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GSS Golden Star Resources

Filed: 11 Mar 21, 4:20pm

Exhibit 99.1

 

 

 

 

 

 

 

 
NI 43-101 Technical Report (March 2021)Wassa Gold Mine

Contents

 

1   Executive Summary16
1.1   Terms of Reference16
1.2   Location and Setting16
1.3   Mineral Tenure, Permits, Royalties and Agreements16
1.4   History17
1.5   Geology and Mineralization17
1.6   Drilling and Sampling18
1.7   Data Verification18
1.8   Metallurgical Test Work19
1.9   Mineral Resource Estimate19
1.10   Mineral Reserve Estimate21
1.11   Mining Methods21
1.12   Recovery Methods24
1.13   Infrastructure26
1.14   Environmental Studies, Permitting and Social or Community Impact27
1.15   Capital and Operating Costs28
1.16   Economic Analysis29
1.17   Preliminary Economic Assessment of the Southern Extension Zone30
1.18   Conclusions and Interpretations32
1.19   Recommendations36
2   Introduction38
2.1   Terms of Reference38
2.2   Wassa Gold Mine38
2.3   Principal Sources of Information39
2.4   Qualified Persons39
2.5   Effective Dates39
2.6   Previous Technical Report40
3   Reliance on Other Experts41
4   Property Description and Location42
4.1   Location of Mineral Concessions42
4.2   Mineral Rights45
4.3   Royalties and Other Payments; Encumbrances46
4.4   Historic Environmental Liability and Indemnity46
4.5   Permits and Authorization47
5   Accessibility, Climate, Local Resources, Infrastructure and Physiography49
5.1   Accessibility49
5.2   Physiography and Vegetation49
5.3   Land Use and Proximity to Local Population Centres49
5.4   Local Resources and Infrastructure50
5.5   Climate and Length of Operating Season50
6   History52
6.1   Wassa52
6.2   Hwini Butre, Benso and Chichiwelli52
6.3   Production History, Previously Declared Resources and Reserves53
7   Geological Setting and Mineralization55

 

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7.1   Regional Geology55
7.2   Local Geology and Mineralization57
8   Deposit Types71
8.1   Wassa71
8.2   Hwini Butre72
8.3   Hwini Butre74
8.4   Chichiwelli75
9   Exploration76
9.1   Wassa76
9.2   Hwini Butre78
9.3   Benso and Chichiwelli79
10   Drilling80
10.1   Surface Drilling80
10.2   Underground Drilling81
10.3   Sampling81
11   Sample Preparation, Analyses and Security84
11.1   Sample Preparation84
11.2   Sample Dispatch and Security84
11.3   Laboratory Procedures84
11.4   Quality Control and Quality Assurance87
11.5   Specific Gravity Data104
12   Data Verification106
12.1   Drilling Database106
12.2   Other Verifications by the Qualified Person106
13   Mineral Processing and Metallurgical Testing107
13.1   Early Metallurgical Test Work107
13.2   2015 Test Work Program107
13.3   Test Work Findings109
14   Mineral Resources119
14.1   Introduction119
14.2   Mineral Resource Estimation Procedures120
14.3   Mineral Resource Database121
14.4   Grade Shell Modelling123
14.5   Statistical Analysis and Variography132
14.6   Block Model and Grade Estimation151
14.7   Model Validation and Sensitivity156
14.8   Mineral Resource Classification165
14.9   Mineral Resource Statement171
14.10   Mineral Resource Risks173
15   Mineral Reserves174
15.1   Cut-off Grade174
15.2   Modifying Factors175
15.3   Mineral Reserve Statement175
15.4   Mineral Reserve Risks176

 

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16   Mining Methods177
16.1   Mineral Resources Considered in Mining Plan177
16.2   Mining Locations178
16.3   Current and Upper Mining Zones (Panels 1-3)180
17   Recovery Methods207
17.1   Processing History207
17.2   Flow Sheet Description207
17.3   Processing Schedule210
18   Infrastructure213
18.1   Electrical Infrastructure215
18.2   Surface Water Management217
18.3   Workshops and Other Site Buildings218
18.4   Site Accommodation218
18.5   Waste Rock Storage219
18.6   Tailings Storage221
19   Market Studies and Contracts227
19.1   Market Studies227
19.2   Contracts227
20   Environmental Studies, Permitting and Social or Community Impact228
20.1   Relevant Legislation and Required Approvals228
20.2   International Requirements232
20.3   Environmental and Social Setting233
20.4   Environmental and Social Management246
20.5   Environmental and Social Issues249
20.6   Closure Planning252
21   Capital and Operating Costs254
21.1   Introduction254
21.2   Capital Costs254
21.3   Operating Costs257
21.4   Closure Costs259
22   Economic Analysis260
22.1   Assumptions260
22.2   Stream, Taxes and Royalty261
22.3   Economic Results, Base Case261
22.4   Economic Results, Consensus Case263
22.5   Sensitivity265
23   Adjacent Properties267
24   Other Relevant Data and Information268
24.1   Southern Extension PEA Introduction268
24.2   Mineral Resources used in the PEA270
24.3   Mining Methods271
24.4   Metallurgical Testing302
24.5   Recovery Methods309
24.6   Infrastructure311
24.7   Environmental, Permitting and Social and Community Impact311

 

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24.8   Closure Planning312
24.9   Capital and Operating Costs312
24.10   Economic Analysis317
24.11   Conclusions and Interpretations323
25   Conclusions and Interpretations328
25.1   Conclusions328
25.2   Risks332
25.3   Opportunities334
26   Recommendations336
26.1   Current Operations336
26.2   Southern Extension Zone337
27   References341
28   Date and Signatures345

 

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NI 43-101 Technical Report (March 2021)Wassa Gold Mine

LIST OF FIGURES

Figure 1-1 Underground Production History and Mineral Reserve plan24
Figure 1-2 Processing Production History and Mineral Reserve plan25
Figure 1-3 Gold Production History and Mineral Reserve plan26
Figure 1-4 Processing schedule for Southern Extension PEA31
Figure 1-5 Gold Production Schedule for Southern Extension PEA31
Figure 1-6 Project Execution Plan, Southern Extension Panels 4 and 537
Figure 4-1 Wassa Mine Location in Ghana, West Africa (United Nations, 2018)42
Figure 4-2 Wassa Mine Location in Ghana, West Africa (GSR, 2021)43
Figure 4-3 Location of operations and infrastructure and concession boundaries (GSR, 2021)44
Figure 7-1 Location of Wassa on the Ashanti Belt (Perrouty et al 2012)56
Figure 7-2 Total magnetic intensity reduced to pole, of the Ashanti Belt (modified from Perrouty et al, 2012)58
Figure 7-3 Compilation of geochronology dating from the Ashanti Belt (Perrouty et al, 2012)59
Figure 7-4 Regional geology of the Ashanti belt, showing Wassa, GSR tenure and major deposits (GSR, 2020)61
Figure 7-5 Wassa mine-scale geology (modified from Bourassa, 2003 and Perrouty et al, 2013)62
Figure 7-6 Vertical section through Nose of deposit-scale F4 fold, Wassa Main deposit63
Figure 7-7 Eburnean folds and foliations from Wassa mine, Starter pit64
Figure 7-8 Eburnean folds and foliations from Wassa mine, B-Shoot pit65
Figure 7-9 Wassa section through 19,650 mN showing high-grade zones, F3 closures, parasitic folding66
Figure 7-10 Wassa section through 19,925 mN showing interpretation with tight-spaced drilling67
Figure 7-11 Wassa section through 18,900 mN showing interpretation and wide spaced (surface) drilling67
Figure 7-12 Hwini Butre section through 33,100 mN69
Figure 8-1 Syn-Eoeburnean veins from B-Shoot, 242 and South-east zones (modified from Perrouty et al, 2013)72
Figure 8-2 Mineralization exposure in Father Brown pit, smoky quartz vein73
Figure 8-3 Mineralization exposure in Adoikrom pit, potassic alteration73
Figure 8-4 Mineralization exposure in Subriso West pit, sheared volcanics74
Figure 8-5 Mineralization exposure in Subriso East pit, fine grained pyrite74
Figure 8-6 Mineralization at Chichiwelli East, hydrothermal veins75
Figure 8-7 Mineralization at Chichiwelli West, shear hosted75
Figure 9-1 Wassa soil geochemistry and anomalies (GSR, 2018)76
Figure 9-2 Wassa airborne magnetic coverage (GSR, 2004)77
Figure 11-1 Transworld Laboratories sample processing flow sheet85
Figure 11-2 Intertek sample processing flow sheet86
Figure 11-3 HARD plot comparing fire assay and BLEG for field duplicates88
Figure 11-4 HARD plot of all coarse rejects (2011) from SGS89
Figure 11-5 HARD plot of all coarse rejects (2012) from SGS89
Figure 11-6 HARD plot of all coarse rejects (2013) from SGS90
Figure 11-7 HARD plot of all Surface Drilling coarse rejects (2014) from SGS91
Figure 11-8 HARD plot of all Surface Drilling coarse rejects (2015) from SGS92
Figure 11-9 HARD plot of all Surface Drilling coarse rejects (2016) from SGS92
Figure 11-10 HARD plot of all Surface Drilling coarse rejects (2017) from SGS and Intertek93
Figure 11-11 HARD plot of all Surface Drilling coarse rejects (2018) from Intertek93
Figure 11-12 HARD plot of all Surface Drilling coarse rejects (2019) from Intertek94
Figure 11-13 HARD plot of all Surface Drilling coarse rejects (2020 Jan-Aug) from Intertek94
Figure 11-14 HARD plot of all Surface Drilling coarse rejects for 2018-19 for Father Brown & Adoikrom, from Intertek95
Figure 11-15 HARD plot of 2018 Wassa duplicate analysis (Intertek vs SGS)103
Figure 11-16 Wassa duplicates correlation plot (Intertek vs SGS)103
Figure 13-1 West view of metallurgical sample locations (GSR, 2015)108
Figure 13-2 Comparative indicated deportment of gold from diagnostic leach results112
Figure 13-3 Variation of UCS and CWi results with depth (mRL)113
Figure 13-4 2015 Ball Mill Bond Work Index against sample depth (mRL)114
Figure 13-5 2015 Abrasion Index against sample depth (mRL)114
Figure 13-6 Leach recovery kinetic curves116
Figure 14-1 Wassa long-range (grey) and short-range (cyan) Mineral Resource estimation model limits119

 

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Figure 14-2 Wassa LR model structural ‘Form’ surfaces (oblique view looking N up plunge), surfaces show deposit scale F4 fold as well as rolling over of mineralization at depth124
Figure 14-3 North-facing cross sections showing structural form (18950 mN and 19170 mN)125
Figure 14-4 Structural form surfaces used in the SR model125
Figure 14-5 Images showing the structural control surfaces on sections 19,750 mN and 19,635 mN.  The images show the longer, LR model defined control surfaces and the shorter, mine geology defined control surfaces126
Figure 14-6 Short-range isoshell modelling parameters (SRK, 2020)127
Figure 14-7 SE Isometric view of final LR model Leapfrog Isoshells (blue = >0.4 g/t, red = >1.5 g/t)128
Figure 14-8 Long section looking East showing the mineralized and halo domain shells on 39,940E (top image).  Section on 39,940E (lower image) displaying the same data, cut by that section line, and the assay data used to create the domain shells.  RED = mineralized domain; BLUE = halo domain129
Figure 14-9 Model section at U=-28.0 in transformed space generated with 2.0 tolerance in V direction130
Figure 14-10 Mineral Resource wireframes and drill hole locations for the Benso deposits (GSR, 2010)131
Figure 14-11 Mineral Resource wireframes and drillhole locations for Chichiwelli (GSR, 2008)132
Figure 14-12 Probability Plot for LG (left) and HG (right) Domains North of 19400N (top row) and South of 19400N (bottom row) (SRK, 2020)133
Figure 14-13 Histogram showing the uncapped 2m Au composite grade distribution for the mineralized domain134
Figure 14-14 Histogram showing the uncapped 2m composite grade distribution for the halo domain135
Figure 14-15 HG Variogram from anchor point 1 (SRK, 2020)137
Figure 14-16 LG Variogram from anchor point 14 (SRK, 2020)137
Figure 14-17 Variogram for the short-range HG & LG mineralized domains (SRK, 2020)138
Figure 14-18 Gold grade probability plot with outliers and far out thresholds highlighted (RMS, 2020)139
Figure 14-19 Inferred nugget effect for gold grade in each vein unit for FBZ deposit (RMS, 2020)140
Figure 14-20 Fitted experimental variogram points for gold grade in FW for FBZ deposit (RMS, 2020)141
Figure 14-21 Fitted experimental variogram points for gold grade in HG for FBZ deposit (RMS, 2020)142
Figure 14-22 Fitted experimental variogram points for gold grade in HW for FBZ deposit (RMS, 2020)143
Figure 14-23 Inferred nugget effect for gold grade in each vein unit for ADK deposit (RMS, 2020)144
Figure 14-24 Fitted experimental variogram points for gold grade in FW for ADK deposit (RMS, 2020)145
Figure 14-25 Fitted experimental variogram points for gold grade in HG for ADK deposit (RMS, 2020)146
Figure 14-26 Fitted experimental variogram points for gold grade in HW for ADK deposit (RMS, 2020)147
Figure 14-27 South-North Swath Plot Comparing Estimated Grades and Informing Capped Composites (SRK, 2020)157
Figure 14-28 Quantile-Quantile Comparison of Block Model Grades to Declustered Change-of-Support Corrected Gold Distributions for LG (left) and HG (right) domains (SRK, 2020)158
Figure 14-29 SWATH plot in the E-W direction X dimension.  Blue line represents Block model Estimated grades and Red is 2m drill hole composites grades (SRK, 2020)159
Figure 14-30 SWATH plot in the N-S direction Y dimension.  Blue line represents Block model Estimated grades and Red is 2m drill hole composites grades (SRK, 2020)159
Figure 14-31 SWATH plot in the elevation direction Z dimension.  Blue line represents Block model Estimated grades and Red is 2m drill hole composites grades (SRK, 2020)160
Figure 14-32 Measured and estimated gold grades at data locations (RMS, 2020)161
Figure 14-33 Swath plot comparison of composites, nearest neighbor estimates and kriging estimates for uncapped and capped grades in HG for FBZ deposit (RMS, 2020)162
Figure 14-34 Swath plot comparison of composites, nearest neighbor estimates and kriging estimates for uncapped and capped grades in HG for ADK deposit (RMS, 2020)162
Figure 14-35 Wassa LR model Indicated Mineral Resource classification surface and solids.  All blocks above surface and within solid mesh were classified as Indicated Mineral Resources (GSR, 2021)166
Figure 14-36 Estimation metrics associated to Indicated (top) and Inferred (bottom) classified Resources (SRK, 2020)167
Figure 14-37 645m RL section showing resource classification, boundary solid and drill holes (GSR, 2020)169
Figure 14-38 Father Brown and Adoikrom Indicated Mineral Resource surface and Inferred Mineral Resource solids.  All material above Magenta surface was classified as Indicated Mineral Resources all material below surface and within cyan (ADK) and red (FBZ) 3D meshes was classified as Inferred Mineral Resource170
Figure 15-1 Wassa UG cut-off optimization174
Figure 16-1 Mineral Resources considered in Mineral Reserve and models applied177
Figure 16-2 Schematic of Wassa location descriptors178
Figure 16-3 Wassa mine design and asbuilt, plan view (GSR, 2021)179

 

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Figure 16-4 Wassa mine design and asbuilt, longitudinal view180
Figure 16-5 Wassa underground production history180
Figure 16-6 Stope cycle for Panel 2 primary stopes182
Figure 16-7 Panel 2 primary/secondary stope extraction sequence, transverse stopes182
Figure 16-8 Stope cycle for Panel 2 secondary stopes183
Figure 16-9 Oblique view of Wassa Panels 1-3, asbuilt and planned development184
Figure 16-10 Typical level layout, Panels 1-2 570 mRL185
Figure 16-11 Oblique view of Panels 3 242 Area, planned development and stopes185
Figure 16-12 Oblique view of Panels 3 B-Shoot Area, planned development and stopes186
Figure 16-13 Stereonet plant of Wassa joint set database187
Figure 16-14 Principal stress measurement Magnitude vs Depth189
Figure 16-15 Support, Barton’s Q-Index chart (Barton and Grimstad, 1993)190
Figure 16-16 Stope axes measurements192
Figure 16-17 Matthews Stability Graph, transverse stopes (Mathews et al, 1981)193
Figure 16-18 Matthews Stability Graph, longitudinal stopes (Mathews et al, 1981)193
Figure 16-19 B-Shoot Pillars, modelled factors of safety from Phase 2 software, (GSR, 2018)194
Figure 16-20 Wassa paste plant Dec-2020, thickener and storage tank in foreground196
Figure 16-21 Paste fill distribution modelling197
Figure 16-22 Wassa Panels 1 and 2 ventilation circuit to end of life199
Figure 16-23 Wassa Panel 3, B-Shoot Upper ventilation circuit199
Figure 16-24 Wassa Panel 3, 242 ventilation circuit200
Figure 16-25 Underground dewatering longitudinal view201
Figure 16-26 620 mRL main pump station202
Figure 16-27 Lateral development schedule for Mineral Reserve204
Figure 16-28 Ore mining schedule for Mineral Reserve204
Figure 16-29 Underground Production History and Mineral Reserve plan205
Figure 17-1 Wassa processing plant flow sheet209
Figure 17-2 Processing schedule for Mineral Reserve plan211
Figure 17-3 Gold Production schedule for Mineral Reserve plan211
Figure 17-4 Processing Production History and Mineral Reserve Plan212
Figure 17-5 Gold Production History and Mineral Reserve plan212
Figure 18-1 Wassa key infrastructure (GSR, 2018)213
Figure 18-2 Wassa site layout (GSR, 2021)214
Figure 18-3 Wassa site layout and underground workings (GSR, 2021)215
Figure 18-4 Site electrical distribution216
Figure 18-5 Wassa Main pit catchments217
Figure 18-6 Tara Camp219
Figure 18-7 Waste dump locations (Golder, 2016)220
Figure 18-8 Section through nominal waste dump design220
Figure 18-9 Wassa TSF 1 and TSF 2 aerial view (August 2020)221
Figure 18-10 View from north of TSF 1 looking southeast (November 2020)222
Figure 18-11 TSF 1 and TSF 2 layout (Geosystems, 2018)224
Figure 20-1 Pra River basin and location of Wassa234
Figure 20-2 Wassa topography and drainage with sub-catchments235
Figure 20-3 Conceptual underground water flow path model (Golder, 2016)236
Figure 20-4 Conceptual groundwater model (Golder, 2016)237
Figure 20-5 Groundwater contours and flow (Golder, 2016)238
Figure 20-6 Conceptual geo-environmental model, E-W section (Golder, 2016)239
Figure 20-7 Paste pH vs NPR for Open Pit241
Figure 20-8 Paste pH vs NPR for Waste241
Figure 20-9 Paste pH vs NPR for Underground241
Figure 20-10 Paste pH vs NPR for South Inferred241
Figure 20-11 NPR vs S% for Open Pit241
Figure 20-12 NPR vs S% for Waste241
Figure 20-13 NPR vs S% for Underground242

 

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Figure 20-14 NPR vs S% for South Inferred242
Figure 20-15 GSOPP oil palm plantation on TSF 1247
Figure 22-1 Cash Flows by Year for Mineral Reserve – Base case261
Figure 22-2 Cash Flows by Year for Mineral Reserve – Consensus case263
Figure 22-3 Sensitivity analysis of the Mineral Reserve base case ($1,300 /oz)265
Figure 22-4 Sensitivity analysis of the Mineral Reserve consensus case (av.  $1,751 /oz)266
Figure 24-1 Illustration of Wassa location descriptors270
Figure 24-2 Mineral Resources considered in Southern Extension PEA (LR model only)270
Figure 24-3 Longitudinal Section looking east, showing the Southern Extension Panels271
Figure 24-4 Cross sectional view, Southern Extension, highlighting the width and complexity across the deposit272
Figure 24-5 Oblique view of the Southern Extension showing twin decline layout (looking northeast)273
Figure 24-6 Schematic of a 4-lift primary transverse stope (illustration not to scale)274
Figure 24-7 Generic downhole stope activity sequence275
Figure 24-8 Primary/secondary stope extraction sequence, transverse stopes276
Figure 24-9 Pillarless retreat stope extraction sequence, transverse stopes277
Figure 24-10 Generic wide-width mining (illustration not to scale)278
Figure 24-11 East decline, oblique looking north-west280
Figure 24-12 West decline, oblique view looking north-east281
Figure 24-13 Generic production block layout with primary/2ndary sequence and vent flows, longitudinal view283
Figure 24-14 Level layout (295 mRL) showing deposit width and twin decline arrangement, plan view284
Figure 24-15 Haulage level arrangement, 470 mRL285
Figure 24-16 Oblique view, approximate Resource development and infill drilling horizons, looking north west286
Figure 24-17 Geotechnical drill hole data in Southern Extension, plan (left) and longitudinal (right) views (OreTeck, 2020)288
Figure 24-18 Wassa preliminary principal stress gradient with reference mines and regions (OreTeck, 2020)289
Figure 24-19 Geotechnical rock mass model in Southern Extension, showing Q-prime in plan (left) and cross-section (OreTeck, 2020)290
Figure 24-20 Unsupported stable stope spans for expected rock mass conditions and current design hydraulic radii (Mathews, 1981; Potvin, 1988)292
Figure 24-21 Unsupported hanging-wall stable hydraulic radii for Southern Extension, longitudinal view (OreTeck, 2020)293
Figure 24-22 Wassa Panels 4-8, ventilation stages, oblique view295
Figure 24-23 Panels 4-8, production block ventilation flows296
Figure 24-24 Primary fan applied pressure (air density 1.1 kg/m3) and motor power (75% efficiency)297
Figure 24-25 Heat loads and cooling summary for Y12 (SRK, 2021)297
Figure 24-26 Estimated refrigeration capacity over mine life (SRK, 2021)298
Figure 24-27 Lateral development schedule for Southern Extension PEA300
Figure 24-28 ROM material mining schedule for Southern Extension PEA300
Figure 24-29 Electrical distribution with Phase 2 expansion required to support Southern Extension302
Figure 24-30 Metallurgical Sample Drillhole Location304
Figure 24-31   Ball Mill Bond Work Index against sample depth (mRL)305
Figure 24-32 Direct Leach – Grind Sensitivity Summary306
Figure 24-33 Preg-robbing Characterization Summary307
Figure 24-34 Processing schedule for Southern Extension PEA309
Figure 24-35 Gold Production schedule for Southern Extension PEA311
Figure 24-36 Cash Flows by Year for Southern Extension – Base Case318
Figure 24-37 Cash Flows by Year for Southern Extension – Consensus Case320
Figure 24-38 Sensitivity analysis of the Southern Extension PEA base case ($1,300 /oz)322
Figure 24-39 Sensitivity analysis of the Southern Extension PEA consensus case ($1,585 /oz)323
Figure 26-1 Project Execution Plan, Southern Extension Panels 4 and 5340

 

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LIST OF TABLES

Table 1-1 RGLD stream payment structure17
Table 1-2 Wassa Measured and Indicated Mineral Resource, as at 31 December 202020
Table 1-3 Wassa Inferred Mineral Resource, as at 31 December 202020
Table 1-4 Wassa Mineral Reserve, as at 31 December 202021
Table 1-5 Wassa, mine design quantities for Mineral Reserve plan22
Table 1-6 Stable stope dimensions for Mineral Reserve plan22
Table 1-7 Mining schedule quantities for Mineral Reserve plan23
Table 1-8 Mobile fleet purchase schedule for Mineral Reserve plan24
Table 1-9 Processing schedule quantities for Mineral Reserve plan25
Table 1-10 Capital Cost Summary for Mineral Reserve plan28
Table 1-11 Cost Estimate, Operating for Mineral Reserve plan29
Table 1-12 Operating Cost Summary for Mineral Reserve plan29
Table 1-13 Conversion of Inferred Mineral Resource to PEA inventory31
Table 2-1 Qualified persons and site visits39
Table 4-1 Mineral rights held by GSWL45
Table 5-1 Communities neighbouring Wassa Mine50
Table 6-1 Recent Production History, Wassa54
Table 7-1 Deformational history of the Ashanti Belt (Perrouty et al, 2012)60
Table 10-1 Exploration data used for Mineral Resource models81
Table 11-1 Summary of analytical quality control data from 2014 to early 201790
Table 11-2 CRM for 2003-2007 (TWL)96
Table 11-3 Geostats CRM for 2008-2012 (SGS)96
Table 11-4 Gannet CRM for 2008-2012 (SGS)97
Table 11-5 Gannet CRM for 2013 (SGS)97
Table 11-6 Gannet CRM for 2014-2017 (SGS)98
Table 11-7 Gannet CRM for 2014 to 2017 (Wassa Site Lab)98
Table 11-8 Gannet CRM for 2018 (Intertek)98
Table 11-9 Gannet CRM for 2019 Wassa UG (Intertek)99
Table 11-10 Gannet CRM for 2020 Jan-Oct, Wassa UG (Intertek)99
Table 11-11 Gannet CRM for 2019 Wassa surface drilling (Intertek)99
Table 11-12 Gannet CRM for 2018-2019 Father Brown/Adoikrom surface drilling (Intertek)99
Table 11-13 Blank sample summary statistics 2011 to Oct-2020100
Table 11-14 Blank sample summary statistics 2019, Wassa surface drilling (Intertek)100
Table 11-15 Blank sample summary statistics 2018-2019 Father Brown/Adoikrom surface drilling (Intertek)100
Table 11-16 Gannet CRM for quarter core sample analysis (Intertek)101
Table 11-17 Summary HARD plot results for quarter core sample analysis101
Table 11-18 Summary HARD plot results for 2013 round robin program101
Table 11-19 Round-robin descriptive statistics 2012102
Table 11-20 Round-robin descriptive statistics 2017102
Table 11-21 Summary HARD plot results for 2017 round robin program102
Table 11-22 Summary HARD plot results for 2018 round robin program103
Table 11-23 Specific gravity test results, open pit104
Table 11-24 Specific gravity test results, underground drilling 2017104
Table 11-25 Specific gravity test results, underground drilling 2018105
Table 11-26 Specific gravity test results, surface drilling 2018105
Table 11-27 Specific gravity test results, surface drilling 2020105
Table 13-1 Ore zones represented by the variability samples107
Table 13-2 Summary and location of test work samples108
Table 13-3 Screened head assay results109
Table 13-4 Elemental and chemical analysis results110
Table 13-5 Summary of diagnostic leach results111
Table 13-6 Results of Crushability Tests: UCS and CWi112
Table 13-7 Results of 2015 BWi and Ai Tests113

 

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Table 13-8 Gravity Gold Recovery Test Results115
Table 13-9 Whole Ore Leach and CIL test results115
Table 13-10 Leach test results and reagent consumptions116
Table 13-11 Overall gravity leach recoveries117
Table 13-12 Reconciliation of assay and back-calculated head grades from test work117
Table 13-13 Comparative settling test results118
Table 14-1 Wassa LR model drill hole database as at February, 2020121
Table 14-2 Wassa Underground short-range drill hole database as of December 1, 2020122
Table 14-3 Father Brown/Adoikrom drill hole database as of December 2020122
Table 14-4 Benso drill hole database as of December 2012122
Table 14-5 Chichiwelli drill hole database as of 2012122
Table 14-6 Leapfrog trend inputs for creation of 1.5 g/t and 0.4 g/t LR model grade Isoshells124
Table 14-7 Leapfrog trend inputs for creation of 1.5 g/t and 0.4 g/t SR model grade Isoshells126
Table 14-8 LR modelling extents126
Table 14-9 LR Isoshell modelling parameters126
Table 14-10 SR block model extents127
Table 14-11 Summary Gold Statistics of Assays and Composites132
Table 14-12 Comparison of Uncapped and Capped Gold Composite Grades – LR model133
Table 14-13 Comparison of uncapped and capped gold composite grades – SR model134
Table 14-14 Local variogram orientations and anchor point locations136
Table 14-15 Local variogram models by domain136
Table 14-16 Descriptive statistics for Hwini Butre modelled domains (uncapped & capped)139
Table 14-17 Capping values selected from analysis of the probability plot140
Table 14-18 Fitted variogram parameters for gold grade in FW for FBZ deposit144
Table 14-19 Fitted variogram parameters for gold grade in HG for FBZ deposit144
Table 14-20 Fitted variogram parameters for gold grade in HW for FBZ deposit144
Table 14-21 Fitted variogram parameters for gold grade in FW for ADK deposit148
Table 14-22 Fitted variogram parameters for gold grade in HG for ADK deposit148
Table 14-23 Fitted variogram parameters for gold grade in HW for ADK deposit148
Table 14-24 Fitted major variogram directions in original space148
Table 14-25 Descriptive statistics for Benso modelled domains (capped)149
Table 14-26 Descriptive statistics for simplified Benso modelled domains (capped)149
Table 14-27 Variogram parameters for the Benso zones150
Table 14-28 Descriptive statistics for Chichiwelli modelled domains (capped)150
Table 14-29 Chichiwelli high grade capping150
Table 14-30 Variogram parameters for Chichiwelli zones151
Table 14-31 Wassa LR model definitions, upper left hand corner coordinates151
Table 14-32 Average Bulk Density used for LR model151
Table 14-33 Wassa SR model definitions152
Table 14-34 LR model Estimation Parameters153
Table 14-35 SR model estimation parameters153
Table 14-36 Kriging search parameters for each vein unit in each deposit153
Table 14-37 Father Brown block model parameters154
Table 14-38 Adoikrom Zone block model parameters154
Table 14-39 Hwini Butre rock density154
Table 14-40 Benso block model parameters154
Table 14-41 Benso ellipsoid search neighbourhood parameters155
Table 14-42 Benso rock density155
Table 14-43 Chichiwelli block model parameters155
Table 14-44 Chichiwelli ellipsoid search neighbourhood parameters156
Table 14-45 Chichiwelli rock density156
Table 14-46 Global mean comparison between nearest neighbor and kriged thickness models for FBZ.160
Table 14-47 Global mean comparison between nearest neighbor and kriged thickness models for ADK.  Variable160
Table 14-48 List of least reliable estimates FBZ161
Table 14-49 List of least reliable estimates ADK161

 

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Table 14-50 Global mean comparison between nearest neighbor and kriged Gold models for FBZ162
Table 14-51 Global mean comparison between nearest neighbor and kriged Gold models for ADK163
Table 14-52 Global mean comparison between nearest neighbor and kriged Gold models for FBZ within the densely sampled area163
Table 14-53 Global mean comparison between nearest neighbor and kriged Gold models for ADK within the densely sampled area163
Table 14-54 Global mean comparison between nearest neighbor and kriged Gold models for FBZ within the sparsely sampled area164
Table 14-55 Global mean comparison between nearest neighbor and kriged Gold models for ADK within the sparsely sampled area164
Table 14-56 Composition of Classified Blocks for Open Pit Extraction Above a Cut-Off Grade of 0.4 g/t Gold167
Table 14-57 Composition of Classified Blocks for Underground Extraction Above a Cut-Off Grade of  2.1 g/t Gold168
Table 14-58 Wassa Measured and Indicated Mineral Resource, as at 31 December 2020172
Table 14-59 Wassa Inferred Mineral Resource, as at 31 December 2020172
Table 15-1 Wassa UG cut-off grade calculation174
Table 15-2 Wassa Mineral Reserve, as at 31 December 2020175
Table 16-1 Upper mine inventory change, OP to UG181
Table 16-2 Wassa Panels 1-3, design quantities for Mineral Reserve186
Table 16-3 Joint sets used for stope design188
Table 16-4 570 decline stress measurement188
Table 16-5 Wassa rock mass characterization parameters (Barton et al, 1974)189
Table 16-6 Modified Stability Number (N’) for Panels 1-3, transverse stopes (Potvin, 1988)191
Table 16-7 Modified Stability Number (N’) for Panels 1-3, longitudinal stopes (Potvin, 1988)191
Table 16-8 Stable stope dimensions, Panels 1-3192
Table 16-9 Wassa ventilation model calibration, Dec-2020198
Table 16-10 Wassa mining schedule quantities for Mineral Reserve plan203
Table 16-11 Mobile fleet productivity assumption206
Table 16-12 Mobile fleet schedule for Mineral Reserve plan206
Table 17-1 Historic plant production, grades and recoveries207
Table 17-2 Key plant design and operating parameters208
Table 17-3 Processing schedule quantities for Mineral Reserve plan210
Table 18-1 TSF 2 stage design details226
Table 20-1 Primary environmental approvals for mines in Ghana229
Table 20-2 Environmental approvals obtained for Wassa mine231
Table 20-3 Communities around Wassa245
Table 20-4 Closure cost estimates, at Dec-2020253
Table 21-1 Cost estimate, Major Projects for Mineral Reserve plan254
Table 21-2 Mine development capital allocation for Mineral Reserve plan255
Table 21-3 Cost estimate, Minor Projects for Mineral Reserve plan255
Table 21-4 Mobile Fleet, categories256
Table 21-5 Cost estimate, Mobile Fleet replacement schedule for Mineral Reserve plan256
Table 21-6 Capital cost summary for Mineral Reserve plan257
Table 21-7 Cost estimate, Operating for Mineral Reserve plan258
Table 21-8 Operating cost summary for Mineral Reserve plan259
Table 21-9 Closure cost summary for Mineral Reserve plan259
Table 22-1 Key life of mine inputs and assumptions used in the economic model for Mineral Reserve260
Table 22-2 Cash flows, Mineral Reserve economic analysis – Base case261
Table 22-3 Mineral Reserve economic analysis – Base case262
Table 22-4 Cash flows, Mineral Reserve economic analysis – Consensus case263
Table 22-5 Mineral Reserve economic analysis – Consensus case264
Table 22-6 Sensitivity results for the Mineral Reserve at different gold prices and discount rates265
Table 22-7 Sensitivity results of the Mineral Reserve base case ($1,300 /oz)265
Table 22-8 Sensitivity results of the Mineral Reserve consensus case (av.  $1,751 /oz)266
Table 24-1 Stope Modifying Factors contained in the MSO settings279
Table 24-2 Conversion of Inferred Mineral Resource to PEA inventory279

 

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Table 24-3 Proposed Resource diamond drilling quantities relative to stope timing by Panel286
Table 24-4 Wassa Panels 4-8 mine design quantities287
Table 24-5 Modified Stability Number (N’) for Panels 4 and 5, transverse stopes (after Potvin, 1988)291
Table 24-6 Modified Stability Number (N’) for Panels 4 and 5, longitudinal stopes (after Potvin, 1988)291
Table 24-7 Stable stope dimensions, Panels 4 and 5292
Table 24-8 Wassa mining schedule quantities, Southern Extension PEA299
Table 24-9 Mobile fleet schedule, Southern Extension PEA301
Table 24-10 Metallurgical Composite Sample Location303
Table 24-11 Metallurgical Composite Head Assay304
Table 24-12 Bond Ball Work Index Results305
Table 24-13 Gravity Recovery Gold - Summary306
Table 24-14 Reagent Consumption Summary307
Table 24-15 Diagnostic Leach Summary308
Table 24-16 Processing Schedule, Southern Extension PEA310
Table 24-17 Cost estimate, Major Projects for Southern Extension PEA312
Table 24-18 Mine development capital allocation for Southern Extension PEA312
Table 24-19 Cost estimate, Minor Projects for Southern Extension PEA313
Table 24-20 Cost estimate, Mobile Fleet addition/replacement schedule for Southern Extension PEA313
Table 24-21 Capital cost summary for Southern Extension PEA314
Table 24-22 Cost estimate, Operating for Southern Extension PEA315
Table 24-23 Operating cost summary for Southern Extension PEA315
Table 24-24 Closure cost summary for Southern Extension PEA316
Table 24-25 Key life of mine inputs and assumptions used in the economic model, PEA317
Table 24-26 Cash flows, PEA economic analysis – Base Case318
Table 24-27 Economic Analysis – Base Case319
Table 24-28 Cash flows, PEA economic analysis – Consensus Case320
Table 24-29 PEA Economic Analysis – Consensus Case321
Table 24-30 Sensitivity results for the Southern Extension PEA at different gold prices and discount rates322
Table 24-31 Sensitivity results of the Southern Extension PEA base case ($1,300 /oz)322
Table 24-32 Sensitivity results of the Southern Extension PEA consensus case ($1,585 /oz)323
Table 24-33 Geology, Mining and Processing risks for Southern Extension325
Table 24-34 Economic risks for Southern Extension326
Table 24-35 Geological Drilling opportunities for Southern Extension326
Table 24-36 Mine design and productivity opportunities for Southern Extension327
Table 25-1 Geology, Mining and Processing risks for the Mineral Reserve332
Table 25-2 Infrastructure risks for the Mineral Reserve332
Table 25-3 Capital and Operating Cost risks for the Mineral Reserve333
Table 25-4 Environmental and Social Risks333
Table 25-5 Mineral Resource Opportunities334
Table 25-6 Mine design and productivity opportunities for the Mineral Reserve334
Table 25-7 Sustainability Opportunities335
Table 26-1 Cost estimate for 2021 – 2022 Resource definition drilling and technical studies339


 

 

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LIST OF ABBREVIATIONS

2020 PEAPreliminary Economic Assessment of potential expansion of the underground mine to extract the Inferred Mineral Resource in the Southern Extension zone
AASAtomic adsorption spectroscopy (sampling)
ACAir-core (drilling)
ADKAdoikrom (deposit)
AiBond abrasion index (metallurgical testing)
ALSALS Minerals
ARDAcid rock drainage
AROAsset retirement obligations (closure planning)
BDGBD Goldfields (company)
BLEGBulk leach extractable gold (assaying)
BWiBond ball mill work index (metallurgical testing)
CILCarbon in leach (processing method)
CIMCanadian Institute of Mining, Metallurgy and Petroleum
CMCCCommunity Mine Consultative Committee
CRMCertified reference material (sampling QA/QC)
CSLCompacted soil liner (civil construction)
CWiBond low impact crushing work index (metallurgical testing)
CYAPCommunity Youth Apprenticeship Program
DDDiamond core (drilling)
EIAEnvironmental Impact Assessment
EISEnvironmental Impact Statement
EMPEnvironmental Management Plan
EMSEnvironmental and social management system
EPAEnvironmental Protection Agency (Ghana)
ESRExcavation support ratio (geotechnical)
FBZFather Brown (deposit)
FOSFactor of safety
FSFeasibility study
HWFootwall
G&AGeneral and administration
GAIGeochemical abundance index (geochemistry)
GCGrade control
GSIGeological strength index (geotechnical)
GSOPPGolden Star Oil Palm Plantation
GSRGolden Star Resources
GSSTEPGolden Star Skills Training and Employability Program
GSWLGolden Star Wassa Limited
HARDHalf absolute relative difference (statistics)
HBBHwini Butre Benso (deposit group)
HBMHwini Butre Minerals (company)
HGHigh grade
LGLow grade
HLHeap leach (processing method)
HWHanging-wall

 

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ICOLDInternational Committee on Large Dams
ILRIn-line reactor (processing method)
IPInduced polarization
JnJoint number (geotechnical)
JrJoint roughness (geotechnical)
JwJoint alteration (geotechnical)
L.I.Legal Instrument
LGLow grade
LHOSLong hole open stoping (mining method)
LR (model)Long-range model (geological modelling)
LVALocally variable anisotropy (geological modelling)
MOUMemoranda of Understanding
MSGModified Stability Graph (geotechnical)
MSOMineable Stope Optimiser (mine planning)
NAGNot acid generating (geochemistry)
NPVNet present value
OKOrdinary kriging (geological modelling)
PCPPractical closure plan (closure planning)
PVCPoly-vinyl chloride
QA/QCQuality assurance, quality control
QPQualified person
RABRotary air blast (drilling)
RCReverse circulation (drilling)
RGIRyal Gold Inc (company)
RGLDRGLD Gold AG (company)
RLRelative level
RMRRock mass rating (geotechnical)
RMSResource Modelling solutions (company)
ROMRun of mine
RPEEEReasonable prospects for economic extraction
RQDRock quality description (geotechnical)
SGLSatellite Goldfields Limited (company)
SJRSaint Jude Resources (company)
SR (model)Short-range model (geological modelling)
TSFTailings storage facility
UCSUnconfined compressive strength
US$United States dollar/s
VRAVolta River Authority (Ghana)
WSLWassa site laboratory (assaying)
WUCWestern University College, Tarkwa (Ghana)
WUGWassa underground mine
XRDX-ray diffraction
XRFX-ray fluorescence

 

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NI 43-101 Technical Report (March 2021)Wassa Gold Mine
1Executive Summary
1.1Terms of Reference

This Technical Report has been prepared to meet the requirements defined by Form 43-101F1, by and for Golden Star Resources, describing the Wassa gold mine in Ghana. GSR owns a 90% interest in and manages Golden Star Wassa Limited, who’s primary asset is the Wassa gold mine, with the Government of Ghana owning the remaining 10%.

The report provides updated information on the currently operating Wassa Gold Mine:

·Updated Mineral Resource and Mineral Reserve estimates, as at 31 December 2020; and
·Summary of a Preliminary Economic Assessment of potential expansion of the underground mine to extract the Inferred Mineral Resource in the Southern Extension zone (2020 PEA).

The 2020 PEA has no impact on the Mineral Reserves, nor on the key assumptions and parameters supporting the Mineral Reserves, which are current, valid and do not rely on any assumptions in the PEA.

The PEA is a scoping level study which is conceptual in nature and there is no certainty that production and financial outcomes will be realized. It has been prepared within the following framework:

·Production schedules to appropriately consider conversion risk of the Inferred Mineral Resource;
·Methodologies and design quantities based on proven, currently available technologies;
·Mine production constrained within current processing capacity (2.7 Mtpa);
·Costs to reflect current operational experience; and
·Minimise capital demand needed to establish full production.

The intent of the framework is to present a deliverable PEA plan which can be executed with GSR’s current operational and financing capacity. Potential enhancements outside this framework are presented as opportunities outside of the PEA outcomes and can be investigated as part of the forward work plan.

Mineral Resources and Mineral Reserves have been prepared in accordance with the 2014 CIM Definition Standards and 2019 Best Practice Guideline.

Units in the report are metric and monetary units are United States dollars (US$) unless otherwise stated.

1.2Location and Setting

The Wassa Mine is located in a rural setting near the village of Akyempim in the Wassa East District, in Ghana’s Western Region, 80 km north of Cape Coast and 150 km west of the capital Accra.

The climate is classified as wet semi-equatorial with a dry season from November to February. The wettest month is June with an average 241 ± 85 mm and annual average rainfall is 1,996 ± 293 mm.

1.3Mineral Tenure, Permits, Royalties and Agreements

GSWL holds three mining leases (Wassa, Hwini Butre, Benso) and several prospecting leases in the region and full details are presented in Table 4-1.

EIA studies have been undertaken to support permitting and there is considerable background environmental data.

The Wassa Mining Lease stipulates a 5% royalty on gross revenue, paid quarterly to the Government of Ghana and the government holds a 10% free-carried interest, which entitles a pro-rata share of dividends.

GSR is party to a gold purchase and sale agreement with Royal Gold, Inc. through its wholly owned subsidiary RGLD Gold AG (RGLD) which was amended and restated on 30 September 2020. The stream covers all gold produced within GSWL’s mineral concessions. The stream payment structure is outlined in Table 1-1. The stream payments are treated as a revenue adjustment.

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Table 1-1 RGLD stream payment structure

Stream TierAttributable Ounces% of spot paidApplication
Tier 110.5%, until 240,000 oz reached20%119,997 oz sold to end-December 2020
Tier 25.5%30%All production after Tier 1 completed
1.4History

Golden Star acquired the 90% share in Wassa in September 2002 from Standard Bank after the foreclosure of Glencar Mining’s share of Satellite Gold Limited. At the time Wassa was an open pit operation treating 3.0 Mtpa through heap leach, with gold recovery of 55-60%.

The carbon in leach plant was commissioned in 2005 and upgraded in 2013 to treat 2.7 Mtpa of fresh rock feed only. Ore has been mined mostly by open pit, with underground development commencing in 2015 and forming the majority of the ore supply since 2018. Gold production has varied from 184 koz in 2009 to 104 koz in 2016 and has averaged 157 koz/yr for the past three years (2018-2020).

1.5Geology and Mineralization
1.5.1Wassa

The Wassa property lies within the southern portion of the Ashanti Greenstone Belt along the eastern margin and within a volcano-sedimentary assemblage located close to the Tarkwaian basin contact.

Wassa lithology is characterized by lithologies of the Sefwi Group, consisting of intercalated meta-mafic volcanic and meta-diorite dykes with altered meta-mafic volcanic and meta-sediments which are locally characterized as magnetite rich, banded iron formation like horizons (Bourassa, 2003). The sequence is characterized by the presence of multiple ankerite-quartz veins, sub-parallel to the main penetrative foliation and Eoeburnean felsic porphyry intrusions on the south-eastern flank of the Wassa mine fold.

Wassa mineralization is subdivided into a number of domains: F Shoot, B Shoot, 242, South East, Starter, 419, Mid East, and Dead Man’s Hill. Each of these represents discontinuous segments of the main mineralized system. The South- Akyempim deposits are located 2 km southwest of the Wassa Main deposit on the northern end of a mineralized trend parallel to the Wassa Main trend.

Mineralization is hosted in highly altered multi-phased greenstone-hosted quartz-carbonate veins interlaced with sedimentary pelitic units. It is structurally controlled and related to vein densities and sulphide contents.

Wassa mineralization is quite old and has been affected by several phases of deformation since emplacement. Two major folding events were likely emplaced early in the deposit’s deformational history, with Gold mineralization later remobilized into the hinges of a tight folding event and finally, the deposit scale fold which influences the open pit configuration.

Remobilized gold in the hinges of the tight folding event are the primary underground mining targets and B-Shoot and F-Shoot are the two main zones. These zones plunge to the south at approximately 20 degrees with Mineral Resource now defined more than 800 m south of the current underground Mineral Reserve.

1.5.2Regional Deposits

The Hwini Butre concession is underlain by three deposits: Adoikrom, Dabokrom and Father Brown, which are all characterized by different styles of mineralization within the Mpohor mafic complex, which consists mainly of gabbroic and gabbro-dioritic intrusive horizons.

At Father Brown and Dabokrom, mineralization is associated with quartz vein systems that are locally surrounded by extensive, lower grade, disseminated quartz stockwork bodies, especially at Dabokrom. At Adoikrom, the mineralization is shear hosted and characterized by the absence of quartz veins; gold is associated with fine grained pyrite and intense potassic alteration.

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The Benso concession is underlain by four main deposits: Subriso East, Subriso West, G Zone and I Zone which all have a similar style of mineralization.

The Benso deposits are hosted in two dominant rock types, Subriso West and I Zone are hosted within Intermediate feldspar porphyry intrusives and meta-volcanics, whereas Subriso East occurs along the contact between carbonaceous phyllites and meta volcanics. Mineralization at Benso is associated with late deformational stages of the Eburnean orogeny and deposits are shear hosted along subsidiary structures.

The Chichiwelli deposit consists of two sub-parallel mineralized trends hosting two distinct types of mineralization. The Chichiwelli West trend is a shear zone hosted deposit with a quartz, carbonate, sericite and potassic alteration assemblage, the mineralization is associated with pyrite. The Chichiwelli East trend is a quartz vein associated deposit with an ankerite and sericite alteration assemblage. Mineralization is also associated with pyrite along vein selvages and in the wall rocks.

1.6Drilling and Sampling

Drilling is carried out by diamond core (DD), reverse circulation (RC) and RAB/air-core techniques. Surveys are conducted on drill hole collars (by total station) and downhole (by either multi-shot downhole camera, or gyro instrument for deeper holes).

A standardized approach to drilling and sampling is applied where typically, sampling is carried out along the entire mineralized drilled length. Sampled spacing is 1.0 m for RC and for DD samples according to mineralization, alteration or lithology. Core is split into equal parts along a median to the foliation plane to ensure representative samples for assaying. The remaining half core is retained for reference and additional sampling if required.

Sample preparation on site is restricted to core logging and cutting, or RC and RAB sample splitting. Facilities consist of enclosed core and coarse reject storage facilities, covered logging sheds and areas for the splitting of RC and RAB samples (with Jones riffle splitter).

From site, samples are transported by road to the primary laboratory in Tarkwa for sample preparation and chemical analysis. Sample security involves:

·Chain of custody of samples to prevent inadvertent contamination or mixing of samples; and
·Rendering active tampering of samples to be as difficult as practicable.

As the samples are loaded, they are checked and the sample numbers are validated. The sample dispatch forms are signed off by the transport driver (dispatched by the laboratory) and a GSR representative. Sample dispatch dates are recorded in the sample database as well as the date when results are received.

Sample assays are performed at either the Wassa Site Lab, SGS or Intertek (formerly TWL), with both independent labs located in Tarkwa. GSR submits quality control samples to each lab for testing purposes. Both SGS and Intertek laboratories are independent of GSR and are accredited for international certification for testing and analysis.

1.7Data Verification

Core logging and sampling procedures adopted by GSR are consistent with industry standards and validated by external consultants checking logging against the remaining half-core with no major errors identified.

Procedures are in place with several steps to verify the collection of drill hole data and minimize potential for data entry errors. Data entry and database management involves logging of holes directly into an SQL Acquire database via laptop computers linked to the main database, with built-in validation tools designed to eliminate erroneous data entry.

Analytical data is checked for consistency by GSR personnel:

·Upon receipt of digital assay certificates; the assay results, along with the control sample values, are extracted from the certificates and imported into the Acquire database;
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·Failures and potential failures are examined and depending on the nature of the failure, re-assaying is requested from the primary laboratory; and
·Analysis of quality control data is documented, along with relevant comments or actions undertaken to either investigate or mitigate problematic control samples.

GSR relies on the laboratory operators’ QA/QC processes for assaying, as well as GSR’s own independent QA/QC program. The GSR program includes inserting blanks, certified reference materials (aka: standards) and pulp or coarse reject duplicates into sample batches before sample lab submission. QA/QC procedures required >=10 % of the samples submitted to the independent laboratories are check samples.

1.8Metallurgical Test Work

Metallurgical test work for underground ore was completed as part of the 2015 underground feasibility study. The test work program showed general consistency with ores previously treated from the open pits. This has subsequently been confirmed by actual processing results where, since 2015, overall plant recoveries between 94-96% have been achieved.

1.9Mineral Resource Estimate

The MRE has been updated for the on the basis of RC and DD drilling at four properties – Wassa, Hwini Butre, Benso and Chichwelli. At Wassa, 30,067 drill holes for 1,353,740 m used to estimate a Long-Range Model and 2,875 drill holes for 544,233 to estimate a short-range model (SR model), limited to the mine area. The Father Brown/Adoikrom MRE has been updated with 3,736 drill holes for 154,589 m. The Benso MRE was updated using 3,162 drill holes for 130,506 m and Chichiwelli used 506 drill holes for 33,494 m. Data was validated prior to use in estimation and is considered fit for the purposes of Mineral Resource estimation by the Qualified Person (QP).

Mineralization was modelled via grade shells at Wassa. The long-range (LR) model used indicator methodology and the SR model used the raw assay data to interpolate grade shells. The halo domain was interpreted at 0.4 g/t Au and the high grade/mineralization domain was interpreted at 1.5 g/t Au. Both cut offs were determined by visual inspection and separated different grade populations in the data. Structural trend surfaces informed the mineralization interpolation and orientation of search ellipses for both models.

At Father Brown and Adoikrom, GSR and Resource Modelling Solutions (RMS) used a vein modelling technique, where the vein unit is modelled by estimating the position of the vein and thickness, hanging wall, high grade and footwall. At Benso, mineralization and oxidation wireframes were created by GSR. The mineralization zones of Benso are structurally controlled with gold emplacement related to the density of quartz veining and sulphide content. At Chichiwelli, mineralization wireframes were interpreted at a nominal 0.5 g/t Au cut off. Mineralization is structurally controlled with gold emplacement related to the density of quartz veining and sulphide content. The mineralization hosting structures generally trend north-south and dip moderate-steeply to the east at 60°.

Assay data was composited (3 m for Wassa LR model, 2 m for Wassa SR model, 2 m for Benso and Chichiwelli), top cut based on review of population disintegration via probability plots. Variograms were modelled where possible, to characterize the grade continuity in grade estimation.

Grades were estimated using ordinary kriging into parent cells for all deposits. At Wassa (long-range and short-range) using locally oriented search ellipses, based on structural trends. Three search passes with successively larger search ranges were used to estimate grades into blocks. The LR model had a block size of 10 x 10 x 5 m and the SR model had a block size of 1 x 2.5 x 2.5 m. In situ dry bulk density has been estimated to be 2.80 t/m3 based on measurements by GSR and Intertek in 2017 and 2018.

Block sizes at Father Brown and Adiokrom models were 1 x 2 x 2 m (X x Y x Z) chosen to reflect the geometry of the deposit. A density of 2.70 t/m3 was assigned. At Benso, the block size chosen was 12.5 x 25 x 10 m (X x Y x Z) to reflect the average spacing of drill fences along strike. Grade was estimated into mineralization domains with soft boundaries between oxidation units, using four search passes. A density of 1.80 t/m3 for oxide and 2.70 t/m3 for fresh was assigned. A block size of 12.5 x 25 x 8 m (X x Y x Z) was chosen for Chichiwelli, grades were estimated in four search passes and a density of 1.80 t/m3 was assigned to oxide and 2.68 t/m3 to fresh.

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Validation was completed via inspection of swath plots, cross sections, mean grade comparisons between composites and blocks.

The basis of the Mineral Resource classification included confidence in the geological continuity of the mineralized structures, the quality and quantity of the exploration data supporting the estimates, confidence in the density measurements and the geostatistical confidence in the tonnage and grade estimates. Reasonable prospects for eventual economic extraction (RPEEE) was informed via pit shell optimization for open pit Mineral Resources (Benso, Chichiwelli, HBB Other) and cut off grade estimation for underground mineral resources (Wassa, Father Brown, Adoikrom). A gold price of US$1500 /oz and costs from the operations were used in pit shell optimization and cut off grade estimation.

·The Mineral Resource estimate complies with the requirements of National Instrument 43-101 and has been prepared and classified in accordance with the 2014 CIM Definition Standards and 2019 Best Practice Guidelines.
·Measured and Indicated Mineral Resources are reported inclusive of Mineral Reserves;
·Underground deposits within the Mineral Resource are reported at a cut-off grade of 1.4 g/t;
·Open pit deposits within the Mineral Resource are reported at a cut-off grade of 0.55 g/t, within optimized pit shells calculated at a $1,500 /oz gold selling price;
·The Mineral Resource models have been depleted using appropriate topographic surveys;
·Regional OP includes deposits at Benso, Chichiwelli and HBB Others;
·Mineral Resources are reported in-situ without modifying factors;
·No open pit resource has been reported for the Wassa deposit, as engineering studies have determined Wassa will be mined by underground methods only; and
·All figures are rounded to reflect the relative accuracy of the estimate.

Table 1-2 Wassa Measured and Indicated Mineral Resource, as at 31 December 2020

 Measured & Indicated Mineral Resource, at 31 December 2020

Meas. & Ind.
Mineral Resource
at 31 December 2019

 Measured ResourceIndicated ResourceMeas. & Ind.
Mineral Resource
 MtAu g/tkozMtAu g/tkozMtAu g/tkozMtAu g/tkoz
Wassa OP---------29.181.291,206
Wassa UG5.904.4584318.963.552,16224.853.763,00516.203.892,027
Father Brown
/Adoikrom UG
---1.317.963351.317.963350.918.67254
Regional OP---3.101.981973.101.981972.512.32187
TOTAL5.904.4584323.373.592,69429.263.763,53748.812.343,675

 

Table 1-3 Wassa Inferred Mineral Resource, as at 31 December 2020

 

Inferred Mineral Resource

at 31 December 2020

Inferred Mineral Resource

at 31 December 2019

 MtAu g/tkozMtAu g/tkoz
Wassa OP---0.621.3126
Wassa UG70.503.397,68958.823.757,097
Father Brown /Adoikrom UG2.665.304531.886.07367
Regional OP0.871.47410.422.1429
TOTAL74.023.448,18361.743.797,518

 

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1.10Mineral Reserve Estimate

The Mineral Reserve has been calculated with a cut-off grade of 1.9 g/t, which is below 2.4 g/t applied previously. The change has been driven by lower operating costs, achieved through increasing underground mining rates that have been sustained through 2019 and 2020 and validated by an assessment during 2020 which showed the optimal NPV for the Mineral Reserve will be achieved at a cut-off of 1.9 g/t and ore mining rate of 5,000 t/d (1.8 Mtpa).

Modifying factors are applied to stopes at 5.0% diluton and 96.1% recovery and 0.0% diluton and 100.0% recovery for development, which are based on actual performance to end-November 2020.

The Mineral Reserves have been prepared in accordance with the 2014 CIM Definition Standards and 2019 Best Practice Guidelines.

Material is included in the Mineral Reserve as follows:

·The Mineral Reserve estimate complies with the requirements of National Instrument 43-101 and has been prepared and classified in accordance with the 2014 CIM Definition Standards and 2019 Best Practice Guideline.
·The Mineral Reserve is reported at a cut-off grade of 1.9 g/t, calculated at a $1,300 /oz gold selling price;
·Modifying factors are applied as 5.0% dilution and 96.1% recovery for stopes;
·Material based on Measured Mineral Resources are reported as Proven Mineral Reserves;
·Material based on Indicated Mineral Resources are reported as Probable Mineral Reserves;
·Material based on Inferred Mineral Resources are excluded from Mineral Reserve;
·Economic analysis of the Mineral Reserve demonstrates economic viability at $1,300 /oz gold price; and
·All figures are rounded to reflect the relative accuracy of the estimate.

Table 1-4 Wassa Mineral Reserve, as at 31 December 2020

 Mineral Reserve, at 31 December 2020

Mineral Reserve

at 31 December 2019

 Proven ReserveProbable ReserveMineral Reserve
 MtAu g/tkozMtAu g/tkozMtAu g/tkozMtAu g/tkoz
UG, Panels 1 & 24.283.284514.482.994308.753.138817.423.72889
UG, Panel 3---2.062.941952.062.94195---
Open Pit---------9.921.57500
Stockpiles0.690.5813---0.690.58131.060.6221
TOTAL4.972.914646.542.9762511.502.941,08918.412.381,410
1.11Mining Methods
1.11.1Mine Design

Mining is by underground with trackless access by decline (1:7 gradient), operated by GSWL personnel. The method is long hole open stoping (LHOS), mostly with transverse uphole stopes, in 25 m lifts, mined top-down, in primary/secondary sequence. Stable unsupported stope heights are up to 100 m and stopes are mined to full orebody width.

The mine is divided into Panels, which reflect progressive stages of capital development. The Mineral Reserve includes Panels 1, 2 and 3, which lie between 150-650 m depth.

The December 2019 Mineral Reserve proposed to extract material below the B-Main and 242 pits by open pit methods. This has changed to underground extraction (Panel 3) based on trade off studies during 2020 which concluded underground extraction provided improved selectivity and reduced capital demand.

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Panels 1 and 2 contain Proven and Probable Mineral Reserve. Panel 3 contains only Probable Mineral Reserve, plus Inferred Mineral Resource which is excluded from the plan informing the Mineral Reserve.

Panels 1 and 3 leave 6-10 m intact rock pillars between stopes, with ad-hoc waste rock fill to enable pillar recovery or for simple waste disposal.

Panel 2 is planned for paste backfill. Primary stopes have been extracted, with voids awaiting commissioning of the paste fill plant in early 2021. Once primary stopes are filled, secondary stope extraction will commence. Stopes are up to 100 m high and are separated vertically by sill pillars which will be recovered after stopes above and below are extracted and filled.

Table 1-5 Wassa, mine design quantities for Mineral Reserve plan

  

Panels 1-2

B-Shoot

Panel 3

B/F-Shoot

Panel 3

242

Ore Mined, by Panel‘000 t8,7551,245818
 g/t3.132.713.30
 ‘000 oz88110987
Ore Mined, Total‘000 t 10,818 
 g/t 3.09 
 ‘000 oz 1,076 
Development, Totalm 44,173 
Dev’t Capitalm 20,392 
Dev’t Operatingm 23,781 
Vertical Developmentm 2,776 
Mined to Waste‘000 t 2,469 
Paste Backfill‘000 m3 2,967 

 

1.11.2Geotechnical

Geotechnical assessments have been completed and the rock mass quality is classified Very Good, using Barton’s (Barton et al, 1974) classification and Geological Strength Index (GSI) rating systems.

In-situ stress measurements have been taken and mining induced seismicity is not expected to have an impact until mining approaches 1,000 m depth (Mineral Reserve is down to 650 m depth).

Geotechnical design for development has been done using Barton’s Q support classification and development excavations lie within Category 1, “No Support Required”, although and minimum standard pattern is applied, consisting of friction bolts and mesh to the back and upper walls of all headings.

Geotechnical design for stopes has been done using the Stability Graph method to determine the stable stope design geometry and stope geometries contained in Table 1-6.

Table 1-6 Stable stope dimensions for Mineral Reserve plan

Stope DimensionTransverse StopeLongitudinal Stope
MINMAXDesign (m)MINMAXDesign (m)
Heightm25100100<152525
Strike Lengthm252525<607070
Width across Strikem153025<151515
Dip, end/side-walls 65°65°65°65°65°65°

 

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1.11.3Ventilation

The ventilation network for Panels 1 and 2 is planned to provide 50 m3/s per working area for up to 9 working areas with total airflow of 590 m3/s.

Total airflow is currently 440 m3/s and will be increased to the design flows with addition of two new 5.5-6.0 m diameter shafts (one each for intake and exhaust) which are budgeted in 2021 and 2022. Both shafts will be located at the southern end of Panel 2.

The Panel 3 B-Shoot area will connect to the Panel 1 -2 network and will enable increase airflow with an additional exhaust raise planned to be collared in the saddle between the Main and South-east pits.

The Panel 3 242 area will be ventilated with an independent network incorporating an exhaust drive mined from the Main pit connecting to the 242 ramp which will have a portal in the 242 pit. Up to 190 m3/s will flow through the 242 ramp with fans placed in the exhaust adit portal.

1.11.4Mining Schedule

Table 1-7 Mining schedule quantities for Mineral Reserve plan

 

 

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Figure 1-1 Underground Production History and Mineral Reserve plan

1.11.5Mobile Equipment

The current mobile equipment fleet will continue to be used and progressively move toward standardized machine types:

·Development Jumbos: continue current twin-boom machine, nominally Sandvik DD421;
·Production Drills: continue current 89-115 mm top-hammer, nominally Sandvik DL421;
·UG Loaders: current mixed fleet of 18 t units, progress toward 21 t, nominally Sandvik LH621; and
·UG Trucks: current fleet of 40 t units, progress toward 60 t machines, nominally Volvo A60H.

Table 1-8 Mobile fleet purchase schedule for Mineral Reserve plan

 

1.12Recovery Methods

The Wassa processing plant is a conventional CIL plant with a four stage crushing circuit (p80 <8 mm) feeding two independent ball mills of 3 MW each (p80 <75 μm).

The grinding circuit includes gravity recovery by Knelson concentrators and intensive leach in an Acacia reactor. Cyclone underflow has cyanide and oxygen added prior to pumping to leach circuit via the in-line reactor pipeline. The CIL circuit consists of six stages of agitated leach with an residence time of 18-20 h at full capacity.

Loaded carbon is acid washed and stripped and gold is electrowon onto steel mesh prior to smelting to produce doré bars.

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Table 1-9 Processing schedule quantities for Mineral Reserve plan

 

 

 

 

Figure 1-2 Processing Production History and Mineral Reserve plan

 

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Figure 1-3 Gold Production History and Mineral Reserve plan

1.13Infrastructure

Surface infrastructure to support the mining and processing operations is in place and includes:

·Access roads;
·Electrical power supply – access to grid, on-site generation and site distribution;
·Waste storage areas and open pit water storage areas, including water diversion structures;
·Main exhaust fans for underground ventilation;
·Processing facilities for processing up to 2.7 Mtpa;
·4,000 tpd paste plant, recently completed;
·2 Tailings Storage Facilities (TSF) – TSF 1 is being revegetated and TSF 2 is active;
·Maintenance workshops and site electrical distribution infrastructure;
·Site administration buildings; and
·Accommodation camp.

The mining and processing of the Mineral Reserve does not require any major upgrades to the site surface infrastructure, other than for the ventilation outlined in Section 0.

TSF design capacity exceeds requirements for the Mineral Reserve plan, without accounting for tails which will be used for paste backfill.

The design of TSF 2 meets the requirements of the Minerals and Mining (Health, Safety and Technical) Regulations, 2012 (L.I. 2182) and takes due consideration of the recommendations of the International Committee on Large Dams (ICOLD), the Australian Committee on Large Dams (1999) and the Canadian Dam Association guidelines (2007).

The minimum Factor of Safety (FOS) values calculated for all conditions on both the downstream and upstream slopes which were found to meet, and in some conditions exceed, the Minerals and Mining (Health, Safety and Technical) Regulations, 2012 (L.I. 2182) requirements for factors of safety.

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1.14Environmental Studies, Permitting and Social or Community Impact

There is a successful history of permitting, environmental and social risk management at Wassa

1.14.1Sustainability

GSR supports, is subject to and / or incorporates enhanced disclosure on a range of international requirements including:

·Human Rights, via the UN Sustainable Development Goals
·Anti-Corruption, through Ghana’s designation as Extractive Industries Transparency Initiative compliant
·Voluntary codes, including International Cyanide Management Code certification; Responsible Gold Standard and the Responsible Gold Mining Principles
·Resettlement, land acquisition and compensation, through conforming to the International Finance Corporations’ Performance Standard 5 on Land Acquisition and Involuntary Resettlement.

There are corporate assurance processes in place, including independent reviews, audit and/or validation to ensure conformance with the codes and standards.

1.14.2Environmental Considerations

The first Environmental Certificate for Wassa (and all other managed concessions) was received in September 2006 and is maintained in good standing with submission of and Environmental Management Plan (EMP) every three years, with the most recent Environmental Certificate issued in 2020.

The EMP is supported by:

·Surface water management, with an emphasis on diversion to minimize contact water.
·Hydrogeology, which concluded in 2016 and 2019 that expected drawdown will not have significant impact on community ground water boreholes.
·Geochemistry: analysis has consistently shown ore and waste lithologies, are generally not acid generating (NAG) which is validated by over two decades of mining.
·Water Quality:
oSurface water in vicinity of the open pits, conforms to the EPA Effluent Quality Guidelines;
oUnderground drainage studies indicate that water quality guidelines be met with discharges predicted to be generally neutral to alkaline with low concentrations of TDS, sulphate and metals. This is validated by routine sampling results.
·Air Quality is routinely monitored and an array of dust suppression mitigations are employed during dry season conditions.
·Noise is routinely monitored at neighbouring villages and results show noise emanations are not the result of Wassa activity.
·Vibration has been modelled and results conform to regulatory limits, which has been validated by monitoring results.
·Biodiversity surveys have been conducted for baseline establishment and measuring impacts.
oFlora: Species of conservation significance are actively propagated for revegetation use and areas identified to host quality unprotected remnant forest are specifically avoided;
oFauna: A 1996 study found no species of small mammal, bats, birds, herpetofauna, or amphibians of outstanding conservation merit. Several species of large mammal with conservation significance were located but observation required traversing over 10 km into a Forest Reserve to observe, likely due to high hunting pressures and impacts of logging.
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1.14.3Social Considerations

The main areas of focus for socio-economic improvement are: health, education, electricity and water supply, and livelihood opportunities. Social investment initiatives include:

·Golden Star Development Foundation: GSR’s main partnership vehicle to implement a variety of community development projects and programs.
·Golden Star Oil Palm Plantation: GSOPP is a community-based oil palm plantation company established in 2006 as a non-profit subsidiary of GSR.
·Capacity Building and Livelihood Enhancement: The Golden Star Skills Training and Employability Program (GSSTEP) provides skills training in non-mining sectors, increasing economic diversity.
·Local Procurement: The GSWL MOU on Local Employment and Contracts builds on the history of building local procurement capacity around Wassa.

Unauthorized small-scale mining (galamsey) occurs on and around the Wassa leases. In addition to the initiatives above, GSWL conducts engagement programs and collaborates with the Minerals Commission, legal small-scale mining associations and community organizations to cede areas of concessions to facilitate legal small-scale mining, along with appropriate security around active mining areas. With this approach, GSR has the opinion that galamsey around Wassa has little potential to impact current or future operations.

1.14.4Closure Planning

Closure concepts and provisional plans are required for permitting and updated three-yearly in the EMP.

The Asset Retirement Obligation estimate is $19.83 M and Practical Closure is estimated at $14.31 M.

1.15Capital and Operating Costs

Capital and operating cost estimates apply the following bases:

·All costs are in US$ for both historic actuals and forward looking expenditures. Majority of costs (including local Ghana) are aligned to US$, negating the effect of exchange rates.
·Expenditures aligned to physical schedules over the life of the project.
·Forward estimates calibrated to 2020 actual spend (Jan-Dec 2020).
1.15.1Capital Costs

Capital costs have been classified into growth (to expand or increase capacity of the operation from the current established base) and sustaining (sustain the established base) capital.

Table 1-10 Capital Cost Summary for Mineral Reserve plan

 

 

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1.15.2Operating Costs

Operating costs have been estimated as follows:

·Analysis of 2020 actual spend;
·Analysis of share of fixed and variable cost for each activity, at the cost element level (eg: fuel, labour, consumables);
·Calculation of periodic spend, driven by scheduled units of a physical activity; and
·Review of step change fixed costs where higher physical rates are planned.

Table 1-11 Cost Estimate, Operating for Mineral Reserve plan

 

 

Table 1-12 Operating Cost Summary for Mineral Reserve plan

 

1.16Economic Analysis

The Mineral Reserve has been valued using discounted cash flows to determine NPV, as at 31 December 2020. It shows positive cash flow at the $1300 /oz reserve selling price and supports declaration of a Mineral Reserve.

For the Mineral Reserve:

·Growth Capital:                  $47.7 M;
·Development Duration: nil (in production);
·Production Phase Life:   6 years (2021-2026);
·Production Phase Rate:   171 koz/yr;
·All-in Sustaining Cost:      $881 /oz; and
·After-tax NPV5%:
oBase Case ($1,300 /oz):                    $121.2 M (100% basis)
oConsensus Case (av $1,751 /oz): $335.6 M (100% basis)

(Consensus of 27 banks and financial institutions, as at the end of January 2021)

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1.17Preliminary Economic Assessment of the Southern Extension Zone

The PEA is entirely based on an Inferred Mineral Resource and there is no certainty that further geological drilling will result in the determination of higher Mineral Resource classification, nor that production and financial outcomes will be realized. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

The PEA of the Southern Extension zone has been completed with the following limits and scope:

·Inferred Mineral Resource south of 19,240 mN;
·Scoping level mining study:
oMining method selection and methodology;
oStope optimization;
oMine design to determine development quantities which inform cost estimate;
oVentilation design and modelling;
oSimulation of truck haulage to validate production rate;
oDefinition drilling strategy;
oPreliminary scheduling;
·Review of metallurgical test work and processing capacity;
·Review of permitting requirements;
·Estimation of capital and operating cost; and
·Economic analysis.

The PEA shows the Southern Extension project is potentially economically viable with an after-tax NPV at 5% discount rate, of $ 783.5M (100% Basis).

·Growth Capital:                $228.8 M;
·Development Duration: 6 years (Y1-Y6);
·Production Phase Life:   11 years (Y7-Y17);
·Production Phase Rate:   294 koz/yr;
·All-in Sustaining Cost:      $778 /oz; and
·After-tax NPV5%:
oBase Case ($1,300 /oz):             $452.2 M (100% basis)
oConsensus Case ($1,585 /oz): $783.5 M (100% basis)

Conversion risk of the Inferred Mineral Resource has been addressed through application of cut-off grades and modifying factors in the different mining panels. 54% of metal is included in the PEA inventory for Panels 4 and 5 where there is more definition drilling and the inclusion factor decreases to 48% for the deeper panels 7 and 8 where definition drilling is more widely spaced.

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Table 1-13 Conversion of Inferred Mineral Resource to PEA inventory

 UnitsPanel 4Panel 5Panel 6Panel 7Panel 8Total

Inferred Mineral Resource,

in-situ

Mt7.811.58.619.618.666
Au g/t3.03.12.74.03.63.4
Moz0.761.140.742.522.147.3
PEA InventoryMt4.15.53.19.47.830
Au g/t3.33.53.74.33.83.8
Moz0.420.610.371.310.943.6
Conversion to PEA Inventory%Moz54%49%48%50%
Cut-off GradeAu g/t2.3 g/t2.9 g/t-
Modifying Factors, Stopes 

7.5% Dilution

95.0% Recovery

13.0% Dilution

75.0% Recovery

-

 

 

Figure 1-4 Processing schedule for Southern Extension PEA

Figure 1-5 Gold Production Schedule for Southern Extension PEA

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Realization of the production schedule from the PEA carries a number of risks in addition to those defined in Section 1.18.2.

·Geology:

The primary risk is that the PEA is based on an Inferred Mineral Resource. As further definition drilling is completed, current interpretations and estimates of geological continuity, gold grade, and mineralization volumes may not be realized.

·Mining:

The PEA is based on a scoping level study and assumes increased productivity rates which are planned but not yet achieved. The main risks to achieving the PEA outcomes are that geotechnical conditions result in slower production and if operations are unable to achieve development advance and stope turnover rates.

·Processing:

Limited metallurgical test work has been completed for the PEA mining area. Results generally suggest processing will be consistent with current operations, but there is minor variability and further test work may identify that planned recoveries and throughput rates may not be achieved.

·Capital and Operating Costs:

The PEA is based on a scoping level study and with further studies, capital and operating costs, which are based on current costs and increased productivity, may increase.

There are a number of opportunities specific to the PEA plan as studies progress.

·Geology:

The Inferred Mineral Resource which informs the PEA is open to the north, south and up and down-dip. Should further drilling increase the defined mineralization, project life and production rates may be increased. Further drilling may also confirm the materially higher grades and mineralization continuity in the deeper Panels 7-8 so that less conservative modifying factors can be applied.

·Mining:

Stope size and level intervals are consistent with current operations and may be increased as studies progress, which would reduce development quantities and cost. Haulage optimization studies and emerging electrification technology may confirm an alternative to the planned diesel truck system which would result in reduced costs (mostly ventilation) and emissions.

1.18Conclusions and Interpretations
1.18.1Conclusions

The following interpretations and conclusions are made by the Qualified Persons in their respective areas of expertise, based on the review of data contained in this Technical Report.

·Mineral Titles and Agreements, Surface Rights, Royalties and Encumberances:

The required mineral titles, surface and access rights, permits and approvals exist and are in good standing required to support ongoing operations.

There is a 5% royalty on gross revenue payable to the Government of Ghana.

There is a two-tier gold stream to Royal Gold Inc. and royalty payments and tax to government are payable prior to the stream payments.

·Exploration, Driling and Data Collection:

The following are appropriate to support estimation of Mineral Resources and Mineral Reserves:

oUnderstanding of the geological setting, lithologies and structural and alteration controls on the mineralization;
oExploration programs completed to date;
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oSampling methods used to collect raw data;
oSample preparation, analysis and security;
oQuantity and quality of the lithological, structural, collar and down-hole survey data collected during drilling programs; and
oQA/QC programs to address issues of precision and accuracy.
·Metallurgical Test Work

Test work programs have been completed which are reflective of processing plant performance and used samples which reasonably represent the plant feed scheduled in the Mineral Reserve plan.

No significant metallurgical issues were identified and this has been validated by actual plant performance.

·Mineral Resource Estimates

Mineral Resources are estimated as:

oMeasured and Indicated Mineral Resource: 29.3 Mt at 3.76 g/t, containing 3.54 Moz; and
oInferred Mineral Resource: 74.0 Mt at 3.44 g/t, containing 8.18 Moz.

The Mineral Resources have been prepared in accordance with the 2014 CIM Definition Standards and 2019 Best Practice Guideline. Mining is assumed by underground methods at Wassa and Hwini Butre, and open pit methods at all other locations.

Mineral Resources have a reasonable chances for of eventual economic extraction, with estimates constrained as follows, assuming $1,500 /oz gold selling price:

oOpen Pit: constrained by open pit optimization shell based on a $1,500 /oz gold selling price and cut-off grade (0.55 g/t); and
oUnderground: constrained by cut-off grade (1.4 g/t).
·Mineral Reserve Estimate

Proven and Probable Mineral Reserves are estimated as 11.5 Mt at 2.94 g/t, containing 1.09 Moz.

The Mineral Reserves have been prepared in accordance with the 2014 CIM Definition Standards and 2019 Best Practice Guidelines. Mining will be by underground long hole open stoping. The former open pit component of the Mineral Reserve has been replaced by underground extraction.

Mineral Reserves are supported by a positive economic test assuming $1,300 /oz gold selling price.

·Mining Methods

The mine plan and schedule use:

oConventional underground mining practices and equipment to carry out long hole open stoping, consistent with currently employed techniques;
oDemonstrated mining rates based on recent development and stoping performance;

The mine plan includes appropriate consideration of:

oGeotechnical conditions;
oStope modifying factors;
oMine ventilation;
oMine dewatering;
oScheduling interactions and rates; and
oMobile fleet capacities.

The introduction of paste fill will require integrating into the stope cycle sequence to enable secondary stoping to commence.

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·Recovery Methods

Recovery methods in the processing plant and forward recovery assumptions (average 94.1%) are supported by test work and plant history.

The processing plant capacity exceeds mine production in all years for the Mineral Reserve plan.

·Environmental, Permitting and Social Considerations

All required environmental and social regulatory requirements to support ongoing operations are in place and maintained in good standing.

GSR complies with international requirements on environmental and conservation, human rights, and anti-corruption. It has adopted voluntary international codes on corporate responsibility in the areas of cyanide management, TSF design, responsible gold mining and resettlement.

GSWL has posted and periodically updates its reclamation bond ($13.7 M at end of 2020).

For environmental impacts, appropriate studies and surveys have been completed, design features and management practices are established and monitoring programmes are in place for:

oWater quality;
oAir quality;
oNoise and vibration; and
oBiodiversity.

GSR supports a number of community and social initiatives:

oGolden Star Development Foundation (community and social development projects);
oGolden Star Oil Palm Plantation (agribusiness sponsored by GSR which aims to become self-supporting); and
oCapacity building and livelihood enhancement (skills training, local procurement)

These initiatives proactively aim to build capacity and diversify the economy of local communities as well as reduce uptake of small-scale illegal mining.

·Capital and Operating Costs

Capital and operating costs have been estimated based on actual 2020 activity costs and 2021 budget costs, projected through the mine plan.

oThe growth capital cost for the life of mine is $47.7 M; and
oThe sustaining capital cost for the life of mine is $136.5 M.

Unit production costs estimated for the Mineral Reserve are:

oDirect operating cost: $669 /oz;
oAll-in sustaining cost: $902 /oz; and
oAll-in cost: $964 /oz.
·Economic Analysis of the Mineral Reserve

An economic analysis to support the declared Mineral Reserve was prepared. Using the assumptions outlined in this Technical Report, the operations show a positive cash flow at the $1300 /oz reserve selling price and support the declaration of a Mineral Reserve.

oGrowth Capital:                $47.7 M;
oDevelopment Duration: nil (in production);
oProduction Phase:           6 years, averaging 171 koz/yr;
oAll-in Sustaining Cost:     $941 /oz; and
oAfter-tax NPV5%:
§Base Case ($1,300 /oz):             $117.3 M (100% basis)

 

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§Consensus Case (av $1,751 /oz): $331.7 M (100% basis)
1.18.2Risks

Realization of the production schedule from the Mineral Reserve carries a number of risks.

·Geology:

Tightly spaced definition drilling is required which, if not completed sufficiently ahead of production could cause production delays or unexpected grade outcomes and negatively impact production and cash flow.

·Mining:

Delivery of the Mineral Reserve plan requires maintaining current productivity for development and stoping activities. Geotechnical conditions are currently very good, but ongoing review and management is required to ensure adverse geotechnical results do not adversely impact production and cash flow.

·Processing:

No material processing risks were identified for the Mineral Reserve.

·Infrastructure:

Delays to commissioning and/or achieving design capacity of the paste backfill plant and mining system will adversely impact production. The intake and return ventilation shafts are not yet geotechnically assessed and adverse findings may add cost and/or time to complete the upgrade, impacting production rates and cash flow.

·Capital and Operating Costs:

Capital and operating costs may significantly increase, particularly if productivity assumptions are not met or there are adverse movements of major cost components (eg: labour, energy).

·Environmental and Social:

Delivery of the Mineral Reserve plan requires access to personnel outside the local communities and this may be impacted by both regional (competition, security) and global (pandemic, transport) factors. Additionally, modernization of practices and technology may reduce reliance on un/semi-skilled labour, limiting accessibility to jobs for local community members, which may adversely impact community support and/or increase artisanal mining around Wassa with commensurate negative outcomes for closure costs and reputation.

1.18.3Opportunities

A number of opportunities have been identified with potential to add value to the Mineral Reserve plan.

·Mineral Resource:

Upside potential exists for the Mineral Resource from definition drilling to upgrade the large Inferred Mineral Resource which is the Southern Extension zone and various targets to grow the defined mineralisation which are not yet tested.

·Productivity and Mine Design:

Mining practices could deliver improved cost and productivity outcomes through application of technology, geotechnical design optimization and improvements to the paste backfill system after it reaches steady-state operation.

·Sustainability:

Identified opportunities exist for emissions reduction (electrification, haulage optimization and application of renewables), water (efficiency and quality preservation) and energy efficiency (comminution optimization).

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1.19Recommendations

Based on the positive results of the technical and economic analysis of the Mineral Reserve of the Wassa gold mine, the following actions are recommended:

·Continue definition drilling to support production in Panels 1 and 2 and increase geological confidence in Panel 3;
·Complete drilling programs with potential to increase the defined mineralization (main Wassa orebody, local soil sampling anomalies and regional including Father Brown/Adoikrom UG).
·Continue extraction of the Mineral Reserve by underground methods, at the optimized cut-off grade of 1.9 g/t and transition the upper areas previously planned for open pit mining, to more selective underground extraction to improve margins and bring forward production;
·Continue delivery of major capital projects (paste backfill plant and system, ventilation upgrade with two new shafts to surface, development of Panel 3 underground);
·Continue processing using CIL treatment in the Wassa processing plant;
·Continue current governance practices to ensure ongoing statutory compliance and license to operate is maintained, including management systems, social investment programs and corporate responsibility programs; and
·Investigate potential to expedite stoping from the Panel 3 (242 and B-Shoot).

For the Inferred Mineral Resource in the Southern Extension Zone, based on the positive results of the preliminary economic assessment, the risks and opportunities identified, and conclusions made, the following actions are recommended to progress the project:

·Continue definition drilling to increase geological confidence to enable upgrading classification of the Inferred Mineral Resource for Panels 4 and 5;
·Commence technical studies to feasibility study level for the disciplines of metallurgy, geotechnics, ventilation and mine design;
·Complete option and trade-off studies to optimize the project plan, including assessment of alternative haulage options (eg: shaft, conveyor), equipment selection (eg: semi/full automation, battery electric) and mine design (level interval);
·Conduct trials in the current operation to validate the proposed stoping methodology in Panels 4-8;
·Investigate electrification and renewable energy options to reduce emissions;
·Complete site water balance model and assess opportunities to improve water efficiency and reduce discharged contaminants; and
·Review crushing and grinding circuit to optimize comminution efficiency.

The progressive development plan proposed for the Southern Extension zone has the project being developed in three major phases of definition drilling and capital investment.

·Panels 4 and 5: Resource development drilling in progress and studies planned to inform and investment decision at the end of project year 2 (Y2).
·Panels 6 and 7: Resource development drilling in Y6-7, decline development starting Y6 and stope production in Y8.
·Panel 8: Resource development drilling in Y10, decline development starting Y10 and stope production in Y12.

The project execution plan for progressing to production from the Southern Extension zone outlines the activities for only the first phase (Panels 4 and 5) to reach an investment decision with a feasibility level study and the potential timeframe to production. The estimated cost to deliver the feasibility study is $14.0 M, which is mostly for definition drilling ($13.2 M).

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Figure 1-6 Project Execution Plan, Southern Extension Panels 4 and 5

 

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2Introduction
2.1Terms of Reference

This Technical Report has been prepared to meet the requirements defined by Form 43-101F1, by and for Golden Star Resources, describing the Wassa gold mine in Ghana. The report provides updated information on the currently operating mine, including an updated Mineral Resource and Mineral Reserve estimate.

The Report also contains the summary of a Preliminary Economic Assessment (PEA) completed in support of the potential expansion of the underground mine to extract the Inferred Mineral Resource in the Southern Extension zone (2020 PEA).

The PEA has been prepared within the following framework:

·Underground mining rate increased to fully utilise the installed processing capacity (2.7 Mtpa);
·Production schedules to appropriately consider conversion risk of the Inferred Mineral Resource;
·Methodologies and design quantities based on proven, currently available technologies;
·Costs to reflect current operational experience; and
·Minimise capital demand needed to establish full production.

The intent of the framework is to present a deliverable PEA plan which can be executed with GSR’s current operational and financing capacity. Potential enhancements outside this framework are presented as opportunities outside of the PEA outcomes and can be investigated as part of the forward work plan.

The 2020 PEA has no impact on the Mineral Reserves, nor on the key assumptions and parameters supporting the Mineral Reserves. The Mineral Reserves are current, valid and do not rely on any of the assumptions made in the 2020 PEA.

The 2020 PEA is conceptual and outlines a mining inventory which is entirely based on an Inferred Mineral Resource. Inferred is the lowest level of confidence for a Mineral Resource and there is no certainty that further geological drilling will result in the determination of higher Mineral Resource classification, nor that production and financial outcomes will be realized. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

The Mineral Resources and Mineral Reserves have been prepared in accordance with CIM Definition Standards for Mineral Resources and Mineral Reserves, adopted by the CIM Council on May 10, 2014, and the CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines, adopted by CIM Council on November 29, 2019.

References within this report to “GSR” include Golden Star Resources (GSR) and Golden Star Wassa Limited (GWSL) as the context requires.

Golden Star is a Canadian federally-incorporated international gold mining and exploration company, producing gold in Ghana, West Africa. This report has been prepared to satisfy GSR’s obligations as a reporting issuer in Canada.

Units used in the report are metric units unless otherwise stated. Monetary units are in United States dollars (US$) unless otherwise stated.

2.2Wassa Gold Mine

The Wassa Gold Mine is located near the village of Akyempim in the Wassa East District, in the Western Region of Ghana. It is 80 km north of Cape Coast and 150 km west of the capital Accra. The property lies between latitudes 5°25’ and 5°30’ north and between longitudes 1°42’ and 1°46’ east. GSWL owns the rights to mine the Wassa, Benso and Hwini Butre concessions. GSR owns a 90% interest in and manages GSWL with the Government of Ghana owning the remaining 10%.

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2.3Principal Sources of Information

This Technical Report was prepared by GSR. Information for the report was based on published material, as well as data, professional opinions, and unpublished material from work completed by GSR. It includes information provided by and discussions with third party contractors and consultants engaged by GSR. Section 27 contains the list of reports and documents used in preparation of this report.

Major contributions by contractors and consultants have been reviewed and approved by QP’s as follows:

·Environmental impact assessment studies undertaken by Golder Associates (Ghana and South Africa);
·Geological modelling:
oLong-range model resource estimate prepared by SRK (Toronto);
oShort-range model wireframe and resource estimate prepared by SRK (Moscow).
·Geotechnical assessment prepared by OreTeck Mining Solutions (Australia);
·Metallurgy and Processing assessment prepared by MineScope Services (Australia);
·Mine Ventilation assessment prepared by SRK (US);
·Tailings storage facility design and geotechnical assessments by Knight Piésold Ghana (the engineer of record); and
·Paste backfill studies carried out by Outotec (Canada) Ltd.
2.4Qualified Persons

Matt Varvari (not independent) is the Qualified Person (QP) responsible for overall project management of the Technical Report and specifically, Sections 1-3, 6, 13, 15-19 and 21-27 of this report. They are a Fellow of the Australasian Institute of Mining and Metallurgy and have the required qualifications and experience to act as a QP. Matt is based in London, UK and is employed full-time by GSR as Vice President Technical Services.

S. Mitchel Wasel (not independent) is the QP responsible for Sections 7-12 and 14 of this report. They are a Chartered Professional of the Australasian Institute of Mining and Metallurgy and have the required qualifications and experience to act as a QP. Mitch is based in Takoradi, Ghana and is employed full-time by GSR as Vice President Exploration.

Philipa Varris (not independent) is the QP responsible for Sections 4, 5 and 20 of this report. They are a Chartered Professional of the Australasian Institute of Mining and Metallurgy and have the required qualifications and experience to act as a QP. Philipa is based in London, UK and is employed by GSR as Executive Vice President and Head of Sustainability.

All QP’s have conducted sufficient visits to Wassa site, as detailed in Table 2-1.

Table 2-1 Qualified persons and site visits

 

 

 

2.5Effective Dates
·Effective date of the Wassa Mineral Resource: 31 December 2020
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·Effective date of the Wassa Mineral Reserve: 31 December 2020
·Effective date of the Economic Analysis for the Mineral Reserve: 31 December 2020
·Effective date of the Preliminary Economic Assessment: 31 December 2020
2.6Previous Technical Report

Golden Star Resources filed the following Technical Report on Wassa with an effective date of 31 December 2018:

·Raffield, M., Wasel, M. and Varris, P. NI 43-101 Technical Report on Resources and Reserves, Golden Star Resources, Wassa Gold Mine, Ghana
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3Reliance on Other Experts

In the preparation of this Technical Report, the qualified persons in specific instances relied on studies, reports, opinions or statements of experts who are not qualified persons. These include:

·Environmental impact assessment studies:
o“Golden Star (Wassa) Limited; Updated Tailings Storage Facility (TSF) 2 Project Environmental Impact Statement”. Prepared by Golder Associates, November 2016, regarding environmental impacts as reported in Section 18 and 20.
o“Golden Star (Wassa) Limited; Tailings Storage Facility (TSF) 2 Project Environmental Impact Statement”. Prepared by Geosystems Consulting, February 2012, regarding environmental impacts as reported in Section 18 and 20.
o“Wassa Expansion Project Environmental Impact Statement”. Prepared by Geosystems Consulting, September 2015, regarding environmental impacts as reported in Section 18 and 20.
o“Environmental Impact Statement for the Wassa Project”. Prepared by Wexford Goldfields Limited (WGL), 2004, regarding environmental impacts as reported in Section 20.
o“The Wassa Project Environmental Impact Statement”. Prepared by Scott Wilson, 2004, regarding environmental impacts as reported in Section 20.
o“Satellite Goldfields Limited, Wassa Gold Project, Environmental Baseline Study”. Prepared by SGS Laboratory Services (Ghana) Limited, November 1996, regarding environmental baseline as reported in Section 20.
·“Final GSR Mining Title Opinion”. Prepared by REM Law Consultancy (Accra), February 2021, regarding the good standing of the Wassa, Benso and Hwini Butre mining leases as reported in Section 4.
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4Property Description and Location
4.1Location of Mineral Concessions

The Wassa Mine is located near the village of Akyempim in the Wassa East District in the Western Region of Ghana, approximately 80 km north of Cape Coast and 150 km west of the capital, Accra. It lies between latitudes 5°25’ and 5°30’ N and longitudes 1°42’ and 1°46’ E. The location of the Wassa Mine is shown in Figure 1-1.

The Wassa Mine is operated under the Wassa mining lease which was issued on September 17, 1992. The total surface area of the Wassa Mining Lease is 5,289 Ha, with approximately 595 Ha of disturbance from GSWL’s activities. GSWL has applied for a reshape of the concession boundary to comply with recent changes to cadastre grid requirements by the Minerals Commission, which will modify the total area of the concession to 6,496 ha.

 

Figure 4-1 Wassa Mine Location in Ghana, West Africa (United Nations, 2018)

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In addition to the Wassa mining lease, GSWL holds the Hwini Butre and Benso mining leases, and several prospecting licences in the Western Region of Ghana. GSWL’s mineral properties are shown in Figure 4-2.

Figure 4-2 Wassa Mine Location in Ghana, West Africa (GSR, 2021)

 

Figure 4-3 shows the locations of GSWL’s mineral concessions and operations:

·Wassa mining lease: Wassa is an operating underground gold mine comprising the following mineralization domains: F Shoot, 419, B Shoot, 242, Starter, South-East, Mid-East and Dead Man’s Hill. SAK comprises several deposits to the west.
·Benso mining lease: comprising the Subriso East, Subriso West, G-Zone, C-Zone and I-Zone deposits.
·Hwini Butre mining lease: comprising the Father Brown, Adoikrom and Dabokrom deposits.
·Benso (Chichiwelli) exploration property: comprising two mineralized zones, Chichiwelli West and Chichiwelli East.
·Manso exploration property: located east of Benso and Hwini Butre.

The properties and leases are spread along a trend of approximately 80 km southwest of the Wassa mine. There are sufficient access and surface rights for GSWL’s operations.

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Figure 4-3 Location of operations and infrastructure and concession boundaries (GSR, 2021)

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4.2Mineral Rights

The Constitution of Ghana vests title in every mineral in its natural state to the Government of Ghana. The exercise of any mineral right in Ghana requires an appropriate mineral title to be issued by the Government of Ghana acting through the Minister responsible for Lands and Natural Resources. The Minister of Lands and Natural Resources administers, promotes and regulates Ghana’s mineral wealth through the Minerals Commission, a governmental organization designed in accordance with the Minerals Commission Act 1993 and the Minerals and Mining Act, 2006 Act 703 (Minerals and Mining Act).

A person must apply to the Minerals Commission and be granted a mineral right by the Minister of Lands and Natural Resources before they can search, survey, prospect, explore or mine for a mineral anywhere in Ghana. There are different types of licenses (namely, reconnaissance and prospecting licenses, and mining leases) for the different mining activities. Each type of licence details the activities that are permitted.

The Government of Ghana holds a 10% free-carried interest in all companies holding mining leases. The 10% free-carried interest entitles the Government to a pro-rata share of future dividends. The Government has no obligation to contribute development capital or operating expenses.

Table 4-1 sets out the mineral rights held by GSWL (or those in which GSWL has an interest). GSR will from time to time seek a title opinion from its legal counsel in Ghana to confirm its title in its material mineral properties, and the good standing of the underlying mineral rights.

Table 4-1 Mineral rights held by GSWL

Name

Type of

Mineral right

No.Issuing AuthorityIssue DateExpiry Date

Surface

Area

Third-Part OwnershipComments

Wassa

Mining LeaseLVB 87618/94Minerals Commission17/09/199216/09/2022

 

 

52.89 km2

Government of Ghana holds a 10% free-carried interest 

Benso

Mining LeaseLVDGAST 37993462020Minerals Commission25/08/202024/08/2031

 

 

19.45 km2

Government of Ghana holds a 10% free-carried interest 

Hwini Butre

Mining LeaseLVDGAST 38000372020Minerals Commission25/08/202024/08/2031

 

 

43 km2

Government of Ghana holds a 10% free-carried interest 

 Dwaben (Safric)

Reconnaissance LicenceLVB1624/06Minerals Commission02/02/2006(expired in 2020)

 

 

 

 

26.92 km2

 Application to convert reconnaissance licence to prospecting licence submitted to the Minerals Commission in November 2020

 Benso (Chichiwelli)

Prospecting LicencePL.2/1550Minerals Commission27/09/2007-

 

 

 

22.46 km2

 Notice of grant of extension of prospecting licence issued by the Minerals Commission in January 2020

 

Abura

Abura Prospecting LicencePL 2/135Minerals Commission13/12/201812/12/2021

 

 

 

65.10 km2

Subject to option agreement with Bowden Gold Resources Limited 

 Manso 1

Prospecting LicencePL 2/378Minerals Commission10/01/2005-

 

 

101.6 km2

 Application to renew prospecting licence submitted to the Minerals Commission in August 2020

 Manso 2

Prospecting LicencePL 2/337Minerals Commission07/09/2007-

 

 

21.38 km2

Subject to option agreement with Pacific Mining LimitedApplication to renew prospecting licence submitted to the Minerals Commission in September 2020

 

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The Wassa Mine sits within the Wassa mining lease which comprises an area of 52.89 km2 lying to the north and south of latitudes 5°25’ and 5°30’, respectively and bounded to the east and west by longitudes 1°42’ and 1°46’, respectively.

The Wassa mining lease was entered between the Government of Ghana and Satellite Goldfields Limited (SGL) on September 17, 1992 for a term of 30 years, renewable. In 2002, the mining lease was assigned by SGL to GSWL with the written consent of the Government of Ghana. GSWL is the registered legal and beneficial holder of the Wassa mining lease. The Government of Ghana holds 10% of GSWL share capital.

4.3Royalties and Other Payments; Encumbrances

GSWL pays to the Government of Ghana within thirty days from the end of each quarter a royalty at a rate of 5% determined based on the total revenue of minerals produced during the quarter. This royalty is payable prior to any adjustments from the Royal Gold stream (see below). Royalties are paid through the Commissioner of Internal Revenue.

Payment of annual ground rent is made to the owner of the land except in the case of annual ground rent in respect of mineral rights over stool lands, which are paid to the Office of the Administrator of Stool Lands. A holder of a mineral right must also pay to the Minerals Commission an annual mineral right fee determined based on the type of tenure. GSWL pays annual ground rate and annual fees in relation to all the mineral rights it holds.

GSR is party to a gold purchase and sale agreement with Royal Gold, Inc. through its wholly owned subsidiary RGLD Gold AG (RGLD). The agreement was initiated on 6 May 2015, amended on 29 Jun 2018, 17 October 2019 and most recently 30 September 2020. The stream covers all gold produced within GSWL’s mineral concessions and requires GSR to deliver according to two tiers:

·Tier 1: 10.5% of all production to RGLD at a cash purchase price of 20% of spot gold until 240,000 ounces have been delivered; and
·Tier 2: thereafter, to deliver 5.5% of all production to RGLD at a cash purchase price of 30% of spot gold.

Pursuant to the terms of the gold sale and purchase agreement, GSR is restricted from granting encumbrances on the Wassa gold project without RGLD’s consent. In 2019, GSR entered into a credit facility agreement with Macquarie Bank Limited pursuant to which GSWL’s mineral rights were, with the approval of the Minister of Lands and Natural Resources and RGLD, encumbered to secure the repayment of the loan.

At the end of December 2020, the remaining balance of the Tier 1 stream was 120,003 oz. The stream is treated as a revenue adjustment.

4.4Historic Environmental Liability and Indemnity

The Wassa operations were permitted under an environmental impact assessment developed for SGL in 1998. At commencement, Wassa was a heap leach operation fed by the Main pits complex comprising the interconnected South-East, 242, B-Shoot, F-Shoot, South, Main South, and 419 pits. The predominant liabilities of the original SGL operations, including heap leach area and waste dumps, have since been fully encompassed by the GSWL operations.

In 2002, GSR purchased certain assets of SGL and liabilities for the operations transferred. In 2005, GSR acquired St. Jude Resources (Ghana) Limited (SJR) and, with it, the Hwini Butre and Benso (HBB) properties and their associated liabilities. Likewise, the development of the HBB operations by GSWL saw the establishment of infrastructure that fully encompassed the previous areas of disturbance of SJR. The establishment of the reclamation security agreement with the EPA in 2005 and the associated bond with the EPA addresses security for reclamation and closure obligations.

There are no other legacy issues associated with the GSWL site.

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4.5Permits and Authorization

In addition to the mineral rights specified in Table 4-1, GSWL requires certain permits and licenses to carry out its activities, including:

·Mining operating permit:

The Minerals and Mining (Health, Safety and Technical) Regulations, 2012 (L.I. 2182) prescribe technical and health and safety standards for mining operations and require a person who is granted a mining lease to, before the commencement of operation of the mine, obtain a mining operating permit from the Inspectorate Division of the Minerals Commission.

·Environmental permit:

The Environmental Assessment Regulations, 1999 (L.I. 1652) require that all developmental activities likely to impact adversely on the environment be subject to environmental assessments. Pursuant to these regulations, an undertaking which in the opinion of the Environmental Protection Agency (EPA) has or is likely to have an adverse effect on the environment cannot commence unless the undertaking has been registered and an environmental permit has been issued by the EPA. The Minerals and Mining Act requires that all necessary approvals and permits required to be obtained from the Forestry Commission and the EPA for the protection of natural resources, public health and the environment.

The major environmental permits in place for the Wassa mine are:

oWassa operations (EPA/EIA/112) and expansions (EPA/EIA/322) including South Akyempim pits (EPA/EIA/190);
oHwini Butre and Benso operations (EPA/EIA/175) and expansion (EPA/EIA/247).
oWassa TSF 2 (EPA/EIA/383) and renewal (EPA/EIA/442); and
oWassa Expansion project, including Wassa underground, Main pits and waste dump expansion (EPA/EIA/508).
·Licence to export, sell or dispose of minerals:

The exportation, sale or disposal of minerals requires a licence from the Minister for Lands and Natural Resources. Pursuant to section 46 of the Minerals and Mining Act., a mining lease authorizes the holder to, inter alia, “take and remove from the land the specified minerals and to dispose of them in accordance with the holder’s approved marketing plan.” Under the Minerals and Mining (General) Regulations, 2012 (L.I. 2173), an application by a holder of a mining lease for a licence to export, sell or dispose of gold or other precious minerals produced by the holder must be accompanied by a refining contract and a sales and marketing agreement.

·Operating licence and permit for the acquisition, use, transportation and storage of explosives:

Under Regulation 23 of the Minerals and Mining (Explosives) Regulations, 2012 (L.I. 2177), the construction of a building or other structure to be used as a magazine for the storage of explosives is subject to an operating license delivered by the Minerals Commission. As required under Regulation 32 of L.I. 2177, the storage of explosives in a magazine is also subject to a permit from the Minerals Commission; the latter is valid for one year and is renewable on application. Under L.I. 2177, an operating licence is required for the purchase and use or transportation of explosives. There are separate operating licences for the purchase and use of explosives and for transportation. Each is valid for a period of one calendar year and is renewable on application made one month before the end of each year. Additionally, a permit is required for each occasion on which explosives are being transported in respect of which the specific type and quantity of explosives must be indicated.

·Licence to use water resources:

The use of water resources is regulated by the Water Resources Commission Act, 1996 (Act 522) and the Water Use Regulations, 2001. Act 522 provides that no person shall (a) divert, dam, store, abstract or use water resources; or (b) construct or maintain any works for the use of water resources except in accordance with the provisions of the Act. Subject to obtaining the requisite approvals or licences, a holder of a mineral right may, for purposes of or ancillary to the mineral operations, obtain, divert, impound, convey and use water from a river, stream, underground reservoir or watercourse within the land the subject of the mineral right.

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·Fire permit:

The Fire Precaution (Premises) Regulations, 2003 requires that a fire certificate be issued by the Chief Fire Officer in respect of premises used as a place of work or for a purpose which involves access to the premises by members of the public, whether on payment or not. The certificate is valid for 12 months and is renewable.

GSWL conducts its operations in accordance with applicable laws and regulations in Ghana and is in compliance with its permitting obligations in relation to its activities. With regards to environmental matters, GSWL has undertaken environmental impact assessment studies on its concessions to support the permitting of its mining projects and has considerable background data to support required environmental permitting processes.

 

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5Accessibility, Climate, Local Resources, Infrastructure and Physiography
5.1Accessibility

The Wassa Mine is near the village of Akyempim in the Wassa East District in the Western Region of Ghana. It is 62 km north of the district capital, Daboase, and 40 km east of Bogoso, and is 80 km north of Cape Coast and 150 km west of the capital Accra.

The main access to the site is from the east, via the Cape Coast to Twifo-Praso road, then over the combined road-rail bridge on the Pra River. There is also an access road from Takoradi in the south via Mpohor.

The satellite sites of Hwini Butre and Benso are respectively, 60 km and 35 km southeast of the main Wassa site. They are accessible from Wassa via an unsealed access and haulage road. All sites lie within 15 km of sealed public road but generally, the haul road to Wassa is the most reliable access.

Figure 4-3 in Section 4.1 shows a plan of the various locations and access infrastructure.

5.2Physiography and Vegetation

The project area is characterized by gently rolling hills with elevations up to 1100 m RL, incised by an extensive drainage network. The natural vegetation is an ecotone of the moist, semi-deciduous forest and wet rainforest zones. It has been degraded due to anthropogenic activities, giving way to broken forest, thickets of secondary forest, forb re-growth, swamps in the bottom of valleys, and cleared areas.

Extensive subsistence farming occurs throughout the area, with plantain, cassava, pineapple, maize, and cocoyam being the principal crops. Some small-scale cultivation of commercial crops is also carried out, with cocoa, teak, coconut and oil palm the most common. Forest patches are present on the steep slopes and in areas unsuitable for agriculture.

Environmental assessments carried out in the project area over the last two decades (SGS 1996 and 1998, WGL 2004, GSR 2015, Geosystems 2013, and Golder 2016) indicate that the biodiversity of the Wassa operational area is of low ecological significance and conservation status.

5.3Land Use and Proximity to Local Population Centres

The Wassa Mine is located in a rural setting with no major urban settlements within 30 km. It lies in the Wassa East District, part of the Western Region of Ghana, 40 km north of Daboase (district capital), 65 km north of Takoradi (regional capital) and 35 km north-east of the city of Tarkwa.

The nearest villages are Akyempim, Akyempim New Site (formally Akosombo, resettled early in Wassa operations) and Kubekro. The Togbekrom community were resettled to Ateiku.

The Hwini Butre and Benso sites are approximately 35 km and 65 km, respectively, north-northwest of the Port of Takoradi and south-east of Tarkwa. The key communities within and outside the concession are Subriso, Odumase, Ningo, Akyaakrom, Mpohor, Benso, and Anlokrom. The total population of these communities is approximately 10,000. The Benso Township is approximately 5 km from the Benso mine site to the south and the Mpohor Township is approximately 2 km west of the Hwini Butre site.

The population data/estimates for the larger communities located within the Wassa concession boundaries are shown in Table 5-1.

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Table 5-1 Communities neighbouring Wassa Mine

CommunityDivisional Area

Estimated Population

(SGS 1996)

Population

(WEDA 2013)

AkyempimMamponso2,5002,533
AkosomboMamponson/a166
KubrekroAnyinabrem300335
NsadwesoAnyinabrem2,4001,541
TogbekromAnyinabremNot measured674

Land uses in the vicinity of GSWL operations are predominantly rural with agricultural, forestry, agroforestry (palm oil and rubber plantations), and unauthorized small-scale mining operations.

5.4Local Resources and Infrastructure

There are five other significant mining operations within 50 km of Wassa:

·Nsuta manganese mine, Ghana Manganese Company;
·Iduapriem gold mine, AngloGold Ashanti;
·Tarkwa gold mine, Gold Fields Ghana Limited;
·Damang gold mine, Gold Fields Ghana Limited; and
·Bogoso-Prestea gold mine, Future Global Resources.

Wassa Mine is in operation with required services, infrastructure, and community support to continue.

·Access is via public road to site with good overall access. Roads are sealed from Accra to within 15 km from site, then access is via formed, unsealed road. From site, travel by road to Takoradi is generally 1.5-2 hours and 4-5 hours to Accra;
·Electrical infrastructure with access to power through the grid and on-site generation;
·On-site processing plant with capacity up to 2.7 Mtpa;
·On-site tailings storage facilities with sufficient permitted capacity;
·Waste rock storage facilities with sufficient permitted capacity;
·Access/haulage road from Wassa site, to the satellite Hwini Butre, Benso and Chichiwelli sites; and
·Access to skilled labour with the history and scale of mining in Ghana.
5.5Climate and Length of Operating Season

The climate in the project area is classified as wet semi-equatorial. The Intertropical Convergence Zone crosses the area twice a year, resulting in a bi-modal rainfall pattern with peaks in Mar-Jul and Sep-Oct. During the dry season months of Nov-Feb, the climate is heavily influenced by the seasonal Harmattan which brings dry and dusty winds from the Sahara across West Africa. Rainfall is mainly influenced by south-west monsoon winds, which blow from the south-western part of the country.

Analysis of available rainfall data, obtained from the Ateiku Meteorological survey (1944 to 2009) indicates:

·Average annual rainfall is 1,996 ± 293 mm;
·Wettest month is June, with average rainfall of 241 ± 85 mm;
·Driest month is January, with average rainfall of 31 ± 35 mm;
·The wettest month on record is June 2009 with 475 mm of rainfall.

Local measurements taken at Wassa from 1998-2019 (at main site) and 2007-2014 (at TSF 1) are consistent with the large data set from Ateiku. Local stations identified the bi-modal rainfall pattern and recorded an average of approximately 1,659 mm/yr and the wettest month being June 2014 with 512 mm at TSF 1 and 417 mm at the main site.

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Annual potential evapotranspiration is estimated to be approximately 1,337 mm/yr, indicating minimum precipitation balance of +288 mm/year. In 2020, measured evaporation at the TSF site was in the order of 1,300 mm/year. Rainfall exceeds potential evapotranspiration from Mar-Jul and Sep-Oct with groundwater recharge most likely to be prevalent during these periods. Relative humidity is consistent throughout the year, ranging from 88% to 90%.

The climatic conditions mean that, with effective surface water management practices, mining operations can continue year-round with short suspensions to open pit operations during storms, most of which are short duration and can occur throughout the year. Underground mining operations are not directly affected by weather events except where very large, long duration rainfall results in the safe capacity of pit sumps being exceeded, requiring evacuation of the underground workings due to risk of inundation. Normal operations are resumed once sumps are pumped down to safe levels.

 

 

 

 

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6History
6.1Wassa
6.1.1Historic Mining

The Wassa area has experienced local small-scale and colonial mining activity at Wassa since the beginning of the 20th century with numerous small pits and adits evident.

From 1988, the property was operated as a small-scale mining operation with a gravity gold recovery circuit by WMRL, a Ghanaian company.

6.1.2Satellite Goldfields Limited (1993-2002)

In 1993 WMRL formed Satellite Goldfields Limited (SGL) with the Irish companies Glencar Exploration Limited and Moydow Ltd, assigning the Wassa mining lease to SGL.

Extensive satellite imagery and geophysical interpretations were carried out which identified a strong gold target. Exploration drilling commenced in February 1994 and by March 1997 58,709 m of drilling had been completed. Construction of the Wassa Mine was initiated in September 1998, after Glencar secured a $42.5 M debt-financing package from a consortium of banks and institutions.

The operation was originally developed as an open pit mine with heap leach treatment of 3.0 Mtpa and planned production of 100,000 oz/yr. First ore was mined from the open pit in October 1998.

During the first year of production, planned gold recovery of 85% from oxide ore in the heap leach was not achieved due to high clay content of the ore and poor solution management. Attempts were made to increase recovery, including doubling the leach solution application rate but recoveries for the oxide ores above 55-60% could not be achieved.

The low gold recovery resulted in debt servicing issues and Wassa was marketed for sale. GSR started negotiations to purchase Wassa in mid-2000. As part of due diligence, GSR initiated a drilling program in March 2001 to test their geological model and extensions to some of the high grade orebodies.

SGL was placed into receivership in November 2001 and in April 2002, GSR concluded that the mineable reserve at Wassa was 30% below the 648,000 oz stated by SGL. Negotiation continued until September 2002 when the agreement for GSR to purchase the 90% share of Wassa was announced.

6.2Hwini Butre, Benso and Chichiwelli
6.2.1Historic Mining

Early European reports indicate the Dabokrom area, around Hwini Butre, may have been a major source for gold sold to Portuguese explorers when they first arrived in Ghana in the late 1400’s.

European interest grew in the 1800’s with the presence of gold and proximity to Sekondi-Takoradi, which had developed as a port to service mines at Tarkwa, Prestea and Obuasi. Many exploration licences were granted during the gold boom of 1898-1902 and by the 1930’s most of the area was under license to various local and European interests.

At Dabokrom, a shaft was sunk by Oceania Consolidated in the 1930’s to follow the shallow dipping quartz veins. The property was worked for several years but stopped in 1939 at the start of World War 2.

At Chichiwelli a shaft was sunk in 1918 following a quartz vein at the very north end of the Benso concession, close to the Subri River Forest Reserve. Mining progressed to the 260 ft level but was abandoned in 1924 after the mine was flooded.

The whole area has many historic workings which evidence mining activity, mostly from the 1930’s.

 

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6.2.2Modern Exploration (1980’s-2005)
6.2.2.1Hwini Butre

The Dabrokrom concession was acquired by BD Goldfields (BDG) during the 1980’s who invited Danish company Lutz Resources Limited to carry out preliminary exploration on the property. The property transferred to Hwini Butre Minerals (HBM) in the early 1990’s, which was controlled by Lutz.

HBM entered a joint venture with Placer-Outukumpu who drilled several holes around Dabokrom in 1993 to assess potential of the vein systems. They concluded that potential was limited by widely spaced veins and little gold in the diorite host rock. Saint Jude Resources (SJR) acquired Dabokrom in 1994 and explored the area until 2002 when work was suspended due to a legal dispute between SJR, BDG and the Government of Ghana. The matter was resolved in 2005 before acquisition of the project by GSR.

SJR began exploring the concession in February 1995 which represented the first sustained exploration program on the concession. SJR undertook ground geophysical surveys which included magnetic, radiometric and induced polarization surveys; soil geochemical surveys were also completed on the concession area, resulting in the identification of numerous targets. Trenching and pitting were conducted in areas of geophysical and geochemical anomalies and over historical prospects or old workings in an attempt to outline near surface mineralization. Subsequent drilling of the surface targets resulted in the delineation of the Adoikrom, Father Brown and Dabokrom prospects along a combined strike length of 900 m. Further exploration conducted in 2005 identified the Adoikrom North prospect. A total of some 22,100 m over 267 drill holes were completed on the main mineralized zones and the exploration targets.

6.2.2.2Benso and Chichiwelli

Reconnaissance work at Chichiwelli, Subriso, Denerawah and Amantin was conducted by BHP Billiton from 1989-92, on what is now the Benso concession. This identified soil geochemical anomalies and follow-up drilling was completed at Chichiwelli but results did not meet targeting criteria and the concessions were relinquished. Tenure was then acquired by a local company, Architect Co-Partners, with a 150 km2 prospecting concession which covered Amantin, Subriso and Chichiwelli, as well as a large part of the Subriso River Forest Reserve that was closed to exploration from 1996.

Canadian company Fairstar Exploration Limited took over the Benso concession in 1995 and carried out extensive work, particularly at Subriso and Amantin, where considerable drilling was completed but ceased by the end of the decade due to funding constraints. An agreement was reached in 2001 for SJR to take over the exploration work.

In 2001, SJR completed an agreement with Fairstar and took over the exploration work. From early 2002 to about mid-2004, SJR focused mainly on the Subriso area where substantial mineralization was outlined at two prospects, Subriso East and West. Numerous other prospects, namely Subriso Central, I Zone and G Zone were identified and drill tested, as was the Amantin area, which had also been drilled to a considerable extent by Fairstar.

6.3Production History, Previously Declared Resources and Reserves
6.3.1Golden Star Resources (2003-present)

Since acquiring Wassa in 2003 GSR has produced 2.4 million ounces of gold and the mine has a remaining life of six years as defined by the current Mineral Reserve.

Milestones at Wassa under GSR management are:

·2003: definition drilling ahead of feasibility study for CIL plant.
·2004: feasibility study completed and construction commences on CIL plant with open pit mining.
·2005: CIL plant commissioned.
·2006: acquired St Jude Resources (Hwini Butre and Benso concessions). Connected to grid power.
·2007: commenced open pit mining at South Akyempim. Construction of haul road to Hwini Butre.
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·2008: commenced open pit mining at Benso, processing at Wassa.
·2009: commenced open pit mining at Hwini Butre and drilling to test underground potential.
·2011: Hwini Butre mining moves from Adoikrom to Father Brown pit.
·2012: commenced drilling to test underground potential below Wassa.
·2013: upgraded plant to 2.7 Mtpa capacity with fresh ore, consolidated mining at Wassa Main pit.
·2014: released positive Preliminary Economic Assessment for Wassa Underground and completed Hwini Butre mining at Father Brown.
·2015: completed positive Feasibility Study for Wassa UG and commenced development, along with starting construction of TSF 2.
·2016: mined first stope ore from Wassa UG in July and definition drilling continued to define wide zones of mineralization in B-Shoot.
·2017: commercial production declared at Wassa UG and deep definition drilling program defines what was later to become the Southern Extension zone. UG averages 1,865 ore t/d.
·2018: open pit mining of Main pit completed and UG ore mining rate increases to 2,945 t/d. Wassa UG Inferred Mineral Resource reported growth to 5.2 Moz with addition of Southern Extension zone.
·2019: completed positive Feasibility Study for paste backfill commenced development. UG ore mining rate increased to 3,895 t/d (1.4 Mtpa).
·2020: completed construction of paste backfill plant and on-site gas-fired power generation. UG ore mining rate increased to 4,480 t/d (1.6 Mtpa).

Annual production is shown in Table 6-1. Production in 2012 and 2013 includes contributions from Hwini Butre and Benso.

Production peaked in 2013 at 187 koz with the plant operating at full capacity and high grade ore being mined from the Father Brown pit at Hwini Butre. From 2014, open pit ore was sourced solely from the Wassa Main pit until its completion in 2017. Lower grades resulted in production of around 100 koz/yr.

Mining transitioned to underground from 2016, with commercial production realized in 2017 and the underground becoming the main production source by 2018. Since 2018, underground production has steadily increased to maintain and exceed 150 koz/yr, with the addition of minor amounts of low-grade ore from open pit stockpiles.

Table 6-1 Recent Production History, Wassa

YearOpen Pit & StockpileUndergroundTotal

Processed

Mt

Feed Grade

Au g/t

Produced

Au koz

Processed

Mt

Feed Grade

Au g/t

Produced

Au koz

Processed

Mt

Feed Grade

Au g/t

Produced

Au koz

20122.512.09159---2.512.09159
20132.702.29187---2.702.29187
20142.631.41110---2.631.41110
20152.501.46109---2.501.46109
20162.441.27930.182.06112.621.32104
20171.931.27760.693.03612.621.73137
20180.530.76121.014.181371.603.06151
20190.160.6531.393.571531.553.27155
20200.380.7991.643.131562.012.70165

 

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7Geological Setting and Mineralization
7.1Regional Geology

The regional geological setting of the Ashanti belt has been described by several authors previously. The most recent publication describing the geological setting of the sub-region was from Perrouty et al., in Precambrian Research in 2012.

The Ashanti greenstone belt in the Western Region of Ghana is composed primarily of paleoproterozoic metavolcanic and metasedimentary rocks that are divided into the Birimian Supergroup (Sefwi and Kumasi Groups) and the Tarkwa Group. Both units are intruded by abundant granitoids and host numerous hydrothermal gold deposits such as the Wassa, Obuasi, Bogoso and Prestea mines and paleoplacer deposits such as the Tarkwa and Teberebie Mines.

Allibone et al. (2002) separated the Paleoproterozoic Eburnean orogeny into two distinct phases known as Eburnean I and II. This classification was revised by Perrouty et al. in 2012 who proposed two distinct orogenic events, the Eoeburnean orogeny and the Eburnean orogeny. The Eoeburnean orogeny predates the deposition of Tarkwaian sediments and is associated with a major period of magmatism and metamorphism in the Sefwi Group basement. The Eburnean event is associated with significant post-Tarkwaian deformation that affected both the Birimian Supergroup and overlying Tarkwaian sediments. The Eburnean orogeny is associated with major north-west to south-east shortening that developed major thrust faults, including the Ashanti Fault along with isoclinal folds in Birimian metasediments and regional scale open folds in the Tarkwaian sediments. These features are overprinted by phases of sinistral and dextral deformational events that reactivated the existing thrust faults and resulted in shear zones with strong shear fabrics.

The Birimian series was first described by Kitson (1928) based on outcrops located in the Birim River (around 80 km east of the Ashanti Belt). Since this early interpretation, the Birimian stratigraphic column has been revised significantly. Before the application of geochronology, the Birimian super group was divided in an Upper Birimian group composed mainly of metavolcanics and a Lower Birimian group corresponding to metasedimentary basins. Subsequent authors have proposed synchronous deposition of Birimian metavolcanics. Most recently, Samarium/Neodymium and U/Pb analyses have reversed the earlier stratigraphic interpretation with the younger metasediments overlying the older metavolcanics. Proposed ages for the metavolcanics vary between 2,162 ± 6 Ma and 2,266 ± 2 Ma. Detrital zircons in the metasediments indicate the initiation of their deposition between 2,142 ± 24 Ma 2,154 ± 2 Ma. The Kumasi Group was intruded by the late sedimentary Suhuma granodiorite at 2,136 ± 19 Ma (U/Pb on zircon, Adadey et al., 2009).

 

 

 

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Figure 7-1 Location of Wassa on the Ashanti Belt (Perrouty et al 2012)

The Tarkwa super group was first recognized by Kitson (1928) and consists of a succession of clastic sedimentary units, which have been divided in four groups by Whitelaw (1929) and Junner (1940).

The Kawere Group located at the base of the Tarkwaian super group is composed of conglomerates and sandstones with a thickness varying between 250 m and 700 m. The unit is stratigraphically overlain by the Banket Formation, which is characterized by sequences of conglomerates interbedded with cross-bedded sandstone layers, the maximum thickness of this group being 400 m. The conglomerates are principally composed of Birimian quartz pebbles (>90%) and volcanic clasts (Hirdes and Nunoo, 1994) that host the Tarkwa Placer deposits.

The Banket formation is overlain by approximately 400 m of Tarkwa Phyllites.

The uppermost unit of the Tarkwa super group is the Huni Sandstone, comprised of alternating beds of quartzite and phyllite intruded by minor dolerite sills that form a package up to 1,300 m thick (Pigois et al., 2003). U/Pb and Pb/Pb geochronology dating of detrital zircons provide a maximum depositional age of 2,132 ± 2.8 Ma for the Kawere formation and 2,133 ± 3.4 Ma for the Banket formation (Davis et al., 1994; Hirdes and Nunoo, 1994). These ages agree with the study by Pigois et al. (2003) that yielded maximum depositional age of 2,133 ± 4 Ma from 71 concordant zircons of the Banket formation. According to all concordant zircon histograms (161 grains) and their uncertainties, a reasonable estimation for the start of the Tarkwaian sedimentation could be as young as 2,107 Ma.

 

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Abundant granites and granitoids intruded the Birimian and Tarkwaian units during the Paleoproterozoic. Eburnean plutonism in south-west Ghana can be divided into two phases between 2,180 to 2,150 Ma (Eoeburnean) and 2,130 to 2,070 Ma (Eburnean) that is supported by the current database of U/Pb and Pb/Pb zircon ages. Most of the granitoids intruded during both phases correspond to typical Tonalite–Trondhjemite–Granodiorite suites. However, in the southern part of the Ashanti Belt, intrusions within the Mpohor complex have granodioritic, dioritic and gabbroic compositions.

Dolerite dykes oriented north-south and east-northeast to west-southwest that are generally less than 100 m in thickness are abundant across the West African craton where they cross-cut Archean and Paleoproterozoic basement. In south-western Ghana these dykes are well defined in magnetic data where they are characterized by strong magnetic susceptibility. Dolerite dykes are observed to cross-cut undeformed K-feldspar rich granites that formed during the late Eburnean, and are overlain by Volta basin sediments with a maximum depositional age of 950 Ma (Kalsbeek et al., 2008). These relationships constrain dyke emplacement to between 2,000 Ma and 950 Ma. In contrast some older dolerite/gabbro dykes and sills were deformed during the Eburnean orogeny and are dated at 2,102 ± 13 Ma (U/Pb on zircon, Adadey et al., 2009).

With the exception of some late Eburnean granitoids, dolerite dykes and Phanerozoic sediments, all other lithologies have undergone metamorphism that generally does not exceed upper greenschist facies. Studies on amphibole/plagioclase assemblages suggest the peak temperature and pressure was 500 to 650°C and 5 to 6 kbar (John et al., 1999), dated at 2092 ± 3 Ma (Oberthür et al., 1998).

7.2Local Geology and Mineralization

The Wassa property lies within the southern portion of the Ashanti Greenstone Belt along the eastern margin of the belt within a volcano-sedimentary assemblage located at proximity to the Tarkwaian basin contact. The eastern contact between the Tarkwaian basin and the volcano-sedimentary rocks of the Sefwi group is faulted, but the fault is discrete as opposed to the western contact of the Ashanti belt where the Ashanti fault zone can be several hundred meters wide.

Deposition of the Tarkwaian sediments was followed by a period of dilation and the intrusion of late mafic dykes and sills.

The lithologies of the Wassa assemblage are predominantly comprised of mafic to intermediate volcanic flows which are interbedded with minor horizons of volcaniclastics, clastic sediments such as wackes and magnetite rich sedimentary layers, most likely banded iron formations. The volcano-sedimentary sequence is intruded by syn-volcanic mafic intrusives and felsic porphyries.

The magnetic signature of the Ashanti belt is relatively high in comparison to the surrounding Birimian sedimentary basins such as the Kumasi basin to the west of the Ashanti belt and the Akyem Basin to the East as illustrated in Figure 7-2.

Rock assemblages from the southern area of the Ashanti belt were formed between a period spanning from 2,080 to 2,240 Ma as illustrated in Table 7-1, with the Sefwi Group being the oldest rock package and the Tarkwa sediments being the youngest. The Ashanti belt is host to numerous gold occurrences, which are believed to be related to various stages of the Eoeburnean and Eburnean deformational event. Structural evidences and relationships observed in drill core and pits at Wassa would suggest the mineralization to be of Eoeburnean timing while other known deposits in the southern portion of the Ashanti belt such as Chichiwelli, Benso and Hwini Butre are considered to be of Eburnean age.

The Eoeburnean deformation is best observed at Wassa where the deformational event has produced a penetrative foliation with an associated lineation which is defined by mineral alignments. A period of extension occurred between the Eoeburnean and Eburnean deformational events which resulted in the formation of the Akyem Basin (Kumasi Group) to the northeast of the Wassa Mine and the Tarkwa group to the west of the Wassa concession. Both metasedimentary sequences of the Tarkwa and Kumasi group have not been affected by the penetrative foliation observed at Wassa.

 

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The Eburnean deformation is divided in multiple events which vary in number depending on the authors as summarized in Table 7-1. All deposits underlying the Wassa concession have been affected by the Eburnean deformational events, the main penetrative foliation has been affected by at least three Eburnean folding events which have resulted in a large scale refolded synform. The main foliation is sub-vertical and oriented northeast to south-west on the south-eastern flank of the Wassa mine fold whereas it is dipping at around 45° to the south-southeast on the north-west flank of the Wassa mine fold.

 

 

Figure 7-2 Total magnetic intensity reduced to pole, of the Ashanti Belt (modified from Perrouty et al, 2012)

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Figure 7-3 Compilation of geochronology dating from the Ashanti Belt (Perrouty et al, 2012)

 

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Table 7-1 Deformational history of the Ashanti Belt (Perrouty et al, 2012)

  

In Birimian

Obuasi/Bogoso

Allibone et al, 2002

In Tarkwaian

Tarkwa, Damang

Tunks et al, 2004

Regional

 

Eisenlhor et al, 1992

Regional

Feybesse et al, 2006

Milesi et al, 1992

Eoeburnean

2187 - 2158 Ma

Sefwi Group volcanism and sedimentation

Volcanism,

Granitoids intrusion.

Regional Metamorphism.

 

Birimian sediments and volcanics.

Penecontemporan-eous Plutonism (Dixcove type granitoids).

Magmatic accretion.

Plutonism.

Birimian sedimentation.

D1, N-S shortening

Regional scale folding in the Sefwi Group.

Possible Gold mineralization.

Onset of deformation in a “foreland thrust” and Tarkwaian deposition.

D2, Extension Phase

2154-2125 Ma

Kumasi Group sedimentation

D1

S1 parallel to bedding.

Flat-lying bedding parallel shearing.

Eburnean

2125 –

2000 Ma

Tarkwa Basin Formation

(2107-2097 Ma)

D3, NW-SE shortening

Km scale folds in Birimian and Tarkwaian.

S3 Subvertical crenulation cleavage (NE-SE).

Thrust faults (Ashanti, Damang…)

Metamorphism peak (2092Ma).

D2, NW-SE shortening

Isoclinal folds with axial surface parallel to the regional faults and shear zones.

Ashanti thrust fault.

D1, NW-SE shortening

Km scale folds (with subvertical axial surface, S3).

Damang thrust fault.

D1, NW-SE shortening

S1 (NE-SE) subvertical and subparallel to bedding in both Birimian and Tarkwaian Regional folds (tight to isocline).

D1, NW-SE shortening

Thrust faults.

Tarkwaian sediments deposition, Syn D1.

Metamorphism (6kbar/550-650°C).

D3

Low dip axial surface fold at Obuasi.

S3 crenulation cleavage overprinting S2. Final stage of D2?

 

D2, Continuing compression

S2 (NE-SE) fabrics overprint S1 foliation.

S2 is defined by aligned muscovite and elongate recrystallized quartz grains.

 

Metamorphism.

 

Syncrogenic plutonism (Cape Coast type granitoids).

D2/D3, NW-SE shortening

Tarkwaian folds.

Strike-slip faults and shearing.

Gold mineralization.

Metamorphism (2-3kbar/ 200-300°C).

D4, NNW-SSE shortening

Sinistral shear reactivism D3 thrust.

S4 crenulation cleavage ENE-WSW.

Greenschist retrograde metamorphism.

Remobilization and concentration of gold particles along the shear zone and at the base of the Tarkwa Basin.

D4, NNW-SSE shortening

Hm scale fold at Obuasi.

D2, NNWSSE shortening

Thrust faults and minor folds.

D5 or syn-D4

Sinistral strike-slip faults and shearing.

Gold mineralization.

D5

Recumbant folds <m.

Sub-horizontal crenulation cleavage.

Last pyrite/gold mineralization associated with quartz vein.

 

D3, ESE-WNW shortening

Folds with shallowly dipping axial surfaces and mineralized quartz veins, post-dating peak of metamorphism.

K-rich plutonism

(cross-cutting all previous structures).

Late plutonism.

D6, NE-SW shortening

Low amplitude folds + crenulation cleavage ~N320/70 (RH).

Reverse faults oriented NW-SE.

D4

Faults oriented NW-SE.

 

 

 

 

 

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Figure 7-4 Regional geology of the Ashanti belt, showing Wassa, GSR tenure and major deposits (GSR, 2020)

7.2.1Wassa

The Wassa lithological sequence is characterized by lithologies belonging to the Sefwi Group and consisting of intercalated meta-mafic volcanic and meta-diorite dykes with altered meta-mafic volcanic and meta-sediments which are locally characterized as magnetite rich, banded iron formation like horizons (Bourassa, 2003), as illustrated in Figure 7-5. The sequence is characterized by the presence of multiple ankerite-quartz veins which are sub-parallel to the main penetrative foliation. The lithological sequence is also characterized by Eoeburnean felsic porphyry intrusions on the south-eastern flank of the Wassa mine fold.

 

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Figure 7-5 Wassa mine-scale geology (modified from Bourassa, 2003 and Perrouty et al, 2013)

The first deformational event (D1) at Wassa is of Eoeburnean timing and consists of North-South Shortening. This pre-Tarkwaian event resulted in a penetrative foliation which transposed lithological contacts along this main foliation. Early, gold bearing, syn-D1 quartz-ankerite veins were also formed during the Eoeburnean event.

 

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The second event of deformation (D2) is an extension period with no local deformation at the mine scale at Wassa. Regionally, this event separates the Eoeburnean and Eburnean orogeny by an extension period of approximately 40 Ma which resulted in the sedimentation of the Birimian and Tarkwaian basins.

The Eburnean orogeny is divided in three distinct deformational events, D3 is a Northwest-Southeast shortening event which resulted in the inversion of regional detachment faults into thrust faults. At the mine scale, this event generated a second penetrative foliation at Wassa and a first phase of Eburnean folding. The D4 deformational event, a North Northwest-South Southeast shortening event resulted in the sinistral reactivation of earlier faults at the regional scale and severely buckled the Wassa stratigraphic sequence into moderately steeply dipping, tight fold patterns (F4 Fold) and a third penetrative foliation (S4).The last deformational event, D5, is the result of sub-vertical compression which resulted in open recumbent folds at Wassa and a fourth foliation located in the axial plane of the F5 folds and is generally sub-horizontal, shallowly plunging to the South.

The deposit scale F4 fold is shown on a vertical section through the nose of this structure in Figure 7-6.

The various phases of Eburnean deformations and their effect on the host rocks are illustrated in:

·Figure 7-7:
oTop image shows syn-D1 veins and S1 foliation folded by and F3 fold;
oBottom image shows syn-D1 veins, S1 and S3 foliations affected by a mesoscopic F4 fold;
·Figure 7-8:
oTop image shows syn-D1 veins folded and buckled by S5 foliation; and
oBottom image shows syn-D1 veins, affected by both S4 and S5 foliations.

 

Figure 7-6 Vertical section through Nose of deposit-scale F4 fold, Wassa Main deposit

 

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Figure 7-7 Eburnean folds and foliations from Wassa mine, Starter pit

 

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Figure 7-8 Eburnean folds and foliations from Wassa mine, B-Shoot pit

 

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The Wassa mineralization is subdivided into a number of domains: F Shoot, B Shoot, 242, South East, Starter, 419, Mid-East and Dead Man’s Hill. Each of these represents discontinuous segments of the main mineralized system which extends approximately 3.5 km along strike from surface and is still open at depth.

The SAK deposits are located approximately 2 km to the southwest of the Wassa Main deposit on the northern end of a well-defined mineralized trend parallel to the Wassa Main trend. The SAK deposits are also located on the western side of a major regional east dipping fault which separates the Wassa main mineralization from this trend. The mineralization is hosted in highly altered multi-phased greenstone-hosted quartz-carbonate veins interlaced with sedimentary pelitic units. The SAK mineralization is subdivided into a number of domains as well, SAK 1, 2 and 3, which are thought to be associated with tight F3 fold closures which plunge steeply to the South west.

Mineralization within the Wassa Mine is structurally controlled and related to vein densities and sulphide contents. Higher grade mineralization has been interpreted to be associated with tight isoclinal folding (F3). These tight folds often have extenuated limbs that are weakly mineralized, where as the fold closure was the focal point of remobilized fluids and associated gold. Mineralization in the limbs is generally narrower, < 10 m and fold closer thicknesses and exceed 25-30 m thicknesses. Higher grade mineralization has up and down dip extension of approximately 150 to 200 metres with a down plunge extension being drill tested over 2,000 m from where it daylights in the Starter pit area to section 18,500 mN in the south, where it remains open.

Figure 7-9 illustrates the tightly folded nature of the gold mineralization, as depicted by the black dotted line showing high-grade zones associated with F3 fold closures and subsequent parasitic folding. The mineralization is then subsequently folded by the deposit scale parasitic F4 folds.

Three vein generations have been distinguished on the basis of structural evidence, vein mineralogy, textures and associated gold grades. Evidence further relates the majority of gold mineralization to the earliest recognized vein generation which is believed to be syn-Eoeburnean. Gold grades broadly correlate with the presence of quartz-dolomite/ankerite-tourmaline bearing quartz veins and the presence of sulphide minerals (predominantly pyrite) within and around the quartz veins. Gold grades appear to be spatially restricted to the quartz veins, vein selvages and the immediate wall rocks. The alteration haloes developed around the veins and pervasively developed within the core of the deposit scale Wassa fold contain lower grade mineralization.

 

Figure 7-9 Wassa section through 19,650 mN showing high-grade zones, F3 closures, parasitic folding

 

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The combined and overprinted Eburnean deformational events (D3 to D5) render precise prediction of the vein geometries and localities difficult in areas with wider spaced or little drillhole data. However, where drilling density is tighter (12.5 m x 10 m), as with in the immediate underground mining areas it is possible to construct both hanging and footwall contacts of the economic gold mineralization, Figure 7-10. The higher grade zones of gold mineralization are constrained with in broader lower grade mineralized zones that can be defined reasonably well with the wider spaced surface drillhole data, but to delineate the geometry of the higher grade zones tighter underground grade control drilling is required. Figure 7-11, drill cross section 18900N shows a simpler interpretation which is based on wider spaced surface drilling. As indicated above further infill underground drilling is necessary to delineate the geometry of the high-grade gold mineralization.

Figure 7-10 Wassa section through 19,925 mN showing interpretation with tight-spaced drilling

Figure 7-11 Wassa section through 18,900 mN showing interpretation and wide spaced (surface) drilling

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7.2.2Hwini Butre

The Hwini Butre concession is underlain by three main deposits: Adoikrom, Dabokrom and Father Brown, which are hosted within the Mpohor mafic complex, which consists mainly of gabbroic and gabbro-dioritic intrusive horizons as illustrated in Figure 7-4 Regional geology of the Ashanti belt, showing Wassa, GSR tenure and major deposits. Each of the three deposits have different mineralization styles.

The timing of the mineralization at Hwini Butre is considered to be of late to post Eburnean age with the period of hydrothermal activity likely to have spanned a considerable length of time. At Father Brown and Dabokrom, mineralization is associated with quartz vein systems which are locally surrounded by extensive, lower grade, disseminated quartz stockwork bodies, especially at Dabokrom. The Father Brown deposit is characterized by well-developed fault-filled quartz veins which are, as is the case for Dabokrom, light grey with carbonate and mica accessory minerals and minor tourmaline and feldspar. Wall rock alteration is commonly associated with elevated gold grades and consists of silicification with carbonates, muscovite and sericite. Secondary strain fabrics are also present, with mylonitic and cataclastic fabrics common in the heavily altered zones. Visible gold occurs as disseminations in discrete quartz veins and within zones of silicification associated with pyrite. Gold is medium to coarse grained and generally occurs with pyrite and appears to be free milling. As at Benso, arsenopyrite is largely absent from the Hwini Butre deposits.

At Adoikrom, the mineralization is shear hosted and characterized by the absence of quartz veins; gold is associated with fine grained pyrite and intense potassic alteration. The higher grade core of gold mineralization at Adoikrom is constrained within a moderately plunging, South West trending shoot which has been drilled tested to approximately 1000 meters depth where it remains open.

 

 

 

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Figure 7-12 Hwini Butre section through 33,100 mN

7.2.3Benso

The Benso concession is underlain by four main deposits: Subriso East, Subriso West, G Zone and I Zone. All the deposits are characterized by similar style of mineralization. The Benso deposits are hosted in two dominant rock types. Subriso West and I Zone are hosted within Intermediate feldspar porphyry intrusives and meta-volcanics, where Subriso East occurs along the contact between carbonaceous phyllites and meta volcanics. Mineralization at Benso is associated with late deformational stages of the Eburnean orogeny and deposits are shear hosted along subsidiary structures.

Mineralogy is relatively simple with fine grained but visible gold disseminated in the shear fabric and associated with pyrite which can be locally abundant. Zones of intense alteration with chlorite, carbonates and epidote are common. Arsenopyrite is absent from the deposits.

7.2.4Chichiwelli

The Chichiwelli deposit consists of two sub-parallel mineralized trends which hosts two distinct types of mineralization. The Chichiwelli West trend is a shear zone hosted deposit with a quartz, carbonate, sericite and potassic alteration assemblage, the mineralization is associated with pyrite. The Chichiwelli East trend is a quartz vein associated deposit with an ankerite and sericite alteration assemblage. Mineralization is also associated with pyrite along vein selvages and in the wall rocks.

 

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The lithological assemblage at Chichiwelli West consists of mainly fine to medium grained dioritic intrusives with local intercalation of basalt and feldspar porphyritic intrusives. Lithologies are moderately to strongly foliated adjacent to the shear zone, the mineralization is bounded to the shear zone and associated with a strong shear fabric. The shear zone mineralization is characterized locally by boudinage quartz and calcite stringers with fine disseminated sulphides, mainly pyrite, and associated with a sericite and potassium alteration assemblage with minor silicification. The Chichiwelli East lithological sequence is comprised mainly of deformed diorite with local strain zones. The mineralization is characterized by milky white quartz veins associated with potassium alteration and euhedral coarse grained pyrite.

 

 

 

 

 

 

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8Deposit Types
8.1Wassa

The Wassa deposit is located on the eastern flank of the northeast trending Ashanti Belt, a Paleoproterozoic greenstone belt which was formed and deformed, along with the dividing Birimian and Tarkwaian sedimentary basins during the Eoeburnean and Eburnean orogeny. Most deposits found within the Ashanti belt can be classified as lode gold deposits or orogenic mesothermal gold deposits, with the exception of the Tarkwaian paleoplacer deposits which have a sedimentary origin. Orogenic gold deposits are the most common gold systems found within Archean and Paleoproterozoic terrains, in the West African shield, these deposits are typically underlain by geology considered to be of Eburnean age and are generally hosted by volcano-sedimentary sequences.

B. Dubé and P. Gosselin of the Geological Survey of Canada described these deposits as greenstone-hosted quartz-carbonate vein deposits in the 2007 special publication No. 5 entitled Mineral Deposits of Canada. The authors described these deposits as typically occurring in deformed greenstone belts and distributed along major compressional crustal scale fault zones commonly marking the convergent margins between major lithological boundaries. The greenstone-hosted quartz-carbonate vein deposits correspond to structurally controlled complex deposits characterized by networks of gold-bearing, laminated quartz-carbonate fault-fill veins. These veins are hosted by moderately to steeply dipping, compressional brittle-ductile shear zones and faults with locally associated shallow-dipping extensional veins and hydrothermal breccias. In these deposits, gold is mainly confined to the quartz-carbonate veins but can also occur within iron-rich sulphidized wall rocks or within silicified and sulphide-rich replacement zones.

The Ashanti belt is considered prospective for orogenic mesothermal gold deposits and hosts numerous lode gold deposits and paleoplacer deposits. As illustrated by Figure 7-4, several major gold deposits are found within the Ashanti belt which can be classified into six different deposit types:

·Sedimentary hosted shear zones;
·Fault fill quartz veins;
·Paleoplacer;
·Intrusive hosted;
·Late thrust fault quartz veins; and
·Folded veins system.

The sedimentary hosted shear zone deposits are localized principally along a steep to sub-vertical major crustal structures located along the western margin of the Ashanti belt referred to as the Ashanti trend. The Ashanti trend shows a range of mineralization styles associated with graphitic shear zones, which represents the principal displacement zone of a regional-scale shear zone that defines the mineral belt. These styles include highly deformed graphitic shear zones containing disseminations of arsenopyrite as the principal gold bearing phase and disseminations of sulphides in mafic volcanic rocks generally found in the footwall of the main shear zones. The sedimentary hosted shear zone deposits which occur along the Ashanti trend include Bogoso, Obuasi, Prestea and Nzema.

The second type of deposit found within the Ashanti belt are laminated quartz vein deposits containing free gold. Fault filled quartz vein deposits also occur along the Ashanti trend but are only present at Obuasi and Prestea. The third type of deposit are paleo-placer deposits within the Tarkwaian sedimentary basin which are hosted within narrow conglomerate horizons intercalated with sandstone units characterized by iron oxides cross beddings. Paleoplacer deposits occur in the southern portion of the Tarkwa basin and examples include Tarkwa, Teberebie and Iduaprim. The fourth type of deposit found within the Ashanti belt are intrusive hosted deposits which occur along second order structures such as the Akropong trend in the Kumasi basin and the Manso trend in the Southern portion of the Ashanti belt. These deposits can be hosted both within felsic and mafic intrusives and are characterized by a penetrative fabric where gold is associated with pyrite and arsenopyrite. Examples of such deposits include Edikan and Pampe along the Akropong trend and Benso and Hwini Butre along the Manso trend. The fifth type of deposit found within the Ashanti belt is late thrust fault associated quartz vein deposits. The Damang mine which is located just west of Wassa is the only known thrust fault related deposit in the Ashanti belt. The deposit is characterized by low angle; undeformed extensional and tensional veins associated with low angle thrust faults. This type of deposit contrasts with the last type of deposit found with the belt, the multi-phase folded Wassa vein deposit. The Wassa mineralization consists of greenstone-hosted, low sulphide hydrothermal deposits where gold mineralization occurs within folded quartz-carbonate veins, as illustrated in Figure 8-1. The Wassa deposit can therefore be classified as an Eoeburnean folded vein system and is the only such deposit recognized to date within the Ashanti belt.

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Host rocks in the Wassa mine area have been affected by at least four phases of ductile deformation, producing a polyphase fold pattern at the mine scale. Discrete high-strain zones locally dissect this fold system. The structural history of the Wassa area is important in that the various deformational events have been responsible for the emplacement of the gold mineralization as well as the geometry of the zones themselves. Mineralized zones at the Wassa Mine are related to vein swarms and associated sulphides that formed during the Eoeburnean deformational event. All rock types underlying the Wassa Mine appear to be altered to variable degrees with the most common alteration consisting of a carbonate-silica-sulphide assemblage.

 

Figure 8-1 Syn-Eoeburnean veins from B-Shoot, 242 and South-east zones (modified from Perrouty et al, 2013)

8.2Hwini Butre

The Hwini Butre deposits can be characterized as mafic intrusive hosted, orogenic shear zones. The deposits are hosted within diorite and granodiorite intrusive rocks of the Mpohor complex. The Father Brown deposit is characterized by well-developed fault-filled quartz veins (Figure 8-2), whereas the Adoikrom deposit is a shear zone hosted deposit characterized by intense potassium and silica alteration assemblage (Figure 8-3).

Analysis of geophysical surveys and topographical features have identified several north to north-northeast trending regional features running through the area which are tentatively interpreted as boundary faults along the margins of the Ashanti Belt. The Mpohor complex exhibits the underlying north-south trends but also has extensive cross cutting features present particularly in the north-west orientation. These structural features are second order or subsidiary structures splaying from primary structures.

 

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The Adoikrom, Father Brown and Dabokrom deposits occur in the south portion of the Mpohor complex and appear to be controlled by a series of shallow to moderately dipping faults and shear structures with dips varying from 20° to the south at Dabokrom and steepening to 65° to the northwest at Adoikrom. 

 

Figure 8-2 Mineralization exposure in Father Brown pit, smoky quartz vein 

 

 

Figure 8-3 Mineralization exposure in Adoikrom pit, potassic alteration

 

 

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8.3Hwini Butre

The Benso deposits can also be characterized as mafic intrusive hosted, orogenic shear zones deposits, which are hosted by Birimian metavolcanics into which coarse plagioclase porphyry units have intruded and are generally conformable with the volcaniclastic units.

At Subriso East, the metavolcanics host complex quartz vein systems associated with intense shearing and abundant sulphide mineralization (Figure 8-5). At Subriso West, the presence of intermediate porphyry intrusive appears to play a more significant role (Figure 8-4) and quartz veining is less extensive and broad scale silicification is more common. The contacts between metavolcanics and porphyry have been identified as potential targets for higher grade gold mineralization.

The mineralization hosting structures generally dip steeply towards the west with foliation generally parallel to the bedding. The aeromagnetic interpretation reveals a north to north-northeast striking fault system along the course of the Ben River with several other fracture systems also evident with strikes varying between the northwest and northeast. The Subriso East deposit is interpreted to dip less steeply to the west at approximately 50°.

Oxidation associated with weathering is variable but generally limited. The weathering forms a layer of lateritic clay rich material grading into a soft saprolite. The vertical depth is generally 10 m or less but can reach depths of 30 m in places. There is a sharp boundary between oxide and fresh material with a narrow and poorly developed transition zone.


Figure 8-4 Mineralization exposure in Subriso West pit, sheared volcanics

 


Figure 8-5 Mineralization exposure in Subriso East pit, fine grained pyrite

 

 

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8.4Chichiwelli

The Chichiwelli deposits can also be characterized as mafic intrusive hosted, orogenic shear zones, the deposits are hosted within diorite and granodiorite intrusive rocks. The mineralization zones at Chichiwelli are similar to those observed at Benso, with the mineralized hosting structures generally dipping to the east.

The Chichiwelli deposit consists of two sub-parallel mineralized trends which hosts two distinct types of mineralization, as shown in Figure 8-7 and Figure 8-6. Mineralization at the Chichiwelli West zone is shear zone hosted with a carbonate, sericite and potassic alteration assemblage, while mineralization along the Chichiwelli East trend is quartz vein associated with an ankerite and sericite alteration assemblage. Mineralization is spatially associated with pyrite at both deposits.

 

 

Figure 8-6 Mineralization at Chichiwelli East, hydrothermal veins

 

 

Figure 8-7 Mineralization at Chichiwelli West, shear hosted

 

 

 

 

 

 

 

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9Exploration

Extensive exploration work has been conducted on and around the Wassa concession. Previously, several airborne and ground geophysical surveys consisting of aero-magnetics, radiometrics and Induced Polarization (IP) were conducted on the properties. The geophysical surveys targeted geochemical anomalies, which had previously been identified following multiple stream and soil geochemical sampling programs.

9.1Wassa

Modern exploration programs on the Wassa concession began in the early 1990s with satellite imagery and geophysical surveys which identified geophysical lineaments and anomalies over small scale and colonial mining areas. Stream and soil geochemistry sampling programs were conducted over the geophysical anomalies and identified two linear gold in-soil anomalies as illustrated in Figure 9-1.

 

Figure 9-1 Wassa soil geochemistry and anomalies (GSR, 2018)

Exploration drilling commenced in February 1994 and, by March 1997, a total of 58,709 m of RC and DD had been completed. In September 1997, consulting engineers Pincock, Allen and Holt completed a FS. Only minimal exploration work was conducted by SGL between the completion of the FS in 1997 and the 2001 bankruptcy.

In March 2002, GSR started an exploration program as part of a due diligence exercise following the ratification of a confidentiality agreement with the creditor of SGL. The exploration program consisted mainly of pit mapping and drilling below the pits to test the continuity of mineralization at depth. The concession was acquired later that year by GSR following the completion of the due diligence exercise. Exploration drilling resumed in November 2002 under GSR with the aim to increase the quoted reserves and resources for the feasibility, which was completed in 2003.

Simultaneously to the resource drilling program that targeted resource increases in the pit areas, GSR also undertook grass roots exploration along two previously identified mineralized trends. The 419 area was located south of the main pits and the SAK anomaly was a soil target that had never been previously drilled and was located west of the main pits. Deep auger campaigns were also undertaken in the Subri forest Reserve, which is located in the southern portion of the Wassa Mining lease.

 

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Figure 9-2 Wassa airborne magnetic coverage (GSR, 2004)

In March and April 2004, a high resolution, helicopter geophysical survey was carried out over the Wassa Mining Lease and surrounding Prospecting and Reconnaissance Licenses (Figure 9-2). Five different survey types were conducted, namely: Electromagnetic, Resistivity, Magnetic, Radiometric and Magnetic Horizontal Gradient. The surveys consisted of 9,085 km of flown lines covering a total area of 450 km2. Flight lines were flown at various line spacing varying between 50 to 100 m depending on the survey type. The geophysical surveys identified several anomalies with targets being prioritized on the basis of supporting geochemical and geological evidence.

 

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The exploration program in 2005 continued to focus on drill testing anomalies identified by the airborne geophysical survey as well as infill drilling within the pit area to expand the reserve and resource base. The resource definition drilling program focused mainly on SAK, South-East and the 419 area. The following years were subject to more infill and resource definition drilling in the pit areas at Wassa. In 2011, exploration drilling programs shifted towards drilling deep HG targets below the pits; this drilling continued until 2015. Drilling was limited in 2016 with rigs in filling the first planned stoping areas to increase confidence in the resource prior to underground mining. The 2017 drilling programs were two-fold, infilling gaps in the previous drilling with in the proposed expanded open pit as well as testing the B shoot underground mineralization both north and south, up and down plunge respectively. The southern extension drilling initiated in 2017 continued into 2019 and utilized larger drill rigs to conduct directional wedging and downhole motor work to delineate the deeper southern extensions of B and F shoot HG mineralization.

9.2Hwini Butre

GSR acquired SJR and the Hwini Butre concession in late 2005 and commenced exploration work in early 2006. GSR exploration activities concentrated on the previously defined mineralization at Adoikrom North, Adoikrom, Dabokrom and Father Brown. The drilling program focused mainly on infill drilling and extending the continuity of the deposits at depth. The previous drilling by SJR reached a maximum vertical depth of approximately 130 m, whereas GSR extended the modelled mineralization at vertical depths of over 250 m.

GSR also undertook regional exploration programs over the concession by targeting a number of geochemical and geophysical anomalies previously identified by SJR, these anomalies were mainly tested by use of rotary air blast drilling. A combination of 4 m deep auger and shallow auger at a grid spacing of 400 m by 50 m was also carried out to further test the existing gold in soil anomalies and gaps in the geochemistry sampling over the Hwini Butre concessions.

In 2007 and 2008, GSR focused its Hwini Butre exploration activities on the northern portion of the concession where several colonial gold occurrences such as Breminsu, Apotunso, Abada, Whinnie and Guadium are located. Previous soil sampling in these areas identified several anomalies and the follow up programs included deep auger and rotary air blast drilling. A total of 1,384 auger holes and 41 RAB holes totalling 725 m were completed.

In 2009, 5,992 m RC (83 holes) and 2,100 m DD (21 holes) were completed on the Hwini Butre property (Father Brown, Adoikrom and Dabokrom) to test the strike extensions of the zones and also upgrade the existing quoted Mineral Resource. The drilling program also identified potential underground target beneath the Subriso West pit. Induced Polarization geophysical surveys were conducted over the Hwini Butre and Benso concessions in 2009. The program generated targets that were coincidental with lithological trends and gold in soil anomalies.

The resource definition drilling program continued in 2010 at Father Brown, Adoikrom and Dabokrom where 5,075 m of RC drilling (72 holes) and 5,207.3 m of DD drilling (24 holes) were completed. The drilling program also tested the underground potential of the deposits with significant success. A deep auger program totalling 746 m over 205 holes to test IP geophysical anomalies at Essaman was also completed.

In 2011 the deeper targets at Father Brown and Adoikrom were tested to evaluate the underground potential of the deposits. In all, 13 DD holes totalling 3,689.6 m were drilled at Father Brown and Adoikrom. RAB drilling, totalling 2,941 m (174 holes) were undertaken at Semkrom to test IP and aeromagnetic/radiometric anomalies. In 2012, exploration concentrated on Father Brown and Adoikrom infill and step out underground drilling program, with 33 DD holes totalling 10,094 m being completed. In 2018, exploration drilling resumed at Father Brown and Adoikrom to continue evaluating the underground potential. The program combined RC and DD holes totalling 8,236.2 m. The 2018 drilling programs rolled over into 2019 where another 28 holes were completed totaling 14,526.9 m (RC and DD ).

 

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9.3Benso and Chichiwelli

GSR acquired the Benso and Chichiwelli concessions in late 2005 and commenced exploration work in early 2006, with exploration activities focusing on the previously defined mineralization at Subriso East, Subriso West, I Zone and G Zone. The drilling program focused mainly on infill drilling and extending the continuity of the deposits at depth. The 2006 exploration program was also the focus of regional exploration programs over the concession by targeting a number of geochemical and geophysical anomalies previously identified by SJR, these anomalies were mainly tested by use of rotary air blast drilling. A combination of 4 m deep auger and shallow auger at a grid spacing of 400 m by 50 m was also carried out to further test the existing gold in soil anomalies and gaps in the geochemistry sampling over the Hwini Butre concessions.

Exploration on the Benso property in 2007 and 2008 concentrated on drill testing new zones of mineralization delineated by the RAB drilling in 2006. A total of 81 holes and 10,232.3 m of RC and DD drilling was completed at Subriso East, Subriso West, G Zone and I Zone. At Amantin, follow-up programs included deep auger sampling on a 200 by 50 m grid and RAB drilling was undertaken to test the previously defined soil anomalies. A total of 3,717 m of RAB drilling from 178 holes and 1,683.9 m of deep auger drilling over 487 holes were completed at Amantin.

The 2009 exploration program at the Benso concession focused on resource delineation and definition drilling at the Subriso East, Subriso West and G Zone deposits. A total of 3,159 m RC (35 holes) and 2,538.4 m DD were completed. IP geophysical surveys were conducted over the Benso concessions in 2009 and the program generated targets that were coincidental with lithological trends and gold in soil anomalies.

The 2010 exploration activities at Benso included the continuation of the resource delineation and definition drilling in and around the pits and also drilling off the potential underground target at Subriso West. A total of 8,815 m RC (112 holes) and 8,286.2 m DD (18 holes) were completed. A deep auger program totalling 1,114 m over 319 holes was undertaken to test IP targets at Subriso West.

In 2011, 12 DD holes, totalling 4,557 m, were drilled on the Benso property at Subriso West to close up the spacing along strike and down dip of the HG zone of mineralization intersected beneath the pit. At Amantin, a shallow RC program totalling 1,177 m (22 holes) was completed to follow up on widely spaced RAB and RC intersections from earlier drilling programs. A deep auger (6 m) program totalling 907.5 m from 174 holes were completed at K Zone and I Zone to test additional targets generated by IP survey program.

Exploration activity at Benso in 2012 was limited to structural interpretation of the controls on mineralization to determine the underground potential at Subriso West.

 

 

 

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10Drilling

Wassa is an advanced property so details of all drill results are not required in this report. This section provides an overview of drilling and representative plans and cross-sections are shown in Section 7.2.

10.1Surface Drilling

Drilling is carried out by a combination of DD, RC and RAB techniques. In general, RAB is used at early stages for follow up to soil geochemical sampling and, during production, for testing contacts and mineralization extensions around the production areas. RAB has a maximum drilling depth of 30 m.

RC pre-collar with diamond core tails drilling is used as the main method for obtaining samples for Mineral Resource estimation and is carried out along drill lines spaced between 25 and 50 m along prospective structures and anomalies defined from soil geochemistry and RAB drilling. RC drilling is typically extended to depths of in the order of 100-125 m. The DD method is used to provide more detailed geological data and where more structural and geotechnical information is required. Generally, the deeper intersections are also drilled using DD and, as a result, most section lines contain a combination of RC and DD drilling.

RC and DD drilling was conducted with a GSR geologist was on site to align the drill rig and check the drill head dip and azimuth. Downhole surveying was conducted using a single shot camera, for RC and DD holes at the bottom of holes exceeding 30 m depths and then taken progressively every 30 m up hole. The single shot camera recorded the dip and azimuth for each surveys which was validated and recorded by the GSR geologists or was recorded by a Reflex survey instrument and captured in the database as well as being filed in the respective drillhole file folders on site.

Drilling depths at Wassa Main have generally been less than 250 m but with the discovery of higher grades below the Wassa Main pit in late 2011, hole depths have increased. In the 1st half of 2014, two gyro survey instruments were utilized to resurvey several of the deeper holes. In total, 153 holes, drilled during 2012 to 2014, were resurveyed. The gyro survey readings were conducted every 10 m both in and out of the hole and the values were then averaged. The 153 gyro surveyed holes were updated in the database and subsequently used for the resource estimates. The gyro surveys showed that there was some deviation in the holes below 250 m drilled depth. Deviations varied from location to location depending on drill orientation with a general tendency for the hole to steepen and swing to the north.

Drilling of the deeper targets at Wassa has required the use of directional drilling methods. The deeper holes, often exceeding 1000 meters, are drilled from surface using HQ sized core and this initial hole (referred to as the “mother” hole) is drilled to the depth where the first directional hole would be started. The directional hole (or “daughter” hole) is drilled using a smaller core size, NQ and is deviated from the mother hole initially using a casing wedge which is oriented in the direction of the mineralized target. Once the initial deflection has been achieved with the wedge, the hole deviation can be controlled using a down hole directional motor which can change the dip and azimuth of the hole by approximately plus or minus 1.5 degrees over a 10-metre run. The direction of the hole can also be controlled by using various combinations of down hole stabilizers and drill bits. The step out deeper drilling fences typically involve two mother holes with three to four daughter holes from each of these. The deeper holes are surveyed, down hole with either a Reflex multi-shot or gyro survey instrument. The surveys are taken while the hole is being drilled as well as every 10 to 15 meters from the bottom of the hole once it has been completed.

Exploration data used in the Long-Range model for the Mineral Resource is summarized in Table 7-1.

The majority of the drilling has been conducted by GSR, although there are some drillholes completed by previous concession owners that have been used to inform the Wassa long-range model (by SGL) and Hwini Butre and Benso models (by SJR). Where drill data by prior ownership is used the data has be validated and checked to the satisfaction of the QP for inclusion to inform interpretation and grade estimates.

 

 

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Table 10-1 Exploration data used for Mineral Resource models

Data SourcePurposeNo.  HolesDrill Metres
 Grade Control (RC)24,957642,470
Pre-Existing (Dec18 Report)Exploration (RC, DD)3,422500,282
 UG Operational (DD)84793,896
 Grade Control (RC)41112,142
2019-2020Exploration (RC, DD)5948,036
 UG Operational (DD)37156,914
Total 30,0671,353,740

All drillhole collars were surveyed using a Nikon Total Station (DTM-332) or Sokkia Total Station by a GSR surveyor. Individual RC and DD holes are identified and marked in the field with poly-vinyl chloride (PVC) pipes. RAB drill holes were surveyed in the field and identified and marked with wooden pegs.

10.2Underground Drilling

Underground diamond drilling is performed using electric-hydraulic diamond drills utilizing the underground mine’s 1,000 V power supply. Core drilled underground is HQ (63.5mm), NQ (47.6 mm) or NQ2 (50.6 mm) in core size. The final drilling density for classification as Measured Mineral Resource is designed to be 15 m along strike and 13 m down dip, or tighter. With the orebody generally striking north-south (on the mine grid), typical drilling azimuths range +/-30 degrees each side of 090˚ or 270˚ azimuth, depending on whether the drills are set up on the hangingwall or footwall side of the orebody. Dips generally range between +30˚ to -60˚.

Downhole surveying is conducted using a Reflex multi-shot downhole surveying tool. When collaring, a single survey is taken at 10-12m depth. At the first survey, the drill hole orientation must fall within ±2o azimuth and ±1.5o dip tolerance, when compared to design. For any hole where the first survey falls outside of tolerance, the geologist has the discretion to either terminate the drill hole and re-collar at the drilling company’s expense, or to continue the hole. At the completion of the drill hole, multi-shot surveys are collected at 15 m intervals on the way out. All downhole surveys are collected by the underground mine geologists. The drilling crews do not perform the surveys themselves.

Drill hole collar locations are captured by the underground mine surveying team. The surveyors use either a Leica TS15 total station, or a Leica TS16 total station to record the collar position in X, Y, Z location. The total station is accurate to less than two seconds in azimuth. In cases where the mine surveyors cannot identify the drill hole collar site, the designed collar coordinates are recorded in the databases.

The Short-Range grade model used for calculation of the December 2020 Mineral Resource estimate, within the active mining area, was completed in December 2020 and utilized 273 additional underground holes totaling 34,275 meters.

10.3Sampling

A standard approach to drilling and sampling on all GSR projects in Ghana. Sampling is typically carried out along the entire mineralized drilled length.

Sample recovery is good across all deposits drilled to date. Ground conditions are generally good and air drilling techniques (AC/RAB and RC) are avoided below the groundwater table where DD is applied.

For RC drilling, samples are collected every 1 m. Where DD holes have been pre-collared using RC, the individual 1 m RC samples are combined to produce 3 m composites which are then sent for analysis. Should any 3 m composite sample return a significant gold grade assay, the individual 1 m samples are then sent separately along with those from the immediately adjacent samples.

 

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10.3.1Diamond Drillholes

DD samples are collected, logged and split with a diamond rock saw in maximum 1.2 m lengths. The core is cut according to mineralization, alteration or lithology. The core is split into two equal parts along a median to the foliation plane using a core cutter. The sampling concept is to ensure a representative sample of the core is assayed. The remaining half core is retained in the core tray, for reference and additional sampling if required.

10.3.2RC Drillholes

RC sampling protocols were established in 2003. The composite length of 3 m has been established to allow a minimum of at least two composites per drillhole intersection based on experience from exploration drilling and mining. The hangingwall and footwall intersections can generally be easily recognized in core from changes in pyrite content and style of quartz mineralization.

The 3 m composite sampling methodology is:

·A sample of each drilled meter is collected by fitting a plastic bag on the lower rim of the cyclone to prevent leakage of material;
·Bag is removed once the “blow-back” for the meter has been completed and prior to the commencement of drilling the subsequent meter;
·Both the large plastic sample bags and the smaller bags are clearly and accurately labelled with indelible ink marker prior to the commencement of drilling. This is to limit error and confusion of drilling depth while drilling is proceeding;
·3 m composite samples are taken by shaking each of the 1 m samples (approximately 20 kg) and taking equal portions of the 3 consecutive samples into a single plastic bag to form one composite sample (approximately 3 kg);
·Composite samples are taken using tube sampling, which uses a 50 mm diameter PVC tube which has been cut at a low oblique angle at one end to produce a spear of approximately 600 mm length;
·The technique assumes that a sample from the cyclone is stratified in reverse order to the drilled interval. A representative section through the entire length of the collected sample is considered to be representative of the entire drilled interval;
·PVC tube is shuffled from the top to bottom of the sample, collecting material on the way. The “shuffling” approach ensures sample accumulated in the tube does not just push the remaining sample away; and
·Material in the tube is emptied into the appropriately labelled sample bag and in the case of 3 m composite samples, stored separately from the 1 m samples.

The 1 m sample collection methodology is:

·1 m re-sampling of selected mineralized composite zones using the 20 kg field samples is undertaken with a single stage riffle splitter;
·Splitter is clean, dry, free of rust, and damage is used to reduce the 20 kg sample weight to a 3 kg fraction for analysis;
·Care is taken to ensure that the sample is not split when it is transferred to the splitter, and is evenly spread across the riffles;
·When considered necessary, the sample is assisted through the splitter by tapping the sides with a rubber mallet;
·Excessively damp or wet samples are not put through the splitter, but tube-sampled or grab-sampled in an appropriate manner. Alternatively, the sample is dried before splitting. A common sense approach to wet sampling is adopted on a case by case basis;

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·Clods of samples are not forced through the splitter, but apportioned manually in a representative manner; and
·Splitter is thoroughly cleaned between each sample using a brush. Where possible, the splitter is cleaned using an air gun attached to the drill rig compressor.
10.3.3RAB/AC Drilling

RAB and Air Core (AC) drilling is used for exploration but is not used to inform any of the current Mineral Resource estimates.

RAB and AC samples are collected and bagged at 1 m intervals. As the samples are generally smaller in size than the RC samples, 3 m composites are prepared by shaking the samples thoroughly to homogenize the sample, before using the PVC tube to collect a portion of the three individual 1 m samples. After positive results from the 3 m composites, the individual 1 m samples are split to approximately 2 to 3 kg using the Jones riffle splitter and then submitted to the laboratory for analysis.

 

 

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11Sample Preparation, Analyses and Security

The measures implemented by GSR related to sample preparation, analysis and security are considered by the Qualified Person to be consistent with standard industry practice and of sufficient quality to include in the estimation of Mineral Resources.

11.1Sample Preparation

Sample preparation on site is restricted to core logging and core cutting or RC and RAB sample splitting. The facilities consist of enclosed core and coarse reject storage facilities, covered logging sheds and areas for the splitting of RC and RAB samples. Sub-sampling of RC and RAB samples is carried out using a Jones Riffle splitter.

11.2Sample Dispatch and Security

Samples are collated at the mine site after core cutting or sample splitting and then transported to the primary laboratory for the completion of the sample preparation and chemical analysis. Samples are trucked by road to the laboratories in Tarkwa.

Sample security involves two aspects, namely, maintaining the chain of custody of samples to prevent inadvertent contamination or mixing of samples, and rendering active tampering of samples as difficult as possible.

The transport of samples from site to the laboratory is by road using a truck dispatched from the laboratory. As the samples are loaded, they are checked and the sample numbers are validated. The sample dispatch forms are signed off by the driver and a company representative. The sample dispatch dates are recorded in the sample database as well as the date when results are received.

No specific security safeguards have been put in place by GSR to maintain the chain of custody during the transfer of core between drilling sites, the core library, and sample preparation and assaying facilities. Core and rejects from the sample preparation are archived in secure facilities at the core yard and remain available for future testing.

11.3Laboratory Procedures

Sample assays have been performed at either the Wassa Site Lab, SGS or Intertek (formerly named TWL). Both commercial labs are located at Tarkwa. GSR submits quality control samples to each lab for testing purposes.

Both SGS and Intertek laboratories are independent of GSR and are accredited for international certification for testing and analysis.

·SGS, Minerals Division – Tarkwa: ISO 17025 and ISO 9001; and
·Intertek Minerals Ltd, Tarkwa: ISO/IEC 17025.

The sample preparation and analysis processes at the Wassa Site Laboratory (WSL), Intertek, and SGS differ slightly. WSL was used as the primary laboratory for 3 m composite and grade control RC drill samples from July 2007 onwards. The laboratory had previously operated as a metallurgical sample processing laboratory at the Wassa mine site.

11.3.1Wassa Site Laboratory

The sample preparation and analysis process at the WSL is as follows:

·Sample reception, sorting, labelling and loading;
·Dry entire sample (3 kg) at 110°C for between 4 and 8 hours;
·Jaw crush entire sample to 3 mm, and secondary Keegor crusher to 1 mm;
·Split 3 kg sample and pulverize for 3 to 8 minutes to 95% passing 75 µm;
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·Sample homogenization using a mat rolling technique, and sub-sample 1 kg into bulk leach extractable gold (BLEG) roll bottle;
·Bottle roll for 6 hours with LeachWellTM accelerant. Allow to settle for 30 to 60 minutes;
·Filter 20 ml aliquot from bottle;
·Di-isobutyl Ketone extraction and atomic absorption spectroscopy (AAS) determination of gold content; and
·1 in 10 residue samples are retained for gold determination using fire assay.
11.3.2Transworld/Intertek

TWL (now Intertek) was the primary laboratory for core samples until July 2007, when it was discontinued due to the following issues:

·Contamination due to poor dust control in pulverizing area of the laboratory. Use of dust attracting cloth gloves for sample handling. BLEG aliquot preparation area containing dirt and liquids, which may result in sample cross-contamination.
·Large fluctuation in employee numbers (60 to 180), which resulted in a risk of training and quality control issues when increasing employment numbers over a short period of time.
·The use of a manual data tracking and capture system, which increased risk of data entry errors. GSR considered this to be a sub-optimal process for a commercial laboratory.

The sample preparation and analysis process used by TWL is illustrated in Figure 11-1.

 

Figure 11-1 Transworld Laboratories sample processing flow sheet 

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11.3.3SGS Tarkwa

The SGS laboratory (Tarkwa) was used for drill core samples from July 2007, to August 2017, with the sample preparation and analysis process as follows:

·Sample received, entered in LIMS, worksheets, printed and samples sorted;
·Samples emptied into aluminium dishes;
·Dry entire sample at between 105 and 110°C for 8 hours;
·Jaw crush entire sample to 6 mm;
·Split sample using a single stage riffle splitter, to result in a 1.5 kg sub-sample;
·Pulverize sub-sample for 3 to 5 minutes, to give 90% passing 75 µm;
·Sample homogenization using a mat rolling technique, and put 1 kg of sample into the BLEG roll bottle;
·Remaining sample is retained as pulp and crushed sample duplicates;
·Bottle roll for 12 hours with LeachWellTM accelerant. Allow to settle for 2 hours;
·Filter 50 ml of aliquot; and
·Di-isobutyl Ketone and AAS for gold grade determination.

During 2017, GSR discontinued using SGS laboratories and began shipping samples to Intertek Laboratories. The Intertek lab sample flow sheet is shown in Figure 11-2. The reason for the change was poor sample result turn-around time. Since the prior issues with Transworld/Intertek, ownership of TWL had changed to Intertek who had implemented internationally recognised standards with changes in management and procedures.

Figure 11-2 Intertek sample processing flow sheet

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11.4Quality Control and Quality Assurance

Quality control measures are set in place to ensure the reliability and trustworthiness of assay data, and to ensure that it is of sufficient quality for inclusion in the subsequent Mineral Resource estimates. Quality control measures include written field procedures and independent verifications of aspects such as drilling, surveying, sampling and assaying, data management and database integrity. Appropriate documentation of quality control measures and analysis of quality control data are an integral component of a comprehensive quality assurance program and an important safeguard of project data.

The field procedures implemented by GSR are comprehensive and cover all aspects of the data collection process such as surveying, drilling, core and RC cuttings handling, description, sampling and database creation and management. At Wassa, each task is conducted by appropriately qualified personnel under the direct supervision of a qualified geologist. The measures implemented by GSR are considered to be consistent with industry best practice.

The quality controls employed by GSR include:

·Field duplicates used to check sampling precision and deposit variability. Two separate samples are collected at the drill site and bagged separately from which two individual samples are produced. The results of these checks can be useful in highlighting natural variability of the grade distribution.
·Pulp duplicates used as a check of sampling precision and coarse gold in pulps. Two separate pulp samples are prepared from a single coarse reject after sample splitting and on site preparation. The results are useful in indicating problems with sample preparation and splitting.
·Repeats as a check of analytical precision and coarse gold. Two separate aliquots are prepared from separate samples taken from the original coarse reject and the two samples results are compared.
·Blanks for highlighting contamination problems and cross labelling when samples are mislabelled in the laboratory.
·Standards as a check of analytical precision and accuracy.

GSR relies on both the laboratory operators QA/QC processes for assaying, as well as GSR’s own independent QA/QC program. The GSR program includes inserting blanks, certified reference materials (otherwise known as standards), and pulp or coarse reject duplicates into sample batches, before sample submission to the lab. GSR also provides sample dispatch lists to the laboratories, to ensure that all samples dispatched from site are received at the lab.

GSR has supplied QA/QC reports to various consultants over the numerous drilling campaigns since 2004, and a summary of the historical and current QA/QC results is included here.

11.4.1Comparison of Assay Methodologies

In 2003, during open-pit operations, it was recognized that there was a variance between primary and duplicate assay grades of the same sample, as well as a variance between the planned mine grade to the mill reconciled grade. The conventional 50g fire assay being used at the time displayed poor reproducibility between field duplicates. This effect was also evident between pulp duplicates; although not as marked. The conclusion was that a component of coarse gold was present in the samples, and contributed to poor reproducibility between samples. It was recommended to switch to an analytical process that made use of significantly larger sample masses, such as LeachWell™ assays.

To address this, GSR changed the assay procedure from the 50 gram fire assay method to a 1kg BLEG assay, with a LeachWellTM accelerant. Gold grade was determined using an AAS finish. Initially, samples were split by a rotary splitter and leached for six hours. Following the analysis of the leach tailings, the leach time was extended to 12 hours.

Due to time constraints, the use of the rotary splitter was discontinued and a Jones Riffle splitter was used to split sub-samples from the larger RC drill hole samples. The difference between the reproducibility of fire assay versus larger BLEG assays is illustrated in Figure 11-3. It shows a significant improvement with respect to sample reproducibility between the fire assay and the BLEG methodologies. Using BLEG, 80% of pairs report Half Absolute Relative Difference (HARD) precisions of less than 17%, compared to the 35% precision attributable to the fire assay method. SRK recommended that GSR continue to monitor the reproducibility of the sample grades from the paired data analysis.

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Figure 11-3 HARD plot comparing fire assay and BLEG for field duplicates

11.4.2Repeat (Coarse Reject) Duplicates 2011 to 2013

From 2011, GSR discontinued the use of pulp samples for determining repeatability. Instead, coarse reject material (leftover material from the laboratory primary crush stage) was used as duplicate sample material.

During the sample prep stage, after the drill core passed through primary crushing, the excess coarse reject material was collected and returned to Wassa. This material was then re-numbered and re-submitted to the laboratory for repeat analysis. Coarse reject duplicates were used to monitor the sample preparation processes of the laboratory.

The HARD plot of all coarse rejects for 2011 is presented in Figure 11-4. The results of this HARD analysis show that approximately 89% of the 369 coarse duplicate samples fall within approximately 20% error and 76% fall within 10% error. This is acceptable for gold deposits of this type.

The HARD plot of all coarse rejects for 2012 is presented in Figure 11-5. The results of this HARD analysis show that approximately 83% of the 2,173 coarse duplicate samples fall within approximately 20% error and 60% fall within 10% error. This is acceptable for gold deposits like Wassa.

The HARD plot of all coarse rejects for 2013 is presented in Figure 11-6. The results of this HARD analysis show that approximately 82% of the 2,962 coarse duplicate samples fall within approximately 20% error and 56% fall within 10% error. This is considered to be acceptable for Wassa.

 

 

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Figure 11-4 HARD plot of all coarse rejects (2011) from SGS

 

 

Figure 11-5 HARD plot of all coarse rejects (2012) from SGS

 

 

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Figure 11-6 HARD plot of all coarse rejects (2013) from SGS

11.4.3QA/QC Data Summary 2014 to early 2017

The analytical quality control data produced between 2014 and early 2017 is summarized in Table 11-1. The data represents approximately 16% of the total number of entire samples for this period.

Table 11-1 Summary of analytical quality control data from 2014 to early 2017

 SGSWGSTotalComment
no.%no.%no.%
Sample Count61,943 96,596 158,539  
Blanks6221.0%6,1596.4%6,7814.28%Coarse Sand
QC Samples4,5647.4%4,3024.5%8,8665.6% 
    ST074/9453575 766 1,341 0.21 g/t
    ST14/9501- 405 405 0.43 g/t
    ST16/9487264 419 683 0.49 g/t
    ST626664 - 664 0.51 g/t
    ST06/948189 280 369 1.02 g/t
    ST06/7384167 - 167 1.08 g/t
    ST588- 763 763 1.60 g/t
    ST39/6373- 168 168 1.67 g/t
    ST602324 - 324 1.91 g/t
    ST482635 516 1,151 1.94 g/t
    ST575476 - 476 2.43 g/t
    G914-214 - 14 2.45 g/t
    ST59661 - 61 2.51g/t
    ST37/6374- 30 30 3.33 g/t
    ST43/7370- 955 955 3.37 g/t
    G910-312 - 12 4.03 g/t
    ST48/8462175 - 175 4.82 g/t
    ST5171,108 - 1,108 5.23 g/t
GC Field Duplicates6,56710.6%- 6,5674.1% 
Coarse Reject Duplicates--3,8023.9%3,8022.4% 
Total QC Samples11,75319.0%14,26314.8%26,01616.4% 

 

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11.4.4Repeat (Coarse Reject) Duplicates 2014 to October 2020

Coarse reject samples from SGS and Intertek sample splits were re-numbered and re-submitted for repeat analyses. Coarse reject duplicates were used to monitor the sample preparation stage at a laboratory.

Analysis of the HARD plots of coarse reject duplicates processed by SGS and Intertek suggested that approximately 56% to 69% of gold assay samples had a HARD below 10% error. Approximately 77% to 96% coarse duplicate samples fell within approximately 20% error. This variance is typical of coarse reject duplicate pairs in gold deposits; indicating that SGS and Intertek can reasonably reproduce this type of paired data.

The HARD plot of all coarse rejects for 2014 is presented in Figure 11-7. The results of this HARD analysis show that approximately 83% of the 2,145 coarse duplicate samples fall within approximately 20% error and 56% fall within 10% error. This is considered to be acceptable for Wassa.

 

Figure 11-7 HARD plot of all Surface Drilling coarse rejects (2014) from SGS

 

 

 

 

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The HARD plot of all surface drilling coarse rejects for 2015 is presented in Figure 11-8. The results of this HARD analysis show that approximately 88% of the 641 coarse duplicate samples fall within approximately 20% error and 69% fall within 10% error. This is considered to be acceptable for Wassa.

 

Figure 11-8 HARD plot of all Surface Drilling coarse rejects (2015) from SGS

 

The HARD plot of all coarse rejects for 2016 is presented in Figure 11-9. The results of this 83% of the 355 coarse duplicate samples fall within HARD analysis show that approximately 20% error and 61% fall within 10% error. This is considered to be acceptable for Wassa.

Figure 11-9 HARD plot of all Surface Drilling coarse rejects (2016) from SGS

 

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The HARD plot of all surface drilling coarse rejects for 2017 is presented in Figure 11-10. The results show that approximately 85% of the 750 coarse duplicate samples fall within approximately 20% error and 62% fall within 10% error. This is considered to be acceptable for Wassa. 

 

Figure 11-10 HARD plot of all Surface Drilling coarse rejects (2017) from SGS and Intertek

 

The HARD plot of all surface coarse rejects for 2018 is presented in Figure 11-11. The results show that approximately 77% of the 2,399 coarse duplicate samples fall within approximately 20% error and 56% fall within 10% error. This is considered to be acceptable for Wassa.

Figure 11-11 HARD plot of all Surface Drilling coarse rejects (2018) from Intertek

 

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The HARD plot of all Surface drilling coarse rejects for 2019 (plus some 2018 results which were not included in 2018 results) is presented in Figure 11-12. The results show that approximately 84% of the 4,079 coarse duplicate samples fall within approximately 20% error and 64% fall within 10% error. This is considered to be acceptable for a Wassa.

 

Figure 11-12 HARD plot of all Surface Drilling coarse rejects (2019) from Intertek

The HARD plot of Underground drill core coarse rejects values for January to October 2020 is presented in Figure 11-13. The results show that approximately 87% of the 1,280 coarse duplicate samples fall within approximately 20% error and 63% fall within 10% error. This is considered to be acceptable for Wassa.

 

Figure 11-13 HARD plot of all Surface Drilling coarse rejects (2020 Jan-Aug) from Intertek

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The HARD plot of coarse rejects values for the Father Brown and Adoikrom (HBB) drilling for 2018 and 2019 surface drilling is presented in Figure 11-14. The results of this HARD analysis show that approximately 86% of the 946 coarse duplicate samples fall within approximately 20% error and 70% fall within 10% error. This is considered to be acceptable for these deposits. 

 

Figure 11-14 HARD plot of all Surface Drilling coarse rejects for 2018-19 for Father Brown & Adoikrom, from Intertek

11.4.5Certified Reference Material

CRM material (otherwise known as standards) are used to monitor the accuracy, precision, and reproducibility of the assay results. CRM materials were sourced from Geostats Pty Ltd. , and Gannet Holdings Pty Ltd. Although the CRM material can be easily identified by the laboratory, the grade of the standard is difficult to determine due to the large number of different standards used. Standards in use between January 2003 and October 2020 are shown in Table 11-2 through to Table 11-12.

A total of 16,100 standards were submitted to SGS between 2008 and 2017. The standards submitted largely performed within expected ranges and mean grades, similar to the expected values. Results indicate that SGS reported assay values both higher and lower than the certified mean value, with some variation to the detection limit. That said, 96% or more of the determinations typically fell within +/–5% of the mean value. Standards submitted to SGS from 2014 to 2017 performed much better with 100% of the determinations falling within +/–3% of the mean value and 75% falling within +/- 2% of the mean value.

A total of 4,320 standards were submitted to the Wassa site laboratory between 2014 and 2017. Standards analyzed by Wassa site lab performed marginally worse with some individual samples beyond two standard deviations of the expected value. These results could possibly be due to the mislabeling of samples. Due to the lower accuracy of the Wassa site lab, along with the slowing down of open-pit operations, in-house assaying was phased out during 2017.

In 2018, GSR began using Intertek Laboratory, as its primary laboratory. A total of 400 CRM were submitted in 2018, with the UG drill core sample batches, with 89 percent or more of the determinations typically falling within +/–2% of the expected value. Between January, 2019 – October, 2020, a total of 20,053 CRMs were submitted to Intertek with the underground core samples batches, with 77 percent of the determinations typically falling within +/–2% of the expected value.

 

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The surface drilling programs in 2019 submitted an additional 2,584 CRM’s with the RC and diamond core samples. These samples were also submitted to Intertek laboratories in Tarkwa and 100% of the samples returned determinations falling with +/–2% of their certified value.

The HBB drilling program also utilized Intertek Laboratories, with 867 CRM’s being submitted with the RC and core samples sent to the lab. The CRM’s returned 79 percent of the determinations typically falling within +/–2% of the certified value.

In general, the performance of the standards inserted with samples submitted for assaying at SGS, Intertek and Wassa site laboratories is acceptable. The majority of the failures appear to be caused by the mislabelling of samples.

Table 11-2 CRM for 2003-2007 (TWL)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
Gannet A0.221960.220%
Gannet B2.521852.57+2%
Gannet C3.46213.53+2%
Gannet D3.40753.400%
Gannet E2.36772.45+4%
Gannet F0.78470.75-4%
Gannet G3.22823.02-6%
Gannet M1.181591.28+2%
Gannet N0.501710.49-2%

 

Table 11-3 Geostats CRM for 2008-2012 (SGS)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
G901-100.48820.51+6%
G305-30.71140.66-7%
G901-21.70321.54-9%
G906-41.901371.99+5%
G999-42.30362.40+4%
G302-22.44702.50+2%
G901-12.50382.38-5%
G396-92.60292.39-8%
G900-73.191933.22+1%

 

 

 

 

 

 

 

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Table 11-4 Gannet CRM for 2008-2012 (SGS)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
ST07/94530.214760.21+2%
ST14/95010.434470.42-3%
ST16/94870.491100.51+3%
ST16/53570.526540.520%
ST4860.571240.54-5%
ST17/22900.78140.79+2%
ST4811.02321.05+3%
ST06/53561.041151.06+2%
ST3221.04181.07+3%
ST06/73841.0818811.04-4%
ST3841.081731.06-2%
ST39/63731.671171.74+4%
ST09/73821.932051.87-3%
ST4821.946951.98+2%
ST53552.371452.39+1%
ST05/94512.455382.53+3%
ST05/63722.461682.44-1%
ST05/22972.56782.49-3%
ST4862.63492.59-5%
ST10/92983.221323.30+3%
ST37/63743.331293.08-7%
ST43/73703.378343.33+1%
ST53593.911313.97+1%
ST3593.93873.96+1%
ST48/84624.825084.89+1%

 

Table 11-5 Gannet CRM for 2013 (SGS)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
ST07/94530.216450.22+4%
ST14/95010.434020.50+17%
ST06/73841.08391.05-3%
ST4821.945281.99+2%
ST05/63722.466652.48+1%
ST37/63743.335793.29-1%
ST48/84624.821874.89+1%

 

 

 

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Table 11-6 Gannet CRM for 2014-2017 (SGS)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
ST07/94530.215750.210%
ST16/94870.492640.490%
ST6260.516640.50-2%
ST06/94811.02891.03+1%
ST06/73841.081671.05-3%
ST6021.913241.97+3%
ST4821.946352.00+3%
ST5752.434762.440%
G914-22.45142.460%
ST5962.51612.510%
G910-34.03123.96-2%
ST48/84624.821754.91+2%
ST5175.2311085.20-1%

 

Table 11-7 Gannet CRM for 2014 to 2017 (Wassa Site Lab)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
ST07/94530.217660.210%
ST14/95010.434050.430%
ST16/94870.494190.50+2%
ST06/94811.022801.00-2%
ST5881.67631.61+1%
ST4821.945161.95+1%
ST37/63743.33303.31-1%
ST43/73703.379553.360%
ST39/63731.671681.66-1%

 

Table 11-8 Gannet CRM for 2018 (Intertek)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
G913-107.10146.94-2%
G915-39.22238.98-3%
G911-42.45322.450%
G316-75.79225.780%
G314-55.30425.23-1%
G314-36.6876.59-1%
ST5881.6691.600%
ST5752.43402.48+2%
ST37/63743.33223.15-6%
ST43/73703.37343.29-2%
ST73-82811.52951.510%

 

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Table 11-9 Gannet CRM for 2019 Wassa UG (Intertek)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
ST6385.292,6525.14-3%
ST6754.913,3084.99+2%
ST43/73703.371,3903.30-2%
ST5752.431,7172.39-2%
G912-32.12542.06-2%
ST6012.092,7342.06-1%
ST6021.911,0101.97+3%
ST5881.63,4881.59-1%

 

Table 11-10 Gannet CRM for 2020 Jan-Oct, Wassa UG (Intertek)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
ST6385.298535.07-4%
ST6754.913354.99+2%
ST43/73703.374323.33-1%
ST5752.434552.39-2%
ST6012.097072.100%
ST6021.911712.00+5%
ST5881.65471.600%

 

Table 11-11 Gannet CRM for 2019 Wassa surface drilling (Intertek)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
ST5881.604531.58-2%
ST6012.095232.06-1%
ST5752.433882.38-2%
ST43/73703.372573.29-2%
ST6754.915845.02+2%
ST6385.293795.18-2%

 

Table 11-12 Gannet CRM for 2018-2019 Father Brown/Adoikrom surface drilling (Intertek)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
ST5881.601001.57-2%
ST6021.911611.97+3%
ST6012.0982.00-4%
G912-32.101482.05-2%
ST5752.433402.39-2%
ST43/73703.3773.24-4%
ST6754.91985.03+2%
G915-39.2258.72-5%

 

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11.4.6Blanks

Blank samples are routinely inserted into the sample stream to check for possible sample contamination during the preparation and assaying process.

Pre-August 2018, the blank material used by GSR, Wassa consisted of coarse sand. The blank samples sent to SGS laboratory consistently yielded values at or below the detection limit, with zero samples yielding a value over 10 times the detection limit of gold. With no failures, the sample blanks performed extremely well and indicated minimal, if any, sample contamination during assaying. From August 2018 to date, coarse aggregate material (crushed granite) from the Winneba belt has been used as blank material and inserted into the sample stream. These blanks have performed extremely well over the years as a check for cross-contamination during sample preparation at Intertek.

From 2014 to 2017, blank material processed at the Wassa site laboratory performed more poorly, with some samples yielding values close to, or above, 10 times the detection limit of gold. Over time, from 2014 to 2016, the blank samples’ performance noticeably declined. Further investigation of anomalously high values indicated contamination in the sample preparation process in some cases. The Wassa site laboratory was primarily used for assaying open pit grade control samples, with a very limited number (13) of underground diamond drill holes processed by the Wassa site lab during 2017.

External lab analysed blank assay data from 2011 to Oct 2020 included 15,073 assays, all assayed by either SGS or Intertek. Summary statistics for the blank material assays are shown in Table 11-13, Table 11-14 and Table 11-15.

Table 11-13 Blank sample summary statistics 2011 to Oct-2020

Sample TypeYearNumberMinimum
(Au g/t)
Maximum
(Au g/t)
Median
(Au g/t)
Mean 
(Au g/t)
Blanks20112780.010.010.010.01
Blanks20121940.010.270.010.01
Blanks20132100.010.110.010.01
Blanks2014560.010.070.010.02
Blanks2015690.010.030.010.01
Blanks20165530.010.030.010.01
Blanks20179300.010.040.010.01
Blanks20174980.010.030.010.01
Blanks20181,0890.0050.660.010.01
Blanks20197,1770.0050.080.010.01
Blanks20202,0740.0050.080.010.01

 

Table 11-14 Blank sample summary statistics 2019, Wassa surface drilling (Intertek)

Sample TypeYearNumberMinimum
(Au g/t)
Maximum
(Au g/t)
Median
(Au g/t)
Mean
(Au g/t)
Blanks201914190.010.030.010.01
Blanks2020150.010.050.010.01

 

Table 11-15 Blank sample summary statistics 2018-2019 Father Brown/Adoikrom surface drilling (Intertek)

Sample TypeYearNumberMinimum
(Au g/t)
Maximum
(Au g/t)
Median
(Au g/t)
Mean
(Au g/t)
Blanks20195110.010.080.010.01

 

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11.4.7Umpire Laboratory Performance

Laboratory checks are performed to check on the reliability of the primary laboratory. In 2013 and 2014, “round-robin” sample check studies were conducted using SGS, TWL (now Intertek) and the Wassa site laboratory. SGS laboratory was the primary laboratory during this period. No additional studies have been completed since then, but there are plans to implement another program in 2021.

In 2014, 252 quarter core samples were selected from drilling conducted between 2012 and mid-2014. The intersections selected were high-grade intervals which averaged approximately 17 g/t Au. GSR has previously conducted similar quarter core sampling studies on other GSR owned deposits. The repeatability of the original results is often poor due to the change in sample size going to half the volume from the original sample. The 2014 Wassa quarter core sampling study produced the same results, with good repeatability between the original sample and the corresponding coarse sample reject, and much poorer repeatability with the quarter core sample. The average grade for both the original assay and the coarse sample reject duplicate compared well at 17 g/t Au, whereas the quarter core sample was less at 12 g/t Au. However, control sample standards that were submitted with these sample batches consistently showed a negative bias, as seen in Table 11-16, so this can partially account for the lower average. The HARD plots shown in Table 11-17 show the good correlation between the original assay value and the coarse sample reject duplicate, but not when comparing the original assay to the quarter core samples analysed at TWL Laboratory. Although the negative lab bias and the smaller sample volume attributes to poor repeatability, the Wassa deposit has a high nugget gold distribution which alone will result in poor repeatability. The variability of the gold distribution was recognized and GSR has put in sample protocols to help reduce the variability, i.e. larger sample volumes, BLEG leach well analysis.

Table 11-16 Gannet CRM for quarter core sample analysis (Intertek)

Standard

Certified Mean

(Au g/t)

Samples Submitted

(no.)

Mean Assay Grade

(Au g/t)

Laboratory Bias
ST5175.2354.98-5%
ST4821.9491.78-8%
ST16/94870.49140.46-6%

 

Table 11-17 Summary HARD plot results for quarter core sample analysis

LaboratorySamples    (no.)<10% HARD<15% HARD<20% HARDCorrelation Coefficient
Original SGS vs Check SGS25265%81%90%0.94
Original SGS vs Check Intertek25232%45%57%0.60
Check SGS vs Check Intertek25229%44%55%0.45

In 2013, 120 RC samples were split into three samples which were sent to each of the laboratories for gold analysis. The sample batches also contained control samples to monitor the precision of the individual laboratories.

The three laboratories all performed well with the best correlation being between SGS and the Wassa site laboratory. The HARD plots for the laboratory comparisons are shown below in Table 11-18.

Table 11-18 Summary HARD plot results for 2013 round robin program

Laboratory <10% HARD<15% HARD<20% HARDCorrelation Coefficient
SGS vs Wassa 65%84%92%0.97
SGS vs Intertek 68%84%88%0.97
Wassa vs Intertek 71%84%90%0.98

 

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The <20% HARD correlation for all the labs demonstrates how the larger RC chip samples provide a better representation of grade. Approximately 90% of the 120 RC samples submitted for this study show a 20% error, compared to the coarse reject core samples submitted in 2013 and 2014 which show a correlation of approximately 80% of the data set with 20% error.

In 2012, a “round-robin” exercise was undertaken to check the reliability of Au assay results from the primary laboratory, SGS. A total of 10% of all assays from the 1m samples received each month were randomly picked for reanalysis. The data was grouped into six separate ranges, namely 0.00 to 0.50 g/t, 0.50 to 0.90 g/t, 0.90 to 1.20 g/t, 1.20 to 2.00 g/t, 2.00 to 2.50 g/t and greater than 2.50 g/t. The selection in each range was manipulated until the 10% is achieved with a bias towards the mineralized intervals.

Three samples, each weighing about 3 kg were prepared from each original sample bag using the one-stage riffle splitter. Four batches of 175 samples including duplicates and standards were dispatched to SGS, WSL, TWL (Intertek), and ALS Minerals in Ghana-Kumasi (ALS). All samples were labeled with the same identification numbers. A total of 157 assays were returned by each laboratory for analysis.

Statistical comparison of the data indicates that ALS returned lower grades and variance than SGS, WSL and TWL (Intertek). SGS and TWL (Intertek) correlated well with similar minimum and maximum grades, and standard deviation population distribution. The descriptive statistics from the round robin exercise are included in Table 11-19.

Table 11-19 Round-robin descriptive statistics 2012

LaboratorySamples
(no.)
Minimum
(g/t)
Maximum
(g/t)
Mean
(g/t)
VarianceStandard Deviation
SGS1570.0112.01.331.751.32
WSL1570.018.91.091.471.21
TWL/Intertek1570.0111.681.151.681.30
ALS1570.019.321.021.311.15

In 2017, GSR submitted 578 samples to both SGS and Intertek laboratories, inclusive of CRM. Statistical comparison of the data indicated that TWL (Intertek) returned slightly lower grades and variance than SGS. SGS and TWL (Intertek) correlated well with similar minimum and maximum grades, and standard deviation population distribution. The descriptive statistics from the round-robin exercise are included in Table 11-20.

Table 11-20 Round-robin descriptive statistics 2017

LaboratorySamples
(no.)
Minimum
(g/t)
Maximum
(g/t)
Mean
(g/t)
VarianceStandard Deviation
SGS5840.01113.003.3360.797.80
TWL/Intertek5840.01109.203.2155.817.47

When comparing the results from the two laboratories, the HARD analysis showed that approximately 84% of the 584 repeat samples fell within approximately 20% error and 69% fell within 10% error. This was a good correlation between the two laboratories and the decision was made to switch from SGS to TWL (Intertek). The HARD results for the comparison between the two laboratories are shown in Table 11-21.

Table 11-21 Summary HARD plot results for 2017 round robin program

Laboratory <10% HARD<15% HARD<20% HARDCorrelation Coefficient
SGS vs TWL/Intertek 69%78%84%0.97

 In 2018, GSWL changed its primary laboratory from SGS to Intertek Ltd. As part of the QA/QC protocols, to check the reliability of the new primary laboratory, coarse rejects of samples analysed by Intertek were re-bagged with different identification numbers and submitted to SGS as checks, 761 samples were submitted to SGS between January to August of 2020.

 

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Sample checks showed a strong correlation between the analytical methods of both laboratories. The HARD results for the comparison between the two laboratories are shown in Table 11-22. Figure 11-15 and Figure 11-16 show the HARD and corelation plots respectively.

Table 11-22 Summary HARD plot results for 2018 round robin program

Laboratory <10% HARD<15% HARD<20% HARDCorrelation Coefficient
Intertek vs SGS 73%84%90%0.99

  

 

Figure 11-15 HARD plot of 2018 Wassa duplicate analysis (Intertek vs SGS)

 

 

Figure 11-16 Wassa duplicates correlation plot (Intertek vs SGS)

 

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11.5Specific Gravity Data
11.5.1Open Pit

At Wassa SG determinations were carried out by GSR. SG was measured on representative core samples from drill runs. This ensured representative SG data across all rock types irrespective of gold grade.

SG was measured at the core facility using a water immersion method. For each sample in the dataset, the sample was weighed in air, then coated in wax and weighed in air and immersed in water. Historically, a total of 606 determinations were collected on core samples.

The water immersion methodology was considered to provide accurate estimates of variations in bulk SG throughout the Wassa gold deposits. After testing, each sample was carefully replaced at its original location in the core box.

Samples were selected from all the different lithologies. The sampling procedure was guided by pit location, lithology, depth, quartz contents (in oxide) and the oxidation state. A total of nineteen holes from Dead Man’s Hill, South East, Starter, 419, 242, B-shoot and F-Shoot were selected with the results presented in Table 11-23.

Table 11-23 Specific gravity test results, open pit

MaterialNo.  SamplesSG Value (g/cm3)Standard Error
Oxide2131.802%
Transition422.193%
Fresh3272.701%
Quartz Vein242.561%

An additional 13 samples consisting of oxide (9), trans (1), fresh (2) and quartz (1) were sent to the Western University College (WUC, Tarkwa) as independent checks. The average results were 1.76, 2.29, 2.73 and 2.59 g/cm3 respectively.

The SG determinations were considered accurate as the reconciliations between the mined tonnages and those estimated from the resource models reconcile well.

11.5.2Underground, 2017

In 2017 an SG study was completed to test whether higher grade mineralization being mined from underground was heavier than waste rock and the lower grade material mined from the open pits. A total of 40 samples were selected from four underground drill holes and were sent to Intertek Laboratories for wax immersion SG determinations. The results from this study indicated that the higher grade underground mineralization is heavier than the lower grade open pit material. Gold mineralization at Wassa is directly related to the percentage of pyrite associated with quartz veining; in general, the higher percentage of pyrite the higher the gold grades. The underground mining exploits these higher grade areas of the mineralization with associated higher percentages of sulfides which in turn accounts for the heavier mass of this material. The results of the study are summarized in Table 11-24.

Table 11-24 Specific gravity test results, underground drilling 2017

Hole ID

From

(m)

To

(m)

Interval Length

(m)

Au Grade

(avg. g/t)

SG

(avg. g/cm3)

BS17-670-2770.598.55.535.532.98
BS17-670-1180.2110.24.304.302.83
BS17-645-5100.5125.625.142.622.94
BS17-670-2361.878.24.854.853.03

 

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11.5.3Southern Extension, 2018

In June 2018, GSR conducted a specific gravity measurement program. 723 samples from surface drill core and 966 samples from underground core were assessed (Table 11-25 and Table 11-26). Results showed average SG for the underground fresh ore is 2.8. The water displacement method to measure density was employed, using paraffin sealed core samples.

Table 11-25 Specific gravity test results, underground drilling 2018

Rock TypeNo.  Determinations

SG

(avg. g/cm3)

Banded Magnetic Mudstone673.02
Diorite7252.83
Felsic Intrusive--
Phyllite672.74
Quartz Vein1072.65
Total9662.81

 

Table 11-26 Specific gravity test results, surface drilling 2018

Rock TypeNo.  Determinations

SG

(avg. g/cm3)

Banded Magnetic Mudstone322.70
Diorite4702.69
Felsic Intrusive412.59
Phyllite1312.63
Quartz Vein492.57
Total7232.63

 

11.5.42020 Drilling

In 2020, 58 check SG determinations were conducted on the limited surface drilling. No further SG determinations were conducted on the UG core as lithologies have essentially remained the same. These results have confirmed the overall density of 2.8 which has been used in the Mineral Resource Estimations. These results are summarized in Table 11-5 below.

Table 11-27 Specific gravity test results, surface drilling 2020

Rock TypeNo.  Determinations

SG

(avg. g/cm3)

Banded Magnetic Mudstone52.75
Diorite272.80
Felsic Intrusive72.78
Phyllite132.87
Quartz Vein62.67
Total582.77

 

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12Data Verification

The measures implemented by GSR related to data verification are considered by the Qualified Person to be consistent with standard industry practice and of sufficient quality to include in the estimation of Mineral Resources.

Core logging and sampling procedures are considered consistent with industry standards. The Qualified Person has supervised work completed by consultants to assess and validate the logging against the halved drill core with no major errors identified.

“Blind” test samples are frequently sent to the laboratory and monthly batch results are analysed. Any anomalous results are queried immediately. A small number of anomalous and/or poor results have been noted over the years, but these have been identified and the reasons for the results fall into two main categories, namely:

·Mislabelling of individual samples, standards, and blanks; and/or
·Individual batch issues corresponding to changes in the laboratory setup or calibration. In these cases, re-assaying has been carried out.
12.1Drilling Database

The procedures implemented by GSR involve several steps designed to verify the collection of drill hole data and to minimize the potential for data entry errors. At Wassa, data entry and database management involves two steps. Drill hole logs are captured directly into an SQL Acquire database via laptop computers, which are linked to the main database. Acquire has built-in validation tools and drop-down menus, designed to eliminate erroneous data entry during the core logging process.

Analytical data is checked for consistency by GSR personnel with oversight by the Qualified Person. Upon reception of digital assay certificates; the assay results, along with the control sample values, are extracted from the certificates and imported into the Acquire database. Failures and potential failures are examined and, depending on the nature of the failure, re-assaying is requested from the primary laboratory. Analysis of quality control data is documented, along with relevant comments or actions undertaken to either investigate or mitigate problematic control samples.

12.2Other Verifications by the Qualified Person

The QP for the Mineral Resource estimate is Mitch Wasel has been involved with the project since 2003 and data verification since then includes:

·Regular site visits to oversee and supervise drill programs, including adherence to procedures, and oversight as it relates to quality control;
·Verification of core logging;
·Spot checks on the database to ensure it’s representative of hard copy data;
·Review and interpretation of QA/QC results (described in Section 11);
·Audits of laboratories;
·Comparison of RC and DD assays;
·Review of analytical methods, results of which informed the decision to move from AAS to BLEG analysis (described in Section 11);
·Confirmation in the field of collar coordinates to verify drill hole locations; and
·Confirmation in the field of downhole surveys.

 

 

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13Mineral Processing and Metallurgical Testing
13.1Early Metallurgical Test Work

On obtaining ownership of the Project in 2002, GSR commissioned a feasibility study (FS) for a CIL operation with the process engineering component undertaken by Metallurgical Process Development Pty Ltd. (now known as MDM). The FS was completed in 2003. The metallurgical test work conducted in support of the MDM FS was conducted on samples from the Wassa area. Samples were originally sent to SGS Lakefield in Johannesburg for both variability and bulk sample test work. Further variability test work was conducted at AMMTEC in Perth.

A total of 24 variability samples were tested; 10 of fresh mineralized material, six of oxide, and 8 samples taken from the existing (now decommissioned and reclaimed) HL operation. Four bulk samples were also tested, representing fresh, oxide, HL phase 1 and HL phase 2. The samples were all taken from the Wassa Main area.

At a grind size of 75% -75 µm, and a 24-hour leach time, the fresh bulk sample achieved a leach recovery of 92%. The Bond Ball Mill Work Index (BWi) for this sample was 14.8 kWh/t. Under the same conditions, the oxide bulk sample achieved a leach recovery of 93%. The BWi for this sample was reported as 8 kWh/t. Minor preg-robbing behaviour was noted, and gravity recovery test work indicated that plant recoveries of 30 to 40% could be expected from a gravity circuit.

13.22015 Test Work Program

In 2015 as part of the Wassa Underground feasibility study, further metallurgical test work was completed. The test work evaluated the performance of feed from underground with a series of half-core samples from definition drilling. The physical characteristics and metallurgical response of these samples were compared to those of a reference sample of current plant feed from that time (open pit sourced).

At the time of test work an exploration decline and bulk sample was obtained from underground, which was expected to be representative of the underground feed material. The benefit of bulk sample treatment through the plant resulted in a reduced test work program that included a series of six variability and four crushability samples that were compared to a reference sample taken from the current open pit ore feed.

The metallurgical test work was undertaken by SGS in Cornwall, UK and the samples were delivered and logged in the middle of December 2014 with this initial phase of test work completed and the draft report issued in early April 2015.

13.2.1Metallurgical Variability, Crushability and Reference Samples

For the purpose of the metallurgical program, the material planned for future processing was differentiated spatially by GSR into six underground domains or zones which are depicted in Figure 13-1 with further details presented in Table 13-1.

Table 13-1 Ore zones represented by the variability samples

ZoneNorthingRelative Level

 Tonnes

(‘000 t)

 Grade
(g/t Au)
cont.Au (koz)Tonnes share %Metal share %
from mNto mNfrom mRLto mRL
Zone 1 upper20,20019937.58576825984.7491.215%14%
Zone 1 lower20,20019937.56826077076.78154.118%23%
Zone 2 upper19,937.5196907826327236.28146.118%22%
Zone 2 lower19,937.5196906325075384.3274.714%11%
Zone 3 upper19,690195006575577725.02124.620%19%
Zone 3 lower19690195005574826134.282.816%12%
Total Processing Inventory  3,9525.3673.5100%100%

 

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Figure 13-1 West view of metallurgical sample locations (GSR, 2015)

Six variability samples were selected, one for each zone from available HQ and NQ half cores. These core sections were further cut in half, with one section used for the metallurgical test work and the remaining quarter core sections retained for reference. Each sample of quarter cores weighed between 50 and 60 kg.

Four full core samples (with a segment removed for assay purposes) were selected for the crushability tests. Each crushability sample consisted of 7 lengths of HQ drill core each approximately 200 mm in length. From these, three samples were prepared for the UCS tests with the remaining core sections and material from UCS testing prepared for the Bond crushability index (low energy crushing) tests.

A single reference sample was also obtained by hand selection from the workings in the Starter pit area at around the 910 m level. Around 100 kg of material was taken and this sample was used for both metallurgical and crushability test work.

Table 13-2 Summary and location of test work samples

Sample TypeDetailNorthingEastingRelative Level

Sub-samples / Intersections

no.

fromtofromtofromtoavg.
mNmNmEmEmmm
Reference 20,42020,39640,00439,9749109109106
VariabilityZ1U19,97220,04340,11339,9848286827636
VariabilityZ1L19,94719,98839,99439,9126786156647
VariabilityZ2U19,77019,84640,08439,9307536537135
VariabilityZ2L19,70019,75740,07939,9316025305756
VariabilityZ3U19,53119,57640,02339,9796025625854
VariabilityZ3L19,49719,56540,04039,9455555105335
Crushability 1BSDD347MET19,49219,48940,02439,9995875145538
Crushability 2WMET420,05320,05040,01439,9997677487538
Crushability 3WMET520,03620,03639,98039,9757227137198
Crushability 4WMET620,01720,01639,97639,9647166527008

 

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The locations of the reference, variability and crushability samples are presented in Table 13-2 along with the nominal ore zones selected. Some of the crushability samples selected were adjacent to rather than completely within the representative ore zone. Crushability 1 was from depth to the south of Zone 3 Lower while the other crushability samples were from different depth within Zone 1 Upper and Zone 1 Lower. The reference sample was taken from the current workings in the starter pit area above and to the north of Zone 1 Lower at 910 mRL.

13.2.2Details of Metallurgical Test work

The metallurgical evaluation test work program included the following investigations:

·Scope of work for reference and variability samples:
oelemental scan: ICP multi-element analysis;
oanalysis of sulphide and total sulphur;
oanalysis of carbonate and graphitic carbon;
odiagnostic leach (gold deportment tests);
oBWi; and
oBond abrasion index (Ai).
·Standard flowsheet treatment tests – to confirm recoveries and reagent additions / consumptions:
ogrind calibration tests;
ogravity concentration;
ocyanide leaching of the gravity tails with pre-aeration; and
osettling tests.
·Scope of work for crushability and reference samples:
ounconfined compressive strength (UCS);
oBond low impact crushing work index (CWi);
oBWi; and
oAi.
13.3Test Work Findings
13.3.1Head Grade and Elemental / Chemical Analyses

The gold and silver head grades were determined by milling and screening at 106 µm with fire assay of the two screen fractions. The results are summarized in Table 13-3.

Table 13-3 Screened head assay results

SampleOverall GradeSize fractionGold DistributionSilver Distribution
+106 micron-106 micron
Au g/tAg g/tShareAu g/tAg g/tAu g/tAg g/t+106μm-106μm+106μm-106μm
Reference1.530.11.9%11.320.21.140.113.9%86.1%3.7%96.3%
Zone 1 Upper6.510.42.4%28.291.67.030.410.3%89.7%8.8%91.2%
Zone 1 Lower7.990.62.3%42.294.27.310.612.0%88.0%15.0%85.0%
Zone 2 Upper5.110.41.3%17.261.04.380.34.2%95.8%3.5%96.5%
Zone 2 Lower4.640.22.4%9.940.84.520.25.0%95.0%8.8%91.2%
Zone 3 Upper4.070.51.6%9.450.64.420.53.6%96.4%2.1%97.9%
Zone 3 Lower5.260.62.2%25.32.85.290.510.3%89.7%10.9%89.1%

 

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In all samples, the gold and silver analyses in the coarse fraction (+106 µm) is higher than for the finer fraction (-106 µm).

An ICP elemental scan was undertaken on the reference and variability samples; in addition, the total carbon and organic carbon as well as the total sulphur and sulphide sulphur were analysed using the Leco method. Results are presented in Table 13-4.

Table 13-4 Elemental and chemical analysis results

Sample1010A2008A3008A4008A5008A6007A7007A
(%)REF1Z1UZ1LZ2UZ2LZ3UZ3L
Cu0.0030.0190.0110.010.010.0080.01
Pb<0.0010.002<0.001<0.001<0.001<0.0010.002
Zn0.0060.0090.010.0080.0090.0080.007
As<0.0010.0010.0030.0010.0010.001<0.001
Cd<0.00010.00030.00030.00030.00020.00020.0002
Ni0.0020.0040.0040.0020.0020.0050.003
Co<0.0010.0030.0040.0040.0030.0030.003
Mn0.070.140.180.20.150.10.13
Bi<0.0001<0.0001<0.0001<0.0001<0.0001<0.0001<0.0001
Sb<0.0001<0.0001<0.0001<0.0001<0.0001<0.0001<0.0001
Hg<0.0001<0.0001<0.0001<0.0001<0.0001<0.0001<0.0001
Te<0.0001<0.0001<0.0001<0.0001<0.0001<0.0001<0.0001
Se<0.0001<0.0001<0.0001<0.0001<0.0001<0.0001<0.0001
SiO278.4674.9665.3966.5159.4265.3957.55
Al3.323.484.464.375.224.655.24
Fe2.835.576.465.484.673.924.62
Mg0.740.881.091.271.531.471.8
Cr0.030.060.050.030.020.010.01
Ca1.821.11.812.143.472.713.77
S0.460.861.560.981.31.170.9
Na0.920.961.461.931.981.572.16
K1.361.71.791.571.382.111.6
% S (total)0.460.861.560.981.31.170.9
% S (soluble)0.020.030.040.040.040.030.03
% S (sulphide)0.440.831.520.941.261.140.87
% C (total)1.41.421.691.992.221.862.52
% C (organic)0.030.020.030.020.030.020.02
% C (CO3)1.371.41.661.972.191.842.5

The level of sulphide sulphur was higher in the higher grade variability samples than in the reference sample. Similarly, the level of iron and other base metals was higher; however, the levels of the other base metals is relatively low.

 

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13.3.2Diagnostic Leach

Diagnostic leaching is a method of quantifying the indicated deportment of gold in a sample and the relative ease or difficulty with which the gold can be recovered. The sample is prepared by grinding to a typical grind size likely to be employed (75% < 75 µm was selected) and is subject to a cyanide leach to dissolve the free gold. The solids from the initial cyanide leach test are then sequentially pre-treated with more aggressive acids to dissolve minerals that could be encapsulating the residual gold. Following each pre-treatment stage the sample is again treated by cyanide leaching. As the level of sulphide minerals was indicated to be higher in the higher grade underground material from the geological interpretation of the core samples and confirmed from the elemental analyses presented in Table 13-5 the aim was to determine whether the increased level of sulphide minerals was resulting in the samples being more refractory to treatment for the recovery of gold.

In the diagnostic leach procedure, the samples are subject the following leach and pre-leach treatments:

·Direct cyanidation: recovers free and exposed gold.
·Hydrochloric acid pre-treatment: liberates gold encapsulated in carbonates, pyrrhotite, galena and iron hydroxide minerals.
·Sulphuric acid (oxidative) pre-treatment: liberates gold encapsulated in sphalerite, labile copper sulphate and labile base metal sulphide minerals.
·Nitric acid pre-treatment: liberates gold encapsulated in pyrite, arsenopyrite and marcasite.
·Carbon combustion: burns off any organic carbon releasing gold that had previously been adsorbed by the carbon and not therefore amenable to recovery by cyanide leaching.

Residual gold and silver present after the above tests represent gold encapsulated in silica and other non-reactive gangue minerals.

Results of the diagnostic leach tests for gold are summarized in Table 13-5 and represented graphically showing the deportment of gold in the samples in Figure 13-2.

Table 13-5 Summary of diagnostic leach results

Gold DeportmentSample Reference
Ref 1Z1UZ1LZ2UZ2LZ3UZ3L
%%%%%%%
Cyanide Soluble91.996.8297.0593.1386.9289.585.34
In Carbonates / Pyrrhotite1.380.881.11.78.832.372.99
In Sphalerite and Labile Sulphides0.660.580.230.731.220.972.18
In Pyrite and Arsenopyrite2.531.221.263.31.914.017.01
In Graphitic Carbon0.590.270.10.350.380.450.4
Residual Gold2.930.230.250.790.742.712.08
TOTAL100100100100100100100


The results generally indicated that the mineralogy and metallurgy of the samples are somewhat different with some samples appearing to have potentially more gold locked or associated with different sulphide minerals and others less when compared to the reference sample. Two samples (Z3U and Z3L) showed potentially higher levels of gold encapsulated in pyrite while sample Z2L showed higher levels of gold potentially associated with more reactive minerals such as pyrrhotite. The results were not seen to be completely consistent with the gravity leach results discussed later.

 

 

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Low levels of preg-robbing potential were indicated from the gold liberated in the burn off stage. It should be noted that due to assay detection limits some of the lower deportments may be marginally inaccurate. Given a detection limit of 0.01 g/t Au, measurements below this level were assigned a nominal assay of 0.005 g/t Au; hence on the lower levels the deportment in these fractions could be slightly overstated.

It was reported in the diagnostic leach tests during the hydrochloric acid digestion that a reasonably vigorous reaction took place on the majority of the variability samples with the generation of green foam. This would tend to indicate a high level of carbonate and also acid soluble iron, possibly pyrrhotite.

 

Figure 13-2 Comparative indicated deportment of gold from diagnostic leach results

13.3.3Crushability

Two separate tests were undertaken into the material strength and crushability by measuring the UCS and the CWi test, which indicates a material’s resistance to crushing. In the UCS test, a sample is prepared by cutting to pre-set dimensions (re-coring) and this is then subject to a compressive load to measure the strength at which the sample fails. The Bond CWi test, also known as the low impact energy test, involves two swinging weighted pendulums which are allowed to fall and impact simultaneously on the sample in order to measure from what height the pendulum needs to fall to crush the sample. Both tests are undertaken on multiple individual samples; 3 prepared samples in the case of the UCS tests and around 20 sample pieces for the Bond CWi test. The results of the tests are presented in Table 13-6.

Table 13-6 Results of Crushability Tests: UCS and CWi

 DensityDepthUCS Result (Mpa)CWi (kwh/t)Depth
t/m3RL mAverageMaxMinAverageStd Devm RL
Reference2.6791059.573.741.89.81.6910
Crushability 12.9355064.776.954.39.71.3550
Crushability 22.8775353.994.431.111.11.2753
Crushability 32.71720167.424490.7112.1720
Crushability 42.8469982.49068.912.32.9699

The UCS test results are seen to be variable, with a relatively large variation between the maximum and minimum measurements on the different samples which mainly appear to relate to the sample tested rather than the depth of the material. Results were generally in the 30 to 95 MPa range, indicating that the materials tested were medium strong to strong, although one sample (Crushability 3) indicated to consist of quartzite (massive quartz vein), rather than schist identified for the majority of the other samples tested, recorded a very strong measurement of around 240 MPa. The other sample of the same type of material measured 90 MPa, while a third sample shattered during preparation and cutting and failed to produce the required test sample.

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The CWi test results are in the easy to medium classification. Similar to the UCS results, the CWi test results are also relatively variable with the reference sample (910 mRL) generally indicating results towards the lower end of those measures; however, no real correlation can be see between the CWi results and relative level of the sample tested as shown in Figure 13-3.

 

Figure 13-3 Variation of UCS and CWi results with depth (mRL)

13.3.4Ball Milling Bond Work Index and Abrasion Index

For the 2003 FS into the treatment of the Wassa material by milling and CIL, test work was undertaken on representative samples of primary ore, oxides and spent HL material. The BWi for the primary and oxide ores were reported to be in the region of 14.6 and 8 kWh/t, respectively.

More recent investigations suggest that the BWi is generally noted to be increasing with depth. Based on samples tested from three different drillholes from the Wassa starter pit area, SE Area and MSN Area, BWi measurements, though somewhat inconsistent, appeared to indicate that the BWi was increasing with depth.

From 2015, with fresh open pit ore feed, the unit power draw presented for the two ball mills is shown to be between 14.5 and 16.5 kWh/t treated. This results in a calculated BWi of around 14 – 16 kWh/t, based on the reported mill feed and product sizes and power draw on the ball mills. An allowance has been included in the calculations for mechanical and other losses between the drive motor and mill. In recent years with the blend of underground and open pit the BWi continues to remain within the 14 – 16 kWh/t range.

The findings of the BWi and Ai investigations from the 2015 tests are presented in Table 13-7 and are shown as a function of the average sample depth Figure 13-4 and Figure 13-5, respectively.

The BWi tests were undertaken at a closing screen size of 106 µm to give a mill product of around 75-80% < 75 µm.

In summary, the findings of the latest test work generally did not support the suspected increasing BWi with further depth with the reference sample (910 m RL) showing the highest BWi reading.

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Table 13-7 Results of 2015 BWi and Ai Tests

Sample DescriptionBWiAiAvg.  Depth
kWh/tRL m
Reference15.70.394910
Z1U Zone 1 Upper15.30.33763
Z1L Zone 1 Lower14.70.276664
Z2U Zone 2 Upper14.90.228713
Z2L Zone 2 Lower14.50.175575
Z3U Zone 3 Upper14.40.229585
Z3L Zone 3 Lower13.90.152533
Crushability 1 (347MET)140.182553
Crushability 2 (MET4)150.205753
Crushability 3 (MET5)14.80.398719
Crushability 4 (MET6)14.80.326700

 

 

Figure 13-4 2015 Ball Mill Bond Work Index against sample depth (mRL)

 

 

Figure 13-5 2015 Abrasion Index against sample depth (mRL)

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The abrasion index is a measure of the anticipated wear on components and consumables in the comminution circuit and is applicable to wear in both crushers and mills (media and liners). Ai is generally shown not to be increasing with depth and it appears that Ai is slightly lower on deeper samples. With the exception of one sample (MET 5) of massive quartz vein, measured Ai is for the reference sample is higher than all the other variability and crushability samples tested. This lower indicated abrasion index with depth may result in the reduced consumption of grinding media and mill crusher liners as mining proceeds deeper into the underground mining areas. All the samples fall into the slightly abrasive classification.

13.3.5Gravity Gold and Leaching Tests
13.3.5.1Gravity Tests

Gravity tests were undertaken by grinding a 1 kg sample to approximately 75% passing 75 µm and then passing the sample through a Falcon centrifugal concentrator. The primary concentrate from the Falcon was further processed on a Mozley shaking table, with the final concentrate weighed and sent for assay. Tailings from the centrifugal concentrator and shaking table were subject to cyanide leach tests.

The results of the gravity concentration tests are presented in Table 13-8.

Table 13-8 Gravity Gold Recovery Test Results

Sample RefGravity Con MassAssayMetal Recovery to Gravity Con
gWt %Au (g/t)Ag (g/t)% Fe% S (total)Au %Ag %
Ref13.30.3384.338.019.5915.6118.1926.4
Z1U2.10.21322.618.837.2821.8610.419.18
Z1L4.90.49322.319.338.2426.9219.7715.01
Z2U2.50.25324.326.337.0531.0915.8718.26
Z2L3.00.30211.613.435.8444.1513.6819.14
Z3U2.70.27199.214.834.3138.3213.218.88
Z3L2.40.24282.824.128.8029.7612.9010.52

Gravity recoveries were lower than previously reported. This is probably a function of the laboratory tests which, for this stage of the investigation, were not optimized to maximize gravity gold recovery. It can also be seen that the recovery from the reference sample is generally higher than on the variability samples.

In all gravity tests, concentrates contained a magnetic component that was readily picked up by a strong rare earth magnet but not an iron magnet. This magnetic component was suspected to be pyrrhotite and this was reported by SGS to be supported by the sulphur to iron ratios measured in the feed analyses.

13.3.5.2Whole Ore Leach and CIL Evaluation Test

In order to investigate the effective leach parameters for the comparative leaching tests, a single leach test was undertaken on the reference sample with and without carbon to confirm whether any preg-robbing effect was evident. The results are presented in Table 13-9.

Table 13-9 Whole Ore Leach and CIL test results

 Solution (24h/48h)Solid tailsGold on CarbonOverall
Recovery
Back Calc.  
Head Grade
 Au g/tAg g/tAu g/tAg g/tAu g/tAg g/tAu %Au %Au g/tAg g/t
Leach Test1.130.080.1050.05----1.550.15
Distribution93.2%67.0%6.8%33.0%--93.2%67.0%--
CIL Test0.140.010.10.0593.412.7--1.210.19
Distribution14.3%6.5%8.3%26.4%77.4%67.1%91.7%73.6%--

 

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The results generally indicated that no preg-robbing effect was evident with the recoveries without carbon addition higher than those with carbon added to the leach (CIL test), although the gold reconciliation was seen to be worse on the CIL test with a back-calculated gold head grade of 1.21 g/t Au compared to the screened analysis head grade and leach test back-calculated head grade of 1.53 and 1.55 g/t Au respectively.

13.3.5.3Gravity Tails Leach Test Results

Leach tests were undertaken on the combined gravity tails from the centrifugal concentrator and concentrate cleaning table. From the gravity tests and one of the diagnostic tests there was potential that pyrrhotite could be present so the gravity tails samples were adjusted to pH 10.5 - 11 using lime and aerated until the pH and dissolved oxygen levels stabilized generally in line with the plant practice of injecting oxygen into the transfer lime from milling to CIL. Pyrrhotite is highly reactive and can result in high consumptions of oxygen and cyanide in leach if not preconditioned.

Leach tests were conducted for 48 hours with samples taken at 2, 4, 6, 24 and 48h and analysed for gold and silver in solution. An initial cyanide level of 1 g/l was used and cyanide levels in solution were maintained at >0.5 g/l by dosing of additional cyanide as required. The tails solids were analysed for silver and gold. No lead nitrate was added in the leach tests.

Leach test results of the gravity tails are presented in Table 13-10.

Table 13-10 Leach test results and reagent consumptions

Sample ReferenceGold Recovery %Assayed TailsConsumption kg/t
24h48hg/t AuNaCN 24hNaCN 48hLime as CaO
Ref177.2288.690.090.431.310.88
Z1U90.6987.350.440.511.480.89
Z1L86.7287.640.680.401.150.75
Z2U92.8193.800.200.431.050.92
Z2L87.8188.060.420.150.910.88
Z3U92.9591.330.230.630.891.16
Z3L94.5793.250.180.631.011.11


Figure 13-6 Leach recovery kinetic curves

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It can be seen that in some tests, recoveries based on 48h leach solution analyses were lower than for the those based on the 24h leach solution assays. This could be caused by analytical discrepancies or errors based on solutions analysed or possibly some adsorption of dissolved gold onto the fine milled solids. As no appreciable preg-robbing potential or effect was indicated in the diagnostic leach and comparative leach and CIL tests, this is not considered to be a major concern as any weakly adsorbed gold would be recovered on the plant due to the presence of activated carbon in the leach circuit.

The leach curves on the gravity tails appear to be relatively consistent with the exception of that for the reference samples which shows slower kinetic especially at 24h, although results in similar overall recoveries at 48h, shown in Figure 13-6.

13.3.5.4Overall Gravity / Leach Recoveries

The overall recoveries from the gravity / leach test work are presented in Table 13-11. These are based on the maximum leach recovery at either 24 or 48h and on the back-calculated head grade from the recovered gold and tailings assays.

Table 13-11 Overall gravity leach recoveries

Sample
Reference
Gold Recovery %
GravityLeachOverall
Ref126.4188.6991.68
Z1U16.3890.6992.22
Z1L22.6987.6490.44
Z2U20.1993.8095.05
Z2L15.3788.0689.90
Z3U16.9192.9594.15
Z3L20.4194.5795.68

In the gravity / leach tests, poor reconciliations were achieved between the back-calculated head grade and the assay head grades from the screened analyses on the master samples with the back-calculated head grades consistently being considerable lower than the head assay results by as much as 35% in two tests.

The comparison of the assay head compared to the back-calculated head grade for both the gravity leach and diagnostic leach results are presented in Table 13-12.

Table 13-12 Reconciliation of assay and back-calculated head grades from test work

SampleAssay Head GradeFrom DiagnosticsFrom Gravity / Leach
GradeGrade
g/t Aug/t Agg/t Aug/t Agg/t Aug/t Ag
Reference1.530.101.350.221.080.13
Z1U6.510.436.740.894.180.31
Z1L7.990.638.711.067.080.46
Z2U5.110.365.100.624.030.38
Z2L4.640.214.840.464.140.32
Z3U4.070.454.120.333.200.30
Z3L5.260.554.280.273.360.29

The correlations were better in the diagnostic leach tests compared to the gravity / leach tests with both positive and negative discrepancies. Differences varied between -10% and +18% resulting in an overall difference of only -2%.

 

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13.3.6Settling Tests

Comparative settling tests were undertaken on the reference sample and one selected variability sample (Z1L). Initial scoping tests were undertaken using five different flocculants with the settling tests undertaken using Nasaco anion flocculants N2132 and N2326. The results show very similar settling performance on the reference samples and one variability sample selected.

The settling test results are presented in Table 13-13.

Table 13-13 Comparative settling test results

SampleFeed
Solids
pHFlocculantFlocculant DosageInitial Settling RateFinal
Solids Content
Thickener Underflow Unit Area
%g/tm3/m2/day%m2/t/d
Reference Test 19.4310.5N213250.041335.26590.235
Reference Test 210.0810.5N232646.622897.8661.80.261
Z1L Test 19.0410.5N213252.212414.8856.50.225
Z1L Test 29.1310.6N232651.692637.7956.90.223

 

 

 

 

 

 

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14Mineral Resources
14.1Introduction

The Mineral Resource Statement presented herein represents an estimate for the Wassa Main deposit and the satellite deposits Chichiwelli, Benso and Hwini Butre. The Mineral Resource Statement is presented in accordance with the guidelines of NI 43-101.

The GSR exploration team was responsible for preparation of the long-range model (LR model) for the Wassa Mineral Resource modelling exercise which included all topographic surfaces, weathering surfaces, structural control lines and resulting Leapfrog Isoshells. SRK (Toronto), utilizing the inputs from the GSR geologists, estimated gold grades for the LR model, whereas the short-range model (SR model), used within the active mining area, Figure 14-1, was created by the GSR mine geologists with assistance from SRK (Moscow). The Father Brown and Adoikrom Mineral Resource Estimates were created in a similar manner with GSR providing drill hole intervals for HW, HG and FW mineralized zones to Resource Modeling solutions (RMS) who provided 2D estimates of the grades and thicknesses. The Benso and Chichiwelli Mineral Resource Estimates are historical models created by GSR geologists and SRK (Cardiff). The Mineral Resource classification and statement was conducted by GSR under the supervision of S. Mitchel Wasel, a QP.

For the SR model, the site’s mine geology group was responsible for the generation of the structural control lines used to influence the grade interpolation in the model. SRK (Moscow) then performed the domain generation and grade interpolation steps, as well as the depletion and validation of the block model. SRK (Moscow) depleted the block model with asbuilt and CMS volumes provided by GSR. The completed block model was also independently validated by the Wassa mine geology group.

 

Figure 14-1 Wassa long-range (grey) and short-range (cyan) Mineral Resource estimation model limits

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This section describes the Mineral Resource estimation methodology and summarizes the key assumptions considered for the estimate. The Mineral Resource estimate reported herein is a reasonable representation of the global gold Mineral Resource found at the Wassa Main and satellite deposits given the current level of sampling. The Mineral Resources have been estimated in conformity with generally accepted CIM “Estimation of Mineral Resource and Mineral Reserves Best Practices” guidelines and are reported in accordance with NI 43-101. Mineral Resources are not Mineral Reserves and do not have demonstrated economic viability. There is no certainty that all or any part of the Mineral Resource will be converted into Mineral Reserve.

The databases used to estimate the Mineral Resources were audited internally by GSR. In the opinion of the GSR QP, S. Mitchel Wasel, the current drilling information is sufficiently reliable to interpret with confidence the boundaries for gold mineralization and that the assay data are sufficiently reliable to support Mineral Resource estimation.

14.2Mineral Resource Estimation Procedures

The Mineral Resource evaluation methodology involved a database compilation and internal validation exercise by GSR. At Wassa, GSR was responsible for structural control lines, grade wireframes, topographic and weathering surfaces. GSR provided SRK with borehole databases, structural control lines, grade wireframes, topographic surfaces and weathering surfaces. At HBB, GSR was responsible for the HW, HG and LG drillhole intervals, topographic and weathering surfaces and RMS estimated the gold grades and mineralized zone thickness.

Prior to initiating the modelling and Mineral Resource estimation process, SRK reviewed the databases for the Wassa project. The Father Brown and Adoikrom data was reviewed by RMS prior to gold grade estimations.

After evaluating the available database, SRK proceeded with (Wassa), the data conditioning (compositing and capping) for geostatistical analysis and variography. At Wassa, the grade wireframe modelling was completed in Leapfrog Geo 4.4 under following the guidelines that GSR and SRK have established together. The grade interpolation methodology was discussed between GSR and SRK, it was decided to use Ordinary Kriging (OK) with local varying angles and local variograms for the estimation of gold grades based on the structural complexity and folded nature of the deposit.

For the SR model, database compilation and internal validation checks were performed by GSR. The database was then passed to SRK (Moscow) who performed a second data validation prior to commencing modelling work. GSR was responsible for the generation of the structural control strings, which were used to influence the domain shell creation and the grade estimation.

SRK (Moscow) then:

·Created the structural control meshes from the structural control strings;
·Created the domain shells;
·Coded the sub-block model with the domain information;
·Estimated grade using OK with locally variable anisotropy;
·Depleted the model with development and stope surveys; and
·Validated the model.

On completion, SRK (Moscow) submitted the Surpac and Leapfrog models to GSR Mine Geology for independent validation.

The classification of the LR model and preparation of the Mineral Resource Statement utilizing both long-range and short-range estimates were conducted by GSR under the supervision of GSR’s QP, Mr. S. Mitchel Wasel.

 

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14.3Mineral Resource Database
14.3.1Wassa

The Wassa database is made up of five individual drillhole databases, namely:

·GSR Wassa exploration database, which contains exploration drilling conducted by GSR since 2002;
·GSR Underground drilling database, which contains all of the diamond drill holes drilled from underground;
·the All Wassa (AW) exploration database, which contains historical exploration drill holes from SGL;
·Satellite Open Pit grade control database, which is dominantly blast holes; and
·GSR Open Pit grade control database, which is dominantly RC holes.

The Satellite grade control database was not included in the Mineral Resource estimate as the blast holes samples are considered not to be of a sufficient quality for use in the Mineral Resource estimate.

A completion date cut-off was applied to the GSR exploration and Underground drill hole databases. For the 2020 year-end LR model estimate the cut off for surface and underground drilling was 31 January 2020 and for the SR model estimate the drilling results were up to 30 November 2020. These are the data sets utilized for the subsequent Mineral Resources shown in Table 14-1.

Table 14-1 Wassa LR model drill hole database as at February, 2020

DatabaseTotal
TypePurposeHolesMetres
Pre-ExistingGrade Control24,957642,470
Exploration3,422500,282
UG Exploration84793,896
At February 20Grade Control41112,142
Exploration5948,036
UG Exploration37156,914
Total 30,0671,353,740

The later 30 November cut-off allowed addition of 273 underground drill holes totaling 34,275 m to be included in the SR model estimate.

The borehole databases contain: collar details; downhole deviation surveys; gold assays; lithological descriptions; alteration; structural data; major structures and vein descriptions. GSR and SRK have performed validation routines to the database. Based on this assessment and checks described in Section 12, it is the opinion of the QP that the database is appropriate to inform the Mineral Resource estimate.

For the SR model, all data was initially stored in the GSR master acQuire database. The relevant Wassa Mine geology data was then exported from acQuire in .csv format, to a Surpac-linkedMicrosoft Access database. The database contained all Wassa related surface exploration drilling, all underground Mineral Resource definition drilling, all underground channel sampling, all underground chip sampling, and all underground sludge sampling data.

For the purpose of the short-range block modelling, the chip, channel and sludge data was excluded from the estimation runs, leaving only the surface and underground diamond drill core and surface RC assay data. GSR made this decision because, after statistical review of the data, SRK (Moscow) concluded that the risk of biases existing in the chip, channel and sludge data exceeded the benefits from allowing these additional samples to inform the estimation. Table 14-2 summarizes the number of drill holes contained in the Surpac Access database, that were subsequently used for the short-range Mineral Resource estimate. The Surpac Database only utilized surface drill holes that are in the immediate vicinity of underground mine, whereas the LR model has a much larger extent and used all of the validated drill holes available.

 

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Table 14-2 Wassa Underground short-range drill hole database as of December 1, 2020

LocationTypeNo.  HolesDrill Metres
Wassa UGDD1,491185,086
Wassa SurfaceDD841302,253
Wassa SurfaceRC54356,894
Total DrillingDD + RC2,875544,233
14.3.2Hwini Butre

The Father Brown/Adoikrom database is made up of Exploration DD and RC holes as well as RC grade control drilling data.

The 2020 year-end Mineral Resource estimate utilized all of the drilling data that was available at the end of 2019 which essentially remained unchanged as at the end of 2020, as summarized in Table 14-3.

Table 14-3 Father Brown/Adoikrom drill hole database as of December 2020

LocationTypeNo.  HolesDrill Metres
Father Brown/AdoikromDD Exploration43566,229
RC Exploration21416,323
RC Grade Control3,08772,037

The borehole databases contain information including collar information, downhole deviation surveys, gold assays, lithological descriptions, alteration, structural data, major structures and vein descriptions.

GSR has performed validation routines to the Mineral Resource database. Based on this assessment, and the checks described in Section 12, it is the opinion of the QPs that the borehole database is appropriate to form the basis of the Mineral Resource estimate.

14.3.3Benso

SRK was provided with a Gemcom project directory containing the drilling data (Table 14-4) as audited by GSR along with the geological wireframes, oxidation and topographic surfaces and block model parameters. Additional information was provided as Excel spreadsheets documenting QA/QC data and results of density determinations.

Table 14-4 Benso drill hole database as of December 2012

LocationTypeNo.  HolesDrill Metres
BensoRC46533,276
DD32137,623
Geotech141,637
GC (RC)2,36257,970
14.3.4Chichiwelli

SRK was provided with a Gemcom project directory containing the drilling data (Table 14-5) as audited by GSR and the geological models subsequently produced by GSR including geological wireframes, oxidation and topographic surfaces and block model parameters. Additional information was provided as Excel spreadsheets documenting QA/QC data and results of density determinations.

Table 14-5 Chichiwelli drill hole database as of 2012

LocationTypeNo.  HolesDrill Metres
ChichiwelliRC48329,802
DD233,692
Geotech--
GC (RC)--

 

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As no mining has taken place at Chichiwelli, the topographic survey used for the 2010 Mineral Resource statement remains current.

The “HBB other” tonnes and grade in the Mineral Resource statements in sub-section 14.9 also includes three small deposits located in the Manso and Hwini Butre Prospecting and Mining licence areas. These deposits are Abada, Adoikrom South and C3PR. The techniques used to estimate these deposits are consistent with those reported for Chichiwelli.

14.4Grade Shell Modelling
14.4.1Wassa Mineralization Wireframes

The LR model wireframe modelling was carried out by GSR geologists using Leapfrog Geo 6.0 software. Mineralized wireframes at Wassa are modelled using an indicator approach which uses a 0.4 g/t cut-off for the low grade (LG) envelopes and a 1.5 g/t cut off for high grade (HG). Visual inspection of assay data suggests that these respective lower cut-off levels are reasonable to separate barren from auriferous sections intersected by each borehole. Mineralized shells are created using this indicator approach combined with structural trend surfaces created by the site geologists and reviewed by SRK.

For the SR model, the cut-offs used to define LG or “halo” domain and HG or “mineralized” domain were the same as those used for the LR model but the methodology used in by SRK using Leapfrog Geo 6.0 software was different in that grade thresholds were used instead of indicator. As per the LR model, structural trend surfaces were used to influence the shape of the domain shells.

14.4.2Wassa Indicator Interpolants - Background

An indicator interpolant works in a similar way to a grade shell, but rather than interpolating the raw grade, all data above the given indicator grade value is assigned a value of 1 and all data below the indicator grade value are assigned a value of 0. A shell is then generated at a defined iso-value, between 0 and 1. This helps to remove the impact of very high grades which can result in “blow-outs” or unrealistic volumes that can result from standard grade shell modelling of highly skewed data populations.

The indicator interpolant is influenced by an anisotropic structural trend, which is based on form surfaces. The form surfaces represent vectors of grade continuity, where grade continuity is high along the modelled form, and low across it. Due to the significantly deformed nature of the gold mineralization, this type of 3-dimensional structural trend is vital to produce a geologically realistic shape of the indicator interpolants.

The SR model did not use the indicator interpolant method employed in the LR model. Rather, the domain shells for the SR model were generated by directly interpolating the Au grade values above a nominated threshold. The reason for the difference between the two modelling methodologies comes down to the difference in drill spacing between the two models – wide spacing in the LR model and tighter drill spacing in the SR model. To avoid “blow-outs” related to high grade values in widely spaced drill data, the LR model seeks to limit the high grade values by restricting the volume of the domain shell generated around that high grade assay. However, for the SR model, the drill spacing is much tighter, which reduces the impact of any isolated high grade values. GSR has periodically reviewed the decision to use two different modelling methodologies. At this time, GSR believes that this approach still provides the best estimate of Au grades in each model.

14.4.3Wassa Structural Trend

The structural HG mineralization trends at Wassa (Figure 14-2) has been created utilizing underground mapping, open pit grade control data and available downhole diamond core structural data. A structural consultant was engaged to assist with the creation of these trends in the southern portion of the Wassa deposit, where there are currently Inferred Mineral Resources.

The structural trend lines were created on sections and then used to create a series of form surfaces, which in turn were used to guide the interpolation of the grade isoshells and populate the local block angles for search ellipse orientation and grade estimation. These form surfaces represent the broad F4 folding event, a plunging synclinal feature which affects grade distribution at the mine scale, with some subtly different internal orientations attributed to the largest features associated with an earlier high strain folding event (F3). In areas where tight underground drilling and mapping are available it is possible to create structural controls surfaces that reflect the local mineralization geometry, often associated with smaller parasitic F4 and F3 folds. 

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Figure 14-2 Wassa LR model structural ‘Form’ surfaces (oblique view looking N up plunge), surfaces show deposit scale F4 fold as well as rolling over of mineralization at depth

A total of 51 structural ‘form’ surfaces have been used for the creation of the LR model isoshells and local varying block angles. Table 14-6 summarizes the parameters used to define the Structural Trend for each model.

Table 14-6 Leapfrog trend inputs for creation of 1.5 g/t and 0.4 g/t LR model grade Isoshells

Trend TypeCompatibilityTrend InputsStrengthGlobal Mean Trend
Strongest Along InputsVersion 2All 51 modelled structural ‘form’ surfaces7.0 to 10.0N/A

An example of the resulting Structural Trend is presented in Figure 14-3.

 

 

 

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Figure 14-3 North-facing cross sections showing structural form (18950 mN and 19170 mN)

For the SR model, in-mine and near mine structural control surfaces were created using oriented core structural measurements, combined with underground structural mapping. This structural information was used to create structural trend surfaces (Figure 14-4 and Figure 14-5) which represented the structural geometry of the Au mineralization in the Wassa Mine area. A total of 68 surfaces were used in the Short-Range model and represented the same fold geometries as seen in the LR model, but with greater local definition.

 

Figure 14-4 Structural form surfaces used in the SR model

Table 14-7 summarizes the parameters used to define the Structural Trend for each model.

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Table 14-7 Leapfrog trend inputs for creation of 1.5 g/t and 0.4 g/t SR model grade Isoshells

Trend TypeCompatibilityTrend InputsRangeStrengthGlobal Mean
Trend Direction
Global Mean Trend Ellipsoid Ratios
Strongest Along InputsVersion 268 modelled structural ‘form’ surfaces151550 dip, 270 azi,
20 pitch
Maximum = 3
Intermediate = 2
Minimum = 1

 

Figure 14-5 Images showing the structural control surfaces on sections 19,750 mN and 19,635 mN. The images show the longer, LR model defined control surfaces and the shorter, mine geology defined control surfaces

14.4.4Wassa Indicator Interpolants - Process

Prior to generating the indicator interpolant shells, the raw assay file was composited to 3m, with a minimum end composite length of 1.5m. Any composites less than the end composite length of 1.5 m were not utilized in the interpolation. Indicator interpolants were defined at 0.4 g/t Au and 1.5 g/t Au threshold.

The indicator interpolants were restricted to be within a bounding box defined by the coordinates provided in Table 14-8.

Table 14-8 LR modelling extents

AxisMinimum extentMaximum extent
X39 050 E40 850 E
Y18 200 N20 800 N
Z-775 Z1 100 Z

Table 14-9 summarizes the parameters that were applied to both the 1.5 g/t Au and 0.4 g/t Au models.

Table 14-9 LR Isoshell modelling parameters

Interpolant TypeRangeNuggetIso-ValueResolutionVolumes Excluded
Spheroidal1000.50.352.5m<5000m³

 

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In order to better reflect the geometry of the Au mineralization at a local scale, and also to improve continuity in areas of wider spaced drilling, GSR edited the indicator interpolant shells using indicator polylines. Indicator polylines are digitized and editable strings that carry an associated numeric value which is added to the assay data points on which the interpolant is based. In this instance, indicator polylines with values of 1 (inside), 0 (outside) and indicator iso-value were added to the interpolant. The “iso-value” indicator polylines allow the specific position of the outer limit of the shell to be locally edited. This helps to influence continuity orientations at a smaller scale, ensuring F3 continuity and geometry could be reflected in the resultant domain wireframes in well drilled areas, and assisted in improving the continuity of the model in some of the more sparely drilled areas.

For the SR model, the raw assay intervals were composited down hole to a fixed 2m length. Any residual lengths less than 1m were discarded. A composite cap of 30 g/t Au was applied to the data prior to grade shell contouring.

Grade thresholds for the “mineralized” and the “halo” domains were defined at 1.5 g/t Au and 0.4 g/t Au thresholds. The 1.5 g/t Au cut-off threshold was selected on the basis of a statistical and visual evaluation of the grade distribution. This threshold has been periodically reviewed by GSR and is considered appropriate. The low grade 0.4 g/t threshold has been in previous models Wassa and is considered appropriate to define the low grade material that surrounds the higher grade core at Wassa.

The raw data interpolants were restricted within a bounding box defined by the coordinates provided in Table 14-10.

Table 14-10 SR block model extents

AxisMinimum extentMaximum extent
X39 580 E40 300 E
Y19 330 N20 520 N
Z300 Z1 070 Z

The Leapfrog numeric model used the interpolant parameters shown in Figure 14-6.

Figure 14-6 Short-range isoshell modelling parameters (SRK, 2020)

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14.4.5Wassa Long-Range Model

The final model, displayed in Figure 14-7, constitutes the following:

·>0.4 g/t Au – the iso-surfaced indicator interpolant;
·>1.5 g/t Au - the iso-surfaced indicator interpolant.

Figure 14-7 SE Isometric view of final LR model Leapfrog Isoshells (blue = >0.4 g/t, red = >1.5 g/t)

The two Mineral Resource Isoshells (wireframes) were constructed by GSR geologists with inputs from structural consultants and SRK. These comprise a LG shell and HG shell, corresponding to a 0.4 g/t gold and a 1.5 g/t gold threshold, respectively.

Figure 14-8 shows the two domain shells generated for the SR model. 0.4 g/t Au was used for the “halo” domain threshold, and 1.5 g/t Au was used for the “mineralized” domain.

 

 

 

 

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Figure 14-8 Long section looking East showing the mineralized and halo domain shells on 39,940E (top image). Section on 39,940E (lower image) displaying the same data, cut by that section line, and the assay data used to create the domain shells. RED = mineralized domain; BLUE = halo domain

14.4.6Hwini Butre

The 2020 Mineral Resource estimates for Father Brown (FBZ) and Adoikrom (ADK) were a combined effort by GSR and Resource Modelling Solutions (RMS). A different approach was taken compared to previous and both of these deposits were modelled using a “vein modelling” technique, estimating both vein thickness and grade.

GSR provided RMS with drill hole intercepts indicating hanging wall (HW) main mineralized zone or vein (ADK or FBZ) (HG annotation) or footwall (FW) from and to intervals.

Each vein unit is modelled by estimating the position of the vein and each one of the thicknesses, HW, HG and FW. The position of the vein is defined by the intercept with the top of HW unit, the first thickness is the difference between the intercept of the contact between HW and HG with the top of HW, the second thickness is defined by the difference between the contact HW and HG and the contact between HG and FW and the third thickness is defined by the base of the FW contact.

Intercepts within a horizontal distance tolerance of 2.0m are used to calculate position and thicknesses in order to check any possible relationship between these variables and determine whether or not independent modelling is adequate for the modelling of each vein unit. The scatterplots between each variable showed no significant correlation between the variables, therefore, the independent modelling of each one of these variables in a stepwise manner is deemed appropriate.

 

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The final vein model is defined by stacking the modelled thicknesses below the vein position model. A cross section at U = -28.0m in transformed space is shown in Figure 14-9. Drill hole FBRGC0980075 in FBZ in Figure 14-9 show discrepancies with surrounding data and are challenging to match, however, this modelling workflow works very well on average. Note that these sections are shown in transformed coordinates, after gold estimations models were rotated back to the original Easting-Northing-Elevation coordinates.

 

Figure 14-9 Model section at U=-28.0 in transformed space generated with 2.0 tolerance in V direction

14.4.7Benso

Geology and mineralization domaining was undertaken by GSR. Mineralized wireframes were constructed on 25 meter sections with the 2D polylines being snapped to drill hole grade intercepts using a nominal grade cut-off of 0.5 g/t Au. The 2D polylines were then tied together to create a 3D mesh that was subsequently used for volume and grade estimates. An oxidation surface was created in a similar manner with the depth of weathering being delineated by a polyline on 25 m spaced drill sections and then used to create a mesh surface. This oxidation surface was then used to code the subsequent block models, distinguishing weathered from fresh rock.

The mineralization zones of Benso are structurally controlled with gold emplacement related to the density of quartz veining and sulphide content.

Four estimation domains subdivided by oxidation state have been modelled for Benso, as follows:

·Subriso East (SE);
·Subriso West (SW);
·G-Zone; and
·I Zone.

The SE domain is physically separated from the others and strikes to the north with a dip to the west of between 55-60°. The SW, G Zone and I Zone domains occur in sub-parallel structures and strike to the north-west (320°) with a steep dip of 75-80° to the south-west. Because of this, it was decided to treat the SE deposit as separate for the purposes of grade interpolation.

Only DD and RC drilling has been used for the subsequent grade estimation. The Mineral Resource wireframes and drillholes are shown in Figure 14-10.

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Figure 14-10 Mineral Resource wireframes and drill hole locations for the Benso deposits (GSR, 2010)

 

14.4.8Chichiwelli

Geology and mineralization domaining was undertaken by GSR and the mineralized zone modelling was conducted in a method similar to the Benso deposits.

The mineralization zones of Chichiwelli are structurally controlled with gold emplacement related to the density of quartz veining and sulphide content. The mineralization hosting structures generally trend north-south and dip moderate-steeply to the east at 60°.

Two estimation domains have been modelled for Chichiwelli as follows:

·East Domain; and
·West Domain.

The East and West domains comprise some 10 individually separated wireframe solids.

Wireframes are based on a roughly 0.5 g/t Au grade value. In places composite grades fall below this threshold value but have been included for the sake of maintaining continuity of the wireframe. The style of mineralization seen at Chichiwelli is analogous to deposits observed elsewhere in the Wassa region and, typically for shear zone hosted gold deposits, the mineralization grades tend to pinch and swell within the defined mineralized bearing structures. The Mineral Resource wireframes and drillholes are shown in Figure 14-11.

 

 

 

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Figure 14-11 Mineral Resource wireframes and drillhole locations for Chichiwelli (GSR, 2008)

14.5Statistical Analysis and Variography

As with the previous LR model Mineral Resource estimates for Wassa, GSR contracted SRK (Toronto) to conduct all of the Statistical Analysis and Variography. GSR provided SRK with all of the relevant DD assay data, structural control lines and high and low grade mineralized wire frames (Leapfrog Isoshells).

Table 14-11 summarizes the gold statistics of the assays tagged by mineralized domains provided by GSR. In May 2017, SRK evaluated four drill hole databases for the Wassa Gold Mine, and after discussions with GSR, agreed to combine these databases as conditioning data to be used in grade estimation. This decision has not since been revisited, and all four databases were combined once again.

For consistency with previous models, SRK chose to composite at 3.0-m lengths within the solid wireframes. Unlike previous Mineral Resource models where all composites with length greater than 0.3 m were kept in the estimation database, SRK chose to remove all composites smaller than 1.5 m (or 50% of the composite length). Summary statistics for these composites are also provided in Table 14-11. There is a slight change in the mean grade between assays and composites; however, assay statistics are length weighted, while composite statistics are unweighted since length weights will not be used during grade estimation.

Table 14-11 Summary Gold Statistics of Assays and Composites

ZoneAssaysComposites**
CountMean

Std

Dev*

Min*Max*CoV*CountMean

Std

Dev*

Min*Max*CoV*
HG + LG264,1661.695.840.0011547.973.4593,3871.713.840.001185.072.24
HG59,8604.2711.150.0011547.972.6120,0944.326.950.001185.071.61
LG204,3060.992.710.001442.202.7373,2931.001.780.001152.751.78

* StdDev = Standard Deviation; Min = Minimum; Max = Maximum; CoV = Coefficient of Variation

** Less 1.5 m residual composites, composite statistics are not length-weighted

 

 

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In collaboration with GSR, SRK selected the capping value by comparing probability plots of gold composites on a by-domain basis and plotting the mean grade and the number of affected data by the chosen cap value shown in Figure 14-12. In recognition of the differences in drill density north and south of 19400N, and the potential impact of this spacing on high grade smearing, SRK separated the database into a northern and southern dataset at 19,400 mN.

 

Figure 14-12 Probability Plot for LG (left) and HG (right) Domains North of 19400N (top row) and South of 19400N (bottom row) (SRK, 2020)

In the northern area, SRK chose to cap HG composites at 50 g/t gold and LG composites at 22 g/t gold. These capping thresholds are slightly higher than those chosen in the January 2019 Mineral Resource model; however, given the data density, SRK does not foresee any overestimation issues. In the south, SRK capped the HG composites at 20 g/t gold and LG composites at 15 g/t gold. Table 14-12 compares the statistics for uncapped and capped composite gold grades.

Table 14-12 Comparison of Uncapped and Capped Gold Composite Grades – LR model

ZoneCompositesCapped Composites
CountMeanStdDev*Max*CoV*MeanStdDev*Max*CoV*
HG + LG93,3871.713.840.001185.061.683.20501.90
HG20,0944.326.950.001185.064.215.67501.35
LG73,2931.001.780.001152.750.981.42221.45

* StdDev = Standard Deviation; Min = Minimum; Max = Maximum; CoV = Coefficient of Variation

As with the previous short-range Mineral Resource estimates for Wassa, GSR contracted SRK (Moscow) to conduct all of the mineralized wireframe modelling, Statistical Analysis and Variography. GSR provided SRK with all of the relevant DD assay data and structural control lines .

 

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The composite assay statistics for the SR model have been provided in Table 14-13. The target composite length was 2.0m, but with small length adjustments made to avoid creating shorter composites at the end of intersections. In the few cases where composites with a length less than 1m were generated, these were discarded.

Table 14-13 Comparison of uncapped and capped gold composite grades – SR model

ZoneCompositesCapped Composites
CountMeanStdDev*Max*CoV*MeanStdDev*Max*CoV*
HG24,0234.928.16217.641.664.344.44201.02
LG39,0050.680.6990.261.020.670.56200.85

* StdDev = Standard Deviation; Min = Minimum; Max = Maximum; CoV = Coefficient of Variation

During the kriging estimate, the 2m Au composites were capped at 20g/t Au. This capping value has been reviewed by GSR at various times throughout the Wassa Underground mine’s history. Mine reconciliation data has shown that the 20g/t Au cap value has reconciled an acceptable level, with the capping considered to have a conservative to neutral effect on the grade estimation, depending on which part of the Wassa Underground deposit is being estimated. As such, the capping value remained unchanged during the latest SR model creation. Histograms in Figure 14-13 and Figure 14-14 have been provided, showing the uncapped 2m composite grade distribution for the two domains.

 

Figure 14-13 Histogram showing the uncapped 2m Au composite grade distribution for the mineralized domain

 

 

 

 

 

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Figure 14-14 Histogram showing the uncapped 2m composite grade distribution for the halo domain

14.5.1Wassa Local Angle Models (LR Model)

SRK generated local angles derived from triangulated facets of the structural trend surfaces provided by GSR. This was achieved using Datamine Studio RM, and an initial angle data set for both dip and dip directions. As before, the structural trend surfaces were generated using Leapfrog, and the mesh resolution provided a smooth variation of the dip and dip direction angles.

The angles data set was then used to interpolate a block model of dip and dip directions, which was later called upon for local estimation. The estimation of angles used inverse distance estimation with a power of three, using an isotropic range of 500 m with up to six conditioning angle data. This is consistent with the Mineral Resource models built since 2017.

14.5.2Local Variogram Models

The local estimation approach chosen for the Wassa Gold Mine required the specification of local variogram models. SRK assessed and modelled local variograms for the HG and LG domains, centred about each anchor point. Anchor point locations were reviewed by GSR prior to finalization of their locations. Table 14-14 summarizes the anchor point locations and their local orientations for variogram calculation and modelling. The modelled local variograms for these anchor points are tabulated in Table 14-15. For the LG domain, SRK relied on variograms based on the combined LG and HG capped composites due to the challenges of inferring reliable variograms based solely on LG composites. One reason for the inference challenge may be related to the spatial voids in the database where the HG domain resides. For anchor points 6, 8 and 13, SRK used the HG domain variograms for the LG domain and adjusted the ranges wherever possible to reflect the combined domain variograms.

For each domain (LG and HG), the local variogram parameters (Table 14-15) were then estimated to the block model grid to be read into the grade estimation. In general, the local variograms should be smoothly transitioning within the series. Abrupt changes in grade continuity, within a zone and between anchor point locations, were not expected. Highly localized changes were addressed by the selection of anchor point locations. To ensure smoothness of the local variograms parameters and consistency with the 2015 model, SRK used global kriging with a continuous spherical variogram with ranges of 1,000 by 750 by 500 metres.

 

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An example of the variograms for the HG and LG zones at two anchor points, 1 and 14 are shown in Figure 14-15 and in Figure 14-16.

Table 14-14 Local variogram orientations and anchor point locations

Anchor PointGSLibGemsCoordinates
ANG1ANG2ANG3AzmDipAzmXYZ
175-88075-88-154026318575136.5
279-75079-75-114016418923-322
3270-690270-691804013519155438.5
4285-740285-7419540245195751005.5
5259-770259-771694000519555603.5
6272-610272-611824001519705873.5
7270-750270-751803997519925945.5
8273-850273-851834024519925876.5
9255-770255-771653994519925624.5
10257-370257-371673997520165780.5
11271-750271-751814025520155951.5
12249-650249-651594004520425975.5
13154-410154-41643983520425966.5
14141-460141-46513955520125906.5

Table 14-15 Local variogram models by domain

DomainAPNuggetStructure 1 (Exp)Structure 2 (Sph)
EffectCCAhmaxAhminAhvertCCAhmaxAhminAhvert
LG10.250.58101080.17404025
20.250.5811118.50.17454525
30.20.55241040.2524146
40.20.35102050.4516025021
50.20.6461780.16459022
60.20.358830.45355013
70.20.55148100.25605020
80.20.451220150.3517517555
90.20.67106.50.2509022
100.30.529.51160.18286025
110.20.58252580.2218518555
120.250.65151512.50.1909012.5
130.20.6881280.12205010
140.20.688860.12404020
HG10.250.5810106.50.17353525
20.250.6811118.50.07252515
30.30.25242450.45323214
40.20.35102050.4516025021
50.20.6781880.135510022
60.20.45101030.35302213
70.20.525880.3605512
80.20.35353550.4517517555
90.20.656.51160.153010035
100.20.626.54.570.18255540
110.20.533131120.2718518555
120.250.3815156.50.37110257
130.30.68101080.02757530
140.30.68101080.02757530

 

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Figure 14-15 HG Variogram from anchor point 1 (SRK, 2020)

 

Figure 14-16 LG Variogram from anchor point 14 (SRK, 2020)

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Local variogram models were not used in the SR model. The SR model used a different locally-variable-anisotropy (LVA) technique. Each location in the block model was assigned azimuth, dip and plunge information based on the structural control surfaces and the overall variogram model. This orientation information was then called upon during the estimation process to set the orientation of the variogram model and search neighbourhood each block grade estimate.

The variogram parameters for the HG & LG mineralized domains were set as shown in Figure 14-17.


Figure 14-17 Variogram for the short-range HG & LG mineralized domains (SRK, 2020)

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14.5.3Hwini Butre Father Brown – Adoikrom

The statistics were preformed on each of the domains for each individual zone, FBZ (HW, HG and FW) and ADK (HW, HG and FW)

The descriptive statistics for the individual modelled domains are summarized in Table 14-16.

Table 14-16 Descriptive statistics for Hwini Butre modelled domains (uncapped & capped)

ZoneDomainCappingCountMinimumMaximumMeanstdevCOV
AdoikromHWUncapped8070.0013.421.311.2650.965
Capped8070.005.001.000.7930.791
HGUncapped9460.09136.387.5810.3001.359
Capped9460.0023.005.905.0880.863
FWUncapped8550.0012.011.191.0580.893
Capped8550.005.000.950.8020.843
Father Brown ZoneHWUncapped1,1300.0137.620.811.7962.214
Capped1,1300.015.000.600.7411.231
HGUncapped1,2070.01253.0011.4117.2721.513
Capped1,2070.0146.009.2810.8271.167
FWUncapped1,1200.0040.101.022.4582.402
Capped1,1200.005.000.740.7891.068

Probability plots for each vein unit for each domain are generated and shown in Figure 14-18. The capping values selected from the probability plots are summarized in Table 14-17.

 

Figure 14-18 Gold grade probability plot with outliers and far out thresholds highlighted (RMS, 2020)

 

 

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Table 14-17 Capping values selected from analysis of the probability plot

DepositVein unitCapped Values used
FBZFW5
FBZHG46
FBZHW5
ADKFW5
ADKHG23
ADKHW5

The variography was performed for each deposit and for each vein unit (FBZ & ADK) using capped composites. Experimental variograms are calculated for full range of possible azimuths with steps of 15 degrees totaling 24 directions. The direction with most continuous experimental points from visual inspection of the 24 directions were utilized for nugget inference. These directions may not coincide with the final major continuity direction when considering all experimental points for final model fit. The nugget is inferred by fitting a single structure spherical variogram to the first few (up to three) experimental variogram points.

The variogram nugget inference for all vein units in FBZ are shown in Figure 14-19. The directions utilized for nugget inference are detailed in each plot in Figure 14-19.

 


Figure 14-19 Inferred nugget effect for gold grade in each vein unit for FBZ deposit (RMS, 2020)

 

The experimental variogram and fitted model for FW unit in FBZ is shown in Figure 14-20. The parameters of the fitted model are summarized in Table 14-18. The experimental variogram and fitted model for HG unit in FBZ is shown in Figure 14-21. The parameters of the fitted model are summarized in Table 14-19. The experimental variogram and fitted model for HW unit in FBZ is shown in Figure 14-22. The parameters of the fitted model are summarized in Table 14-20.

 

 

 

 

 

 

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Figure 14-20 Fitted experimental variogram points for gold grade in FW for FBZ deposit (RMS, 2020)

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Figure 14-21 Fitted experimental variogram points for gold grade in HG for FBZ deposit (RMS, 2020)

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Figure 14-22 Fitted experimental variogram points for gold grade in HW for FBZ deposit (RMS, 2020)

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Table 14-18 Fitted variogram parameters for gold grade in FW for FBZ deposit

 NuggetStructure 1Structure 2
Contribution0.0000.8510.149
Model Shape exponentialexponential
Angle 1 43.343.3
Angle 2 0.00.0
Angle 3 0.00.0
Range 1 10.0155.9
Range 2 16.310.0
Range 3 1.01.0

Table 14-19 Fitted variogram parameters for gold grade in HG for FBZ deposit

 NuggetStructure 1Structure 2
Contribution0.2500.4160.334
Model Shape exponentialexponential
Angle 1 29.129.1
Angle 2 0.00.0
Angle 3 0.00.0
Range 1 10.077.3
Range 2 10.050.0
Range 3 1.01.0

Table 14-20 Fitted variogram parameters for gold grade in HW for FBZ deposit

 NuggetStructure 1
Contribution0.2830.717
Model Shape exponential
Angle 1 72.6
Angle 2 0.0
Angle 3 0.0
Range 1 10.0
Range 2 10.0
Range 3 1.0

The variogram nugget inference for all vein units in ADK is shown in Figure 14-23. The directions utilized for nugget inference are detailed in each plot in Figure 14-23 and the experimental variogram and fitted model for FW unit in ADK is in Figure 14-24. The parameters of the fitted model are summarized in Table 14-21. The experimental variogram and fitted model for HG unit in ADK is shown in Figure 14-25 and the parameters of the fitted model in Table 14-22. The experimental variogram and fitted model for HW unit in ADK is shown in Figure 14-26. The parameters of the fitted model are summarized in Table 14-23.

 

Figure 14-23 Inferred nugget effect for gold grade in each vein unit for ADK deposit (RMS, 2020)

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Figure 14-24 Fitted experimental variogram points for gold grade in FW for ADK deposit (RMS, 2020)

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Figure 14-25 Fitted experimental variogram points for gold grade in HG for ADK deposit (RMS, 2020)

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Figure 14-26 Fitted experimental variogram points for gold grade in HW for ADK deposit (RMS, 2020)

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Table 14-21 Fitted variogram parameters for gold grade in FW for ADK deposit

 NuggetStructure 1Structure 2
Contribution0.0000.0620.938
Model Shape exponentialexponential
Angle 1 77.877.8
Angle 2 0.00.0
Angle 3 0.00.0
Range 1 15.015.0
Range 2 929.515.0
Range 3 1.01.0

Table 14-22 Fitted variogram parameters for gold grade in HG for ADK deposit

 NuggetStructure 1Structure 2
Contribution0.0970.4600.443
Model Shape exponentialexponential
Angle 1 13.113.1
Angle 2 0.00.0
Angle 3 0.00.0
Range 1 32.821.9
Range 2 38.910.0
Range 3 1.01.0

Table 14-23 Fitted variogram parameters for gold grade in HW for ADK deposit

 NuggetStructure 1Structure 2
Contribution0.0000.6500.350
Model Shape exponentialexponential
Angle 1 73.173.1
Angle 2 0.00.0
Angle 3 0.00.0
Range 1 15.058.4
Range 2 15.099.9
Range 3 1.01.0

The major direction of continuity for each variogram model for each deposit is inferred from the weighted ranges of each variogram structure utilizing their contribution as weights. The major direction is rotated back to original space and the results are summarized in Table 14-24.

Table 14-24 Fitted major variogram directions in original space

DepositDomainAzimuthDipWeighted Anisotropy
FBZFW118.926.52.1
FBZHG131.618.51.4
FBZHW86.938.51.0
ADKFW172.910.94.8
ADKHG4.111.71.1
ADKHW170.715.11.5

 

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14.5.4Benso

The statistics are based on composited assay values within the wireframes modelled by GSR; the data was composited to 2 m lengths within the mineralized zones, and composites of less than 1.50 m were removed.

The descriptive statistics for the individual modelled domains, split by oxidation state, are summarized below in Table 14-25. The transition zone is relatively thin, and so has not been analysed separately. For all datasets, zero values were checked in the database, and were set to 0.001 g/t.

Table 14-25 Descriptive statistics for Benso modelled domains (capped)

DomainOxidationCountMinimumMaximumMeanVarianceCOV
Subriso EastOxide2660.00130.812.1115.181.85
Fresh6490.00151.582.5425.491.99
Total 9150.00151.582.4222.511.96
Subriso WestOxide360.4115.863.1414.111.20
Fresh5710.001223.833.88147.393.13
Total 6070.001223.833.83139.483.08
G ZoneOxide440.00121.152.7618.691.57
Fresh5700.00152.332.0411.11.63
Total 6140.00152.332.0911.641.63
I ZoneOxide110.211.510.970.230.49
Fresh860.1118.182.7210.961.22
Total 970.1118.182.5210.041.26

The four areas were combined into two areas for estimation purposes; namely Subriso East, and Subriso West, G Zone and I Zone combined. The Subriso East domain is separated from the Subriso West, G Zone and I Zone areas, and strikes roughly north-south, with a dip to the west of between 55 and 60°. The Subriso West, G Zone and I Zone areas lie in sub-parallel structures, striking roughly to the north-west (320°), with a steep dip of 75 to 80° towards the south-west. The descriptive statistics for the two separate estimation domains are shown below in Table 14-26.

Table 14-26 Descriptive statistics for simplified Benso modelled domains (capped)

DomainCountMinimumMaximumMeanVarianceCOV  
Subriso East9150.00151.582.4222.511.96  
  
Subriso West, G Zone and I Zone13180.001223.832.9371.052.88  
  

Statistical distributions for the two domains are similar, with the histograms indicating that the distribution is not normal, being highly negatively skewed. The log transformed gold grade data demonstrates there may be several populations within the distribution and that the distribution approached log-normality.

HG caps were applied to the composite data as follows:

·Subriso East: 40 g/t cap; and
·Subriso West, G Zone, I Zone: 60g/t cap.

The estimation data sets noted above were used to derive variograms for estimation. In all cases, the grade block model for each individual modelled solid was estimated using only the composites inside that solid.

Variography was undertaken on the log transformed data, with a short lag, omnidirectional, downhole variogram used to derive the nugget effect. Directional variograms were then calculated within a rotated plane aligned with the strike and dip of the modelled solids. The variogram parameters derived from the modelled variograms are shown in Table 14-27. Variograms were back transformed before use in OK.

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Table 14-27 Variogram parameters for the Benso zones

ParameterSubriso EastSubriso West, G Zone and I Zone
Co0.250.19
C10.410.54
C20.340.27
a1 (strike)2020
a1 (dip)158
a2 (strike)5050
a2 (dip)4030
14.5.5Chichiwelli

The statistics are based on composited assay values domained within the mineralization wireframes described previously, with sample data composited to 2 m lengths within the mineralized zones.

The statistics presented here are based on all drilling data that intersect the wireframes. The composites inside the modelled bodies were also split into oxidation states, but as there was little information for the transition zone, SRK combined the three oxidation states and used the combined oxidations datasets throughout the statistical and geostatistical studies, and the subsequent grade estimation.

The descriptive statistics for the two separate estimation domains are shown below in Table 14-28.

Table 14-28 Descriptive statistics for Chichiwelli modelled domains (capped)

DomainCountMinimumMaximumMeanVarianceCOV
East4180.00141.11.7517.642.41
West5590.00146.31.6910.141.89

HG capping was applied to both the East and West domains. The HG caps were determined on the basis of the shape of the tail of the log histogram and the log probability plots. Capping reduces the extreme values to a nominated capped value, which affects the mean grades of the 2.0 m composites, as indicated by Table 14-29.

Table 14-29 Chichiwelli high grade capping

DomainCap AppliedMean Grade
before Cap
Mean Grade after CapPercentage difference
(g/t)(g/t)(g/t)(%)
East251.751.65-6.06
West151.691.59-6.29

The estimation data sets noted above were used to derive variograms for estimation. In all cases, the grade block model for each individual modelled solid was estimated using only the composites inside that solid.

Variograms were modelled for the East and West domains separately. Variography was attempted for the individual solids, but the resultant variograms were unable to be modelled. Raw variography resulted in difficult to model variograms, and so a Gaussian transformation was applied to the data. The first stage was to define the nugget effect from a short-lag omnidirectional variogram, which is calculated along the drillhole, and then to model the variogram ranges from directional variograms from along strike, down-dip and across dip directions. The directional variograms are then back transformed into “raw” space and used for subsequent estimation. The back transformed variograms and resultant variogram parameters are included in Table 14-30.

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Table 14-30 Variogram parameters for Chichiwelli zones

ParameterEastWest
Co7.943.06
C13.601.59
Nugget Effect (%)68.865.81
Range (m)  
a1 (strike)2540
a1 (dip)2535
a1 (normal to strike)84.7
14.6Block Model and Grade Estimation
14.6.1Wassa Long-Range Block Model

A 3D block model including rock type, gold, percent mineralization, density and class was constructed for each of the modelling areas, Wassa short-range and long-range. The selection of the block size was driven by the borehole spacing and mainly by the geometry of the auriferous zones, but also based on mining parameters and in accordance with the previous Mineral Resource estimate. The LR models block size was set at 10 x 10 x 5 m in the northing, easting and elevation directions, respectively along the mine grid. The block model origins can be seen in Table 14-31.

Table 14-31 Wassa LR model definitions, upper left hand corner coordinates

 Block SizeOrigin*Block Count
(m)(m)
X1039,050180
Y1018,200260
Z51,100375

* Coordinates relative to mine grid.

A percent block model was used to evaluate tonnages. Tonnage for each respective block was obtained by weighting volumes corresponding to the interpreted auriferous zones and the respective mean SG defined by weathering profile.

The block model bulk density data was coded based on weathering surface which was built to define oxide material from fresh material. The weathering surface defined the ‘top of fresh’ material; all blocks above the ‘top of fresh’ surface were designated as oxide and material below the surface as ‘fresh’. The bulk density values assigned to the block model were based on series of measurements made over the various exploration phases going back to the initial GSR exploration program in 2002. The density values used for the tonnage estimate were provided by GSR and are detailed below in Table 14-32.

Table 14-32 Average Bulk Density used for LR model

Weathering TypeAvg Bulk Density t/m3
Oxide1.8
Fresh2.8

For the SR model, the Surpac 3D block model contained the 20 g/t Au capped composite grade estimate, the bulk density coding, the domain coding, the depletion coding, and the Mineral Resource classification coding. The block size for estimation is effectively 5m cubes, with sub-blocking to 2.5 mN, 1.0 mE and 2.5 mRL. The block size was based on the Measured Mineral Resource drill spacing at Wassa. The Surpac block model is structured with a 2.5 mN, 1.0 mE, and 2.5 mRL block size, and no sub-blocking, in order to facilitate transfer of the block model between different software. The Surpac block model contained a total of 9,041,540 blocks, with the origins outlined in Table 14-33.

 

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Table 14-33 Wassa SR model definitions

CoordinateOriginBlock Size (m)No.  of Blocks
X39,6001272
Y19,3502.5460
Z3202.5292

As the Wassa Underground mine exists well below the bottom of the partial oxidation layer, the bulk density in the model has been set to 2.8 (fresh rock). The 2.8 value comes from test work performed by both GSR and Intertek during 2017 and 2018.

14.6.2Wassa Mineral Resource Estimation Methodology

For the LR model, SRK implemented the same methodology since 2018 to construct the Mineral Resource model, using ordinary kriging with local varying angles and local variograms for the estimation of gold grades. The general steps required to implement the approach are:

·Construct locally varying angles models for dip and dip direction;
·Calculate and model local variograms for each series and interpolate these local variograms to construct a model of local variogram model parameters;
·Estimate gold grades using ordinary kriging, calling upon the local models of dip, dip direction, and variogram models; and
·Check estimated model using qualitative and quantitative methods.

Table 14-31 summarizes the block model definition used for the model area using the mine grid. No rotation was applied. GSR opted to change the vertical size of a block from 3 m to 5 m to improve alignment of the model to mine elevations. The vertical extent of the model has increased to encompass the mineralization delineated by the deeper southern exploration boreholes.

The following sections summarize the method(s) used, assumptions made, and results obtained for each of the four modelling steps.

For the SR model, the Mineral Resource estimation methodology involved:

·The construction of structural control surfaces that represented the orientation of the mineralized structures in the local area;
·The generation of domains shells from the composited assay data, influenced by the structural control surfaces;
·Coding of the block model framework with the domain shell information;
·Estimation of the Au grade into the block model, using the azimuth and dip information collected from the structural control surfaces;
·The structural information was used to rotate the variogram model into the local mineralized trend orientation. Estimation was performed using ordinary kriging;
·Coding of the model using solids created by the Wassa Mine Geology group to define Measured and Indicated Mineral Resources. The classification was based on the drill density observed on a section by section basis; and
·Depletions and validation of the model.
14.6.3Wassa Grade Estimation

Using the models of local angles and local variograms, SRK performed the grade estimation using ordinary kriging methodology. The LG and HG estimation used the parameters in Table 14-34. The parameters differ from previous models and are based on an estimation sensitivity analysis conducted in January 2020. The selection of an appropriate set of estimation parameters was based on ensuring a good quantile-quantile comparison of the resultant estimated grades distribution to change of support corrected distributions for each of the LG and HG domains. This should ensure an appropriate level of smoothness.

 

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Table 14-34 LR model Estimation Parameters

PassCompositesMaximum Composites per BoreholeSearch EllipseGSLIB  
  
Min.MaxSvxSvySvzA1A2A3  
(m)(m)(m)  
1382505050270-500  
22124909050270-500  
3115-200200200270-500  

The first estimation run required at least two holes with the aim to localize grade estimates. The second pass was slightly more relaxed, requiring fewer samples found within a larger search radius. The third estimation pass considered search ellipses sized at least twice the variogram ranges, with the aim of estimating most of the blocks unvisited by the first two passes. As the estimation considered a stationary search ellipsoid, these ranges were selected to ensure that local estimation yielded estimates that conformed to the local anisotropy and local variograms.

After estimation of each domain, the LG and HG domain grades were then combined into a single block grade based on a percentage weighted average of the estimated grade based on fill volume of the respective Mineral Resource wireframes. These single block grades were used to generate the swath plots.

For the SR model, SRK (Moscow) performed the grade estimate using same estimation methodology employed at the underground mine since the start of production. Hard boundaries were used to prevent mineralized domain and halo domain information from mixing. Composites from inside the halo domain were only allowed to influence the grade estimation inside the halo wireframe, whilst composites from inside the mineralized domain were only allowed to influence the grade estimation inside the mineralized wireframe. Grade estimation was performed using Ordinary Kriging. The parameters for the estimation have been provided in Table 14-35.

Table 14-35 SR model estimation parameters

PassCompositesMaximum composites per BoreholeSearch Ellipse  
  
MinMax

Long

mX

Intermediate

mY

Short

mZ

  
16185603015  
2424No Limit1206030  
31245200200100  

The variogram model was oriented along 270° azimuth, 50° dip, 20° plunge to the south.

14.6.4Hwini Butre Father Brown – Adoikrom

The estimation is performed using ordinary kriging with uncapped and capped gold grades. The number of composites and maximum search radius utilized for each vein unit in each deposit are shown in Table 14-36. The influence of outliers is visually evident in the HG unit.

Table 14-36 Kriging search parameters for each vein unit in each deposit

DepositVein UnitMaximum Search (m)Maximum Composites
FBZHW2508
FBZHG5004
FBZFW5004
ADKHW2504
ADKHG100024
ADKFW10002

 

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Two block models were produced one for FBZ the other for ADK. No rotation was applied to the models. Block sizes were chosen to reflect the geometry of the deposits. Grade data for each of the modelled units was interpolated into the individual structures only. Block model parameters for FBZ and ADK are summarized in Table 14-37 and Table 14-38.

Table 14-37 Father Brown block model parameters

CoordinateOriginBoundary sizeBlock Size (m)
X175681.4714431
Y32345.736832
Z1176.724822

 

Table 14-38 Adoikrom Zone block model parameters

CoordinateOriginBoundary sizeBlock Size (m)
X175731.387181
Y32394.438042
Z1271.617212

The density values used for the tonnage estimate were provided by GSR, and are detailed in Table 14-39.

Table 14-39 Hwini Butre rock density

Oxidation StateValue (t/m3)
Fresh2.7

 

14.6.5Benso

A block model was produced for the whole Benso area. No rotation was applied to the model. Block sizes were chosen to reflect the average spacing of drill lines along the strike. Grade data for each of the modelled units was interpolated into the individual structures only, with soft boundaries between oxidation states, and subsequently reported as oxide or fresh. Block model parameters for Benso are summarized in Table 14-40.

Table 14-40 Benso block model parameters

CoordinateOriginBlock Size (m)No.  of Blocks
X17375012.5300
Y5600025160
Z12051060

Block grades for each of the mineralized zones were estimated using OK. OK was carried out in four passes for each mineralized zone, and the search parameters for the individual domains are shown in Table 14-41. The discretization grid was set at 5 x 2 x 1 (xyz) in all cases. The search ellipsoids are relatively large compared to the variogram ranges, but as there is quite a high data density the blocks were usually estimated with data significantly closer than the edges of the ellipsoid.

 

 

 

 

 

 

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Table 14-41 Benso ellipsoid search neighbourhood parameters

DomainSearch 1Search 2
Subriso EastX100200
Y80160
Z2040
Min.  Samples44
Max.  Samples3636
Subriso West, G Zone and I ZoneX100200
Y80160
Z2040
Min.  Samples44
Max.  Samples3636

GSR modelled the oxidation surface to determine the boundary between oxide and fresh material. No transition zone was modelled. The density values used for the tonnage estimate were provided by GSR and are detailed in Table 14-42.

Table 14-42 Benso rock density

Oxidation StateValue (t/m3)
Oxide1.8
Fresh2.7
14.6.6Chichiwelli

A block model was produced for the whole Chichiwelli area. No rotation was applied to the model. Block sizes were chosen to reflect the average spacing of drill lines along the strike. Grade data for each of the modelled units was interpolated into the individual structures only, with soft boundaries between oxidation states, and subsequently reported as oxide or fresh. Block model parameters for Chichiwelli are summarized in Table 14-43.

Table 14-43 Chichiwelli block model parameters

CoordinateOriginBlock Size (m)No.  of Blocks
X631,093.6412.5100
Y580,787.202560
Z1216 (max)865

Block grades for each of the mineralized zones were estimated using OK. OK was carried out in four passes for each mineralized zone, and the search parameters for the individual domains shown below in Table 14-44. The discretization grid was set at 5x2x1 (xyz) in all cases. The search ellipsoids are relatively large compared to the variogram ranges, but as there is quite a high data density, the blocks were usually estimated with data significantly closer than the edges of the ellipsoid. Octants were used on the 1st and 2nd pass searches with three consecutive empty sectors, however they were not applied on the 3rd search pass, hence the same number of minimum and maximum samples for the 2nd and 3rd searches.

 

 

 

 

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Table 14-44 Chichiwelli ellipsoid search neighbourhood parameters

Domain Search 1Search 2Search 3Rotation Parameters
EastX60120120Azimuth: 20
Y60120120Dip: 60
Z204040-
Min.  Samples333-
Max.  Samples808080-
WestX80160160Azimuth: 20
Y80160160Dip: 60
Z102020-
Min.  Samples333-
Max.  Samples808080-

GSR modelled the oxidation surface to determine the boundary between oxide and fresh material. No transition zone was modelled. The density values used for the tonnage estimate were provided by GSR and are detailed in Table 14-45.

Table 14-45 Chichiwelli rock density

Oxidation StateValue (t/m3)
Oxide1.8
Fresh2.68
14.7Model Validation and Sensitivity
14.7.1Wassa

SRK checked the resultant LR block model by considering: (1) visual comparisons of block grades and nearby composites via a sectional approach; (2) swath plots for the combined LG and HG domains along northing, easting and a vertical swath; and (3) change of support checks. Sectional checks showed good consistency between the informing data and local estimated blocks, and also good conformity of grade trends to the local folds in the mineralization. In general, the swath plots showed that in areas of abundant data, the model matches well with the composite grades in that moving average window. Mismatches in the informing composites and the average block grades are attributed to those regions of the model that are sparsely sampled, specifically in the southern extent of the mineralized zone as shown in Figure 14-27.

 

 

 

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Figure 14-27 South-North Swath Plot Comparing Estimated Grades and Informing Capped Composites (SRK, 2020)

Histogram bars correspond to block tonnages in the February 2020 long-range Mineral Resource model South of 19,400 North, the estimated grades and the composite grade profiles are more erratically behaved. This is attributed to the presence of fewer composites with some very high-grade intersections, and the continuity of the grade shells in the southern area. Mineral Resources in the southern portion of the estimate have all been classified as Inferred Mineral Resources to reflect the lower confidence.

SRK anticipates that additional drilling in this area may impact the continuity of the grade domains and dampen the influence of these higher-grade intervals. SRK understand that GSR have conducted subsequent studies comparing 2018 to 2020 long- and short-range models; these comparisons demonstrate that year on year, the impact of additional drilling has historically increased the Inferred and Indicated Mineral Resources.

SRK also compared the ordinary kriging block model distribution with the declustered, change-of-support corrected distribution of the informing composites for the LG and HG domains (Figure 14-28). Declustering mitigates the influence of preferential sampling of borehole data; this often results in a distribution of composites whose mean statistic is often comparable to that of the estimated model. Further, a change-of-support correction is applied to account for the volume difference between the composite scale and the final block volume scale. Figure 14-28 shows the quantile-quantile comparison of the gold distribution from the block model and the expected grade distribution following declustering and change-of support corrections for the LG and HG domains. Overall, the mean grades from the block model are reasonably close to those predicted from declustering. The quantile-quantile plot shows that the block model is comparable to that predicted by the change-of-support for the HG domain, and slightly smoother than predicted for the LG domain.

The preliminary block model was delivered on February 28, 2020 for further review by GSR. The sub-blocked Surpac model was delivered on March 3, 2020.

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Figure 14-28 Quantile-Quantile Comparison of Block Model Grades to Declustered Change-of-Support Corrected Gold Distributions for LG (left) and HG (right) domains (SRK, 2020)

For the SR model, validation included the following processes:

·Sectional visual check to ensure that the mineralized domain was completely enclosed by the halo domain;
·Sectional visual check to ensure that the mineralized domain enclosed continuous high grade assay values and that the halo domain enclosed the lower grade isolated assay values;
·Sectional visual check to ensure that the domain shells were influenced by the structural control surfaces;
·Sectional visual check to ensure that composite assay values and block model grades were consistent with each other;
·Volume check, to ensure coded block model volumes are close to original wireframe volumes;
·Check that capping of the 2m composite assay values was performed correctly;
·Check between the block model and the 2m capped composite values statistics for similar values;
·Swath plots of the block grades versus the capped 2m composite values in the Easting, Northing and Elevation directions to ensure that the block grade estimate was valid; and
·Comparison between the latest short-range block model tonnes and grade and the previous short-range block model tonnes and grade. Any difference in tonnes and grade was attributed to the addition of new drilling.

Swath plots reporting the comparison between the block model estimated grade and the capped 2m composite grade has been provided in Figure 14-29, Figure 14-30 and Figure 14-31.

 

 

 

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Figure 14-29 SWATH plot in the E-W direction X dimension. Blue line represents Block model Estimated grades and Red is 2m drill hole composites grades (SRK, 2020)

Figure 14-30 SWATH plot in the N-S direction Y dimension. Blue line represents Block model Estimated grades and Red is 2m drill hole composites grades (SRK, 2020)

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Figure 14-31 SWATH plot in the elevation direction Z dimension. Blue line represents Block model Estimated grades and Red is 2m drill hole composites grades (SRK, 2020)

The data, as presented in the swath plots, showed that the block estimation respected the underlying drill hole composite dataset and that the block model estimate was valid.

14.7.2Hwni-Butre
14.7.2.1Thickness models

In order to validate the thickness estimates a nearest neighbor model is generated for the thickness of each vein unit of for each deposit. Swath plots comparing the data distribution, nearest neighbor estimates and kriging estimates are generated. Table 14-46 shows a summary comparing the global mean of each model for FBZ. The difference ranges from -6.34% to 7.67%.

Table 14-46 Global mean comparison between nearest neighbor and kriged thickness models for FBZ.

FBZ VariableNN Mean
(m)
Kriging Mean
(m)
% Difference
HW Thickness1.541.45-6.34
HG Thickness1.281.397.67
FW Thickness1.231.251.89

Table 14-47 shows a summary comparing the global mean of each model for ADK. The difference ranges from 0.93% to 12.41%.

Table 14-47 Global mean comparison between nearest neighbor and kriged thickness models for ADK. Variable

ADK VariableNN Mean (m)Kriging Mean (m)% Difference
HW Thickness1.922.087.63
HG Thickness1.931.950.93
FW Thickness1.151.3112.41

 

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14.7.2.2Gold models

The gold grade data reproduction is also checked. The scatterplot of measured and estimated gold grades at data locations for FBZ and ADK are shown in Figure 14-32. The least reliable estimates for FBZ are listed in Table 14-48. The least reliable estimates for ADK are listed in Table 14-49. The scatter plots for both FBZ and ADK indicate good data reproduction, Figure 14-32. Although the estimates, when compared to the raw (uncapped) drill hole assay data are less accurate than the drill assay, all but one is under stating the grade. This means the estimate is conservative in these areas.

 

Figure 14-32 Measured and estimated gold grades at data locations (RMS, 2020)

 

Table 14-48 List of least reliable estimates FBZ

Hole idMin ZonesAUModifiedEstimateError
A FBRGC0950072HG166.7288.26-78.46
B FBRGC0980109HG192.00122.63-69.37
C FBZDD059HG253.00196.52-56.48
D FBRGC0950039HG140.6092.76-47.84
E FBRGC0980164HG91.9251.63-40.29

 

Table 14-49 List of least reliable estimates ADK

Hole idMin ZonesAUModifiedEstimateError
A ADKGC164HG8.7157.9749.26
B ADKGC042HG106.7159.87-46.84
C ADK-62HG136.38101.67-34.71
D ADKGC0960037HG104.9273.22-31.70
E ADK-71HG8.1733.4825.30

In addition to the validation above a nearest neighbor model was generated for each vein unit for each deposit for uncapped and capped gold grades. Swath plots comparing the data distribution, nearest neighbor estimates and kriging estimates were generated. The swath plots for HG units in the FBZ zone are shown in Figure 14-33. Table 14-50 shows a summary comparing the global mean of each model for FBZ. The difference ranges from -3.8% to 8.3%.

 

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Figure 14-33 Swath plot comparison of composites, nearest neighbor estimates and kriging estimates for uncapped and capped grades in HG for FBZ deposit (RMS, 2020)

 

Table 14-50 Global mean comparison between nearest neighbor and kriged Gold models for FBZ

VariableVein Unit NNMeanKriging Mean% Difference
AUModifiedFW0.330.368.28
AUModified3FW0.300.301.72
AUModified5FW0.310.323.38
AUModifiedHG4.204.516.92
AUModified30HG3.714.007.39
AUModified46HG3.834.147.41
AUModifiedHW0.330.32-3.17
AUModified3HW0.330.31-3.79
AUModified5HW0.330.32-3.48

A nearest neighbor model was generated for the ADK estimate for each deposit for uncapped and capped gold grades. Swath plots comparing the data distribution, nearest neighbor estimates and kriging estimates were generated and plots for HG units are shown in Figure 14-34. Table 14-51 shows a summary comparing the global mean of each model for ADK. The difference ranges from -043% to 3.09%.

 

Figure 14-34 Swath plot comparison of composites, nearest neighbor estimates and kriging estimates for uncapped and capped grades in HG for ADK deposit (RMS, 2020)

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Table 14-51 Global mean comparison between nearest neighbor and kriged Gold models for ADK

VariableVein Unit NNMeanKriging Mean% Difference
AUModifiedFW0.580.581.14
AUModified3FW0.570.581.28
AUModified5FW0.580.581.13
AUModifiedHG5.035.193.09
AUModified16HG4.774.75-0.43
AUModified23HG4.874.931.26
AUModifiedHW0.910.920.85
AUModified4HW0.90.910.93
AUModified5HW0.90.910.91

Swath plots comparing the nearest neighbor estimates and kriging estimates for the densely sampled area for FW, HG and HW units in FBZ deposit were created. The swath plot restricted to the densely sampled areas does not show bias when compared to the nearest neighbor estimates for any of the vein units in the FBZ deposit. Table 14-52 shows a summary comparing the global mean of each model for FBZ. The difference ranges from -0.9% to 2.2%.

Table 14-52 Global mean comparison between nearest neighbor and kriged Gold models for FBZ within the densely sampled area

VariableVein UnitNN MeanKriging Mean% Difference
AUModifiedFW0.770.781.23
AUModified3FW0.710.721.29
AUModified5FW0.730.741.4
AUModifiedHG10.3910.29-0.92
AUModified30HG8.588.580.04
AUModified46HG9.419.38-0.29
AUModifiedHW0.620.642.24
AUModified3HW0.570.580.57
AUModified5HW0.590.590.89

Swath plots comparing the nearest neighbor estimates and kriging estimates for the densely sampled area for FW, HG and HW units in ADK deposit were also constructed. The swath plot restricted to the densely sampled areas does not show bias when compared to the nearest neighbor estimates for any of the vein units in the ADK deposit. Table 14-53 shows a summary comparing the global mean of each model for ADK. The difference ranges from -1.9% to 0.5%.

Table 14-53 Global mean comparison between nearest neighbor and kriged Gold models for ADK within the densely sampled area

VariableVein Unit NNMeanKriging Mean% Difference
AUModifiedFW0.990.990.51
AUModified3FW0.960.960.11
AUModified5FW0.970.970.33
AUModifiedHG6.396.38-0.13
AUModified16HG5.525.530.08
AUModified23HG5.845.840.04
AUModifiedHW1.111.09-1.89
AUModified4HW1.061.05-0.86
AUModified5HW1.071.06-0.99

 

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In order to determine how the estimate performed in areas of sparse drill hole data Swath plots comparing the nearest neighbor estimates and kriging estimates for FW, HG and HW units in FBZ were created. The swath plot restricted to the sparsely sampled areas is consistently lower for nearest neighbor estimates for the FBZ deposit except for the HG unit at high U coordinate. High U coordinate correspond to extrapolation at deeper area of the deposit. Table 14-54 shows a summary comparing the global mean of each model for FBZ. The difference ranges from -4.8% to 9.6%.

Table 14-54 Global mean comparison between nearest neighbor and kriged Gold models for FBZ within the sparsely sampled area

VariableVein Unit NNMeanKriging Mean% Difference
AUModifiedFW0.290.329.57
AUModified3FW0.270.271.80
AUModified5FW0.270.283.77
AUModifiedHG3.734.078.43
AUModified30HG3.343.658.71
AUModified46HG3.413.748.88
AUModifiedHW0.310.30-4.46
AUModified3HW0.310.29-4.79
AUModified5HW0.310.30-4.50

Swath plots comparing the nearest neighbor estimates and kriging estimates for the sparsely sampled area for FW, HG and HW units in ADK deposit were also created. The swath plot restricted to the sparsely sampled area shows consistently lower nearest neighbor estimates for the HW unit for the ADK deposit. The nearest neighbor estimate is mostly higher for the FW unit. The HG unit for ADK show reasonable match with exception of high U coordinates. The high U coordinates correspond to the deeper zones of the deposit. Table 14-55 shows a summary comparing the global mean of each model for ADK. The difference ranges from -0.5% to 1.4%.

Table 14-55 Global mean comparison between nearest neighbor and kriged Gold models for ADK within the sparsely sampled area

VariableVein Unit NNMeanKriging Mean% Difference
AUModifiedFW0.550.561.20
AUModified3FW0.550.551.39
AUModified5FW0.550.561.21
AUModifiedHG4.955.123.31
AUModified16HG4.734.71-0.46
AUModified23HG4.824.881.34
AUModifiedHW0.900.911.04
AUModified4HW0.900.911.04
AUModified5HW0.900.911.04

It is the opinion of the QP that the validation exercises of the block model above show that the estimate is robust and accurate with errors within acceptable ranges.

14.7.3Benso

The block models were validated by comparing the block model mean grades with the declustered composite mean grades and through validation slices through the block models.

The mean grades for each of the estimated block models were compared to the declustered mean grade for the composite input data. Each of the modelled zones was compared separately. The differences between the declustered mean composite grades and the block grades are relatively small, indicating that the model is similar to the input data on a global scale.

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The block model was also compared to the composite grades within defined sectional criteria in a series of validation slices, the results of which are displayed on graphs to check for visual discrepancies between grades along the defined coordinate line. The expected outcome of the estimation process is to observe a relative smoothing of block model grades around the composite values.

Overall, the estimation of the Benso domains is robust and the results have been verified to a reasonable degree of confidence. Globally, the block model average grade is relatively similar to that of the declustered input data, indicating that no biases have been introduced.

The sectional validation slices show a reasonable correlation between the composite grades and the block model grades and it appears that a reasonable degree of smoothing has taken place for the majority of the domains.

14.7.4Chichiwelli

The block models were validated by comparing the block model mean grades with the declustered composite mean grades and through validation slices through the block models.

The mean grades for each of the estimated block models were compared to the declustered mean grade for the composite input data. Each of the modelled zones was compared separately. The differences between the declustered mean composite grades and the block grades are relatively small with the largest differences up to 10% for a few of the less well sampled domains, indicating that the model is similar to the input data on a global scale.

The block model was also compared to the composite grades within defined sectional criteria in a series of validation slices, the results of which are displayed on graphs to check for visual discrepancies between grades along the defined coordinate line. The expected outcome of the estimation process is to observe a relative smoothing of block model grades around the composite values.

Overall, the estimation of the Chichiwell domains is robust and the results have been verified to a reasonable degree of confidence. Globally, the block model average grade is relatively similar to that of the de-clustered input data, indicating that no biases have been introduced.

The sectional validation slices show a reasonable correlation between the composite grades and the block model grades and it appears that a reasonable degree of smoothing has taken place for the majority of the domains.

14.8Mineral Resource Classification

Block model quantities and grade estimates for the Wassa HBB Project were classified according to the CIM Definition Standards for Mineral Resources and Mineral Reserves (10 May 2014).

Mineral Resource classification is typically a subjective concept. Mineral Resource classification should consider the confidence in the geological continuity of the mineralized structures, the quality and quantity of exploration data supporting the estimates and the geostatistical confidence in the tonnage and grade estimates. Appropriate classification criteria should aim at integrating all concepts to delineate regular areas at similar Mineral Resource classification.

The geological modelling honors the current geological information and knowledge. The location of the samples and the assay data are sufficiently reliable to support Mineral Resource evaluation.

The sampling information was acquired primarily by diamond core and RC drilling on sections spaced at variable distances between the different deposit areas.

In situ dry bulk density has been estimated to a sufficient level to inform tonnages.

Using the above criteria a 3D surface and solid were created to separated areas of higher confidence (Indicated Mineral Resources) from those of less confidence (Inferred Mineral Resources).

 

 

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14.8.1Wassa

GSR classified the long-range Mineral Resource model and SRK used these to code an interim model constructed in January 2020, Figure 14-35. The GSR classification surfaces and solids did not change from the January 2020 Interim model and were also used to classify the final model created in February 2020. All blocks above the Indicated Mineral Resource surface and within the solid mesh were classified as Indicated Mineral Resources. No Mineral Resources were classified as Measured in the long-range Resource model.

Indicated Mineral Resources were classified where drilling was up to 50 m spacing, with Inferred Mineral Resources being classified where drill spacing was greater than 50 m.

SRK noted that blocks classified as Indicated blocks are informed by composites within an average distance less than 22 m from the estimated block, with more than 4 holes on average (Figure 14-36). Inferred Mineral Resources were supported by informing composites within an average distance of 56 m of the estimated block from an average of 2 holes.

Figure 14-35 Wassa LR model Indicated Mineral Resource classification surface and solids. All blocks above surface and within solid mesh were classified as Indicated Mineral Resources (GSR, 2021)

To provide quantitative support analyses for the classification scheme adopted by GSR, SRK extracted some statistics pertaining to the classified blocks based on an optimized pit generated by GSR. For this analysis, SRK used a cut-off grade of 0.4 g/t gold and 2.1 g/t gold for open pit and underground Mineral Resources, respectively.

Table 14-56 shows the breakdown of the Open Pit blocks above 0.4 g/t gold cut-off grade, based on Mineral Resource category, domains and also distance metrics to the nearest 3 holes. The Indicated blocks account for 99% of the contained metal within the pit, of which 38% comes from the HG and 61% comes from the LG domain. Overall, the open pit blocks are based on an average of 15 m distance to the closest 3 holes (or equivalently 25 metre drillhole spacing) and are mostly estimated in the first estimation pass.

Table 14-57 shows a similar breakdown for underground blocks above 2.1 g/t cut-off, based on Mineral Resource category, domains and also distance metrics to the nearest 3 holes. Unlike the open pit blocks, only 22% of underground blocks are classified as Indicated with the remaining 78% Inferred. Indicated blocks are largely supported by data from 4 or more holes estimated in the first pass, found within 21 m of the block and corresponding to drillhole spacing that is 25 m or less. Inferred blocks comprise 78% of metal content from underground blocks, of which 71% is from the HG domain and 7% from LG. Statistical analysis showed that approximately 60% of the metal within Inferred blocks are in areas of less than 100 m drillhole spacing, 30% in areas of 100-150 m spacing, 5% from 150- 170 m and 5% from greater than 170 m spacing.

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Figure 14-36 Estimation metrics associated to Indicated (top) and Inferred (bottom) classified Resources (SRK, 2020)

 

 

Table 14-56 Composition of Classified Blocks for Open Pit Extraction Above a Cut-Off Grade of 0.4 g/t Gold

Category3 Holes within
(m)
Avg.
Pass
Avg. No. 
Holes
Avg.
Data
Dist for
Estimate
(m)
Avg.  
Dist to 3
holes
(m)
Max.  
Dist to 3
holes
(m)
TonnageContained Metal
(oz)
% Metal  
  
Indicated 1.04.620.315.3113.627,629,0171,303,91899%  
HG 1.04.620.713.441.33,760,615504,88738%  
unlimited1.04.527.40.00.01,8012170%  
251.04.915.18.316.81,814,004226,67117%  
501.04.426.418.532.61,815,578258,56220%  
1401.13.431.626.541.3129,23119,4371%  
LG 1.04.720.215.6113.623,868,402799,03161%  
unlimited1.34.326.128.4113.6223120%  
251.04.814.19.617.010,825,800354,56227%  
501.04.623.418.733.911,577,012391,62530%  
1401.04.132.929.561.21,465,36652,8324%  
Inferred 1.23.737.737.266.1148,2159,5321%  
LG 1.23.737.737.266.1148,2159,5321%  
501.04.726.322.029.420,3391,9350%  
1401.23.539.739.966.1127,8767,5981%  
Total 1.04.620.315.4113.627,777,2311,313,451100%  

 

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Table 14-57 Composition of Classified Blocks for Underground Extraction Above a Cut-Off Grade of 2.1 g/t Gold

Category3 Holes within (m)Avg.  PassAvg.  No.  HolesAvg.  Data Dist for Estimate (m)Avg.  Dist to 3 holes (m)Max.  Dist to 3 holes (m)TonnageContained Metal (oz)% Metal  
  
Indicated 1.04.321.116.074.513,176,3851,873,29122%  

HG

Domain 4800

 1.04.320.815.251.811,116,4481,673,51620%  
unlimited1.03.933.60.00.08,7831,1190%  
251.04.614.08.617.26,579,6541,085,56913%  
501.04.326.619.734.02,484,389324,5104%  
1401.13.235.431.651.82,043,622262,3193%  

LG

Domain 8800

 1.04.223.019.974.52,059,937199,7742%  
unlimited2.01.017.273.074.51,161800%  
251.04.713.59.716.4764,42179,3491%  
501.04.524.019.534.3646,06961,0591%  
1401.03.333.232.555.8648,28659,2871%  
Inferred 1.92.054.166.2199.051,682,5946,678,10878%  
Domain 4800 1.92.054.265.2190.945,714,9506,083,75471%  
unlimited2.31.467.290.0190.915,774,7282,113,51425%  
251.13.919.310.416.170,9458,3920%  
501.03.428.622.034.01,059,985124,3261%  
1401.72.348.153.498.828,809,2933,837,52245%  
Domain 8800 1.81.953.373.1199.05,967,644594,3547%  
unlimited2.11.358.596.4199.03,172,757338,1504%  
251.04.317.911.214.812,2189180%  
501.04.127.621.832.499,4188,4400%  
1401.72.450.054.997.32,683,251246,8463%  
Total 1.62.644.852.0199.064,858,9798,551,399100%  

For the SR model, Mineral Resource classification was performed by wireframing the Measured and Indicated Mineral Resources, based on drill spacing displayed on section. For sections between 20,500 mN and 19,725 mN, sections were spaced every 12.5 m along northing. For sections 19,725 mN to 19,350 mN, sections were spaced every 15.0 m along northing.

For Measured Mineral Resources:

·between 20,500N and 19,725N, defined in areas where the drill intercepts were consistently no greater than 10m apart, up dip or down dip, along the mineralized structures, on each 12.5mN spaced section; and
·Between 19,725N and 19,350N, defined in areas where the drill intercepts were consistently no greater than 13m apart, up dip or down dip, along the mineralized structures, on each 15.0mN spaced section.

For Indicated Mineral Resources:

·Indicated Mineral Resources were classified for blocks in the model that were not classified as Measured Mineral Resources but were within a domain shell, trimmed against a boundary solid, used to define the limits of the use of the SR model in the final model. Outside of this boundary, the LR model is relied upon. The boundary solid typically extended outwards from the tightly defined Measured resource a maximum distance of approximately 100 to 120 meters vertically or horizontally.
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An example demonstrating the classification of resources on the 645m RL has been provided in Figure 14-37. The image shows the mineralization classified as either Measured or Indicated Mineral Resources in the SR model. The image also shows mineralization that is located outside the boundary solid. Mineral Resources and classification for this material would be informed by the LR model.

Figure 14-37 645m RL section showing resource classification, boundary solid and drill holes (GSR, 2020)

 

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14.8.2Hwini Butre – Father Brown – Adoikrom

Mineral Resource Classification for Hwini Butre generally follows the same general principles as those applied at Wassa. Classification has been assigned using a combination of drillhole spacing, geological and confidence in mineralization interpretation, as well as slope of regression values from the estimation process. The classification was modelled visually by digitizing a wireframe in order to define contiguous zones of confidence.

The surface used to define Mineral Resource classification was extended approximately half the drill hole spacing on section, as this is where confidence in the geological interpretation was considered to reduce. Indicated Mineral Resources have been defined in the areas of Father Brown and Adoikrom where drilling is sufficient to demonstrate geological and grade continuity to a reasonable level. The Inferred Mineral Resources have been constrained by two 3D solids that have included the wider spaced drilling at depth (100 to 200m spacing), shown in Figure 14-38. All other material outside of the 3D mesh/surface constraints remained unclassified.

Figure 14-38 Father Brown and Adoikrom Indicated Mineral Resource surface and Inferred Mineral Resource solids. All material above Magenta surface was classified as Indicated Mineral Resources all material below surface and within cyan (ADK) and red (FBZ) 3D meshes was classified as Inferred Mineral Resource

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14.8.3Benso

Classification for Benso generally follows the same principles applied at Wassa and Hwini Butre. Classification has been carried out using a combination of drillhole spacing, geological and wireframe confidence and was modelled by digitizing a wireframe.

The Indicated Mineral Resource wireframe was extended approximately half the drill hole spacing on section, as this is where confidence in the geological interpretation was considered to reduce. Indicated Mineral Resources have been defined in the Subriso East, Subriso West and G Zone areas of Benso where drilling is sufficient to demonstrate geological and grade continuity to a reasonable level (nom. 25 x 25 m).

14.8.4Chichiwelli

Classification for Chichiwelli generally follows the same general principles as those applied at Wassa, Hwini Butre and Benso, with classification carried out using a combination of drillhole spacing, geological and wireframe confidence, and was modelled visually by digitizing a wireframe.

Wireframes were digitized for East Domain and West Domain, with the areas inside the modelled solids considered to be Indicated Mineral Resources, and outside, Inferred Mineral Resources.

The majority of the Chichiwelli Mineral Resource has been classified as Indicated Mineral Resources. For the three additional deposits covered by the Chichiwelli MRE, an Inferred classification has been applied.

14.9Mineral Resource Statement

The Mineral Resources have been prepared in accordance with CIM Definition Standards for Mineral Resources and Mineral Reserves, adopted by the CIM Council on May 10, 2014, and the CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines, adopted by CIM Council on November 29, 2019.

Mineral Resources are reported inclusive of Mineral Reserves.

The Wassa Mineral Resource Estimates are a combination of the long-range (LR) and short-range (SR) models.

The “reasonable prospects for eventual economic extraction” (RPEEE) requirement implies that the quantity and grade estimates meet certain economic thresholds and that the Mineral Resources are reported at an appropriate COG, taking into account extraction scenarios and processing recoveries.

In order to determine the quantities of material offering “reasonable prospects for economic extraction” by open pit mining, GSR used a pit optimizer and reasonable mining assumptions to evaluate the proportions of the block model (Indicated and Inferred blocks) that could be “reasonably expected” to be mined from an open pit. The assumptions of open pit mining were only assumed for the Benso, Chichiwelli and HBB other prospects. No open pit Mineral Resource are reported herein for Wassa.

The optimization parameters are based on actual costs from the operations. The reader is cautioned that the results from the pit optimization are used solely for the purpose of testing the “reasonable prospects for economic extraction” by an open pit and do not represent an attempt to estimate Mineral Reserves.

GSR considers that the blocks located within the conceptual pit shells show “reasonable prospects for economic extraction” and can be reported as a Mineral Resource.

The underground Mineral Resources were reported above an economic cut off based on a $1500/ ounce gold price and mining, processing and general administrative costs that were adjusted from actual costs at the Wassa underground operation. Table 14-58 and Table 14-59 shows the Mineral Resource statements for the Wassa main and HBB deposits.

  

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Table 14-58 Wassa Measured and Indicated Mineral Resource, as at 31 December 2020

 Measured & Indicated Mineral Resource, at 31 December 2020

Meas. & Ind.

Mineral Resource

at 31 December 2019

 Measured ResourceIndicated ResourceMeas.  & Ind.   
Mineral Resource
 MtAu g/tkozMtAu g/tkozMtAu g/tkozMtAu g/tkoz
Wassa OP---------29.181.291,206
Wassa UG5.904.4584318.963.552,16224.853.763,00516.203.892,027
Father Brown /Adoikrom UG---1.317.963351.317.963350.918.67254
Benso OP---1.382.501111.382.50111   
Chichiwelli OP---1.111.75621.111.7562   
HBB Other OP---0.621.21240.621.21242.512.32187
TOTAL5.904.4584323.373.592,69429.263.763,53748.812.343,675

 

Table 14-59 Wassa Inferred Mineral Resource, as at 31 December 2020

 

Inferred Mineral Resource

at 31 December 2020

Inferred Mineral Resource

at 31 December 2019

 MtAu g/tkozMtAu g/tkoz
Wassa OP---0.621.3126
Wassa UG70.503.397,68958.823.757,097
Father Brown /Adoikrom UG2.665.304541.886.08367
Benso OP0.053.375---
Chichiwelli OP0.052.224---
HBB Other OP0.771.31320.422.1429
TOTAL74.023.448,18361.743.797,519

Notes to the Mineral Resource estimate:

·The Mineral Resource estimate complies with the requirements of National Instrument 43-101 and has been prepared and classified in accordance with the CIM Definition Standards for Mineral Resources and Mineral Reserves, adopted by the CIM Council on May 10, 2014, and the CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines, adopted by CIM Council on November 29, 2019;
·Measured and Indicated Mineral Resources are reported inclusive of Mineral Reserves;
·Underground deposits within the Mineral Resource are reported at a cut-off grade of 1.4 g/t;
·Open pit deposits within the Mineral Resource are reported at a cut-off grade of 0.55 g/t, within optimized pit shells calculated at a $1,500 /oz gold selling price;
·The Mineral Resource models have been depleted using appropriate topographic surveys;
·Mineral Resources are reported in-situ without modifying factors;
·No open pit resource has been reported for the Wassa deposit, as engineering studies have determined Wassa will be mined by underground methods only; and
·All figures are rounded to reflect the relative accuracy of the estimate.

 

 

 

 

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14.10  Mineral Resource Risks

During estimation of the Mineral Resources, the following risks were identified:

·At Wassa, the geometry of gold mineralization is complex and will require tight spaced drilling prior to extraction.
·The Inferred Mineral Resource in the southern portion of the Wassa deposit is informed by wide spaced drilling. It is the opinion of the QP that the global estimate in this area is within the accuracy limits to be classified as an Inferred Mineral Resource but the geometry of the mineralized zones will change with additional definition drilling.
·In the Southern portion of the Wassa LR model, additional definition drilling may impact the continuity of the grade domains and dampen the influence of higher-grade intervals. GSR have conducted subsequent studies comparing 2018 to 2020 long and short-range models. The comparisons demonstrate that historically, the addition of more drilling has resulted in larger Inferred and Indicated Resource estimates.
·Reporting of the Wassa underground resource at 1.4 g/t within the modelled 1.5 g/t isoshell may result in tonnages being underestimated and grades overstated. During 2021, the modelling parameters will be reviewed as to ensure the estimate is appropriate for the cut-off grade.
·The Inferred Mineral Resources at FBZ and ADK have been classified based on drill hole spacing in excess of 100m in some cases and there is risk associated with the grade estimates in these areas. The wider spaced drilling has however demonstrated the continuity of the mineralized structure and through further drilling the average grade of the inferred resource should be realized.

Beyond the risks disclosed here and in Section 25.2 not material risks have been identified.

 

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15Mineral Reserves
15.1Cut-off Grade

The cut-off grade applied for the Wassa UG Mineral Reserve is 1.9 g/t for stoping and development. This is a decrease from 2.4 g/t for the previous declaration, with the change driven by lower operating costs, achieved from increasing underground mining rates that have been sustained through 2019 and 2020.

Table 15-1 Wassa UG cut-off grade calculation

 

An assessment was completed during 2020 for cut-offs from 1.5-3.0 g/t. Stope shapes were generated for each cut-off, indicative schedules were developed by applying vertical advance benchmarks and costs were estimated using fixed and variable rates (lower $/t at higher rates).

Preliminary NPV’s (pre-tax) were calculated at $1,300 /oz and results showed peak NPV generated across the range of 1.6-2.0 g/t . 1.9 g/t was selected as the cut-off for calculation of the Reserve as the associated 5,000 t/d ore mining rate (1.8 Mtpa), is considered close to full capacity of the current mining system.

 

Figure 15-1 Wassa UG cut-off optimization

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15.2Modifying Factors

Modifying factors were applied to calculate the Wassa UG Mineral Reserve as follows:

·Stopes: based on back analysis of actual stope performance from 2020.
oMining Dilution: 5.0%
oMining Recovery: 96.1%
·Development: based on back analysis of actual stope performance from 2020.
oMining Dilution: 0.0%
oMining Recovery: 100.0%

Diluting material is assumed to contain no gold.

The effective modifying factors for the combined stope and development ore yield:

o100.8% of in-situ ore tonnes;
o95.8% of in-situ grade; and
o96.6% of in-situ contained ounces.

Modifying factors were determined from analysis of the stope performance for 2020 to end of November and are a more conservative approach than the previous declaration which assumed 0% dilution and 100% stope recovery. The change is mostly due to improving systems for monitoring and tracking stope excavation performance.

15.3Mineral Reserve Statement

The Mineral Reserves have been prepared in accordance with CIM Definition Standards for Mineral Resources and Mineral Reserves, adopted by the CIM Council on May 10, 2014, and the CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines, adopted by CIM Council on November 29, 2019.

Table 15-2 Wassa Mineral Reserve, as at 31 December 2020

 Mineral Reserve, at 31 December 2020

Mineral Reserve
at 31 December 2019

 Proven ReserveProbable ReserveMineral Reserve
 MtAu g/tkozMtAu g/tkozMtAu g/tkozMtAu g/tkoz
UG, Panels 1 & 24.283.284514.482.994308.753.138817.423.72889
UG, Panel 3---2.062.941952.062.94195---
Open Pit---------9.921.57500
Stockpiles0.690.5813---0.690.58131.060.6221
TOTAL4.972.914646.542.9762511.502.941,08918.412.381,410

Notes to the Mineral Reserve estimate:

·The Mineral Reserve estimate complies with the requirements of National Instrument 43-101 and has been prepared and classified in accordance with the CIM Definition Standards for Mineral Resources and Mineral Reserves, adopted by the CIM Council on May 10, 2014, and the CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines, adopted by CIM Council on November 29, 2019;
·The Mineral Reserve is reported at a cut-off grade of 1.9 g/t, calculated at a $1,300 /oz gold selling price;
·Modifying factors are applied as 5.0% dilution and 96.1% recovery for stopes;
·Material based on Measured Mineral Resources are reported as Proven Mineral Reserves;
·Material based on Indicated Mineral Resources are reported as Probable Mineral Reserves;

 

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·Material based on Inferred Mineral Resources are excluded from Mineral Reserve;
·Economic analysis of the Mineral Reserve demonstrates economic viability at $1,300 /oz gold price; and
·All figures are rounded to reflect the relative accuracy of the estimate.
15.4Mineral Reserve Risks

The Mineral Reserve estimate could be materially affected should assumptions not be realized for:

·Underground mining productivity and unit costs;
·Geotechnical conditions requiring a material change to the mine design;
·Processing performance (throughput and recovery) and unit costs; and
·Failure to maintain operating permits in good standing.

 

 

 

 

 

 

 

 

 

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16Mining Methods
16.1Mineral Resources Considered in Mining Plan

The Wassa Underground Mine (WUG) commenced development in 2015 and declared commercial production in January 2017.

The Wassa property has an established record of successful permitting applications from project commencement in 1998 to present. These are detailed in Section 20.1.2.

16.1.1Mineral Resource Inclusions

Mineral Resources considered in this assessment are as at December 2020 and consist of two geological models:

·Short-Range (SR) Model (bm201201_v4):

Estimate for mineralization proximal to current underground mine. SR model is applied north of 19,350 mN and below 745 mRL to the 350 mRL. It contains material classified as Measured and Indicated Resource.

·Long-Range (LR) Model (srkwasmar20e):

Estimate for mineralization in all areas not defined by the SR model and is applied from 19,240 to 19,350 mN and above 745 mRL. It contains material classified as Measured, Indicated and Inferred Mineral Resource and Inferred Mineral Resources are excluded from consideration.

 

Figure 16-1 Mineral Resources considered in Mineral Reserve and models applied

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16.1.2Definitions

Figure 16-2 illustrates the following definitions used to describe the different mining quantities.

·Panel: Each panel defines a progressive phase of definition drilling and capital development. The definition can be flexible but new panels are defined by their requirement for new access infrastructure (eg: Panel 3 vs Panel 1-2), a change in mining method (Panel 2 vs Panel 1) or a new phase of definition drilling followed by an investment decision (eg: Panels 4, to 5, to 6, etc).
·Area: Areas are semi/continuous zones of mineralization, which require extraction in a connected sequence but are geotechnically independent from other areas within the panel. Panels can extend across multiple areas where the panel boundaries are designed to permit sequence independence.
·Stope: A stope is a single production excavation which follows a defined sequence to complete the production cycle (eg: development, drilling, blasting, loading, filling).
·Lift: Stopes across multiple levels are mined in a series of lifts, as they progress through each level, i.e. a four-lift stope is four levels high.

 

Figure 16-2 Schematic of Wassa location descriptors

16.2Mining Locations

The underground mine has been divided into 3 zones which are shown in the plan in Figure 16-3 and longitudinally in Figure 16-4.

·Panels 1 & 2:

Current zones of mining, include B-Shoot, F-Shoot and Hanging-wall. Panel 1 is from 20,400 mN, south to 19,730 mN (+/-10m) and vertically from 745 mRL to 520 mRL. Panel 2 lies further south, from 19,700 mN (+/-10m) to 19 240 mN and vertically from 695 to 345 mRL. Mineral Resource in the mine plan for Panels 1 and 2 is classified as Measured and Indicated. Natural surface level is nominally 1,000 mRL meaning Panels 1 and 2 range from approximately 250 to 650 m depth.

·Panel 3:

The upper mine zones of B-Shoot, F-Shoot and 242, were included in the December 2019 Mineral Reserve Statement to be mined by open pit methods. This assessment now proposes a change to underground extraction (refer 16.3.1.1.1 for discussion).

Panel 3 runs from 20,200 mN, south to 19,700 mN (+/-10m) and vertically from 945 down to 745 mRL. Mineral Resources in the plan for Panel 3 are all Indicated. The zone also contains Inferred Mineral Resource that is not included in the mine plan. Panel 3 ranges from approximately 50 to 250 m depth.

 

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Figure 16-3 Wassa mine design and asbuilt, plan view (GSR, 2021)

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Figure 16-4 Wassa mine design and asbuilt, longitudinal view

16.3Current and Upper Mining Zones (Panels 1-3)

This section covers the extraction of ore north of 19,240 mN, being Panels 1, 2 and 3.

Wassa commenced underground development in 2015 and stoping production in 2017. During the establishment of the underground mine, open pit mining was occurring in parallel to deplete the Main and 242 pits. Open pit mining was completed in 2017.

Figure 16-5 Wassa underground production history

 

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16.3.1Mine Design

Panels 1, 2 and 3 will continue to be mined by underground methods using the Long Hole Open Stoping (LHOS) method, with 25 m level spacing.

16.3.1.1Stoping Methodology
16.3.1.1.1Upper Zones Method Change

The December 2019 Mineral Reserve included mineralization above the current underground workings and below the Main pit, as well as below the 242 pit, to be extracted by open pit methods. These upper zones are now collectively referred to as Panel 3.

During 2020, trade-off reviews were completed to determine the optimal extraction method.

The trade-offs showed that the optimal extraction method for the upper zones was by underground, rather than open pit. The advantages of underground extraction being:

·Improved selectivity: enables focus on extraction of high margin mineralization. This has resulted in a decrease in total ounces, but metal removed from the inventory is the higher unit cost, lower grade material.
·Reduced capital demand: the smaller scale underground plan will have a lower upfront capital demand than the large cutbacks required for open pit extraction. This further enables bringing forward production from the upper zones.
·Elimination of interactions: between active open pit and underground operations.

Table 16-1 Upper mine inventory change, OP to UG

 

16.3.1.1.2Panels 1 and 3
·Stope length is 25 m along strike with 6-10 m pillars between.
·Stope width is full width of the orebody which can be up to 35 m but is usually 15-25 m.
·Stopes are mined with uphole blastholes drilled from below. The stope lifts are extracted in a top-down sequence; each stope lift is extracted below the open stope void above. Up to four stope lifts are extracted to create a continuous excavation up to 100 m high.
·In Panels 1 and 3, the 100 m height limit usually enables full extraction of the orebody without the need for sill pillars. Where this is not the case, a sill pillar is left in the level between the sets of stopes.
·Mined voids are generally left open with some loose rock fill to dispose of waste or for opportunistic pillar recovery.
·Narrower ore zones (<15 m) are mined as longitudinal stopes with progressive placement of rock fill to minimize ore loss in pillars.
·Stopes at the south end of Panel 1 will utilize paste backfill within approximately 100 m of the panel boundary. Use of paste further north is limited by distribution pressures.
16.3.1.1.3Panel 2
·The introduction of paste backfill permits a change to increase the extraction ratio.
·Primary stope length is 20 m with 20 m pillars left between, which will then be mined as secondary stopes after filling and curing of the primaries.

 

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·Stope width is full width of the orebody which can approach 50 m but is usually 15-25 m.
·Primary stopes are mined with uphole blastholes drilled from below. The stope lifts are extracted in a top-down sequence; each stope lift is extracted below the open stope void above, shown in Figure 16-6. Up to four stope lifts are extracted to create a continuous excavation up to 100 m high, which are filled with paste fill as a single fill mass.
·Secondary stopes are mined with blastholes drilled from below with each stope lift extracted in a bottom-up sequence, shown in Figure 16-7. To minimize paste exposure in the side-walls, each stope lift is planned to be filled before extracting the lift above. As experience with paste fill increases, there is an opportunity that the secondary stopes could be extracted in multiple lifts before filling. This would deliver a more productive mine schedule but this plan assumes the more conservative approach.
·Combined stope excavations up to 100 m high are planned before a sill pillar is introduced. The first sill pillar in Panel 2 lies between the 520 and 545 mRL levels.
·Sill pillars can be extracted after the secondary stopes are backfilled and sill pillar extraction assumes 60% recovery of the full stope.

 

 Figure 16-7 Panel 2 primary/secondary stope extraction sequence, transverse stopes

The primary / secondary sequence in Panel 2 extracts the first pass of stopes to full design height (4-lifts, 100 m), mining every second stope along strike (the primary stopes), with pillars left between. The pillars are extracted as secondary stopes after sufficient primary stope voids complete paste backfilling.

In Panel 2, extraction of the first pass of primary stopes is well progressed. Paste backfilling and subsequent mining of the first secondary stopes is planned in 2021.

Figure 16-8 shows progression of the generic primary/secondary sequence, which includes:

·The first stope is generally located in the centre of the block and the mining front radiates from the centre toward the peripheries, with primary stopes mined to full height of the block (4-lifts, 100 m).
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·Secondary stopes follow the primary stope front with a lag distance of 120-140m along strike to create a sufficient buffer from active primary stopes and development.
·Secondary stopes will initially be constrained to one lift per stope, to limit exposure dimensions of the paste fill mass in stope walls.
·Extraction of the sill pillar commences when there is a sufficient distance from secondary stopes in the blocks above and below. Each crosscut into the sill pillar is scheduled to be redeveloped in time for the uphole stopes to be mined.
·This resulting sequence has primary stope extraction almost, if not fully, complete before the first secondary stope is mined. It creates a production profile which is high in the early years of primary stoping, then slows as secondary stopes are mined and becomes low when the block is only producing from stopes in the sill pillar.

 

Figure 16-8 Stope cycle for Panel 2 secondary stopes

16.3.1.2Stope Design

Mine design for Panels 1-3 was completed by the Wassa mine technical team in January 2020 as part the planning for the December 2020 update of the Mineral Reserve.

Optimal stope shapes were developed from the Mineral Resource block model using Datamine Mineable Shape Optimizer (MSO) software. MSO is a design algorithm which processes a geological block model against user defined geometrical parameters to produce optimized stope shapes. MSO inputs were:

·Cut off Grade: 1.9 g/t
·Stoping width (minimum/maximum): 5 – 100 m
·Minimum pillar between adjacent stopes: 10 m
·Minimum hanging/foot-wall dip angle: 80°

Optimization shapes were validated by manual checks to remove outliers and updated with production designs where applicable.

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16.3.1.3Development Design

The methodology used for the development design was:

·Operating development, Levels: both in ore and waste, designed to provide access for drilling and loading of stopes;
·Capital development, Decline: linking of level accesses with inclined development (1:7 grade), including stockpiles to facilitate development;
·Capital development, Levels: infrastructure on each level located at the required access position along strike, to enable the stope extraction sequence (access, stockpiles, electrical cuddy, paste fill access, footwall accesses); and
·Capital development – Infrastructure: attachment of infrastructure to the main decline and level accesses, including: ventilation network extensions, sumps and drainage system, and emergency egress (escapeway) network.

Panels 1 and 2 are accessed via the Daniel Owiredu Portal (formerly Portal 1), located in the Starter Pit. The Main decline is positioned east of B-Shoot which has variable dip. Maximum ramp grade is 1:7 and follows the plunge of the deposit south toward the deeper levels of Panel 2. Levels are accessed every 25 metres through level access drives connecting the ramp to each level’s footwall drive.

The upper zones will be mined as Panel 3 with two new decline accesses due to their spatial distance from the Main decline:

·Upper B-Shoot decline portal will be in the southern end of the Main pit and connects the Main decline at the 760 mRL. Duplicating the main decline enables the B-Shoot Upper material to be extracted in parallel with Panels 1 and 2, plus also forms part of a haulage loop system.
·242 decline will be mined in the footwall of the 242 shoot which is north of the mineralized zone. It is remote from any B-Shoot infrastructure and will be an independent ramp.

Figure 16-9 shows an isometric view of the asbuilt and planned underground development for Panels 1 to 3. Figure 16-10 shows a typical level layout for Panels 1/2.

 

Figure 16-9 Oblique view of Wassa Panels 1-3, asbuilt and planned development

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Figure 16-10 Typical level layout, Panels 1-2 570 mRL

 

Figure 16-11 Oblique view of Panels 3 242 Area, planned development and stopes

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Figure 16-12 Oblique view of Panels 3 B-Shoot Area, planned development and stopes

16.3.1.4Design Quantities

The design quantities defining the Mineral Reserve (as at 31 December 2020) and are summarized in Table 16-2.

Table 16-2 Wassa Panels 1-3, design quantities for Mineral Reserve

  

Panels 1-2

B-Shoot

Panel 3

B/F-Shoot

Panel 3

242

Ore Mined, Development‘000 t1,14316693
 g/t2.982.853.70
 ‘000 oz1101511
 share% ozXXX
Ore Mined, Stopes‘000 t7,6111,079724
 g/t3.152.713.25
 ‘000 oz7719376
Ore Mined, Total‘000 t8.7551,245818
 g/t3.132.713.30
 ‘000 oz88110987
Ore Mined, Total‘000 t 10,818 
 g/t 3.09 
 ‘000 oz 1,076 
Development, Totalm 44,173 
Dev’t Capitalm 20,392 
Dev’t Operatingm 23,781 
Vertical Developmentm 2,776 
Mined to Waste‘000 t 2,469 
Paste Backfill‘000 m3 2,967 

 

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16.3.2Geotechnical

Geotechnical characterization and design parameters used in the mine design for Panels 1-3 are based on:

·Wall mapping in permanent openings and ore drives to define the structural discontinuities;
·Geotechnical data available in from surface exploration and underground drilling logs;
·Logs of underground boreholes that had been subject to detailed geotechnical logging;
·A limited set of laboratory strength and deformation test results;
·Empirical support classification assessment to determine the support requirement for the permanent drives;
·Empirical Stability Graph (Mathews et al, 1981) assessments to determine the maximum stable spans of the stopes; and
·Numerical modelling to assess the stability and stress distributions around the stope spans and the crown pillar.

Geotechnical characterization has been done using Q classification values (Barton et al, 1974), for input to the Empirical Stability Graph Method, and Geological Strength Index (GSI, Marinos et al 2007) classification values.

16.3.2.1Structural Data

A number of faults have been identified orientated at right angles to the limbs of the fold. These are normal faults with downthrows of up to 5 m. They are characterized as fairly tight with little to no evidence of shearing adjacent to the contacts. Some were identified to contain in-filling material.

The rock mass structure has little variance between each lithology. The small scale discontinuities can be related to the major scale deformational processes that have affected the deposit. Several joint sets have been identified from the relevant data sources and have been considered for the mine design criteria.

Based on the structural assessment of the geotechnical mapping, the joint sets presented in the stereonet plot shown in Figure 16-13 and summarized in Table 16-3 were used for the stope stability assessment.

 

Figure 16-13 Stereonet plant of Wassa joint set database

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The dominant discontinuity sets in all domains, shown in Figure 16-13 are:

·The orebody parallel foliation (F);
·A moderately north-east dipping joint set (J1);
·A joint set trending south-east (J2);
·Joint set J3 is a moderately developed sub-horizontal set; and
·Joint set 4 (J4) is steeply dipping, north trending.

In this analysis, the mean discontinuity orientations presented in Table 16-3 have been used.

Table 16-3 Joint sets used for stope design

Discontinuity SetDipDip DirectionComments
Foliation16.5311.61Tightly healed foliation planes
J1--Set of tightly healed North-east trending joints
J21.641.33Set of tightly healed South-east trending joints
J31.661.37Set of sub-Horizontal North-west trending joints
J419.8314.31Set of steeply dipping North trending joints

 

16.3.2.2In-Situ Stress

Over-coring stress measurements were completed in September 2019 to measure in-situ stress levels in the mine.

The measurements were taken at two sites using the CSIRO HI-Cell method. The sites are located in the hanging-wall at 645-DD7 and in the footwall at the 570 decline.

The stress measurement was rated as Excellent for the 570 decline according to the stress test rating system used by the service supplier (industry standard). The 645-DD7 stress measurement was disregarded due to the test site being considered within the mining induced stress zone of excavations and not a valid reflection of the virgin in-situ stress field.

The results from the-570 decline are shown in Table 16-4 and the interpreted depth gradient is shown in Figure 16-14.

Table 16-4 570 decline stress measurement

Principal Stress

Magnitude

(Mpa)

Depth

(m)

Ratio

Gradient

(MPa/m)

DipDirection
Major26.54302.230.062339°
Intermediate18.94301.590.04469°
Minor11.94301.000.02883°219°

 

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Figure 16-14 Principal stress measurement Magnitude vs Depth

16.3.2.3Rock Quality

Based on the structural assessment of the geotechnical mapping, the rock mass quality is classified Very Good, using Barton’s (Barton et al, 1974) classification and Geological Strength Index (GSI) rating systems. Table 16-5 contains the rock mass condition data of the Wassa geotechnical domains, which were used for the stope stability assessment and ground support design.

Table 16-5 Wassa rock mass characterization parameters (Barton et al, 1974)

ParameterFootwall / Orebody / Hanging-wallSource
MINMAXAverage
Rock Quality Desc.RQD%859085Geotechnical and mapping
Joint NumberJn696Borehole structural data and mapping
Joint RoughnessJr343Detailed geotechnical logs & mapping
Joint AlterationJa10.751Detailed geotechnical logs & mapping
Q’ 435348 
Rock Mass Quality Very GoodVery GoodVery Good 
Geol.  Strength IndexGSI788078Underground mapping & inspections
Unconfined Compressive StrengthUCS Mpa110160135Rocklab laboratory test result
Unconfined Tensile StrengthUTS MPa161817Rocklab laboratory test result
Young’s ModulusGPa7080.575.3Rocklab laboratory test result
Poisson’s Ratio 0.280.320.3Rocklab laboratory test result
Densityt/m32.792.812.8Rocklab laboratory test result

 

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16.3.2.4Geotechnical Design, Development

Barton’s Q support classification system was applied to estimate the support requirement for the development headings, which are nominally designed at 5.5 m wide.

Using and Excavation Support Ratio (ESR) of 1.3, as defined by Barton and Grimstad (1993) for permanent mine openings, results are plotted in Figure 16-15. The development excavations are plotted in red and are assessed to be within Category 1 (No Support Required). This aligns with observations of a Very Good rock mass with little or no fallout and spalling.

Notwithstanding the results of the analysis, GSR applies a standard reinforcement pattern of friction bolts and surface mesh to the back and upper walls of all development headings. This level of support plots in Category 3 (Systematic Support) in Figure 16-15.

 

Figure 16-15 Support, Barton’s Q-Index chart (Barton and Grimstad, 1993)

16.3.2.5Geotechnical Design, Stopes
16.3.2.5.1Modified Stability Number

The Q’ value derived from the geotechnical characterization (Barton et al, 1974) has been used, along with the stability graph parameters A, B and C to determine the Modified Stability Number (N’) (Potvin, 1988) for stope back, side-walls (hanging/foot) and end-walls.

The stress parameter A was estimated by calculating the gravitational stress generated from the weight of the overburden rock above the mining. The structural parameters B and C were derived from an assessment of the interaction of the dominant joint sets with the stope boundaries.

Calculated N’ for the Q’ value derived from the rock mass characterization for both the longitudinal and transverse stopes are presented in Table 16-6 and Table 16-7.

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Table 16-6 Modified Stability Number (N’) for Panels 1-3, transverse stopes (Potvin, 1988)

ParameterStope Wall, TransverseComments
BackSideEnd
Q’ 47.947.947.9From mapping and core RQD, Jn, Jr, Ja
UCS,  Sigma CMpa130130130Average intact rock strength
Depthm500500500Average depth below surface, Panel 2
Max.  Principal Stress, Sigma 1Mpa13.513.513.5Estimated overburden stress
Stress : Strength Ratio1:9.69.69.6 
Factor A 1.01.01.0 
Angle between Stope Face & Daylighting Joint 15°15°45°Critical Joint for all back and side-walls is J3, end-wall is J4.
Factor B 0.20.20.5 
Potential Failure Mode GravitySlabbingSlabbingGravity or Slabbing
Dip of Stope Face 90°64° 
Factor C 285 
N = Q’ x A x B x C 19.276.7119.8N-value for all stopes >=64° slope

 

Table 16-7 Modified Stability Number (N’) for Panels 1-3, longitudinal stopes (Potvin, 1988)

ParameterStope Wall, LongitudinalComments
BackSideEnd
Q’ 47.947.947.9From mapping and core RQD, Jn, Jr, Ja
UCS,  Sigma CMpa130130130Average intact rock strength
Depthm500500500Average depth below surface, Panel 2
Max.  Principal Stress, Sigma 1Mpa14.614.614.6Estimated overburden stress
Stress : Strength Ratio1:8.18.18.1 
Factor A 0.90.90.9 
Angle between Stope Face & Daylighting Joint 15°45°15°Critical Joint for all back and side-walls is J3, end-wall is J4.
Factor B 0.20.20.5 
Potential Failure Mode GravitySlabbingSlabbingGravity or Slabbing
Dip of Stope Face 64°90° 
Factor C 24.98 
N = Q’ x A x B x C 17.3106.369.0N-value for all stopes >=64° slope

 

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16.3.2.5.2Stable Slope Design Geometry

Table 16-8 summarizes the calculated range of stable stope and design geometries for the expected rock mass conditions for Transverse and Longitudinal Open Stoping Methods in Panels 1-3. The orientation of the measurement axes is shown in Figure 16-16.

Table 16-8 Stable stope dimensions, Panels 1-3

Stope DimensionTransverse StopeLongitudinal Stope
MINMAXDesign (m)MINMAXDesign (m)
Heightm25100100<152525
Strike Lengthm252525<607070
Width across Strikem153025<151515
Dip, end/side-walls 65°65°65°65°65°65°

 

Figure 16-16 Stope axes measurements

The nominal designs for transverse and longitudinal stopes were plotted on the Matthews Stability graph and are shown in Figure 16-17 and Figure 16-18. All faces plot in the stable portion of the graph, without additional support.

This indicates that for a 30 m wide ore zone, stopes of 20 m width by 100 m height will be stable which is confirmed by field observations of current excavations.

 

 

 

 

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Figure 16-17 Matthews Stability Graph, transverse stopes (Mathews et al, 1981)

 

Figure 16-18 Matthews Stability Graph, longitudinal stopes (Mathews et al, 1981)

 

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16.3.2.6Geotechnical Design, Major Pillars
16.3.2.6.1Existing Pillars, B-Shoot

A crown pillar exists between the Main pit where B-Shoot has been mined below from underground. In addition, there are a number of sill pillars remaining between mined stopes underground, which are a mix of open void and loose rock fill.

Stability analysis was conducted using Phase 2 software which calculated factors of safety:

·Crown Pillar, B-Shoot Main pit and 720-N1 stope = 1.58
·Sill Pillar, 720-N1 and 745-S1 = 1.58

This indicates that in both situations a stable pillar can be maintained, Figure 16-19.


Figure 16-19 B-Shoot Pillars, modelled factors of safety from Phase 2 software, (GSR, 2018)

16.3.2.6.2Future Pillars, Panel 3

Panel 3 stopes are proposed to be excavated close to the 242 and Main B-shoot pits.

At the time of modelling the B-Shoot pillars, stope designs for Panel 3 were not complete, so no assessment was completed.

The indicative minimum design thickness of the crown pillars is approximately 20 m which is considered reasonable at this early stage but further geotechnical investigation will be required to confirm stability prior to mining.

 

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16.3.2.7Geotechnical Design, Ventilation Shafts

Ventilation shafts are planned for Panels 2 and 3 which will be excavated by raisebore with some possible use of drilling and blasting. The proposed locations of the ventilation shafts have not been geotechnically assessed and will require geotechnical data collection (drill holes) and evaluation to assess ground conditions to determine the unsupported diameter.

The first major ventilation shaft is planned to be raisebored in 2021 and includes drilling of a diamond drill hole for geotechnical logging prior as part of the work program.

16.3.3Hydrogeology

Hydrogeological investigations were undertaken in 2016 and 2019. Key conclusions for the Panel 1-3 were:

·Inflow of groundwater occurs along discreet zones of faulting and fracturing.
·Hydraulic testing of the underground mining regions to depths of 800 m below surface showed generally the formation is not water bearing and has generally very low permeability, although localized high permeability zones have been identified at depth with potential permeability of up to 2 l/s measured.
·Pit sumps in B-Shoot (Starter and Main) and 242 are possibly hydrogeologically connected to the underground workings. They have dual-use as surge sumps to manage surface runoff and staging points for the underground dewatering system. This functionality means that generally, sumps aren’t full but do hold sufficient water that there is potentially a recirculating groundwater load.

A main dewatering system was constructed in 2020 and planned for commissioning in 2021. Addition of the new pump station will allow the Starter pit sump to be dedicated for collection of incident rainfall and surface runoff, which will reduce sump inventory and likely reduce flows from hydraulic connection.

16.3.4Backfill

Different backfill systems and stoping methods are used across Panels 1-3.

16.3.4.1Rock Fill

In Panels 1 and 3, the nominal stope design mines stope voids, left unfilled, with ore pillars between. Uncemented rock fill is applied in irregular locations to increase recovery of mineralization by avoiding the creation of, or enabling the recovery of, ore pillars. Rock fill is sourced from development waste and tipped directly into stope voids by truck, or rehandled by loader from stockpile.

16.3.4.2Paste Fill

A feasibility study for the application of paste backfill at Wassa was completed by Outotec in 2018. Plant construction was completed at the end of 2020, Figure 16-20, and the full system is planned for commissioning early in 2021. Design capacity is 4,000 t/d of dry tailings processed to produce 120 cu.m/h of cemented paste fill. Depending on utilization, comparable plants support mining rates of 1.5-2.6 Mtpa.

16.3.4.2.1Test Work

The feasibility study program tested material characterization, rheology and strength and concluded Wassa tailings to be suitable for production of paste fill:

·Dewatering, including thickening and vacuum filtration, was achieved through proven unit processes typical of most backfill plants;
·Typical primary stopes sizes (20 mL x 20 mW x 25 mH) will require 4.5% cement to achieve the required strength of 270 kPa. Secondary stopes will require 3% cement to achieve the minimum threshold strength of 150kPa; and
·Underground distribution is amenable to gravity distribution (rather than pumping) with the location of the surface plant relative to the underground stopes.

 

 

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16.3.4.2.2  Filter and Mixing Plant

Paste fill will be produced in a filtration and mixing plant with the following processes:

·Tailings will be pumped as the full-stream discharge from the CIL plant in batches with automated changeover and flushing between the alternate discharge to the TSF. The pipeline corridor has secondary containment over the 3 km length.
·Tailings are processed in thickener and underflow thickened tailings are held in an agitated storage tank which which creates buffer capacity between the batches of tailings pumped from CIL and continuous filtration.
·Ceramic disc filters produced tailings cake which is delivered by conveyor to a paste mixer where binder (cement) and thickened tailings are added to achieve the required density and binder content.
·Surplus water (thickener overflow) is recycled to the process water network for re-use in the main processing plant.
·Mixed paste is transferred into a hopper which discharges to a borehole which supplies the underground distribution network.

The paste plant is operated from a dedicated control room with access to monitoring data from pressure sensors in the underground distribution network. In addition, the CIL plant control room can monitor operations and alarms in the paste plant, with both control rooms able to operate the tailing pumping processes from the CIL to paste plants. The paste plant is shown in Figure 16-20.

Figure 16-20 Wassa paste plant Dec-2020, thickener and storage tank in foreground

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16.3.4.2.3  Underground Distribution

Underground distribution is via a borehole to the 620 mRL, approximately 440 m long, shown in Figure 16-21. The borehole is duplicated for redundancy and to allow periodic cleaning to prevent blockage.

Paste will be distributed by 8 inch steel pipes and can be distributed by gravity to all stopes in Panel 2. Stopes in the south end of Panel 1 can also be reached by modifying the fill mix to include more moisture, which will require offsetting increased binder addition.

 

Figure 16-21 Paste fill distribution modelling

16.3.5Ventilation

Primary ventilation flows at Wassa are modelled using VentSim software with the model validated using results of volume and pressure surveys through the mine.

16.3.5.1Design Criteria

Ghanaian mining regulations prescribe:

·Maximum velocity of 6 m/s in travelling roadways;
·Minimum flow 0.06 m3/kW/s for diesel engine capacity;
·Minimum velocity of 0.2 m/s in headings, 0.1 m/s in large openings;
·32.5ºC wet bulb maximum working temperature; and
·Carbon monoxide must be continuously monitored in return airways and information transmitted to surface.

The ventilation system is designed to meet these regulations as a minimum, with airflow volumes in Panels 1-3 determined based on the following criteria:

·Up to 9 working areas at any time;
·50 m3/s, per working area.

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16.3.5.2Network Design, Panels 1 and 2

The current installed ventilation network, servicing Panels 1 and 2 has surface connections clustered north of the operating mining areas.

Primary intake points are:

·Daniel Owideru Portal (Portal 1) collared in the Starter Pit; and
·Portal 3, 670 mRL intake shaft and 695 mRL waste pass, collared in the Main Pit.

Primary exhaust points are:

·Portal 2, collared in the Starter Pit with 4 x 132 kW fans in parallel at the portal entrance; and
·695 mRL exhaust shaft, collared in the Main Pit with 2 x 280 kW fans in series on the 695 mRL level.

A ventilation review was completed by SRK(US) in 2020 which included updating the ventilation model for current and future operations. Surveyed and modelled airflows as measured in October 2020 are shown in Table 16-9. Model accuracy is considered adequate by SRK.

Table 16-9 Wassa ventilation model calibration, Dec-2020

 

To adequately ventilate the complete extraction of Panels 1 and 2, additional main airways and other changes to the current circuit are required to provide intake and exhaust capacity south of the current working areas:

·Construction of a 5.5 m diameter exhaust shaft from surface to 570 mRL, with installation of new primary fans;
·Construction of a 5.5 m diameter intake shaft from surface to 575 mRL; and
·Removal of the exhaust fans in Portal 2 and reversing airflow to become intake.

The new circuit will increase total airflow to approximately 590 m3/s, with 190 m3/s exhausting via the existing 695 mRL fans and 400 m3/s via the new southern exhaust shaft (RAR1). This will provide sufficient flow to operate 9 working areas with 50 m3/s per location. Figure 16-22 shows the Panel 1 and 2 ventilation circuit to end of life.

Construction of the two new shafts and installation of the new primary exhaust fans are budgeted to commence in 2021 and be completed early 2022.

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Figure 16-22 Wassa Panels 1 and 2 ventilation circuit to end of life

16.3.5.3Panel 3
16.3.5.3.1B-Shoot Upper

The Upper B ventilation circuit integrate with Panel 2 with connections for both exhaust and intake (via decline) at 745 mRL, shown in Figure 16-23. An optional exhaust shaft will be included in the mine plan to provide ventilation independence (approximately 100 m3 /s), enabling activity in B-Shoot to be ventilated without compromizing capacity to the main Panel 2 production area.

The ‘Extended Haulage Ramp’ shown in Figure 16-23 is a ramp to create a haulage loop in Panels 1 and 2.

 

Figure 16-23 Wassa Panel 3, B-Shoot Upper ventilation circuit

16.3.5.3.2242

The 242 mining area is spatially separated from the rest of the underground mine and has an independent ventilation network. Preliminary ventilation designs are preliminary only but reflect a conservative design approach to ventilate the area.

The current design is shown in Figure 16-24 with the following features:

·Exhaust via a connecting ramp to the Main pit. As well as exhaust ventilation, this drive will be used for definition drilling and as a second egress.
·Intake via the access decline.

 

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Figure 16-24 Wassa Panel 3, 242 ventilation circuit

The two exhaust fan positions in Panel 3 plan to re-use the four 132 kW fans currently located in Portal 2. Although detailed models are not yet completed, it is anticipated that, given the similarity to their current duty, with two fans in parallel each position provides approximately 100 m3/s to ventilate two working areas in each area.

16.3.6Mining Services
16.3.6.1Electrical

The underground electrical system has been designed and installed according to Ghanaian mining regulations and to efficient mining standards and will have high availability, medium utilization and low operating maintenance. The high voltage circuit above 570 level is 6.6 kV and 1 kV for low voltage. The high voltage circuit below 570 level in Panel 2 will be 11kV, with 1kV outlets for mining equipment. The total underground feed from both the 6.6kV and 11kV circuits is approximately 13.0 MVA., split 5.0MVA to the 6.6kV circuit via the Starter Pit and 8.0MVA to the 11kV circuit via a borehole to 570 level.

Panel 3’s electrical circuit will use the 6.6kV circuit because of Upper B and 242’s proximity to the already installed infrastructure. Panel 2 below 570 level will use the 11kV circuit. Refer to Figure 18-4 for the site’s basic electrical layout line diagram.

16.3.6.2Compressed Air

The compressed air system comprises 2 x 90 kW compressors located on surface at the Starter Pit portal. Compressed air is distributed underground via a 110 mm poly pipe down the main decline. Due to pressure drop along the reticulation and incremental increases in duty, an additional compressor is planned.

16.3.6.3Service Water

A 30,000 litre water tank is installed above the portal area to supply the underground mine with service water for drilling, dust suppression and general use. The service water tank is filled using the 90 kW Flygt pump that is permanently installed in the Starter pit sump. Service water is reticulated throughout the mine by 110 mm HDPE lines installed in the primary headings and reducing to 63 mm HDPE for supply to end use locations.

16.3.6.4Underground Dewatering

The underground mine dewatering system is designed and installed to remove both ground water and service water (collectively called mine water), including up to 10% by volume solid particles.

The mine dewatering system contains the following staging in the upper part of the mine:

  • 18kW decline face pumps which pump to sumps on operating levels;
  • 37kW pumps transfer initially settled water to either mono pumps or 90 kW Flygt pumps, which pump to the starter pit sump outside the main portal.

 

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The system dewaters to the Starter Pit sump at a rate of 35 l/s and, where required, up to 65 l/s. The F Shoot mining area will continue to use this system; the rest of the B Shoot mining areas will use a recently installed permanent pump station on 620 mRL.

The permanent pump station at the 620 mRL level can pump 80 l/s over a 440m total dynamic head. The pump station pumps directly to the surface via a borehole to surface settling and discharge routes. The station uses cascading settling sumps to drop out as many solids as possible prior to pumping to surface. The station uses multistage pumps to meet the total dynamic head and flow rate required. A borehole connects the pump station to surface with a 200 mm NB steel rising main installed.

As the mine progresses at depth beyond the 620 mRL level, additional staging pumps will be utilized and directed to the 620 mRL pump station. These pumps will be similar to existing pumps with sublevel 37 kW Flygt pump and for main dewatering at depth a 90 kW Flygt pump and, if required, supported by a 55 kW Mono pump. A reduced dewatering system above the 620 mRL pump station will remain in place to intersect inflow at higher levels and dewater to the Starter Pit sump on surface. The dewatering system is shown in Figure 16-25.

The recently installed 620 mRL pump station is shown in Figure 16-26.

 

Figure 16-25 Underground dewatering longitudinal view

 

 

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Figure 16-26 620 mRL main pump station

16.3.7Mining Schedule

Mining quantities for Panels 1-3 were scheduled using MineSched software, with spatial links between development and stoping, and capacity constraints which reflect the methodology and sequence outlined above.

The mine schedule calculates all physical quantities which are input into the cost estimate.

·Definition Drilling: geological diamond drilling required to define the mineralization. Drilling is capitalized where the material being targeted is not yet classified as Mineral Reserve or production is planned two or more years after drilling.
·Development: Lateral and ramp development to access and support stoping. Heading types which support production from a number of stopes are classified as capital (decline, stockpiles, level access, footwall drives, vent access, orepass access, dewatering) and access for production from one stope are operating (stope crosscut, ore drives).
·Vertical Development: Vertical development (long hole raise or raisebore) for ventilation, egress, orepass or other infrastructure, which are all capitalized. Raising required for stope blasting is not quantified and is included in the $/t unit cost for stope blasting.
·Backfill: Paste fill volume placed.
·Waste Material: waste generated from lateral and vertical development activities, including any material which may be placed as rock fill.
·ROM Material: Material generated from development and stoping and sent for processing.
·Haulage: Estimated haulage quantities, calculated from planned tonnes mined and average one-way lead distance.

 

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The key scheduling constraints applied were:

·Maximum advance rate from a single heading: 4.5 m/d for decline, 6.0 m/d for other lateral;
·Maximum total advance rate: 27 m/d (820 m/mth);
·Maximum ore tonnage from a single stope: 3,000 t/d; and
·Maximum total tonnage: 6,200 t/d in 2021 and 6,800 t/d thereafter when Panel 3 development commences.

Milestones assumed in the scheduling of Panels 1-3 are:

·Continued extraction from Panels 1 and 2 from 2021;
·Initial development from Panel 3 in mid-2022;
·Stoping commences from Panel 3 in 2025;
·Sill pillar extraction in Panel 2 commences 2024;
·Development completed 2025;
·Primary and secondary stopes completed in 2025; and
·Sill pillar extraction (and all other activity) completed in 2026.

The schedule quantities are shown in Table 16-10.

Table 16-10 Wassa mining schedule quantities for Mineral Reserve plan

 

 

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Figure 16-27 Lateral development schedule for Mineral Reserve

Figure 16-28 Ore mining schedule for Mineral Reserve

 

 

 

 

 

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Figure 16-29 Underground Production History and Mineral Reserve plan

 

16.3.8Mobile Equipment

The current mining fleet at Wassa is a mixture of the original low-cost, pre-owned fleet from when the mine was established and new units which have been commissioned as the project has modernized and productivity has grown. The forward plan at Wassa assumes that this cycling out of older units with new equipment will continue:

·Development Jumbos: Current fleet is four Sandvik DD421 twin-boom jumbos which continue as the standardized development machine.
·Production Drills: Current fleet is three rigs, one each of Sandvik DL411/421/431 which are the same class machine (89-115 mm top-hammer) with different boom configurations. This machine will continue as the standardized blasthole drilling machine, possibly with different boom configuration or replacement of a longhole drill with a small raisebore/boxhole rig for stope slots.
·UG Loaders: Current fleet is four 18 t class LHD’s (two older Cat R2900G and one Sandvik LH517) and two 21 t class machines (Sandvik LH621), which will be standardized to the LH621. Fleet numbers in early years have been adjusted to reflect operation of the smaller units until they are cycled out of the fleet.
·UG Truck: Current fleet is eight 40 t class articulated trucks (Volvo A45G) which are planned to upgrade to 60 t class machines (Volvo A60H), with the first larger truck budgeted for 2021. Fleet numbers in early years are adjusted to reflect the smaller units until they are cycled out of the fleet.

Machine numbers for the mobile equipment fleet categories were estimated using the productivity assumptions shown in Table 16-11 and the resulting fleet schedule is shown in Table 16-12.

 

 

 

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Table 16-11 Mobile fleet productivity assumption

 

 

Table 16-12 Mobile fleet schedule for Mineral Reserve plan

 

 

 

 

 

 

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17Recovery Methods
17.1Processing History

Wassa started industrial scale processing in 1998 utilizing a heap leach (HL) to recover gold from the ore. The process involved crushing, screening and agglomeration of the mined feed material before being stacked on leach pads which were irrigated with a weak cyanide solution to recover the gold. The solution was processed through carbon columns, stripped from the loaded carbon and smelted through to gold doré bars. Actual recoveries of 55-60% did not achieve planned recovery of 85% which led to the suspension of operations in 2001.

In 2003 a feasibility study commenced to evaluate construction of the current CIL plant. The study results were positive and the plant was constructed in 2004 and commissioned in 2005. The CIL plant uses crushing, milling and CIL and was designed to process 3.5 Mtpa from a feed blend comprising 45% fresh material, 25% oxidized material and 30% reclaimed spent HL material. Spent HL material reclaimed from the pads was added to the mill feed via a scrubber until this material was depleted in 2014. After that, mill feed consisted of fresh material from the open pit until 2016 when underground material was introduced to the feed. Open pit mining was completed in 2018 and since then, the predominant feed has been underground ore with supplementary addition of open pit stockpiles fresh, low grade ore.

Table 17-1 Historic plant production, grades and recoveries

 

17.2Flow Sheet Description

Gold recovery is achieved using conventional CIL technology, although the plant itself contains a few atypical features due to its history and development.

The plant flowsheet has transitioned from the historical HL processing and currently consists of the following operations:

·A four-stage fine crushing circuit incorporating an open circuit primary jaw crusher followed by secondary, tertiary and quaternary cone crushers with the secondary and tertiary crushers operated in closed circuit with sizing screens. A single secondary, two tertiary and four quaternary crushers give a nominal crushed product size from the crushing circuit of 80% <8 mm.
·Two independent milling circuits, each comprising a 5.03 m diameter x 6.7 m long ball mill fitted with 3 MW motors feeding individual clusters of classifying cyclones. Reported mill product size is around 80% <75 µm.
·Two separate gravity gold recovery circuits using 48” Knelson centrifugal concentrators process a portion of the classifying cyclone feed in each mill circuit.

 

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·The gravity concentrate from the Knelson concentrators is leached using an intensive leach reactor in combination with an electrowinning cell to recover the precious metals as a sludge prior to refining. The tails from the centrifugal concentrators are returned to the milling circuits.
·Classifying cyclones and pre-leach thickener. The thickener underflow feeds a transfer vessel together with the secondary cyclone underflow where cyanide is added before the slurry is transferred to the CIL circuit. Oxygen is injected into the transfer line after the transfer pumps. The transfer pipeline acts as an In-Line Reactor (ILR), where most leaching occurs.
·Adsorption occurs in the counter current CIL circuit, consisting of six stages of agitated vessel each of 2500 m3, providing an overall residence time of 18-20 hours at a 7,400 t/d mill capacity. Hydrogen peroxide is added periodically to CIL tank 1 to maintain the dissolved oxygen level. Activated carbon is retained in each tank using interstage basket screens and is moved counter-current to the slurry flow using submerged vertical spindle pumps in each tank. Loaded carbon is recovered from the first CIL stage.
·Loaded carbon is acid washed and then stripped of gold using caustic soda in an 11.5 t pressure Zadra elution system with the gold electrowon onto steel mesh before smelting.
·Eluted carbon is thermally regenerated and returned to the last stage of the CIL circuit.
·The gravity gold concentrate and electrowon gold are smelted separately to produce doré bars.
·Additional supporting facilities include:
oTwo, 2.1 t/d capacity pressure swing absorption oxygen plant located in the milling area;
oemergency diesel powered generators.

The key plant design and operating parameters are shown in Table 17-2 and a schematic flowsheet for the Wassa plant is presented in Table 17-2. The schematic incorporates the new densifying cyclone and thickening circuit currently being installed.

The Wassa process plant is currently operating below design capacity due to limited feed supply. 2.0 Mt of ore was processed in 2020 compared to nameplate capacity of 2.7 Mtpa.

The Wassa process operation achieved certification with the International Cyanide Management Code in early November 2009 and was recertified in 2017 and again in 2020.

Table 17-2 Key plant design and operating parameters

ParameterUnitFresh Ore Feed
DesignCurrent Operations
Nominal throughputMtpa2.651.60
Crushing Circuit Product% passing80%< 8mm80%< 8mm
Crushing Circuit Utilization%7575
Plant Design Availability%9292
Mill product grind% passing70%<75 micron70%<75 micron
CIL Feed Density, Design/Current% Solids4040 (CIL tanks - measured)
CIL Feed Density, with thickener% Solids44-4644-46
CIL Retention Time (calculated)h (total)2033

 

 

 

 

 

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Figure 17-1 Wassa processing plant flow sheet

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17.2.1Plant Accounting

Plant throughput is reported based on the belt weighers installed on the conveyors feeding the two ball mills from the crushed material stockpile. There is also a belt weigher installed on the crushing circuit product to the crushed material stockpile.

Plant performance and accounting is assessed based on samples of feed and tailings taken automatically using inline slurry samplers, which are composited into 12 h shift accounting samples. The feed sample is taken after the milling and gravity circuit before transfer to the CIL circuit and the gold recovered by gravity and smelted separately is added to calculate the plant feed grade. The feed and tail slurry samples are analysed using bottle roll laboratory tests to assess the BLEG tests.

Slurry samples are filtered and washed and solids are pulverized to 95% <75 µm before being subject to BLEG (bulk leach extractable gold) bottle roll extraction. BLEG tests are run for 8 hours at high cyanide concentration. Solutions from BLEG test and slurry filtrate are analysed by gold extraction into an organic phase and then measured by atomic adsorption spectroscopy (AAS). Extended BLEG tests are also done to confirm that all the recoverable gold has been extracted during the standard BLEG leach period. The BLEG tails are periodically fire assayed to determine residual gold in the samples not recovered in the BLEG tests (gold potentially locked in silica, pyrite or other sulphide minerals). Initially, a BLEG factor was used in assessment of the total gold in the plant tails to determine the overall plant gold recoveries. However, GSR has continuously improved its BLEG testing process so that the tests achieve complete gold dissolution. Periodic fire assays on the residue of the tails from the BLEG tests confirm the efficiency of the BLEG tests.

The gold recovered by gravity is leached, electrowon and smelted separately and this is added to the gold in the mill product sample to determine the gold grade in the feed. A sample is taken of crushed ore from the feed to the ball mills and this is used as a check measurement on the plant feed grade although is not used for accounting purposes.

Reconciliation is undertaken monthly between the gold produced and the gold present in the feed and tails. This also considers the changing gold inventory on the plant from month start to month end. Based on the reconciliation the reported head grade is adjusted to correlate with the monthly gold production.

17.3Processing Schedule

Table 17-3 shows the processing schedule quantities for the Mineral Reserve plan.

Table 17-3 Processing schedule quantities for Mineral Reserve plan

 

 

 

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Figure 17-2 Processing schedule for Mineral Reserve plan

Figure 17-3 Gold Production schedule for Mineral Reserve plan

 

 

 

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Figure 17-4 Processing Production History and Mineral Reserve Plan

 

Figure 17-5 Gold Production History and Mineral Reserve plan

 

 

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18Infrastructure

The locations of mining areas and major infrastructure at Wassa are shown in Figure 18-2, including:

·Main roads, towns and power lines;
·Open pit voids and waste storage areas;
·Processing facilities;
·Tailings storage facilities; and
·Site accommodation.

Key infrastructure locations around the main site area are shown in Figure 18-1 and a site layout in Figure 18-2, which shows both the local mine grid and UTM grid (WGS84 30N). Figure 18-3 shows the same site layout image plus the underground workings surveyed at the end of December 2020.

 

Figure 18-1 Wassa key infrastructure (GSR, 2018)

 

 

 

 

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Figure 18-2 Wassa site layout (GSR, 2021)

 

 

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Figure 18-3 Wassa site layout and underground workings (GSR, 2021)

18.1Electrical Infrastructure
18.1.1Power Supply

Wassa has two power supply sources. The site is connected to the national grid, along with on-site power generation.

Grid power from the national power supplier (VRA) via a network operated by GridCo comes from a 161 kV line to local substation where power is transformed down through a 33 MVA transformer to 34.5 kV. The grid connection has been the primary site power supply since commissioning in 2006.

An on-site power station was constructed during 2020 to improve long-term reliability of the power supply. The plant is owned and operated by Genser under an agreement and contains two 34.5kV, 16.5MW gas turbines. The plant was commissioned in early 2021 and now supplies all site power except for the site accommodation camp.

With on-site generation, the grid connection is retained, permitting use of the grid for standby supply.

 

 

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18.1.2Site Distribution

Power is supplied to the main site substations at 34.5 kV with a major upgrade completed in 2020.

·The processing plant and other surface facilities are fed via the GSR substations, which are 16 MVA and 18 MVA capacity and distribute at 6.6 kV. They are sized for the plant to operate at the full 2.7 Mtpa throughput rate.

 

Figure 18-4 Site electrical distribution

·The underground mine is supplied by three 34.5/6.6 kV substations:
oOne 5.0 MVA capacity transformer with two 2000 kVA 400 V diesel generators with switching and transformers to distribute at 6.6 kV to distribution substations in the underground mine and associated locations, where it is locally stepped down as required to 1000 V, 415 V and 240 V. Spare switches are available for future requirements.
oTwo 4.0 MVA transformers were added in 2020 when distribution was expanded to provide capacity for the paste plant and ventilation fans required to extract the Reserve.

Distribution voltage was increased from 6.6kV to 11kV and the project included installing a new switch yard to split the feed between the 34.5kV/6.6kV circuit servicing mine via the Portal bench and the new 34.5kV/11kV circuit which connects to a ring main unit at 570 level via a single point suspended 185mm2 XLPE cable in a raisebored service hole.

The 11kV project has capacity to be readily expanded and Figure 18-4 shows a simplified line diagram, with the conceptual expanded HV circuit shown in yellow shading.

 

 

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18.2Surface Water Management

Water diversion structures are installed as required to prevent inflow of surface runoff from the surrounding topography. Within the pit crests, water inflow is a combination of rainfall and groundwater.

Each of the four catchments within the Main pit complex (Figure 18-5) have a storm water collection sump which is designed for a 1:100 year, 24 hour duration event (241 mm). Catchment modifications have been completed to manage flow directions and capacity requirements at each sump. The dewatering discharge from the pit sumps is used across the site (eg: process plant, dust suppression) and excess water is directed to settling and drainage systems prior to release.

 

Figure 18-5 Wassa Main pit catchments

·Starter Pit: has direct connectivity to underground workings via the two portals at 905 mRL. The sump below the portal entrances has 115% of the capacity required for the design rainfall event.

Service and emergency pump systems are installed to maintain low operating levels and provide surge capacity to draw down the sump during and after a rainfall event.

oService – 45 l/s (electric submersible pump with 160 mm HDPE pipe); and
oEmergency – 165 l/s (diesel pump set with 2x 160 mm HDPE pipes).
·B-Shoot Pit: has direct connectivity to underground workings via the portal, waste rock pass and vent shaft breakthoughs at the 844 mRL. The sump in the bottom of the B-Shoot pit has 200% of the capacity required for the design event.

A single diesel pump set with 2x 160 mm HDPE pipes is installed with 150 l/s capacity.

·242 Pit: currently has limited connectivity to underground workings through groundwater seepage only. However, this will change with underground development of the Panel 3, 242 area, although the surface water management plan will remain consistent with the current strategy to catch water in the 242 pit sump to prevent entry into underground workings.

 

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Current sump capacity is 360% of the design requirement and will reduce to 138% when the new portal at 940 mRL is cut to establish the underground workings.

A single diesel pump set with 2x 160 mm HDPE pipes is installed with 150 l/s capacity.

·South East Pit: currently has limited connectivity to underground workings through groundwater seepage only. Sump capacity is 240% of the design requirement.

A single diesel pump set with 2x 160 mm HDPE pipes is installed with 150 l/s capacity.

·The pump network is maintained regularly and sufficient spare equipment is on site including one of the electric submersible pumps and two of the diesel pump sets.
18.3Workshops and Other Site Buildings

The following engineering workshops are in place to support site operations:

·Processing Fixed Plant: located near the processing plant to support its activity.
·Surface Mobile Equipment: located between the administration area and equipped with offices, overhead cranage, services and welding bay to support the former open pit mining fleet.
·Underground has three workshop areas to support the underground mining fleet:
oSurface, located near the underground offices with offices, services and 1000 V power supply for equipment testing;
oStarter pit portal bench: lube and service bay, with 1000 V test panel; and
oUnderground workshop at 595 mRL: this facility is in the final stages of construction (excavation, support, concreting and water management are complete) and is planned for completion in 2021-Q2. It will be used to service drilling equipment and loaders and reduce tramming time to surface.
·Light Vehicles: located near the warehouse.

Other buildings on site include:

·Administration offices;
·Kitchen and messing facilities;
·Diesel fuel storage;
·Warehouse and dry goods storage;
·Metallurgical laboratory; and
·Core processing and logging facility.
18.4Site Accommodation

Employees reside both on-site, or in surrounding towns and villages.

On-site accommodation is located at the Tara Camp 3 km northwest of the mine site as well as at Camp 2 located within the Akyempim village.

Facilities include:

·Accommodation for both single employees with some houses for families;
·Company medical and health clinic including primary care, laboratory, pharmacy, radiology ambulance and detention services;
·Kitchen and messing facilities;
·Recreation facilities including gymnasium, tennis court, swimming pool and bar; and
·Commissary.

 

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Accommodation is currently being expanded to provide additional capacity to permit more personnel to reside on site during their roster-cycle and reduce risk of pandemic exposure. The accommodation expansion project is shown top-left of Figure 18-6.

Figure 18-6 Tara Camp

18.5Waste Rock Storage

The waste dumps are located adjacent to the Main and South Akyempim pit complexes. Waste from underground operations is either placed in underground stope voids or hauled to the waste dump locations shown in Figure 18-7.

Waste dumps were designed and then permitted in 2017, to allow an additional 88 Mt of storage of waste from the underground and Main pits Cut 3 cutback. The 419 waste rock dump, south of the Main pit, is the currently active waste placement area. The permitted storage at the site is sufficient for the waste volume scheduled to be mined alongside the Mineral Reserve.

 

 

 

 

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Figure 18-7 Waste dump locations (Golder, 2016)

Dumps are designed with 10 m bench height with 10 m wide berms and consider operational and rehabilitation phases. For operations, as-dumped designs have 37° batter angles (natural angle of repose) with wider berms and final rehabilitation applies 25° batters to achieve the overall slope of 22°. A nominal dump section is shown in Figure 18-8.

Asbuilt and designed dumps include the following features:

·Adequate drainage to ensure that any discharge from the waste dump is contained for settlement and/or monitoring, to enable compliance with the EPA effluent discharge limits.
·The top surface of the dump, and any berms partway up the dump slopes, are constructed to shed water away from the surface of the dump.
·Water collecting drains are constructed around the perimeter of the dump to route discharges and runoffs into settlement and monitoring ponds.

 

Figure 18-8 Section through nominal waste dump design

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18.6Tailings Storage

There are two tailings storage facilities at Wassa which are described below and shown in Figure 18-9:

·TSF 1:

TSF 1 is located northwest of the processing plant at the head of a southerly draining valley and immediately adjacent to the historical leach pad area. Ground levels range from 1000 mRL on the valley floor to above 1060 mRL on the surrounding hills.

It is a cross valley impoundment created by the construction of a main embankment in the south with confining saddle embankments at the north of the facility. Containment to the east and west is provided by natural ridges. Access is via unsealed access road west of the plant site area.

The catchment area of TSF 1 is estimated to be approximately 140 Ha, of which 124 ha is covered with tailings as the facility proceeds through closure revegetation trials.

Deposition into TSF 1 ceased in 2019 with paddock deposition completed to achieve the approximate closure surface topography requirements of the closure landform.

Re-vegetation trials commenced in 2017 towards the next land use and by the end of 2020 revegetation planting was mostly complete.

·TSF 2:

TSF 2 is located in the valley system that trends eastward from the north embankment of TSF 1. It is approximately 2.5 km from the processing plant and 1.3 km downstream of TSF 1 Saddle Dam 5.

TSF 2 has a footprint of 260 ha, of which 72 ha have been developed to date, and lies within a total project area of 340 ha including buffer zones.

The remaining capacity of TSF 2 is well in excess of that required for processing of ore defined by the Mineral Reserve, both before and after allowing for use of tails solids in paste backfill.

 

Figure 18-9 Wassa TSF 1 and TSF 2 aerial view (August 2020)

Figure 18-10 is a photograph taken from the north of TSF 1, looking southeast in November 2020. In the image the following features can be seen:

·Revegetated TSF 1 to the right;
·Active deposition into TSF 2, Cell 1 in background left; and
·Basin preparations of TSF 2, Cell 2 in the foreground.

 

 

 

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Figure 18-10 View from north of TSF 1 looking southeast (November 2020)

18.6.1History

TSF 1 was commissioned August 2004 to meet the tailings storage needs for the mine life associated with construction of the CIL plant. Since starter embankment construction, embankments raises have been designed, permitted and constructed up to the final elevation at 1039 mRL.

During planning for TSF 2 in 2009 and 2010, alternatives were considered for additional tailings capacity, with four potentially feasible options:

·Two different locations for new TSF’s;
·Increasing elevation of TSF 1 to 1049.5 mRL; and
·Increasing elevation TSF 1 whilst progressing a new TSF.

The selected option was to construct a new facility (TSF 2). The new TSF required development of a Resettlement Action Plan (RAP) which ultimately determined the RAP scope to include resettlement of the entire community of Togbekrom.

In compliance with the requirements of the EPA’s Environmental Assessments Regulations, 1999 (L.I. 1652), GSWL registered a new TSF project with the EPA in May 2010 and obtained authorization to proceed to permitting in July 2010. An Environmental Scoping Report was submitted to the EPA in March 2011 and later, and EIS was submitted for the construction and operation of the proposed TSF 2. The EIS was approved by the EPA in April 2013 (EPA/EIA/383) and conditions of the EIA permit led to GSWL re-designing the TSF 2 facility to accommodate a geomembrane liner.

While conducting the impact assessments and the preparation of the EIS, GSWL sought permission to raise the TSF 1 by an additional 5 m and for continued deposition between August 2011 and May 2015. All embankments have subsequently been constructed to the final permitted elevation of 1039 mRL.

 

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In March 2015, GSWL obtained permitting to expand TSF 1 into the disused heap leach area that was located directly east of TSF 1. The 16.2 ha extension provided 2.09 Mt additional storage capacity conventional deposition (embankment spigotting) and in over 2.17 Mt of capacity, primarily through paddock deposition (spigotting from day walls) across the entire TSF 1, to achieve the optimal drainage design, ahead of TSF 1 closure.

GSWL applied to the EPA in July 2014 for the renewal of the TSF 2 permit in compliance with the requirements of the EPA Environmental Assessment Regulations, 1999 (L.I. 1652) and Section 3.7 of the EPA Permit (EPA/EIA/383) after the facility was re-designed to accommodate a geomembrane liner. The EIS for TSF 2 was updated in January 2014, following advice from the EPA permitting was issued in January 2016 with an effective date of November 2015 (EPA/EIA/442).

The development of TSF 2 necessitated resettlement of some 105 households within the Togbekrom and surrounding hamlets to New Ateiku, approximately 10 km north. All the affected people affected by the project were successfully relocated to their new homes in Q1 of 2013. The RAP has been successfully completed.

TSF 2 has current design capacity of some 41 Mt of tailings, which provides approximately 15 years capacity at 2.7 Mtpa throughput. It will be constructed in three cells and 11 stages. The cellular design provides flexibility to modify stage raises and enable dry season construction for various throughput rates.

At the time of permit renewal, the TSF 2 design had been revised to a cellular arrangement with lining of the entire basin with HDPE geomembrane. In February 2016, the Mines Inspectorate Division of the Minerals Commission directed that, as per the Minerals and Mining Regulations, 2012 (L.I. 2182), the TSF 2 design be constructed with a compacted soil liner (CSL). As GSWL was well advanced with development of TSF 2 at the time, dispensation was granted for HDPE lining of TSF 2 Cell 1, with all future cells and stage raises to incorporate a compacted soil liner.

TSF 2 Cell 1 construction commenced in July 2016. Verbal approval from the Inspectorate Division to start deposition was given in February 2017 and EPA approval followed in April. Deposition started May 2017.

In July 2017, the Mines Inspectorate Division had completed their review and recommended the re-design of TSF 2 with a compacted soil liner to the Chief Inspector of Mines for approval.

In 2017, GSWL commenced an EIA to support the submission of a Supplementary EIS for TSF 2 Cell 2. The Supplementary EIS was submitted to the EPA in October 2018 with the compacted soil liner design (Figure 18-11). The TSF 2 Cell 2 Supplementary EIS was permitted in August 2020 and Environmental Permit issued in November 2020 (EPA/EIA/533).

 

 

 

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Figure 18-11 TSF 1 and TSF 2 layout (Geosystems, 2018)

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18.6.2TSF 2 Details
18.6.2.1Geotechnical Characterization

A detailed geotechnical investigation comprising sub-soil, in-situ and laboratory testing of soils of the TSF 2 basin was carried out by Knight Piésold Consulting Ltd using test pitting, cable percussion drilling, standard penetration testing, permeability testing, moisture content, grading, Atterberg Limits tests, consolidation tests, triaxial testing on undisturbed soil samples and falling head permeability tests.

Results established the soil profile of the basin, the strength of the foundation soils, and the permeability of the different soil types, to inform in the design of the TSF embankments, base and environmental protection features.

The TSF 2 basin is characterized by a rugged and dissected ground profile that defines the soil profiles in the area according to topographical location. Two main soil types are found in the TSF footprint:

·Alluvial soils formed by deposition of eroded materials from the surrounding hills; and
·Residual soils formed in-situ from the chemical weathering of the underlying base rocks.

Soils can be classified as either of:

·High ground and side-slope soils that are found along slopes and crests of hills, plateau and other high ground that characterizes the TSF footprint; or
·Basin valley and embankment foundation soils that dominate the valleys and low-lying areas.

Guelph permeability tests conducted on nearby surface soils in the valley floor indicated that in some areas the soils have very low permeability (lower than 1.0 x 10-8 m/s). In-situ falling head permeability tests showed that the residual soils, at depths greater than 1.0 m, have a relatively high permeability. Laboratory falling head permeability tests corroborated the field studies and showed that in the valley floor, very low permeability strata exists to approximately 1.0 m depth.

18.6.2.2TSF 2 Design

The TSF 2 design comprises three cells separated by embankments, a temporary embankment and a series of perimeter saddle dams, providing primary containment to ensure that tailings are contained within the valley basin. Other key environmental protection features of the design to enable efficient and appropriate water management include:

·Lining of the base with geomembrane and/or compacted soil liner;
·Spillway;
·Decant barge;
·Secondary confinement;
·Ground water drains; and
·Basin under-drains.

The TSF 2 design assumed a processing rate of 2.7 Mtpa. The facility is designed for a storm capacity of:

·Containment of a 1:100 year, 24 hour duration event with allowance for wave run-up and no flow through the spillway; and
·Safe discharge of a 1:1000 year, 24 hour duration event.

The design of the TSF 2 meets the requirements of the Minerals and Mining (Health, Safety and Technical) Regulations, 2012 (L.I. 2182) and takes due consideration of the recommendations of the International Committee on Large Dams (ICOLD), the Australian Committee on Large Dams (1999) and the Canadian Dam Association guidelines (2007).

  

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Following release of the Global Industry Standard on Tailings Management in August 2020, GSR engaged its Engineer of Record to complete a gap analysis against the standard which is expected to be completed in 2021.

TSF 2 is being constructed in stages and stage storage capacities are presented in Table 18-1. Alternative stage raises have been evaluated to facilitate annual raising during suitable construction weather conditions (Knight Piésold, 2017).

The useful life of TSF 2 will be prolonged by the commencement of paste fill in 2021. On average 40-60% of tails solids are estimated to be used in paste fill and deposited underground.

Beach and bathymetry surveys conducted in September 2020 indicate the current design allows for approximately 33 Mt of further capacity. The Mineral Reserve plan requires processing of 10.8 Mt of ore and schedules estimate 5.0 Mt of the resulting tails solids will be used for paste backfill. Even with the most conservative of paste fill estimates, TSF 2 has more than sufficient design capacity to support extraction of the Mineral Reserve.

Table 18-1 TSF 2 stage design details

 

18.6.2.3Stability Analysis

Stability analyses were conducted for static and seismic loading conditions and static post liquefaction conditions for critical embankments and stages using SLOPE/W® and the Morgenstern-Price method of analysis, which considers force and moments equilibrium of circular slips.

A conservative peak seismic design horizontal ground acceleration of 0.1 g, obtained from “Seismicity of Southern Ghana: Causes, Engineering Implications and Mitigation Strategies” by N.K. Kumapley (1996), was employed in the pseudo static analyses.

For the stability analyses on the upstream slopes, the worst-case scenario was considered, where no tailings are present in front of each embankment stage. For the stability analyses of the downstream slopes, the worst-case scenario was also considered, where the TSF was full to capacity in front of each stage raise (1 m below crest). Modelling scenarios assessed drained and undrained conditions and worst case-phreatic conditions. The assumed conditions combine to present a conservative analysis.

The minimum Factor of Safety (FOS) values calculated for all conditions on both the downstream and upstream slopes were found to meet, and in some conditions exceed the Minerals and Mining (Health, Safety and Technical) Regulations, 2012 (L.I. 2182) requirements for factors of safety.

Stability of the facility was also assessed under the condition where, following the design seismic event, the tailings may be subjected to liquefaction. Seismic stability assessment of the various embankments was conducted in the undrained condition for upstream failure and static drained condition for downstream failure. Tailings were modelled with a residual post-liquefied undrained strength but with no earthquake loading. The minimum FOS values calculated for the post-liquefied condition of the downstream and upstream slopes meet and, in some conditions exceed, the regulatory requirements.

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19Market Studies and Contracts
19.1Market Studies

All gold from Wassa is shipped to a South African gold refinery under a long-term sales contract. Shipping is in the form of doré bars, which average approximately 90% gold by weight with the remaining portion being silver and other metals. The sale price is generally set with reference to the London p.m. fix on the day of the shipment to the refinery.

Gold is a freely traded commodity on the world market and whilst the selling price is subject to fluctuation, the volume of gold produced at Wassa will not be material to the supply/demand balance and will not influence the selling price.

This report considers two gold price assumptions:

·Base Case: for Mineral Reserve estimation and economic test – $1,300 /oz flat; and
·Consensus Case: consensus long-term forecast of 27 banks and financial institutions, as at the end of January 2021:
o2021 – $1,944.26 /oz;
o2022 – $1,879.70 /oz;
o2023 – $1,772.87 /oz;
o2024 – $1,715.61 /oz; and
o2025 and beyond (long-term) – $1,584.68 /oz.
19.2Contracts

The following major contracts are in place to support the Wassa operations:

·Gold sales contract is in-place with Rand Refinery in South Africa;
·Electricity supply (on-site gas generation) – Genser Energy Ghana;
·Fuel and lubricants supply – Ghana Oil Company;
·Electricity transmission from VRA – GRIDCo;
·Electricity supply – VRA;
·Explosives and associated systems – AEL (AECI Ghana Ltd);
·Medical services – International SOS;
·Bulk lime supply – Carmeuse Lime Products (Ghana);
·Site security – Magnum Force Security; and
·Freight forwarding and logistics – Racing Link Express.

All contracts are currently valid and in good standing. Terms, rates and charges of contracts are considered consistent with industry norms. Contract management processes are in place and resourced so that contracts re-tendered and/or renewed as they approach expiry.

 

 

 

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20Environmental Studies, Permitting and Social or Community Impact
20.1Relevant Legislation and Required Approvals

The minerals and mining sector in Ghana is governed by Act 703. It requires mines to obtain environmental approvals from relevant agencies as outlined in Table 20-1. Ghanaian environmental legislation is well developed and enforced by the Environmental Protection Agency (EPA).

20.1.1Permitting Requirements in Ghana
20.1.1.1Environmental Assessment Requirements

Environmental aspects in Ghana are regulated by the EPA Act, 1994 (Act 490). The EPA’s primary legislation for regulation and monitoring of mineral operations are the Environmental Assessment Regulations, Legal Instrument 1652 of 1999 (L.I. 1652), which cover requirements for:

·Environmental permitting;
·Environmental Impact Assessment (EIA);
·Preparation of preliminary environmental reports and environmental impact statements (EIS);
·Environmental certificates;
·Environmental Management Plan (EMP); and
·Reclamation bonding.

The EPA grants environmental approval to projects through an Environmental Permit, which is issued subject to the findings of an EIA, which is documented in an EIS and also covers social aspects. For a mine, an EIS must include a reclamation plan and a provisional EMP. Prior to formal review by the EPA, the EIS may be subject to public exhibition and hearing, with responses from regulators and community to be incorporated into the EIS before an Environmental Permit is granted.

Two years from receipt of an Environmental Permit, an Environmental Certificate is required from the EPA to confirm:

·Commencement of operations;
·Acquisition of all permits and approvals;
·Compliance with mitigation commitments in the EIS and/or EMP; and
·Submission of annual reports to EPA as required.

Within 18 months of commencing operations an EMP must be submitted to and be approved by the EPA. A provisional EMP is included in the EIS which is then updated and incorporated into the mine’s active EMP which is updated every three years over the mine’s life. EMP’s are submitted to and approved by the EPA.

Mines in Ghana are required to have a reclamation plan (Regulation 14 of L.I. 1652) and mining operations submit annual environmental reports (Regulation 25 of L.I. 1652) and monthly environmental monitoring results to EPA, with commentary where values exceed limits and response plans as required.

Relevant guidelines and standards are provided under Act 490, including the Mining and Environmental Guidelines (1994) which provide guidance for: EIS and EMP contents; reclamation plans; EIA procedures; effluent and emission standards; ambient quality and noise levels; and economic instruments.

The EPA conducts routine monitoring of environmental parameters for mines and the results obtained are cross-checked with the monthly results submitted by operations and compared to relevant standards.

The EPA is empowered to suspend, cancel, or revoke Environmental Permits in the event of a breach of L.I. 1652, the permit conditions or the mitigation commitments in the EMP.

 

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Table 20-1 Primary environmental approvals for mines in Ghana

Regulatory institutionApprovals & PermitsCompliance

Environmental Protection Agency

Established under the Environmental Protection Agency Act, 1994 (Act 490), responsible for enforcement of environmental regulations.

Environmental Permit

Under Section 18 of the Mining Act, 2006 (Act 703), and the Environmental Assessment Regulations, 1999 (L.I. 1652), of the EPA, the holder of a mineral right requires an Environmental Permit from the EPA in to undertake any mineral operations.

Approved Environmental Management Plan

EMP to be submitted within 18 months of commencement of operations and updated every three years (Regulation 24 of L.I. 1652).

Environmental Certificate

Must be obtained from EPA within 24 months of commencement of an approved undertaking (Regulation 22 of L.I. 1652).

Reclamation plan

Mine closure and decommissioning plans to be prepared and approved by the EPA (Regulation 14 of L.I. 1652).

Reclamation bond

Mines must post a reclamation bond based on an approved reclamation plan (Regulation 22 of L.I. 1652).

Reporting

Mines submit monthly returns and annual environmental reports to the EPA.

Inspections

EPA undertakes regular inspections to ensure compliance.

Enforcement

EPA may suspend, cancel or revoke an Environmental Permit or certificate and prosecute breaches.

Minerals Commission and Mines Inspectorate Division

Established under the Minerals and Mining Act, 2006 (Act 703), the Minerals Commission administrate mineral rights in trust for the people of Ghana.

Exploration and mining operating plans

Operating Permit from Inspectorate Division required to commence operations. Changes to operating plans to be approved by the Chief Inspector of Mines.

Emergency response plan

An approved emergency response plan must be in place.

Resettlement plan

Resettlement plans to be approved by the district planning authority, according to requirements for compensation & resettlement in L.I. 2175.

Closure Plan

Closure plan to comply with Regulations 273 to 277.

Other

A number of other minor permits and licences are required to support operations (eg: explosives).

Reporting

Mines submit monthly and quarterly returns.

Inspections

Mines Inspectorate undertakes regular inspections to ensure compliance.

Enforcement

Regulations 21 and 22 allow the Mines Inspectorate to issue improvement and/or prohibition notices for contraventions of the Regulations.

Water Resources Commission

Established under the Water Resources Commission Act, 1996 (Act 522), WRC is responsible for regulation and management of the use of water resources.

Approvals for water usage

Under Section 17 of the Mining Act, 2006 (Act 703), the holder of a mineral right may obtain, divert, impound, convey and use water from a watercourse or underground reservoir on the land of the subject of the mineral right, subject to obtaining the requisite approvals under Act 522.

The Water Use Regulations, 2001 (L.I. 1692), regulate and monitor the use of water.

Reporting

Holders submit quarterly and annual reports to the WRC.

Inspection

WRC can inspect works and ascertain abstraction volumes.

Enforcement

Act 522 and L.I. 1692 prescribe sanctions for breaches.

Forestry Commission and Forestry Services DivisionIn accordance with Section 18 of the Mining Act, 2006 (Act 703), a holder of a mining right must obtain necessary approvals from the Forestry Commission. 
20.1.1.2Minerals and Mining Requirements

Act 703 establishes laws on the process for obtaining mineral rights, the administration and management of these rights and protection of the environment. Supporting Act 703 are the Minerals and Mining Regulations, 2012 which cover:

·General aspects (L.I. 2173);
·Compensation and resettlement (L.I. 2175);
·Explosives (L.I. 2177);
·Support services (L.I. 2174); and
·Health, safety and technical requirements (L.I. 2182).

 

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The following regulations have particular relevance to environmental and social management:

·Minerals and Mining (Health, Safety and Technical) Regulations 2012 (L.I. 2182): requirements for approval of mine closure plans and TSF hazard classes outlining requirements for embankment design, factors of safety, impoundments, freeboard, discharge systems, safety arrangements, monitoring, planning, auditing and closure.
·Mining General Regulations 2012 (L.I. 2173): promote preferential employment of Ghanaians and procurement from Ghanaian suppliers. Mines prepare localization plans to achieve this and submit periodic reports (monthly, six-monthly and annual) detailing Ghanaian and expatriate staff numbers, payments of salaries and wages, royalty and corporate tax.
·Mines (Support Services) Regulations, 2012 (L.I. 2174): extend the requirement to preferentially employ Ghanaians to providers of services to mines.
·Mines (Compensation & Resettlement) Regulations, 2012 (L.I. 2175): require that people displaced to conduct mining operations are resettled to suitable alternative land and that livelihoods and living standards are improved. The resettlement plan must be approved by the district planning authority and then given effect by the Minister responsible for Mines.

GSWL has submitted its localization plan to the Minerals Commission covering expatriate staff and the company remains in full compliance with the regulatory requirements.

GSR is listed on the Ghana stock exchange and continues to submit its annual financial reports as required.

20.1.1.3Water Resource Legislation Requirements

The Water Resources Commission Act, 1996 (Act 552) establishes the Water Resources Commission (WRC) and sets requirements regulating the use of water resources. The Water Use Regulations, 2001 (L.I. 1692), and Drilling Licence and Groundwater Development Regulations, 2006 (L.I. 1827), complement the Act by specifying the requirements for obtaining permits for water use, water rights, and priorities for water use; and water drilling licences, and well construction requirements; respectively.

20.1.2Permitting of Existing Operations

A summary of environmental approvals held by GSWL is provided in Table 20-2 and can be summarized as:

·1998 – Satellite Goldfields Limited (SGL): EIS effected approval for development of Wassa, including the original extent of the Main pits complex (South East, 242, F-Shoot, B-Shoot, South, Main South and 419 pits) and processing via a heap leach operation.
·Sep-2002: GSR purchased the project, recommencing operations under Wexford Goldfields Limited (WGL), with ownership 90% GSR and 10% Government of Ghana.
·2004 – WGL: EIS effected approval to construct and commence processing via carbon-in-leach (CIL) and establishment of the tailings storage facility (TSF 1).
·2005: GSR acquired St Jude Resources (SJR) and with it, Hwini Butre and Benso properties to the south.
·2006: Permitting to extend open pit mining to South Akyempim.
·2007: Hwini Butre and Benso (HBB) EIS permitted expansion of open pits with processing at Wassa.
·2010: GSWL (Wassa) Pits Expansion EIS permitted cutbacks at 242, South, Main South, F & B-Shoots. A further EIS effected approval of the Benso G-Zone waste rock dump.
·2011, 2012, 2013: TSF 1 stage raises to 1035.5 mRL, 1037 mRL and 1039 mRL respectively.
·2015: TSF 1 extension, establishment of TSF 2 and permitting of underground exploration.
·2016: TSF 2 re-design and re-permitting.
·2017: Permitting to incorporate underground mining, pit cut back and waste dump extensions.
·2020: Permitting of TSF 2 Cell 2.

 

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Table 20-2 Environmental approvals obtained for Wassa mine

ApprovalPermit No.Date of IssueExpiry DateComments
Environmental Protection Agency – Environmental Permits to commence operations
Approval of the Satellite Goldfields Limited Wassa project EISn/a1998 There are no formal approval documents on record
EIA and EMP for Exploration in Subri River Forest Reserven/a2004 There are no formal approval documents on record
Environmental Permit for the Wassa Power ProjectForm D (0010335)7-May-2004n/aBased on Volta River Authority Wexford Power Project 161 kV Power Transmission Line Bogoso to Akyempim Environmental Scoping Report (2003)
Environmental Permit to pursue operationsEPA/EIA/11218-Mar-2004n/aBased on Wexford Goldfields Limited Wassa project EIS (2004)
Hwini Butre PermitEPA/EIA/17524-Feb-2006n/aSt Jude Resources (Ghana) Limited based on Hwini Butre EIS and Subriso EIS
Benso Subriso Permit
Detox Plant and Discharge to Kubekro Creek ApprovalLetter23-Dec-2005n/a 
South Akyempim Environmental PermitEPA/EIA/1902-Jun-2006n/aBased on EIS on South Akyempim Project (2005)
Hwini Butre/Benso Project Environmental PermitEPA/EIA/2472-Oct-2007n/aBased on the Hwini Butre and Benso EIS (2005)
Wassa Pits Expansion Project Environmental PermitEPA/EIA/32220-Dec-2010n/aBased on Wassa Pits Expansion EIS (2010)
G-Zone Waste Rock Dump Environmental PermitEPA/EIA/32313-Dec-2010n/aBased on Supplementary EIS for G-Zone Waste Dump (2010)
TSF 1 embankment raise to 1035.5 mRLLetter4-Aug-2011n/a 
TSF 1 embankment raise to 1037 mRLLetter9-May-2012n/a 
Environmental Permit for Mineral Exploration (Manso)EPA/PR/PN/7704-Sep-20123-Sep-2014New permit not presently required
TSF 2 PermitEPA/EIA/3835-Apr-20134-Oct-2014Based on corresponding EIS (2013)
TSF 1 embankment raise to 1039 mRLLetter12-Apr-2013n/a 
Father Brown/Dabokrom Supplementary EISLetterInvoiced 14-Jan-2014 Based on Father Brown/Dabokrom Impact Prediction Study (2012)
TSF 1 extension Environmental PermitEPA/EIA/41913-Mar-2015n/aBased on TSF 1 extension EIS (2014)
TSF 2 (re-design) Environmental PermitEPA/EIA/44225-Nov-2015n/aBased on TSF 2 EIS (2015)
Wassa Underground Exploration PermitEPA/PR/PN/9293-Jul-20154-Jul-2017Transitioned to EPA/EIA/508
Wassa Expansion Project Environmental PermitEPA/EIA/50830-Oct-2017n/aBased on Wassa Expansion EIS (2016)
TSF 2 Cell 2EPA/EIA/53328-Aug-2020n/aBased on TSF 2 Cell 2 SEIS (2018)
Environmental Protection Agency – Environmental Certificate
Environmental CertificateEPA/EMP/055Sep-2006Sep-20092006-2009 EMP
Environmental CertificateEPA/EMP/093Apr-2011Apr-2014

Submitted as required by law in 2010.

EPA approved for period 2011-2014

Environmental Certificate Invoiced 2014 Submitted as required by law in 2013.  2014-17 EMP renewal processed by EPA in 2014
Environmental CertificateEPA/EMP/221Jun-2020Dec-2021

Submitted as required by law in 2017

EPA approved for period 2020-2021

Water Resources Commission
Permission to divert Adehesu creek at South Akyempimn/a6-Dec-2006n/a 
Water Use Permit Diversion of Ben and Subri Streamsn/a27-Mar-2008n/a 
Water Use Permit (C Zone fish cages)GSWLID455/1727-Jun-201726-Jun-2020Application for renewal submitted Aug-19 and permit issuance pending.  No activity underway.
Water Use Permit (Mpohor)GSWLID212/191-Jan-201931-Dec-2021 
Water Use Permit (Benso)GSWLID193/191-Jan-201931-Dec-2021 
Water Use Permit (Akyempim)GSWLID134/1/201-Jan-202031-Dec-2022 
Water Use Permit (dewater Wassa Main and Starter)GSWLID134/2/201-Jan-202031-Dec-2022 
Water Use Permit (bores and 242)GSWLID134/3/201-Jan-202031-Dec-2022 
District Assembly
Togbekrom Resettlement PlanWEDA/DEV 159-Jan-2013n/aWassa East District Assembly
Awunakrom Resettlement PlanAWDA/DEV 214-Mar-2013n/aAhanta West District Assembly

 

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20.1.3Environmental Certificate and EMP for Overall Operations

GSWL (then WGL) received the first Environmental Certificate for Wassa for the period September 2006 to September 2009. Since that time GSWL has routinely submitted the 3-yearly EMP as required to maintain the Environmental Certificate in good standing.

The most recent renewal was initiated by submitting the updated EMP to the EPA in December 2017. Following review by the EPA, the Environmental Certificate was invoiced in June 2018, the EMP was finalized and resubmitted, then the Environmental Certificate was issued in 2020.

The Environmental Certificate and the EMP cover all concessions managed by GSWL including Wassa, Hwini Butre (suspended), Benso (suspended) and associated infrastructure including the HBB access road.

20.1.4Notable Conditions of Approval

The Environmental Permit and EIS require a reclamation bond to be posted within one year of commencing operations. The initial reclamation bond for Wassa was posted in November 2004 has been updated periodically as new projects or changes are approved. At the end of 2020 the GSWL bond was $13,672,231.

The mining leases contain conditions relevant to environmental management. The Wassa Mining Lease (LVB7618/94), Benso Mining Lease (LVB26871/07), and Hwini Butre Mining Lease (LVB1714/08) contain conditions to limit encroachment of mining activities on community infrastructure, disturbance of vegetation, conservation of resources, reclamation of land and prevention of water pollution.

20.1.5Permitting of Future Operations

Future changes to the plan which would likely trigger the need for a new permitting are:

·Increasing processing capacity above 2.7 Mtpa;
·Introduction of infrastructure or a new activity outside the permitted footprint, specifically construction of a hoisting shaft to service the Southern Extension area, although infrastructure that can be located within the current open pit excavation is not expected to be subject to EIS/EIA.

If required, the indicative approval timeline for an EIS/EIA process is approximately 2.5 years.

20.2International Requirements
20.2.1Environment and Conservation

The Government of Ghana is party to a number of international treaties relating to the environment:

·Ramsar Convention on Wetlands of International Importance (there are five designated Ramsar sites along the coast of Ghana but none in the Wassa project area).
·Convention of International Trade in Endangered Species.
·United Nations Framework Convention on Climate Change.

Ghana has more than 1,000 IUCN-management protected areas, including 317 forest reserves (EarthTrends, 2003). There are two forest reserves near the Wassa project area; the Bonsa River Forest Reserve and the Subri River Forest Reserve. Approximately 12 km of the Hwini Butre Benso access road traverses the Subri River Forest Reserve.

20.2.2Human Rights

In 2005 GSR, with full support of its Board of Directors wrote to the UN Secretary General as a statement of commitment to adopt the United Nations Global Compact (www.unglobalcompact.org) and GSR continues to integrate the Global Compact principles in its business activities.

GSR’s 2019 Corporate Responsibility Report (formerly Sustainable Development Report) is the 14th public report on how the company is contributing to advance Ghana’s performance against the Sustainable Development Goals in the Global Compact.

 

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The 2019 report incorporated enhanced disclosures including:

·Global Reporting Initiative (GRI) standards;
·Sustainability Accounting Standards Board (SASB), Metals and Mining Sustainability Accounting Standards;
·Mining Local Procurement Reporting Mechanism (Mining LPRM);
·World Gold Council, Conflict Free Gold Standard disclosure;
·Investor Mining and Tailings Safety Initiative (IMTSI) disclosure; and
·World Gold Council, Responsible Gold Mining Principles (RGMPs) disclosure.
20.2.3Anti-Corruption

The Government of Ghana was designated Extractive Industries Transparency Initiative compliant in 2010. To support this GSR provides annual public reports, declaring payments to the Government of Ghana, with significant contributions made by GSR businesses to the end of 2019:

·GSWL payment of more than $259 M over the previous decade; and
·In 2019, the expected royalty distributions from GSWL operations including those to the Office of the Administrator of Stool Lands, Traditional Authorities, Stool Lands and District Assemblies, was over $10.2 M.

GSR are registered in the US and Canada, so are subject to the US Dodd-Frank Wall Street Reform and Consumer Protection Act, the US Foreign Corrupt Practices Act and the Canadian Corruption of Foreign Public Officials Act. Internal GSR policies address these items for GSR management.

20.2.4Voluntary Codes

GSR has adopted a number of voluntary international codes and standards pertaining to corporate responsibility and apply to the Wassa operations:

·Cyanide management – full certification to International Cyanide Management Code since 2010;
·TSFs – current TSF 1 and TSF 2 designs align with the ICOLD requirements;
·Gold mining and processing – as a member of the World Gold Council, GSR ascribes to the Responsible Gold Standard and the Responsible Gold Mining Principles; and
·Resettlement, land acquisition, and compensation – since 2009, GSR has ensured all resettlement projects conform to the International Finance Corporation’s Performance Standard 5 on Land Acquisition and Involuntary Resettlement.

GSR has corporate assurance processes which include independent review, audit and/or validation to ensure conformance of the principles ascribed in these codes and standards.

20.3Environmental and Social Setting
20.3.1Biophysical Setting

The concession area falls within the wet semi-equatorial climatic zone of Ghana and is characterized by an annual double maxima rainfall pattern occurring in the months of May to July and September to October. Average annual rainfall measured at the nearest meteorological station (Ateiku) is 1,996 ± 293 mm. Average annual rainfall measured at the Wassa weather station is approximately 1,750 mm.

20.3.2Hydrology
20.3.2.1Existing Catchment and Flow Paths

The Wassa operations lie within the Pra River basin which is one the two major rivers draining south-western Ghana. The Pra Basin is located in south central Ghana (Figure 20-1) and is extensive, with several river systems traversing the basin.

 

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Topographic elevation of the Pra basin is from sea level, up to 800 m. The highest elevations are to the north and at the eastern edges of the basin, where elevations of 800 m are common. The southern sections are relatively flat to slightly undulating and there are a few peaks in the central regions. The nature and orientation of the highlands determine the flow direction of the drainage network in the basin.

The Wassa mining lease area is drained by tributaries of the Pra, namely the Toe to the far south, Kubekro to the east and the Petetwum to the north. The Petetwum River flows directly into the Pra River and is fed by the Petetwum, Nankadam, and Kumue streams. The Subiri River, locally known as Subri, which drains the western end of the concession, is a tributary of the Bonsa.

 

Figure 20-1 Pra River basin and location of Wassa

The topography of the Wassa site area is generally undulating, dissected by steep-sided valleys and incised by an extensive, largely dendritic drainage network (Figure 20-2).

The project site is located high in its local catchments such that most surface water comes from rainfall runoff, rather than stream flows entering the site.

 

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Figure 20-2 Wassa topography and drainage with sub-catchments

20.3.2.2Surface Water Management

Multiple surface water studies have been undertaken over the project’s life, the most relevant to the current project being:

·2015 feasibility study to establish underground mining (SRK); and
·Various studies in 2010-2014 for initial TSF 2 permitting and subsequent redesign and permitting (Knight Piesold and Geosystems Consult).

Established surface water management features on site include a stream diversion around the processing plant and administration area, TSF drainage diversions and a French drain to prevent water inflow to the southern end of the Main pit area. These water management features were engineered and constructed as permitted through EIS/EIA processes.

In addition to these main features, secondary drainage works are in place around site to direct water runoff from dumps and roads, away from active mining areas or toward dewatering infrastructure.

Surface water management features are maintained by the surface earthworks and underground mining workgroups as needed.

20.3.3Hydrogeology

The Wassa site falls in within the Birimian Province and it is characterized by aquifers of the Birimian metasediments and metavolcanics and the Tarkwaian aquifers.

Extensive baseline studies were conducted in 1995 and 1996 by Minerex Environmental Limited prior to the development of Wassa operations which were extended by subsequent EIA/EIS processes.

In 2015 and 2019 expanded hydrogeological field studies were conducted by Golder Associates to testing hydrogeological conditions in the underground Mineral Resource. Hydraulic (packer) testing was conducted on selected exploration holes to provide data for conceptual and numerical hydrogeological models. The two programs found the main lithological units to be saprolite, saprock, fractured and fresh bedrock.

 

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Groundwater typically occurs in a shallow, weathered zone aquifer and a fractured deeper bedrock aquifer. The aquifers have overall low permeability (4 x 10-7 m/d), significantly enhanced along tabular zones where interconnected joint sets, faulting and/or quartz vein occur (6 x 10-2 m/d).

Inflow to the underground workings occurs along these discreet zones of faulting and fracturing (Figure 20-3), however these higher permeability areas are isolated, forming a very small percentage of the overall rock mass and therefore only localized higher inflow in the underground workings (Figure 20-4).

 

Figure 20-3 Conceptual underground water flow path model (Golder, 2016)

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Figure 20-4 Conceptual groundwater model (Golder, 2016)

Ground water elevation contours show that generally the groundwater flows in a south-westerly direction following the major topographical features, with the site primarily drained by the Kubekro catchment network of streams. Near the active open pits, the prevailing hydraulic head is towards the open pit in response to the active pit dewatering (Figure 20-5).

 

 

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Figure 20-5 Groundwater contours and flow (Golder, 2016)

 

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Figure 20-6 Conceptual geo-environmental model, E-W section (Golder, 2016)

At Wassa shallower gold deposits have been extracted from pits with voids now empty or in various stages of backfilling with waste rock. Waste rock not used for backfilling is stockpiled in waste rock dumps adjacent to or west of the Main pit complex (Figure 20-6). Precipitation on these catchments and storm water inflow into mining areas also contributes to underground water make.

To understand these influences environmental stable isotope (ESI) signatures were investigated by Golder in 2017 to identify the origin of water ingress into the underground workings. ESI analysis provides the measurable variation of the stable isotopes against the natural values (Dansgaard, 1964). The analysis found that key pits share a common water source with the underground water bearing zones which suggests the pits and underground workings are hydraulically connected.

Due to the potential connection between the surface and underground workings, water volumes in the pits are monitored to ensure water inventory is managed. These controls complement the surface water management systems described in Section 18.2.

Data from the various hydrogeological studies have informed numerical groundwater model (FEFLOW) development. The model suggests that the permeability will decrease with depth in the saprolite to act as a confining horizon (Golder, 2016). Groundwater recharge is considered low (approximately 10 mm/a or 0.5-1.0%). The reported dewatering rates pumped to the underground mine range from 5-10 l/s and from the underground mine ranges from 45-120 l/s. The implications are:

·The drawdown (dewatering cone) from the operations is not expected to have any significant effect on existing (community) groundwater boreholes;
·Groundwater in the shallow weathered zone and deeper bedrock aquifers flows from the elevated areas towards the rivers, following the topography;
·Preliminary modelled underground water production is within current permitted abstraction; and
·The receiving environment will not receive underground mine leachate in the recovered state and no decant is expected to occur. Additionally, leachate from mine waste rock dumps is controlled by the cone of depression and is not expected to impact on the receiving environment.

 

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