PRELIMINARY ECONOMIC ASSESSMENT
on the
BIG HILL DEVELOPMENT PROJECT
STAWELL GOLD MINE
Victoria Australia
For
CROCODILE GOLD CORPORATION
Job No. 1723_M Mining One Pty Ltd Doc No. 3438.doc Level 9, 50 Market Street Date: January 2013 Melbourne VIC 3000 Prepared by: Mark Van Leuven Ph: 03 9600 3588 Fax: 03 9600 3944
FINAL REPORT PRELIMINARY ECONOMIC ASSESSMENT
STAWELL GOLD MINE
TABLE OF CONTENTS 1 SUMMARY ...... 2 1.1 Introduction and Terms of Reference ...... 2 1.2 Property Description and Ownership ...... 2 1.3 Geology and Mineralisation ...... 3 1.4 Status of Exploration, Development and Operations ...... 4 1.5 Processing ...... 4 1.6 Mineral Resource and Mineral Reserve Estimates ...... 5 1.7 Conclusions and Recommendations ...... 7 2 INTRODUCTION ...... 8 2.1 Certificate of Qualified Persons (Ni43-101) ...... 9 3 RELIANCE ON OTHER EXPERTS ...... 11 3.1 External Consultant Reports and Reviews ...... 12 4 PROPERTY DESCRIPTION AND LOCATION ...... 13 4.1 Property Description ...... 14 4.2 Legislation and Permits ...... 15 4.3 SGM Local Survey Grid Reference ...... 17 5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE & PHYSIOGRAPHY 19 6 HISTORY ...... 20 6.1 Historical and Modern Production ...... 21 7 GEOLOGICAL SETTING AND MINERALISATION ...... 22 7.1 Regional Geology ...... 22 7.2 Local Geology ...... 23 7.2.1 Stratigraphy at Stawell ...... 23 7.2.2 Magdala Basalt...... 23 7.2.3 Leviathan and Albion Formations...... 23 7.2.4 Magdala Facies ...... 25 7.2.5 Felsic Intrusions ...... 25 7.2.6 Stawell Granite ...... 25 7.2.7 Lamprophyric Intrusions ...... 25 7.2.8 Structural History at Stawell ...... 25 7.2.9 Early, Ductile Deformation ...... 26 7.2.10 Late, Brittle Deformation ...... 26 7.2.11 Stawell Mine Geological Architecture ...... 28 7.3 Mineralisation ...... 32 8 DEPOSIT TYPES ...... 34 8.1 General Deposit Types and Mineralisation ...... 34 8.2 Big Hill Deposit Types ...... 35 8.2.1 Mariner’s ...... 37 8.2.2 Allen’s Zone ...... 38 8.2.3 Iron Duke Zone ...... 39 8.2.4 Magdala Flanks ...... 39 8.2.5 Scotchman’s Fault: ...... 39 8.2.6 Cross Course Fault: ...... 40 8.2.7 Lower Cross Course Fault: ...... 40 8.2.8 Basalts: ...... 40 P:\1723_M\.3438_Final_w_signatures_Rev6.docx i
PRELIMINARY ECONOMIC ASSESSMENT
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8.2.9 Foliation and shearing ...... 40 9 EXPLORATION ...... 42 9.1 Current Exploration ...... 42 9.2 Expenditure ...... 42 9.3 Exploration and Mineral Tenure...... 43 9.4 Kiska Metals Corporation Joint Venture ...... 45 9.5 Interest in Navarre Minerals ...... 46 10 DRILLING ...... 47 10.1 Stawell Gold Mines Mineral Resource Definition Process ...... 47 10.2 Drilling Process ...... 50 10.2.1 Big Hill Drilling ...... 52 10.2.2 Reverse Circulation Drilling ...... 55 10.2.3 Diamond Drilling ...... 55 10.2.4 Drill hole Orientation ...... 56 10.2.5 Collar Survey Control ...... 58 10.2.6 Down hole Survey Control ...... 59 10.2.7 Down hole Survey Quality Control ...... 62 10.2.8 Diamond Drill Core Processing ...... 64 10.2.9 Logging ...... 66 10.2.10 Core Recovery ...... 66 10.2.11 Diamond Drill Core Sampling ...... 67 10.2.12 Big Hill Diamond Drill Core Sample Intervals ...... 67 10.2.13 Big Hill RC Sampling ...... 68 10.2.14 Reliability of Samples ...... 68 11 SAMPLE PREPARATION, ANALYSIS AND SECURITY ...... 70 11.1 Assay Laboratories ...... 70 11.1.1 Sample Preparation ...... 71 11.1.2 Sample Transport and Security ...... 72 11.1.3 Assay Methods ...... 73 11.1.4 Big Hill Re-logging and Re-assaying...... 74 11.2 Database Storage and Integrity ...... 76 11.3 Big Hill Assay QA/QC Process ...... 77 11.3.1 Big Hill Historical QAQC process ...... 77 11.3.2 Big Hill 2012 QAQC process ...... 79 11.3.3 QAQC Checks and actions ...... 80 11.3.4 Standards ...... 80 11.3.5 Lab Duplicates and Repeats ...... 84 11.4 QAQC Results for the Current Report ...... 89 12 DATA VERIFICATION ...... 91 12.1 Collar Locations ...... 91 12.2 Down hole Survey ...... 92 12.3 Core photos and logging ...... 92 12.4 EOH ...... 92 12.5 Assay Records ...... 92 13 MINERAL PROCESSING AND METALLURGICAL TESTING ...... 93 13.1 Metallurgical Test Work ...... 93 13.2 Davis Pit Area Test Work ...... 93 13.3 Big Hill Area Test Work ...... 94 P:\1723_M\.3438_Final_w_signatures_Rev6.docx ii
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13.4 Davis Pit Historical Treatment ...... 94 13.5 Expected Recovery ...... 95 14 MINERAL RESOURCE ESTIMATES ...... 96 14.1 SGM Mineral Resource Estimate Dec 2011 ...... 96 14.2 Introduction and Scope of Big Hill Mineral Resource Estimate ...... 96 14.3 Big Hill Resource Summary ...... 98 14.4 Pit Shell ...... 98 14.4.1 Void Modelling ...... 99 14.4.2 Weathering model ...... 102 14.4.3 Bulk Density ...... 102 14.4.4 Moisture Determinations: ...... 103 14.4.5 Further Resource Work and Recommendations ...... 103 14.5 Geological Modelling ...... 103 14.6 Big Hill Summary of Geological Modelling ...... 105 14.6.1 Block Modelling ...... 108 14.6.2 Block Model Dimensions ...... 108 14.6.3 Block Model Coding ...... 109 14.6.4 Drill hole Coding ...... 109 14.6.5 Compositing ...... 109 14.6.6 Geostatistical Parameters ...... 109 14.6.7 Resource Classification ...... 113 14.7 External Factors Effecting Extraction of Mineral Resources and Reserves ...... 113 14.7.1 Magdala Surface Mineral Resources ...... 113 15 MINERAL RESERVES ESTIMATES ...... 114 16 MINING METHODS ...... 115 16.1 Underground Mining Methods ...... 115 16.1.1 Open Pit Methods ...... 115 16.1.2 Underground Access from Open Pit ...... 115 16.2 Whittle Optimisation of Open Pit ...... 116 16.2.1 Block Model Preparation ...... 116 16.2.3 Optimisation Methodology ...... 117 16.2.4 Optimisation Parameters ...... 117 16.2.5 Whittle Optimisation Results ...... 118 16.3 Pit and Dump Design ...... 124 16.4 Pit Schedule ...... 128 16.4.1 20Mt Case ...... 129 16.4.2 10Mt Case ...... 131 16.4.3 100m Buffer Case ...... 134 17 RECOVERY METHODS ...... 137 17.1 Mineral Processing ...... 137 18 PROJECT INFRASTRUCTURE ...... 139 18.1 Surface Infrastructure ...... 139 18.2 Tailings Storage Facilities ...... 139 18.3 Power Update to Big Hill ...... 140 18.4 Water ...... 141 19 MARKET STUDIES AND CONTRACTS ...... 142 20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT ...... 143
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20.1 Environmental Issues and Studies ...... 144 20.2 Waste Management ...... 148 20.3 Permitting Requirements ...... 151 20.4 Heritage, Social Impact & Community Engagement ...... 151 20.5 Mine Closure (remediation and reclamation) Requirements and Costs ...... 157 21 CAPITAL AND OPERATING COSTS ...... 159 21.1 Capital Estimate ...... 159 21.2 PFS Studies -$1.1M ...... 159 21.3 Geotechnical Drilling - $0.5M ...... 159 21.4 Equipment Mobilisation - $1.8M ...... 159 21.5 Plant sustaining capital - $0.5 ...... 159 21.6 Communications tower $1.05M ...... 159 21.7 Dams- $3.0M ...... 160 21.8 Noise Barrier- $3.5M ...... 160 21.9 Noise and dust monitoring -$ 0.10M ...... 160 21.10 Purchase of houses - $ 0.5M ...... 160 21.11 Permitting and Tech. Studies - $ 2.5M ...... 160 21.12 Compensation for houses -$1.0M ...... 161 21.13 Big Hill & Davis Oxide Studies-$0.2M ...... 161 21.14 Wall raise of TSF2 - $1.1M ...... 161 21.15 Project Management - $0.3M ...... 161 21.16 Explanation of the differential in capital expenditure of the different Pits...... 164 21.17 Operating Cost Estimate ...... 165 22 ECONOMIC ANALYSIS ...... 166 22.1 Taxes ...... 168 22.2 Royalties ...... 169 23 ADJACENT PROPERTIES ...... 170 24 OTHER RELEVANT DATA AND INFORMATION ...... 171 25 INTERPRETATION AND CONCLUSIONS ...... 172 26 RECOMMENDATIONS ...... 173 27 REFERENCES ...... 175
TABLE INDEX Table 1-1 Big Hill Mineral Resources November 2012 estimate ...... 5 Table 1-2 Summary of Mining Physicals and Economics ...... 6 Table 3-1 Site Experts who contributed to the Technical Report ...... 11 Table 8-1 Diamond Core Intercepts ...... 37 Table 9-1 Regional Tenement Information ...... 44 Table 10-1 SGM geological processes and approximate drill spacings ...... 49 Table 10-2 Summary of Big Hill Drilling ...... 53 Table 10-3 Diamond drill core sample interval statistics for samples used in November – 2012 Resource Estimate ...... 68
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Table 11-1 Laboratory assay method codes, descriptions and limits of detection...... 74 Table 11-2 Range of standards used at SGM ...... 80 Table 11-3 SGM assay standards and certified values reported ...... 89 Table 12-1 Data verification work completed for the 2012 Mineral Resource update ...... 91 Table 13-1 Davis Pit Test work ...... 93 Table 14-1 December 2011 Final Mineral Resource Calculation for Stawell Gold Mines ...... 96 Table 14-2 Big Hill Mineral Resource Estimate 1998 and November 2012 comparisons ...... 97 Table 14-3 Big Hill Drilling Summary ...... 97 Table 14-4 Big Hill Mineral Resources November 2012 estimate ...... 98 Table 14-5 Geological Domains at Big Hill 2012 ...... 106 Table 14-6 Block sizes utilised in SGM local area block models ...... 108 Table 14-7 Big Hill variography summary ...... 110 Table 14-8 Big Hill Search Parameter summary ...... 111 Table 16-1 Block Model Geometry ...... 116 Table 16-2 Additional Attributes ...... 116 Table 16-3 Summary of Optimisation Parameters ...... 118 Table 16-4 $1,400/ounce - Optimisation Parameters and Results, 3 cases ...... 120 Table 16-5 Pit Design Inventory ...... 124 Table 16-6 Pit Inventories of Mines ...... 128 Table 16-7 Summary of Key Physicals for 20Mt Case ...... 129 Table 16-8 Summary of Key Physicals for 10Mt Case ...... 132 Table 16-9 Summary of Key Physicals for 100m Buffer Case ...... 134 Table 20-1 Estimated costs to rehabilitate the Big Hill Project area if the project proceeds ...... 158 Table 21-1 Capital for 20Mt Case ...... 162 Table 21-2 Capital for 10Mt Case ...... 163 Table 21-3 Capital for 100m Buffer Case ...... 164 Table 22-1 Whittle Pit Optimisation Results at $1400/oz ...... 166 Table 22-2 Pit Schedule Results ...... 167
FIGURE INDEX Figure 4-1 Map Highlighting Location of SGM Figure 4-2: Location Map of the MIN5260 and MIN5520 Lease. The Grid is Latitude and Longitude (GDA94) Figure 4-3: Aerial View of the MIN5260 and MIN5520 Lease indicating Underground and Pit Locations
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PRELIMINARY ECONOMIC ASSESSMENT
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Figure 4-4: Local Mine Grid at SGM Figure 6-1: Annual Gold Production since 1984 Figure 6-2: Annual Processing Plant Production since 1984 Figure 7-1: Image showing Lachlan Fold Belt, Locating Stawell on the Western Boundary Figure 7-2: Mine Geology Cross-section highlighting Architecture of the Magdala and Golden Gift Ore Bodies Figure 7-3: D1 to D5 Ductile and Brittle Evolution of the Stawell System Steronets. Figure 7-4: Evolution of the Stawell system from 420 to 380 Ma (modified from Miller & Wilson 2004a). Figure 7-5: Plan View Geological Interpretation of the Stawell Structural and Stratigraphic Architecture at 1000mRL. Figure 7-6: SGM Longitudinal Projection showing the Location of the Mineralised ore block. The geological and spatial relationship between the Magdala, Golden Gift and Wonga deposits can clearly be seen. Figure 7-7: Figure Example of Central Lode Mineralisation Figure 7-8: Example of Basalt Contact Mineralisation Figure 8-1: Section 5690 through Big Hill North, showing the Cross Course Fault (blue), Mariners ore domain (green) and Allen’s ore domains (pink). Figure 8-2: Section 5510 through Big Hill South, showing Scotchman’s Fault (upper blue surface), Iron’s Duke ore domains (purple), Magdala ore domains (red) and basalt waste domain (yellow). Figure 8-3: Section 5090 through Davis area, showing Magdala ore domains (purple) and basalt waste domain (yellow). Note the previously mined Davis depletion. Figure 8-4 Section through Mariner’s lode showing structural complexity and offsetting Figure 9-1 Mine Geology Longitudinal section outlining near mine exploration from September 2008- October 2012 Figure 9-2 SGM regional Victorian tenement areas Figure 10-1 SGM Diamond drilling process flowsheet Figure 10-2 An example of a daily drill record from Big Hill 1998 surface drill programme. Figure 10-3 Plan showing Big Hill pit used for Resource reporting, drill coverage and drill orientations. Noting recent 2012 RC infill drilling in red. Used in both the Resource estimate and void modelling Figure 10-4 3D view looking towards the NE, note displayed drill hole traces are holes before MD1000, all are drilled from underground and spatially occur outside the pit shell used for Resource reporting. Figure 10-5 Plan view (top) and cross section views (A-B and C-D) ) (Note: Potential Pit = grey, Surface topography including previously mined Davis pit = red, Davis lode = light green, Iron Duke lode = dark green, Allen’s lode = light blue and Mariner’s lode = dark blue. Figure 10-6 Collar survey information and drill hole survey information checklist 1997 to 98 programme Figure 10-7 Collar survey information and drill hole survey information checklist, modern process For 2008 and 2012 programme
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PRELIMINARY ECONOMIC ASSESSMENT
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Figure 10-8 Plot of drillhole survey method (SE = Eastman Single Shot in purple, SD = Electronic survey instrument in blue) by time. Post 2001, the standard survey instrument used has been an electronic single shot downhole survey tool Figure 10-9 Magnetic declination correction as currently applied to SGM drill hole data Figure 10-10 Survey record sheet, example from Big Hill 1998 surface programme Figure 10-11 An example of drillhole collar survey information collated for geologist review. Note some corrections to erroneous data have been made as indicated Figure 10-12 Example of the check plots used to correct down hole survey information Figure 10-13 Photograph of the SGM survey camera test bed. Figure 10-14 SGM core photography installation Figure 10-15 Diamond drillcore processing flow sheet Figure 10-16 An example of Big Hill diamond drill core photographs stored digitally from scanned slides, 1998 drill programme Figure 11.1.1-1 SGM drill core sample preparation, assay and QA/QC flowsheet Figure 11.1.4-1 Diamond Core re-logging Programme Figure 11.1.4-2 Aqua Regia vs Fire Assay grades for re-logged core Figure 11.1.4-3 Aqua Regia vs Fire Assay grades – dilution errors Figure 11.1.4-4 Showing a Log/log graphical representation of the field assay repeats from the July 1998 model update drill holes. Figure 11.1.4-5 Showing a Log/log graphical representation of the monitored lab duplicate results from the July 1998 model update drill holes. Figure 11.1.4-6 Showing a Log/log graphical representation of assay repeats which is used to analyse the lab assay repeatability. Figure 11.1.4-7 Assay standard performance for Big Hill 2012 RC drill programme and Nov 2012 resource model update. Figure 11.1.4-8 OR15h standard performance for Big Hill 2012 RC drill programme and Nov 2012 resource model update. Figure 11.1.4-9 OR15g standard performance for Big Hill 2012 RC drill programme and Nov 2012 resource model update. Figure 11.1.4-10 Precision of the lab duplicate for Big Hill 2012 RC drill programme and Nov 2012 resource model update. Figure 11.1.4-11 Assay duplicate comparison for Big Hill 2012 RC drill programme and Nov 2012 resource model update. Figure 11.1.4-12 Log Log graphical representation of the assay repeats form Big Hill 2012 RC drill programme and Nov 2012 resource model update. Figure 11.1.4-13 SGM QA/QC review and actions flow sheet. Figure 11.1.4-14 QA/QC diary entries for 2012 programme Figure 12-1 Rounding discrepancy
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Figure 14-1 Plan view, showing Big Hill pit optimisation (black) used for current Resource estimate reporting and model boundaries (orange) previously used for reporting in 1998 Figure 14-2 Long Section view looking NE, showing Big Hill pit optimisation (black) used for current Resource estimate reporting and 130mRL (orange) previously used for reporting above Figure 14-3 Plan view showing potential Big Hill pit and underground voids (light green) and Davis Pit void as mined (dark green) Figure 14-4 Long Section view looking NE, showing potential Big Hill pit and underground voids (light green) and Davis Pit void as mined (dark green) Figure 14-5 Bulk density of material Figure 14-6 Section 5690 through Big Hill North, showing the Cross Course Fault (blue), Mariner’s domain (green) and Allen’s domains (pink). Figure 14-7 Section 5510 through Big Hill South, showing Scotchman’s Fault (upper blue surface), Iron Duke domains (purple), Magdala domains (red) and basalt waste domain (yellow). Figure 14-8 Section 5090 through Davis area, showing Magdala domains (purple) and basalt waste domain (yellow). Note the previously mined Davis depletion. Figure 16.4-1: Davis and Big Hill Optimisation Process Figure 16.4-2: Selected Whittle Shells, 20Mt Case and 10Mt Case Figure 16.4-3: Selected Whittle Shells, 100m Buffer Case Figure 16.4-4: Whittle Shell A$1,400/oz – 20 Mt Case Dimensions Figure 16.4-5: Whittle Shell A$1,400/ounce - 20Mt Case (Looking North-East) Figure 16.4-6: Whittle Shell A$1,400/oz – 10 Mt Case Dimensions Figure 16.4-7: Whittle Shell A$1,400/ounce - 10Mt Case (Looking North-East) Figure 16.4-8: Whittle Shell A$1,400/ounce – 100m Buffer Case Dimensions Figure 16.4-9: Whittle Shell A$1,400/oz – 100m Buffer Case (Looking North-East) Figure 16.4-10: Pit Design A$1,400/oz – 20Mt Case Dimensions Figure 16.4-11: 20Mt Pit Design & Mineralised Material (Looking North-East) Figure 16.4-12: Pit Design A$1,400/oz – 10Mt Case Dimensions Figure 16.4-13: 10Mt Pit Design & Mineralised Material (Looking North-East) Figure 16.4-14: Pit Design A$1,400/oz – 100m Buffer Case Dimensions Figure 16.4-15: 100m Buffer Pit design & Mineralised Material (Looking North-East) Figure 16.4-16 Dump Design and location Figure 16.4-17: 20Mt Case – Mill Feed Au Figure 16.4-18: 20Mt Case – Total Movement Figure 16.4-19: 20Mt Case – Open Pit Au Ounces by Pit Figure 16.4-20: 20Mt Case – Individual Open Pit Total Movement Figure 16.4-21: 20Mt Case – Individual Waste Dump Movement
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Figure 16.4-22: 10Mt Case – Mill Feed Au Figure 16.4-23: 10Mt Case – Total Movement Figure 16.4-24: 10Mt Case – Open Pit Au Ounces by Pit Figure 16.4-25: 10Mt Case – Individual Open Pit Total Movement Figure 16.4-26: 10Mt Case – Individual Waste Dump Movement Figure 16.4-27: 100 Buffer Case – Mill Feed Au Figure 16.4-28: 100 Buffer Case – Total Movement Figure 16.4-29: 100 Buffer Case – Open Pit Au Ounces by Pit Figure 16.4-30: 100 Buffer Case – Individual Open Pit Total Movement Figure 16.4-31: 100 Buffer Case – Individual Waste Dump Movement Figure 17-1 SGM treatment plant flow sheet Figure 18.2-1 Plan showing the location of MIN 5260, Stawell Gold Mines operational infrastructure. Figure 20-1 Survey area for the Big Hill project EES Figure 20-2 Known registered Aboriginal archaeological sites around the Stawell area Figure 22-1: Annual and Cumulative Discounted Cashflow Figure 26-1: Study and Project Execution Timeline
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Term Description
Au gold
AUD Australian dollars
g/t grams per metric tonne
m metre
km kilometre
Ha Hectare (10,000 sq m)
Ma Million Years
mRL Mine grid reduced level (metres)
ML Mega litres
Crocodile Crocodile Gold Corp
SGM Stawell Gold Mines
Shotcrete sprayed concrete slurry for support of underground mine openings
CRF Cemented Rock Fill used as stope backfill
C7 Central Lode 7 ore block
U3 Upper South Fault ore block
GG1 Golden Gift 1 ore block
GG2 Golden Gift 2 ore block
GG3 Golden Gift 3 ore block
GG5 Golden Gift 5 ore block
GG5L Golden Gift 5 Lower ore block
GG6 Golden Gift 6 ore block
GG South Golden Gift South ore block
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1 SUMMARY
1.1 Introduction and Terms of Reference The authors have prepared this Technical Report for Crocodile Gold in respect of its Big Hill Project located in Stawell, Western Victoria, Australia. Big Hill is the up dip extension of the Magdala system which is currently being mined from an underground perspective. The underground operation has been in continual operation since 1981 and has produced in excess of 2 million ounces during this period. Crocodile Gold acquired Stawell Gold Mines including the Big Hill project in May 2012 through an acquisition of AuRico Gold’s Australian assets. Crocodile Gold owns a 100% interest in the Big Hill project with a AUD$2.00 per gold ounce royalty payable to Minerals Ventures of Australia. This Technical Report represents a Preliminary Economic Assessment (PEA) of the Big Hill project. It outlines a possible open pit operation based on an updated Resource estimation, optimisation study and design works carried out during the latter half of 2012. The PEA is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorised as mineral reserves, and there is no certainty that the PEA will be realized. This Technical Report conforms to the CIM Definition Standards on Mineral Resources and Mineral Reserves referred to in National Instrument 43-101 – Standards of Disclosure for Mineral Projects (“NI 43-101”). The qualified persons (each, a “QP”) who supervised the preparation of this Technical Report are: Mark Van Leuven BEng in Mining Engineering, MBA (Technology Management), FAusIMM(CP), Principal Mining Engineer for Mining One consultants of, Melbourne, Victoria Stuart Hutchin BSc, Applied Geology, MAIG, MAusIMM, Geology Manager for Mining One consultants of, Melbourne, Victoria, Australia Both QP’s are independent of the issuer as defined by NI 43-101.
1.2 Property Description and Ownership The Big Hill Enhanced Development Project is located adjacent to the Stawell township in western Victoria and is confined to an area within the existing Stawell Gold Mines (SGM) licenced boundary located to the north-east of Stawell. The Project site is bound by Crowlands Road to the north, Leviathan Road to the east, Albion Road and Fisher Street to the west and Main Street to the north- west. The Project is proposed to be located on publicly accessible Crown land which is currently occupied by a Pioneer Memorial rotunda, car park, Scenic Road, Reefs Road, CFA tower and communications tower, water reservoir 1, 4 and 6, water tanks , water supply system memorials as well as on land currently utilised by SGM operations (former Davis Pit). The Project site is located on Crown land which is partly leased to SGM. The northern part of the site is accessible to the public as informal recreational reserve, whilst the eastern portion of the site is primarily utilised by SGM. The Project site is predominately Crown land under the control of DSE, with three exceptions, namely:
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¾ CA 11B is a Public Purpose Reserve – Committee of Management is Northern Grampians Shire Council; ¾ CA 10A & 11A comprise a Reservoir Reserve – Committee of Management is GWM-Water; and ¾ CA 10F is freehold land owned by GWM-Water. Within the Crown land, apart from the mine ventilation shaft, aboveground mine development and some memorials and a picnic facility, four small areas of land are developed for specific uses, namely: ¾ CA 10B a former Forest Commission Reserve, containing a DSE Fire Tower. ¾ CA 10C, a former Municipal Purposes Reserve leased to Shire of Stawell is now leased to Vencorp and houses the organisation’s radio communication facility. ¾ CA 10D a former Police & Emergency Services Reserve, is leased to the Victorian Police. It houses the State Mobile Radio Network (Telstra). ¾ CA 10G is leased by Optus Communications and contains a mobile telephone tower and three buildings. It houses the mobile telephone facilities for Optus, Telstra and Vodaphone.
1.3 Geology and Mineralisation The Stawell Goldfield is located in the western Stawell Zone of the Lachlan Fold Belt. The Stawell Zone is a belt of predominantly deformed meta-sedimentary rocks representing the lower parts of the Cambro-Ordovician Lachlan Fold Belt stratigraphy bound to the west by the Moyston Fault and to the east by the Coongee Break (Vandenberg et al. 2000). The stratigraphy at Stawell is divided into three principal units: Magdala Basalt; Albion Formation; Leviathan Formation. Intruded into this sequence are the Stawell Granite and a number of felsic and mafic intrusions. Squire and Wilson (2005) interpret that the rock unit previously termed the Magdala Volcanogenics (Watchorn and Wilson, 1989) is an alteration facies that locally occurs adjacent to the basalt. Big Hill is the up dip extension of the Magdala system which has been historically mined from underground. It is located in the western Stawell Zone of the Lachlan Fold Belt. It contains Basalt Contact mineralisation, Central Lode mineralisation and Stockwork mineralisation, all typically seen in the Magdala system. Big Hill geology and mineralisation can be broken into 4 main domains Mariner’s, Allen’s, Iron Duke and Magdala Flank. All except Mariner’s and Allen’s are separated by faults. Mariner’s has a consistent width and orientation, averaging around 14m thick and dipping to the northwest (mine grid). It constitutes a shear package, with massive quartz veining. Allen’s is a series of massive quartz veined zones very steeply dipping to vertical to stockwork style veining seen near the top of the zone. Much of the geology consists of well bedded, foliated volcanogenic sediments which are unpredictably mineralized. The Iron Duke zone is a wedge of geology occurring between the Scotchman’s Fault and the Lower Cross Course Fault. It is the up-dip extension of the volcanogenic package/shear zone which has been offset from the main Magdala shear system by reverse movement on the Lower Cross Course Fault. Magdala Flank describes the package of lodes lying to the west of and continuing up dip from the nose of the Magdala basalt. It consist of volcanogenic sediments, stock worked sediments and quartz lodes.
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In general oxidation can be seen down to around 90m below surface with little to no sulphides in this area. As the mineralised zones transition and more primary materials are apparent gold is found associated with associated pyrite, arsenopyrite and recrystallised pyrrhotite.
1.4 Status of Exploration, Development and Operations Significant exploration progress and successes has been made over the life of the Stawell project including the discovery of the current mineral resources and mineral reserves. In the period 2008 to 2012, of note is the conversion of the Golden Gift 6, discovered in 2007, into reserve and the Golden Gift 6 Lower, discovered in 2008, into inferred resource. Scoping the northern extension of the Magdala dome has also added to the C7 (Dukes) reserve. Exploration work in 2012 is forecast to be $2.4 million and is primarily directed toward regional soil sampling and drill targets. 2013 will see a continuation of regional soil campaigns and the evaluation of these for drill targeting. Historically a number of exploration campaigns have been undertaken over the Big Hill area. Drilling was on and off up to 1997 when a major programme of drilling was undertaken to define the potential of the Big Hill project. During 1997 to 1998, 217 reverse circulation drill holes were drilled for 19,500m, and 30 cored diamond holes for 2,369m were also drilled. A later 2008 RC programme was completed over the Davis area or the southern end of Big Hill to better constrain the grade in this area. The most recent work was a RC programme completed over the Davis portion of Big Hill in early 2012. A total of 34 holes were completed in this programme for 2,238 meters. Underground mining operations at Stawell Gold Mines has been continuous since 1981 with the extraction of gold bearing ore from the Magdala and Golden Gift systems to a vertical depth of - 1200mRL and -1600mRL respectively. Underground mine production is scheduled for completion in 2013 as mineable reserves are exhausted at depth. The present underground mine production environments are immediate to the Wildcat Fault where exploration programmes determined a significant offset to the Magdala and Golden Systems which are considered uneconomically viable given the magnitude of structural offset. Remnant mining activity has been explored over the past years as strengthened gold pricing has enabled the revisiting of marginal resources vertically accessed. The proposed Big Hill Development project will be considered prior to exercising mine closure conditions at Stawell Gold Mines in full. Processing of both remnant low grade oxide stockpiles and the remaining underground ore will be undertaken in 2013 whilst this consideration takes place.
1.5 Processing The processing plant at SGM has been in operation since 1984 and has a well-established flow path to accommodate the treatment of both oxidized and fresh ore types. The flow path consists of five unit processes, namely: ¾ Size Reduction (Crushing and Milling) ¾ Gravity gold recovery ¾ Flotation/ultrafine milling ¾ Leach-adsorption, and ¾ Gold recovery
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Both of the Flotation and Gravity sections of the plant can be isolated to allow tailored treatment of oxide, transitional and fresh ore types to optimize gold recovery and minimize treatment costs. Ore from the Davis Open Cut has been previously processed at SGM with close to 500kt milled, with an overall gold recovery of 90% achieved. Testing of the mineralisation in Big Hill has demonstrated that the mineralisation elsewhere in the deposit will behave in a similar manner to the Davis Open Cut mineralisation previously processed. A gold recovery of 90% is predicted. The tailings facility at SGM has a current approved work plan to allow an additional 3 metre wall lift. The requirements of the Big Hill project will likely require a two metre lift to accommodate the resultant tailings.
1.6 Mineral Resource and Mineral Reserve Estimates This technical report provides a Preliminary Economic Assessment (PEA) of the Big Hill Project of SGM. It outlines a possible open pit operation. This PEA is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically and technically to have the economic considerations applied to them that would enable them to be categorised as mineral reserves, and there is no certainty that the PEA will be realized. The Mineral resources reported and used in this Preliminary Assessment have been prepared according to the guidelines set out under the requirements of National Instrument 43-101. No Mineral Reserves have been estimated. Mineral resources that are not mineral Reserves do not have demonstrated economic viability. The mineral resource model used in this PEA was updated and is current as at 30th November 2012. This represents the first updated resource estimate for the Big Hill area since 1998 and includes additional RC drilling during the 2008 and 2012 periods. Drill spacing over the estimate area is largely 20m x 25m. In addition to the new RC drilling this estimate incorporates an updated geological interpretation by site staff, update resource estimation, review and update of historical void models, utilization of a pit shell at AUD$1400 gold price and an economic cut off for reporting of 0.44g/t. The November 2012 Resource estimate is presented below in Table 1-1.
Table 1-1 Big Hill Mineral Resources November 2012 estimate
Indicated Inferred
tonnes grade ounces tonnes grade ounces Surface (,000's) g/t Au (,000's) (,000's) g/t Au (,000's)
Big Hill 2,830 1.84 167 46 1.15 2
The mine life is anticipate to be in the order of 4 to 5 years and is anticipated to use a mining fleet consisting of between one and three 100t excavators and between eight to ten 90t trucks. Some drill and blast is anticipated however, a significant portion of the material is expected to be “free dig”. Existing underground workings in the area is expected to effectively “dewater” the pits although some perched water is anticipated. Mining operating costs have been benchmarked against the costs of operating mines of a similar size and nature. The restricted hours of operation (5 days per week / daylight hours only) is
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expected to attract a cost premium, this is yet to be quantified. However, the estimate used here is believed to be indicative and aligned with the intended targeted study accuracy. The cut-off grade applied (0.44g/t) in this instance is estimated based on administration and processing costs only. Sustaining capital for the mill has not been included to estimate the cut-off; this is partly due to the underutilisation of Mill. Current plans indicate that there is excess capacity in the mill and the plan assumes that the decision to continue operations is given. Therefore, while there continues to be excess capacity in the Mill, any grade above a cut-off grade based on variable costs only will contribute to any fixed operating or sustaining capital requirements. Open Pit Mining Capital requirements are mainly associated with impacted infrastructure and are estimated to be in the order of $22M. These are inclusive of further studies, test work and all required pre-start capital as at the time of writing this report. A quarterly mine schedule with associated cash flow estimated has been generated to outline anticipated cash flow drawdown commitments. Three separate cases were evaluated in order to determine a best case scenario for the Big Hill Project. A summary of the open pit mining physicals and preliminary economics estimated for each of these cases is shown in the table below:
Table 1-2 Summary of Mining Physicals and Economics
100m Buffer Description Unit 20Mt Case 10Mt Case Case Mill Feed t 3,470,546 2,277,200 2,317,587 Input Grade Au g/t 1.52 1.65 1.39 Mill Input Ounces Au oz 170,530 120,590 104,142 Mill Recovered Ounces AU oz 153,477 108,531 93,728
Mining Cost $ -73,371,063 -36,582,481 -42,424,778 Mill Cost $ -52,058,161 -34,158,006 -34,763,778 Rehandle Cost $ -20,788,996 -13,873,640 -25,335,771
Revenue Au $ 214,867,986 151,943,087 131,219,072 Undiscounted Cash flow $ 44,685,448 45,614,643 7,955,408 Discounted Cash flow (Inclusive of capital)* $ 36,748,767 39,647,998 6,419,168 IRR % 664 812 69 *Please note that a 10% discount rate has been applied to this project
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1.7 Conclusions and Recommendations The Big Hill mineral resource has been re-evaluated in terms of mine ability and profitability taking cognisance of constraints imposed by the locality of the resource. Several concepts were investigated with regards to foot print, rehabilitation and infrastructure without jeopardising optimum profitability. Three pit shells were selected for financial evaluation to determine the financial balance of optimum pit size and material handling. The selection criteria was determined by modelling numerous pit shells, financial evaluation and subsequently, identified by incremental step changes. The three pit shells identified were: a twenty million tonnes pit shell, a ten million tonnes pit shell and a hundred metre exclusion zone pit shell. The hundred metres exclusion zone pit shell was selected to test the financial viability with regards to requirements to obtain permission from neighbouring land owners. The base capital estimation included cost to complete the next level of study and high level estimation of capital required prior to execution. Operating cost used, is comparable to similar size operation for evaluation purposes and in line for this level of study. All other parameters used, are from historical established values. Pit designs were based on standard practices with batter angles designed at fifty degrees and single lane haul ramps. Considering a gold price of $1,400/oz the financial results for the three pits yield an Internal Rate of Return of 664% (20Mt case), 812% (10Mt case) and 69% (100m exclusion case). The discounted cash flows are $37M (20Mt case), $40M (10Mt case) and $6M (100m exclusion case). It was concluded that the 10Mt case demonstrates the most viable proposition and it is recommended based on the financial evaluation, to progress this project to a prefeasibility phase.
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2 INTRODUCTION This technical report has been prepared for Crocodile Gold Corporation (Crocodile); the beneficial owner of Stawell Gold Mines (SGM). Crocodile is listed on the Toronto Stock Exchange and acquired Stawell Gold Mines in May 2012. This technical report provides a Preliminary Economic Assessment (PEA) of the Big Hill Project of SGM. It outlines a possible open pit operation. This PEA is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorised as mineral reserves, and there is no certainty that the PEA will be realized. This report has been prepared in accordance with the requirements of the National Instrument 43- 101 (NI 43-101) standard and updates material changes to the Mineral Resource and Mineral Reserve position of the Big Hill project, Stawell Gold Mine as of December 2012. The Mineral Resources estimate for Big Hill is a summation of a number of individual estimates for various ore bodies or various geographically constrained areas. All of these estimates are contained within the Mining Lease MIN5260. Details of the locations and geographical constraints of the various ore body components as of December 2011 are given in Section 9. The Stawell gold deposit was discovered in the mid 1850’s during the Victorian gold rush which also saw the discovery and exploitation of significant deposits at Bendigo and Ballarat. Mining activity eventually ceased in the 1920’s and after a prolonged period of sporadic exploration mining operation recommenced in 1981. Mining operations and various levels of exploration and resource development activities have been continuous since 1981 and as such the project has significant past production and development history, which is discussed in this report and is also utilised during the compilation of the Mineral Resource and Mineral Reserve estimates. This report has been prepared by a number of the SGM site personnel with the assistance of an independent geological consultant. The report utilises information available within Company technical reports, published geological papers and internal Mineral Resource and Mineral Reserve documents completed by members of the SGM mine geological and mine engineering teams. Mark Van Leuven of Mining One Pty Ltd is a qualified person as defined by NI 43-101 and accepts overall responsibility for the preparation of all sections of this report. All information presented in this report was prepared in accordance with the requirements of NI 43-101, Standards of Disclosure for Mineral Projects and is in the format prescribed by that instrument. Mark Van Leuven visited the Big Hill Project at Stawell Gold Mines on 15 October 2012. During that visit, he inspected the site of the proposed Big Hill pits and the immediate surroundings. He also had discussions with Johan Booyse, the Mine Manager, and Mark Haydon, the Geology Manager at SGM.
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2.1 Certificate of Qualified Persons (Ni43-101) Mark Van Leuven BEng in Mining Engineering, MBA (Technology Management), FAusIMM(CP) I, Mark Van Leuven, BEng, Level 9, 50 Market Street, Melbourne, Victoriaa, Australia, do hereby certify that: a) I am employed as a Principal Mininng Engineer for Mining One conssultants of, Melbourne, Victoria, Australia, specializing in miniing engineering consulting, contracting for a wide variety of clients, and have been so employed since 2007. b) This certificate applies to the Reserve related sections of the technicaal report titled “Post Nov 26th Big Bill PEA V1_Final Model_SL” covering the preliminary economic assessment of the Big Hill open pit project at the Stawell Gold Project (“the Technical Report”), with a date of December 7th , 2012. c) I graduated with a Bachelor of Engineering in Mining Engineering degree from the South Australian Institute of Technology, now called the University of South Australia, Australia. I am a current Fellow of the Australasian Institute of Mining and Metallurgy (AusIMM). I have read the current definition of a “Qualified Person” (QP) as set out in NI 43-1101, and state by virtue of my education, membership in professional associations, and past relevant work experience, which spans over 27 years in the field of mine design, planning, reserves, technical and operaational management with multiple companies and projects, I fulfil the requirements to be a QP for the purposes of NI 43-101. d) I have conducted a personal site visit to the Stawell site on 15 October 2012, in conjunction with the issuance of this Report. e) I am responsible for Sections 15-16 and 21-22 of this Report. I have reviewed areas covered by these Sections. f) I am independent of the issuer, as defined by the test in section 1.5 of NI 43-101. g) I have had no prior involvement with the property that is the subject of this Technical Report. h) I have read NI 43-101 and Form 43-1101F1 and all of the Sections of the Technical Report for which I am responsible, and those sections have been prepared in compliance therewith. i) As of the date of this certificate, to the best of my knowledge, infoormation and belief, the sections referenced above contain all scientific and technical information that is required to be discloosed to make the Technical Report not misleading.
Effective Date: December 07, 2012 Signed:
Mark Van Leuven, BEng, FAusIMM(CP)
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Stuart Hutchin BSc, Applied Geology, MAIG, MAusIMM I, Stuart Hutchin, BSc, Level 9, 50 Market Street, Melbourne, Victoria, Australia, do hereby certify that: a) I am employed as a Geology Manager for Mining One consultants of, Melbourne, Victoria, Australia, specializing in geological consulting and exploration contracting for a wide variety of clients, and have been so employed since 2011. b) This certificate applies to the Resource related sections of the technical report titled “Post Nov 26th Big Bill PEA V1_Final Model_SL” covering the preliminary economic assessment of the Big Hill open pit project at the Stawell Gold Project (“the Technical Report”), with a date of December 7th , 2012. c) I graduated with a Bachelor of Science degree in Applied Geology from the University of South Australia, Australia. I am a current member of the Australian Institute of Geoscientists (AIG) and the Australasian Institute of Mining and Metallurgy (AusIMM). I have read the current definition of a “Qualified Person” (“QP”) as set out in NI 43-101, and state by virtue of my education, membership in professional associations, and past relevant work experience, which spans over 17 years in the field of geology and mineral exploration with multiple companies and projects, I fulfil the requirements to be a QP for the purposes of NI 43-101. d) I have conducted a personal site visit to the Stawell site on December 4th and 5th, 2012, in conjunction with the issuance of this Report. e) I am responsible for Section 14 of this Report (Mineral Resource Estimates). I have reviewed areas covered by this Section. f) I am independent of the issuer, as defined by the test in section 1.5 of NI 43-101. g) I have had no prior involvement with the property that is the subject of this Technical Report. h) I have read NI 43-101 and Form 43-101F1 and all of the Sections of the Technical Report for which I am responsible, and those sections have been prepared in compliance therewith. i) As of the date of this certificate, to the best of my knowledge, information and belief, the sections referenced above contain all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.
Effective Date: December 07, 2012 Signed:
Stuart Hutchin (BSc Geology, MAIG, MAusIMM)
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3 RELIANCE ON OTHER EXPERTS The Author of this report is a Qualified Person, and has relied on various datasets, report, and financial models that were provided by SGM to support the compilation of this report. The author has relied on other experts from both SGM and the independent consulting group Mining One to compile this report. The SGM experts are listed in Table 3-1. It is the view of the authors that the data collection, storage, and analysis methods utilised in estimating and compiling Mineral Resource estimates at SGM are of sufficient quality to ensure the information is reliable and suitable for compilation of this Technical Report. The principal Author is not aware of any critical data that has been omitted so as to be detrimental to the objectives of this report. There was sufficient data provided to enable credible interpretations to be made in respect of the data. The principal author believes that no information that might influence the conclusion of the present report has been withheld from the study.
Table 3-1 Site Experts who contributed to the Technical Report
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3.1 External Consultant Reports and Reviews No experts have been used by Mining One to assess the potential of the area except as reported by other companies or as published in the research literature. Where there is reliance on other experts they are referenced in Section 27. SGM initiated a full external review as part of the Resource model finalisation. This work was undertaken by Quantitative Group’s Principal Consultant, Mike Stewart. Mike Stewart spent 4 days on site from the 19th November 2012 during which time he reviewed and validated the Big Hill resource model and completed an independent report “SGM, Review of Big Hill Resource Estimate, CRG21204”. Quantitative Group have stated in this report “After thorough review of inputs, method, implementation and outputs of resource estimation at Big Hill, QG are of the opinion that the 2012 resource estimate is a fair and reasonable representation of the in-situ resource, and that this constitutes a sound basis for mine planning”
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4 PROPERTY DESCRIPTION AND LOCATION The SGM is located in the State of Victoria, Australia and is 250 kilometres northwest of the city of Melbourne and two kilometres from the township of Stawell (Figure 4-1). Stawell is a rural township of approximately 6,500 people and is within the Northern Grampians Shire. The mine’s principal approval is its Mining Lease (MIN5260) issued by the Victorian State Government under the Mineral Resources Sustainable Development Act 1990. This approval was first issued on the 31st May 1985 as MIN1219 and has been amended on at least six occasions since as a result of approved Work Plan variations. The current mining licence approval is active until 2020. The MIN5260 lease (centroid coordinates of 142.80° E and 37.06° S, GDA94) encompasses both the Magdala and Wonga mines and is located both under and around the Township of Stawell with an area of 1000.58 Ha. The mining lease is comprised of private and crown land including designated crown land reserves. Designated crown land reserves require particular consideration in accordance with the Mineral Resources Development Act 1990 and National Parks (Box Ironbark and Other Parks) Act 2002. Within MIN5260 there is an A$2.00 per ounce gold royalty payable to Mineral Ventures of Australia (MVA). This royalty agreement came into place in February of 2004 and is in place until the earlier of 15 year of production or 2.5 million ounces of gold produced. Furthermore this royalty agreement extends to Victorian tenements held by Leviathan Resources Ltd (now a wholly-owned subsidiary of Crocodile Gold Corporation) which included MIN5260. Rehabilitation for the project is ongoing and the Corporation’s wholly owned subsidiary, SGM entered into a cooperative research project with both Curtin and Melbourne Universities in the year 2000 to conduct rehabilitation trials to prepare the rehabilitation programme for eventual closure of the operations. These trials were extensive and undertaken over many years with works ongoing with O’Kane Consultancy regarding Tailings Storage Facility capping design. Other than the rehabilitation bond the project is not subject to any other environmental liabilities. Stawell Gold Mines principal approval, MIN5260, is the applicable “right to mine” title over this land and is current to 2020. Attached to this title are a series of licence conditions that must be met and are the controlling conditions upon which an annual Work Plan and Work Plan variations are filed with the regulatory authority. Stawell Gold Mines is operating under a Work Plan submitted as required under Section 3 of the General Licence Conditions of MIN5260. A key component of this Work Plan is an Environmental Management Plan (EMP) the most recent of which was approved by the Department of Primary Industries Victoria in April 2011 in conjunction with the 2011 Wonga Pit Cut Back Work Plan Variation. A requirement of the General Licence Conditions of Mining Licence is to maintain the EMP and will be updated accordingly as work plan variations are presented. Stawell Gold Mines management reviews applicable regulatory requirements on an ongoing basis. Stawell Gold Mines currently holds and maintains all statutory approvals required to conduct its mining operations.
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Figure 4-1 Map Highlighting Location of SGM
4.1 Property Description Stawell Gold Mines principal approval is its Mining Lease MIN5260, (Figure 4-2) issued by the Victorian State Government under the Mineral Resources (Sustainable Development) Act 1990. This approval was first issued on the 31st May 1985 as ML1219 and has been amended on at least six occasions as a result of approved Work Plan variations. The current Mining Licence approval is active until 2020. Additionally Stawell Gold Mines applied for MIN5520 late 2009 with the licence being granted in March 2010. No mining activity is currently underway on this licence. MIN5520 was applied to cover the northern extension of the Magdala lodes at depth with further exploration required to scope the potential of this area. As such a work plan to explore only has been applied for and granted. The MIN5520 mining Licence is active until 2030. Big Hill locally refers to the area approximately 1km north of the current underground mine and milling operations at SGM. With respect to this PEA Big Hill collectively includes the Davis (South Pit) and Big Hill (North Pit) areas see Figure 4-3.
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Figure 4-2: Location Map of the MIN5260 and MIN5520 Lease. The Grid is Latitude and Longitude (GDA94)
Figure 4-3: Aerial View of the MIN52600 and MIN5520 Lease indicating Underground and Pit Locations
4.2 Legislation and Permits Stawell Gold Mines principal approval, ML5260, is the applicable “right to mine” title over this land and is current to 2020. Attached to this title are a series of licence conditionns that must be met and are the controlling conditions upon which an annual Work Plan and all associated Work Plan variations are filed with the regulatory authority.
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Apart from the primary mining legislation the Mineral Resources (Sustainable Development) Act 1990 (Vic), operations on MIN5260 are subject to the additional following legislation and regulations for which all appropriate permits and approvals have been obtained. Acts: ¾ Extractive Industry Development Act 1995 (Vic) ¾ Environment Protection Act 1970 ¾ Mines Act 1958 ¾ Planning and Environment Act 1987 ¾ Environmental Protection and Biodiversity Conservation Act 1999 ¾ National Environment Protection Council (Vic) Acts 1995 ¾ Flora and Fauna Guarantee Act 1988 ¾ Catchment and Land Protection Act 1994 ¾ Archaeological and Aboriginal Relics Preservation Act 1972 ¾ Heritage Act 1995 ¾ Forest Act 1958 ¾ Dangerous Goods Act 1985 ¾ Mines Safety and Inspection Act 1994 ¾ Drugs, Poisons and Controlled Substances Act 1981 ¾ Health Act 1958 ¾ Water Act 1989 ¾ Crown Land (Reserves) Act 1978 ¾ Radiation Act 2005 ¾ Sustainability Victoria Act 2005 ¾ Country Fire Authority Act 1958 ¾ Conservation, Forests and Lands Act 1987 ¾ Wildlife Act 1975 Regulations: ¾ Dangerous Good (Explosives) Regulations 2011 ¾ Dangerous Good (Storage and Handling) Regulations 2000 ¾ Dangerous Goods (HCDG) Regulations 2005 ¾ Mines safety and Inspection Regulations 1995 ¾ Forest Fire Regulations 1992 ¾ Biodiversity Conservation Regulation 2000 ¾ Drugs, Poisons and Controlled Substances 9Commonwealth Standard) Regulations 2001 ¾ Mineral Resources (Infringements) Regulations 1991 ¾ Environmental Protection (Vehicle Emissions) Regulations 2003
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Stawell Gold Mines is operating under a Work Plan submitted as required under Section 3 of the General Licence Conditions of Mining Licence (MIN5260). A key component of the Work Plan is an Environmental Management Plan (EMP), the most recent of which was approved by the Department of Primary Industries Victoria in April 2011. A requirement of the General Licence Conditions of Mining Licence is to maintain the EMP. Regular reviews of legislation and regulation requirements have been completed and SGM maintains all required statutory approvals to continue with mining operations. An environmental bond for the project of AUD$4,803,000 is lodged with the Department of Primary Industries Victoria. Rehabilitation for the project is ongoing. SGM entered into a cooperative research project with both Curtin and Melbourne Universities in the year 2000 to conduct rehabilitation trials to prepare the rehabilitation programme for eventual closure of the operations. These trials were extensive and undertaken over many years with works ongoing with O’Kane Consultancy regarding Tailings Storage Facility capping design. Other than the rehabilitation bond the project is not subject to any other environmental liabilities.
4.3 SGM Local Survey Grid Reference All survey data on MIN5260 is collected and stored using modified AMG co-ordinates, based on Australian Map Grid AGD 66 (Zone 54). The convention is to drop the first digit of the northing, so 5896000N becomes 896000N. The Easting value is unchanged. The mine RL is calculated as -300m RL AHD (Australian Height Datum) and displayed as a negative number below surface. The RL origin is at 303.60 AHD, measured at a trigonometric station located adjacent to the mine on Big Hill, Stawell. The principal local grid in use within the Mine Lease is the SGM Grid (also referred to as the “45 degree grid”) as shown in Figure 4-4. This grid is orientated 45 degrees west of AMG north and has its origin at 5890137.479N and 659498.820E. It is convention to divide the northing by 20 and refer to this as the section northing line, i.e. northing 6200 becomes the 310 section line. Additional local grids are used as required for presentation of geological information as required.
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Figure 4-4: Local Mine Grid at SGM
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5 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE & PHYSIOGRAPHY Stawell Gold Mines is easily accessible from Melbourne via the Western Highway. Access closer to the mine site is provided through a network of sealed bitumen government roads. Roads within the mine site are unsealed and regularly maintained. The main Melbourne to Adelaide rail line passes through the town of Stawell, which is also serviced by a local sealed airfield. The town of Stawell is located within the southern part of the Wimmera where the climate is described as semi-arid, allowing for exploration and mining activities all year round. Stawell weather for the past 20years has recorded an annual daily average temperature of 19.7°C. A daily high mean maximum of 28.1oC in February and low daily maximum mean of 12oC occurring in July. Mean annual rainfall is 562.1 millimetres with 80.4 days per year on average recorded as having rain. Stawell Gold Mines has been in operation for over 25 years, developing a highly experienced workforce. Many contractors, also having a long association with the mine, are available in the township of Stawell and surrounding regions. Due to the mines close location to the town of Stawell many facilities are available. Within the township area is a police station, hospital, schools and shops. Mains Electricity and water is also accessible. Stawell Gold Mines facilities are extensive and representative of a modern gold mining operation. On the surface facilities include the gold processing plant, offices, core shed, laboratory and workshops. Larger infrastructure onsite includes a tailings dam, covering 96 Ha and receiving 100% of gold tailings from the processing plant. Three freshwater dams occur throughout the mine lease. The mine purchases electric power under a two year contract from Origin Energy Australia that expires in December 2012. Water supply is from harvested rainfall runoff, mine dewatering, recycling of process water from the tailings facility, and by way of a 1ML/day raw water right entitlement and urban customer access to potable supply from Lake Bellfield located in the Grampians Mountains. The capacity of the site water storages is approximately 690ML. The Lake Bellfield water is potable and is preferentially used in the processing operations as it improves gold recovery. The area surrounding Stawell is composed of flat to gently undulating farmland with the Grampians Mountain range and National Park 20 kilometres to the southwest. Close to the centre of Stawell is Big Hill, the town’s highest point at a height of 303.6 metres above mean sea level. Stawell Gold mine is situated on the southern slope of Big Hill. Parts of the area adjacent to the mine are covered by Iron bark forest.
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6 HISTORY Stawell is a historic goldfield that produced 2.7 million ounces of gold between 1853 and 1926 from both alluvial and hard rock sources. There was little mining activity in the Stawell area from 1926 to 11th March 1976 when Western Mining Company (WMC) Resources Ltd was granted an exploration licence over the Stawell Goldfield. In 1981 SGM was reopened by the WMC/Central Norseman Gold joint venture with commencement of the Magdala decline. By 1984 the operation had expanded with the construction of a processing facility and subsequent commencement of an open cut operation at the Wonga mine (2 kilometres south of Magdala). A number of historical tailing dumps were retreated during this period. Towards the end of mining of the Wonga open cut (1987) the Davis open cut operation was commenced. The Davis open cut exploited the oxide material on up dip projection of the Magdala deposit. The Wonga Open Cut operated from 1984 to 1987 and produced 778,847 tonnes recovering 69,159 ounces of gold. The Davis Open Cut operated from 1987 to 1989 and produced 154,525 tonnes for 8,992 recovered ounces of gold. In December 1992 the operation was acquired in a 50/50 joint venture by Mining Project Investors (MPI) Pty Ltd and Pittston Mineral Ventures (Pittston). At this stage the Magdala decline was at 410m RL, while the Wonga decline was at 180-200m RL. With the acquisition there was a clear direction to increase expenditure on resource definition drilling and near mine exploration. The joint venture continued until 2004 during which time there was a record of continued exploration success with discovery of additional mineralized deposits that were subsequently mined. A study to establish an open pit mining operation on the Big Hill section of the deposit was completed and a detailed Environmental Effects Statement (EES) prepared. The EES was submitted to the Victorian State Government in 1999 seeking approval to commence a mining operation within the Big Hill area. In November 2000 the proposal was rejected by the Minister for Planning and the project was not able to proceed due to ministerial concern in the areas of social impact, heritage, unfilled mining void, habitat impact and a tenuous link between the proposed project of the time and continuation of underground mining operations. Stawell Gold Mines believes there is strong potential to re-examine the project with all areas of ministerial concern addressed in present project consideration. In February 2004 MPI acquired the Pittston 50% share of the project. Exploration continued in the Golden Gift area during 2004 with the commencement of the Golden Gift South surface exploration programme. In November 2004 a de-merger of the MPI gold business came into effect, and Leviathan Resources Ltd was floated in December 2004. The resource drilling into the Golden Gift initially identified seven areas of mineralisation offset from each other due to late faulting. Conversion of these areas of mineralisation into ore blocks wasn’t universal but was successful in a majority of cases. The further drilling of the fault blocks also identified other mineralised surfaces previously unknown due to the faulted nature of the Golden Gift. From the increased geological understanding of the Golden Gift deposit, it was clear in the mine planning process that two declines were required, the GG5 and GG3 declines, to access the ore zones for continuity of supply. In January 2007 Perseverance acquired Leviathan Resources Ltd. Perseverance was acquired by Northgate on the 18th of February 2008. Northgate was acquired by AuRico in October, 2011. Crocodile Gold completed their acquisition of Stawell Gold Mines from AuRico on May 4th, 2012. Production of both the Magdala and Golden Gift Ore bodies has remained continuous with workings now reached to a depth of -1200mRL and -1600mRL respectively.
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6.1 Historical and Modern Production Stawell is a historic goldfield having produced approximately 2.7Moz of gold between 1853 and 1926 from both alluvial and hard rock sources. Since the commencement of mining in the modern period, 1984, until December 2011, over 2.2 million ounces have been prooduced from the Stawell ore body. A summary of annual gold production is shown in Figure 6-2. Ore treated and treated head grade are shown in Figure 6-1.
Figure 6-1: Annual Gold Production since 1984
Figure 6-2: Annual Processing Plant Production since 1984
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7 GEOLOGICAL SETTING AND MINERALISATION For details on the geological setting of the Stawell Project the reader is referred to the previous technical report entitled “Technical Report on Stawell Gold Mine, Victoriia Australia (28 March 2008)”.
7.1 Regional Geology
The Stawell Goldfield is located in the western Stawell Zone of the Lachlan Fold Belt Figure 7-1. The Stawell Zone is a belt of predominantly deformed meta-sedimentary rocks representing the lower parts of the Cambro-Ordovician Lachlan Foold Belt stratigraphy bound to thhe west by the Moyston Fault and to the east by the Coongee Break (Vandenberg et al. 2000).
Figure 7-1: Image showing Lachlan Fold Belt, Locating Stawell on thee Western Boundary
Interpretations from the Victorian Geological Survey present a thin skinned tectonics model where the Moyston Fault is an east dipping basal detachment which has juxtaposed higher metamorphic grade rocks of the Stawell Zone against lower grade Cambrian rocks of tthe Delamarian Glenelg Zone. The west dipping Stawell Fault, Coongee Break and other paralllel west dipping faults
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represent back thrusts from the Moyston Fault. These back thrusts have progressively emplaced deeper stratigraphy against shallower stratigraphy with a generally west over east sense. An apparent anomaly in this sequence is the presence of deeper magnetic stratigraphy in the Stawell- Wildwood corridor. Vandenberg et al. 2002 interprets that the Pleasant Creek Fault, to the west of the Stawell Fault, actually dips east and has an east over west sense - similar to the Moyston Fault. The Stawell-Wildwood corridor therefore represents a significant structural high in an up-thrown block of deeper stratigraphy between the Coongee Break and Pleasant Creek Fault.
7.2 Local Geology There are three separate ore bodies defined at Stawell; the Magdala, Golden Gift and Wonga. All have differing characteristics but the same local geology is relevant to the genesis of them all.
7.2.1 Stratigraphy at Stawell The stratigraphy at Stawell is divided into three principal units: Magdala Basalt; Albion Formation; Leviathan Formation, see Figure 7-2 (Squire and Wilson, 2005). Intruded into this sequence are the Stawell Granite and a number of felsic and mafic intrusions. Squire and Wilson (2005) interpret that the rock unit previously termed the Magdala Volcanogenics (Watchorn and Wilson, 1989) is an alteration facies that locally occurs adjacent to the basalt.
7.2.2 Magdala Basalt The Cambrian Magdala Basalt is composed of subaqueous low-K tholeiitic lavas that exhibit an aphyric to sparsely plagioclase-phyric texture (Watchorn and Wilson, 1989; Squire and Wilson, in review). The Basalt body comprises flows ranging from 0.5m to 50m thick, pillows basalts with pillows ranging in size from 0.1 to 2 metres in size (Watchorn and Wilson, 1989) and monomictic basalt breccias varying proportions (Pritchard, 2001; Squire and Wilson, 2005). The basalts are interpreted to form part of the Victorian Cambrian greenstone sequences, which are the oldest known rocks in the Palaeozoic Lachlan Orogen, and have inferred ages of 516-514Ma (Squire and Wilson, 2005). The basalts near Stawell occur as dome-like units in the footwall and hangingwall of major faults (Miller and Wilson, 2002). The Magdala Basalt which closely resemble typical back-arc basin basalts (Kaufman, 2003; Crawford, 1988) has been interpreted to represent the medial to distal facies on the flank of a large basalt edifice upward of 500m thick (Squire and Wilson, 2005) and has similar magmatic affinities to the known basalt bodies north of Stawell, Wildwood and Kewell Basalts (Kaufman, 2003; Jupp, 2003).
7.2.3 Leviathan and Albion Formations Overlying the Magdala Basalt is a 200-300m thick sequence of non-fossiliferous turbidites (Squire and Wilson, 2005). The turbidite sequence has been subdivided into two different lithologies: the Albion Formation; and the Leviathan Formation. The differences between the two lithologies was first recognised but not explored by Gane (1998). He recognised the sediments on the western side of the Magdala Basalt graded from predominantly mud-rich to more sand-rich away from the basalt. The Albion Formation is the lowest clastic sequence to the west of the Magdala Basalt. The unit varies in thickness with the top of the unit defined by a 20-100m sequence of black mudstone. Within the Albion Formation there are a number of facies which along with black mudstone include calcareous sandstone, siliceous siltstone and sulphidic black mudstone. Squire and Wilson (2005) suggested that the sediments were deposited predominantly due to suspension settling in a sediment-starved sedimentary basin. There were short-lived periods of oxygen-rich conditions shortly after volcanism recognised by the presence of siliceous siltstone but the dominance of black
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mudstone within the Albion formation indicates the basin of deposition was predominantly anoxic (Squire and Wilson, 2005). The provenance for the Albion Formation sediments has been identified from detrital compositions to be a low-grade metamorphic terrane (Cas, 1983).
Figure 7-2: Mine Geology Cross-section highlighting Architecture of the Magdala and Golden Gift Ore Bodies
The Leviathan Formation overlies the Albion Formation and is dominated by fine-to medium-grained quartz-rich sandstones (Squire and Wilson, 2005). The contact between the two formations is gradational and conformable. Although the Leviathan Formation was deposited in a higher-energy environmeent than the underlying Albion Formation, the detrital compositions indicate little change in the provenance between the two formations (Squire and Wilson, 2005).
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The Leviathan and Albion Formations are not segregated by the mine or exploration geologists at SGM and are referred to by the local name of ‘mineschist’.
7.2.4 Magdala Facies The Magdala Facies, termed Magdala Volcanogenics at SGM, distinguished by its dark green colour, is a result of intense chloritic alteration of mudstone and or shales located at the base of the Albion formation and immediately above the Magdala basalt. The Magdala Facies is the primary and most important host rock for the sulphide replacement style of gold mineralisation at Stawell. Major mineralisation sulphides include arsenopyrite, pyrite and pyrrhotite, the latter two commonly occurring along cleavage planes and concentrated within shear zones (Robinson, 2005).
7.2.5 Felsic Intrusions Quartz±feldspar-phyric felsic intrusions crosscut the turbidite sequence. The quartz±feldspar-phyric felsic intrusions vary in thickness from 50cm to 12m wide and show chilled margins. They are predominantly composed of quartz and plagioclase with phenocrysts up to 3mm in size. The feldspar phenocrysts have euhedral shapes and display multiple twinning while the quartz phenocrysts had a cloud-like appearance and were rimmed with fibrous quartz (Gedge, 1997). The groundmass is composed of ~80% quartz in anhedral grains and display undulose extinction (Gedge, 1997). The felsic intrusions tend to follow northwest-trending shear zones (Wilson et al., 1992) and the emplacement of the quartz±feldspar-phyric felsic intrusions post-dates the main Magdala mineralisation event. The intrusions have been dated at 413±3Ma (Arne et al., 1998).
7.2.6 Stawell Granite The Stawell Granite was emplaced during the early Devonian, 401±4Ma (Arne et al., 1998), and is located about 2km south of the Magdala deposit (Xu et al., 1994) and adjacent to the Wonga deposit. The pluton is approximately 20km wide and 13km long and intrudes the sandstone and shale units of the turbidite sequence. The pluton is an asymmetrically zoned, medium grained intrusion with contains diorites, granodiorites and magnetite-rich felsic granites (Wilson et al., 1992). There is a 0.5km to 1.0km contact aureole surrounding the Stawell Granite (Xu et al., 1994).
7.2.7 Lamprophyric Intrusions
Lamprophyres intrude all the above lithologies and are hosted in D4 shears (Gedge, 1997). The lamprophyre intrusions can range in thickness from 1.0cm to 3m (Wilson et al., 1992). The dykes vary in colour from dark grey to chocolate brown and display conchoidal fracture patterns (Gedge, 1997). The compositions of the intrusions vary from monchiquite (olivine bearing) to fourchite (augite and no olivine) (Wilson et al., 1992). The lamprophyre dykes typical mineralogy is composed of Na- rich plagioclase (albite), clinopyroxenes, biotite, sulphides, ilmenite and Ti-rich magnetite (Gedge, 1997).
7.2.8 Structural History at Stawell At least seven deformation events have been recognised at Stawell (Wilson et al., 1992). These deformation events can be broadly split into two categories; early, ductile deformation (D1 to D4), and late brittle deformation (D4 and later). A description of both categories of deformation and related structures are given below.
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7.2.9 Early, Ductile Deformation The ductile deformation events all occurred under a northeast-southwest shortening direction (see
Figure 7-3) (Miller and Wilson, 2002). The earliest ductile event recognised, D1, is thought to be thrust-related with early shearing along detachment surfaces that produced a fabric, S1, parallel to bedding (Wilson et al., 1992). This event has been suggested to occur at about 510-504Ma (Squire and Wilson, 2005). The second ductile deformation event, D2, produced the most dominant ductile fabrics at Stawell and occurred at about 496-494Ma (Squire, 2004). D2 refolded F1 closures and fabric into tight F2 folds causing S1 to appear predominantly parallel to S2 (Miller and Wilson, 2002). The mesoscopic F2 folds trend to the northwest and are generally asymmetric with hinges varying in size from centimetres to tens of metres (Wilson et al., 1992). Peak metamorphism at Stawell is considered pre- to syn- D2 and reached mid-greenschist grades (Miller and Wilson, 2002, Wilson et al., 1992). Overprinting both of the earlier fabrics at Stawell is an asymmetric differentiated crenulation cleavage, S3. D3 is a result of developing west over east shearing and folding (Watchorn and Wilson, 1989). This event is date at approximately 494-492Ma (Squire and Wilson, 2005). The crenulation foliation is generally sub-horizontal and is associated with cleavage-parallel veins. There is evidence of a fourth ductile fabric at Stawell which is associated with D4 and is interpreted as a ductile-brittle event (Miller and Wilson, 2002).
7.2.10 Late, Brittle Deformation
Superimposed on the ductile fabrics (D1-D4) are a number of brittle structures (refer Figure 7-3 and Figure 7-4). The geometry and style of the brittle deformation is strongly dependent on the pre- existing geometry of the basalt (Miller and Wilson, 2002). The initiation of the brittle deformation occurred during late D4 when the shortening direction changed from a northeast-southwest to an east-west orientation (Watchorn and Wilson, 1989; Miller and Wilson, 2002). The early D4 shear zones have a northwest trend and dip to the southwest between 20˚ to 60˚ with a reverse sense of movement (Watchorn and Wilson, 1989). There was a change in the regional stress field within the Lachlan Orogen during the Late Silurian which is expressed at Stawell as a change from east-west shortening to sinistral wrenching along pre-existing faults (Miller and Wilson, 2004a). Reactivation of the D4 shears by sinistral wrenching is termed D5 (Mapani and Wilson, 1994). The sinistral wrenching was followed by another change in the regional stress field with the shortening direction changing to northwest-southeast. A set of major faults oblique to the earlier structural trends associated with this change in shortening direction is termed ‘early South Fault’ structures (Miller and Wilson, 2004a). The last major deformation event was associated with a final change to a northeast-southwest shortening. Faults associated with this event dip northwest and have a dip-slip sense of movement (Miller and Wilson, 2004a).
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Figure 7-3: D1 to D5 Ductile and Brittle Evolution of the Stawell System Steronets.
Stereonets represent hangingwall transport direction calculated at pole to fauult with the circle centre of each arrow representing a single fault pole (Miller & Wilson 2004a). These transport directions are the inferred maximum resolved shear stress along a fault for an applied stresss tensor. A change in the hangingwall transport direction for similarly oriented faults represents a change in stress tensor. From Milleer et al. 2006.
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Figure 7-4: Evolution of the Stawell system from 420 to 380 Ma (modified from Miller & Wilson 2004a).
Stereonets represent hangingwall transport direction calculated at pole to fauult (Miller & Wilson 2004a). Map symbol are the same as those in Figure 7 5. From Miller et al. 20006.
7.2.11 Stawell Mine Geological Architecture The dominant feature at Stawell is the 1..2km wide doubly plunging northhwest striking Magdala Basalt dome. The Magdala Basalt is made up of a series of basalt nosess, interrupted to be flow
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sheets (Squire and Wilson, 2005), which dip to the southwest and plunge to the northwest. Areas of sedimentation are present between the basalt noses (interpreted flow sheets) and are locally termed ‘waterloos’. The Magdala Basalt has been drilled and identified to a depth of 1.7km and interpreted from existing drill information and from geophysical modelling to extend along strike at least 5km. Surrounding the basalt dome is the turbidite sequences of the Albion and Leviathan Formations (mineschist) which young to the west. The contact between the mineschist and Magdala Basalt on the western side is marked by the alteration package of the Magdala Volcanogenics. The Magdala Volcanogenics is weakly developed on the eastern surface of the Magdala Basalt. This Magdala geology has been faulted and offset by later brittle deformation; the most notable of these offsets is the South Fault which has a northeast over southwest sense of transport (refer Figure 7-2 and Figure 7-5). Above the South Fault is the Magdala ore body which contains limited offsets due to late faulting. The Basalt surface in the Magdala ore body dips to the west and strikes towards 340˚. Beneath the South Fault is the Golden Gift ore body which is heavily offset by late faulting creating isolated ore blocks. Unlike the Magdala ore body the basalt in the Golden Gift dips to the east and strikes towards 315˚. The late faulting as well as creating isolated ore blocks also complicates the ore geometry within the each block. To the south of the Magdala Basalt is the Stawell Granite, which structurally is situated below the South Fault (Figure 7-6). Located close to an embayment in the Stawell Granite are a series of brittle structures. One of the structures termed the Hangingwall Structure strikes towards 350˚ and dips between 25˚ and 50˚ towards the east, and the other structural set, termed Link Structures, generally trend toward 240˚ and dip between 40˚ and 70˚ to the southeast (Xu et al., 1994). Cross cutting these late brittle structures are a series of late felsic intrusive. This fault system hosts the Wonga ore body.
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Figure 7-5: Plan View Geological Interpretation of the Stawell Structural and Stratigraphic Architecture at 1000mRL.
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Big Hill
Figure 7-6: SGM Longitudinal Projection showing the Location of the Mineralised ore block. The geological and spatial relationship between the Magdala, Golden Gift and Wonga deposits can clearly be seen.
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7.3 Mineralisation There are three different ore bodies at Stawell, being Magdala, Golden Gift and Wonga. Each of the differing ore and mineralisation types are summarised below. Both the Magddala and Golden Gift ore types are hosted within the Magdala Volcaanogenics. Within the Magdala ddeposit there are three main ore types; Central Lode, Basalt Contacct Lodes, and Magdala Stockwork Lodes. These are the lodes that are present in the Big Hill area. Central Lode mineralisation was a significant production source from Magdala early in the mine history. It is a quartz-rich shear lode ranging from 0.5m to 10m in width and generally dips 55 - 65° to the west with a total strike length of 4km and a down dip extend of one kilometre (Figure 7-7). The overall structure is mineralised shoots vary from 20m to 30m in strike up tto 200 – 350 metres in strike. Free gold in the quartz is associated with pyrite, arsenopyrite and recrystallised pyrrhotite. The average mined grade for Central Lode is 4-7g/t Au.
Figure 7-7: Figure Example of Central Lode Mineralisation
Basalt Conntact Lodes are located parallel to the Magdala Basalt and in ‘waterloo’ or re-entrant positions. They are typically 2m wide and are represented by arrays of quartz sulphide tension veins immediately adjacent to the volcanogenic Basalt contacts (Figure 7-8). Sulphides include pyrrhotite, arsenopyrite and pyrite and occur as alteration selvages on tension vein margins. The main alteration mineral is stilpnomelane, resulting in its dark colour. The mineralissation is isolated to the Magdala Volcanogenic package with none present in the adjacent Magdala Basalt. Ore shoot lengths range between 50m and 450m. Thhe average mined grade for Basalt Contact Lodes is 4 – 9g/t Au.
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Figure 7-8: Example of Basalt Contact Mineralisation
The Magdala Stockwork Lodes are situated above major basalt noses and can be described as a hybrid between Central and Basalt Contact Lodes. They consist of large quartz tension vein arrays with arsenopyrite and pyrrhotite dominant sulphide mineralisation. The strike extent is limited to 40m to 50m and limited vertically to between 30m and 50m. The average mined grade for Magdala Stockwork Lodes is 4 – 7g/t Au. Unlike the Magdala deposit there is only one identifiable ore type in the Golden Gift and this is termed the Golden Gift Stockworks. Though there is only one discernible ore type in the Golden Gift, the Golden Gift Stockworks contain a spectrum of all Magdala styles. Typical widths range from 8- 12m up to 30m and the strike extents of shoots range between 150m and 400m. Areas of highest grades and largest widths are situated above major basalt noses which are present in most ore bodies. Quartz content is generally below 25%. Mineralisation includes abundant recrystallised pyrrhotite and coarse grained arsenopyrite, pyrite and visible gold. The average mined grade is 4 – 10g/t Au. The Wonga deposit is hosted within the locally termed Wonga Schist that is part of the Leviathan Formation along two main fault systems. The Wonga Schist has undergone contact metamorphism during the emplacement of the Stawell Granite and undergone three ductile deformation events similar to other areas of the Stawell region. The two fault systems controlling the mineralisation are the Hangingwall Structure, which strikes towards 350° and dips between 25° and 50° towards the east and the Link Structures which generally trend toward 240° and dip between 40° and 70° to the southeast. The mineralisation is represented by arsenopyrite disseminations in quartz veins within these structures. The main ore minerals present are anhedral fine grained pyrrhotite and arsenopyrite. The higher grade ore zones often show andalusite and sericite alteration with rutile and ilmenite associations. Production grades from 4 – 6g/t Au were common for the Wonga ore.
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8 DEPOSIT TYPES For details on the deposit types and mineralisation of the Stawell deposits the reader is referred to the previous technical report entitled “Technical Report for Stawell Gold Mine, Victoria Australia (28 March 2008)”.
8.1 General Deposit Types and Mineralisation Victorian mineralisation episodes have been dated to occur during the Devonian and Silurian periods with no gold mineralisation occurring prior to 440Ma (Miller and Wilson, 2002). A description of the mineralisation episodes in Western Victoria is described below. The largest and most significant mineralisation event in the western Lachlan Orogen occurred at ca. 440Ma (Foster et al., 1998). It occurred contemporaneous in the Stawell and in the Bendigo-Ballarat zones with the mineralisation occurring during late D4 deformation in the Stawell zone and occurring during late D1 deformation in the Bendigo-Ballarat zone. The mineralisation is hosted in D4 brittle structures associated with east-over-west movement at Stawell while in the Bendigo-Ballarat zone the mineralisation occurs in saddle reefs in the hinges of D1 folds and in reverse faults created via D1 fold lock-up (Miller and Wilson, 2002; Schaubs and Wilson, 2002). This mineralisation event produced the largest endowments of gold within the western Lachlan Orogen (Miller and Wilson, 2002). The next episode of gold mineralisation occurred at about 426-420Ma (Foster et al., 1998) and is associated with fault reactivation throughout western Victoria (Miller and Wilson, 2002). This episode of gold mineralisation produced significantly smaller endowments than the 440 Ma event (Miller and
Wilson, 2002). The late Silurian mineralisation is associated with the D5 sinistral wrenching at Stawell and has been recognised at the Percydale fields in the Stawell zone and at Tarnagulla in the Bendigo-Ballarat zone (Miller and Wilson, 2002). The final episode of mineralisation recognised in western Victoria is the Wonga mineralisation at Stawell (Miller and Wilson, 2004a). The mineralisation at Wonga over prints the quartz- and felsic- rich intrusions and is overprinted by the Stawell Granite contact areole. Watchorn and Wilson (1989) suggested that this mineralisation is temporally and spatially related to the granites emplacement. Miller and Wilson (2004b) advocate the mineralisation event at Wonga formed at ca. 400Ma (Foster et al., 1998). The Wonga mineralisation occurred during a late-stage magmatic event within a long- lived orogenic system at shallow crustal levels (Miller and Wilson, 2004b). This mineralisation occurred in a series of brittle structures dependent on pre-existing weakness which are related to a fluid over-pressure event after the lockup of major structures (Miller et al., 2004). There are three different ore bodies at Stawell; the Magdala, Golden Gift and Wonga. Of the three only the Magdala system presents into the Bill Hill area. Both the Magdala and Golden Gift ore types are hosted within the Magdala Volcanogenics. Underground ore types and deposits mined at SGM have previously been described in “Crocodile Gold NI 43-101 Report for Stawell Gold Mine, Victoria, Australia, May 2012”.
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8.2 Big Hill Deposit Types Big Hill is the up dip extension of the Magdala system. It contains Basalt Contact mineralisation, Central Lode mineralisation and Stockwork mineralisation, all typically seen in the Magdala system and mined historically from an underground perspective. Big Hill can be broken into 4 different geological ore domains: ¾ Mariners ¾ Allen’s ¾ Iron Duke ¾ Magdala Flank All except Mariner’s and Allen’s are separated by faults. Figure 8-1 to Figure 8-3 show modelled geological domains by area.
Figure 8-1: Section 5690 through Big Hill North, showing the Cross Course Fault (blue), Mariners ore domain (green) and Allen’s ore domains (pink).
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Figure 8-2: Section 5510 through Big Hill South, showing Scotchman’s Fault (upper blue surface), Iron’s Duke ore domains (purple), Magdala ore domains (red) and basalt waste domain (yellow).
Figure 8-3: Section 5090 through Davis area, showing Magdala ore domains (purple) and basalt waste domain (yellow). Note the previously mined Davis depletion.
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8.2.1 Mariner’s Mariner’s is the least problematic geological unit to define. It has a consistent width and orientation, averaging around 14m thick and dipping to the northwest (mine grid). It constitutes a shear package, with massive quartz veining. Less consistent is the Mariner’s spur (domain 111) which is modelled on the upper surface (sections 289/290N). This is only defined by 8 (3 are twinned) drill holes and the geology in this area will be more complex than modelled. Offset extensions of the Mariner’s shear have also been modelled down dip (Figure 8-4). The original model identified one offset extension; however two further offset domains were identified in this 2012 model update. Recent underground drilling of Mariners Lower in early 2012, drove an increased understanding of the structural framework, identifying increased faulting complexity and offset lodes which now connect the underground Mariners resource with the Big Hill model. The main Mariners Lode is offset by Fault 3 which strikes at 145 degrees (Mine Grid) and dipping at around 35-40 to the NE. It has an apparent reverse displacement of between 5 and 25m, decreasing northwards. This section of the lode (Mariner’s L1) is then truncated and offset at depth by the Cross Course Fault. In this model update the Cross Course Fault has been remodelled as two fault surfaces which bounds another lower lode offset (Mariners L2). A third, less horizontally extensive offset (Mariner L3) was identified and was modelled to be bounded by the cross course fault lower surface and the Scotchmans Fault. The Mariners L3 has a greater offset 15m-25m and plunges further to the north than the above two offset lodes, which is likely due to a greater influence from the Scotchman’s faulting. In reality these offset sections are likely to be more complexly faulted than modelled. Diamond core intercepts for Mariner’s are presented in Table 8-1.
Table 8-1 Diamond Core Intercepts
Drill hole Intersection Comment
SD540 36-45m sheared to massive vein quartz
SD549 29-35m massive to sheared vein quartz + sulphides
SD541 12-17.5m clay altered siliceous volcs, strong FeO cementation, ? sheared 17.5-33m sheared quartz + boxwork after sulphide
SD550 47-71m strongly sheared qrtz lode + volcs
SD544 12-25m quartz lode 25-30m siliceous volcanogenics, sheared, smokey vein quartz 30-36.5m siliceous volcs, pelitic, foliated, haem alteration
SD552 0-6m Massive quartz lode
Beneath the 404 Fault the general definition of the geology is hampered by a lack of diamond core information. The principal reason for modelling a fault in this location is to explain the lack of semi- planar continuity of the otherwise predictable Mariner’s structure. Without a fault, it would be
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necessary to invoke a 30-40° change in the orientation of the shear. Note that the geological confidence of the interpretation around these faults is lower than the more planar section of Mariner’s, because of the greater degree of structural complexity and uncertainty in the location of the faults, see Figure 8-4 below.
Figure 8-4 Section through Mariner’s lode showing structural complexity and offsetting
8.2.2 Allen’s Zone Interpretation of the Allen’s zone is somewhat problematic and the interpretation of the geology in this zone is of lower confidence than other sections of the ore body. Exposures in the Allen’s open pit show lithological layering (S0 and later S2/3 foliation), as well as massive veining to be very steeply dipping to vertical. Diamond core shows the same. Two massive veins exposed in the Allen’s Adit strike approximately N-S mine grid and dip vertically. Some stockwork style veining is seen near the top of the zone (immediately beneath the Mariner’s structure), but much of the geology consists of well bedded, foliated volcanogenic sediments which are unpredictably mineralised. Much of the gold occurs in these sediments. Because of the highly oxidised state of these rocks the form of sulphide mineralisation is largely obliterated. There is a poor correlation between quartz percentage and gold mineralisation. Accordingly the wire-framed structures which have been interpreted in the Allen’s zone should not be regarded as having hard boundaries for ore/waste selection. Rather they have been used to constrain a tonnage that should reflect the tonnage of ore grade material adequately. They have dimensions that reflect the attitude of mappable geology, which in turn will probably exert the main influence on gold mineralisation. However, the locations of the wire-frames are controlled as much by distribution of gold grades as by definable geological contacts/zones.
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It is worth noting that variography using quartz percentage (based on an indicator set to 1 above and 0 below 65% estimated quartz by volume) showed a strong continuity towards mine grid north (plunge of 20°, range of 80m to sill), while the principal direction identified for gold grades was towards 345° (65m to sill) i.e. both gold grade and geology show similar orientations of maximum continuity. Because of the angle of drilling insufficient data is available to identify the ranges in the vertical direction. It is suspected that a degree of remobilisation of gold within the weathering profile may exist, although no distinct enrichment horizons could be identified. The geology is of such variety that it is very difficult to separate any secondary effects from primary variation. The relationship between the Mariner’s and Allen’s zones is also still problematic. Little overlap exists beneath the main flatter section of Mariner’s and the Allen’s zone, which dies out rapidly northwards beneath Mariner’s. No clear evidence is seen of the timing relationships between these structures, the Allen’s mineralisation appears to hang as a pendant of dilational veining beneath a change in orientation of the Mariner’s shear structure.
8.2.3 Iron Duke Zone The Iron Duke zone is a wedge of geology occurring between the Scotsman’s Fault and the Lower Cross Course Fault. It is the up-dip extension of the volcanogenic package/shear zone which has been offset from the main Magdala shear system by reverse movement on the Lower Cross Course Fault. It is truncated above by the Scotchman’s Fault zone, which again displaces the ore body westward. Interpretation of the geology in this zone is hampered by the strongly oxidised nature of the rocks. It has been interpreted as a series of constrained stockwork style ore zones (same as Allen’s), where the envelopes have been used to constrain a tonnage. In reality the margins will have gradational boundaries.
8.2.4 Magdala Flanks “Magdala Flanks” is the term used to describe the package of volcanogenic rocks lying to the west of and continuing up dip from the nose of, the basalt antiform. The western margin of this zone is the contact between the volcanogenics and the mine schist, which often plays host to the Hangingwall shear. The eastern margin is generally the contact with the basalts, but above the basalt noses the eastern margin is a transitional boundary to siliceous eastern schists, usually marked by a clear grade boundary. The Hangingwall shear was most recently air-leg mined from the –180 and –145 levels (and from sub-levels up to –109), between 269 and 278 N sections in the early 2000’s. Significant old stoping on this shear was encountered up dip of those recent air-leg workings.
8.2.5 Scotchman’s Fault: The Scotchman’s Fault zone is a major structure which lies across the top of the basalt antiform and displaces the Magdala shear zone. It has been proven in underground development driven northwards at the –290 level. Its position at surface is pinned down by an exposure at 4730mE, 5500mN (Mine Grid). The general position of the fault can be clearly identified by the upward termination of anomalous and ore grade structures. Below the base of Transition Zone the fault can generally be easily identified in both RC chips and diamond core by the presence of graphite and sheared textures. The logged intercepts do not fall onto a single, neat plane but are dispersed over a zone. In the weathered profile identification of the fault is difficult. As is seen underground there is some significant mineralisation associated with the Scotchman’s Fault but this is of a highly variable
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and poddy nature. Hole SP298 encountered [email protected]/t Au before hitting a stope. However, other holes drilled to establish the continuity of this zone failed to encounter any significant mineralisation.
8.2.6 Cross Course Fault: Four lines of evidence support the presence of the Cross Course Fault zone: 1) The Mariner’s shear structure is truncated at depth. 2) Surface mapping picked up evidence of sheared textures with an appropriate strike and dip at two locations, which project to the interpreted position of faulting. 3) The Hanging Wall lodes (Scotchman’s Vertical and Cross Vertical) are offset by a fault – shown on old sections. This is where the name “Cross Course” fault is derived. Offsets by faults dipping in this orientation are also shown in a 1900 section of the Mariner’s and Sloan’s shafts. These also project to approximately the correct position. 4) Two of four diamond holes show good evidence of fault disruption and core loss which the upper and lower boundaries of the fault zone have been pushed through. The other two cored holes show little evidence of faulting, although zones of weaker core and core loss are found at the appropriate positions. As noted earlier, the Cross Course Fault was separated into two separate surfaces, and upper and a lower for the 2012 model update. In reality this is likely to be a zone of repeated faulting, but for the purpose of separating out a geologically consistent ore domain it has been modelled as such.
8.2.7 Lower Cross Course Fault: The Lower Cross Course Fault is a newly recognised fault. It was interpreted independently of the Cross Course Fault and it wasn’t until the two were viewed together that it was realised that they are co-planar. Again no definitive plane of faulting can be identified but the evidence of offset of the ore zones is compelling. It has only been modelled as a single plane rather than as a zone, principally because of a lack of drilling control – but also because the zone appears more discrete. The timing relationship of the Cross Course to the Scotchman’s is unknown. Below –80mRL there is a dogleg in the Scotchman’s Fault zone coincident with the intersection of the Cross Course Fault, which suggests the Cross Course may in fact post-date the Scotsman.
8.2.8 Basalts: As with previous interpretations basalts were interpreted as a series of thrust stacked antiformal noses. Detached noses floating above the main termination of the basalt are also present. These occur primarily within the weathered zone so deciphering the complex shearing around these noses is difficult.
8.2.9 Foliation and shearing Two dominant foliation fabrics are recognised. S2 is a pervasive schistosity accompanying tight to isoclinal folding. It is generally north-south striking and steep to vertical in orientation, although variation does occur. Dips generally flatten slightly away from the basalt nose/ shear zone area. S3 is a less pervasively developed crenulation cleavage. Regionally it has a shallow to moderate dip to the north or north east (Mine grid) although locally dip is to the northwest or west. In the Davis pit area the dominant foliation is S2, dipping vertically to steeply east. On the east wall of Davis Pit a set of D4 shears dip at a shallow angle to the northeast sub-parallel to the plunge of the basalt noses. A shear of this orientation appears to have been involved in the
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documented failure which occurred during the life of the Davis Pit. Shears of similar orientation have been interpreted in the deeper sections of the ore body.
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9 EXPLORATION
9.1 Current Exploration Significant exploration progress and successes has been made over the liffe of the Stawell project including the discovery of the current mineral resources and mineral reserves. In the period September 2008 - October 2012 (Figure 9-1), of note is the conversion of the Golden Gift 6, discovered in 2007, into reserve and the Golden Gift 6 Lower, discovered in 2008, into inferred resource. Scoping the northern extension oof the Magdala dome has also added to the C7 (Dukes) reserve. Exploration in 2011 focussed on big picture targets in the near mine environs and has previously been described in “Crocodile Gold NI 43-101 Report for Stawell Gold Mine, Victoria, Australia, May 2012”. Exploration work in 2012 is forecast to be $2.4 million and is primarily directed toward regional soil sampling and Wildwood drill Targets. 2013 will see a continuation of regional soil programmes through to a stage ready for drill targeting. An RC programme was completed over the Davis portion of Big Hill in early 2012. A total of 34 holes were completed for 2,238 meters. These holes have been included in the November 2012 Resource estimate, for grade estimation, geological interpretation and for void modelling.
Figure 9-1 Mine Geology Longitudinal section outlining near mine exploration from September 2008- October 2012
9.2 Expenditure Total geological expenditure (both near minne and regional) year to date, to October 2012 is $1.88M. Past and current mine site and regional SGM programmes have previously been described in “Crocodile Gold NI 43-101 Report for Stawell Gold Mine, Victoria, Australia, May 2012”. No further exploration is planned around the Big Hill pit areas. Further infill grade control drilling will be required to better constrain mineralisation and historical voids; however, the area is currently drilled to a 20 meter x 25 meter spacing.
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9.3 Exploration and Mineral Tenure Stawell Gold Mines comprises a package of tenements covering the Stawell Belt and areas south of Ballarat covering an area of approximately 1,986km2. A number of these tenements are under farm in or JV agreements (Figure 9-2 and Table 9-1). Under the Victorian Minerals Resources (Sustainable Development) Act 1990 ongoing relinquishment of tenements is required at regular intervals. These being 25% of the original tenement size at the end of year 2 and 35% of the original tenement size at the end of year 4. Amendments to the act which came into effect on the 1st Feb 2012 require an additional 20% of the original tenement size at the end of year 7 and 10% of the original tenement size at the end of year 10 (leaving 10% of the original tenement area). Exploration Licences that are more than 10 years old may be renewed for up to an additional 2 years. Further renewal may be given for a period not exceeding 2 years but only in exceptional circumstances and where it can be demonstrated that there is likelihood of identifying a mineral resource in the term. Following this term no further renewals are allowed. Tenements that are greater than 10 years old will have the 7 year and 10 year relinquishments applied (for a total of 30%) at their next renewal. An outline of current SGM tenement can be found in, noting a number of Stawell Gold Mines tenements ages are at, exceeding or coming up to their 7 and 10 year anniversaries and will require additional relinquishment of area or the tenement. EL3008 is considered a “strategic” exploration tenement under the Act amendments and may be renewed for up to 5 years with no additional relinquishment requirements. Further renewals for a period up to 2 years may be given only in exceptional circumstances and where it can be demonstrated that there is likelihood of identifying a mineral resource in that term. Following this term no further renewals are allowed. MIN5520 is a licence that does not currently have any mining activity being undertaken. Exploration has tested the northern extent of the Magdala system that this lease covers and has shown there to be low prospectively in this area. Under the Victorian Minerals Resources (Sustainable Development) Act 1990 a MIN may have up to 2 years of exploration activity before a work plan to mine must be submitted or the licence reverted back to an existing EL. It is planned to revert MIN5520 back into EL3008 in early 2013.
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Table 9-1 Regional Tenement Information
Annual Expiry / Name Number Area (km2) Expenditure Grant Date Renewal Comments commitment Date
Murtoa EL3941 318 $269,000 17/12/1996 17/12/2012 Kiska Metals JV Glenorchy EL4279 24 $39,000 14/05/1998 21/06/2013 SGM Managed Wildwood EL3008 363 $105,000 16/12/1988 20/06/2013 SGM Managed Stawell MIN5520 10 $337,600 31/03/2010 30/03/2030 SGM Managed SGM MIN5260 4 $899,730 31/05/1985 30/05/2020 SGM Managed Ararat Sth EL4695 409 $205,500 17/04/2003 16/04/2013 SGM Managed Dundonnell EL4730 197 $113,500 19/11/2003 18/11/2013 SGM Managed Barrabool ELA5443 496 $95,500 Application only SGM Managed
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Figure 9-2 SGM regional Victorian tenement areas
9.4 Kiska Metals Corporation Joint Venture Kiska Minerals Corporation, a Canadian listed company (TSX.V: KSK), had the right to earn a 50% interest inn one or all of the following properties EL3941, EL4695 and EL 4730, by making expenditures of $500,000 by 30th June 2012 for each tenement in which the JV is sought. If Leviathan elects not to continual to contrribute on the property at this point KSK may earn the remaining 50% by spending an additional $2,000,000 on exploration and ddevelopment within the next 4 yeaars. Otherwise KSK can elect to earn an additional 10% interest in the property by spending an additional $1,500,000 on exploration and development per property within the following 3 years with a minimum expenditure of $300,000 each year. Under the aggreement KSK assumes responsibillity for management of the tenements including meeting annual expenditure requirements and reporting. In November 2012 Kiska acquired a 50% interest in ELL3941 by meeting their expenditure commitments under the agreement. Expenditure commitments ffor EL4695 and EL4730 were not met and KSK have relinquished theeir interest in these tenements.
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9.5 Interest in Navarre Minerals Navarre Minerals is an Australian listed company (ASX: NML). Navarre Minerals is a Victorian based mineral exploration company searching for gold and base mineral deposits. Through a tenement sale agreement Leviathan Resources Pty Ltd has at the time of reporting 55,829,603 fully paid ordinary shares, a proportion of which are escrowed 24 months from listing which equates to an approximate 9.3% shareholding in Navarre. Additionally, Leviathan holds a 2% Net Smelter royalty on EL4897. This is subject to a buy back clause where the holder of the project will have the right to buy back 1% of the royalty for a payment of $2M at any time within 4 years of signing the formal variation of tenement sale agreement.
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10 DRILLING
10.1 Stawell Gold Mines Mineral Resource Definition Process The Mineral Resource definition process at SGM is an ongoing activity. Current Mineral Resources extend from surface to 1800mRL (effectively 1800m below surface). Geological information is collected by a variety of methods with the objective of improving the confidence of the Mineral Resource estimates prior to and during the mining process including grade control drilling for development and stope definition. Given the continuous and ongoing nature of the Mineral Resource process the data utilised varies as the mining operations develop towards the resource area (Table 10-1). A summary of the main drilling methodologies employed and data types utilised is as follows with specific details on the sampling and assaying methodologies given below. Surface Reverse Circulation (RC) Drilling ¾ Utilised for definition of near surface resources where diamond drilling is not required for detailed structural definition ¾ Utilised as a method for pre collaring deeper diamond drill holes ¾ Drilling completed using 5 ¼ ” Face Sampling Hammers unless otherwise specified ¾ Samples collected from cyclone discharge ¾ Hole depths vary but are generally less than 200m ¾ Samples collected at 1m sampling interval, bit pulled back and flushed between intervals ¾ Sample tipped into three tier splitter from crate to ensure equal quantities available to all vanes ¾ Splitter cleaned by shaking/banging/brush/air compressor as necessary between samples ¾ Cyclone cleaned at regular intervals by banging/checked by hand/arm ¾ Around 99.5% of drilling was conducted dry. Big Hill has been very effectively de-watered by the underground mining operation. In a handful of holes, samples were damp after a rod change. Where a sample could not be effectively riffle split it was spear sampled using a PVC spear. ¾ Sample volumes were not routinely recorded pre 1999. An initial programme in which samples were weighed showed little variation in recovered volume by depth or geology and it was decided that routine weighing was unnecessary. ¾ Drilling is conducted dry or with sufficient air to ensure collected samples are dry ¾ In some project areas casing advancing technologies have been utilised to ensure drilling through fill produces reliable samples ¾ Undertaken by contract drilling personnel under the supervision of SGM Geology Surface Diamond Drilling ¾ Primarily used in initial exploration programmes, near surface resource definition and to provide structural and geological information in near surface RC drilling programmes ¾ Drilling by wireline methods. ¾ Hole sizes PQ3, HQ3, NQ3, HQ2, NQ2, BQ2
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¾ Hole depths vary from <100m to >2000m ¾ Directional drilling utilised for specific tasks ¾ Core orientation devices often utilised to aid in structural interpretation ¾ Undertaken by contract drilling personnel under the supervision of SGM Geology Underground Diamond drilling ¾ Utilised at all stages of the geological process, Exploration, Resource Definition and Grade Control ¾ Drilling by conventional and wireline methods ¾ Hole sizes HQ3, HQ2, NQ2, BQ2, LTK60, LTK48 (Not used post 1997) ¾ Hole depths vary from <50m to 1200m ¾ Directional drilling utilised for specific tasks ¾ Undertaken by contract drilling personnel under the supervision of SGM Geology Open Hole Percussion Sampling “Sludge Sampling” ¾ Utilised after development of ore drives for final stope definition ¾ Hole sizes 89mm open hole ¾ Hole depths vary from 5m – 25m ¾ Samples of cutting of variable length are collected primarily for geological logging of the chips to identify major faults and geological contacts ¾ Undertaken by SGM production blasthole rigs
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Table 10-1 SGM geological processes and approximate drill spacings
Resource Activity Area Target Type Criteria Geological Data available Classification
• Geophysics • Conceptual geological model • Mapping • Geophysical anomaly Exploration Conceptual Targets • Wide spaced Exploration Drilling • Geochemical anomaly • Conceptual geological models
• Wide Spaced Exploration drilling • Geological model confirmed by drilling • Assay information Confirmed Targets • Mineralisation confirmed by drilling • Drill logs
• Broad spaced Grid Drilling • Geological continuity established • Detailed cross sectional interpretations • Ore grade intersections established Pre resource, • Assay information Scoped targets • Preliminary geological model established Resource Target • Drill logs • Drill spacing 160 m X 120 m • Geological models
• Regular Grid Drilling • Geological continuity confirmed • Detailed cross sectional and 3D interpretations • Ore grade intersections continuous • Assay information Resource • Geological interpretation modelled Inferred Resource • Drill logs Definition • Geostatistical model established • Geological models • Drill spacing 80m X 60m • Geostatistical model • QA/QC analysis • Regular Grid Drilling • Geological continuity confirmed • Detailed 3D modelling and interpretations • Ore grade intersections confirmed • Assay information • Geological interpretation modelled • Drill logs • Geostatistical model Indicated Resource • Geological models • Economic analysis • Geostatistical model • Drill spacing 40m X 40m to 30m X 30m • QA/QC analysis • Close Grid Drilling • Confident Geological continuity • Detailed 3D modelling and interpretations • Ore grade intersections confirmed • Assay information • Geological interpretation modelled • Drill logs Grade Control • Geostatistical model Indicated Resource • Geological models • Economic analysis • Geostatistical model • Drill spacing 20m X 20m • QA/QC analysis • Close Grid Drilling • Confident Geological continuity • Detailed 3D modelling and interpretations • Ore grade intersections confirmed • Assay information • Geological interpretation modelled • Drill logs • Geostatistical model • Geological models • Economic analysis Measured Resource • Geostatistical model • Level Development above and below • QA/QC analysis • Drill spacing 20m X 20m to 10m X 10m • Development face mapping sheets and ore runs • Open-hole sludge drilling • Open-hole sludge drill geological data where required
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10.2 Drilling Process A flowsheet of the diamond drill process from design to implementation is shown in Figure 10-1. The diamond drill contract personnel provide a daily record of drilling activities for all drill rigs. A copy of a daily drill record sheet is shown in Figure 10-2. Data from the daily record sheet is entered daily to a site database for tracking of drilling production and to enable tracking of drilling progress interrogation at a later date. Geological personnel track the drill hole path and maintain in control of the daily activities of all drill rigs including which drillers were responsible for various sections of the hole should there be issues with core presentation or down hole depths that require clarification. A regime of regular rig audits and inspections are also used to assist with maintaining the high level of core presentation and sample quality. These drill records are kept indefinitely enabling review of drillhole information many years after completion of drilling.
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Figure 10-1 SGM Diamond drilling process flowsheet
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Figure 10-2 An example of a daily drill record from Big Hill 1998 surface drill programme.
10.2.1 Big Hill Drilling Drilling over the Big Hill area is shown in Figure 10-3 and summarised in Table 10-2 below. Note all drilling up to and including the 2012 RC drilling has been incorporated innto the November 2012 updated rresource estimate for the Big Hill area. Additionally the 2012 RC holes have been utilised to update void models used in this estimatee. Historically a number of campaigns have been undertaken to target this area, see Table 10-2 below. Drilling was on and off up to 19977 when a major programme of drilling was undertaken to define the potential of the Big Hill area. During 1997/98, 217 reverse circulation drill holes were drilled for 19,500m, and 30 cored diamond holes for 2,369m were also drilled. This drilling took the Big Hill area to a hole spacing of nominally 20m North by 25m RL. Two later RC programmes were completed over the Davis area or the southern end of Big Hill to better constrain the grade in this area.
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Table 10-2 Summary of Big Hill Drilling
Big Hill Drilling Summary
2012 programme 2008 programme 1997 to 1998 Pre May 1997/8 Total programme
# metres # metres # holes metres # metres # Metres holes holes holes holes
Surface Percussion 34 2238 7 589 217 19,537 63 5,916 321 28,280
Surface Diamond 18 1,972 81 * 99 20,370
Diamond Tails 12 397 12 397
U/G 145-180 167 14,920 167 14,920 diamond levels
-212 38 2,242 0 2,242 upholes
66,209
*Although 18,300m of older diamond core drilling falls within the PCF limits, most of this drilling targeted deeper sections of the ore body and lies within mine schist west of the ore zone.
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Figure 10-3 Plan showing Big Hill pit used for Resource reportinng, drill coverage and drill orientations. Notiing recent 2012 RC infill drilling in red. Used in both the Resource estimate and void modelling
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10.2.2 Reverse Circulation Drilling RC drilling was conducted in five programmes over a number of decades. In summary: Holes SP1-SP81 was drilled by Western Mining Corporation between 1978 and 1993. Rig type is unknown but it is guessed that they were probably drilled using a 4.5” cross-over sub hammer. The sampling protocol for these holes is unknown, but probably only consisted of grab or spear samples and assays should be treated with caution. Holes SP137-148 were drilled by SGM during Mar/April 1997, using a 5¼” face sampling hammer on a track mounted drill rig. Holes SP180-447 have been drilled by SGM during 1997/8. All of this drilling was carried out with two similar UDR 1000 drill rigs, using a 5¼” face sampling hammer. Holes SP577 to SP583 were drilling by SGM during May 2008, using a 5¼” face sampling hammer. Holes SP606 to SP638 were drilling by SGM between February 2012 and April 2012, using a 5¼” face sampling hammer on a track mounted drill rig.
10.2.3 Diamond Drilling Surface diamond holes are of two generations. Pre SD500 were drilled by Western Mining (and previously in some cases). In many cases these were rotary pre-collared to around 100m. Core was generally ½ core (split or sawn) sampled. Assay method was mostly by AAS. Post SD500 was drilled by SGM during the course of this exploration programme. All of this core was HQ in size, full core sampled and fire assayed as noted in section 11 Underground diamond holes can be broken into two sets also: Holes after MD1000 were full core sampled and assayed by Fire Assay, with the entire volcanogenic zone sampled. In addition holes over MD 1300 were logged in Datcol or acQuire and all data was entered to database as set out in section 11 Holes drilled before MD1000 were mostly ½ core sampled (split or sawn) and assayed using Aqua Regia digestion. Only selected sections of the Volcanogenic zone were sampled. A programme of re-logging and re-assaying as previously noted has minimised the impact of these lower confidence assay techniques. Additionally these holes were focused toward underground targets and spatially occur outside the current optimised pit shell, again minimising the opportunity for these to create variation, see Figure 10-4.
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Figure 10-4 3D view looking towards the NE, note displayed drill hole traces are holes before MD1000, all are drilled from underground and spatially occur outside the pit shell used for Resource reporting.
10.2.4 Drill hole Orientation Where possible, drilling is oriented perpendicular to the structures being tested. The nature of the mineralisation around Big Hill and the availability of suitable surface drilling platforms due to community constraints will always result in compromises in the ability to obtain near perpendicular tests of the mineralisation. Overall the Big Hill Resource area has been drilled a regular 20m along strike intervals with the drill holes oriented perpendicular to the strike and dip of the main mineralisation system with an up and down dip spacing of 25m, see Figure 10-5.
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Figure 10-5 Plan view (top) and cross section views (A-B and C-D) ) (Note: Potential Pit = grey, Surface topography including previously mined Davis pit = red, Davis lode = light green, Iron Duke lode = dark green, Allen’s lode = light blue and Mariner’s lode = dark blue.
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10.2.5 Collar Survey Control All survey control foor Big Hill surface and underground drilling programmees were established by SGM survey personnel. Survey control pooints are maintained in a number of surface locations and in the underground decline by SGM survey personnel and these locations provide the control for all mark out and pick-up surveying that is conducted. On conclusion of underground drilling and drill hole grouting, diamond drilling personnel will insert a wooden wedge labellled with the drill hole ID into the collar of the hole. This provides permanent identification of the drill hole collar to ensure matching of surveying information to the correct drill hole collars. On concllusion of surface drilling and after the collar has been surveyed SGM personnel cap the hole and rehabilitate the site, effectively removing the collar at surface. These holes can be re-accessed if required depending on the age of the hole. The collar surveey information is entered in the SGM database by data managerss. An example of the information supplied by SGM surveyors and the checklist utilised to ensure appropriate information is collated is shown in Figure 10-6 to Figure 10-7. The coordinate system in use at Stawell is a modified version of AMG which is discussed in Section 4.3. All survey data pertaining to the mining operation is stored in this coordinate system.
Figure 10-6 Collar survey information and drill hole survey information checklist 1997 to 98 programme
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Figure 10-7 Collar survey information and drill hole survey information checklist, modern process For 2008 and 2012 programme
10.2.6 Down hole Survey Control Downhole survey control is managed by utilising down hole cameras to survey the drill hole path. Electronic single shot instruments (REFLEX® and RANGER® tools) have progressively been used in preference to the Eastman® mechanical cameras since 2002 at SGM as shown in Figure 10-8. The vast majority of the down hole surveys of diamond drill holes that are utilised in the estimation of the Mineral Resource estimate detailed in this report have been made with Eastman® mechanical single shot cameras. Some of the deeper surface diamond drill holes have been surveyed using a North Seeking Gyro instrument however this is less relevant to Big Hill as while these holes may be collared in the area they target deeper areas of the underground mining operations. Down hole survey instruments routinely measure azimuth relative to magnetic north and declination (dip) relative to the horizontal. A correction is applied to convert Magnetic North to Grid North. The detail of how this correction is currently applied is shown in Figure 10-9.
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Downhole survey method SE ‐ Surveyed:Eastman SD ‐ Surveyed:Digital MD6500
MD6000
MD5500
MD5000
MD4500
MD4000 HoleID MD3500
MD3000
MD2500
MD2000
MD1500 8 8 8 9 9 9 0 0 0 1 1 1 2 2 2 3 3 3 4 4 4 5 5 5 6 6 6 7 7 7 8 8 8 9 9 9 0 0 0 1 1 1 2 9 9 9 9 9 9 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 0 01 01 01 01 01 01 01 /1 /1 /1 /1 /1 /1 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 /2 01 05 09 01 05 09 01 05 09 01 05 09 01 05 09 01 05 09 01 05 09 01 05 09 01 05 09 01 05 09 01 05 09 01 05 09 01 05 09 01 05 09 01 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ 1/ Date
Figure 10-8 Plot of drillhole survey method (SE = Eastman Single Shot in purple, SD = Electronic survey instrument in blue) by time. Post 2001, the standard survey instrument used has been an electronic single shot downhole survey tool Contract drilling personnel are responsible for providing survey information at pre-determined spacings down the drillhole. For diamond drilling the first survey is taken at 15m downhole and is effectively used as the collar survey. This depth is used as it reduces the influence of magnetic effects associated with the drill rig and support equipment. Subsequent surveys are taken at 30m spacing or at closer intervals where deemed necessary by the supervising geologist. The contract personnel record survey details on the daily drilling record sheet and also on a separate survey record sheet (Figure 10-10) from which the information is entered to the acQuire database system. The electronic instruments provide a direct reading of the magnetic field intensity at the survey locations. This reading can be used to determine if survey readings have been influenced by magnetic material down hole.
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Figure 10-9 Magnetic declination correction as currently applied to SGM drill hole data
RC drill holes were surveyed using an Eastman camera at the completion of the hole. Most holes had open-hole surveys beginning at 20m aand an end of hole survey. A smaall number of holes had in-rod surveys only, which due to magnettic influence of the rods gives a dip measurement only. This method was tested for accuracy in Occtober 1997 with an electronic multi-shot (EMS) and was found to be accurate to within acceptable limits, internal memo Mike Stewart “Down Hole Survey Test, Surface RC Holes, 1997”.
Figure 10-10 Survey record sheet, example from Big Hill 1998 surface programme
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10.2.7 Down hole Survey Quality Control Currently at SGM several quality control and quality assurance processes are in place to ensure that appropriate survey (down hole and collar) information is stored to the database. Apart from the Database Managers checklist, a review sheet for the down hole survey information is provided to the responsible geologists such that this information can be validated and where required adjustments made to the survey information. A copy of this sheet is shown in Figure 10-11. It is unclear what the methods for survey control were during the post 1999 period, however, visual checks of the drill holes do not show any spurious results in 3D space and holes generally have a consistent trend. Spot checks of Eastman surveys correlate well with values in the database and from internal reports regular check of the Eastman instrument were performed on a test bed.
Figure 10-11 An example of drillhole collar survey information collated for geologist review. Note some corrections to erroneous data have been made as indicated
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For longer drill hole traces the survey information is plotted to provide a graphical review of the information method is utilised where adjustments to the survey information can be made using the overall trend of the drill hole trace, Figure 10-12.
Downhole Survey Checks ‐ Example 316.0 0.0 ‐2.0 314.0 ‐4.0 312.0 ‐6.0 ‐8.0 Planned 310.0 Azi ‐10.0 Dip Actual Azi Azimuth 308.0 ‐12.0 Planned ‐14.0 Dip 306.0 ‐16.0 304.0 ‐18.0 15 45 75 105 135 165 185 Depth
Figure 10-12 Example of the check plots used to correct down hole survey information
Where clear discrepancies have been identified with the validity of the survey information and adjusted surveys entered, the original surveys are given a lower priority in the database system. A record of survey methods and or adjustments are maintained in the main acQuire database as part of the audit trail. Stawell Gold Mines personnel utilise a survey camera test bed with known azimuth and dip to routinely check the accuracy of the down hole survey cameras. This test bed (Figure 10-13) is located on the surface well away from any potential magnetic sources and is utilised by contract drilling personnel to routinely check camera performance and determine if equipment requires servicing or re-calibration.
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Figure 10-13 Photograph of the SGM survey camera test bed.
10.2.8 Diamond Drill Core Processing All diamond drill core was delivered to the SGM core processing facilitty by the diamond drill contractor. Diamond drill core is washed to remove grease and inddividual core trays are photographed in a light controlled installation (Figure 10-14) prior to lay out on benches ready for logging by the site geologists. As part of the standard SGM geological procedures, all core collected from diamond drill holes is photographed and a complete record of digital core photographs is available to assist in the geological interpretation process. Pre mid 2001 digital core photography was not in place. Photographic slides are available in most cases and SGM has undertaken a programme to begin converting slides to digital formats. A detailed flow sheet of core processing activities is shown in Figure 100-15. Prior to drill hole number MD2678 and SD607 the core photography was taken on film and sttored as prints. An example of the style of core photography that is available for all diamond drilling is shown in Figure 10-16.
Figure 10-14 SGM core photography installation
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Figure 10-15 Diamond drillcore processing flow sheet
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Figure 10-16 An example of Big Hill diamond drill core photographs stored digitally from scanned slides, 1998 drill programme
10.2.9 Logging All diamond drill core was logged by the site geological teams using a standardised logging methodology. The data is captured electronically at the point of collection using either a barcode logging “Datcol” software system or and acQuire logging system. The “Datcol” system was developed on site in the mid 1990’s and has remained the standard process since that time where the key tables for lithology, alteration, structure, and geotechnical information are populated during the logging process. During 2009 the acQuire logging system was developed to replace the “Datcol” system. The acQuire logging system utilises the same standard key tables for the lithology, alteration, structure, and geotechnical information, which are populated during the logging process. RC logging is recorded first on paper logs utilising a RC logging template. The logs are than transcribed into an excel database, validated by the geologist responsible for the programme and uploaded into the acquire database.
10.2.10 Core Recovery During the logging process any lost core is estimated and logged as lost core with a specific start and end interval. A review of the database indicates good core recovery throughout the deposit particularly adjacent to the major mineralised zones. Where core is lost it is usually associated with significant faulting or historical voids. Lost core is identified in the logging as “LOST” and as such there are very few if any assay intervals utilised in the Mineral Resource estimate where core recovery is less than 100%.
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10.2.11 Diamond Drill Core Sampling During the logging process the geologist will mark up the intervals of core required for sampling. Not all diamond core is sampled. Historical sampling has identified the key lithological and structural units that will host mineralisation and the selection of units for sampling follows the protocols shown below. All Magdala Facies, also known as ‘Magdala Volcanogenics’ are sampled for assay A minimum of 2.0m into the hangingwall and/or footwall is sampled Fault zones and zones of sulphide are sampled at the geologists discretion Magdala Basalt and Albion Formation units are sampled at the discretion of the logging geologist Not all diamond core is cut in half prior to sampling. Sampling of diamond drill core follows one of two methods as detailed below. 1) Exploration and Resource Definition – HQ or NQ drill programmes ¾ Core is logged and geological derived intervals are marked up for sampling ¾ Sample intervals are matched to geological boundaries (structural or lithological) and fall within the range of 0.10m to 2.0m. The average sample interval is approximately 1.0m ¾ For Resource definition drilling programmes 1 in 5 holes is cut with a diamond saw prior to sampling and one half of the core is sent for assay. The remaining half core is retained as a record within the core library. All other drill holes are sampled as whole core, which is sent for sample preparation and assay as per the flow sheet shown in Figure 10-15. All drill holes deemed to be for the purpose of exploration are ½ core sampled and the entire remaining core retained in storage on site at SGM. 2) Grade control diamond drill programmes – NQ or LTK60
Core is logged and geological derived intervals are marked up for sampling Sample intervals are matched to geological boundaries (structural or lithological) and fall within the range of 0.10m to 2.0m. The average sample interval is approximately 1.0m For grade control drilling programmes drill holes are sampled as whole core with the entire sample sent for sample preparation and assay as per the flow sheet shown in Figure 10-15. Detailed operating procedures for sampling of diamond drill core are used at Stawell Gold Mines to ensure uniformity of process and prevent errors.
10.2.12 Big Hill Diamond Drill Core Sample Intervals The sample interval statistics for Big Hill are shown in Table 10-3. Some sample methodology variation exists with the historical drilling.
Surface holes pre SD500 holes were drilled by Western Mining. In many cases these were rotary pre-collared to around 100m. Core was generally ½ core by splitting or core saw sampled. Post hole number SD500 were drilled by SGM during the course of the 1997 to 1998 exploration programme. All of the core was HQ, full core sampled and processed as described above.
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Earlier underground diamond holes, holes drilled before MD1000 were mostly ½ core sampled by splitting or core saw sampled. Holes after MD1000 to MD 1300 were full core. Holes over MD 1300 were processed as described above.
Table 10-3 Diamond drill core sample interval statistics for samples used in November – 2012 Resource Estimate
Total 13,124.12
Min 0.2m
Max 2.2m
Ave 0.93m
10.2.13 Big Hill RC Sampling A summary of the RC sampling protocols that were utilised for this programme and for other RC drilling that has been conducted on site are given below: All SGM RC sampling was carried out using the following protocol Generally the entire hole was sampled from the collar unless it was recognized as recent fill or material associated with the construction of the drill pads Samples were collected at 1.0m sampling interval, bit pulled back and flushed between intervals Samples discharged into tightly fitting plastic sample bag from cyclone Sample transferred to a rectangular plastic tub the same size as the splitter Sample tipped into three tier splitter from plastic tub to ensure equal quantities available to all vanes 1/12 sub split (nominally 3kg) collected in calico sample bag, tied and placed in lots of five into plastic bags Residue sample collected in original sample bag and transported to bag farm for storage Splitter cleaned by shaking/banging/brush/air compressor as necessary between samples Cyclone cleaned at regular intervals (completion of each hole minimum) by banging/checked by hand/arm Every 20th sample re-split to give a 2nd 3kg sample, field splits dispatched along with first splits Samples dispatched to Laboratory for analysis by Fire Assay. In the case of the Big Hill Mineral resource estimate all assaying up to and including 1998 was conducted at Aminya Laboratories Ballarat
10.2.14 Reliability of Samples It is the opinion of the Qualified Person that the drilling and sampling methodologies employed by SGM are of a high standard and provide representative tests of the mineralisation suitable for the estimation of Mineral Resources. Older sampling or sampling up to and including 1998 are difficult to validate, however, it is believed similar processes were in place at this time and have been
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STAWELL GOLD MINE reported on in “Technical Report for Stawell Gold Mine, Victoria Australia, 28 March 2008”. Standard drill spacing’s adopted by SGM for Big Hill are appropriate for the various stages of Mineral Resource development and whilst other factors also contribute to decisions regarding classification of the Mineral Resources the drill spacing’s discussed in this section enable appropriate geological interpretation and Mineral Resource classification decisions to be made.
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11 SAMPLE PREPARATION, ANALYSIS AND SECURITY The Qualified Person has reviewed the sample preparation, assay and sample security processes utilised at SGM a summary of which is included in this section. All sampling and sample preparation is currently completed by employees of SGM. The employees follow appropriate written procedures as documented in the section. Older sampling and assaying or sampling and assaying up to and including 1998 is difficult to validate, however, it is believed similar processes were in place at this time and have been reported on in “Technical Report for Stawell Gold Mine, Victoria Australia, 28 March 2008” and a summary of variations is included within this section.
11.1 Assay Laboratories During the life of the SGM a number of laboratories have been utilised for routine assaying of diamond drill core and RC samples. The details of the laboratories and the periods for which assaying has been conducted are as follows; ¾ SGM site laboratory. Utilised intermittently prior to 1995 for assaying of diamond drill core and RC samples ‐ Non accredited company assay laboratory ‐ Assay method was 10g Aqua Regia with pre-digest roasting and AAS finish. ‐ Assaying of diamond drill core and RC samples was discontinued in 1995 and the laboratory sample preparation and assay methods updated to industry standard practice ‐ Now utilise a 25g Aqua Regia method with pre-digest roasting and AAS finish ‐ Utilised for metallurgical assaying, UG Face sample and other geological grade control sampling ‐ Undertake routine sample preparation of diamond drill core samples – post 2004. ¾ WMC Ballarat Assay laboratory utilised prior to 1995 for assaying of diamond drill core and RC samples ‐ Non accredited company assay laboratory ‐ Assay method was 10g Aqua Regia with pre-digest roasting and AAS finish. ‐ Assaying of diamond drill core and RC samples was discontinued in 1995 and the laboratory sample preparation and assay methods updated to industry standard practice. ¾ AMDEL Laboratories - Adelaide SA ‐ ISO 9001 accredited ‐ Utilised intermittently from 1995 through to present day ‐ Primary supplier of assay services during from 2004 – mid 2007 ‐ Ongoing utilization for check assays ‐ SGM reduced reliance on AMDEL mid 2007 as a result of very slow turnaround of assays results ¾ AMINYA Laboratories – Ballarat VIC ‐ Not accredited ‐ Primary supplier of Assay services for Diamond drill core and RC samples period 1995 – 2004 ‐ Discontinued in 2004
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¾ Intertek Genalysis formally Genalysis Laboratory Services – Perth WA ‐ Genalysis is a NATA accredited laboratory to ISO 17025. ‐ Provider of assay services during the period 2004 - 2006 ¾ ALS Laboratory Group – Orange NSW ‐ ALS is accredited to ISO 9001 and ISO 17025 ‐ Primary provider of assay services to SGM from August 2007 to present day
11.1.1 Sample Preparation The sample preparation protocol for diamond drill core is shown in the flow sheet given in Figure 11-1. This sample preparation flow sheet was developed in 1995 and has been in operation for all SGM diamond core and RC samples since that time. During the period 1995 to 2004 all sample preparation was conducted by the assay laboratory facilities. In 2004 it was decided by site personnel to complete this task on site at the SGM laboratory facility. The sample preparation follows the same process utilising modern sample preparation equipment. ¾ Daily Primary Crusher and Pulveriser size fraction calibrations are reported in the weekly SGM site laboratory report to ensure that the size fractions are meeting the set standards. The size fraction calibration for the Crusher is 75% of material must past through a 2mm screen (LABSOP-060 Boyd Crusher Size Fraction Analysis). For the Pulveriser the size fraction calibration is 90% of material must pass through a 75µm screen (LABSOP-061 LM5 Pulveriser Size Fraction Analysis). ¾ A quartz flush is inserted, at a 1:5 ratio, at the crushing stage for all Diamond and RC drill core. If visible gold is identified at the logging stage then a quartz flush is inserted after every sample within that mineralized zone. ¾ Splitting using a vibrating feed cone splitter. ¾ Pulverising to 95% passing 75um using Labtechnics LM5 pulverising mills. ¾ A quartz flush is inserted after every sample at the pulverising stage for all Diamond and RC drill core. By retaining responsibility for this work through the existing site based facility SGM has flexibility in sending the pulps only to a variety of assaying laboratories and also retain the coarse rejects on site for ongoing metallurgical test work programmes.
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Figure 11-1 SGM drill core sample preparation, assay and QA/QC flowsheet
11.1.2 Sample Transport and Security All drill core and RC samples are delivered directly to the mine site based core farm facility. This is on a shift by shift basis for underground drill core and daily basis for all others. Access to the mine site is restricted to authorized personnel only or as a visitor under full supervision by SGM personnel. A single site access point exists which is manned 24:7 by security personnel.
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Security of drill core and samples is managed by maintaining records throughout the complete process from drilling, core processing, logging, sampling, sample preparation and assaying through to return of results. Key record keeping utilised in managing sample and data security are; ¾ Daily drilling records are entered to the database which provides records of drill core produced. ¾ Core is photographed within 24 hours of being delivered to the core processing facility. ¾ The SGM sample processing facility is located on the Mine Lease within a security fenced area. All core stored here is only able to be accessed by SGM personnel. ¾ At the conclusion of logging a sample requisition sheet is generated listing sample numbers, assay standard insertion and assay requirements. This is loaded directly to the acQuire database enabling tracking of samples after this process. ¾ SGM personnel are trained in appropriate procedures for logging and sampling of the diamond drill core and generate an Analytical Request sheet outlining sample ID and assay requirements. ¾ The production of carefully labeled sample pulps for dispatch by registered posts. The pulps are dispatched from the SGM prep laboratory to the assay laboratories using registered post or courier services. Consignments travelling by registered post or courier services are required to be signed off by each leg of the postage route on arrival and can be tracked online. The assay laboratories are also required to send a statement informing SGM that the pulps have arrived and that the samples, as detailed on the analytical request sheet can be accounted for.
11.1.3 Assay Methods A summary of the laboratory methods utilised by the various laboratories is given in following Table 11-1. The majority of assaying for gold that are utilised in the Mineral Resource estimates have been completed by fire assay method (30 – 50g charge weights) with an AAS finish. A small number of holes drilled from underground operations potentially have aqua regia assays (due to their age no assay type is recorded in the database). These are from historical programmes, occur outside the pit areas and likely have little influence on the resource estimate, see Section 10.2.3 & Figure 10-4.
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Samples from the 1997 to 1998 Big Hill programmes were dispatched to Aminya Laboratories (Ballarat) for analysis by fire assay. Samples from the 2008 Big Hill programme onwards were prepared on site at SGM and dispatched to ALS Chemex.
Table 11-1 Laboratory assay method codes, descriptions and limits of detection Lab Method Limit of detection Laboratory Description Code (ppm) Aminya PE01S 50g Fire Assay with AAS finish 0.01
AMDEL FA1 40g Fire Assay with AAS finish 0.01
Intertek FA25 AAS 25g Fire Assay with AAS finish (used for 0.01 Genalysis repeats only) Intertek FA50 AAS 50g Fire Assay with AAS finish 0.01 Genalysis (Standard method used) ALS Au-AA26 50g Fire Assay with AAS finish 0.01
ALS Au-AA25 30g Fire Assay with AAS finish – Used 0.01 post January 2008
For samples reporting below LLD, a value of 0.5xLLD is utilised as standard in resource estimation.
11.1.4 Big Hill Re-logging and Re-assaying A programme of re-logging carried out in 1997 showed a very poor correlation existed between original Aqua Regia assays and subsequent fire assays. A summary of the re-logging programme is shown in Figure 11-2. In addition, assaying of previously unsampled volcanogenics showed not all mineralised zones had been sampled. In addition all original logs were paper based, and had been logged to an earlier WMC logging scheme. Consequently in 1997 it was decided to embark upon a substantial programme of re-logging/re- assaying of old core. The bulk of the underground drill holes collared on the -180 and -145mRL levels were re-logged and re-assayed by site staff, G. Masur with assistance from G. McDermott. If the previous interval had been re-split, the remaining ½ core was assayed. If it had not been previously sampled the remaining full core was sampled. Re-sampling did not mix ½ and full core samples.
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Diamond Core Re-logging Programme, 1998
Surface Diamond 6 1,209 6 1,215
U/G 145-180 levels 71 5,922 71 5,993 diamond
-212 upholes 14 837 14 851
8,059
Figure 11-2 Diamond Core re-logging Programme A scattergram of original (aqua regia) vs new (fire) assays, composited to equivalent 1m intervals, again shows a very poor correlation. Check assaying of lab duplicates during the current programme showed good repeatability of fire assay grades, see Figure 11-3.
Figure 11-3 Aqua Regia vs Fire Assay grades for re-logged core
A second assaying problem became apparent during this programme, arouund 20 holes in the low MD300’s series had a problem possibly relating to incorrect dilution erroors. The original assays showed long runs of elevated values in the range 0.5-1.2g/t Au. Re-assays of these holes showed that the intervals actually have background grades in the range 0.1-0.15g//t Au. The problem can be clearly identified visually on section and is clearly shown in Figure 11-4. All 20 holes were re- assayed via fire assay in 1997, see also Section 10.2.3.
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Figure 11-4 Aqua Regia vs Fire Assay grades – dilution errors
11.2 Database Storage and Integrityy All SGM drilling data is stored within the “acQuire” Database Management System. The acQuire database has been in place at SGM since November 1998. Prior to acQuire site data was stored in a Surpac database and Paradox database with a Surpac front end. Alll data was migrated to acQuire in November 1998. In general this migrated data contains the raw information only i.e. assay results, with little to no additional information on QAQC or assay method or type. The current database operates in an SQL Server framework and data security is established by having various levels of user access rights. Stawell Gold Mines maintain a security access system where looading and manipulation of data is only conducted by one of two data managers. All geological personnel have access to the database for read only purposes. Analysis results are received from laboratory in fixed digital format. The load routine imports assay data matching against sample ID created during the logging procedure. In March 2010, the acQuire database was updated to the CorpAssay ADM (acQuire Data Model) and QAQC functionality “Assay Pending” has been implemented on analysiis results reported since then. Assay results are first loaded to the database with an initial priority indicating that they have not passed through QAQC. When subsequently reviewed for QAQC, the prriority is changed as the result is accepted or rejected. Only dataa with a priority of 1 is visible in the MineSight drill hole views ensuring QAQC of drill data in the model has been approved.
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Data validation occurs during upload of data to database using the acQuire DBMS. Checks include: ¾ All alphanumeric codes (e.g., lithology) are valid and not duplicated ¾ All numeric fields are within acceptable limits and not duplicated ¾ Sample from-to depths cannot be greater than the maximum hole depth ¾ Checks are performed for overlapping samples ¾ Analysis results are received from laboratory in fixed digital format. The load routine imports assay data matching against sample ID created during logging procedure. Alpha analysis codes are stored as logged and/or reported e.g. NS (Not Sampled), IS (Insufficient Sample), <0.01. The database MetaAssayExport table records equivalent values which are substituted by client software (e.g. MineSight). The convention for defined values is a numeric value half detection limit for results at LLD, and for all other codes, -1 is substituted. After data compilation is complete, it is critically reviewed by geologists with on-going scrutiny using logs, section/plan plotting and 3D modeling. After data compilation is complete, a geologist critically reviews it with on-going scrutiny using logs, section/plan plotting and 3D modelling.
11.3 Big Hill Assay QA/QC Process The general process for QAQC for the Big Hill can be divided into two main groups or phases; 1) Historical data including drilling up to and including the 1998 programmes. Standard reference material for QAQC purposes only began at SGM in 2001. Prior to this time duplicates and repeats were the main forms of QAQC. QAQC data prior to 1998 was not captured electronically, corresponding to changing DMBS to acQuire in November 1998. This process is outlined in Section 11.6.1 below 2) Modern 2008 drill programme consisting of 6 Reverse Circulation holes. No QAQC data can be located for this drill programme, and the reason for no QAQC data is unclear, however the operation was going through a wind down phase and it may be QAQC was not completed. These holes have been visually assessed in the sites 3D Minesight software. They appear to correlate well with surrounding holes and do not appear visually to contain any spurious results.
11.3.1 Big Hill Historical QAQC process The following is an excerpt from the “Technical Report for Stawell Gold Mine, Victoria Australia, 28 March 2008” and provides a summary of QAQC undertaken on assaying up to and including 1998 drilling. Repeatability of field sample splitting was tested by re-splitting every 20th sample and submitting this for assay using consecutive sample numbering. The lab was not aware of this practice and in general repeatability was excellent. In addition lab duplicates and repeats of all sample types were tracked. While only July 1998 data is shown here, this data was tracked monthly and showed consistently high repeatability.
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July - RC field replicates
10
1 1st split
0.1 2nd split
0.01 0.01 0.1 1 10 1st split g/t
Figure 11-5 Showing a Log/log graphical representation of the field assay repeats from the July 1998 model update drill holes.
The graph presented is an indication of the homogeneity of the sample after the splitter stage in the field. This graph indicates that the field sample duplicate variation more frequently occurs in the lower grade range, which is of lesser interest. The higher grade duplicates are all relatively constant and within acceptable tolerance.
July 1998, Lab duplicates 100
10 Au1 1
pg Aud 0.1
0.01 0.01 0.1 1 10 100 1st assay g/t
Figure 11-6 Showing a Log/log graphical representation of the monitored lab duplicate results from the July 1998 model update drill holes.
The graph presented is an indication of the homogeneity of the sample after the pulverisation stage, used to analyse the lab duplicate outliers. This graph indicates that the duplicate variation
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STAWELL GOLD MINE more frequently occurs in the lower grade range which is of lesser interest. The higher grade duplicates are all relatively constant and within acceptable tolerance.
Figure 11-7 Showing a Log/log graphical representation of assay repeats which is used to analyse the lab assay repeatability.
Repeated assays are compared to initial assays and as duplicate asssays are a measure of laboratory precision any large discrepanccy is followed up and discussed with the internal and external laboratory managers. This graph indicates that the repeat vaariation more frequently occurs in the lower grade range which is of lesser interest. The higher grade duplicates are all relatively constant and within acceptable tolerance.
11.3.2 Big Hill 2012 QAQC process The general flow sheet for the sample preeparation and assaying including the Quality Assurance and Quaality Control (QA/QC) samples submitted to ensure this compliance is shown in Figure 11-1. ¾ Exploration diamond drill core is routinely Half Core sampled ¾ Mine diamond drill core is mostly Full Core sampled, with approximately 1 in every 5 holes being half core sampled. Samples are crushed and pulverised ¾ Routine assaying of diamond drill ccore samples has been undertaken utilising fire assay methodologies (30 gram charges) with an AAS finish ¾ Screen Fire Assay is also completed on samples in which visible Au was observed during the logging process ¾ Assays are reported to an LLD of 0.01ppm. For sample reporting below LLD, a value of 0.5 x LLD is used in resource estimationn
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¾ Sample QA/QC procedure incorporates routine check assays including repeats (re-assay or repeat assay), duplicate (second sample taken after pulverisation), splits (half coarse sample split, at a ratio of approximately 1:10 samples), and standards ¾ The SGM QAQC process was independently reviewed by Quantitative Group consultants in July 2011 and February 2012. QG found the system in place to be a robust and appropriate process for ensuring quality assay returns (Stewart July 2011 and Stewart Feb 2012).
11.3.3 QAQC Checks and actions A range of checks and resulting actions are in place to monitor the QA/QC of the SGM data set as set out in the QA/QC flow sheet shown in Figure 11-14. When monitoring these checks, the following guidelines are followed:
11.3.4 Standards A range of standards (Table 11-2) are regularly inserted at the sampling stage (1:20 ratio) to monitor assay analysis accuracy. Figure 11-8 shows the assay standard performance for 2012 RC programme used in the November 2012 Resource Model update.
Table 11-2 Range of standards used at SGM
Start/End Standard ID Gating Values Comment Date Used
Lower Recommended Upper Limit Value Limit 2σ 2σ
SGM LowA June 2001 3.18 3.34 3.5 - May 2010
SGM LowB June 2001 3.24 3.54 3.64 Discontinued - no - May 2005 sample left
SGM HighA June 2001 3.885 4.2 4.515 Discontinued - - Oct 2008 performance issues
SGM HighB June 2001 4.18 4.54 4.9 Discontinued - no - Feb 2009 sample left
SGM High Feb 2006 - 8.77 9.36 9.95 May 2010
SGM1 Nov 2008 / 3.49 3.63 3.77 current
SGM2 Nov 2008 / 6.77 7.15 7.53 current
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Start/End Standard ID Gating Values Comment Date Used
SGM3 Aug 2009 / 2.06 2.21 2.36 May 2011
OR2Pd Apr 2011 / 0.058 0.89 0.943 Discontinued - when Jan 2012 supply out, replaced with matrix matched standards
OR54Pa Apr 2011 / 2.68 2.9 3.12 Discontinued - when Jan 2012 supply out, replaced with matrix matched standards
OR15h Apr 2011 / 0.97 1.02 1.068 current
OR10c Apr 2011 / 6.27 6.6 6.92 current
OR62d Apr 2011 / 9.84 10.36 11.16 current
OR12a Apr 2011 / 11.31 11.79 12.27 current
OR17c Jan 2012 / 2.87 3.04 3.21 current
OR15g Jan 2012 / 0.481 0.527 0.573 current
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Figure 11-8 Assay standard performance for Big Hill 2012 RC drill programme and Nov 2012 resource model update.
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If the sstandard assay is outside ±3 standard deviations from the knownn value then ~50% of that batch is automatically repeated, similarly if two standards report outside ±2 standard deviations then ~50% of that batch will be repeated. Entry is made intto the onsite QA/QC diary to indicate which sample IDs have been repeated. The standard performance over time for each standard used for the model update is individually graphed and reviewed. This is to understand if non-complliant standards were well spread across all standard material in use and if it is likely a result of the tightness of the confidence intervals. The spread of the standard results for each standard are reviewed to identifyfy any potential positive or negative bias. Figure 11-9 and Figure 11-10 show examples of the November 2012 model analysis of the standard reference materials OR15h and OR15g respectively.
Figure 11-9 OR15h standard performance for Big Hill 2012 RC drill programme and Nov 2012 resource model update.
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Figure 11-10 OR15g standard performance for Big Hill 2012 RC drill programme and Nov 2012 resource model update.
11.3.5 Lab Duplicates and Repeats Lab splits and duplicates are taken to monitor sample precision. Three sets of duplicate data are collected for SGM samples. ¾ The first duplicate data set is collected at the split stage of the sample preparation, before pulverisation; every tenth sample is split with the split portion becoming the B sample (1:10 ratio). This duplicate is referred to as the lab-split at SGM and monitors the precision of the crush and spliit stages of the sample preparatioon. ¾ The second duplicate data set is collected at the fire assay stage; randomly, within a 11:20 ratio, a second scoop of the pulp is taken and processed in the same batch to the original sample. This duplicate is referred to as the duplicate at SGM and monitors the precision of the pulverisation sammple preparation stage. ¾ The third duplicate data set is collected at the fire assay stage; randomly a second scoop of the pulp is taken and processed in a different batch to the original sample. This duplicate is referred to as the repeat at SGM and monitors the precision of the pulverisation sample preparation stage as well as across batch reepeatability. This type of duplicate can also be requested if the is a non-compliant standarrd within the batch.
Where duplicates and repeats are outside of 10% of the original sample additional duplicate repeats are requested to determine if it is a laboratory issue or associated with coarse gold within the sample. Figure 11-11 shows graphically the hommogeneity of the sample after the pulverization stage for the 2012 Big Hill RC programme. This analysis shows more than 95% of the duplicate
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STAWELL GOLD MINE assays have a relative difference of less than 10% to the original assay value indicating an appropriate level of precision through this stage in the sample preparatioon. Figure 11-12 is a Log/Log graphical representation of the same data which is also used to analyse the lab duplicate outliers. This graph indicates that the dupllicate variation more frequently occurs in the lower grade range (yellow highlight) which is off lesser interest. The higher grade duplicates are all relatively constant and within acceptable ttolerance.
Figure 11-11 Precision of the lab duplicate for Big Hill 2012 RC drill programme and Nov 2012 resource model update.
Figure 11-12 Assay duplicate comparison for Big Hill 2012 RC drill programme and Nov 2012 resource model update.
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Figure 11-13 Log Log graphical representation of the assay repeats form Big Hill 2012 RC drill progrraamme and Nov 2012 resource model update.
Figure 11-13 is a Log graphical representation of assay repeats which is used to analyse the lab assay repeatability. Repeated assays are compared to initial assays and as duplicate assays are a measure of laboratory precision any large discrepancyy is followed up and discussed with the internal and external laboratory managers. Monitoring of these checks is done within two days of the sample batch return and actioned generally no later than seven days after the return date. Any actionns taken during the monitoring process are recorded in the SGM QA/QC diary, which was set up in September 2007 (Figure 11-15).
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Figure 11-14 SGM QA/QC review and actions flow sheet.
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Stawell Gold Mines
QAQC Request Record
Date Requested HoleID Orebody Non-compliant Lab Lab Repeated samples Date prelim Date Final Repeats Date rpts Reviewed comments Action requested Actioned Action Lab Job No. loaded to DB by Re-assay reques by sampleID Name Job No. Re-assays reported Re-assays reported by date 1 stds outside 3SD. Reassays compliant, minor change, 29/03/2012 JT -> JC SP608 Davis L047547 ALS OR12048373 L047529‐L047592 11/04/2012 OR12069201 19/04/2012 JT Change database value to reassay. CD 19/04/2012 17/04/2012 field blanks consistant. 1 standard well outside 3 st. Deviations . Standard compliant on reassay, however their was a moderate 01/05/2012 JT -> JC SP609 Davis L047616 ALS OR12048374 L047594‐L047623 04/05/2012 07/05/2012 OR12090876 10/05/2012 JT Accept re-assays into database CD 10/05/2012 change in the reassay results, but not enough resampled to indicate an ALS fail. 1 standard well outside 3σ. Reassays compliant with 02/05/2012 JT -> JC SP636 Davis L049444 ALS OR12077602 L049425‐L049445 08/05/2012 10/05/2012 OR12090877 10/05/2012 JT Change database value to re-assay CD 10/05/2012 minimal change 1 standard well outside 3σ. Reassays compliant with 03/05/2012 JT -> JC SP626 Davis L049031 ALS OR12084681 L049010‐L049041 07/05/2012 10/05/2012 OR12090878 10/05/2012 JT Change database value to re-assay CD 10/05/2012 minimal change 1 standard well outside 3σ. First reassay batch sent through, with big differences and ALS made comment that there was a sequencing issue so they chose to redo the reassays at their own cost. ALS did the reassay of the Change data base to reassay values for the 04/05/2012 JT -> JC SP623 Davis L048733 ALS OR12066127 L048712‐L048734 09/05/2012 10/05/2012 OR12093820 29/05/2012 JT CD 29/05/2012 whole batch with significant differences showing their was whole batch. a squencing isssue -very good thing that reassay was requested. Laurynda please discuss at next meeting how this occurred. 1 standard well outside 3σ. Reassays non compliant so Do not accept reassays as further reassays 09/05/2012 JT -> JC SP627 Davis L049095 ALS OR12071862 L049110‐L049122 11/05/2012 17/05/2012 OR12099766 13/06/2012 JT CD 13/06/2012 further reassays requested with a standard included requested. Do not accept original either! 1 standard well outside 3σ , 1 standard outside 2σ. Reassay returned compliant results but analysis shows 09/05/2012 JT -> JC SP619 Davis L048255, L048294 ALS OR12066120 L048292‐L048308 11/05/2012 17/05/2012 OR12099765 29/05/2012 JT Change database value to re-assay CD 29/05/2012 that 80% was not with 20% HARD so ALS pays for reassays. Reassays requested as originals were non complaint and I 16/05/2012 JT -> JC SP627 Davis L049095 ALS OR12071862 L049094-L049122 not received 08/06/2012 OR12104192 13/06/2012 JT did not get a standard reassayed so I have no confidence Please accept second batch of reassays. CD 13/06/2012 which set of assays are compliant. 2nd batch compliant
Figure 11-15 QA/QC diary entries for 2012 programme
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SGM Standard Reference Material Eight standard reference materials are currently being used by SGM, and as per standard procedure are inserted at frequencies of 1:20-1:25 samples to monitor accuracy. All of the current standards, except for OR62d are matrix matched. All standards are commercially made by Ore Research and Exploration Pty Ltd (ORE) of Melbourne and certification certificates are available in SGM records. The details of the standards and when they were introduced to the system are shown in Table 11-3.
Table 11-3 SGM assay standards and certified values reported
StandardID Start/End Date UsedGating Values Comment Lower Limit Upper Limit Recommended Value 2σ 2σ SGM LowA June 2001 ‐ May 2010 3.18 3.34 3.5 SGM LowB June 2001 ‐ May 2005 3.24 3.54 3.64 Discontinued ‐ no sample left SGM HighA June 2001 ‐ Oct 2008 3.885 4.2 4.515 Discontinued ‐ performance issues SGM HighB June 2001 ‐ Feb 2009 4.18 4.54 4.9 Discontinued ‐ no sample left SGM High Feb 2006 ‐ May 2010 8.77 9.36 9.95 SGM1 Nov 2008 / current 3.49 3.63 3.77 SGM2 Nov 2008 / current 6.77 7.15 7.53 SGM3 Aug 2009 / May 2011 2.06 2.21 2.36 Discontinued ‐ when supply out, OR2Pd Apr 2011 / Jan 2012 0.058 0.89 0.943 replaced with matrix matched standards Discontinued ‐ when supply out, OR54Pa Apr 2011 / Jan 2012 2.68 2.9 3.12 replaced with matrix matched standards OR15h Apr 2011 / current 0.97 1.02 1.068 OR10c Apr 2011 / current 6.27 6.6 6.92 OR62d Apr 2011 / current 9.84 10.36 11.16 OR12a Apr 2011 / current 11.31 11.79 12.27 OR17c Jan 2012 / current 2.87 3.04 3.21 OR15g Jan 2012 / current 0.481 0.527 0.573
11.4 QAQC Results for the Current Report While historical QAQC practices for the Big Hill area are not as exhaustive as current SGM practices results appear reasonable overall. Further work should be done on uploading to the database the original duplicate and repeated data to allow for further validate of this information. Additionally a number of twinned holes could be undertaken to confirm repeatability of results and original assaying work. The author believes the level of work undertaken is sufficient for the current Resource classification. Stawell Gold Mines has retained the services of several key consulting groups over time to ensure mineral resource estimation processes have been maintained to a high standard. Quantitative Geosciences (QG) has provided ongoing coaching, training, mentoring, and mineral resource estimation services on an as required basis since the late 1990’s. This has included a number of reviews of the Big Hill estimate and ensured consistency of process over this time as key site personnel have changed. ¾ Internal QG Report PSV01810 “Comments on Big Hill Resource Estimate”, September 2007 ¾ Internal QG memo “Review of resource modeling practices, SGM”, July 2011 ¾ Internal QG Report “SGM, Review of Big Hill Resource Estimate, CRG21204”, Nov 2012
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For more recent work, specifically the 2012 RC programme, an analysis of the QA/QC data returned for the current resource estimate has been detail in the above section and summarised below. Table 11-4 This analysis encompasses all QA/QC data returned to SGM for the current period and serves to demonstrate that a responsible and ongoing approach to managing assay data quality is maintained at SGM and that assaying information is of a good quality for Mineral Resource estimation. As a result of QA/QC monitoring processes carried out during the 2012 period, 229 samples were re-assayed from an initial 2,360 assays.
Table 11-4 QAQC assay results
Au OR15h OR15g Population # 58 29 29 # Outside 2σ 12 0 0 % Outside 2σ 21% 0% 0% # Outside 3σ 5 0 0 % Outside 3σ 9% 0% 0% % positive 14% 17% 10% % negative 86% 83% 90% % at zero 0 0 0 Total Bias -3.7% -3.5% -4.3%
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12 DATA VERIFICATION Given the age (dominantly 1997/1998) at which the data was collected the Qualified Persons, have not been personally involved in the collection, QAQC and processing of geological data and information pertaining to this updated estimation of mineral resources. Data for this estimate has been reviewed visually in 3D mine planning software and spot checks have been made to validate the data used. Drilling and sampling methodologies are virtually unchanged to present day and are considered by the Qualified Persons to be robust and aligned with industry best practice. Hard copies of collar survey pickups, down hole survey disks and core photos are all available. QAQC processes, while not as exhaustive as SGM’s current process appears reasonable overall and have been undertaken with the guidance from consulting groups throughout SGM history, maintaining a high standard. To further test the integrity of the information kept within the SGM database, six recent drill holes were randomly selected from the Big Hill resource area: Table 12-1. Checks were made to locate the original data for collar locations, down hole surveys, core photographs, and end of hole and assay records for each of the drill holes.
Table 12-1 Data verification work completed for the 2012 Mineral Resource update
Date Collar Downhole Core Core or Chips Hole Logging EOH Assay Records Drilled Locations Survey Photos available (stored) Davis Area No additional lab info SP210 1997/1998 verified verified NA ‐RC hole yes no verified recorded in database before acQuire Verified ‐ Co‐ SD555 ords noted on Method recorded as F, No Yes ‐ digital additional lab info paperwork as (diamond drilled 1998 verified scanned slides yes yes verified recorded in database tail of SP310) being that of before acQuire SP310 Big Hill South Area No additional lab info SP353 1998 verified verified NA ‐RC hole yes no verified recorded in database before acQuire Yes ‐ digital no – full core No additional lab info SD570 1998 verified verified yes verified recorded in database scanned slides sampled before acQuire Big Hill North Area No additional lab info SP272 1998 verified verified NA ‐RC hole yes no verified recorded in database before acQuire Method recorded as F, No Yes ‐ digital no – full core additional lab info SD543 1998 verified verified yes verified scanned slides sampled recorded in database before acQuire
12.1 Collar Locations All Eastings and Northings entered into the database match the surveyor register exactly. RL was surveyed as AHD, entered into the database as Mine RL which is AHD–300m. Two holes have rounding discrepancy at the second decimal place, see Figure 12-1.
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HoleID Database Calculated from survey register entry
SD555 (SP3310 -24.2 -24.19 collar)
SP353 -9.97 -9.98
Figure 12-1 Rounding discrepancy
12.2 Down hole Survey All original surveys, being Eastman camera discs, have been checked against database records. One disc (SP210, 20m) can no longer be read due to excessive chemical residue. One hole (SP272) has the first survey recorded as 0m, for the other holes in this selection list, the first survey varies, with 10m the minimum.
12.3 Core photos and logging Core photos were sourced and reviewed. All were of a high standard and quality. Logging and geological information was available in the database for all holes.
12.4 EOH No drillers plods were entered into the database pre acQuire. EOH depths of geology and collar have been compared and all match exactly.
12.5 Assay Records Two holes, SD555 and SD543, have method recorded as “F”, no method recorded for other holes. No additional lab info recorded in database before acQuire.
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13 MINERAL PROCESSING AND METALLURGICAL TESTING
13.1 Metallurgical Test Work An extensive test work programme has been completed for both the Big Hill northern and southern (Davis Extension) pit areas. This test work was based on the leaching of a large number of RC drill intercepts during the period of 1998 to early 1999. Electronic versions of this test work do not appear to be available, however, a small subset of this work was summarised for the purposes of evaluating a cut back of the existing Davis Pit in 2005. The remainder of the work appears now to be available as hard copy only.
13.2 Davis Pit Area Test Work Of the original Big Hill test work completed in 1998, a total of 45 intercepts from 27 drill holes were tested from within a later proposed extension of the existing Davis Pit. This subset of the data is summarised in Table 13-1 The test work consisted of basic bottle roll leaching and preg rob testing with several intercepts tested by amalgamation for free gold content. Grind size was as per standard SGM milling output, with all leach conditions standard to SGM treatment of oxide mineralisation types. Overall, the gold recovery from testing averaged 91.7%, with recovery consistent with depth, level of weathering of the rock and mine section. It was concluded that based on the results achieved, a recovery of 90% would be achieved in the processing plant. Upon further review of the test work results, it is recommended that the SGM gravity recovery circuit be in operation whilst treating this mineralisation, as free gold contents ranging from 2% to 48% were seen during testing.
Table 13-1 Davis Pit Test work
SECTION SUMMARY.
METRES INTERVALS GRADE g/t Au PREG REC % SECTION TESTED TESTED CALC COMP ROB O'ALL 248 28 4 1.56 1.56 0.3 89.9 250 56 8 3.09 3.10 0.5 95.2 253 36 7 1.72 1.60 0.3 86.6 255-256 47 7 3.09 2.69 0.5 89.1 259 22 3 2.86 1.74 0.5 92.9 262 47 6 1.75 1.80 0.6 90.6 264 22 4 2.08 2.61 0.7 95.5 267 35 6 1.84 1.82 0.3 90.1 OVERALL 293 45 2.32 2.20 0.5 91.7
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WEATHERING SUMMARY.
METRES INTERVALS GRADE g/t Au PREG REC % WEATHERING TESTED TESTED CALC COMP ROB O'ALL EXTREME - HEAVY 50 8 2.77 2.81 0.3 91.1 MODERATE 179 26 2.13 1.93 0.5 92.7 SLIGHT 43 7 2.41 2.40 0.5 87.1 FRESH 21 4 2.66 2.65 0.7 95.2 OVERALL 293 45 2.32 2.20 0.5 91.7
DEPTH SUMMARY.
METRES INTERVALS GRADE g/t Au PREG REC % DEPTH TESTED TESTED CALC COMP ROB O'ALL 0 - 20 METRES 15 3 1.51 1.54 0.4 78.0 20 - 30 METRES 11 2 1.23 1.16 0.3 90.5 30 - 40 METRES 42 5 1.84 1.81 0.2 93.7 40 - 50 METRES 39 7 3.00 3.09 0.5 94.7 50 - 60 METRES 60 9 2.19 1.83 0.6 93.6 60 - 70 METRES 81 12 2.73 2.57 0.4 92.0 70 - 90 METRES 45 7 2.15 2.11 0.6 86.5 OVERALL 293 45 2.32 2.20 0.5 91.7
13.3 Big Hill Area Test Work The bulk of the test work completed on Big Hill RC drill intercepts is now available in hard copy only. The summary below is extracted from the “Big Hill Project Open Pit Feasibility Study (Draft for Comment)” authored by Karl Guilfoyle and Jeff Moncrieff in May 1999. The works summarised below are inclusive of the “Davis Extension” data discussed above. The testing comprised laboratory leaching of 127 RC drill intercepts, with amalgamation works on 22 of the intercepts. The testing indicated an expected recovery overall of 86.7%, with a breakdown of recovery listed as; Oxide Ore – 92%, Transition Ore – 90% and Sulphide (or fresh) Ore – 84%. The recovery from the Oxide and Transitional components of the ore are comparable to the Davis Extension works above and also to historical treatment of Davis oxide ore types (see below).
13.4 Davis Pit Historical Treatment Davis Pit ore treated prior to September 1993 was blended with all other ore types and no reliable records of grade and recovery can be summarised. In September 1993, the decision was made to campaign mill ores from the two underground mines, the Wonga and the Magdala, as well split the
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ore types from the Magdala to better reflect ore characteristics. From this point Davis ore was treated as a blend with low preg robbing “Basalt” (from the Magdala mine) and “Wonga” ores types, making compilation of results from the Davis Pit possible. Davis Pit ore treated from September 1993 to its completion in February 1996 was 187,176 tonnes at 1.32g/t Au and 92.4% gold recovery. In 2006, a higher gold price led to the treatment of the “Davis Mullock” stockpile over the following 5 years. In all 273,049 tonnes were treated at 0.92g/t Au at 87.0% recovery. These two treatment periods saw the combined treatment of 460,225 tonnes of Davis Pit oxide material at 1.08g/t Au and 89.7% recovery. This result is consistent with expectations for recovery of oxide ores from the Davis Pit area of the Big Hill project determined by the test work programme described above.
13.5 Expected Recovery Based on all test work completed to date and on historical treatment of Davis oxide ore types, a recovery of 90% is considered achievable in the SGM processing plant.
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14 MINERAL RESOURCE ESTIMATES
14.1 SGM Mineral Resource Estimate Dec 2011 Mineral Resources for all SGM areas have previous been reported as at December 2011 in “Crocodile Gold NI 43-101 Report for Stawell Gold Mine, Victoria, Australia, May 2012” and are presented again in Table 14-1 below. Note this report is not re-stating the entire SGM Resource but intends to update the Big Hill section only, stated as Magdala Surface in the table below (or 2,985,000 tonnes at 2.15g/t Au for 206,000 ounces of Indicated Resource).
Table 14-1 December 2011 Final Mineral Resource Calculation for Stawell Gold Mines
Mineral Resources exclusive of Mineral Reserve Indicated Inferred tonnes grade ounces tonnes grade ounces (,000's) g/t Au (,000's) (,000's) g/t Au (,000's) Underground above Magdala 552 3.91 69 509 4.11 67 1250mRL
above Golden Gift 46 3.29 5 243 5.73 45 1650mRL
above Wonga 164 5.83 31 1000mRL Sub-total U/G 598 3.86 74 916 4.85 143 Surface Magdala 2,985 2.15 206 Wonga 149 2.52 12 39 1.57 2 Sub-total 3,134 2.16 218 39 1.57 2 Surface TOTAL 3,732 2.44 292 955 4.72 145
14.2 Introduction and Scope of Big Hill Mineral Resource Estimate Table 14-2 below outlines the Big Hill Mineral Resource estimate for this report, November 2012 and compares it to the previous December 2011 estimate which was a historical estimate carried forward from 1998. A number of key changes have been incorporated into the current estimate including, ¾ Reporting within a pit shell as opposed to the 1998 estimate which reported above the 130MRL ¾ Reporting above a 0.44g/t Au cutoff as opposed to the 1998 estimate of 0.80g/t Au
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¾ Historical voids were reviewed and incorporated into the estimate as modelled voids as opposed to the 1998 estimate which used a number of factors for voids. ¾ Addition of 41 RC holes over the Davis area from the 2008 and 2012 programmes ¾ Updated geological interpretation and geostatistical modelling with a number of unconstrained domains modelled as individual lodes to reduce grade smearing
Table 14-2 Big Hill Mineral Resource Estimate 1998 and November 2012 comparisons
Surface tonnes grade ounces Resource Comments (,000's) g/t Au (,000's) Classification Big Hill area Resource completed 1998, see Previous Magdala “Technical Report for Stawell Gold Mine, (historical 1998 2,985 2.15 206 Indicated Victoria Australia, 28 March 2008” and resource) “Crocodile Gold NI 43-101 Report for Stawell Gold Mine, Victoria, Australia, May 2012” Updated Big Hill & Davis, update void Magdala update 2,830 1.84 167 Indicated models, updated drilling, updated and current report 46 1.15 2 Inferred interpretation and geostatistical modeling, Pit (November 2012) shells utilized for reporting
Drilling utilised in this estimate includes a number of surface programmes targeting Big Hill with the main programme in 1997 and 1998 and underground holes from the upper levels of the underground operation. Over the 2008 and 2012 periods, 589 metres of RC drilling was completed for 7 holes and 2,238 metres of RC drilling was completed for 34 holes, respectively, see Table 14-3.
Table 14-3 Big Hill Drilling Summary
Big Hill Drilling Summary
1997 to 1998 2012 programme 2008 programme Pre May 1997/8 Total programme
# holes metres # holes metres # holes metres # holes metres # holes Metres
Surface Percussion 34 2238 7 589 217 19,537 63 5,916 321 28,280
Surface Diamond 18 1,972 81 * 99 20,370
Diamond Tails 12 397 12 397
145-180 U/G diamond 167 14,920 167 14,920 levels
-212 38 2,242 0 2,242 upholes
66,209
* Although 18,300m of older diamond core drilling falls within the PCF limits, most of this was targeted at deeper sections of the orebody and lies within mine schist west of the ore zone.
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14.3 Big Hill Resource Summary Table 14-4outlines the Big Hill Resource as at November 2012 and represents an update to the previously reported Big Hill Resource, “Technical Report for Stawell Gold Mine, Victoria Australia, 28 March 2008” and “Crocodile Gold NI 43- 101 Report for Stawell Gold Mine, Victoria, Australia, May 2012” , The previous estimate was completed in 1998, with the new estimate incorporating new drilling from 2008 and 2012 RC programmes, an updated geological interpretation and Resource estimation, historical void model review and utilisation of a pit shell at A$1400 gold price for reporting within. The resource is reported above a 0.44g/t Au cut off and is reported below in Table 14-4.
Table 14-4 Big Hill Mineral Resources November 2012 estimate
Indicated Inferred
tonnes grade ounces tonnes grade ounces Surface (,000's) g/t Au (,000's) (,000's) g/t Au (,000's)
Big Hill 2,830 1.84 167 46 1.15 2
14.4 Pit Shell The previous 1998 Resource for Big Hill was reported above the 130mRL bounded to the south by 4900N or mine section 245N and to the north by 5380N or mine section 269N. Reporting within a Pit shell is considered a more robust method for Resource reporting than was previously considered and better meets the requirement for the consideration under NI43-101 for “reasonable prospects for economic extraction”. Pit Shells were used for reporting grade above a 0.8g/t Au cut off and have been constructed by consultants “Mining One” using a gold price of A$1400. Details of assumptions are listed in Section 16 and 22 of this report. Shells have been constrained where possible to within 20m of housing and major infrastructure, with around 51 residences occurring within 100m of the pit boundary. A plan view of the shells is depicted in Figure 14-1 and a long section in Figure 14-2.
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Figure 14-1 Plan view, showing Big Hill pit optimisation (black) used for current Resource estimate reporting and model boundaries (orange) previously used for reporting in 1998
Figure 14-2 Long Section view looking NE, showing Big Hill pit optimisation (black) used for current Resource estimate reporting and 130mRL (orange) previously used for reporting above
14.4.1 Void Modelling The previous 1998 Resource for Big Hill was reported using a number of void factors for different areas based on a visual assessment of know voids and drill holes that encounter voids. These factors were then used to downgrade the resource. Factors by area were, ¾ Mariners 7.5% void ¾ Magdala Flanks 30% void ¾ Iron Duke 2.5% void
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This method can understate the actual depleted material as an average grade is used to deplete the model when in reality it is usually the higher grade zones that have been removed. A number of programmes have been undertaken since 1998 to better define historical and underground voids and resulted in the creation of a 3D void model in 2012. This included, ¾ April 1999, importation and validation of modern development and stoping from the MS-Stopes database above the 212 level. This included all jumbo and air-leg development. Reconstruction of major air-leg, narrow mining and long-hole stopes was undertaken where required. This work identified four major contributors to voids within the model. ‐ Jumbo development on the 145 and 180 levels ‐ 212/282 long-hole stope ‐ 180/272 air-leg stope and associated development ‐ 145/272 narrow long-hole and associated development ¾ April 1999, a data search and reconstruction of development and stoping carried out prior to WMC/SGM mining operations. This work proved extensive workings were historically mined along the ridge and especially in and around the current Davis pit area. However virtually no positional data existed and an interpretation of the likely position and geometry of the most significant working were made using available plans, records and drill information. ¾ June and September 2012 involved the incorporation of the historical work above with the 2008 RC and 2012 RC programmes. Additional voids were modelled for Big Hill or updated based on this information. New voids were based on voids intersected in drill holes and from indications of historical mining. Where additional voids were intersected in the drilling an estimate of the possible stope shapes has been made by joining stopes between intercepts, and projecting ½ drill spacing along interpreted structures. Voids were not projected further than 10m past any particular drill hole intersect. In a number of areas it appeared the original digitized void was larger than originally thought with recent drill intersections. This is most likely due to subsequent failure over time and the original shapes were updated to incorporate the new data. The resulting underground void model was used in conjunction with the historically mined Davis pit void model to deplete the Resource estimate prior to reporting. Figure 14-3 below shows a plan view of the underground void model and Davis pit void model used in the current Resource estimate. A long section is depicted in Figure 14-4.
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Figure 14-3 Plan view showing potential Big Hill pit and underground voids (light green) and Davis Pit void as mined (dark green)
Figure 14-4 Long Section view looking NE, showing potential Big Hill pit and underground voids (light green) and Davis Pit void as mined (dark green)
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14.4.2 Weathering model A model of weathering state was constructed using the following criteria: ¾ Base of total oxidation was defined as the point down hole at which weathering is not completely pervasive of the rock mass and where remnant fresh rock first appears. ¾ Base of transition zone was defined as the point down hole at which oxidised material becomes an insignificant volume component (<1% by volume). In diamond core, these boundaries were identified from logging and core photographs. In RC chips identification was principally made from colour and mineralogy changes. Many of the holes were re- logged specifically for identification of these boundaries. Both surfaces are highly variable in reality due to uneven fracture density and fluid flows, and holes may pass in and out of totally weathered, partially weathered and fresh material. The surfaces are therefore necessarily smoothed.
14.4.3 Bulk Density Specific gravity (dry bulk density) was calculated on the basis of averaged core weights. ¾ All core weighed to +/- 10gm prior to despatch, core weights recorded on sample sheets ¾ Core weights entered to database ¾ Specific Gravity values calculated as SG=measured weight/volume (=length*area*recovery %) ¾ SG values imported to file 11 (bh9811.den) and coded with geology ¾ Mean SG values by geological domain and weathering type reported excluding obvious outlier values (both high and low). Stat reports are appended. ¾ Core weights were of air dried core. Test work shows that in-situ moisture contents vary between 2 and 20%, but that these have uniformly dropped to 0.5-2.5% after air drying. Mean values used in the model are presented in Figure 14-5.
Domain Oxide Transition Sulphide
Mariner’s 2.3 2.5 2.85
Allen’s Constrained 2.15 2.3 2.85
Volcanogenic 2.1 2.3 2.85
Iron Duke Constrained 2.15 2.3 2.85
Volcanogenic 2.1 2.3 2.85
Magdala Constrained 2.1 2.3 2.7
Volcanogenic 2.0 2.3 2.85
Figure 14-5 Bulk density of material
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This method was deemed to be the most effective way of accounting for voids and fractures which cannot be accounted for in wrapped core determinations and which provide the largest variable component in oxidised core.
14.4.4 Moisture Determinations: Moisture determinations performed on RC chips immediately upon delivery to surface showed in-situ moisture contents to vary between 2 and 20%. After air-drying for two days (average exposure of core to air before sampling) moisture contents varied from 0.5-2.5%. In-situ moisture is directly related to clay percentage (and therefore parent lithology) and is not directly linked to depth.
14.4.5 Further Resource Work and Recommendations Ongoing and further work should include, ¾ Further infill drilling should be undertaken in the North pit areas to confirm and constrain void models and improve grade confidence in this area. The Big Hill North area is particularly important as the Resource through this area is a stock-working style of mineralisation and can be discontinuous and variable in nature ¾ Further structural work should be undertaken to understand the structural setting of the Big Hill area and define any un-modelled structures as they relate to the continuity of mineralisation and geotechnical considerations ¾ SG test work should be completed on any further drilling to confirm historical test work. This should be incorporated into the infill drilling and any future geotechnical drilling programmes ¾ Twinned hole analysis should be completed to validate historical assay and sampling results. A number of twinned holes RC to Diamond occur in the project area which this analysis should be undertaken on ¾ Reconciliation on Davis as mined with current Resource model. While this may have to be a broad comparison as detailed production records are not available it would provide a good indication of compliance from model to what was historically mined in Davis
14.5 Geological Modelling Post 1997, 3D block modelling methodologies have progressively become the standard method for estimating Mineral Resources and have provided the basis for detailed mine design and the estimation of Mineral Reserves. This work is carried out in the MineSight software suite which is an industry standard geology and mine planning software package. Detailed Mineral Resource information is reported via an internal document or Resource report and at the time of this report was in progress. The description below describes the general process that is common to each model at SGM and applies to the current Big Hill Resource estimate. Geological modelling is carried out on individual areas by the geological team at SGM for the Davis (South area) and Big Hill (North area) areas of the model. The general process is as follows; ¾ Wireframe models of major geological units are interpreted and created using MineSight software. ¾ Less advanced deposit areas are first interpreted on paper while more advanced deposits are often interpreted directly within the software as is the case for the current Bill Hill model.
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¾ All available geological information is utilised in the interpretation process. ‐ Diamond drill core logs, core photographs, face mapping and photographs, and sludge drilling geology logging. ‐ The drillhole logging information available to the mine geologists includes; ‐ Lithology ‐ Alteration ‐ Quartz veining percentage and veining style ‐ Sulphide percentages, type and style ‐ Location and orientation of lithological contacts, shears and fault structures ‐ Core texture – indicating faulting, shearing etc ‐ Core photographs ¾ Geological modelling is an ongoing process and models are progressively updated to reflect the most up to date information. Model updates are generally made upon completion of infill drilling programmes and completion of development levels, where this information results in a material change in the amount and quality of data involved and/or material changes in geological understanding. This was an update to work previously completed in 1998 for the Big Hill North and South areas and 2009 in the Davis area. The gross geological architecture of the mineralisation systems is well understood and described in detail in Section 7, 8 and this section of the report. Mineralisation is hosted by relatively distinctive and variable predictable geological units that are modeled by the area geologists. The key units that are modeled in each area are; ¾ Magdala Basalt – unmineralised. Geometry of the basalt is required to estimate the quantity of dilution that will be incorporated into the mine designs. ¾ Mine Schist – unmineralised. Also important for estimating the quantity of dilution that will be incorporated into the mine designs. ¾ Weakly mineralised volcanogenic – important to estimate dilution in the mine designs. ¾ Mineralised domains. For individual model areas these may be either in the Central Lode position, Basalt Contact positions and contain zones of stockwork mineralisation. ¾ Key fault structures are modelled as they can have a significant impact on the shape of the mineralised domains (particularly in Mariners domain area). ¾ Wireframes are where possible snapped to drillhole intervals. The key control on mineral resource estimation is accurate definition of the constraining geological models. Estimation of grade within the domains, whilst still very important, is of secondary importance to the first order geological domaining. The data available and used in the modelling included, ¾ Surface and underground drilling as previously discussed ¾ Solos data, recent and old long-holes were imported to a drill hole set and used to help condition the interpretation but were not used for grade interpolation ¾ Channel data, imported and used to help condition the interpretation but not used for grade interpolation
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¾ 1998 Mapping data, ‐ Surface mapping along the length of Big Hill was carried out by Mike Gane who mapped the northern section of the hill (roughly the Davis fence line northwards, while Mike Stewart mapped the southern section. Mapping of the northern section is compiled onto a 1:500 plan, while mapping of the southern section is compiled directly on a 1:1000 aerial ortho- photo. Location of this data is unclear at the time of this report. ‐ Underground mapping was taken from 1:500 scale level plans and sections. Much of this mapping was carried out during Western Mining ownership. Quick checking of some features was undertaken but no significant remapping was carried out underground.
14.6 Big Hill Summary of Geological Modelling The package of Magdala Volcanogenics, as a whole, plays host to the bulk of mineralisation. Where continuous mineralised structures can be defined with a reasonable degree of confidence, these structures have been constrained within wire-frame solids. Where it has not been possible to confidently define the geological control to mineralisation this material has been left unconstrained. Table 14-5 along with Figure 14-6 to Figure 14-8, outline the solids that have been used in construction of the 3D solid model and for coding of geology to the 3D Block Model.
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Table 14-5 Geological Domains at Big Hill 2012
Geology Wireframes Geology Final Solids Domain Block modelBlock % SLOPE Au Item FAULTS Fault 2 (scotchman) 11_Fault SM 11 FF%FAU do not use this fault domain in reserve calculatoins Fault 3 - no volume Fault 3 Fault 4 - no volume Fault 4 Fault cc Lower - no volume Fault cc Lower Fault cc Upper - no volume Fault cc Upper Fault Scotchmans Splay -no volume Fault Scotchmans Splay
BIG HILL Iron Dukes 107_Iron Duke 107 BH BH% BHAU subdomained into 171, 172, 173, 174 Allens 108_Allens 108 BH BH% BHAU subdomained into 181, 182, 183, 184, 185, 186, 187, 188, 189 Mariners 109_Mariners 109 BH BH% BHAU Mariners Lower1 109_Mariners L1 109 BH BH% BHAU Mariners Lower2 109_Mariners L2 109 BH BH% BHAU Mariners Lower3 110_Mariners L3 110 BH BH% BHAU Mariners Splay 111_Mariners Splay 111 BH BH% BHAU LG Volc Allens 501_LG Volc Allens 501 LL%LAU LG Volc Iron Duke 502_LG Volc Iron Duke 502 LL%LAU LG Volc Mariners 503_LG Volc Mariners 503 LL%LAU LG Volc 1 503_LG Volc 1 503 LL%LAU LG Volc 2 503_LG Volc 2 503 LL%LAU LG Volc 3 504_LG Volc 3 504 LL%LAU
DAVIS Domain - Brown 601 Brown 601 D D% DAU Domain - Orange 602 Orange 602 D D% DAU Domain - Blue 603 Blue 603 D D% DAU Domain - Green 604 Green 604 D D% DAU Domain - Purple 605 Purple 605 D D% DAU Domain - Grey 606 Grey 606 D D% DAU Domain - Teal Links 607 Teal Links 607 D D% DAU subdomained into 671, 672, 673, 674 Domain - Stockworks cyan 608 cyan 608 D D% DAU Domain - North Yellow 610 Yellow 610 D D% DAU Domain- Pink Stocks 611 Pink Stocks 611 DD%DAU LG Volc Davis 505_LG Volc Davis 505 LL%LAU
BASALT WASTE Basalt Above Scotchmans Basalt - Above SM 4 B B% BAU Basalt - Lower Basalt - Lower 4 B B% BAU Basalt - Upper Basalt - Upper 4 B B% BAU
topo
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Figure 14-6 Section 5690 through Big Hill North, showing the Cross Course Fault (blue), Mariner’s domain (green) and Allen’s domains (pink).
Figure 14-7 Section 5510 through Big Hill South, showing Scotchman’s Fault (upper blue surface), Iron Duke domains (purple), Magdala domains (red) and basalt waste domain (yellow).
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Figure 14-8 Section 5090 through Davis area, showing Magdala domains (purple) and basalt waste domain (yellow). Note the previously mined Davis depletion.
14.6.1 Block Modelling The process of block modelling at SGM is relatively standard across all areas. Stawell Gold Mines has retained the services of several key consulting groups over time to ensure mineral resource estimation processes have been maintained to a high standard. Quantitative Geosciences (QG) has provided ongoing coaching, training, mentoring, and mineral resource estimation services on an as required basis since the late 1990’s. This has ensured consistency of process over this time as key site personnel have changed and supported by the Qualified Person.
14.6.2 Block Model Dimensions Block model dimensions vary on an area by area basis as indicated in Table 14-6. Block model dimensions are varied based upon the density of the available drilling data and also on the overall geometry of the mineralised structures.
Table 14-6 Block sizes utilised in SGM local area block models
Model Limits
Number of Block size Composite Model Method X min X max Y Min Y max Z min Z max Blocks (x- (x-y-z) Length y-z)
Big Hill November 2012 OK 5m-10m-5m 2m 4500 5000 4400 6000 -220 20 100-160-48
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14.6.3 Block Model Coding Block models are coded with the key structural/mineralisation domains. This process is completed using the MineSight software. The details of the coding also vary by Mineral Resource area and are determined based on the number and geometry of the mineralisation domains. Given the block sizes utilised by SGM relative to the individual structures, multiple domain codes and percentages are stored in each block.
14.6.4 Drill hole Coding A diamond drill hole set is coded with the key structural/mineralisation domains. Importantly this coding is checked manually by the area geologists to ensure that they match the wireframes created. Where required some manual adjustment is made and if suspect drill hole locations are noted. These drill holes may be excluded from the estimate. These details are included in the documentation for each Mineral Resource area
14.6.5 Compositing The drill hole files are composited down hole to fixed length composite intervals. The composites are matched to honour the geological domains as coded in the drill hole files. Composite intervals appear in Table 14-6. The choice of composite intervals is made on a model by model basis that will best reflect the block size, data available and geometry of domains.
14.6.6 Geostatistical Parameters All Mineral Resource areas modelled in 3D have gold grades (Au ppm) estimated by Ordinary Kriging. Over time this methodology, when coupled with detailed and robust geological models, provided reliable estimates of in situ gold grade. The key geostatistical parameters are modelled for each project separately. Variography studies are completed using MineSight MSDA software. Individual variogram studies are conducted for each domain and modelled separately. A summary of these parameters is presented in Table 14-7 Key variogram parameters, nugget and sill, and variogram ranges are modelled on either normal variograms or Gaussian transformed data. Whilst the estimated values for the nugget and ranges vary for each modelled area they are generally relatively consistent. The variograms studies have been performed to optimise the kriging neighbourhood. This enables quantitative evaluation of the results of the kriging to be performed and, as well as enabling the search neighbourhoods to be optimised, provides numerical outputs from the kriging runs (Slope of Regression of the Estimate), a summary of these parameters is presented in Table 14-8. The Slope of Regression is used in part to aid in classification of the Mineral Resource estimates as per the definitions contained within NI 43-101.
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Table 14-7 Big Hill variography summary REA A DOM axis 1 axis 2 axis 3 Nugget pri. axis (x) sec. axis (y) ter. axis (z) variance pri. axis (x) sec. axis (y) ter. axis (z) variance Total Variance Nugget rotation in MineSight
DAVIS 601 315 0 52 2.6 8 8 2 0.9 30 30 20 1.4 4.90 53% DAVIS 602 330 0 41 2.5 12 12 8 2.8 30 30 20 1 6.30 40% DAVIS 603 330 34 53 1.7 10 10 2 0.9 50 30 20 1.4 4.00 43% DAVIS 604 326 0 44 2.6 8 8 2 0.9 30 30 20 1.1 4.60 57% DAVIS 605 317 0 36 2.6 8 8 2 0.9 30 30 20 1.1 4.60 57% DAVIS 606 291 0 46 0.45 8 8 4 0.3 40 45 10 0.3 1.05 43% DAVIS 671-674 300 0 25 2.6 8 8 2 0.9 30 30 20 1.1 4.60 57% DAVIS 608 310 0 67 2.6 8 8 2 0.9 30 30 20 1.1 4.60 57% DAVIS 610 326 0 46 2.8 30 25 20 2.5 5.30 53% DAVIS 611 325 0 45 2.6 2.6 8 8 2 0.9 30 30 20 24.60 11% DAVIS 505 315 0 52 2.6 2.6 8 8 2 0.9 30 30 20 24.60 11%
BIG HILL 171-174 320 0 40 2.69 8 8 4 0.56 30 30 10 0.56 3.81 71% BIG HILL 181-189 305 0 -85 1.2 10 10 4 1.8 30 30 10 1.7 4.70 26% BIG HILL 109 347 0 46 3 8 8 3 2.5 50 50 25 1.9 7.40 41% BIG HILL 110 326 0 71 3 8 8 3 2.5 50 50 25 1.9 7.40 41% BIG HILL 111 365 0 70 3 8 8 3 2.5 50 50 25 1.9 7.40 41% BIG HILL 11 76 0 28 0.5 50 40 10 0.7 1.20 42% BIG HILL 501 305 0 -85 1.2 10 10 4 1.8 30 30 10 1.7 4.70 26% BIG HILL 502 320 0 40 2.69 8 8 4 0.56 30 30 10 0.56 3.81 71% BIG HILL 503 347 0 46 3 8 8 3 2.5 50 50 25 1.9 7.40 41% BIG HILL 504 319 0 46 3 8 8 3 2.5 50 50 25 1.9 7.40 41%
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Table 14-8 Big Hill Search Parameter summary hole hole max axis 1 1 axis 2 axis 3 axis AREA search search search distance distance DOMAIN DOMAIN Quadrant Quadrant Search Type Search Type outlier restriction outlier restriction min. samples 1st min. samples Outlier restriction Outlier restriction max. samples 1st max samples per samples max max. samples per x axis - 1st search search - 1st x axis search - 1st y axis search - 1st z axis
search distance rotation sample number definition
DAVIS 601 75 75 20 300 0 52 5 32 4 DAVIS 602 75 75 20 330 0 41 5 32 4 DAVIS 603 75 75 20 306 38 60 5 28 4 10 10 DAVIS 604 75 75 20 326 0 44 5 32 4 DAVIS 605 75 75 20 317 0 36 5 32 4 DAVIS 606 75 75 20 291 0 46 5 32 4 DAVIS 671 50 75 20 300 0 49 5 32 4 DAVIS 672 75 75 20 320 0 25 5 32 4 DAVIS 673 75 75 20 320 0 25 5 32 4 DAVIS 674 75 75 20 320 0 25 5 32 4 DAVIS 608 75 75 20 310 0 67 5 32 4 DAVIS 610 75 75 20 326 0 46 5 32 quad 9 4 DAVIS 611 75 75 20 325 0 45 5 32 4 DAVIS 505 75 75 20 316 0 52 5 32 4
BIG HILL 171 75 75 20 320 0 40 5 32 4 BIG HILL 172 75 75 20 320 0 40 5 32 4 BIG HILL 173 75 75 20 320 0 40 5 32 4 BIG HILL 174 75 75 20 320 0 40 5 32 4 BIG HILL 181 75 75 20 305 0 -90 5 32 4
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BIG HILL 182 75 75 20 305 0 -83 5 32 4 BIG HILL 183 75 75 20 305 0 -83 5 32 4 8 10 BIG HILL 184 75 75 20 305 0 -83 5 32 4 7 10 BIG HILL 185 75 75 20 305 0 -83 5 32 4 BIG HILL 186 75 75 20 305 0 -90 5 32 4 BIG HILL 187 75 75 20 305 0 -75 5 32 4 BIG HILL 188 75 75 20 305 0 -75 5 32 4 BIG HILL 189 75 75 20 305 0 -85 5 32 4 9 10 BIG HILL 109 75 75 20 347 0 46 5 32 4 BIG HILL 110 75 75 20 319 0 46 5 32 4 BIG HILL 111 75 75 20 365 0 70 5 32 4 15 10 BIG HILL 11 75 75 20 76 0 28 5 32 4 BIG HILL 501 75 75 20 305 0 -85 5 32 4 BIG HILL 502 75 75 20 320 0 40 5 32 4 BIG HILL 503 75 75 20 347 0 46 5 32 4 BIG HILL 504 75 75 20 347 0 46 5 32 4
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14.6.7 Resource Classification The classification of Indicated and Inferred material for Big Hill is based on geological confidence, Slope of Regression analysis and model validation results. The practice adopted at SGM uses general guidelines for classification that utilise the following information; ¾ Drilling density ¾ Demonstrated geological continuity of structures and mineralised domains ¾ Slope of Regression of the Estimate analysis (calculated value during the Kriging Process) ¾ Average distance statistics of composites used in the block estimation is within the modeled variogram ranges Classification as applied to the Mineral Resources disclosed in this document have been reviewed in detail by internal SGM personnel and external consultants and are considered to be appropriate and within the guidelines.
14.7 External Factors Effecting Extraction of Mineral Resources and Reserves
14.7.1 Magdala Surface Mineral Resources The Magdala Surface Mineral Resources are located within a segment of Crown Land called the Big Hill Ridge which defines the up-dip portion of the Magdala mineralised system. This area is bordered by residential areas on two sides. An extensive RC and Diamond drilling programme was completed during 1997 and 1998 to define the extent of the mineralised system and define the Mineral Resource. A study to establish an open pit mining operation on this section of the deposit was completed and a detailed Environmental Effects Statement (EES) prepared. The EES was submitted to the Victorian State Government in 1999 seeking approval to commence a mining operation within the Big Hill area. In November 2000 the proposal was rejected by the Minister for Planning and the project was not able to proceed due to ministerial concern in the areas of social impact, heritage, unfilled mining void, habitat impact and a tenuous link between the proposed project of the time and continuation of underground mining operations. Stawell Gold Mines believes there is strong potential to re-examine the project with all areas of ministerial concern addressed in present project consideration. The methodologies that were proposed at the time and the current review and proposal are defined and reported in this document. From 2007 to the present the waste dump created from the initial mining of the Davis open pit was progressively reclaimed and treated through the SGM processing facility. As such there is a precedent for mining operations in this area. Whilst there is currently proximity constraints to the Stawell community, which restricts open pit mining of the mineral resource in full, there is potentially significant economic benefit if this resource could be exploited by a combination of surface and underground mining activity. Extensions of mineralisation and resource beneath the designed pit will be evaluated as part of the ongoing underground operation for full economic benefit realisation of the proposed project.
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15 MINERAL RESERVES ESTIMATES
The current level of study is of insufficient detail to support a Mineral Reserve Estimate. Mining One Pty Ltd (Mining One, M1) was approached by Johan Booyse, Mining Manager, of Stawell Gold Mines (Stawell) to carry out a scoping level study involving a pit and dump design on the Davis and Big Hill deposits, to provide a preliminary economic analysis. M1 conducted a ‘Whittle’ pit optimisation based on the input parameters and block model data provided by Stawell. The gold price considered was A$1,400 per ounce. Due to community proximity constraints, a pit optimisation has been conducted for 3 (three) cases considering relative impacts on the local community and the environment: ¾ 20Mt Case ‐ This case refers to the option with the highest potentially economic pit mining case while honouring proposed stand-off distances and maximising any potential to extract resources using a combination of underground and open pit methods in the base of the northern and southern pits. ¾ 10Mt Case ‐ This case further considers environmental and community impacts by targeting higher margin material, preferentially mining “more profitable targets”. The result is a reduced pit size, shorter mine life and ultimately a less impact on the community and the environment. It may also be possible to utilise a combined open pit and underground mining method in the base of the southern pit to improve resource extraction. ¾ 100m buffer Case ‐ This case focuses on the application of a minimum 100m stand-off distance from mine workings to the community. This option then explores any remaining opportunity. The results of the PEA are discussed in Sections 16-22 of this report. As this is a preliminary economic assessment and not a prefeasibility study, no Mineral Reserves are presented. This technical report provides a Preliminary Economic Assessment (PEA) of the Big Hill Project of SGM. It outlines a possible open pit operation. This PEA is preliminary in nature, it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorised as mineral reserves, and there is no certainty that the PEA will be realized.
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16 MINING METHODS
16.1 Underground Mining Methods The Magdala Mine is accessed by a decline from a portal located adjacent to the mill. The mine access development and services are located mainly within basalt. Ground conditions are good and there is no history of major seismic activity. Development follows the Magdala Lode system down plunge and between 470RL and 786RL the decline splits into a north and a south decline to access the Golden Gift ore bodies. To facilitate ore access, extraction levels are developed at approximately 20m to 25m vertical intervals. The mining areas currently extend over approximately 3km of strike to more than 1,600m below surface, measured from the top of Big Hill. The mining method used in the Magdala mine was bench stoping with cemented rock fill pillars in primary stopes, and rock filled secondary stopes. The mining method used in the Golden Gift and narrow Magdala ore zones is retreat open stoping with either cemented rock fill (CRF) if full extraction or combinations of CRF and rock fill or all rock fill stope with pillars. In the Magdala ore body, stope sizes typically range from 2,000 to 10,000 tonnes. In the Golden Gift area where the width and tenor of the reserves have so far been of higher quality, generally larger stopes up to 15,000 tonnes have been mined Stope ore is recovered using loaders under direct or remote control of an operator, with haulage by 60 tonne trucks. The access decline is used as an intake airway and a chilled water plant delivers conditioned air via an intake shaft. Local spot refrigeration plants are also used for decline advance. Exhaust air is drawn through the workings by a series of ventilation rises and drives by two primary ventilation fan installations located at the northern and southern ends of the mine. The mine is relatively dry. Water pumped from the workings is recycled for use in the mine or the treatment process.
16.1.1 Open Pit Methods Open pit mining will utilize conventional benching techniques of drill and blast or free dig in the oxidized zones. Loose overburden will be removed with either a combination of ripping and excavator or conventional drill, blast, excavate and truck. The discrete nature of open pits will necessitate the use of small scale equipment. Ramp widths are kept to a safe minimum of one lane at 12 m as the mining fleet will consist of only a few trucks. Ore will be hauled to the mill during day light hours, 5 days per week. Mine design assumptions consider a ramp gradient of 1:9 (with a final ramp section at 1:8), a berm interval of 20m vertical metres has been assumed. Batter slope angles are typically 50 degrees in weathered material and reach 60 degrees in fresh rock. Batter slope angles will be determined on a location basis, using geotechnical advice determined from diamond drill holes and test work data. Lerchs Grossmann pit optimisation modelling of the resource is used as a guide to a potential economic pit shape, with mining recovery factors of 100% and a dilution factor of 20% assumed. Work shifts will be typically Monday to Friday during day light hours to minimize impacts to the surrounding community and environment.
16.1.2 Underground Access from Open Pit In order to maximise the potential to extract resources, a combination of open pit and underground mining methods has been considered. In order to reduce the extent to which
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16.2 Whittle Optimisation of Open Pit
16.2.1 Block Model Preparation The block model data was received from Stawell in November 2012. The data was imported into Datamine Studio 3 software. Further modifications to the block model were applied based on the latest mine topography and underground void information provided by Stawell site personnel. These modifications were: ¾ Application of provided void and fill information; ¾ Addition of community buffer zones; ¾ Inclusion of additional attributes for import into Whittle 4X including; ‐ SG – addition of SG to the entire model; ‐ Cost adjustment factors; and ‐ Mineralised zones. The geometry for the final combined block model is shown in Table 16-1 below:
Table 16-1 Block Model Geometry
Parameter Min Max Block dimension Sub block # Blocks
X Origin 4,500 5,000 5 - 100 Y Origin 4,400 6,000 10 - 160 Z Origin -220 20 5 - 48
The additional attributes are shown in Table 16-2 below:
Table 16-2 Additional Attributes
Variable Default Type Description
MCAF 1 numeric Mining cost adjusted factor value
PCAF 1 numeric Processing cost adjusted factor value
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16.2.3 Optimisation Methodology The ‘Whittle Optimisation’, process was generally conducted as illustrated in Figure 16-1 The ‘Whittle Optimisation’ process involves the evaluation of a deposit by considering a potential economic pit. The shell is generated from the in situ resource block model considering mining, geometric and economic parameters to estimate a potential economic pit. The overall slope angle is used to approximate the slope angle used in Whittle to determine an ultimate pit shell. This angle is represented by a derivative of block model parent cell blocks and is therefore only an approximation of the slope angle provided.
Davis & Big Hill ‐ OPTIMISATION PROCESS OVERALL MILL PROCESSING SLOPE (deg) COST ($/t‐milled) DISCOUNT RATE (%) MINING RECOVERY SELL PRICE ($/UNIT MINING/PROCESSING MINING COST (%) ADMIN PROCESSING PRODUCT) RATE (MTPA) MINING RECOVERY (%) DILUTION PARAMETERS ($/t‐mined) (%) COST ($/t‐milled)
WASTE
MINING OPTIMISATION
WHITTLE PROCESSING TOTAL COST PIT SELECTION ORE CONCENTRATE SALES NPV PLANT REVENUE CASH FLOW Figure 16-1: Davis and Big Hill Optimisation Process
Whittle uses the Lerchs Grossman algorithm optimisation technique to generate a series of nested shells based on the aforementioned parameters. These nested pits are generated considering a range of revenue factors or gold prices, in this instance, to determine pits of lower to higher cost per ounce of gold. This is done on a current value basis which provides a series of nested shells for further analysis. Subsequent analysis of these shells is carried out to determine which pit may provide the best economic result for a given set of economic conditions, such as the mining sequence, mining and milling rates, bench advance rates, the appropriate discount rate, etc. This analysis will determine the best of the aforementioned shells for a given set of economic and mining criteria. A practical mine and dump design has been generated for this study. The mining cost has not been re-estimated based on the mine schedule and mine design (i.e. original Whittle Mining Cost Adjustment Factors have been assumed to cost the schedule).
16.2.4 Optimisation Parameters A summary of the Whittle optimisation input parameters is presented in Table 16-3.
A base cost of A$3.40/t has been assumed and depth penalty cost adjustment factors applied in the Whittle optimisation. These mining costs were estimated using costs from a similar project of size and nature. The mining cost includes contract mining, grade control, load and haul and mining related site operational staff. It is Mining One’s opinion that the average total mining costs would be expected to be in the range of A$3.30/t - $4.50/t.
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Table 16-3 Summary of Optimisation Parameters INPUT PARAMETERS Value Unit Notes Sources BLOCK MODEL INPUTS Density SG t/m3 Specific Gravity Stawell Au Grade AU g/t Au grade Stawell Zone ZONE numeric Stawell Material Classification WTYPE alpha Stawell Mining Cost Adjustment MCAF numeric M1 Factor Processing Cost Adjustment PCAF numeric M1 Factor GEOTECHNICAL/PIT PARAMATERS Ramp Width 12 metre M1 Ramp Grade 1:9 M1 Batter High 20 metre M1 Batter Berm 5 metre M1 Batter Slope – All area 50 degrees M1 Overall Slope – All area 41 degrees M1 MINING PARAMETERS Mining Recovery 100 % Stawell Dilution 20 % Stawell $/tonne M1 Mining Cost 1.0 Base Mining Cost mined Estimate M1 estimate for $3.60 Mining Cost Adjustment M1 – 3.80 /t total mining Factor (Depth Penalty) Estimate cost PROCESSING PLANT PARAMETERS $/tonne Excluding sustaining Total Processing Cost 15.00 Stawell milled capital Mill Recovery Au 90 % Stawell SCHEDULE PARAMETERS Mining Limit 4,600,000 tpa Stawell FINANCIAL PARAMATERS Sell Price A$ 1,400 $/ounce Stawell Royalty $2 $/ounce Stawell Discounted Rate 10 % M1 CONVERSION FACTORS ounces ‐> grams 31.10348
16.2.5 Whittle Optimisation Results Considering the proximity of the Big Hill open pit project to the local community and infrastructure, the decision regarding the selection of the ultimate pit shell includes, but is not limited to, the following criteria:
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¾ Sustaining Capital Cost ¾ Environmental rehabilitation costs ¾ Potential Cash flow ¾ Net present value ¾ Internal rate of return ¾ Project duration, including rehabilitation ¾ Open pit foot print / open pit size ¾ Proximity to community and infrastructure Three (3) potential cases were selected for design and assessment. These cases were selected based on maximum cash flow potential, community impact (100m buffer case) and to maximise internal rate of return. The cases are referred to as follows: ¾ Maximum cash flow potential – 20Mt Case ¾ Maximum IRR potential – 10Mt Case ¾ Reduced community impact - 100m Buffer Case As can be seen in Figure 16-2 below, the pit size for the maximum cash case is approximately 17.5Mt. The 10Mt pit was selected due to both the maximum cash and pit size variations between the selected pit and the pit which is immediately smaller. This selected pit yields a significant maximum cash amount ($100M) and also has a designable pit size. The 100m Buffer case has been based on open pit foot print restrictions, and the pit contains approximately 11Mt or material. The Whittle Optimisation is presented in Figure 16-3 below.
20Mt Case
10Mt Case
Figure 16-2: Selected Whittle Shells, 20Mt Case and 10Mt Case
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100m Buffer
Figure 16-3: Selected Whittle Shells, 100m Buffer Case
The 20Mt case pit shell (not designed) results in an undiscounted operating cash flow of approximately $116M and contains approximately 3.7Mt of mineralised material with an average grade of 1.59g/t. The 10Mt case pit shell results in an undiscounted cash flow of approximately $100M and contains approximately 2.5 Mt of mineralised material with an average grade of 1.71g/t. The 100m Buffer case pit shell results in an undiscounted cash flow of approximately $59M and contains approximately 2.4 Mt of mineralised material with an average grade of 1.40g/t. The results of this pit optimisation are shown in Table 16-4 below. The Whittle optimisation pit shells are presented in Figure 16-3, Figure 16-7 and Figure 16-9 below. Note that these shells are not pit designs and are yet to consider haul ramps and minimum mining widths.
Table 16-4 $1,400/ounce - Optimisation Parameters and Results, 3 cases Description Unit 20Mt Case 10Mt Case 100m Buffer Case Gold Price A$/ounce 1,400 1,400 1,400 Royalty % revenue 0 0 0 Wall angle All area 41° 41° 41° Mining Dilution % 20 20 20 Mining Recovery % 100 100 100 Processing Recovery % - All 90 90 90 Processing Cost Total ($) -15.00 -15.00 -15.00 Mineralised Tonnes 3,731,028 2,455,936 2,464,949 Au Grade g/t 1.59 1.71 1.40 Waste Tonnes 13,936,475 6,564,192 9,196,734 Total Rock Tonnes 17,667,503 9,020,128 11,661,683 Recovered Ounces Ozs 171,656 121,520 99,855 Stripping Ratio 3.73 2.67 3.73 Total Undiscounted A$ 116,236,981 100,709,465 59,792,190 Value
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Figure 16-4: Whittle Shell A$1,400/oz – 20 Mt Case Dimensions
Figure 16-5: Whittle Shell A$1,400/oz - 20Mt Case (Looking North-East)
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Figure 16-6: Whittle Shell A$1,400/oz – 10 Mt Case Dimensions
Figure 16-7: Whittle Shell A$1,400/oz - 10Mt Case (Looking North-East)
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Figure 16-8: Whittle Shell A$1,400/oz – 100m Buffer Case Dimensions
Figure 16-9: Whittle Shell A$1,400/oz – 100m Buffer Case (Looking North-East)
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16.3 Pit and Dump Design The pit has been designed to comply with the overall slope criteria, while at the same time incorporating Whittle optimisation factors in order to minimise waste tonnes and maximise mineralised material. The pits have been designed using the following specifications: ¾ 12 m wide single lane ramps with a gradient of 1:9 ¾ A batter height of 20 m with a batter angle of 50º ¾ A berm width of 5 m These design parameters assumed 90t haul trucks and a 100t excavator for use as the primary mining fleet. Following the inclusion of the haul ramps into the pit design, the mining inventory was determined. The inventory is summarised in Table 16-5. The conceptual pit designs are presented in Figure 16-13, and Figure 16-15. Note that these conceptual designs include integrated haul ramps.
Table 16-5 Pit Design Inventory
Description 20Mt Case 10Mt Case 100m Buffer Case
Mineralised Tonnes 3,470,546 2,277,200 2,317,587 Au Grade (g/t) 1.52 1.65 1.39 Waste Tonnes 15,892,388 7,825,097 9,446,563 Total Tonnes 19,362,934 10,102,297 11,764,150 Recovered Ounces 152,642 108,722 93,215 Stripping Ratio 4.5 3.4 4.07
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Figure 16-10: Pit Design A$1,400/oz – 20Mt Case Dimensions
Figure 16-11: 20Mt Pit Design & Mineralised Material (Looking North-East)
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Figure 16-12: Pit Design A$1,400/oz – 10Mt Case Dimensions
Figure 16-13: 10Mt Pit Design & Mineralised Material (Looking North-East)
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Figure 16-14: Pit Design A$1,400/oz – 100m Buffer Case Dimensions
Figure 16-15: 100m Buffer Pit design & Mineralised Material (Looking North-East)
The dumps have been designed based on environmental and community restrictions and with the intention of restoring the Big Hill to its original landform. The bulking factor chosen for dumped waste material is 1.15 based on Stawell Gold Mine site experience. The dumps have been designed using available data related to material properties based on experience
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STAWELL GOLD MINE from other mining areas. Some dumping will occur in the Wonga Pit (previously mined pit). The volume of material dumped into the Wonga Pit is dependent on the case assessed. The Wonga pit backfill is not shown in the figure below. The basic design parameters are: 12 metre wide single lane ramps with a gradient of 1:9 A batter height of 20 metres The location of the conceptual dump design is presented in Figure 16-16 below.
Temporary Waste Rock Stockpile
Figure 16-16 Dump Design and location
16.4 Pit Schedule For pit scheduling purposes, Davis and Big Hill deposits were divided into separate mining areas described as either “Big Hill, Middle and Davis” or “North and South”, depending on the size of the final pit. Where it was not practical to separate the pit into mining areas, a single inventory is presented. A summary of the pit inventories of each mine area for each case is presented in Table 16-6.
Table 16-6 Pit Inventories of Mines
20Mt Case 10Mt Case 100m Description Buffer Big Hill Middle Davis North South Case
Mineralised Tonnes 933,615 1,622,939 913,992 1,055,007 1,222,193 2,317,587 Au Grade (g/t) 1.79 1.48 1.33 1.87 1.44 1.39 Waste Tonnes 2,100,269 8,138,650 5,653,469 2,165,159 5,659,938 9,446,563 Total Tonnes 3,033,884 9,760,589 6,567,461 3,220,166 6,882,131 11,764,150 Recovered Ounces 48,356 69,502 35,174 57,086 50,926 93,215 Stripping Ratio 2.25 5.01 6.18 2.05 4.63 4.07
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High level schedules were conducted based a North to South mining sequence for each of the cases. The North to South Mining sequence was selected considering community impact, waste dump sequencing and minimising negative cash drawdown. The aforementioned scenarios assume a 5 day per week, day shift only roster and a maximum bench turn-over rate of 80 m per year. The mine life is anticipated to be in the order of 4 to 5 years and is assumed to use a mining fleet consisting of between one and three 100t excavators and between eight to ten 90t trucks. Some drill and blast is anticipated, however, a significant portion of the material is expected to be “free dig”. Existing underground workings in the area are expected to effectively “dewater” the pits although some perched water is anticipated. The larger underground workings are generally well known from historic plans and subsequent drilling in the area. These voids will be targeted to source further information and managed as part of operations. Mining operating costs have been benchmarked against the costs of operating mines of a similar size and nature. The restricted hours of operation (5 days per week / daylight hours only) is expected to attract a cost premium, this is yet to be quantified. However, the estimate used here is believed to be indicative and aligned with the intended targeted study accuracy. The cut-off grade applied (0.44g/t) in this instance is estimated based on administration and processing costs only. Sustaining capital for the mill has not been included to estimate the cut-off grade. This is partly due to the underutilisation of the mill. Current plans indicate that there is excess capacity in the mill and the plan assumes that the decision to continue operations is given. Therefore, while there continues to be excess capacity in the Mill, any grade above a cut-off grade, based on variable costs only, will contribute to any fixed operating costs or sustaining capital requirements. A quarterly mine schedule with associated cash flow estimated has been generated to outline anticipated cash flow drawdown commitments. A discount rate of 10% has been used in the mine schedules.
16.4.1 20Mt Case High level schedules were completed for the 20Mt case. The key physicals and anticipated cash flow for the base case is summarised in Table 16-7 Table 16-7 Summary of Key Physicals for 20Mt Case
Description Unit Inventory Mill Feed t 3,470,546 Input Grade Au g/t 1.52 Mill Input Ounces Au oz 170,530 Mill Recovered Ounces AU oz 152,642 Mining Cost $ -73,371,063 Mill Cost $ -52,058,161 Rehandle Cost $ -20,788,996 Revenue Au $ 214,867,986 Undiscounted Cash flow $ 44,685,448 Discounted Cash flow (Inclusive capital) $ 36,748,767 IRR % 664
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20Mt Case Schedule: The following key physicals are presented in their respective figures: ¾ Quarterly mill feed tonnes (t) and grade (g/t Au) in Figure 16-17; ¾ Quarterly total open pit movement (t) in Figure 16-18; ¾ Quarterly gold ounces mined from each pit (oz) in Figure 16-19; ¾ Quarterly open pit movement from each pit (t) in Figure 16-20; and ¾ Quarterly individual and cumulative waste dump movement (m3) in Figure 16-21.
Figure 16-17: 20Mt Case – Mill Feed Au
Figure 16-18: 20Mt Case – Total Movement
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Figure 16-19: 20Mt Case – Open Pit Au Ounces by Pit
Figure 16-20: 20Mt Case – Individual Open Pit Total Movement
Figure 16-21: 20Mt Case – Individual Waste Dump Movement
16.4.2 10Mt Case High level schedules were completed for the 10Mt case. The key physicals and anticipated cash flow for the base case is summarised in Table 16-8
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Table 16-8 Summary of Key Physicals for 10Mt Case Description Unit Inventory Mill Feed t 2,277,200 Input Grade Au g/t 1.65 Mill Input Ounces Au oz 120,590 Mill Recovered Ounces AU oz 108,722
Mining Cost $ ‐36,582,481 Mill Cost $ ‐34,158,006 Rehandle Cost $ ‐13,873,640
Revenue Au $ 151,943,087 Undiscounted Cash flow $ 45,614,643 Discounted Cash flow (Inclusive capital) $ 39,647,998 IRR % 812
10Mt Case Schedule: The following key physicals are presented in their respective figures: ¾ Quarterly mill feed tonnes (t) and grade (g/t Au) in Figure 16-22; ¾ Quarterly total open pit movement (t) in Figure 16-23; ¾ Quarterly gold ounces mined from each pit (oz) in Figure 16-24; ¾ Quarterly open pit movement from each pit (t) in Figure 16-25; and ¾ Quarterly individual and cumulative waste dump movement (m3) in Figure 16-26.
Figure 16-22: 10Mt Case – Mill Feed Au
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Figure 16-23: 10Mt Case – Total Movement
Figure 16-24: 10Mt Case – Open Pit Au Ounces by Pit
Figure 16-25: 10Mt Case – Individual Open Pit Total Movement
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Figure 16-26: 10Mt Case – Individual Waste Dump Movement
16.4.3 100m Buffer Case High level schedules were completed for the 100m Buffer case. The key physicals and anticipated cash flow for the base case is summarised in Table 16-9
Table 16-9 Summary of Key Physicals for 100m Buffer Case Description Unit Inventory Mill Feed t 2,317,587 Input Grade Au g/t 1.39 Mill Input Ounces Au oz 104,142 Mill Recovered Ounces AU oz 93,215
Mining Cost $ -42,424,778 Mill Cost $ -34,763,778 Rehandle Cost $ -25,335,771
Revenue Au $ 131,219,072 Undiscounted Cash flow $ 7,955,408 Discounted Cash flow (Inclusive capital) $ 6,419,168 IRR % 69
100m Buffer Case Schedule: The following key physicals are presented in their respective figures: ¾ Quarterly mill feed tonnes (t) and grade (g/t Au) in Figure 16-27; ¾ Quarterly total open pit movement (t) in Figure 16-28; ¾ Quarterly gold ounces mined from each pit (oz) in Figure 16-29; ¾ Quarterly open pit movement from each pit (t) in Figure 16-30; and ¾ Quarterly individual and cumulative waste dump movement (m3) in Figure 16-31.
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Figure 16-27: 100 Buffer Case – Mill Feed Au
Figure 16-28: 100 Buffer Case – Total Movement
Figure 16-29: 100 Buffer Case – Open Pit Au Ounces by Pit
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Figure 16-30: 100 Buffer Case – Individual Open Pit Total Movement
Figure 16-31: 100 Buffer Case – Individual Waste Dump Movement
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17 RECOVERY METHODS
17.1 Mineral Processing The gold processing facilities utilised at SGM comprise a standard Carbon-In-Leach (CIL) gold recovery circuit following crushing and grinding and sulphide flotation. The treatment plant consists of five unit processes. These are: ¾ size reduction (crushing and milling) ¾ gravity gold recovery ¾ flotation/ultra-fine grinding ¾ leach-adsorption, and ¾ gold recovery Geographically the plant can be split up into five main areas. These are: ¾ the primary crushing circuit ¾ the milling circuit ¾ the flotation/ultra-fine grinding circuit ¾ the leach-adsorption circuit, and ¾ the elution/electrowinning circuit Testing of samples from both the Big Hill Northern and Southern (Davis Pit extension) areas has demonstrated that there will be no requirement for flotation/regrind of the mineralisation as the oxide nature of the mineralisation renders this process unnecessary. A current processing flow sheet is shown in Figure 17-1. This, with the flotation section bypassed, is the flow sheet for stockpiled oxide ores currently treated, and represents the flow path for Big Hill oxide mineralisation treatment.
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Figure 17-1 SGM treatment plant flow sheet
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18 PROJECT INFRASTRUCTURE
18.1 Surface Infrastructure Stawell Gold Mines facilities are extensive and representative of a modern gold mining operation (Figure 18-1). The main site location comprises; ¾ Office and administration complex ¾ Store and storage facilities ¾ Heavy underground equipment workshop and Light vehicle workshop ¾ Surface Run of Mine stockpiles ¾ Gold Processing Plant and associated facilities ¾ On Site Assay and Metallurgical test work Laboratory ¾ Four freshwater storage dams to store rainfall run-off and mine dewatering which is used in the plant or around the mine site. ¾ Power for the plant is fed from a main transformer located adjacent to the administration complex. ¾ A batch plant for preparing shotcrete for underground support ¾ Core farm and core processing facility Surface facilities include the gold processing plant, offices, core shed, laboratory and workshops. Larger infrastructure onsite includes tailings dams covering 96 ha and receiving all tailings from the processing plant. Four freshwater dams occur throughout the mine lease.
18.2 Tailings Storage Facilities Since operations began in 1984, three tailings dams have been constructed and operated, two of which have since been decommissioned; ¾ Reserve Tailings Dam. This has been decommissioned and rehabilitated to a Clay Target Shooting Complex ¾ No 1 Tailings Dam. This has been decommissioned and partially rehabilitated ¾ No 2 Tailings Dam remains in operational All dams were constructed as earthen embankments with upstream sub-aerial deposition and are subject to annual integrity and operational review by an independent industry expert. No 2 Tails Dam or TSF2 currently has an approved Work Plan Variation enabling the completion of a wall lift to 253RL (the current wall level is 250RL). Review of remaining underground and Mt Micke oxide stockpile stocks by the dam designer (as per the above independent industry expert) has shown that to treat the mineralisation expected from the Big Hill pits, a lift to 252RL will be required.
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Figure 18-1 Plan showing the location of MIN 5260, Stawell Gold Mines operational infrastructure.
18.3 Power Update to Big Hill Stawell Gold Mines purchases power under contract from Origin Energgy Australia. This 2 year contract expires in Dec 2012. Supply from the National Electricity Grid to SGM is via high voltage iinstallations in two locations; ¾ Moonlight Substation (10 Mega Waatts feed) which supplies the Magdala Underground operation, and ¾ Reefs Road (7 Mega Watts feeed) that supplies the Gold Processing Plant, administration, workshop facilities and parts of the upper levells of the Magdala Underground Mine. Power to underground from the Moonlight substation is supplied through a 990m steel cased borehole and that from the Reefs Road substation via the Magdala decline. The total facilities for electrical power input distribution to site is 5 833 6000 kwh/month The Main underground demands are outtlined below, 1) Primary ventilation fan sizes are ¾ Surface 250kw ¾ 482L 1000kw ¾ 435L 1120kw 2) Compressor sizes are ¾ 1300 cfm x 4(250kw x 4 units) ¾ 1000 cfm x 1(160kw x 1 unit underrgground) ¾ 650 cfm x 3(110kw x 3 units) The total volume of ventilation supplied to the mine is 8,150 cfm.
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Stawell Gold Mines management have provided information that the current power availability is sufficient to meet the needs of the current Life of Mine operating plan and future project requirement.
18.4 Water Water supply is from harvested rainfall runoff, mine dewatering, recycling of process water from the tailings facility, and by way of a 1ML/day raw water right entitlement and urban customer access to potable supply from Lake Bellfield located in the Grampians Mountains. The capacity of the site water storages is approximately 690ML. The Lake Bellfield water is supplied as raw or potable and is preferentially used, when required in the processing operations as it improves gold recovery.
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19 MARKET STUDIES AND CONTRACTS Stawell Gold Mines produces gold doré bars at the SGM, which are transported to AGR Matthey in Western Australia and refined to produce gold bullion. The Corporation sells the gold bullion over the counter according to its treasury policy through either AGR Matthey or an Australian based bank. Representative mining costs have been used by Mining One to conduct the pit optimisation. No mining contracts have been prepared or reviewed for the PEA. The author is not aware of any agreements that are not within market parameters.
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20 ENVIRONMENTAL STUDIES, PERMITTING AND SOCIAL OR COMMUNITY IMPACT SGM is committed to maintaining effective management systems with respect to environmental matters at the Stawell Gold Mine. At SGM, site management monitors and regularly reviews the environmental and social impacts of the operations, such as water quality, air quality, blast vibration and noise. Monitoring and site environmental performance results are shared with regulatory authorities and local communities and are reported quarterly to all stakeholders. SGM promotes responsible environmental behaviour among all employees and contractors. Operations are managed in accordance with SGM's Environmental Management Plan which provides the standards, procedures and guidelines required to achieve that aim. An Environment Effects Statement (EES) for the Big Hill project was conducted in 1998 at a cost of five million dollars with the intention to progress the project to a prefeasibility study. The EES documents were presented at panel hearings in January 2000. In November 2000 the Minister for Planning released the decision which recommended against the project but after considerations, the EES Panel endorsed the project subject to certain enhancements. The Minister’s recommendation was that the “project did not provide an acceptable balance of economic, social and environmental outcomes’ and outlined social impact, loss of Box Ironbark and Heathy forest and an un-filled southern void were significant issues. During the initial Big Hill Project proposal process, SGM worked successfully with the Northern Grampians Shire Council (NGSC) to resolve the 42 points or issues raised by the NGSC in the original EES enquiry. The NGSC passed a motion in support of SGM’s Big Hill development proposal and has stated that the issues were largely addressed by SGM’s enhanced proposal consideration in 2003. Now subjected to favourable gold prices and after again reviewing the Panel’s comments, along with issues identified by the NGSC and comments in the Minister for Planning’s recommendation, SGM now seeks to re-address the Big Hill Development Project. The major physical enhancements proposed for the project are: ¾ The North pit void will be backfilled to re-establish and reshape the Big Hill landform. The north void will be backfilled by Q1 year 3 of mining to reduce community exposure. ¾ The southern void will be backfilled. Remnant Mt Micke stocks will be used to complete the backfilling and profiling of the southern void. Utilising Mt Micke material will allow for the Mt Micke stockpile footprint near the Wonga Pit to be returned to original topography for final rehabilitation work. ¾ Waste dump design now encompassed within existing disturbed mining footprints therefore no further disturbance to Box Iron Bark habitat is required. Pit volume reductions have down sized the requirement for additional waste rock dump space. ¾ The pit design is considerate of both reservoir preservation and northern community exposure. Reduced capital costs of approximately $5 million and present infrastructure impacts. ¾ Increased oxide treatment leading up to the close of mining operation to aid in further enhancing TSF2 capping material for closure design and facility closure.
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¾ Historic mining voids rehabilitated will remove existing community safety exposure allowing the area to be safe and stable for perpetuity; ¾ Completion of the Big Hill landform to be incorporated into the cultural heritage gold trail including a community facility such as a mining museum with views of the Grampians and the Pyrenees Range which is Tourism and Goldfields Heritage Scheme aligned. The project will follow environmental management guidelines outlined in the License Conditions that applies with the current operation on ML5260 and documented in the sites Environmental Management Plan (EMP). The noise limits have been derived from procedures outlined in the EPA State Environment Protection Policy (Control of Noise from Commerce, Industry & Trade) No. N-1. Water monitoring will be undertaken to ensure compliance with the SEPP (Waters of Victoria), Work Plan Conditions and Mining License Conditions and dust monitoring will follow SEPP Air Quality Management (AQM) and National Environmental Protection (Air Quality) Measures (NEPM) guidelines. Since the initial EES process in 1998 SGM has continued to work to increase its level of communication and interaction with the community by increasing its focus in the areas of community consultation, community involvement and research and development. Specific elements of the project could be subject to change as SGM progresses with the community consultation process.
20.1 Environmental Issues and Studies An initial assessment of the project has determined the highest key environmental impacts will be amenity impacts from air and noise emissions followed by landscape impact (visual) and flora and fauna. Amenity - Air, Noise, Light and Vibration Impacts The project will be developed to minimise the amenity impacts on surrounding residents. SGM has committed to completing the project in the quickest possible timeframe by working the North pit first then allowing progressive backfilling and rehabilitation while excavating the South pit and by minimising hours of operation to minimise impacts on surrounding residents. Engineering measures have also been built into the project to minimise amenity impacts such as a noise attenuation barrier to be installed after noise assessments are complete. SGM will be undertaking a comprehensive air, noise, light and vibration assessment to assess the impacts of the project on surrounding residents and develop management and monitoring actions to address identified issues. The proposed Big Hill Development project is envisaged to require a noise attenuation wall between the project and the community of Stawell, with restricted hours of operational also applied for the purpose of managing to compliance limits. These considerations have been applied within the economic assessment, although remain subjective to Noise Attenuation Study determinations. Human Health Risk Assessment Based on outputs from the various investigations into matters such as air quality, dust, noise and vibration, a Human Health Risk Assessment will be conducted to properly identify potential effects and mitigation measures that may be required.
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Landscape Impact The topographic feature known as ‘Big Hill’ will be excavated, processed and backfilled and landscaped to the original topographic shape as part of this project. This will result in a temporary change in the landscape that will be visible from within the Stawell township and surrounding local area for the duration of the project.
The visual impact will be minimised wherever possible, and progressive backfilling and rehabilitation works will be undertaken as the Big Hill landform is reinstated. Where required, a noise attenuation wall will be placed in consideration for preservation of vegetation and visual amenity / screening, as well as noise reduction.
The community will have the opportunity to be involved in developing the final landscape form of Big Hill.
Native Vegetation, Flora and Fauna Ecology and Heritage Partners Pty Ltd undertook a preliminary flora, fauna and net gain assessment in March 2012 based on a January 2012 survey.
These assessments comprise two reports titled as follows:
¾ ‘Flora, Fauna and Net Gain Assessment at the Big Hill Site for the proposed extension of the Stawell Gold Mine, Stawell, Victoria.’– December 2012; ¾ ‘Flora, Fauna and Net Gain Assessment at the Davis Site for the proposed extension of the Stawell Gold Mine, Stawell, Victoria.’– December 2012; The aims of these assessments were to identify the type, quality and quantity of native vegetation and fauna habitat present within the study area. The January 2012 flora and fauna assessment included the majority of the pit area. Additional survey is required to assess the remaining pit footprint to the south, TWRS and haul road footprints, however initial survey data indicates that the project impacts on flora and fauna is unlikely to be significant. The majority of the project site is highly modified and contains a high diversity of exotic species mixed with indigenous over-storey. The fauna and fauna survey conducted for the pit area has concluded the following: Flora ¾ One Ecological Vegetation Class (EVC) is present within the study area, Box-Ironbark Forest - which is listed as depleted within the Goldfields bioregion; ¾ One state significant species is present in the survey area Clasping Goodenia Goodeniabenthamiana– which is rare in Victoria. Fauna ¾ No nationally significant fauna species were recorded during the survey, however 14 nationally significant fauna species have previously been recorded in the local area including the Swift Parrot and the Australian Painted Snipe which may utilize the woodlands and dams onsite on rare occasions. However the site is unlikely to provide permanent and/or important habitat for these species;
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¾ One species of state significance Brown Treecreeper Climacterispicumnusvictoriae andone of regional significance Bearded Dragon Pogonabarbata were identified during survey. Suitable habitat exists for both of these species within the study area. EPBC Act listed species No listed flora or fauna species under the EPBC Act 1999 are considered likely to occur within the area, as there is no suitable habitat present. FFG Act listed species Several flora species including Hedge Wattle Acacia paradoxa, Fireweeds Senecio sp., Drooping CassiniaCassiniaarcuata and Common Everlasting Chrysocephalumapiculatum which are listed as protected were recorded during the field assessment. The primary effects to flora and fauna which would occur as a result of the vegetation and habitat removal include: ¾ Loss of nesting and foraging habitat; ¾ Potential removal of significant flora; ¾ Removal or modification of habitat for significant flora species. Adequacy of the 1999 EES Study Studies were undertaken for the previous Big Hill proposal in 1999, however, due to legislation changes, modelling techniques and changes to the environment, a gap analysis has been conducted. and further studies are required. These additional studies will be undertaken as a part of the permitting requirements for the project. Legislation changes which will require further investigation and subsequent actions are: Air Quality ¾ New State Environment Protection Policy – Air Quality Management (AQM) was promulgated in 2001 and State Environment Protection Policy (Ambient Air Quality) in 1999; ¾ The PEM: Mining and Extractive Industry (EPA Pub 1191, Dec 2007) is an incorporated document which supports the interpretation of the SEPP; ¾ Although background information from the 1999 Air Quality assessment may be utilised, a new Air Quality Assessment is likely to be required to satisfy the new SEPP objectives and incorporated PEM and other guidelines; ¾ The Air Quality Assessment will need to be updated to incorporate the revised mining extraction and rehabilitation proposals including impact assessment of potentially new sources and receptors from filling the southern void which was not assessed as part of the 1999 EES. Noise Since the 1999 Noise Assessment was undertaken SEPP (N-1) has been subject to variations and the introduction of guidelines for Noise control in Rural Victoria. The 2012 Big Hill project is likely to require re-assessment of background noise levels and modelling of potentially new sources and receptors associated with the revised mining and rehabilitation proposal against the following updated policy and guidance published since the 1999 EES:
¾ EPA Publication 1411 Noise from Industry in Regional Victoria (October 2011);
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¾ EPA publication 1254 Noise Control Guidelines (2008); ¾ Department of Primary Industry Environmental Guidelines, Ground Vibration and Air blast Blasting Limits for Blasting in Mines and Quarries, 2001; ¾ Updates to Australian Standards AS1055:1997 “Acoustics –Description and Measurement of Environmental Noise” as relevant.
Blast Effects A new Blasting Study is required for the Project, given the depth and dimensions of the pits have been changed. Furthermore, SGM have collected substantially more information on ground vibration over the last 12 years which will allow more accurate calibration of any modelling. It is envisaged that blasting requirements will be limited to the deeper sections of the pit activity as free dig and ripping techniques become restrictive to transitional zone rock types. The proposed pits are approximately -90MRL deep with transitional rock zones which may require blasting presenting from about -80mRL.
Visual Impact A new Visual Impact study is required for the 2012 project due to the following changes which will have an impact on the study conclusions: ¾ the development footprint has changed; ¾ the southern void is now proposed to be filled which substantially improves the landscape impact; ¾ the vegetation has grown in the past 12 years since the 1999 proposal which may reduce visual impacts; ¾ the size of the waste rock emplacement is substantially reduced from the 1999 proposal. The assessment needs to consider any regional policy regarding landscape values that may have been introduced since the 1999 Landscape Assessment was completed.
Socio - Economics The economic study considers the Stawell economy and mining operations in 1999. The underground operations are currently in a transitional phase towards closure and an economical re-assessment has to be conducted. Surface Water The surface water management assessment would need to be redeveloped to consider mine water management i.e. the implications of progressively backfilling the pit voids, the management of water inflow into the pits and finally, sediment and erosion control. Hazard and Risk A Hazard and Risk Assessment study needs to be carried out to assess the impacts and risks of the proposed project activities, to provide early stage recommendations to the technical departments for consideration in the design. Since 1999 the approach to environmental risk assessment has advanced considerably and become less qualitative, more logical, robust and integrated with technical studies and therefore require a comprehensive risk assessment. In relation to environmental
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STAWELL GOLD MINE assessment, regulators prefer a structured risk based framework that applies a semi quantitative approach. The approach to risk assessment involves: ¾ Clear articulation of the risk assessment context and objectives; ¾ Identification of risk events, their likelihoods and consequences using workshops with key stakeholders and technical specialists. Greenhouse Gas Greenhouse gas is now considered an environmental impact which requires assessment as part of any major development proposal. A greenhouse gas assessment has not been part of the 1999 EES study. A greenhouse study for this project must include the following: ¾ Review of baseline greenhouse gas/energy data for existing operations; ¾ Review of proposed energy use for construction activities; ¾ Estimation of energy uses for construction and operating phase; ¾ Identification of means of reducing energy usage and greenhouse gas during construction; and ¾ Review of proposed ongoing energy use of the facility and how can it be minimised.
20.2 Waste Management Tailings Disposal If the Big Hill resource is mined there will be a requirement to raise the wall of the current tailings storage facility (TSF2) by between one and two metres. It has been identified that additional lifts to the tailings facility could potentially add to existing environmental impacts regarding the tailings facility. SGM has an extensive water monitoring programme on site with includes a monitoring bore network and surface water receivers. The monitoring programme has detected localised impacts on the groundwater in some shallow monitoring bores around TSF2. It has also been identified that the groundwater in several areas is also elevated due to hydraulic pressure from the storage facility. To investigate management options SGM has undertaken multiple studies of the area and the impacts from the seepage (Coffey 2008, LanePiper, 2008, Rockwater, 2010, NQ Groundwater and Environment 2011 and NQ Groundwater and Environment 2012). An Environmental Audit of groundwater (LanePiper, 2008) was completed for the entire site. The conclusions of the audit were that the primary impacts are related to salinisation of soil from elevated groundwater levels and cyanide compounds. During mining operations, to manage elevated water levels and seepage from TSF2, rock filled cut off seepage trenches have been installed in several locations around the facility. A number of 50mm seepage interception bores have also been installed around TSF2. In January 2012 an additional five 150mm seepage interception bores were installed and fitted with solar powered pumps in Q2 2012. Additional 50mm monitoring bores were also installed to correlate the extent of seepage. In a review conducted by NQ Groundwater and Environment in 2012 during the installation of the five new seepage bores it has been concluded that due to a very high clay content and high weathering rates , seepage
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STAWELL GOLD MINE mitigation from TSF2 is very slow ranging from 0.0016m/day to 0.067m/day. Increasing distance, dilution and absorption of cyanide of geological materials down grade will slow migration. It is expected that with capping and closure of TSF2 the tailings will progressively de- saturate, with the affect that groundwater levels will slowly re-establish to near pre-TSF levels. This is supported by the modelling works undertaken by O’Kane consultants. To manage seepage post closure, a wet land system using native wetland plant species to filter and polish water is being investigated, which is to be incorporated into the final capping design of TSF2. SGM is committed to sustainable development and providing long term environmental protection, for the community of Stawell and the state of Victoria. To be able to do this, SGM has established a research alliance with Melbourne and Curtin Universities, which have undertaken cutting edge research in the following areas: ¾ Innovative Rehabilitation Strategies for Sulphidic Tailings Dams is a project involving a number of PhD students and numerous honours students studying botany and geochemistry; ¾ The ultimate objective of this research is to devise cost effective rehabilitation/revegetation protocols for sulphide-bearing gold tailings that results in land of high economic and/or environmental value, which will provide an economic asset to the Stawell community once the mine has ceased to operate; ¾ Investigations into the success of rehabilitation strategies used at Stawell. This work involves carrying out flora and fauna surveys in areas rehabilitated over 10 years ago by the mine’s owners and compares the results to that for natural bushland. The aim is to either validate the measures used or to adjust them in light of the longer term results to ensure disturbed areas are truly “rehabilitated”; and ¾ In 2000, SGM established a Tailings Experimental Research Facility (TERF) on tailings dam 1 as a field trial for different closure mechanisms for the tailings dams. This research has shown that this material will provide the best net percolation reduction for the capping medium preventing seepage from Tailings Dam 2 reaching Salt Creek. The University of Melbourne and Curtin University have been undertaking research in this field to determine the impacts of thin cover designs on a number of aspects including: ‐ Acid Mine Drainage; ‐ Ecotoxicology; ‐ Soil stabilization; ‐ Arsenic uptake; and ‐ Phytostabilisation using trees. An area was set up on TSF1 for research to be carried out by O’Kane Consultants. O’Kane Consultants Pty Ltd (OKC) was engaged in 2009 to assist with the design of a cover system for the TSF2 closure planning. The main objective of this project was to assess the research undertaken on the Tailings Experimental Research Facility (TERF) and develop an enhanced store-and-release cover system for TSF2. To achieve these objectives several tasks were completed as part of the cover system design process:
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¾ A comprehensive field sampling and material testing programme to assess potential cover materials was undertaken; ¾ Cover system criteria in terms of impacts of solutes on downstream receptors was assessed; and ¾ Several cover system design alternatives were investigated using the information determined during this study and from background information. The recommendation following the research conducted is to rehabilitate the tailings dams by placement of 0.3m processed oxide cap with additional 0.1m topsoil. If the processing of the material is not economical viable, the tailings dam should be capped with a 0.3m oxide cap and 0.1m of topsoil. The 0.3m cap was found to perform at a rate suitable to prevent downstream impacts on water systems. The Big Hill Development project will present more oxidised mineralisation for treatment at close of operations and further enhance the capping design. Research works pertaining to Big Hill mineralogy are to be modelled by O’Kane Consultancy in final TSF capping design. The mineralogy is believed to be similar in nature to the Wonga Pit Oxides (Mt Micke) and should not be an area of concern. Waste Management The Temporary Waste Rock Site (TWRS) site is located approximately 200 metres south of town water reservoir number 7. The site until recently stockpiled waste rock from the former Davis Pit and is currently waiting for rehabilitation. The Davis Pit waste rock has progressively been mined over the last 6 years and processed due to its gold grades becoming economic. The TWRS is located entirely on previously disturbed site. The proposed footprint proposes minimal vegetation removal which is likely to be low quality regrowth. This vegetation will be assessed to determine its quality, quantity and need for net gain offset calculation. An option for the stockpiling of additional waste materials on adjacent cleared land may be considered pending requirements of GWM-Water reservoir holdings and air borne dust considerations. This option would offset the imposition of heightened materials, rehandling to the Wonga Pit and further enhance the Project economics with no impacts to environment and a commitment for rehabilitation and re-establishment of grazing grasslands post project. Design of the TWRS will be undertaken to ensure that all run off is directed back towards the open pit or sumps located on site and the water will be handled within the site catchment. All water captured on site is sent via the processing facility to the TSF. Pit water and any sumps created as a part of the project will be incorporated into the current water monitoring programme.
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20.3 Permitting Requirements The Big Hill Enhanced Development Project will require a variety of regulatory approvals at the State, Local and potentially Commonwealth level. Potential approvals required include: ¾ Planning Permit required under the local Planning Scheme as the site is located in three zones – Public Park and Recreation Zone (PPRZ), Special Use Zone 1 (SUZ1) and Public Use Zone 1 (PUZ1). ¾ Flora and Fauna Guarantee Act approval – if Project involves native vegetation removal. ¾ Environment Protection and Biodiversity Conservation Act (Commonwealth) approval – if the Project impacts on Matters of National Environmental Significance (NES). Available information confirms there are limited matters of NES on the Project site. It is highly unlikely the Project will require approval through the EPBC Act. However the need for referral will be reviewed as flora, fauna and heritage studies progress. ¾ Mineral Resources (Sustainable Development) Act – Work Plan. The project also requires the formal consent of all owners of a dwelling within the 100m buffer of the pit under the Mineral Resources Sustainable Development Act (MRSDA) section 45 Prohibition of work near dwellings and certain places and site. The current operation operates on the mining licence MIN5260 which is current until 30th May 2020. MIN5260 covers all current activities including processing (including tailings disposal) and the underground mining operation. The current tailings storage facility TSF2 is also permitted to have another 3m lift if required under the 2009 TSF2 Lift Work Plan Variation. SGM regularly updates its estimate of closure expenditures. In November 2010 SGM submitted the Closure Plan to the Department of Primary Industry (DPI). On review of that document in 2011 the DPI increased the site bond by $255,000, lifting the security bond for the site to $A4.803 Million. This estimate is based on an additional cutback conducted at the Wonga Pit, available information including preliminary closure plans, alternatives and applicable regulations. It is envisaged that there will be a review of the bond to align with the proposed project should it be permitted.
20.4 Heritage, Social Impact & Community Engagement Heritage Heritage studies have been undertaken with the project and surrounding area during the life of the mine. These cover aboriginal and European heritage and each time no finding have occurred for aboriginal or European heritage sites. These studies will be further updated to assess any more recent discoveries. Aboriginal Cultural Heritage Over a number of years the Aboriginal occupation and archaeology of the Stawell region has been the subject of numerous surveys (eg: Gunn 1983, Snoek 1983, Russell 1992 & 1995, Gunn& Parsons 1997a & b). During 1981 Stuart. I (1981) conducted a survey for Western Mining around the project area and found no evidence of Aboriginal occupation but evidence of historic mining practices
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STAWELL GOLD MINE was prevalent. His assessment was that mining practices would have destroyed any possible evidence of Aboriginal occupation. In 1983 a survey by Snoek (1983) found a single quartz flake but nothing else. His assessment of Aboriginal occupation and mining practices was the same as that of Stuart (1981). Additional surveys conducted by Snoek (1988) and Gunn and Skurrie (1998) also found no additional evidence of Aboriginal occupation. In 1998 Stawell Gold Mines commissioned Resource Strategies to compile an EES for the Big Hill Project which incorporated the area associated with the Davis Pit. An archaeological and heritage impact assessment was undertaken by R.G. Gunn and A.Burns (1998) for the project. Stuart Harrington and senior Aboriginal elder Jack Kennedy, from the Goolum- Goolum Aboriginal Cooperative were contacted to discuss the significance of Big Hill (Kobram). Mr Kennedy, while personally unhappy with the thought of the removal of Big Hill he was not aware of any religious significance of the hill or the immediate surrounding area. Aboriginal Affairs Victoria and the coordinator of the south-west and Wimmera Heritage Programme were also notified of the intent to undertake the survey. The survey area was subdivided into 6 units, Blocks A-F; with a seventh (X) being the remaining areas that did not warrant detailed foot survey (Figure 20-1).
Figure 20-1 Survey area for the Big Hill project EES
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As with other archaeological surveys conducted in the past, the survey conducted by R.G. Gunn and A.Burns in 1998 recorded no Aboriginal archaeological sites and they imposed no restrictions to the proposed development (from an Aboriginal Heritage perspective), however, within a 20 km radius of Stawell, 441 sites have been recorded to date. Figure 20-2 shows known registered Aboriginal archaeological sites around the Stawell area.
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Figure 20-2 Known registered Aboriginal archaeological sites around the Stawell area
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European Cultural Heritage A study of the mining sites in Stawell Mining Division undertaken by David Bannear in 1994 identified no sites of significance within the Big Hill Development Project Area. The sites identified near Big Hill were the Four Jacks Mine, Magdala-cum-Moonlight, North Magdala, Oriental Company and Leviathan Cyanide Works, each of which was assessed as being of regional significance (that is, significant in a regional context, but not of sufficient significance to be entered in the Victorian Heritage Register). In 1998 SGM commissioned Resource Strategies to compile an EES for the Big Hill Development Project which incorporated the area associated with the Davis Pit. An archaeological and heritage impact assessment was undertaken by R.G. Gunn and A.Burns (1998) for the project. During their survey they identified two areas that contain evidence of the early mining period of Stawell. The mineshaft and building foundations in area C (Figure 20-1) by Albion Rd are clearly historic features and should be recorded and assessed. The relief features in the south-western section of Block E (Figure 22-1) probably represent remnant mine workings also from the earlier Mining 1 phase and should be similarly recorded and assessed. They recommended that these areas be recorded and assessed as per the European Heritage Study which was to be conducted as part of the EES. Michael Pearson from Heritage Management Consultants Pty Ltd (1998) conducted a European Heritage Study for the Bill Hill Development Project EES which included the Davis Pit area. His findings during his assessment identified at that time, there were no places associated with the Big Hill Development Project entered in the Register of the National Estate and also there were no sites entered into the Victorian Heritage Register. He also identified that no sites had yet been entered in the Inventory of Archaeological Sites maintained by Heritage Victoria in relation to the operation of the archaeological controls within the Heritage Act (as at December 1998). A new heritage study has been undertaken by the (NGSC) as part of the development of the heritage layer of the planning scheme for Northern Grampians Shire. The results of this have yet to be published by the NGSC. Community Engagement SGM states as one of its key cultural behaviours that its employees will “respect the environment and community in which we work”. The Company is a signatory to the Minerals Council of Australia’s “Enduring Value” and supports its core principles including the importance of strengthening relationships with the community. As signatories to the MCA’s Enduring Value Framework, SGM are committed to regular, open and honest communication with the community. As per SGM’s community policy, the community will be kept informed of all programmes associated with the mine’s development. Communication methods used include updates in the local newspaper, the quarterly Environmental Review Committee, SGM’s quarterly newsletter, direct mail outs and door knocking. As with all mine operations, community members will be encouraged to provide feedback at any time throughout the programmes to a nominated SGM contact person. Stawell Gold Mine is closely associated with a small regional country town and is surrounded by farming land – grazing and cropping – to the north and east, State Forest to the west and the township of Stawell to the south and east. Due to the long history of mining activities in the area many attitudes have been formed on past activities by numerous administrations. Community attitudes to the mining operation
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STAWELL GOLD MINE are quite diverse, ranging from low interest to being fully engaged in working to overcome potential areas of concern with the Company. The expectation by the local community is that the SGM will continue to operate in accordance with its license conditions and modern community standards and the local community expects and appreciates being informed of changes to the operations. SGM has engaged external consultants in community and public relations and is currently developing an Engagement and Community Strategy for the project. The purpose of this document will be to set out the process and requirement for consultation for landholders, businesses and stakeholders. To date only high level consultation has been undertaken with approval agencies and agencies directly involved with that; ¾ Northern Grampians Shire Council (NGSC); ¾ Department of Primary Industries (DPI); ¾ Department of Planning and Community Development (DPCD); ¾ Department of Business and Innovation (DBI); ¾ Grampians Wimmera Mallee Water (GWM). The social impacts of the project will be assessed through a full Social Impact Assessment (SIA) and will be conducted covering a range of issues and include mitigation measures for: ¾ Social amenity and disruption; ¾ Air quality and Dust; ¾ Human health risk; ¾ Noise; ¾ Night Lighting; ¾ Blasting and vibration; ¾ Rehabilitated land form; and ¾ Visual Impact during mining and after rehabilitation. Some of the areas the community engagement will focus and address are; ¾ Current designs for the project mining operations will come within 100m of 51 dwellings where formal consent will be required from the landholders. ¾ Two residents located to the west of the proposed pit outline located on Main Street will be displaced by this project; ¾ Big Hill forms a natural lookout over the town of Stawell and is utilised by locals and tourists; ¾ The level of impact on the adjacent residents and wider community for the duration of the project has not yet been determined; ¾ The environmental impacts (noise, dust) due to the mining excavation activities in close proximity to the residences on Fisher and Upper Main Street. This will be confirmed through an air quality study and noise assessment of the project prior to commencement. Once the community and stakeholders have been consulted to identify potential project impacts and benefits, mitigation will be addressed by:
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¾ Assessing potential negative impacts according to likelihood and consequence; ¾ Identifying impacts requiring mitigation; ¾ Assessing the opportunity for positive socio-economic contributions, and steps needed to realise these; and ¾ Assess the cumulative impact of relevant projects and developments in the affected area.
20.5 Mine Closure (remediation and reclamation) Requirements and Costs The proposed area of project activity is within the current mining licence ML5260 as is all the associated components (processing and tailings storage). MIN5260 is a combination of crown land managed by the Department of Sustainability and Environment (DSE) and privately owned SGM land. The total mining license area is approximately 1400ha of which approximately 288ha has been disturbed. This includes the open pits, tailings dams, processing facility and other associated infrastructure. All rehabilitation work will be in line with approved work plans and guidance from regulatory bodies and relevant stakeholders. Final rehabilitation will be in accordance with the approved work plan and any additional requirements as directed by DPI after the expiration of the final term of the licence or if the licence is declared void for any reason whatsoever, the holder shall continue to be bound by the rehabilitation provision of the licence and the approved work plan. The Provisions of Amendment 21 to the Town of Stawell Planning Scheme gives detail of the requirements for landscaping and rehabilitation. A Landscaping plan is to be submitted to the council prior to the commencement of construction activity and the revegetation and rehabilitation of the site is to be completed within two years of the cessation of mining operations. All derelict and redundant plant, vehicles machinery and equipment that cannot be sold will be removed from the licensed area and deposited at an appropriate waste disposal site. SGM is committed to a progressive rehabilitation plan and will seek opportunity where presented. This consideration has been applied within the Preliminary Economic Assessment of the Big Hill Development project with mining extraction sequences enabling progressive rehabilitation and minimal waste stockpile footprint. Further consideration has been applied to materials movement to the disused Wonga Pit and may assist rehabilitative and closure requirements of that area of operations. The proposed Enhanced Big Hill Development project also intends on utilising the Wonga Pit waste dump (Mt Micke) for the purpose of filling remaining Big Hill Development pit voids. This will also enhance the closure and rehabilitation of the Mt Micke stockpile site. The costs for back filling Big Hill are included in the operating costs for the project, additional costs will occur from the rehabilitation of the site to native bushland. These costs are presented in Table 20.1.
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Table 20-1 Estimated costs to rehabilitate the Big Hill Project area if the project proceeds
Project estimated costs do not include stakeholder negotiations. Material movement for mining and relocation of material back into pit voids is covered in project mining costs. All costs are subject to change until final design, contractor costs, operational requirements and community consultation are set.
Item Requirement Cost ($) Comments
Waste movement Waste material to fill - Waste movement costs north and south pit incorporated into mining costs.
Drainage Works Fill in drain lines and 36,000 Approximately 3,000m of sumps drainage lines and 2 sumps.
Growth medium May require 111,000 430,000 May require a cover of up to 0.3m material m3 of growth medium cover for plant growth before material (Mt Micke) topsoil applied.
Top soil movement Will require 143,000 Top soil will come from current approximately stockpiles 37,000 m3
Earth works Final contour and 59,000 Approx. $1600 per hectare scarifying
Haul roads Deep rip and cover 10,000 Approx. 650m x 7m with topsoil
Plants Approx. 1000 plants 130,000 Approx. $3.5 per seedling planted per hectare (37,000 plants)
Spray seeding Approx. 37ha 279,000 Spray seeding to maintain stabilised relocated soils
Weed/land Ongoing weed 93,000 Management of rehab area will management management be required for a 10 year growth period.
Davis Area Earthworks and 161,000 Push in soil walls and reshape rehabilitation surface. Oxide cover, topsoil application, plant and seed.
Total 1,341,000
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21 CAPITAL AND OPERATING COSTS
21.1 Capital Estimate Open Pit Mining Capital requirements are mainly associated with impacted infrastructure and are estimated to be in the order of $22M, inclusive of sustaining capital. These are inclusive of further studies, test work and all required pre-start capital as at the time of writing this report.
21.2 PFS Studies -$1.1M Provision has been made for consultants to progress the Preliminary Economic Assessment to the next level i.e. prefeasibility study.
21.3 Geotechnical Drilling - $0.5M In August 1999 the Kevin Rosengren and Associates compiled a Preliminary Geotechnical Review of the proposed Big Hill open pit located in Stawell. This report was based upon data extracted from nine (9) vertical drill holes and exposures in the nearby Davis Pit. The report was designed to identify key geotechnical domains to and provide design recommendations to prevent the most likely failure mechanisms. At this level of study the geotechnical advice is currently at 30% of confidence. Mining One engineers have incorporated these slope design recommendations and designed a pit compliant with the overall slope angle (OSA), to represent a practical pit design. As such, this has enabled the project to move forward to Feasibility stages. The current design is not final, Crocodile Gold currently has a Feasibility investigation study planned which includes a significant geotechnical drilling programme at close spacing. This will include laboratory testing of site specific samples and is targeted to increase the design recommendation confidence to 50-75%
21.4 Equipment Mobilisation - $1.8M Equipment mobilisation will be established once a trade-off has been conducted to determine whether the mining will be executed by a contractor or whether the mining will be done by Crocodile. An estimated site establishment cost of $1.8M was included in the capital for this purpose.
21.5 Plant sustaining capital - $0.5 Provision of $0.5M was included in the capital for plant sustaining and the amount will be refined in the prefeasibility study.
21.6 Communications tower $1.05M The communications tower is currently located on top of Big Hill and services two telecommunications providers as well as emergency services. A new telecommunications tower would be required to be erected in a suitable location and be functioning before the current facility can be removed. The current design of the pits allows for a ridge to be maintained between the two pits that contains the telecommunication tower, this will need to be assessed geotechnically to decide whether it is suitable for infrastructure to remain.
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21.7 Dams- $3.0M There is a possibility that two of GWM Water’s containment reservoirs would be required to be decommissioned and replaced with a single dam 300m to the south-east on GWM Waters land. These two dams are located directly to the east of the north and south pit. It has been proposed to GWM Water that the containment reservoirs are taken offline during the project to lessen the geotechnical instability and perched water risk on the closest pit wall. This option will not require the capital expenditure and construction of a new reservoir.
21.8 Noise Barrier- $3.5M A noise barrier is proposed to be erected between the Big Hill Development project and the Stawell community prior to the commencement of mining works. This intent remains subject to noise attenuation studies.
21.9 Noise and dust monitoring -$ 0.10M Due to the proximity to residences the sensitivity for noise and dust, there would be a need to install fixed real time noise and dust monitors to ensure compliance from mining activities.
21.10 Purchase of houses - $ 0.5M Two properties to the north (Upper Main St) are in immediate proximity to the proposed north pit outline. These properties would be required to be purchased for the project to proceed.
21.11 Permitting and Tech. Studies - $ 2.5M A proposed project referral has been submitted to the Victorian Government planning department regarding the project and coupled to this and obligatory through either option the planning department chooses (EES or special permitting) a host of environmental, social and heritage studies will be required to be undertaken. These studies include: ¾ Air quality ¾ Noise ¾ Blast Effects ¾ Visual Effects ¾ Heritage studies ¾ Socio-Economics ¾ Public Infrastructure ¾ Traffic ¾ Waste Management ¾ Hazard & Risk ¾ Health Effects ¾ Social Impact Assessment ¾ Flora & Fauna ¾ Planning ¾ Greenhouse Gas Assessment
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21.12 Compensation for houses -$1.0M Allowances for compensation have been considered within the Preliminary Economic Assessment given the likelihood of residential inconvenience immediate to the project. The extent to which compensation is considered will be decided through a risk based approach after the completion of both noise and dust assessments, in addition to property valuation and impact likelihood.
21.13 Big Hill & Davis Oxide Studies-$0.2M Closure and capping studies have been completed on the Mt Micke stockpile for the use on the tailings storage facilities for closure. Should the proposed project go ahead the Mt Micke stockpile will be depleted and the project material will be placed on top of TSF2. The studies will analyse whether the Big Hill and Davis oxides are the same geochemically and suitable for capping the tailings storage facility. If there are differences either chemically or structurally to the project oxides then a redesign and modelling of the proposed cap is required.
21.14 Wall raise of TSF2 - $1.1M There will be a requirement to raise the wall of the current tailings storage facility (TSF2) by between one and two metres.
21.15 Project Management - $0.3M Project management will be undertaken during the proposal and permitting aspect of the project to co-ordinate the permitting and technical studies as well as community consultation and public relations. The project management will especially focus on the schedule, scope, legislation and legal requirements of permitting for the department of planning. The timing of the capital expenditure for each of the cases has been estimated and the estimates are presented below in Table 21-1, Table 21-2, and Table 21-3
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Table 21-1 Capital for 20Mt Case
Capital Pre-Start Capital Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Total Capital Communications tower -1,050,000 -1,050,000 Fire lookout tower -99,999 -99,999 Dams -3,000,000 -3,000,000 Noise Barrier -3,497,650.2 -3,497,650.2 Noise & Dust monitoring equipment -100,000 -100,000 Houses -500,000 Compensation (home owners) -1,000,000 -1,000,000 -1,000,000 -1,000,000 -1,000,000 -250,000 -5,250,000 Referral, permitting & tech studies -2,500,000 Roads -500,000 -500,000 Big Hill / Davis Oxide studies -200,000 Big Hill / Davis Oxide TSF2 lift -1,100,000 -1,100,000 PF study -1,100,000 Geotechnical drilling -500,000 Site machinery (contractor or owner / operator) -1,800,000 -1,800,000 Project management team for set up -291,668 Plant sustaining capital -500,000 -500,000 -500,000 -500,000 -425,000 -50,000 -2,475,000 Capital Total -5,091,668 -12,047,650.2 -2,000,000 -1,500,000 -1,500,000 -1,425,000 -399,999 -23,964,317
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Table 21-2 Capital for 10Mt Case
Pre-Start Capital Year 1 Year 2 Year 3 Year 4 Total Capital Capital Communications -1,050,000 -1,050,000 tower Fire lookout tower -99,999 -99,999 Dams -3,000,000 -3,000,000 Noise Barrier -3,497,650.2 -3,497,650.2 Noise & Dust monitoring -100,000 -100,000 equipment Houses -500,000 0 Compensation (home -1,000,000 -1,000,000 -1,000,000 -750,000 -3,750,000 owners) Referral, permitting & -2,500,000 tech studies Roads 0 -500,000 0 0 -500,000 Big Hill / Davis Oxide -200,000 studies Big Hill / Davis Oxide -1,100,000 -1,100,000 TSF2 lift PF study -1,100,000 Geotechnical drilling -500,000 Site machinery (contractor or owner / -1,800,000 -1,800,000 operator) Project management -291,668 team for set up Plant sustaining -500,000 -500,000 -500,000 -225,000 -1,725,000 capital Capital Total -5,091,668 -12,047,650.2 -2,000,000 -1,500,000 -1,074,999 -21,714,317
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Table 21-3 Capital for 100m Buffer Case
Pre-Start Capital Capital Year 1 Year 2 Year 3 Total Capital Communications tower -$1,050,000 -1,050,000 Fire lookout tower -99,999 -99,999 Dams -3,000,000 -3,000,000 Noise Barrier -3,497,650 -3,497,650 Noise & Dust monitoring equipment -100,000 -100,000 Houses -500,000 Compensation (home owners) -1,000,000 -1,000,000 -1,000,000 -3,000,000 Referral, permitting & tech studies -2,500,000 Roads -500,000 -500,000 Big Hill / Davis Oxide studies -200,000 Big Hill / Davis Oxide TSF2 lift -1,100,000 -1,100,000 PF study -1,100,000 Geotechnical drilling -500,000 Site machinery (contractor or owner/operator) -1,800,000 -1,800,000 Project management team for set up -291,668 Plant sustaining capital -500,000 -500,000 -500,000 -1,500,000 Capital Total -5,091,668 -12,047,650.2 -2,000,000 -1,599,999 -20,739,317
21.16 Explanation of the differential in capital expenditure of the different Pits. The main differences between the capital required for the different cases is associated with the compensation for houses, estimated to be $250,000 per quarter; and Mill sustaining capital, estimated to be $125,000 per quarter. The respective difference in the mine life for the cases, directly impacts the amount of housing compensation and Mill sustaining capital required to execute each case. The respective mine life and capital differences are summarised as follows: • 20Mt case – 6 year mine life o Housing compensation $5.3M o Plant Sustaining Capital $2.5M • 10Mt case – 4 year mine life o Housing compensation $3.8M o Plant Sustaining Capital $1.7M • 100m buffer case – 3 year mine life o Housing compensation $3.0M o Plant Sustaining Capital $1.5M
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21.17 Operating Cost Estimate Mining operating costs have been benchmarked against the costs of operating mines of a similar size and nature. The restricted hours of operation (5 days per week / daylight hours only) is expected to attract a cost premium, this is yet to be quantified. However, the estimate used here is believed to be indicative and aligned with the intended targeted study accuracy. The cut-off grade applied (0.44g/t) in this instance is estimated based on administration and processing costs only. Sustaining capital for the mill has not been included to estimate the cut- off grade. This is partly due to the underutilisation of the mill. Current plans indicate that there is excess capacity in the mill and the plan assumes that the decision to continue operations is given. Therefore, while there continues to be excess capacity in the Mill, any grade above a cut- off grade based on variable costs only will contribute to any fixed operating or sustaining capital requirements.
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22 ECONOMIC ANALYSIS Mining One Pty Ltd (Mining One, M1) was approached by Johan Booyse of Stawell Gold Mines to carry out a pit and dump design on the Davis and Big Hill deposits, Victoria. As outlined in Section 16 of this report, the Whittle Optimisation is essentially concerned with maximising the operating cash flow from the mine. A subsequent cash flow model is required to further evaluate the project consider the following: ¾ Sustaining Capital Cost ¾ Environmental rehabilitation costs ¾ Potential Cash flow ¾ Net present value ¾ Internal rate of return ¾ Project duration, including rehabilitation ¾ Open pit foot print / open pit size ¾ Proximity to community and infrastructure Three (3) potential cases were identified from the Whittle Optimisation. The results of these potential cases are presented in Table 22-1 below (Pit shells not designed pits):
Table 22-1 Whittle Pit Optimisation Results at $1400/oz Description Unit 20Mt Case 10Mt Case 100m Buffer Case Mineralised Tonnes 3,731,028 2,455,936 2,464,949 Au Grade g/t 1.59 1.71 1.40 Waste Tonnes 13,936,475 6,564,192 9,196,734 Total Rock Tonnes 17,667,503 9,020,128 11,661,683 Stripping Ratio 3.73 2.67 3.73 Total Undiscounted Value* A$ 116,236,981 100,709,465 59,792,190 *Excludes required capital High level schedules were conducted based on a North to South mining sequence for each of the cases. The North to South Mining sequence was selected considering community impact, waste dump sequencing and minimising negative cash drawdown. The aforementioned scenarios assume a 5 day per week, day shift only roster and a maximum bench turn-over rate of 80 m per year. As outlined in Section 16 the pit shells are designed and subsequently scheduled. It is not uncommon to lose apparent value during the design process. This is primarily due to additional practical considerations required to extract the resources (i.e. minimum mining width, ramps, etc.). Furthermore, critical to the project returns is the capital required to extract the resources. Capital requirements have been estimated as follows: ¾ Initial Capital before and during project start-up - : ¾ Communications tower $1.05M
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¾ Dams- $3.0M ¾ Noise Barrier- $3.5M ¾ Noise and dust monitoring -$ 0.10M ¾ Purchase of houses - $ 0.5M ¾ Permitting and Tech. Studies - $ 2.5M ¾ Compensation for houses -$1.0M ¾ Big Hill & Davis Oxide Studies-$0.2M ¾ PFS Studies -$1.1M ¾ Wall raise of TSF2 - $1.1M ¾ Geotechnical Drilling - $0.5M ¾ Equipment Mobilisation - $1.8M ¾ Project Management - $0.3M ¾ Plant sustaining capital - $0.5 These values give a cost of $17.15M before the project starts. The following additional capital will be required during and at completion of the project: ¾ Plant sustaining capital of $0.5M per year of operation. ¾ Houses compensation $1.0M per year of operation. ¾ Re-establishment of roads on Big Hill at end of year two - $0.5M. ¾ Fire outlook tower at end of project -$.1M The total capital required for the Big Hill project is $21.71M. The pit (design), post capital schedule results are shown in Table 22-2 below: Table 22-2 Pit Schedule Results
100m Buffer Description Unit 20Mt Case 10Mt Case Case Mill Feed t 3,470,546 2,277,200 2,317,587 Input Grade Au g/t 1.52 1.65 1.39 Mill Input Ounces Au oz 170,530 120,590 104,142 Mill Recovered Ounces AU oz 153,477 108,531 93,728
Mining Cost $ -73,371,063 -36,582,481 -42,424,778 Mill Cost $ -52,058,161 -34,158,006 -34,763,778 Rehandle Cost $ -20,788,996 -13,873,640 -25,335,771
Revenue Au $ 214,867,986 151,943,087 131,219,072 Undiscounted Cash flow $ 44,685,448 45,614,643 7,955,408 Discounted Cash flow (Inclusive of capital)* $ 36,748,767 39,647,998 6,419,168 IRR % 664 812 69 *Please note that a 10% discount rate has been applied to this project
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As can be seen in Table 22-2 above, the 10Mt case yields the highest discounted cash flow ($40M). This value is inclusive of capital requirements, and minimise both community and environmental exposure. The 10Mt case has the following advantages over the 20Mt case with respect to the environmental and community issues: ¾ Smaller open pit foot print, resulting in less encroachment on residential areas; ¾ Shorter project duration; ¾ Less waste management considerations, as the northern pit may be used as a waste dump facility; ¾ Rehabilitation may commence earlier due to the northern pit being backfilled during mining of the southern pit; and ¾ Less stockpiling considerations. This case also assists with reducing waste rehandle costs.
Discounted Cash Flow 15 60,000
50,000 10 40,000 5 30,000 (000s)
(000s)
Cash
‐ 20,000 Flow
10,000
Cash ‐5,000
0 Cumulative ‐10,000 ‐10,000
‐15,000 ‐20,000
Discounted Cash Flow Cum. Discounted Cash Flow
Figure 22-1: Annual and Cumulative Discounted Cashflow
Due to the level of confidence associated with the input parameters for the study, the results should only be used as an indication of a potential open pit opportunity available due to the exploitation of the Big Hill Project.
22.1 Taxes Stawell Gold Mines is currently subject to the following taxes and duties: ¾ Victorian Payroll tax of payable at the rate of 4.90% calculated on wages paid by SGM to its employees. Stawell Gold Mines is liable for Victorian Payroll Tax when its total Australian wages exceeds the Victorian general deduction threshold of AUD$45,833 a
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month or AUD$550,000 over a full financial year and/or when grouped with other businesses of the Corporation, the combined Australian wages of the group exceed the general deduction threshold level. ¾ Minor taxes and duties including land tax, insurance duty, mortgages duty, motor vehicle duty debits tax and local council rates. The Corporation is currently subject to the following taxes: ¾ Federal Income tax is levied on the taxable income of the Corporation at a rate of 30%. In general terms, taxable income is calculated on assessable income less any allowable deductions. ¾ Capital Gains Tax (CGT) is paid on any capital gain that the Corporation includes in its annual income tax return. CGT is not a separate tax; it’s a component of income tax. ¾ Fringe Benefits Tax (FBT) is payable for benefits paid to an employee or the employee’s associate by the Corporation. FBT is separate from income tax and is based on the taxable value of the various benefits provided. ¾ Goods and Services Tax (GST) is a broad-based tax of 10 per cent on the sale of most goods and services and other things in Australia. Being registered for GST enables the Corporation to claim input tax credits for the GST included in the purchase price of goods and services used in the business. ¾ Victorian Payroll Tax of payable at the rate of 4.90% calculated on wages paid it and its subsidiaries to its employees. Corporation is liable for Victorian Payroll Tax when its total Australian wages exceeds the Victorian general deduction threshold of AUD$45,833 a month or AUD$550,000 over a full financial year and/or when grouped with other businesses, the combined Australian wages of the group exceed the general deduction threshold level. ¾ Minor taxes and duties including land tax, insurance duty, mortgages duty, motor vehicle duty debits tax and local council rates. ¾ Stawell Gold Mine is part of the Northgate Australian Venture Corporation Pty Ltd tax consolidation group, and will therefore be able to access significant non-capital losses to reduce taxable income and any taxes payable.
22.2 Royalties Within M5260 there is an AUD$2.00 per gold ounce royalty payable to Mineral Ventures of Australia (MVA). This royalty agreement came into place in February of 2004 and is in place until the earlier of 15 years of production or 2.5 million ounces of gold produced. Furthermore, this royalty agreement extends to Victorian tenements held by Leviathan Resources in February 2004 which included MIN5260. These royalties have not been considered in the economic analysis considered here.
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23 ADJACENT PROPERTIES There are no adjacent properties to the SGM.
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24 OTHER RELEVANT DATA AND INFORMATION Not Applicable
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25 INTERPRETATION AND CONCLUSIONS Stawell Gold Mines has in place rigorous processes for the estimation of Mineral Resources and Mineral Reserves. These processes are managed on site and appropriate resources (personnel and drilling) are allocated to this function. The conclusion of the Qualified Person as defined in this report is that the Mineral Resources as stated in this document are valid and supported by appropriate data collection, sampling, processing, interpretation and estimation methodologies and conform to NI 43-101 guidelines. The pit optimisations that were conducted for this preliminary economic assessment are preliminary in nature, they include inferred mineral resources and are considered too speculative geologically and technically to have the economic considerations applied to them that would enable them to be categorised as mineral reserves. There is no certainty that this preliminary economic assessment will be realised. The Stawell project has significant production history and the site personnel significant knowledge of the ore body and the mining process which adds additional support to the methodologies used to estimate the Mineral Resources and Mineral Reserves.
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26 RECOMMENDATIONS It is recommended that specified case (10Mt) proceed to prefeasibility as it provides the best economic outcome while prudently considering the potential impacts on the community. Specific focus of the feasibility study should be on, but not limited to: • Geotechnical confidence; • Operating productivities and work hour restrictions; • Operating costs; • Community impact; • Rehabilitation effort and final landform requirements. Mining One have generated estimates for the Geotechnical, Hydrogeological and Mining study components for a prefeasibility study. It is envisaged that the process from initiation of the study to the commencement of Mining Operations would take approximately 12 months covering disciplines such as Geotechnical, Hydrogeological and Mining technical studies. Provision ($4.6M) has been made for costs including but not limited to; drilling costs, government approvals and permitting costs and infrastructure relocation and removal study costs. Figure 26-1 is an estimate of what sort of timing might be expected to study, approve and execute start-up of the project.
YR1 -QTR1 YR1 -QTR2 YR1 -QTR3 YR1 -QTR4 YR2 -QTR1 Drilling Geotechnical Studies Hydrogeological Studies Mining Studies Approvals Contract Award Equipment Mobilisation Site Establishment Commencement of Mining Figure 26-1: Study and Project Execution Timeline It is recommended that a complete risk assessment to be conducted including the assessment of financial, community and legislative risk. Such an approach may resemble the following: 1) Confirm mining costs through a formal Request for Pricing (RFQ) process sourcing quotes from prospective local contractors based on the restrictions to operational hours per week; 2) Initiate community impact and cost estimate studies to better understand potential land purchase and civil infrastructure relocation costs; 3) Initiate geotechnical and hydrogeological studies to better understand the pit crest location and impact on surrounding infrastructure; 4) Initiate work to better understand / confirm the underground voids in the Big Hill and Davis areas;
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5) Initiate work to model and better understand the potential impact of dilution and mineralised material loss on the project; 6) Cultural, environmental, and government factors should be further studied to asses any impact on a potential pit.
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27 REFERENCES Arne, D.C., Bierlein, F.P., McNaughton, N., Wilson, C.J.L., Morland, V.J. 1998. Timing of gold mineralisation in western and central Victoria, Australia: New constraints from SHRIMP II analysis of zircon grains from felsic intrusive rocks. Ore Geology Review, 13, 251 273 Bird, P.J., (1986), The tectonic and intrusive geological history of the Great Western - Stawell Area: Central Western Victoria. Honours Thesis (unpublished). Latrobe University, 121p Cas, R.A.F. 1983. A review of the palaeogeographic and tectonic development of the Palaeozoic Lachlan Fold Belt of southeastern Australia. Geological Society of Australia, Special Publication, 10 Coffey Environments (2008), Stawell Gold Mines – Stage 4(a) – Hydrogeological Assessment, Tailings Storage Facility No 2 Factual Report (Draft). 621.0/10/1.
Crawford, A. J. 1988. The Cambrian system in Victoria. Geology of Victoria. J. G. Douglas, and Ferguson, J.A. Melbourne, Geology Society of Australia, Victoria Division Special Publications: 37 62 Doronila, A., (2006). Phytostabilisation of arsenic – rich sulphidic gold mine tailings in Victorian goldfields, Australia. Ph.D. Thesis submitted to the School of Botany, University of Melbourne. Melbourne, Australia.
Dowsley. K., (2001). Characterisation of a partially oxidised sulphidic tailings dam. Honours Thesis (unpublished). School of Earth Sciences. University of Melbourne.
Foster D.A., Gray, D.R., Kwak, T.A.P., Bucher, M. 1998. Chronology and tectonic framework of turbidite hosted gold deposits in western Lachlan fold belt, Victoria: 40Ar 39Ar results: Ore Geology Review, 13, 229 250 Gane, M.J., (1998) Gold mineralisation within the basalt contact ore zones, Magdala Mine, Stawell, Victoria. Masters Thesis (unpublished). The University of Melbourne, 208p Gardner, K. 2008. Magdala block “S600” Resource Model Report, May 2008, unpublished internal company report. Gedge, L. 1997. The relationships between structure, gold mineralisation and intrusive events at Stawell, Victoria. Unpublished thesis, Melbourne, Australia, The University of Melbourne, 95p Fredericksen, D., Miller, G., Dincer, T., Technical Report for Stawell Gold Mine, Victoria Australia (28 March 2008), unpublished internal company report. Hamilton, C. 2011. Wonga Stage 3 – Feasibility Investigation, internal unpublished AuRico Gold company report. Jupp, B. 2003. Hydrothermal alteration and lithogeochemistry of the Kewell and Wallup prospects and their comparison with the Magdala gold, Stawell, Victoria. Unpublished thesis, Melbourne, Australia, The University of Melbourne, 193p Kaufman, A. 2003. The volcano sedimentary and structural evolution of the Wildwood prospect, Western Lachlan Orogen. Unpublished thesis, Melbourne, Australia, The University of Melbourne, 85p
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LanePiper. (2008). 207119Report 01.2 Environmental Audit Report Stawell Gold Mine Leviathan Rd, Stawell Vic, January 2008.
Mapani, S.E., Wilson, C.J.L. 1994. Structural evolution and gold mineralization in the Scotchmans Fault Zone, Magdala Gold Mine, Stawell, Western Victoria, Australia. Economic Geology, 89, 566 583 Miller, J.McL., Wilson, C.J.L, Dugdale, L.J. 2006. Stawell gold deposit: a key to unrasvelling Cambrian to Early Devonian structural evolution of the western Victorian goldfields, Australian Journal of Earth Sciences, 53, 677 695 Miller, J.McL., Wilson, C.J.L. 2004a. Structural Analysis of faults related to a heterogeneous stress history: reconstruction of a dismembered gold deposit, Stawell, western Lachlan Fold Belt, Australia. Journal of Structural Geology, 26, 1231 1256 Miller, J.McL., Wilson, C.J.L. 2004b. Stress controls on intrusion related gold lodes: Wonga gold mine, Stawell, Western Lachlan Fold Belt, southeastern Australia. Economic Geology, 99, 941 963 Miller, J. McL., Wilson, C.J.L. 2002. The magdala lode system, Stawell, Southeastern Australia: Structural style and relationship to gold mineralization across the western Lachlan Fold Belt. Economic Geology, 97, 325 349 NQ, Groundwater & Environment (2011). Modelling of physical groundwater control options and cost estimates for ozone remediation at Stawell Gold Mines, July 2011.
NQ, Groundwater & Environment (2012). Review of Conceptual Groundwater Model with reference to the detection of SCN in Bore PS585.
Oldmeadow, D., (2008). Geochemical evolution of experimental tailings rehabilitations systems at Stawell Gold Mine, Victoria, Australia: implications for the use of thin composite cover in storage of sulphidic gold mine tailings. Thesis, Curtin University of Technology.
Pritchard, E.G. 2001. The Magdala Basalt in the east exploration decline, Magdala gold mine, Stawell Victoria: Petrography, alteration paragenesis and structural style. School of Earth Sciences, The University of Melbourne, 68p Quick, D., 1988. The Stawell Goldfield, in Jones, D.G. (ed.), Central Victorian Gold Deposits. Bicentennial Gold 88, Excursion No. 2 Guide. The Geology Department and University Extension, The University of Western Australia, Publication No. 13, 55-60 Richards, J.R. Singleton, O.P., (1981) Palaeozoic Victoria, Australia: igneous rocks, ages and their interpretation. Journal of the Geological Society of Australia, 28, 395-421 Robinson, J.A., Wilson, C.J.L, Rawling T.J. 2006. Numerical modelling of an evolving gold lode system: structural and lithological controls on ore shoot formation in the Magdala goldmine, western Victoria, Australian Journal of Earth Sciences, 54, 799 – 823. Robinson, J.A. 2005 Nature of the mineralised (ore shoot) environment within the Magdala gold deposit, western Lachlan Fold Belt, Australia, Unpublished thesis, Melbourne, Australia, The University of Melbourne, 373p Rockwater., (2010). Summary Report on Hydrogeological Assessment and Modelling of Possible Seepage from Tailings Storage Facility No2. 621.0/101, March 2010.
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Schaubs, P.M., Wilson, C.J.L. 2002. The relative roles of folding and faulting in controlling gold mineralisation along the Deborah anticline, Bendigo, Victoria, Australia. Economic Geology, 97, 351 370 Squire, R.J., 2004. Stawell Au deposit – ARC Linckage Project: June 2004 progress report, Unpublished Squire, R.J., Wilson, C.J.L. 2005. Tectonic responses to super continent formation: correlation of Cambrian geological events along proto Pacific margin of East Gondwana. Journal of the Geological Society, London, 162, 749 761 Stewart, M. 2012. Stawell Gold Mines, Review of Big Hill Resource Estimate, CRG21204, unpublished internal Quantitative Group Consultant report Vandenberg, A.H.M., Willman, C.E., Maher, S., Simons, B.A., Cayley, R.A., Taylor, D.H., Morand, V.J., More, D.H., and Radojkovic A. 2000. The Tasman Fold Belt System in Victoria, Geological Survey of Victoria Special Publication Vann, J, Jackson, S, and Bertoli, O. 2003. Quantitative Kriging Neighbourhood Analysis for the Mining Geologist — A Description of the Method with worked case examples, in Proceedings of the, 5th International Mining Geology Conference (Australasian Institute on Mining and Metallurgy, Melbourne) Watchorn, R.B., Wilson, C.J.L 1989. Structural setting of the gold mineralisation at Stawell, Victoria, Australia. Economic Geology Monographs, 6, 292 309 Wilson, C.J.L., Will, T.M., Cayley, R.A., Chen, S. 1992. Geologic framework and tectonic evolution in Western Victoria, Australia. Tectonophysics, 214, 93 127 Wonga Geology and Resource Model Report May 2009, finished April 2012, unpublished internal company report. Wonga North Resource Model Report, April 2012, unpublished internal company report. Xu, G., Powell, R., Wilson, C.J.L., Will, T.M. 1994. Contact metamorphism around the Stawell granite, Victoria, Australia. Journal of Metamorphic Geology, 12, 609 624
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DOCUMENT INFORMATION Status Final
Version 6
Print Date 25th January 2013
Author((s) Dean Basile
Reviewed By Bill Frazer, Mark Van Leuven
Pathname P:\1723_M - Stawell Long Term Planning AuRico Gold\WPO\3438_Final_w_signatures_Rev6.docx
File Name 3438.doc
Job No 1723_M
Distribution PDF emailed to client
DOCUMENT CHANGE CONTROL Version Description of changes/amendments Author (s) Date
1 Dean Baasile 20/12/12
5 Final Dean Basile 16/1/13
6 Following added to tables 1-22 and 22-2 *Please note Dean Basile 25/1/13 that a 10% discount rate has been applied to this project
DOCUMENT REVIEW AND SIGN OFF Version Reviewer Position Signature Date
5 Bill Frazer Director 16/1/13
5 Mark Van Leuven Principal Mining Engineer 16/1/13
5 Mark Van Leuven Principal Mining Engineer 25/1/13
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