species, plant heights, productivity)
Demonstration that water quality objectives are met
Assessment of the adequacy and performance of drainage structures and sediment control systems.

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Closure and post-closure monitoring and control programs will be conducted twice per year (dry and wet seasons) for a period of five years after closure has been completed. In the event that deficiencies in vegetation establishment are identified, appropriate mitigation measures will be taken to correct these deficiencies.

Final reclamation of the pit, waste rock facilities, yards, and roads is estimated to cost approximately $18,900/ha, for a total of approximately $6.3 million.

Reclamation of the TMF, which includes the costs for the placement of a 40 cm cover, an evaporation pond, and the closure spillway, is estimated to be approximately $7.2 million. Revegetation is estimated to be approximately $0.5 million.

No salvage credits have been assumed for the closure costing.

Demolition and reclamation of the plant and other infrastructure is estimated to be approximately $4.6 million. Closure and post closure monitoring costs for a period of 10 years (five years of TMF drain-down and five years of post-closure monitoring) is estimated at $0.7 million.

The total reclamation cost is estimated at $29.9 million including administrative costs and a 15% contingency (this includes the cost of reclaiming both the WRFs).

20.5 Permitting

Permitting is discussed in Section 4.7.

The Bisha Project SEIA and Harena Project SEIA Addendum were conducted to comply with Eritrean requirements and with the International Finance Corporation Performance Standards on Social and Environmental Sustainability (IFC Performance Standards, April 2006) where the latter are more stringent or comprehensive than national requirements. As noted in Section 4.6, the Bisha SEIA report was submitted in December 2006 and following review by the appropriate Eritrean Government agencies, a Mining Licence was issued in May 2008, signifying that environmental approval had been granted. Subsequently, BMSC indicated that various environmental studies were proceeding in order to provide more information for the operational management of the Project, and an SEIA update was issued in early 2009. Subsequent to this, the Social and Environmental Management Plans (SEMP) have been developed and accepted by the Ministry of Environment. This document is considered “live” and undergoes regular review and updates, the latest having been during 2013. The Harena SEIA Addendum was submitted December 2011 with the mining licence granted in July 2012. The SEIA Addendum is still under review with additional requested studies complete and as per MoM and MoE requirements.

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20.6 Considerations of Social and Community Impacts

Since exploration and environmental baseline data collection began, considerable effort was spent developing support for the Project by fostering local relationships, developing a strong local workforce, educating stakeholders about the Project and mining in general and providing stakeholders with regular Project updates and, where appropriate, site visits.

The key socio-economic issues assessed by the SEIA study and addressed in the proposed social management and related plans are as follows:

Direct footprint disturbance of 442 ha with associated potential for displacement of people and their customary use of the land (although it is noted that the affected area is sparsely populated and only lightly used).

Influx of people seeking employment with associated potential issues, including pressure on existing social infrastructure.

Inward investment and creation of direct and indirect employment opportunity.

20.7 Discussion on Risks to Mineral Resources and Mineral Reserves

The QPs consider the environmental, permitting, and social risks to the Mineral Resources and Mineral Reserves to be minimal. Other than the provisional mining licence grated for Harena, all permits and licences are in order. Work is underway to fulfill the remaining provisions

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21 Capital and Operating Costs

The statement of capital and operating costs represent forward-looking information that are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here.

21.1 Capital Cost Estimates

21.1.1 Introduction

For the Bisha mine, the majority of the capital cost has already been spent and recovered on the gold CIL plant and the copper flotation plant. The copper plant was commissioned in mid-2013. From 2009 through end of 2013, the capital costs including gold plant, copper plant, sustaining capital, and mobile fleet expansion capital was $440 million.

Future capital expenditures are as follows:

21.1.2 Process Capital Costs

The zinc phase project is currently estimated at $89 million, including a contingency of $14 million. The testwork to support the design of the flotation circuit is underway at the time of writing. The cost breakdown by per period is shown in Table 21‑2.

21.1.3 Tailings Management Facility Expansion

The existing TMF will require periodic dam lifts during the remaining life of the project, the total cost of which is currently estimated at $40 million. The cost per period is shown in Table 21‑2.

21.1.4 Mine Capital Costs

Mine capital costs are limited to mobile fleet and dewatering replacements and additions. The equipment requirements are discussed in Section 16-11. Unit costs are based on recent purchase invoices by the mine for major equipment with recent quotations for support equipment. Mining capital requirements are $39 million during the mine life. The cost per period is shown in Table 21‑1.

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Table 21‑1: Mining Replacement and Additions Cost ($ ‘000s)

	Total	2014	2015	2016	2017	2018	2019	2020	2021	2022	2023	2024
Blasthole Drills	2,190	-	1,460	-	730	-	-	-	-	-	-	-
Trucks	23,830	2,860	-	5,720	5,720	3,810	2,860	2,860	-	-	-	-
Dozer	3,760	750	-	2,260	-	-	750	-	-	-	-	-
Front End Loader	1,340	-	-	-	-	1,340	-	-	-	-	-	-
Grader	950	-	-	-	-	-	-	950	-	-	-	-
Prime Excavator	5,900	1,180	-	1,180	1,180	1,180	-	1,180	-	-	-	-
Dewatering Equipment	900	200	-	100	200	-	-	200	-	-	200	-
Total	38,870	4,989	1,460	9,260	7,830	6,330	3,610	5,190	-	-	200	-

21.1.5 Sustaining Mill, Camp, and Port Capital

A total of $31 million in sustaining capital is estimated for replacements and upgrade for the mill, camp and port facilities.

A summary of the anticipated LOM capital expenditures is shown Table 21‑2.

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Table 21‑2: Capital Cost Summary ($ ‘000s)

Description	Total ($'000)	2014	2015	2016	2017	2018	2019	2020	2021	2022	2023	2024	2025
Phase III Zinc Plant Capital	89,000	30,000	40,000	19,000	-	-	-	-	-	-	-	-	-
TMF Dam Raise Costs	40,000	4,000	-	-	10,000	-	-	12,000	-	14,000	-	-	-
Mining Fleet Replacements/Additions	38,900	5,000	1,500	9,300	7,800	6,300	3,600	5,200	-	-	200	-	-
Mill, Camp, and Port	31,100	3,300	9,500	4,500	2,700	4,500	1,600	1,600	1,450	1,300	550	100	-
Total	199,000	42,300	51,000	32,800	20,500	10,800	5,200	18,800	1,450	15,300	750	100	-

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21.2 Operating Cost Estimates

21.2.1 Basis of Estimate

The costs presented in this section reflect Q4 2013 dollars. Inflation has not been in included. Labour rates costs reflect current pricing staffing and labour rates at the mine site, as used in the BMSC 2014 budget. The diesel price used for 2014 budget cost estimation was $1.19/L, delivered to site.

21.2.2 Mine Operating Costs

The mine operating costs are based on BMSC’s 2014 budget mining cost estimate. The mining cost (inclusive of operations, maintenance, grade control, geology, and lab services) was $2.73/t, plus an appropriate incremental haulage cost per bench of $0.01/t/5 m. This is exclusive of a sustaining capital allowance that was added for cutoff and pit optimization purposes. The LOM average cost with incremental haul additions is $2.87/t mined. Ore from the Harena satellite deposit has an additional $2.63/t overland haul cost, which was treated as an ore based cost for cutoff and pit optimization purposes.

21.2.3 Process Operating Costs

The process cost is based on BMSC’s 2014 budget process cost estimate, which is $28.74/t, which includes maintenance, power, labour, and consumables. Note this is exclusive of a sustaining capital allowance that was added for cutoff and pit optimization purposes.

21.2.4 G&A Operating Costs

G&A costs are based on the 2014 Budget estimate $10.04/t milled.

21.2.5 Downstream Costs

Downstream costs or costs incurred while transporting sellable end-products to their destinations plus smelting and refining charges are $288.66/t of copper concentrate, $405.37/t of zinc concentrate and $5.12/oz of gold doré.

21.3 Operating Cost Summary

The annual mining and ore based costs are shown in Table 21‑3.

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Table 21‑3: Annual Operating Costs ($ ‘000S)

	Units	Totals	2014	2015	2016	2017	2018	2019	2020	2021	2022	2023	2024	2025
Total Mined	kt	177,500	20,428	19,837	21,119	21,160	20,465	19,273	17,656	14,201	12,180	7,681	3,500	-
Total Ore Milled	kt	27,464	2,330	2,446	2,457	2,509	2,458	2,500	2,471	2,464	2,464	2,464	2,442	461
Mining Cost	$ ‘000s	509,968	56,177	54,822	58,795	59,731	58,645	56,121	52,470	42,873	36,858	22,799	10,677	-
Mining Cost	$/t mined	-	2.75	2.76	2.78	2.82	2.87	2.91	2.97	3.02	3.03	2.97	3.05	-
Process + G&A	$ ‘000s	1,068,323	90,366	94,836	95,288	97,293	95,327	96,952	95,806	95,544	95,545	96,855	96,637	17,874
Process + G&A	$/t milled	-	38.78	38.78	38.78	38.78	38.78	38.78	38.78	38.78	38.78	39.31	39.58	38.78

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22 Economic Analysis

NI 43-101 regulations exempt producing issuers from the requirement to disclose Economic Analysis on properties currently in production, unless the technical report prepared by the issuer includes a material expansion of current production. Nevsun is a producing issuer, the Bisha mine is currently in production, and a material expansion is not included in the current Bisha LOM Plan. The QP has performed an economic analysis using the Mineral Reserves and Life-of-Mine Plan presented in this report, and confirms that the outcome is a positive cash flow that supports the statement of Mineral Reserves.

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23 Adjacent Properties

Some 10 km to the north of Bisha, Chalice Gold Mines (Chalice) hold the Mogoraib North exploration licence, which is effectively adjacent to Bisha. This relationship is shown in Figure 23‑1 . In October 2012, Chalice reported a drill intersection of 10 m of massive pyritic sulphide mineralization containing approximately 1% Cu and 2% Zn (from portable XRF analysis) at the T209 discovery shown in Figure 23‑1, (Source¹). In their 2013 Annual Report, (Source²) Chalice state that similar intersections were obtained in their follow-up drilling program, but the “mineralisation discovered….remains uneconomic due to low grades and thicknesses identified.” Chalice additionally report that their focus for Mogoraib North will be to “….assess the best way forward for the Mogoraib North JV (with ENAMCO).”

There are no other properties immediately adjacent to Bisha or its project areas.

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Figure 23‑1: Chalice Gold Mines Licence Holding and Bisha Properties

Source:	www.chalice.com/projects, accessed March 2104

Source:	¹: http://www.prnewswire.com/news-releases/chalice-intersects-massive-sulphides-at-mogoraib-north-176587371.html, accessed March 2014.

Source:	http://chalicegold.com/upload/documents/InvestorRelations/releases/20131025ASXAnnouncementAnnua lReport%28glossy%29.pdf, accessed March 2014.

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24 Other Relevant Data and Information

There are no other data that are relevant to the Report.

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25 Interpretation and Conclusions

Following evaluation of the Project, the QPs are of the opinion that:

Based on information provided by BMSC to AGP, the mining tenure held is valid, and sufficient to support declaration of Mineral Resources and Mineral Reserves.
BMSC holds sufficient surface rights to support mining operations over the planned life-of-mine, and the declaration of Mineral Resources and Mineral Reserves.
Royalties payable include an Eritrean Government royalty of 5.0% of precious metal NSR and 3.5% of base metal NSR.
Permits obtained by BMSC to operate the mine and undertake Project development are sufficient to ensure that mining activities are conducted within the regulatory framework required by the Eritrean government. BMSC has a fully approved mining license for Bisha Main; the Mining Licence for Harena remains conditional although BMSC has met all its obligations providing detailed geotechnical design, hydrogeological and waste rock characterization studies to the Eritrean government by the end of December 2012.
At the effective date of this report, environmental liabilities are considered to be typical of a mine that will produce base metals, and include an open pit, tailings and waste rock facilities, mining infrastructure, and roads.
Environmental impacts were assessed during the permitting process for Bisha and Harena. The required environmental permits for the Bisha operation have been issued and environmental bond has been lodged.
Closure requirements were assessed by Knight Piésold during 2010 and have been escalated to include Harena closure requirements.
The existing and planned infrastructure, availability of staff, the existing power, water, and communications facilities, logistics, and any planned modifications or supporting studies are well-established, or the requirements to establish such, are understood by BMSC.
The geologic understanding of the deposit settings, lithologies, and structural and alteration controls on mineralization has been enhanced from experience gained in production and is sufficient to support estimation of Mineral Resources.
The setting of the Bisha, Harena, Northwest, and Hambok deposits are well understood and the exploration programs and drilling completed to date are appropriate to the VMS style of mineralization. The quantity and quality of the lithological, geotechnical, collar, and downhole survey data collected in the exploration, delineation, and grade control programs are sufficient to support Mineral Resource and Mineral Reserve estimation. Data collection, management and QC monitoring of historical and new data are completed by a dedicated team using industry standard software (acQuire).

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Sampling methods are acceptable, meet industry-standard practice, and are adequate for estimating Mineral Resources.
The quality of the gold, silver and base metal analytical data are reliable. Sample preparation, analysis, and security are currently performed in accordance with industry exploration practices and industry standards.
The quantity and quality of bulk density sampling is adequate for estimating Mineral Resources.
The estimate and classification of Mineral Resources for Bisha Main and Harena conform to industry best practices, and meet the requirements of CIM (2010). The Mineral Resources are adequate to support mine planning. Initial reconciliation from mining of the supergene mineralized zone indicates a good correlation with predicted values.
Estimations of Mineral Reserves for the Project conform to industry best practices, and meet the requirements of CIM (2010). Reviews of the environmental, permitting, legal, title, taxation, socio-economic, marketing, and political factors and constraints for the operation support the declaration of Mineral Reserves using the set of assumptions outlined.
The mine plans are appropriate for the style of mineralization.
The main risk to achieving the open pit slope designs is ineffective depressurizing the rock mass of the excavation. Additional pumping wells and horizontal drains are required. The detailed design of the depressurization system will require further analysis based on the updates to the open pits.
Opportunities to increase the inter-ramp angles used in the current work include:

- The mitigation of rock falls at the bench scale with good blasting and scaling may allow the catch bench widths to be reduced by 1 m; this could result in up to a 2 to 3° increase in the inter-ramp angles in design sectors where the bench geometry controls the design inter-ramp angle.

- Good rock quality at depth and partially healed or wide spacing between breaks along foliation may eliminate toppling as a credible failure mode for some slopes. With appropriate slope monitoring and depressurization, these conditions could result in increases in the inter-ramp angles of up to 7° in certain design sectors.

Further optimization of the mine plan is underway to investigate opportunities to defer construction of the Zinc plant. This work involves a trade-off between processing the Supergene material over a shorter period of time and deferring the zinc plant capital expenditure versus adding additional mine capacity and stockpiling more primary ore for longer periods of time. The results from additional ongoing variability metallurgical test work will feed into this analysis.
It is the QPs’ opinion that the metallurgical test work completed to date on the Bisha Main and Harena deposits have established appropriate metallurgical recoveries and processing routes for the different mineralization styles in the various deposits to a level sufficient to support Mineral Reserves declaration.

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The economic analysis of the combined Bisha Main and Harena is positive under the set of assumptions used, indicating a robust mining operation.

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26 Recommendations

The QPs recommendations for the Bisha operation, satellite deposits, and regional work are described below.

26.1 Mineral Resources

The recommendations for each deposit are as described below.

26.1.1 Bisha

Continue in-house grade control drilling to provide additional detail in delineation of supergene mineralized domains and identification of the zinc depletion surface that indicates the start of the primary zone.
Drilling of the down dip extension of the main massive sulphide zone to the west and below the current optimal pit shape limiting the Mineral Resource in order to gain better understanding of the potential for underground mining. The envisaged program of eight holes for 4,800 m would cost in the order of $1.75 million; continuation of such a program would be based on results.
Continue in-house in-situ density testing of the supergene mineralized zone.
Complete an in-house desktop study of underground potential with a view to increasing Resource inventory and understanding of future direction for mining. See also recommendations for Reserves below.

26.1.2 Harena

Drilling for potential strike and dip extension in the centre and south of the deposit. A program of eight holes for 2,400 m is envisaged to determine potential for relatively shallow mineralization costing an estimated $800,000.

26.1.3 Northwest

Drilling of the Eastern zone to better define and upgrade Resource category; identify potential for shallow mineralization to be added to the current Resource. A program of 16 holes for 1,500 m at an estimated cost of $0.6 million is envisaged.
RC drilling of oxide and supergene gold cap to improve sample recovery and better delineate these zones with accompanying upgrade to Indicated category. A program of 120 short holes for 7,200 m at a cost of approximately $250,000 is envisaged.

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Down dip drilling of the primary massive sulphides to identify potential as follow-up to the successful deeper drilling completed in 2013. A program of 21 holes for 6,000 m at an approximate cost of $2.1 million is envisaged.

26.1.4 Hambok

Complete Resource infill program within the current Resource area to extend on strike and better delineate near surface primary mineralization. A program of eight holes for 2,850 m is envisaged at a cost of approximately $1.0 million.
RC drilling of the oxide zone to examine potential for shallow mineralization with strike extension, particularly to the south, and for better delineation. An initial program of 30 holes for 2,250 m is recommended at an approximate cost of $800,000.

26.2 Mineral Reserves

The recommendations for each deposit are described below.

26.2.1 Bisha

Complete an in-house trade-off study examining possible underground mining versus ultimate pit expansion. Approximate cost (limited input from external consultants) is $50,000.
Geotechnical:

- BMSC needs to examine blasting practices to develop a method of incorporating geotechnical information into the blast designs, since rock quality varies considerably within domains.

- Continue development of Slope Management Plan

Hydrogeological:

- The Knight Piésold hydrogeological report (2014) identified three aquifers that will influence the seepage of groundwater into the pit, of which the middle Fractured Bedrock aquifer is the most significant as it will be encountered throughout the life of mine. This aquifer however needs to be correlated with the major structures and the geotechnical domains, which are expected to have different hydraulic conductivities and will change as mining progresses. The FRK (fresh rock unit) has very good RQDs and should have a lower hydraulic conductivity than the overlying WRK (weathered rock unit) in which most mining is currently taking place.

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- An ongoing program of targeted horizontal depressurization holes into the pit slopes needs to be implemented as mining progresses to relieve the pore pressures that build up behind the pit walls.

Reconcilation

- Confirm process plant production data vs. concentrate shipment outturn data once a consistent pattern of concentrate movements has been established.

Confirm metallurgical response of transition material identified between supergene and primary zones

- Additional testwork to examine processing characteristics for the transition zone of mixed copper minerals, with zinc, at the top of the primary zone; approximate approximate cost is $200,000.

26.2.2 Harena

Additional geotechnical drilling targeting the primary sulphide Harena open pit should be completed to provide data for rock mass characterization, hydrogeological characterization, and structural geology assessments. Geotechnical core logging, packer testing, and core orientation should be completed in 3 to 5 holes.

Metallurgical testwork on supergene material, as for primary transition mineralization above, is dependent on potential quantity, which will be assessed.

26.2.3 Northwest

Complete a prefeasibility study leading to Mineral Reserve estimate from existing data for improved LOM planning options; approximate cost is $200,000.

26.2.4 Hambok

In line with Eritrean government’s requirements:

Complete an in-house study to better understand project economics; approximate cost (limited input from external consultants) is $25,000.

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Preparatory to a prefeasibility study leading to a Mineral Reserve estimate:

- complete a program of six geotechnical holes, testing, analysis and reporting; approximate cost is $1.0 million

- complete a program of three metallurgical test holes with associated testwork, analysis and reporting; approximate cost is $500,000.

- complete a program of four hydrogeological test holes, with associated pump testing, analysis and reporting; approximate cost is $700,000

- complete a program of waste rock characterization; approximate cost is $200,000

Complete a prefeasibility study leading to Mineral Reserve estimate for improved LOM planning options; approximate cost is $20,000.

26.3 Metallurgical

Additional metallurgical studies for Bisha Main, Harena and the Northwest zone include:
Additional batch and locked cycle flotation testwork is recommended to support the future design of the primary ore flotation circuit and support associated metals recovery factors.
Investigate the effects of oxidation on supergene and primary flotation performance recognizing that this issue is common to other volcanogenic massive sulphide deposits.
Investigate the occurrence and deportment of gold and silver in Bisha primary mineralization to improve payability.

These metallurgical recommendations have an approximate cost of $1,500,000.

26.4 Exploration

Additional exploration work is envisaged both regionally and around the Bisha/Harena/Northwest deposits and programs of diamond drilling, geophysics, geochemistry, and associated works are planned.

Within the Mogoraib EL, drilling will be along the Hambok mineralized trend and related targets.

The approximate cost of exploration activities for 2014 is 2.85 million.

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28 Certificate of Authors

The qualified persons’ certificates are included overleaf.

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Jay Melnyk, P.Eng.
AGP Mining Consultants, Inc.
92 Caplan Avenue, Suite 246, Barrie, Ontario, L4N 0Z7

CERTIFICATE OF QUALIFIED PERSON

I, Jay Melnyk, P.Eng., am employed as Principal Mining Engineer with AGP Mining Consultants, Inc.

This certificate applies to the technical report entitled “Bisha Mine, Eritrea, Africa, NI 43-101 Technical Report” (the “Technical Report”) with an effective date of December 31, 2013 and a report date of 24 March 2014.

I am a Professional Mining Engineer. I graduated from the Montana Tech of the University of Montana with a Bachelor of Mining Engineering degree in 1988 and from the British Columbia Institute of Technology with a Diploma in Mining Technology in 1984.

I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (Registration #25975) and the Association of Professional Engineers and Geoscientists of Newfoundland and Labrador (Registration #06438).

I have practiced my profession for 25 years. I have been directly involved in open pit mining operations, and design of open pit mining operations in Argentina, Canada, Chile, Eritrea, Indonesia, Iran, Mexico, Perú, and the United States. As a result of my experience and qualifications, I am a Qualified Person as defined in NI 43–101.

I visited the mine most recently from 9 to 15 November 2013.

I am responsible for the Sections 3, 5, 15, 16, 18, 20, 21, 22, and portions of Sections 1, 24, 25, 26, and 27.

I am independent of the Issuer (Nevsun Resources Ltd.) as defined by Section 1.5 of the Instrument.

I have co-authored the two previous Technical Reports on this property.

I have read NI 43-101 and the sections of the Technical Report which I am responsible for have been prepared in compliance with NI 43‑101.

At the effective date of the Technical Report, to the best of my knowledge, information, and belief, the portions of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 24^th day of March, 2014.

“Original signed and sealed”

Jay Melnyk, P. Eng.

MINERALURGY PTY. LTD.

A.C.N. 094 603 166

Consultants to the mining, metallurgical and process industries

Tel +61-(0)7 3870 7024
Fax +61-(0)7 3870 7606
PO Box 818
Toowong
Queensland 4066
AUSTRALIA

I, Peter Munro, FAusIMM, of Taringa, Queensland , Australia, as a qualified person (QP) of this Technical Report, Bisha Mine, Eritrea Africa, NI 43-101 Technical Report with an effective date of December 31, 2013 and a report date of 24 March 2014, do hereby certify the following statements:

I am a Senior Principal Consulting Engineer with Mineralurgy Pty Ltd, with a business address at Unit 2, 42 Morrow Street, Taringa, Queensland, 4068, Australia.
I am a graduate from the University of Adelaide with a BAppSc in Applied Chemistry in 1970. In addition I have obtained a BEcon degree in 1975 and a BComm degree in 1979 from the University of Queensland.
I am a Fellow of the Australasian Institute of Mining and Metallurgy (member number 104257); Member of Engineers Australia; Member of the Institution of Materials, Mining and Metallurgy (U.K.); Member of the Society of Mining Metallurgy and Exploration (USA); Member of the Canadian Institute of Mining, Metallurgy and Petroleum and Member of the Southern African Institute of Mining and Metallurgy.
I have worked as a metallurgist for a total of 44 years since graduation, predominantly in base metals and gold mineral processing operations, process development and project development. I have been directly involved in mineral processing operations in Argentina, Australia, Canada, Chile, Eritrea, Lao PDR and Papua New Guinea.
As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101).
I have not visited the project site.
I am responsible for the Sections 13, 17, 19, 26 (part) and 27 (part) in the technical report titled “Bisha Mine, Eritrea, Africa, NI 43-101 Technical Report”, dated 24 March 2014.
I have read the definition of “qualified person” set out in National Instrument 43 101 (NI 43 101) and certify that, by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purpose of NI 43-101.

As of the date of this Certificate, to my knowledge, information, and belief, this technical report contains all scientific and technical information that is required to be disclosed to make the technical report not misleading.
I am independent of the Issuer (Nevsun Resources Ltd. (Nevsun)) as defined by Section 1.5 of the Instrument.
I have co-authored the previous Technical Report on this property.
I have read NI 43-101 and the sections of the Technical Report which I am responsible for have been prepared in compliance with NI 43‑101 and Form 43-101F1.

Signed and dated this 24^th day of March 2014, at Taringa, Queensland, Australia.

“Original Signed and Sealed”

Peter Munro FAusIMM

Paul Gribble C.Eng., FIMMM
Bisha Mining Share Company
P.O. Box 4276, Asmara, Eritrea

CERTIFICATE OF QUALIFIED PERSON

I, Paul Gribble C.Eng., FIMMM, am employed as Chief Resource Geologist, Bisha Mining Share Company.

I am a Professional Geologist. I graduated from University College, University of London, with a Bachelor of Science degree in geology in 1977.

I am a member in good standing of the Institute of Minerals, Mining and Materials, membership number 50145 and Chartered Engineer with the Engineering Council, UK, registration number 518788;

I have practiced my profession for 35 years. I have been directly involved in geology and Resource estimation in precious and base metals in operations and projects in a range of geological environments; with particular reference to copper/zinc/gold mineralisation, in Botswana, Eritrea, and Ecuador. As a result of my experience and qualifications, I am a Qualified Person as defined in NI 43–101.

I am an employee of Bisha Mining Share Company based at the mine site in western Eritrea since December 2012.

I am responsible for the Sections 2, 4, 6, 7, 8, 9, 10, 11, 12, 14, 23, and portions of Sections 1, 24, 25, 26, and 27.

I am not independent of the Issuer (Nevsun Resources Ltd.) as defined by Section 1.5 of the Instrument.

I had no previous involvement with the Property prior to my employment.

I have read NI 43-101, and the sections of the Technical Report which I am responsible for have been prepared in compliance with NI 43‑101.

Dated this 24^th day of March, 2014.

“Original signed and sealed”

Paul Gribble C.Eng., FIMMM