NI 43-101 TECHNICAL REPORT FOR THE KWANIKA …Prepared for Serengeti Resources Inc. Page 1 of 139 NI...

Serengeti Resources Inc. Kwanika Project

NI43-101 Technical Report for the Kwanika Project Prepared for Serengeti Resources Inc.

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NI 43-101 TECHNICAL REPORT FOR THE

KWANIKA PROPERTY

PRELIMINARY ECONOMIC ASSESSMENT 2013

Near Fort St. James, British Columbia, Canada

Centred at 55’ 31’ N and 125’ 20’ W

Submitted to:

Serengeti Resources Inc.

March 4, 2013

Moose Mountain Technical Services James H. Gray, P.Eng.

Heather M. Robillard, P.Eng.

1975 1st Ave. South

Cranbrook, B.C. V1C 6Y3 Canada

Tel: 250.489.1212



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TABLE OF CONTENTS

1.0 Summary .......................................................................................................................................9 1.1 Property Description and Location ............................................................................................................ 9 1.2 Property Ownership ................................................................................................................................... 9 1.3 Accessibility, Climate, Local Resources, Infrastructure and Physiography ............................................ 10 1.4 History ..................................................................................................................................................... 10 1.5 Geological Setting and Mineralization .................................................................................................... 10

1.5.1 Geology ............................................................................................................................................... 10 1.5.2 Mineralization ..................................................................................................................................... 11

1.6 Deposit Type ........................................................................................................................................... 11 1.7 Exploration .............................................................................................................................................. 11 1.8 Drilling .................................................................................................................................................... 12 1.9 Sample Preparation, Analyses and Security ............................................................................................ 12 1.10 Data Verification ..................................................................................................................................... 12 1.11 Mineral Processing and Metallurgical Testing ........................................................................................ 12 1.12 Mineral Resource Estimates .................................................................................................................... 13

1.12.1 Central Zone ....................................................................................................................................... 13 1.12.2 South Zone .......................................................................................................................................... 14 1.12.3 Open Pit Mine Planning ..................................................................................................................... 15 1.12.4 Underground Mine Planning .............................................................................................................. 16

1.13 Mineral Reserve Estimates ...................................................................................................................... 17 1.14 Mining Methods ...................................................................................................................................... 17

1.14.1 Rock Storage Facility .......................................................................................................................... 18 1.14.2 Mining Operations .............................................................................................................................. 18

1.15 Recovery Methods ................................................................................................................................... 21 1.16 Project Infrastructure ............................................................................................................................... 21

1.16.1 Access ................................................................................................................................................. 21 1.16.2 On-Site Transport Road ...................................................................................................................... 21 1.16.3 Main Facilities .................................................................................................................................... 21 1.16.4 Power Supply and Distribution ........................................................................................................... 21 1.16.5 Mine Rock and Tailing Storage Facilities .......................................................................................... 21 1.16.6 Mine Area Water Management ........................................................................................................... 22 1.16.7 Mine Area Closure Plan ..................................................................................................................... 22

1.17 Market Studies and Contracts .................................................................................................................. 24 1.18 Environmental Studies, Permitting and Social or Community Impact .................................................... 24

1.18.1 Regulatory Framework ....................................................................................................................... 24 1.18.2 Regional Land Use Processes ............................................................................................................. 24 1.18.3 Programs Already in Progress ........................................................................................................... 24

1.19 Capital and Operating Costs .................................................................................................................... 25 1.19.1 Capital Costs ...................................................................................................................................... 25 1.19.2 Operating Costs .................................................................................................................................. 25

1.20 Economic Analysis .................................................................................................................................. 26 1.20.1 Sensitivity Analysis ............................................................................................................................. 28

1.21 Adjacent Properties ................................................................................................................................. 29 1.21.1 Regional .............................................................................................................................................. 29 1.21.2 Local District ...................................................................................................................................... 29

1.22 Interpretation and Conclusions ................................................................................................................ 30 1.22.1 Geology and Resource Modeling ........................................................................................................ 30 1.22.2 Metallurgy .......................................................................................................................................... 30 1.22.3 Underground Mine Plan ..................................................................................................................... 31



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1.22.4 Open Pit Mine Plan ............................................................................................................................ 31 1.22.5 General ............................................................................................................................................... 31 1.22.6 Regional and Other On-sight Opportunities ....................................................................................... 31

1.23 Recommendations ................................................................................................................................... 32 1.23.1 Optimization Studies ........................................................................................................................... 32 1.23.2 Pre-Feasibility Study .......................................................................................................................... 32 1.23.3 Study Costs.......................................................................................................................................... 33

2.0 Introduction ................................................................................................................................ 34 3.0 Reliance on Other Experts ........................................................................................................ 35 4.0 Property Location and Description .......................................................................................... 36 5.0 Accessibility, Climate, Local Resources, Infrastructure and Physiography ........................ 39 6.0 History ......................................................................................................................................... 40 7.0 Geological Setting and Mineralization ..................................................................................... 41

7.1 Geology ................................................................................................................................................... 41 7.2 Mineralization .......................................................................................................................................... 41

8.0 Deposit Type ............................................................................................................................... 43 9.0 Exploration ................................................................................................................................. 44 10.0 Drilling ........................................................................................................................................ 45 11.0 Sample Preparation, Analyses and Security ............................................................................ 46 12.0 Data Verification ........................................................................................................................ 47 13.0 Mineral Processing and Metallurgical Testing ....................................................................... 48 14.0 Mineral Resource Estimates ..................................................................................................... 50

14.1 Central Zone ............................................................................................................................................ 50 14.2 South Zone .............................................................................................................................................. 51 14.3 Economic Pit Limit and Pit Designs ........................................................................................................ 53

14.3.1 Pit Optimization Method ..................................................................................................................... 53 14.3.2 Economic Pit Limit Design Basis ....................................................................................................... 53 14.3.3 Pit Slope Angles .................................................................................................................................. 54 14.3.4 Process Recoveries ............................................................................................................................. 54 14.3.5 Metal Prices ........................................................................................................................................ 54 14.3.6 LG Economic Limits ........................................................................................................................... 55 14.3.7 Underground Mine Design ................................................................................................................. 61 14.3.8 Further Work towards Open Pit Design ............................................................................................. 62 14.3.9 Further Work towards Underground Design ...................................................................................... 62

15.0 Mineral Reserve Estimates ........................................................................................................ 64 16.0 Mining Methods ......................................................................................................................... 65

16.1 Summary ................................................................................................................................................. 65 16.2 Introduction ............................................................................................................................................. 65 16.3 Mining Datum ......................................................................................................................................... 66 16.4 Production Rate ....................................................................................................................................... 66 16.5 Mine Planning 3D Block Model .............................................................................................................. 66

16.5.1 Net Smelter Return .............................................................................................................................. 69 16.5.2 Mining Loss and Dilution ................................................................................................................... 70 16.5.3 Detailed Pit Designs ........................................................................................................................... 71 16.5.4 Haul Road Widths ............................................................................................................................... 71 16.5.5 Design Standards ................................................................................................................................ 74 16.5.6 LG Phase Selection ............................................................................................................................. 74 16.5.7 Kwanika Detailed Pit Phase Designs ................................................................................................. 75 16.5.8 Kwanika Central Underground Mining .............................................................................................. 81 16.5.9 Pit Delineated Resources .................................................................................................................... 81

16.6 Mine Plan ................................................................................................................................................ 81



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16.6.1 LOM Production Schedule .................................................................................................................. 81 16.6.2 Rock Storage Facilities ....................................................................................................................... 86 16.6.3 Mine Pre-Production Detail ............................................................................................................... 87 16.6.4 Mine Production Detail ...................................................................................................................... 89

16.7 Mine Operations ...................................................................................................................................... 94 16.8 General Organization ............................................................................................................................... 94

16.8.1 Direct Mining Area ............................................................................................................................. 95 16.8.2 Mine Maintenance Area ...................................................................................................................... 98 16.8.3 General Mine Expense Area ............................................................................................................... 98

16.9 Mine Closure and Reclamation ............................................................................................................... 99 16.10 Mine Equipment .................................................................................................................................... 100

16.10.1 Drilling Equipment ........................................................................................................................... 101 16.10.2 Blasting Equipment Facilities ........................................................................................................... 102 16.10.3 Loading and Hauling Equipment ...................................................................................................... 102 16.10.4 Dewatering Equipment ..................................................................................................................... 102 16.10.5 Mine Ancillary Facilities .................................................................................................................. 103

17.0 Recovery Methods .................................................................................................................... 104 18.0 Project Infrastructure ............................................................................................................. 105

18.1 General Site Geotechnical Investigation ................................................................................................ 105 18.2 Site Layout ............................................................................................................................................ 105 18.3 Site Roads .............................................................................................................................................. 107

18.3.1 Access Road ...................................................................................................................................... 107 18.3.2 On Site Transport Road .................................................................................................................... 107

18.4 Process Plant and Process Related Facilities ......................................................................................... 107 18.5 Ancillary Buildings ............................................................................................................................... 107 18.6 Communications System ....................................................................................................................... 107 18.7 Power Supply......................................................................................................................................... 107 18.8 Power Distribution ................................................................................................................................. 108

18.8.1 Plant Site ........................................................................................................................................... 108 18.8.2 Remote Loads .................................................................................................................................... 108 18.8.3 Ancillary Systems .............................................................................................................................. 108

18.9 Mine Rock, Tailing, and Water Management ........................................................................................ 108 18.9.1 Solids Management ........................................................................................................................... 108 18.9.2 Water Management ........................................................................................................................... 110 18.9.3 Kwanika Creek Diversion ................................................................................................................. 110 18.9.4 Pre-Production Earthworks for Mine Rock and Water Management ............................................... 111 18.9.5 Post-Production Earthworks for Mine Rock and Water Management ............................................. 112

19.0 Market Studies and Contracts ................................................................................................ 113 20.0 Environmental Studies, Permitting and Social or Community Impact .............................. 114

20.1 Regulatory Framework .......................................................................................................................... 114 20.2 Regional Land Use Processes ................................................................................................................ 114 20.3 Environmental and Corporate Social Responsibility Programs Already in Progress ............................ 114 20.4 Fisheries Resources and Permitting Issues ............................................................................................ 115

21.0 Capital and Operating Cost Estimates................................................................................... 117 21.1 Capital Cost Estimate ............................................................................................................................ 117

21.1.1 Exclusions ......................................................................................................................................... 117 21.1.2 Open Pit Mine Capital Costs ............................................................................................................ 118 21.1.3 Underground Mine Capital Costs ..................................................................................................... 118 21.1.4 Mining Basis of Estimate .................................................................................................................. 118 21.1.5 Assumptions ...................................................................................................................................... 119 21.1.6 Contingency ...................................................................................................................................... 119

21.2 Operating Cost Estimate ........................................................................................................................ 119



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21.2.1 Open Pit Mine Operating Costs ........................................................................................................ 120 21.2.2 Mine Fuel Consumption .................................................................................................................... 121 21.2.3 Underground Operating Costs ......................................................................................................... 121 21.2.4 Process Operating Costs .................................................................................................................. 121 21.2.5 General and Administrative .............................................................................................................. 122

22.0 Economic Analysis ................................................................................................................... 123 22.1 Introduction ........................................................................................................................................... 123 22.2 Pre-Tax Model ....................................................................................................................................... 123

22.2.1 Metal Production Financial Model .................................................................................................. 123 22.2.2 Financial Evaluations of NPV and IRR ............................................................................................ 125 22.2.3 Sensitivity Analysis ........................................................................................................................... 126 22.2.4 Royalties ........................................................................................................................................... 128

23.0 Adjacent Properties ................................................................................................................. 129 23.1 Regional ................................................................................................................................................. 129 23.2 Local District ......................................................................................................................................... 129

24.0 Other Relevant Data and Information ................................................................................... 130 25.0 Interpretations and Conclusions ............................................................................................ 131

25.1 Geology and Resource Modeling .......................................................................................................... 131 25.2 Metallurgy ............................................................................................................................................. 131 25.3 Underground Mine Plan ........................................................................................................................ 132 25.4 Open Pit Mine Plan ............................................................................................................................... 132 25.5 General .................................................................................................................................................. 132 25.6 Opportunities ......................................................................................................................................... 133

26.0 Recommendations .................................................................................................................... 134 26.1 Optimization Studies ............................................................................................................................. 134 26.2 Pre-Feasibility Study ............................................................................................................................. 134 26.3 Future Engineering Study Costs ............................................................................................................ 134

27.0 References ................................................................................................................................. 135 28.0 Certificates of Qualified Persons ............................................................................................ 136

LIST OF APPENDICES

APPENDIX A Block Caveability Assessment

APPENDIX B Property Ownership

APPENDIX C Mining

APPENDIX D Financial Analysis

Note: Appendix B, C, and D are available upon request at the Serengeti Resources Inc. offices in Vancouver.



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LIST OF TABLES

Table 1-1 Kwanika Process Recovery Assumptions .........................................................................13 Table 1-2 Central Zone Mineral Resources – December 31, 2010....................................................14 Table 1-3 South Zone Mineral Resources – December 31, 2010 ......................................................15 Table 1-4 Metal Prices and NSP – 3 Year Trailing Average Sept 1, 2011 ........................................15 Table 1-5 Run of Mine LG Pit Delineated Resource Including Indicated and Inferred Classes .......16 Table 1-6 Run of Mine Underground Delineated Resource Including Indicated and Inferred Classes

...........................................................................................................................................17 Table 1-7 Summarized Indicated and Inferred Pit Mill Feed for Production Scheduling .................17 Table 1-8 Summarized Indicated and Inferred Underground Mill Feed for Production Scheduling 18 Table 1-9 Life of Mine Production Summary ...................................................................................20 Table 1-10 Metal Prices and NSP – 3 Year Trailing Average Oct 15, 2012 .......................................24 Table 1-11 Initial Capital Cost Summary ............................................................................................25 Table 1-12 Sustaining Capital Cost Summary .....................................................................................25 Table 1-13 Operating Cost Summary ..................................................................................................26 Table 1-14 Metal Production from Kwanika .......................................................................................27 Table 1-15 Summary of the Economic Evaluation ..............................................................................27

Table 13-1 Kwanika Process Recovery Assumptions .........................................................................49

Table 14-1 Central Zone Mineral Resources – December 31, 2010....................................................51 Table 14-2 South Zone Mineral Resources – December 31, 2010 ......................................................53 Table 14-3 Kwanika Input Unit Costs for LG Pits ..............................................................................54 Table 14-4 Kwanika Proposed Bench Design Criteria ........................................................................54 Table 14-5 Kwanika Process Recovery Assumptions .........................................................................54 Table 14-6 Metal Prices and NSP Base Case – Sept 2011 ..................................................................55 Table 14-7 ROM LG Pit Delineated Resource Including Indicated and Inferred Resources ..............56 Table 14-8 Run of Mine Pit Delineated Resources from Detailed Designs ........................................57 Table 14-9 Run of Mine Underground Delineated Resource Including Indicated and Inferred Classes

...........................................................................................................................................61

Table 16-1 Summarized Indicated and Inferred Pit and Underground Mill Feed for Production

Scheduling .........................................................................................................................65 Table 16-2 Pit and Underground Delineated Resource by Assurance of Existence Class ..................65 Table 16-3 3DBM Items ......................................................................................................................67 Table 16-4 Metal Prices and NSP from September 2011 ....................................................................70 Table 16-5 Material Types Defined for MS-SP ..................................................................................82 Table 16-6 Life of Mine Production Summary ...................................................................................84 Table 16-7 Mine Waste Rock Tonnages by Area and Year (kt) .........................................................87 Table 16-8 Blasting Assumptions ........................................................................................................97 Table 16-9 Major Mine Equipment Requirements ............................................................................100 Table 16-10 Production Drilling Assumptions ....................................................................................101

Table 19-1 Metal Prices and NSP – 3 Year Trailing Average Oct 15, 2012 .....................................113

Table 21-1 Initial Capital Cost Summary ..........................................................................................117



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Table 21-2 Sustaining Capital Cost Summary ...................................................................................117 Table 21-3 Open Pit Mine Initial Capital Costs ................................................................................118 Table 21-4 Underground Mine Initial Capital Costs .........................................................................118 Table 21-5 Open Pit Mine Mobile Equipment Initial Capital Schedule ............................................118 Table 21-6 Underground Mine Mobile Equipment Initial Capital Schedule ....................................119 Table 21-7 Operating Cost Summary ................................................................................................120 Table 21-8 Open Pit Mining Costs per Tonne Material Mined .........................................................121

Table 22-1 Metal Prices and NSP – 3 Year Trailing Average Oct 15, 2012 .....................................123 Table 22-2 Metal Production from Kwanika Project .........................................................................124 Table 22-3 Summary of the Economic Evaluation ............................................................................124



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LIST OF FIGURES

Figure 1-1 General Site Layout ...........................................................................................................23 Figure 1-2 Base Case Sensitivity to Pre-Tax NPV @ 5% ..................................................................28 Figure 1-3 Base Case Sensitivity to Pre-Tax IRR ...............................................................................29 Figure 4-1 Kwanika Property Location ..............................................................................................37 Figure 4-2 Kwanika Claim Map .........................................................................................................38 Figure 14-1 Kwanika Central Zone Tonnes above CoG by LG Pit ........................................................55 Figure 14-2 Kwanika South Zone Tonnes above CoG by LG Pit ..........................................................56 Figure 14-3 Plan View of the Kwanika LG Pit Limits ...........................................................................58 Figure 14-4 Kwanika Economic Pit Limits – Orthographic View from the East .................................59 Figure 14-5 Kwanika Central LG & Detailed Pit Limits – NS Section at East 351500 Looking East .60 Figure 14-6 Kwanika South LG & Detailed Pit Limits – NS Section at East 352400 Looking East ...60 Figure 14-7 Kwanika Economic Underground Stopes – Orthographic View ......................................62

Figure 16-1 Kwanika Central Mine Planning Model Limits ................................................................67 Figure 16-2 Kwanika South Mine Planning Model Limits ...................................................................67 Figure 16-3 Kwanika Central Model Area with NSR >$11.90/t Grade Shell – Plan View with 5m

Topography Contours .......................................................................................................68 Figure 16-4 Kwanika South Model Area with NSR >$11.90/t Grade Shell – Plan View with 5m

Topography Contours .......................................................................................................69 Figure 16-5 Dual Lane High Wall Haul Road Cross Section ...............................................................72 Figure 16-6 Dual Lane External Haul Road Cross Section ..................................................................72 Figure 16-7 Single Lane High Wall Haul Road Cross Section .............................................................73 Figure 16-8 Single Lane External Haul Road Cross Section ................................................................73 Figure 16-9 Plan View of Kwanika Central Pit CK615 ........................................................................76 Figure 16-10 Plan View of Kwanika Central Pit CK625 ........................................................................77 Figure 16-11 Plan View of Kwanika South Pit P612 .............................................................................78 Figure 16-12 Plan View of Kwanika South Pit SK622 ...........................................................................79 Figure 16-13 North-South Section View of all Central Pits at East 351500 – Looking East .................80 Figure 16-14 North-South Section View of all South Pits at East 352400 – Looking East ....................80 Figure 16-15 Side View of Kwanika Underground Stopes ....................................................................81 Figure 16-16 ROM Mill Feed Sources and Mill Head Grades for Feed Cu, Au, Ag, and Mo ...............85 Figure 16-17 Cumulative Open Pit Strip Ratio .......................................................................................85 Figure 16-18 EOP – Time 00 End of Pre-production – Start of Production ...........................................91 Figure 16-19 EOP – Year 4 ....................................................................................................................92 Figure 16-20 EOP – LOM ......................................................................................................................93 Figure 16-21 General Organization Chart ..............................................................................................95 Figure 16-22 Haul Truck Fleet Size ......................................................................................................101

Figure 18-1 General Site Layout .........................................................................................................106 Figure 18-2 Proposed Kwanika Creek Diversion Channel – Valley Bottom Diversion .....................111

Figure 22-1 Undiscounted Annual and Cumulative Cash Flow .........................................................125 Figure 22-2 Base Case Sensitivity to Pre-Tax NPV @ 5% ................................................................127 Figure 22-3 Base Case Sensitivity to Pre-Tax IRR .............................................................................128



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1.0 Summary The Serengeti Resources Inc. Kwanika Project (the Project) involved the development of a copper-gold

deposit located near Fort St. James, British Columbia, Canada.

This National Instrument 43-101 (NI 43-101) compliant report on the Project has been prepared by

Moose Mountain Technical Services (MMTS) and is based on work produced by the following

independent consultants:

Roscoe Postle Associates Inc. (RPA), formerly Scott Wilson Roscoe Postle Associates Inc.

Moose Mountain Technical Services

David W. Rennie (P.Eng.) of RPA has visited the Project on May 25 - 27, 2008; January 28, 2010, and

October 30, 2010. He is the Qualified Person (QP) for all matters relating to the Mineral Resource

estimate.

Mr. Jim Gray (P.Eng.) of MMTS visited the Project site on October 18, 2011. He is the QP for all

matters relating to infrastructure, mining, mining capital costs, mining operating costs, financial

evaluation, and overall report preparation.

Mr. Tracey Meintjes (P.Eng.) of MMTS is the QP for matters relating to mineral processing, mineral

processing capital, mineral processing operating costs, and metallurgical testing.

The Preliminary Economic Assessment (PEA) is based on exploration and internal Serengeti studies

since 2004. The Resource model used is most recently described in an NI 43-101 Technical Report dated

March 3, 2011 which is referenced in this report. This PEA adds a mine plan and infrastructure bringing

it to a scoping level of accuracy. The mine plan includes an integrated open pit and underground

production schedule using typical operating parameters. Conceptual evaluation by AMEC has confirmed

that Block Caving is a suitable underground mining method as applied to this study. All dollar figures

presented in this report are stated in Canadian dollars unless otherwise specified.

1.1 Property Description and Location

The Kwanika property in north central British Columbia is situated in the Omineca Mining Division,

approximately 140km northwest (approximately 200km by road) of Fort St. James, located on NTS map

sheets 93N06 and 93N11, at latitude 55º31’ N and longitude 125º20’ W. The property is accessible year-

round by four-wheel-drive vehicle, provided there is active snow removal in winter.

The property is the host of two porphyry-style mineral deposits: the copper-gold-molybdenum-silver

South Zone and the copper-gold-silver Central Zone, both of which encompass current Mineral

Resources.

1.2 Property Ownership

The Kwanika property consists of 28 contiguous unpatented mineral claims covering an area of 8,960.29

ha and is 100% owned by Serengeti. It is not subject to any royalties or other outstanding liabilities.

Serengeti acquired the current extent of the property through staking between 2004 and 2006.



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1.3 Accessibility, Climate, Local Resources, Infrastructure and Physiography

The Kwanika Property is located approximately 75km to the southwest of the Kemess power line, and

B.C. Railway Company maintains an active rail line to Fort St. James. The Kwanika Project is also in

close proximity to the well-serviced communities of Prince George, Smithers, Fort St. James, and

Mackenzie. Access to the Kwanika Property from Fort St. James is via the all-weather Leo Creek and

Driftwood forest service roads (FSR) and the 30km long Tsayta Lake Road. Other infrastructure on the

Kwanika Property consists of dirt logging roads and several kilometres of excavated trails. There is

sufficient water available in the immediate vicinity of the property to support both exploration and

potential mining activities. Infrastructure currently consists of a 30-man exploration camp.

Serengeti has developed a beneficial association with the local Takla Lake First Nation and that there is

community support for the Kwanika Project and the potential employment that it would provide.

The average temperature for this area is 3.1ºC, with a peak average monthly temperature of 21.9°C in

July and an average monthly low of -15.8°C in January. The region receives an average of 295mm of

rainfall and 192cm of snowfall annually, with 138 days per year where precipitation exceeds 0.2mm. The

Kwanika property is snow-covered from late October to May.

1.4 History

Exploration on the Kwanika property dates back to the 1930’s and 40’s. Copper mineralization was first

recognized along Kwanika Creek in 1964 by Hogan Mines. Between 1966 and 1976, exploration was

carried out that included geological, geochemical, and geophysical surveys that resulted in an aggregate

of 5,700m of percussion and diamond drilling. In 1976, a Mineral Resource estimate for the main

(currently referred to as the South Zone) deposit was published.

Between 1981 and 1989, different operators (Placer Developments Ltd., Aume Resources Ltd. and Daren

Resources Ltd., Eastfield Resources Ltd.), conducted geochemical surveys and sampled rock outcrops, as

well as IP and drilling. The claims were allowed to lapse and, in 1995, the property was re-staked by

Discovery Consultants (Discovery) who conducted additional heavy mineral stream sediment and rock

sampling. No more work was done until Serengeti staked the property starting in 2004.

1.5 Geological Setting and Mineralization

1.5.1 Geology

The Kwanika property lies in the northern part of the Upper Triassic to Lower Jurassic Quesnellia

Terrane (Quesnel Trough) which is the host of numerous alkalic and calc-alkalic porphyry copper-gold

deposits within British Columbia. In the area around the Kwanika property, Quesnellia is bounded by the

Pinchi fault on the west and by the Manson fault on the east.

The Kwanika Project consists of two mineralized areas: the Central Zone and the South Zone. The

Central Zone deposit is characterized by the presence of two major and several minor intrusive bodies of

the multi-phase Hogem Batholith that intrude into successions of the Takla Volcanic Group and are host

to high grade copper-gold mineralization. The Takla Group andesites also host lower grade

mineralization throughout the property.



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The geology of the South Zone is also typified by various intrusive phases of the Hogem Batholith, with

deposits occurring within a fault bounded north-south trending structural corridor of strongly altered

intrusive rocks of alkalic to intermediate composition.

1.5.2 Mineralization

Copper and gold mineralization in the Central Zone at Kwanika occurs primarily in potassically and

albitically altered lithologies. Alteration and mineralization grade outwards from a strong to intensely

potassically and albitically altered, strongly mineralized core zone to a variably propylitically altered,

weakly mineralized periphery. Hypogene mineralization is controlled by several generations of quartz +

sulphide veining, with the highest copper and gold grades occurring in areas of quartz stockwork. A

supergene enrichment blanket has been superimposed on the upper surface of the hypogene

mineralization in the Central Zone.

The South Zone is characterized by porphyry style copper + gold + molybdenum + silver mineralization

within monzonite, quartz monzonite, and monzodiorite with primary mineralization comprised of fine to

coarse grained chalcopyrite disseminations and molybdenite mineralization along fractures and quartz

selvages and, less commonly, disseminated blebs associated with pyrite and chalcopyrite. Enrichment is

associated with brecciated zones that have undergone secondary K-feldspar flooding and/or intense pyrite

+ chlorite + silica alteration.

1.6 Deposit Type

The Central Zone deposit is similar in characteristics to both the classic alkali porphyries in that the

mineralization is associated with an intrusive complex of alkali-feldspar-saturated monzonite and the

calc-alkalic porphyry type deposits in that the mineralization is associated with strong quartz stockwork.

The South Zone deposit is a structurally controlled porphyry deposit with quartz monzonitic to quartz

monzodioritic host lithologies.

1.7 Exploration

In 2005, Serengeti carried out 530km airborne magnetic/radiometric surveys and collected eleven rock

samples. In 2006, Serengeti conducted a 26.9km ground magnetic and IP survey and followed up with a

ten hole diamond drill program totalling 1,874.3m. This was followed in 2007 by 320 line-km of

regional airborne magnetic and electromagnetic (EM) surveys. From March 2007 to August 2008, 113

diamond drillholes (53,646.3m) tested what are now referred to as the Central and South zones.

In late 2008, an additional 70 line-km of 100m spaced dipole IP survey was conducted that delineated

several anomalies for further investigation in the southern extent of the Kwanika property. From June to

September 2009, seventeen diamond drillholes were completed with an aggregate depth of 6,249.1m that

primarily tested the South Zone area and assisted in the establishment of structurally controlled porphyry

deposit model. In 2010 Serengeti drilled an additional 28 holes totalling 7,619m. Most of this drilling

was to in-fill and expand the South Zone.

Additional drilling in the Central Zone area subsequent to the 2010 resource estimation was largely

outside of the mineralized envelope and has no material impact on the existing resource.



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1.8 Drilling

Included in the current Resource Estimate is a total of 69,389m of diamond drilling in 168 holes on the

Kwanika Project from July 2006 to September 2010. Drilling on the Central Zone totalled 50,925.7m in

109 holes, while drilling at the South Zone totalled 18,463.6m in 59 holes. The drilling achieved three

main goals that include delineating Indicated and Inferred Mineral Resources on the Central Zone

deposit; delineating Inferred Mineral Resources on the South Zone deposit; testing several geophysical

anomalies on the Kwanika Property to explore for possible extensions of the Central Zone and South

Zone deposits.

1.9 Sample Preparation, Analyses and Security

All drill core was logged for geological and geotechnical characteristics (geotechnical logging included

rock quality designation (RQD), magnetic susceptibility, and specific gravity), and was photographed,

sampled, and split by diamond saw or core splitter on-site.

1.10 Data Verification

An independent assay QA/QC program has been in place throughout the drill campaigns carried out by

Serengeti. The total number of QA/QC insertions into the sample stream comprised approximately 7% of

all determinations.

1.11 Mineral Processing and Metallurgical Testing

All the metallurgical aspects presented in this PEA Report rely on the information available from the

Technical Report on the Kwanika Project, NI 43-101 Report dated March 3, 2011.

Copper-Gold mineralization in Kwanika has been identified as two main zones, Central Zone, and South

Zone. Serengeti conducted preliminary metallurgical testing on samples from the Central Zone.

Metallurgical testing of the South Zone has not been conducted.

The preliminary metallurgical test work concluded that a conventional concentration process would

require a primary grind of 80% passing 75m, and regrinding of the rougher concentrate to 80% passing

26m before feeding a three-stage cleaning flotation circuit. The final copper concentrate from a locked

cycle test recovered 88.5%Cu, and 65.2%Au with a concentrate grade of 27.7%Cu, and 20.9g/t Au. The

final copper concentrate was found to be very clean, and the content of penalty elements such as As, Bi,

Sb, and Hg is very low.

A follow-up test designed to investigate improved gold recovery from tailings indicates that up to 10.5%

additional gold can be recovered by very fine grinding and flotation of rougher tailings.

The test program is only preliminary in scope, and as such, the composite sample taken from the Central

Zone is not likely to be representative of the entire deposit.

The master composite sample has higher grades than the average PEA mine plan mill feed grades. It is

assumed that any recovery reduction in future test work associated with a reduction of head grade will be

offset by recovery improvements from a more detailed process test work program.

Table 1-1 shows the metallurgical recovery assumptions used for the PEA. These assumptions are

preliminary and will vary with future test work.



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Table 1-1 Kwanika Process Recovery Assumptions

Parameter Value

Copper Recovery 89%

Gold Recovery 70%

Silver Recovery 75%

Molybdenum Recovery 60%

Copper Concentrate 24% Cu

1.12 Mineral Resource Estimates

1.12.1 Central Zone

In 2008, RPA carried out an estimate of the Mineral Resources for the Kwanika Central Zone deposit.

RPA has updated this estimate, and included silver in the grade interpolation for the first time. The

estimate was made using a block model constrained by 3D wireframes of principal geological domains.

Grade was interpolated into the block model using Ordinary Kriging. The results of the 2010 estimate

are shown in Table 1-2 below.

Block model validation was carried out by RPA and adjustments were made at the extreme margins of

the drilled area based on comparisons with drillhole composites. A different methodology, Inverse

Distance Cubed (ID3), was used to check the kriging estimate. The ID3 estimate agreed with the kriged

estimate to an acceptable tolerance and confirmed that the interpolation is reasonable.

The Mineral Resources were classified as Indicated within the area of the deposit drilled to a nominal

spacing of 50m. Blocks outside this area, and comprising a reasonably coherent body constrained by the

search parameters and the extent of the drilling, were categorized as Inferred.

A Lerchs-Grossmann pit optimization was run by RPA to test whether the deposit could potentially be

mined by open pit methods as a large scale operation. Only those blocks captured by the pit shell were

included in the Mineral Resources estimate.

For the 2010 update, the cut-off criterion was changed from a CuEq cut-off to a gross dollar value. The

dollar value was calculated using metal prices of US$1,200/oz Au, US$3.50/lb Cu, and US$21.00/oz Ag,

which represents an increase of the prices used in 2008 and also included a provision for metallurgical

recovery.

The addition of silver to the estimate and the changes to the reporting and cut-off criteria resulted in a

significant increase in tonnage and a partially compensating decrease in grade. The resulting metal

contents for the 2010 estimate were marginally higher for the Indicated category and significantly higher

for the Inferred.



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Table 1-2 Central Zone Mineral Resources – December 31, 2010

Category Cut-Off Tonnage Au Cu Ag

(US$/t) (Mt) (g/t) (M oz) (%) (M lb) (g/t) (M oz)

Indicated 20.00 91.0 0.36 1.06 0.38 754 1.08 3.15

17.50 113 0.33 1.18 0.34 849 0.98 3.56

15.00 142 0.29 1.32 0.31 962 0.89 4.07

12.50 178 0.26 1.47 0.27 1,080 0.81 4.61

10.00 213 0.23 1.59 0.25 1,170 0.74 5.08

7.50 244 0.21 1.66 0.23 1,230 0.69 5.39

Inferred 20.00 5.13 0.27 0.044 0.26 29.0 0.65 0.11

17.50 9.18 0.24 0.071 0.23 45.6 0.56 0.16

15.00 16.6 0.21 0.113 0.20 72.4 0.49 0.26

12.50 26.5 0.18 0.156 0.18 103 0.46 0.39

10.00 39.3 0.16 0.202 0.16 136 0.44 0.56

7.50 55.2 0.14 0.245 0.14 168 0.42 0.74 Notes:

1. CIM definitions were followed for Mineral Resources.

2. Mineral Resources are estimated at a cut-off grade of US$7.50/t. The dollar value cut-off was estimated using

provisions for metallurgical recovery and off-site costs, as described in the text of this report.

3. Mineral Resources are estimated using an average long-term price of US$1,200/oz Au, US$3.50/lb Cu, and

US$21.00/oz Ag.

4. A minimum mining width of 5m was used.

5. Metallurgical recovery factors of 89% Cu, 70% Au, and 75% Ag were used to derive the dollar value cut-off.

1.12.2 South Zone

Following completion of a drilling program in 2009, Scott Wilson RPA was retained to estimate the

Mineral Resources for the South Zone, which is located approximately 750m south of the Central Zone.

Serengeti carried out a diamond drilling program in 2010 to confirm and extend the known mineralization

in the South Zone. Following the completion of this program, RPA was retained to update the block

model.

The estimate was prepared using a block model constrained by 3D wireframe models. Grade was

interpolated into the blocks using ID3 weighting.

All Mineral Resources in the South Zone are classed as Inferred owing to the sparseness of drilling and

the preliminary nature of the geological interpretation.

The South Zone Mineral Resource estimate is summarized in Table 1-3 below. The update to the

Mineral Resource estimate resulted in a significant increase in tonnage with a partially offsetting drop in

grade, and an overall increase in metal content.

A Lerchs-Grossmann pit optimization was run by RPA using somewhat optimistic metal prices and cost

parameters in order to test whether the deposit could potentially be mined by open pit methods as a large

scale operation. Only those blocks captured within the pit shell were included in the Mineral Resource

estimate. As stated above, all Mineral Resources are classified as Inferred. The results of the 2010

estimate are shown in Table 1-3 below.



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A dollar value was calculated from the grades, at metal prices of US$3.50/lb Cu, US$1,200/oz Au,

US$17.00/lb Mo, and US$21.00/oz Ag with provision for metallurgical recovery. Metallurgical

recoveries were 89% Cu, 70% Au, 60% Mo, and 75% Ag. The cut-off applied to the block model for

reporting purposes was US$7.50/t, which, in RPA’s opinion, was consistent at the time of writing with

similar deposits in British Columbia.

Table 1-3 South Zone Mineral Resources – December 31, 2010

Category Cut-Off Tonnage Au Cu Mo Ag

(US$/t) (Mt) (g/t) (M oz) (%) (M lb) (%) (M lb) (g/t) (M oz)

Inferred 20.00 74.3 0.12 0.292 0.33 546 0.012 19.4 2.15 5.15

17.50 98.8 0.11 0.365 0.30 656 0.011 22.9 1.99 6.34

15.00 132 0.11 0.457 0.27 785 0.009 27.0 1.84 7.83

12.50 176 0.10 0.554 0.24 928 0.008 31.8 1.68 9.52

10.00 215 0.09 0.623 0.22 1,030 0.008 35.5 1.56 10.8

7.50 240 0.09 0.664 0.20 1,080 0.007 37.6 1.49 11.5 Notes:


2. Mineral Resources are estimated at a cut-off grade of US$7.50/t. The dollar value cut-off was estimated using provisions

for metallurgical recovery and off-site costs, as described in the text of this report.

3. Mineral Resources are estimated using an average long-term price of US$1,200/oz Au, US$3.50/lb Cu, US$17.00/lb Mo,

and US$21.00/oz Ag.


5. Metallurgical recovery factors of 89% Cu, 70% Au, 60% Mo and 75% Ag were used to derive the dollar value cut-off.

1.12.3 Open Pit Mine Planning

MMTS produced a series of Lerchs-Grossman (LG) pit shell optimizations for the Kwanika deposit using

mining, processing, tailings, general and administrative (G&A) costs, and process metal recoveries

estimated from similar studies and from information available for the Kwanika project. Indicated and

Inferred Resource classes are used in the economic pit optimization for this scoping level (PEA) study.

Cut-off Grade (COG) is determined using an estimated Net Smelter Return (NSR) in $/t, which is

calculated using Net Smelter Prices (NSP). The NSR (net of offsite charges and mill recovery) is used as

a cut-off item for break-even economic selection of mineralized material. The NSP includes metal prices,

US$ exchange rate, and off-site transportation, smelting, and refining charges from earlier work

estimated in late 2011. The 3-year trailing average metal prices and resultant NSP are shown in Table

1-4.

Table 1-4 Metal Prices and NSP – 3 Year Trailing Average Sept 1, 2011

Metal Metal Price (US$/unit) NSP ($/unit)

Copper US$3.12/lb $2.84/lb

Gold US$1,178.78/oz $36.30/gm

Silver US$21.19/oz $0.443/gm

Molybdenum US$14.52 $13.85/lb

It has been deemed as not significant to update ultimate pit limit for this PEA study using current metal

prices. The lower prices used from 2011, make the economic limits conservative. Current prices and

parameters have however been included for the cashflow.



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MMTS notes that the economic pit limits are based on the Indicated and Inferred Resource classes as well

as estimated mining unit costs at a PEA level of study. These costs are derived to meet the local

condition for the Project and the specific project arrangements for mine rock management, water

management, environmental, and reclamation to the PEA level of study, as well as certain input

parameters, such as pit slope angles, process recoveries, environmental considerations, and reclamation

requirements. All of these components affect the mining quantities and activities to release the specified

mineralization and, as such, affect the economic pit limits. Some of the elements of this study require

more sampling and evaluation for future prefeasibility level work and changes to the general design

concepts can impact the mine plan. Changes to these design elements and parameters will not only affect

the cost estimates within the plan, but will also impact the economic mining limit in future studies.

The in situ LG pit delineated resources are summarized in Table 1-5 with mining loss, dilution and a cut-

off applied.

Table 1-5 Run of Mine LG Pit Delineated Resource Including Indicated and Inferred

Classes

Zone Category Tonnage Cu Au Ag Mo

(Mt) (%) (g/t) (g/t) (%)

Central Indicated 20.6 0.336 0.291 0.882

Inferred 0.9 0.332 0.209 0.840

South Inferred* 23.95 0.334 0.117 1.890 0.016**

Note: NSR cut-off used is $11.90/tonne with a provision for mining loss of 5% and dilution of 2%

* There is no indicated classified resource present in South Zone

** Moly only present in South Zone

1.12.4 Underground Mine Planning

There are sufficient tonnes and grade below the Central Open Pit to support an underground mine.

Several different mining methods have been evaluated including block caving. AMEC has reviewed the

drill core and the targeted higher grade mineralized zone and considers Block Caving to be a viable

mining method. Accordingly MMTS developed Block Cave stopes and infrastructure based on typical

parameters which include caveability, requirements for rock pre-conditioning, geotechnical aspects of the

rock, development rates and mining loss and dilution factors. These parameters will require refinement

in future studies.

The underground designs include two stope outlines, access development spiraling down from the bottom

of Central Open Pit, ventilation development and undercut, drawbell, cross-cut and extraction level

designs for block cave mining. At this stage, the stope outlines are conservative and planning includes an

estimate of equipment and rate of development from general factors. Optimization will be needed in

future designs to maximize the underground extraction and optimize the ROM head grades from the

underground operations. Subsequent to this evaluation it is noted that there are zones of material adjacent

to the stopes used in this study which have sufficient grade to be included in future studies. The

underground delineated resources are listed in Table 1-6.



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Table 1-6 Run of Mine Underground Delineated Resource Including Indicated and Inferred

Classes


(Mt) (%) (g/t) (g/t)

UG1 Indicated 19.92 0.456 0.467 1.359 -

Inferred 0.05 0.332 0.418 1.788 -

UG2 Indicated 7.82 0.472 0.470 1.273 -

Note: Includes all material within the conceptual stope with a provision for mining loss of 10% and dilution of 15%

1.13 Mineral Reserve Estimates

The current study is at a PEA level and therefore there are currently no Mineral Reserves estimated for

the Kwanika Project.

1.14 Mining Methods

A production schedule based on a 15,000t/d mill feed rate at a preliminary assessment level has been

developed for the Kwanika Project. The pit phases are engineered based on the results of an updated

economic pit limit analysis. The underground stopes are engineered based on the results of a cut-off

grade analysis (refer to 14.3.7). A summary of Indicated and Inferred open pit mill feed for production

scheduling is provided in Table 1-7 using whole block grades with mining dilution and loss varying by

extraction method (open pit versus underground).

The copper (Cu), gold (Au), silver (Ag) and Molybdenum (Mo) grade items used in this section of the

study are based on the resource model provided by RPA as described in Section 14.0.

Table 1-7 Summarized Indicated and Inferred Pit Mill Feed for Production Scheduling

Source In-pit Resource

(Mt) Cu(%)

Au

(g/tonne)

Ag

(g/tonne) Mo (%)*

Mine Rock

(Mt)

Strip Ratio

(t:t)

Pit CK615 11.05 0.37 0.331 0.978 26.38 2.39

Pit CK625 9.75 0.28 0.218 0.722 30.48 3.13

Pit SK612 7.19 0.40 0.059 2.005 0.011 25.70 3.57

Pit SK622 17.88 0.29 0.140 1.821 0.018 75.55 4.23

Note: Open Pit NSR cut-off used is $11.90/tonne with a provision for mining loss of 5% and dilution of 2%.

* Moly only present in South Zone

The underground designs include two stope outlines, access development spiralling down from the

bottom of Central Open Pit, ventilation development and undercut, drawbell, cross-cut and extraction

level designs for block cave mining. At this stage of planning, the stope outlines are conservative and

planning includes an estimate of equipment and rate of development from general factors. Optimization

will be needed in future designs to maximize the underground extraction and optimize the ROM head

grades from the underground operations. The summarized underground mill feed for production

scheduling is listed in Table 1-8.



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Table 1-8 Summarized Indicated and Inferred Underground Mill Feed for Production

Scheduling


(Mt)

Cu

(%)

Au

(g/tonne)

Ag

(g/tonne)

Mo

(%)*

Mine Rock

(Mt)

Strip Ratio

(t:t)

UGTALL 19.97 0.46 0.467 1.360 0.000 0 0

UGWEST 7.82 0.47 0.470 1.273 0.000 0 0


Note - The Open pit and Underground Mineral resources stated in Table 1-7 and Table 1-8 are

preliminary in nature, and include inferred mineral resources that are considered too speculative

geologically to have the economic considerations applied to them that would enable them to be

categorized as mineral reserves in future studies. There is no certainty that the results of the

preliminary economic assessment will be realized.

1.14.1 Rock Storage Facility

The mine rock below cut-off grade is placed in a rock storage facility (RSF) situated east of the Central

Pit area, and is constructed in lifts using a bottom-up method. Foundation preparation is completed, as

required. Allowances are made to address reclamation and post-closure requirements.

1.14.2 Mining Operations

Detailed pit phases have been designed to include the detailed pit slope criteria and high wall roads. The

Ultimate Pit is divided into smaller mining phases, or pushbacks, to enable a low strip ratio starter pit and

to allow for even waste stripping to be scheduled.

The underground and open pit production for mill feed will be mined simultaneously in order to meet the

throughput requirements of the processing plant. However the first few years of operations will be solely

from Central Pit. Once the Central Pit mining has commenced, development of the access ramp and

ventilation raise for the underground block caving operation will start. In order to avoid influence from

the block caving subsidence, the access ramp will be driven from surface and not from the bottom of the

Central Pit. Level development will be in place when Central Pit Phase 2 mining is complete. The

underground mill feed production will be hauled to the top of the underground access ramp, then re-

loaded and hauled to the processing plant.

Pre-stripping waste production for the open pit operations is achieved by starting the mine fleet during

the construction period. Underground access, services, and extraction level development is scheduled

concurrent with open pit operations from Central Pit. Production from the first block cave stope will

commence as Central Pit is completed. The access and extraction level for the deeper stope will progress

as the upper stope is mined. Production will shift to the lower stope and to South Pit towards the end of

the mine life.

The above mining resources (Table 1-7 and Table 1-8) are used to produce a consolidated production

schedule. This includes sequencing the development of the pit phases to even out the stripping

requirements and to run reduced open pit mill feed to supplement the lower capacity underground

mining. A summary of the production schedule is provided in Table 1-9.



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Mining operations, methods, and equipment will be typical of open-pit and underground mining in

Northern British Columbia. The Project will be a medium-capacity operation that utilizes appropriately

sized equipment for the major operating areas in order to generate high productivities, and reduce unit

and overall mining costs.

The mine plan and production schedule will undergo further refinement during the Prefeasibility Study.

Additional information on underground footprint optimization should more accurately determine the

underground stope geometry and projected caving rates. Further details on rock storage management,

water management, and final land use will be developed for the Environmental Assessment application,

the result of which will impact the mine plan. These elements, along with other optimization details, will

be integrated into prefeasibility-stage mine planning.



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Table 1-9 Life of Mine Production Summary



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1.15 Recovery Methods

A conventional copper-gold flotation process is assumed for the Kwanika project including crushing,

grinding, and multi-stage froth flotation to produce a copper-gold concentrate.

1.16 Project Infrastructure

The Project site will be accessible by road from Fort St. James via the all-weather Leo Creek and

Driftwood forest service roads (FSR) and the 30km long Tsayta Lake Road. Access to the Kemess

Power Line is available 75km from the Project site.

The Project site will have open pit and underground mining related facilities, process related facilities,

and a permanent camp. The general site layout is shown in Figure 1-1.

1.16.1 Access

Currently there is an existing forest service road between Fort St. James and the Tsayta Lake Road

providing surface access to the site. For the purposes of this study, it is assumed that the 30km long

Tsayta Lake Road will be upgraded to meet the needs of the operation.

1.16.2 On-Site Transport Road

On-site service roads will be constructed connecting to the Project access road, the explosives storage and

manufacturing facilities, tailings storage facility, processing plant, rock storage facility, and open pits.

1.16.3 Main Facilities

The main facilities for the Project will include:

Primary crushing facility

Main process plant including; primary grinding, flotation, regrinding, and water treatment

Mine rock storage facility

Tailings storage facility

Mine maintenance facilities

Truck shop; including warehouse

Permanent camp; including general offices

Explosive storage facilities

Sewage treatment facility

Potable water system

Electrical Sub-station

1.16.4 Power Supply and Distribution

The selected power supply option for the Project includes extending a connection from the Kemess

Power Line 75km to the Kwanika Project site.

1.16.5 Mine Rock and Tailing Storage Facilities

Mine rock and run-of-mine (ROM) waste products from the mining operations will be stored in separate

storage facilities as described in the following sections.



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Tailings Storage Facility (TSF)

The flotation tailings will be pumped to the TSF to the West of the processing plant and South West of

the Central Pit area. Mine rock will form the primary dam structure and will have a core of lower

permeability material. A starter dam will be constructed with pre-stripped mine rock and till during pre-

production operations at an elevation of 1045m. Approximately 35Mm3 of flotation tailings will be

disposed of in the tailings storage facility during the life of mine operations.

The supernatant water that will accumulate within the tailings storage facility will be reclaimed by

floating barge and pumped to either the process plant or to the water storage pond in a closed system. A

seepage collection system will be included in the design.

Rock Storage Facility (RSF)

Mine rock represents the largest waste stream from the mining operations with an estimated volume of

70Mm3 during the life of mine operations. The RSF will be located to the east of Central Pit and will

occupy an area of 121ha. The mine rock will be directly hauled from the pit and end-dumped in 30-60m

lifts building the dump from the bottom up. Seepage and runoff from this facility will drain southwards

and will be collected and managed within the water storage pond.

1.16.6 Mine Area Water Management

Currently, Kwanika Creek flows through the proposed South Open Pit area and the tailings storage

facility area therefore it will be necessary to divert the creek away from these operations. A diversion

channel has been proposed that routes the stream northward around the tailings storage facility, and

eastward around South Pit to join the existing Kwanika Creek east of the proposed development.

Fresh water diversion will be used where possible to keep ‘non-contact water’ from entering active

mining area. For contact water, it will be necessary to construct a water storage pond at the south end of

the RSF to collect surface runoff from catchment areas and seepage water from the flotation tailings

during mining operations and the initial stages of closure and reclamation. Water from the active open pit

areas and from underground pumping, will also be directed to a water storage pond for re-use or

eventually treated and released to the environment. Details will be addressed as part of the overall site

water balance and site water management plan in future studies.

1.16.7 Mine Area Closure Plan

At the cessation of mining operations, a closure plan will be developed to return the operating area to a

condition that will meet the end land use objectives.

The flotation tailings will be progressively capped and the outer slopes of the RSF will be re-sloped to

blend with the natural landscape and to enable access to wildlife. Natural seepage water collected within

the water storage pond will be pumped and discharged to the open pit until the water quality meets

discharge criteria.

The open pits will be allowed to fill through seepage and surface run-off. Stream run-off may be directed

into the completed mining areas (open pit and Underground) to reduce the ARD and metal leaching

potential as quickly as possible. The Kwanika Creek diversion channels will either continue to operate or

will be decommissioned.



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Figure 1-1 General Site Layout



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1.17 Market Studies and Contracts

An update to the three year trailing average metal prices was performed on October 15, 2012. These

metal prices and resultant Net Smelter Prices are shown in Table 1-10.

Table 1-10 Metal Prices and NSP – 3 Year Trailing Average Oct 15, 2012



Gold US$1,427.00/oz $44.57/gm



NSP – Market price less smelting, refining and off-site charges

Concentrates will be sold into the general market. This will either be to North American, European, or

Asian smelters and refineries.

1.18 Environmental Studies, Permitting and Social or Community Impact

1.18.1 Regulatory Framework

The Serengeti Kwanika Project falls within the category of a “reviewable project” of the British

Columbia Environmental Assessment Act (BCEAA and with a proposed diversion of Kwanika Creek, a

fish-bearing stream, will require a Federal Fisheries Act approval, and will likely trigger the requirements

of the CEAA. Other requirements of Provincial and Federal Acts and Regulations may also apply,

depending upon final design components.

Consideration will be made in ongoing engineering to mitigate issues that will trigger the full CEAA

process.

1.18.2 Regional Land Use Processes

The Project is located within lands that have been dedicated in the Fort St. James Land and Resource

Management Plan, approved by government in 1999. The Project area is within the Multi-Value

Resource Management Zone Land Use designation, where lands are managed to integrate a wide range of

resource values, including mining.

1.18.3 Programs Already in Progress

In support of the exploration programs, Serengeti has been in consultation with the local Takla Lake First

Nations (TLFN), providing jobs as well as starting base line environmental, archeological, weather and

water studies including a project specific Valued Ecosystem Component (VEC) study (see Section 20).

The positive relations to date, as well early baseline data for local environmental and water quality will

be a benefit to the future Pre-Feasibility study



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1.19 Capital and Operating Costs

1.19.1 Capital Costs

All currencies in this section are expressed in Canadian dollars. Costs in this report have been converted

using a fixed currency exchange rate of US$0.95 to CDN$1.00. The expected accuracy range of the

capital cost estimate is +/- 40%.

An initial capital of $363.6M is estimated for the Project.

Initial capital has been designated as all capital expenditures required for producing copper concentrates

for shipment to contract smelters. A summary of the major capital costs is shown in Table 1-11 and

Table 1-12.

This PEA estimate is prepared with a base date of Q4 2012 and does not include any escalation past this

date.

Table 1-11 Initial Capital Cost Summary

Description Capital Cost

($000)

Open Pit Mining – Equipment 56,920

Open Pit Mining – Pre-Production 31,575

Underground Mining - Equipment 10,080

Underground Mining - Development 11,400

Processing Plant 129,000

Site Infrastructure 41,170

Access & Power 24,900

Contingency 58,520

Total Initial Capital Cost 363,565

Table 1-12 Sustaining Capital Cost Summary


($000)

Open Pit Mining – Sustaining 20,790



Reclamation 1,810

Total Sustaining Capital Cost 143,870

1.19.2 Operating Costs

The operating costs for the Project, as shown in Table 1-13, are estimated at an overall cost of $21.20/t of

mill feed. The estimate is based on an average annual process rate of 5,475,000t milled at an average

grade of 0.38% copper, 0.295g/t gold, 1.378g/t silver, and 0.016% molybdenum (present in the South

Zone only) including mine dilution. The cost estimates in this section are based upon budget prices in Q4

2012 or based on the information from the database of MMTS. When required, costs in this report have

been converted using an average currency exchange rate of US$0.95 to CDN$1.00. All costs are



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reflected in 2012 Canadian dollars. The expected accuracy range of the operating cost estimate is +/-

35%.

Table 1-13 Operating Cost Summary

$/tonne Mined $/tonne Milled

Mine (OP) $2.42

Mine (UG) $7.50*

Mine Total $9.14**

TSF $0.18

Mill $10.69

G & A $1.20

Tailing Treatment Included in G&A

Water Treatment Included in G&A

Total $21.20

*Note: UG cost/tonne mined based on total tonnes mined by UG only.

**Note: Total mining cost (OP + UG) divided by total tonnes milled.

The operating costs are defined as the direct operating costs including mining, processing, tailings

handling, water treatment, and G&A. The power is estimated to be $0.05/kWh. The power cost is based

on other studies in the general area of the Project.

1.20 Economic Analysis

A Base Case economic evaluation has been undertaken incorporating historical three-year trailing

averages for metal prices as of October 15, 2012. This approach is consistent with the guidance of the

United States Securities and Exchange Commission, adheres to National Instrument 43-101 and is

consistent with industry practice. The metal production values indicated in Table 1-14 are a summary of

the results of the production schedule, which is used in the cash flow to determine projected revenues.

The pre-tax economic results in Canadian dollars are listed in Table 1-15.



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Table 1-14 Metal Production from Kwanika

Years 1 to 5 LOM

Total Tonnes to Mill (000s) 27,375 73,663

Annual Tonnes to Mill (000s) 5,475 5,475

Average Grade

Copper (%) 0.35 0.377

Gold (g/t) 0.285 0.295

Silver (g/t) 1.057 1.378

Moly (%)* 0.001 0.016

Total Production (after Recovery)

Copper (000s lb) 187,993 544,892

Gold (000s oz) 175.6 489.0

Silver (000s oz) 697.7 2,447.7

Moly (000s lb) 344.6 5,251

Average Annual Production

Copper (000s lb) 37,599 40,512

Gold (000s oz) 35.1 36.4

Silver (000s oz) 139.5 182.0

Moly (000s lb) 68.9 390.4

* South Pit only

Table 1-15 Summary of the Economic Evaluation

Unit Base Case

Metal Price

Copper US$/lb 3.63

Gold US$/oz 1,427.00

Silver US$/oz 27.50

Molybdenum US$/lb 14.45

Exchange Rate US$:CDN$ $0.95

Economic Results (Pre-Tax)

Undiscounted cash flow $ M 567.1

NPV (at 5%) $ M 262.6

NPV (at 8%) $ M 143.3

NPV (at 10%) $ M 81.2

IRR % 13.4

Payback years 7.3



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1.20.1 Sensitivity Analysis

Sensitivity analyses have been carried out on the following parameters:

Copper and gold metal price

Exchange rate

Initial capital expenditure

On-site operating costs

Copper & Gold Recovery

The analyses are presented graphically as financial outcomes in terms of NPV and IRR. Both the Project

NPV and IRR are most sensitive to Copper Price followed closely by Copper & Gold Recovery, with

Capital Cost having the least impact. The NPV and IRR sensitivities can be seen in Figure 1-2 and

Figure 1-3. These results are presented graphically only to show trends for future evaluation. At a scoping

level of engineering and costing the absolute values are not deemed relevant for economic evaluation.

0%

50%

100%

150%

200%

250%

-15% -10% -5% 0% 5% 10% 15% 20%

Sen

siti

vity

to

Bas

e C

ase

% ofBase Case

Input Sensitivities at 5% discount rate

Metal Price Sensitivity

Capital Cost Sensitivity

Operating Cost Sensitivity

Exchange Rate Sensitivity

Copper & Gold Recovery Sensitivity

Copper Price Sensitivity

Gold Price Sensitivity

Figure 1-2 Base Case Sensitivity to Pre-Tax NPV @ 5%



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0%

20%

40%

60%

80%

100%

120%

140%

160%

180%

-15% -10% -5% 0% 5% 10% 15% 20%

Sen

siti

vity

to

Bas

e C

ase

% of Base Case

Input Sensitivities to IRR








Figure 1-3 Base Case Sensitivity to Pre-Tax IRR

Note - The Mineral resources and economic results stated in this Preliminary Economic Assessment

are preliminary in nature, and include inferred mineral resources that are considered too speculative

geologically to have the economic considerations applied to them that would enable them to be

categorized as mineral reserves, and there is no certainty that the results of the preliminary economic

assessment will be realized.

1.21 Adjacent Properties

1.21.1 Regional

The Quesnel Trough is the host to several other porphyry copper ± gold mines and significant deposits.

These deposits include: the Mount Polley Mine, the former Kemess Mine and its related infrastructure

located north of Kwanika, and the Mount Milligan Mine development project located approximately

85km south of Kwanika.

1.21.2 Local District

The adjacent Lustdust claims, owned by Alpha Gold Corporation, are located immediately to the north of

the Kwanika property. The Lustdust property has been the subject of exploration for fifteen years on

various precious and base metal vein and skarn occurrences and contains a small Indicated and Inferred

copper-gold Mineral Resource known as the Canyon Creek Zone. The other significant prospect in the

general vicinity of Kwanika is the Lorraine porphyry copper-gold property jointly controlled by Teck

Corporation and Lorraine Copper Corp. which contains a modest, Indicated and Inferred Mineral

Resource in two deposits. Both of these properties are the subject of current NI-43-101 reports.



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1.22 Interpretation and Conclusions

The Kwanika deposit represents a copper-gold-silver deposit that is amenable to open pit and

underground mining through the use of block caving and conventional milling consisting of flotation

concentration.

1.22.1 Geology and Resource Modeling

RPA carried out estimates of Mineral Resources for the Central Zone in 2008 and South Zones on the

Kwanika Property in 2009 and 2010. The current Mineral Resources estimates are summarized in Table

14-1 and Table 14-2 above.

RPA drew the following conclusions:

Drilling, core handling and sampling and security protocols were appropriate and samples should

be representative of the mineralization.

Conventional assaying techniques were used, sample QA/QC protocols were adequate and

checks at a secondary laboratory were consistent with the primary laboratory results.

Validation by RPA showed the sample database was reasonably free of errors, representative and

appropriate for use in Mineral Resource Estimates.

Due to a lower density of drillholes, geological interpretation, wireframe modelling and

geostatistical analysis in the South Zone are preliminary in nature and all of the resource was

categorized as Inferred.

Due to a sufficient density of drilling in the Central Zone, a block model was constrained by 3D

wire-framed geological domains constructed with grade interpolation by ordinary kriging for the

Central Zone and much of the resource was characterized as Indicated category.

Additional drilling will be required to increase the confidence level Inferred Mineral Resources

to Indicated category particularly in the South Zone.

1.22.2 Metallurgy

The metallurgical test work indicated that the mineralization responds well to the process

consisting of conventional flotation. Copper recovery of 88.5% in the locked cycle test to a

copper concentrate grading 27.7% Cu confirmed earlier results obtained in batch cleaner tests.

The concentrate contained 20.9g/t Au at a gold recovery of 65.2%. The test material responded

well to various collectors, which will allow the least costly reagents to be used in a commercial

operation. A follow-up test designed to investigate improved gold recovery from tailings

indicates that up to 10.5% additional gold can be recovered by very fine grinding and flotation of

rougher tailings.

The test program is only preliminary in scope, and as such, the composite sample taken from the

Central Zone may not be representative of the entire deposit.

With the project now focused on the Central and South pits, and Underground stopes in this PEA

mine plan, Metallurgical samples need to be collected that represent this part of the resource area

and appropriate lab test run.

A mill throughput in the order of 15,000 tonnes per day is indicated as a basis of estimate.

Further evaluation of throughput is required in future studies.

Process design and costing in future studies will be required after the through put optimization is

completed.



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1.22.3 Underground Mine Plan

The mineral zone evaluated for underground mining and the rock characteristics are suitable for

Block Caving. This mining method will provide a low operating cost.

Additional economic underground resources are possible both below and adjacent to the block

cave stope outline used in this report. Optimization of the underground mine plan should add to

these potential economic resources. This should include the use of optimization software to

define the economic footprint and also evaluation of the preferred development/extraction levels.

Further evaluation of the extraction levels could increase the amount of mineable material, and

can also improve project cashflow. Designing an extraction level closer to the open pit bottom

will allow higher grade underground mill feed to be produced earlier in the production schedule

and with less up front capital development costs which will be positive to the project financials.

After the conceptual economic stope limits and preferred extraction levels are optimized,

planning and design can proceed at a Pre-Feasibility level. This will require geotechnical studies

and Engineering to evaluate caving parameters, caving rate, pre-conditioning requirements, and

draw point design details.

1.22.4 Open Pit Mine Plan

The Central and South pits are designed to their break-even economic limit and provide low cost

mining for the near surface material. The lowest benches have the highest incremental strip ratio

and are the higher cost on both a “per tonne mined” and “per tonne milled” basis. Optimization

of the mining method at the open pit/underground interface has the potential to improve the

overall economics of the property. Material on the bottom benches of the pit may produce a

better net return if mined as part of the underground interface. An evaluation is required to verify

this.

Geotechnical field investigations, including drilling, are required and the open pit design needs to

be taken to a Prefeasibility study level. This redesign work will also be based on the

Underground mining limit optimization, equipment optimization studies, and water and waste

rock management details.

1.22.5 General

The financial results from the base case plan used in this report are based on the parameters and

assumptions as described which are preliminary in nature. The financial sensitivity to these parameters

and assumptions has been indicated and as noted there are other areas in the mine plan that need

optimization that could increase or decrease the economic mining limits. To advance the project to a Pre-

Feasibility level will require significant site investigation work and testing, including exploration drilling,

Geotechnical drilling and testing, metallurgical sampling and testing, and base line Environmental

studies. Before this Pre-Feasibility work is done, several optimization studies should be undertaken to

test the project’s sensitivity to key parameters and to guide the scope of a future Pre-Feasibility Study.

1.22.6 Regional and Other On-sight Opportunities

This PEA study provides a scoping level basis for a viable operation with the opportunity to add

more economic resources both on site and in the local area.

There are other properties in the local area that have the potential of using the Kwanika facilities

on a contract or joint venture basis.

The expanded resource can use the facilities and infrastructure from this study. A significant

mineralized resource in the Kwanika deposit surround the resources used in this mine plan. The

potential exists for some of this marginal material to be brought into an economic resource base



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after this proposed operation has met capital payback, or if future expansions can provide a lower

operating cost due to economies of scale.

1.23 Recommendations

MMTS recommends advancing the project in two steps. First is a series of optimization studies to refine

the engineering and costing in the areas with higher economic sensitivity. This work will be based on

more engineering but will require minimal additional field data. The results of these studies will be a

revised PEA. If still justified after the Optimization studies, the project can then advance to a Pre-

Feasibility Study level, which will involve more field data for upgrading the inferred resources,

geotechnical evaluations, metallurgical test work, and field data collection to start the environmental

impact assessment and reclamation planning.

1.23.1 Optimization Studies

Based on the project’s economic sensitivity to various parameters, further engineering assessment work

should be done at a scoping level using existing data to reduce the down side risk and to focus the

planning and design work before extensive expenditures are incurred for a Pre-Feasibility study. These

optimizations should include:

Optimize the underground conceptual plan and add material not included in the current plan.

Investigate the ‘best’ economic mining method for the interface between open pit and

underground.

Refine equipment and planning parameters, for open pit, underground, and processing.

Estimate mining loss and dilution and dilution grades.

Update the capital cost estimates including a project timeline for development expenditure and

more detailed indirect costs.

With revisions to the underground plan, optimize the production schedule to bring higher grade

underground mill feed forward in time and with less up front development capital.

1.23.2 Pre-Feasibility Study

The project should advance to a Pre-Feasibility Study based on the results from the Optimization Studies.

The results of the Optimization studies will not only allow a “go/no go” decision but will also establish

areas requiring more detailed data and analysis for the Pre-Feasibility work.

Recommended Future Metallurgical Testwork

Sampling and metallurgical testing the South Zone. A sample representing the potential mill feed

from the South Zone needs to be obtained and tested for conventional flotation as conducted for

the Central Zone.

The Central Zone flotation process needs further investigation aiming to increase the primary

grinding P80, and increase precious metals recovery.

Future Central Zone process test work should be carried out with samples representing the

anticipated mill feed grade range.

Variability sampling and testing representing the rock type, mineralization, alteration, hardness

and grade variability throughout the deposit. These include additional bond work index to

characterize the hardness variability throughput the deposit, and better define the association of

copper with precious metals in the whole range of the expected mill feed grade, and their

response to flotation.

The presence of impurities in final concentrate requires quantification.



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Flotation products require the following evaluations:

o Settling and filtration of copper-gold concentrate.

o Sedimentation of flotation tailings.

o Potential acid generation from tailings.

Future flotation test work should assess the molybdenum recovery to the bulk copper concentrate

(Cu-Au-Mo), and evaluate molybdenum separation to a separate molybdenum concentrate.

1.23.3 Study Costs

To advance the project the following approximate costs will be incurred:

Optimization Studies $50,000

Pre-Feasibility $3,500,000



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2.0 Introduction The Serengeti Resources Inc. Kwanika Project (the Project) involves the development of a copper-gold

deposit located near Fort St. James, British Columbia, Canada.

This National Instrument 43-101 (NI 43-101) compliant report on the Project has been prepared by

Moose Mountain Technical Services (MMTS) and is based on work produced by the following

independent consultants:

RPA Inc. (RPA), formerly Roscoe Postle Associates Inc.


David W. Rennie (P.Eng.) of RPA has visited the Project on May 25-27, 2008; January 28, 2010, and

October 30, 2010. He is the Qualified Person (QP) for all matters relating to the Mineral Resource

estimate.

Mr. Jim Gray (P.Eng.) of MMTS visited the Project site on October 18, 2011. He is the QP for all

matters relating to infrastructure, mining, mining capital costs, mine operating costs, financial evaluation,

and overall report preparation.

Mr. Tracey Meintjes (P.Eng.) of MMTS is the QP for matters relating to mineral processing, mineral

processing capital, mineral processing operating costs, and metallurgical testing.

The Preliminary Economic Assessment (PEA) is based on exploration and internal Serengeti studies

since 2005. The resource model used is most recently described in an NI 43-101 Technical Report dated

March 3, 2011 which is included in this report. This PEA adds a mine plan and infrastructure bringing it

to a scoping level of accuracy. The mine plan includes an integrated open pit and underground

production schedule using typical operating parameters. A conceptual evaluation by AMEC Americas

Ltd. (AMEC) has confirmed that Block Caving is a suitable underground mining method as applied to

this study. All dollar figures presented in this report are stated in Canadian dollars (CAD) unless

otherwise specified.



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3.0 Reliance on Other Experts AMEC Americas Ltd., of Vancouver, British Columbia reported on matters pertaining to the block

caveability of the Central Zone. (See Appendix A)



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4.0 Property Location and Description The Kwanika Property in north central British Columbia is situated in the Omineca Mining Division,



round by four-wheel-drive vehicle, provided there is active snow removal in winter. Figure 4-1 depicts

the Kwanika Property location.

The Property consists of 28 contiguous unpatented mineral claims covering an area of 8,960.29ha and is

100% owned by Serengeti. It is not subject to any royalties or other outstanding liabilities. Serengeti

acquired the current extent of the property through staking between 2004 and 2006. Figure 4-2 illustrates

the Serengeti Claim boundaries.

The property is the host of two porphyry-style mineral deposits; the copper-gold-molybdenum-silver

South Zone and the copper-gold-silver Central Zone, both of which encompass current Mineral

Resources.



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Figure 4-1 Kwanika Property Location

Figure 4-1



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Figure 4-2 Kwanika Claim Map

Figure 4-2



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5.0 Accessibility, Climate, Local Resources, Infrastructure and

Physiography The Kwanika Property is located approximately 75km to the southwest of the Kemess power line, and

B.C. Railway Company maintains an active rail line to Fort St. James. There is sufficient water available

in the immediate vicinity of the property to support both exploration and potential mining activities.

The average temperature for this area (based on data from Fort St. James) is 3.1ºC, with a peak average

monthly temperature of 21.9°C in July and an average monthly low of -15.8°C in January.

The region receives an average of 295mm of rainfall and 192cm of snowfall annually, with 138 days per

year where precipitation exceeds 0.2mm. The Kwanika property is snow-covered from late October to

May.

The Kwanika Project is in close proximity to the well-serviced communities of Prince George, Smithers,

Fort St. James, and Mackenzie. These established centres can provide skilled labour for mine

construction and operation.

Serengeti has developed a beneficial association with the local Takla Lake First Nation and that there is

community support for the Kwanika Project and the potential employment that it would provide.

Infrastructure currently consists of a 30-man exploration camp. Access to the Kwanika Property from

Fort St. James is via the all-weather Leo Creek and Driftwood forest service roads (FSR) and the 30km

long Tsayta Lake Road. Other infrastructure on the Kwanika Property consists of unpaved logging roads

and several kilometres of excavated trails.



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6.0 History The first exploration on the Kwanika Property occurred in the 1930s and 1940s with the discovery of

mercury at Pinchi Lake. Copper mineralization was first recognized along Kwanika Creek and the

project was tested by a geochemical survey, trenching, and a 26.5m X-ray drill program conducted by

Hogan Minerals Ltd.

Between 1966 and 1969, geological, geochemical, and geophysical surveys were conducted by Canex

Aerial Exploration Ltd. and Great Plains Development Company of Canada. This resulted in an

aggregate of eighteen holes totalling 2,176m drilled by the two companies. An area of low grade copper

mineralization was outlined by this drilling.

In 1972, Bow River Resources Ltd. mapped the property and drilled 549m in six percussion holes. This

was followed by Pechiney Developments Ltd. optioning the property in 1973, conducting induced

polarization (IP) and resistivity surveys and drilling 2,993m in 30 percussion drillholes.

In 1976, a Mineral Resource estimate for the main (currently referred to as the South Zone) deposit was

published.

Between 1981 and 1986, different operators (Placer Developments Ltd., Aume Resources Ltd. and Daren

Resources Ltd.) conducted geochemical surveys and sampled rock outcrops.

Eastfield Resources Ltd. optioned the Swan property (which is situated within the current Kwanika claim

boundary) in 1989 and carried out soil, stream, and rock sampling programs as well as IP. Geological

mapping, prospecting, and resampling of the historic core was followed by a four hole, 549m diamond

drill program in 1991. The claims were allowed to lapse and, in 1995, the property was re-staked by

Discovery Consultants who conducted additional heavy mineral stream sediment and rock sampling.

No additional work was done until Serengeti staked the property starting in 2004. Work done subsequent

to Serengeti’s acquisition of the property is summarized in Item 9.0, Exploration.



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7.0 Geological Setting and Mineralization

7.1 Geology

The Kwanika Property lies in the northern part of the Upper Triassic to Lower Jurassic Quesnellia

Terrane (Quesnel Trough). The Quesnel Trough is the host of numerous alkalic and calc-alkalic

porphyry copper-gold deposits within British Columbia. In the area around the Kwanika Property,

Quesnellia is bounded by the Pinchi fault on the west and by the Manson fault on the east.

Porphyry deposits in the general vicinity of the property are associated with potassically altered diorite,

monzodiorite, monzonite, and syenite plugs and stocks, as well as coeval andesitic volcanic rocks and are

associated with strong aeromagnetic features that trend east-west and southeast-northwest.

The Kwanika Project consists of two mineralized areas: the Central Zone and the South Zone. The

Central Zone deposit is characterized by the presence of two major and several minor intrusive bodies of

the multi-phase Hogem Batholith intruding into successions of andesitic rocks of the Takla Volcanic

Group. All lithologies are constrained to the west by the north-northwest striking, terrane-bounding

Pinchi fault. The most economically significant intrusive body is a north-northeast trending monzonite

stock that dips “shallowly to steeply” to the west and is the dominant host to high grade copper-gold

mineralization. Hypogene mineralization in the Central Zone is characterized by a core zone of strong to

intense, texture destructive albite alteration with variable intensities of potassic, sericite, and hematite

alteration. The albitic and quartz stockwork core zone is surrounded by a broad envelope of weak to

strong, pervasive to vein/fracture controlled potassic alteration.

The geology of the South Zone is typified by various intrusive phases of the Hogem Batholith. The host

lithologies of the South Zone deposits occur within a north-south trending structural corridor. This

structural corridor is bounded by the West Fault to the west and possibly by a similar fault interpreted to

lie along the east boundary of the corridor.

7.2 Mineralization

Copper and gold mineralization in the Central Zone at Kwanika occurs primarily in potassically and

albitically altered lithologies. Alteration and mineralization grade outwards from a strong to intensely

potassically and albitically altered; strongly mineralized core zone to a variably propylitically altered,

weakly mineralized periphery. Stronger mineralization mainly occurs in the monzonite, monzodiorite,

and diorite units, but is also present at generally lower grades within the andesite.

Hypogene mineralization is controlled by several generations of quartz + sulphide veining, with the

highest copper and gold grades occurring in areas of quartz stockwork. Chalcopyrite and bornite are the

most important copper minerals. Chalcopyrite occurs as fine to coarse clots within veins and, less

commonly, as fine to coarse disseminations. Bornite is commonly observed as rims to chalcopyrite

grains and, in many instances, it appears that chalcopyrite has been altered to bornite. Disseminated

mineralization is of lesser importance in the upper part of the deposit but increases in significance at

depth within the system. Pyrite occurs throughout the Central Zone as disseminations within veinlets and

wall rocks.



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A supergene enrichment blanket has been superimposed on the upper surface of the hypogene

mineralization in the Central Zone. Two distinct assemblages of supergene mineralization are observed:

supergene oxide and supergene sulphide. Supergene oxide commonly overlies the supergene sulphide

assemblage. Native copper in the supergene oxide occurs mainly as wires along fractures. Chalcocite

and covellite mineralization in the supergene sulphide phase are thought to have replaced chalcopyrite

based on textural observations.

The South Zone is characterized by porphyry style copper + gold + molybdenum + silver mineralization

within monzonite, quartz monzonite, and monzodiorite. Enrichment is associated with brecciated zones

that have undergone secondary K-feldspar flooding and/or intense pyrite + chlorite + silica alteration.

Primary mineralization in the South Zone is comprised of fine to coarse grained chalcopyrite

disseminations and molybdenite mineralization along fractures and quartz selvages and, less commonly,

disseminated blebs associated with pyrite and chalcopyrite. Carbonate flooding and veining occurs

throughout the sulphide enriched regions. Quartz + sulphide veinlets are minor in the South Zone but are

observed to be more abundant with increased copper-molybdenite grades at depth along the West Fault.



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8.0 Deposit Type The Central Zone deposit is similar to the classic alkali porphyries in that the mineralization is associated

with an intrusive complex of alkali-feldspar-saturated monzonite. On the other hand, it is similar to the

calc-alkalic porphyry type deposits in that the mineralization is associated with strong quartz stockwork,

which is more typical of a calc-alkalic porphyry type deposit. In the opinion of Serengeti geologists, the

Central Zone deposit may in fact be transitional in nature between alkalic and calc-alkalic types.

The South Zone deposit is a structurally controlled porphyry deposit with quartz monzonitic to quartz

monzodioritic host lithologies. Copper-gold-silver-molybdenite mineralization is associated with K-

feldspar ± silica flooding of zones of brittle deformation. Structures bounding the deposit to the east and

west are interpreted as being the cause of brittle deformation and the conduits for fluid flow.



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9.0 Exploration In 2005, Serengeti carried out 530km airborne magnetic/radiometric surveys and collected eleven rock

samples. In 2006, Serengeti conducted a 26.9km ground magnetic and IP survey and followed up with a

ten hole diamond drill program totalling 1,874.3m. This was followed in 2007 by 320 line-km of

regional airborne magnetic and electromagnetic (EM) surveys. From March 2007 to August 2008, 113

diamond drillholes (53,646.3m) tested what is now referred to as the Central and South Zones.

In late 2008, an additional 70 line-km of 100m spaced dipole IP survey was conducted that delineated

several anomalies for further investigation in the southern extent of the Kwanika Property.

From June to September 2009, seventeen diamond drillholes were completed with an aggregate depth of

6,249.1m that primarily tested the South Zone area and assisted in the establishment of structurally

controlled porphyry deposit model.

In 2010 Serengeti drilled an additional 28 holes totalling 7,619m. Most of this drilling was to in-fill and

expand the South Zone.

An additional 3000m has been drilled in the Central Zone area subsequent to the 2010 resource

estimation. While some mineralization was detected, this drilling was largely outside of the mineralized

envelope and has no material impact on the existing resource. All holes from drill section 225 S, to the

northern end of the property are described as being in the Central Zone. All holes south of 225 S are

described as being in the South Zone.



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10.0 Drilling Included in the current Resource Estimate is a total of 69,389m of diamond drilling in 168 holes on the

Kwanika Project from July 2006 to September 2010. Drilling on the Central Zone totalled 50,925.7m in

109 holes, while drilling at the South Zone totalled 18,463.6m in 59 holes.

The drilling achieved three main goals:

delineated Indicated and Inferred Mineral Resources on the Central Zone deposit;

delineated Inferred Mineral Resources on the South Zone deposit;

tested several geophysical anomalies on the Kwanika Property to explore for possible extensions

of the Central Zone and South Zone deposits.



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11.0 Sample Preparation, Analyses and Security All drill core was logged for geological and geotechnical characteristics (geotechnical logging included

rock quality designation (RQD), magnetic susceptibility, and specific gravity), and was photographed,

sampled, and split by diamond saw or core splitter on-site.

All drill core sampled is NQ (4.76cm) size except for the rare occurrence where poor ground conditions

necessitated the reduction of the core diameter to BQ (3.64cm). There are four occurrences in the Central

Zone where holes were collared using HQ size (6.35cm dia.) core.

Assaying of samples was carried out by Global Discovery Labs (GDL) in Vancouver, British Columbia,

and subsequently Acme Analytical Laboratories Ltd. (Acme) after Acme acquired GDL in August 2009.

In RPA’s opinion, the logging and sampling protocols are appropriate for the mineralization style and the

work was carried out to a suitable standard. Sampling methods and approach used by CMG and

Serengeti employees were consistent with common industry best-practice standards. The core boxes and

samples were properly marked and labeled with tags stapled in the boxes. The sampling equipment was

appropriately configured, well maintained, and the core appeared to have been sampled in a correct

manner.

RPA is of the opinion that the samples taken are representative of the mineralization. The average

sample length for all drill programs since 2006 is 1.84m, with a median length of 2m.



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12.0 Data Verification An independent assay QA/QC program has been in place throughout the drill campaigns carried out by

Serengeti. The total number of QA/QC insertions into the sample stream comprised approximately 7% of

all determinations. In RPA’s opinion, the frequency of QA/QC sample insertion met an acceptable

standard.

The data from the Kwanika Project reside in a GIS database compiled by Serengeti personnel. RPA have

carried out a number of validation checks of the database against the original assay reports. Checks were

carried out prior to preparation of the Mineral Resource estimates. The drillhole assay, lithological, and

survey databases were also validated using the GEMCOM utility and found to be free of errors.

In RPA’s opinion, the database is reasonably free of errors and acceptable for use in estimation of

Mineral Resources.



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13.0 Mineral Processing and Metallurgical Testing All the metallurgical aspects presented in this PEA Report rely on the information available from the

Technical Report on the Kwanika Project, NI 43-101 Report dated March 3, 2011.

Copper-Gold mineralization in Kwanika has been identified as two main zones, Central Zone, and South

Zone. Serengeti conducted preliminary metallurgical testing on samples from the Central Zone.

Metallurgical testing of the South Zone has not been conducted.

The preliminary metallurgical test program commenced on November 3, 2008, directed by Scott Wilson

RPA, and executed by SGS Metallurgical Services Ltd. (SGS). A total of 52 samples weighing 186kg

were collected by Serengeti personnel and sent to SGS where equal amounts of each sample were used to

construct a 120kg master composite, with the remaining material being stored for later testing. The

master composite sample assayed 0.66% Cu and 0.76g/t Au.

The Central Zone test work included chemical and mineralogical analyses, Bond Ball Work Index

testing, gravity concentration, batch rougher and cleaner flotation tests, and a locked cycle flotation test.

Eleven batch flotation tests included five rougher tests, and six cleaner tests. Primary grind P80 ranged

from 133m to 75m and regrind P80 ranged from 32m to 20m. Highlights from the Central Zone

test work include the following:

Copper mineralization is mostly chalcopyrite, with minor content of bornite and other copper

sulphides.

Mineralization is finely disseminated, with chalcopyrite 80% liberated at about 25m

Gold appears to be associated with sulphides including pyrite.

Bond Work Index is approximately 16kWh/tonne.

The metallurgical test work concluded that a conventional concentration process would require a primary

grind of 80% passing 75m, and regrinding of the rougher concentrate to 80% passing 26m before

feeding a three-stage cleaning flotation circuit. The final copper concentrate from a locked cycle test

recovered 88.5%Cu, and 65.2%Au with a concentrate grade of 27.7%Cu, and 20.9g/t Au. The final

copper concentrate was found to be very clean, and the content of penalty elements such as As, Bi, Sb,

and Hg is very low.

A follow-up test designed to investigate improved gold recovery from tailings indicates that up to 10.5%

additional gold can be recovered by very fine grinding and flotation of rougher tailings.

The test program is only preliminary in scope, and as such, the composite sample taken from the Central

Zone is not likely to be representative of the entire deposit.

The master composite sample has higher grades than the average PEA mine plan mill feed grades. It is

assumed that any recovery reduction in future test work associated with a reduction of head grade will be

offset by recovery improvements from a more detailed process test work program.

Table 13-1 shows the metallurgical recovery assumptions used for the PEA. These assumptions are

preliminary and will vary with future test work.



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Parameter Value

Copper Recovery 89%

Gold Recovery 70%

Silver Recovery 75%

Molybdenum Recovery 60%

Copper Concentrate 24% Cu



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14.0 Mineral Resource Estimates

14.1 Central Zone

In 2008, RPA carried out an initial estimate of the Mineral Resources for the Kwanika Central Zone

deposit. RPA updated this estimate in 2010, and included silver in the grade interpolation for the first

time. The estimate was made using a block model constrained by 3D wireframes of principal geological

domains. Grade was interpolated into the block model using Ordinary Kriging (OK). The results of the

2010 estimate are shown in Table 14-1 below.

Wireframe models were constructed of some of the principal geological features, as interpreted from the

drilling by Serengeti geologists. Wireframes were also constructed of the topography and the top of the

bedrock.

The estimate was prepared based on a database of diamond drilling results compiled by Serengeti. The

database contained records for 168 drillholes, totaling 69,389m of drilling, of which 109 holes totaling

50,925m were deemed as drilled in or around the Central Zone. Only the sample data from Serengeti’s

holes were used in the Resource Estimate. The cut-off date for the database was October 31, 2010, and

the effective date for this estimate is deemed to be December 31, 2010. The samples were composited to

3.05m (10ft) downhole lengths; with no breaks for lithology.

The search ellipsoids were configured to be identical for all elements within a particular domain, so that

each block that received an estimate for one element would receive an estimate for all. Consequently, the

search ranges tended to be limited by the element with the shortest variogram ranges. The grades were

interpolated in two passes, the first using an octant search and the second using a simple ellipsoidal

search. The block model comprised 10m x 10m x 10m blocks in an array oriented parallel to the UTM

coordinate grid (i.e., no rotation).

Bulk tonnages were estimated using mean density for each domain based on 2,913 measurements

collected on intact core specimens using a water immersion method.

Block model validation was carried out by RPA and adjustments were made at the extreme margins of

the drilled area based on comparisons with drillhole composites. A different methodology, Inverse

Distance Cubed (ID3), was used to check the kriging estimate. The ID3 estimate agreed with the kriged

estimate to an acceptable tolerance and confirmed that the interpolation is reasonable.

The Mineral Resources were classified as Indicated within the area of the deposit drilled to a nominal

spacing of 50m. Blocks outside this area, and comprising a reasonably coherent body constrained by the

search parameters and the extent of the drilling, were categorized as Inferred.

A Lerchs-Grossmann pit optimization was run by RPA to test whether the deposit could potentially be

mined by open pit methods as a large scale operation. Only those blocks captured by the pit shell were

included in the Mineral Resources estimate.

For the 2010 update, the cut-off criterion was changed from a CuEq cut-off to a gross dollar value. The

dollar value was calculated using metal prices of US$1,200/oz Au, US$3.50/lb Cu, and US$21.00/oz Ag,



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which represents an increase of the prices used in 2008 and also included a provision for metallurgical

recovery.

The addition of silver to the estimate and the changes to the reporting and cut-off criteria resulted in a

significant increase in tonnage and a partially compensating decrease in grade. The resulting metal

contents for the 2010 estimate were marginally higher for the Indicated category and significantly higher

for the Inferred.

Table 14-1 Central Zone Mineral Resources – December 31, 2010

Category Cut-Off Tonnage Au Cu Ag

(US$/t) (Mt) (g/t) (M oz) (%) (M lb) (g/t) (M oz)

Indicated 20.00 91.0 0.36 1.06 0.38 754 1.08 3.15

17.50 113 0.33 1.18 0.34 849 0.98 3.56

15.00 142 0.29 1.32 0.31 962 0.89 4.07

12.50 178 0.26 1.47 0.27 1,080 0.81 4.61

10.00 213 0.23 1.59 0.25 1,170 0.74 5.08

7.50 244 0.21 1.66 0.23 1,230 0.69 5.39

Inferred 20.00 5.13 0.27 0.044 0.26 29.0 0.65 0.11

17.50 9.18 0.24 0.071 0.23 45.6 0.56 0.16

15.00 16.6 0.21 0.113 0.20 72.4 0.49 0.26

12.50 26.5 0.18 0.156 0.18 103 0.46 0.39

10.00 39.3 0.16 0.202 0.16 136 0.44 0.56

7.50 55.2 0.14 0.245 0.14 168 0.42 0.74

Notes:


2. Mineral Resources are estimated at a cut-off grade of US$7.50/t. The dollar value cut-off was estimated using

provisions for metallurgical recovery and off-site costs, as described in the text of this report.

3. Mineral Resources are estimated using an average long-term price of US$1,200/oz Au, US$3.50/lb Cu, and

US$21.00/oz Ag.


5. Metallurgical recovery factors of 89% Cu, 70% Au, and 75% Ag were used to derive the dollar value cut-off.

14.2 South Zone

Following completion of a drilling program in 2009, RPA was retained to estimate the Mineral Resources

for the South Zone, which is located approximately 750m south of the Central Zone. Serengeti carried

out a diamond drilling program in 2010 to confirm and extend the known mineralization in the South

Zone. Following the completion of this program, RPA was retained to update the block model.

The estimate was prepared using a block model constrained by 3D wireframe models. Grade was

interpolated into the blocks using ID3 weighting.

The database contained records for 168 drillholes, totaling 69,389m of drilling of which 59 holes,

totalling 18,463.6m, were deemed to have been drilled in and around the South Zone. The cut-off date

for the database was December 31, 2010, and this is the effective date for the estimate.

An additional 58 diamond and percussion holes totaling 5,609m had been drilled by previous operators of

the property. For the pre-Serengeti era holes, the assay data could not be verified, so the older data was

used for interpretive purposes but excluded from the grade interpolation.



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All Mineral Resources in the South Zone are classed as Inferred owing to the sparseness of drilling and

the preliminary nature of the geological interpretation.

The South Zone Mineral Resource estimate is summarized in Table 14-2 below. The update to the

Mineral Resource estimate resulted in a significant increase in tonnage with a partially offsetting drop in

grade, and an overall increase in metal content.

Wireframe models were constructed of some of the principal geological features as interpreted from the

drilling by Serengeti geologists. These features included the bounding East and West faults, along with a

grade shell enclosing the mineralized zone. Wireframes were also constructed of the topography and the

top of the bedrock. These wireframe models were used to assign rock codes to the block model and to

the drill intercepts. For the South Zone estimate, the rock code assignment was whole-block centroid

location, and no percentage factors were applied.

RPA chose to cap the high grades at 1.5% Cu, 1.25g/t Au, 0.1% Mo, and 10g/t Ag. RPA reviewed the

sample lengths and found that over 98% were 3.05m (10ft) or less in length, so a composite length of

3.05m was used.

Attempts to prepare semi-variograms from the composite statistics in the mineralized domain were

inconclusive due to the lack of sufficient numbers of samples in order to provide enough pairs for

variogram generation. Due to the lack of a coherent set of variogram results, ID3 was selected for grade

interpolation over OK. A two pass search strategy, employing a spherical search of radius 200m for Pass

I and a 50m radius in Pass II was used for all components. The first pass required composites from at

least two drillholes in order to estimate a block. The second pass had a smaller search radius, but the

minimum number of samples was reduced to two. This allowed blocks to be estimated by composites

from one drillhole only.

The block model comprised 10m x 10m x 10m blocks in an array oriented parallel to the UTM grid (i.e.,

no rotation)

The average bulk density for the mineralized zone was 2.68t/m3. Bulk density estimates were derived

from 2,913 measurements collected on intact core specimens using a water immersion method. RPA

notes that the density in the South Zone is 3% lower than in the Central Zone.

In RPA’s opinion, local block grade estimates are probably not accurate owing to the comparatively low

data density. Global composite grades agree with global block grades and there does not appear to be

evidence of bias, which means that global block model grades will be reasonable. This is acceptable for

Inferred Mineral Resources, which is the case for the entire South Zone estimate.

A Lerchs-Grossmann pit optimization was run by RPA using somewhat optimistic metal prices and cost

parameters in order to test whether the deposit could potentially be mined by open pit methods as a large

scale operation. Only those blocks captured within the pit shell were included in the Mineral Resource

estimate. As stated above, all Mineral Resources are classified as Inferred.

A dollar value was calculated from the grades, at metal prices of US$3.50/lb Cu, US$1,200/oz Au,

US$17.00/lb Mo, and US$21.00/oz Ag with provision for metallurgical recovery. Metallurgical

recoveries were 89% Cu, 70% Au, 60% Mo, and 75% Ag. The cut-off applied to the block model for



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reporting purposes was US$7.50/t, which, in RPA’s opinion, was consistent at the time of writing with

similar deposits in British Columbia.

Table 14-2 South Zone Mineral Resources – December 31, 2010

Category Cut-Off Tonnage Au Cu Mo Ag

(US$/t) (Mt) (g/t) (M oz) (%) (M lb) (%) (M lb) (g/t) (M oz)

Inferred 20.00 74.3 0.12 0.292 0.33 546 0.012 19.4 2.15 5.15

17.50 98.8 0.11 0.365 0.30 656 0.011 22.9 1.99 6.34

15.00 132 0.11 0.457 0.27 785 0.009 27.0 1.84 7.83

12.50 176 0.10 0.554 0.24 928 0.008 31.8 1.68 9.52

10.00 215 0.09 0.623 0.22 1,030 0.008 35.5 1.56 10.8

7.50 240 0.09 0.664 0.20 1,080 0.007 37.6 1.49 11.5

Notes:


2. Mineral Resources are estimated at a cut-off grade of US$7.50/t. The dollar value cut-off was estimated using provisions

for metallurgical recovery and off-site costs, as described in the text of this report.

3. Mineral Resources are estimated using an average long-term price of US$1,200/oz Au, US$3.50/lb Cu, US$17.00/lb Mo,

and US$21.00/oz Ag.


5. Metallurgical recovery factors of 89% Cu, 70% Au, 60% Mo and 75% Ag were used to derive the dollar value cut-off.

14.3 Economic Pit Limit and Pit Designs

14.3.1 Pit Optimization Method

The economic pit limit is selected after evaluating Lerchs-Grossman (LG) pit cases conducted by MMTS

using Mintec’s MineSight® Economic Planner software (MS-EP). This study is based on the in situ

resource model from RPA as described above. The LG assessment is carried out by generating sets of

LG pit shells by varying revenue assumptions to test the deposit’s geometric/topographic sensitivity.

The economic pit limit is determined by estimating the pit size where an incremental increase in pit size

does not significantly increase the pit resource. Economics of the pit limits are tested by independent

calculations to determine that they are economically viable.

14.3.2 Economic Pit Limit Design Basis

In October 2011, MMTS estimated the mining costs, and process costs, which include process, G&A,

tailings placement, and water treatment costs.

The average unit costs per tonne used in the LG pit program are shown in Table 14-3.



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Table 14-3 Kwanika Input Unit Costs for LG Pits Unit Costs

(CDN$/t Mined)

Unit Costs

(CDN$/t Mill Feed)

Mining $2.30

Process $10.70

G&A $1.10

Water Treatment $0.05

Tailings P lacement $0.05

Total Process Costs $2.30 $11.90

14.3.3 Pit Slope Angles

To date there has not been a geotechnical study to define the pit slope design criteria. As a result, values

were assumed, based on similar projects in the area and are summarized in Table 14-4.

Table 14-4 Kwanika Proposed Bench Design Criteria

Slope Design Element All Open Pits

Bench Face Angle 70°

Pit Slope Angle 40°

Final Bench Height (m) 2x10m

Minimum Catch Bench Width (m) 8m

During prefeasibility studies, additional geotechnical studies on high wall slope parameters should revise

or confirm the pit slopes and the ultimate pit configuration can then be re-designed. Geotechnical studies

will be needed to define the rock strength parameters, structural orientations, discontinuities and faults

and geo-hydrology conditions.

14.3.4 Process Recoveries

Process recovery assumptions used to generate the LG pits are provided by SGS Metallurgical Services

Ltd. MMTS used the same values as RPA and these are shown in Table 14-5.


Metal Process Recovery (%)

Cu 89

Au 70

Ag 75

Mo 60

14.3.5 Metal Prices

The Base Case metal price and NSP is provided in Table 14-6. The NSP is based on three year trailing

average metal prices from September 2011, a US$/CDN$ exchange rate of 0.95, off-site transportation,

refining charges, etc. (the smelter schedule is provided in the Design Basis Memorandum, available in

Appendix C).



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Table 14-6 Metal Prices and NSP Base Case – Sept 2011



Gold US$1,178.78/oz $36.30/gm



LG pits are generated by varying prices in 10% increments in the range from 50–140% of the base NSP.

These prices and resultant pit limits were derived during preliminary pit limit assessment. It has been

deemed as not significant to update the ultimate pit limit for the PEA study using higher current metal

prices. The lower prices used, make the economic limits conservative. Current prices and parameters

have however been included for the cashflow in Section 21.0.

14.3.6 LG Economic Limits

Figure 14-1 and Figure 14-2 summarize Tonnes above cut-off grade (CoG) by LG Pit for the Central and

South Zones. Pit A06 and B06 are using the base case metal price. The CoG used to determine ore vs.

waste within the LG economic limits was NSR=$11.90/tonne.

0

10,000

20,000

30,000

40,000

50,000

60,000

70,000

A01 A02 A03 A04 A05 A06 A07 A08 A09

Cu

mm

Ore

To

nn

es

abo

ve C

oG

Pit Number

Central Zone

Inflection Point

Figure 14-1 Kwanika Central Zone Tonnes above CoG by LG Pit



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-

5,000

10,000

15,000

20,000

25,000

30,000

35,000

40,000

45,000

B04 B05 B06 B07 B08 B09 B10

Cu

mm

Ore

To

nn

es

Ab

ove

Co

G

Pit Number

South Zone

Inflection Points

Figure 14-2 Kwanika South Zone Tonnes above CoG by LG Pit

The inflection points at Pit Case A08 (shown in Figure 14-1) and at Pit Case B06 (shown in Figure 14-2),

occur where an incremental increase in pit size does not significantly increase the pit resource, or an

incremental increase in the pit resource results in only marginal economic return. These inflection points

represent a potential economic pit limit.

Although the inflection point in the Central Pit is at Pit A08, this pit limit is meant for a case generated

with metal prices greater than the base case. The opportunity to mine a larger pit can only be

incrementally justified if more advance engineering show lower costs or economies of scale. Until more

work is done, Pit A06 from Central Zone has been selected as the pit limit. This is a conservative

decision, and is based on the assumption that the underground block cave, directly below Pit A06, will

mine this material anyways.

Pit B06 has been selected as the pit limit for South Zone.

The run of mine resource for the selected LG pit limits are shown in Table 14-7.

Table 14-7 ROM LG Pit Delineated Resource Including Indicated and Inferred Resources


(Mt) (%) (g/t) (g/t) (%)


Inferred 0.9 0.332 0.209 0.840

South Inferred* 23.95 0.334 0.117 1.890 0.016**






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Following the LG economic limit, more detailed pit designs have been developed to include phases and

ramps. The resultant Pit delineated resources suitable for production scheduling are indicated in Table

14-8.

Table 14-8 Run of Mine Pit Delineated Resources from Detailed Designs


(Mt) (%) (g/t) (g/t) (%)


Inferred 0.90 0.332 0.207 0.839

South Inferred* 25.07 0.324 0.120 1.838 0.016**




A plan view, north-south section view and orthographic view of the LG and Detailed pits are shown in

Figure 14-3 to Figure 14-6. Figure 14-3 the red pits are the detailed designs based on the blue LG pits.



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Figure 14-3 Plan View of the Kwanika LG Pit Limits



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Figure 14-4 Kwanika Economic Pit Limits – Orthographic View from the East



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Figure 14-5 Kwanika Central LG & Detailed Pit Limits – NS Section at East 351500 Looking

East

Figure 14-6 Kwanika South LG & Detailed Pit Limits – NS Section at East 352400 Looking

East



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14.3.7 Underground Mine Design

There are sufficient tonnes and grade below the Central Open Pit to support an underground mine.

Several different mining methods have been evaluated including block caving. AMEC has reviewed the

drill core and the targeted higher grade mineralized zone and considers Block Caving to be a viable

mining method. Accordingly MMTS developed Block Cave stopes and infrastructure based on typical

parameters which include caveability, requirements for rock pre-conditioning, geotechnical aspects of the

rock, development rates and mining loss and dilution factors. These parameters will require confirmation

and refinement in future studies.

The underground designs include two stope outlines, access development spiraling down from the bottom

of Central Open Pit, ventilation development and undercut, drawbell, cross-cut and extraction level

designs for block cave mining. At this stage, the stope outlines are conservative and planning includes an

estimate of equipment and rate of development from general factors. Optimization will be needed in

future designs to maximize the underground extraction and optimize the ROM head grades from the

underground operations. Subsequent to this evaluation there are zones of material adjacent to the stopes

used in this study which have sufficient grade to be included in future studies. The underground

delineated resources are listed in Table 14-9.

Table 14-9 Run of Mine Underground Delineated Resource Including Indicated and Inferred

Classes


(Mt) (%) (g/t) (g/t)

UG1 Indicated 19.92 0.456 0.467 1.359 -

Inferred 0.05 0.332 0.418 1.788 -

UG2 Indicated 7.82 0.472 0.470 1.273 -


An orthographic view of the Underground Stopes is shown in Figure 14-7.



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Figure 14-7 Kwanika Economic Underground Stopes – Orthographic View

Note - The Open Pit and Underground Mineral resources stated in Table 14-7 and Table 14-9 are

preliminary in nature, and include inferred mineral resources that are considered too geologically

speculative to have the economic considerations applied to them that would enable them to be

categorized as mineral reserves in future studies. There is no certainty that the results of the preliminary

economic assessment will be realized.

14.3.8 Further Work towards Open Pit Design

The Central and South pits are designed to their break-even economic limit and provide low cost mining

for the near surface material. The lowest benches have the highest incremental strip ratio and are the

higher cost on both a per tonne mined and per tonne milled basis. Optimization of the mining method at

the open pit/underground interface has the potential to improve the overall economics of the property.

Material on the bottom benches of the pit may produce a better net return if mined as part of the

underground instead. An evaluation is required to determine this.

Geotechnical field investigations, including drilling, is required and the open pit design needs to be taken

to a Prefeasibility study level. This redesign work will also be based on the Underground mining limit

optimization, and equipment optimization studies, and water and waste rock management details.

14.3.9 Further Work towards Underground Design

The mineral zone evaluated for underground mining and the rock characteristics are suitable for Block

Caving. This mining method will provide a low operating cost.



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Additional economic underground resources are possible both below and adjacent to the block cave stope

outline used in this report. Optimization of the underground mine plan should add to these potential

economic resources. This should include the use of optimization software to define the economic

footprint and also evaluation of the preferred development/extraction levels.

After the conceptual economic stope limits and preferred extraction levels are optimized, planning and

design can proceed at a Pre-Feasibility level. This will require geotechnical studies and Engineering to

evaluate caving parameters, caving rate, pre-conditioning requirements, and draw point design details.



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15.0 Mineral Reserve Estimates The current study is at a PEA level and therefore there are currently no Mineral Reserves estimated for

the Kwanika Project.



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16.0 Mining Methods

16.1 Summary

A production schedule based on a 15,000t/d mill feed rate at a preliminary assessment level has been

developed for the Kwanika Project. The pit phases are engineered based on the results of an updated

economic pit limit analysis. The underground stopes are engineered based on the results of a cut-off

grade analysis (refer to 14.3.7). A summary of Indicated and Inferred pit resources and underground mill

feed for production scheduling is provided in Table 16-1 using whole block grades with mining dilution

and loss varying by extraction method (open pit versus underground).

The copper (Cu), gold (Au), silver (Ag) and Molybdenum (Mo) grade items used in this section of the

study are based on the resource model provided by RPA as described in Section 14.0 with mining loss

and dilution applied.

Table 16-1 Summarized Indicated and Inferred Pit and Underground Mill Feed for

Production Scheduling


(Mt)

Cu

(%)

Au

(g/tonne)

Ag

(g/tonne)

Mo

(%)*

Mine Rock

(Mt)

Strip Ratio

(t:t)

Pit CK615 11.05 0.37 0.331 0.978 26.38 2.39

Pit CK625 9.75 0.28 0.218 0.722 30.48 3.13

Pit SK612 7.19 0.40 0.059 2.005 0.011 25.70 3.57

Pit SK622 17.88 0.29 0.140 1.821 0.018 75.55 4.23

UGTALL 19.97 0.46 0.467 1.360 0 0

UGWEST 7.82 0.47 0.470 1.273 0 0

Note: Open Pit NSR cut-off used is $11.90/tonne with a provision for mining loss of 5% and dilution of 2%. Underground mill feed has a

provision for mining loss of 10% and dilution of 15%.

* Moly only present in South Zone

The pit and underground delineated resource by assurance of existence class is shown in Table 16-2.

Table 16-2 Pit and Underground Delineated Resource by Assurance of Existence Class

Assurance of

Existence Class

Mill Feed

(Mt)

Cu

(%)

Au

(g/t)

Ag

(g/t)

Mo

(%)

Indicated 47.6 0.41 0.39 1.14 0.000

Inferred 26.0 0.32 0.12 1.84 0.015

16.2 Introduction

The mine planning work for this study is based on the 3D block model (3DBM) created by RPA for the

NI 43-101 published resource model, dated March 2011 and uses MineSight® software, including the

resource model, pit optimization (MS-EP), detailed pit design, and optimized production scheduling

(MineSight® Strategic Planner [MS-SP]).



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In addition to the geological information used for the block model, other data used for mine planning

included the base economic parameters, mining cost data derived from supplier estimates and data from

other projects in the local area, and technical information from other consultants for recommended pit

slope angles, projected project metallurgical recoveries, plant costs and throughput rates.

16.3 Mining Datum

The project design work is based on NAD83 coordinates. The historical drillhole information is based on

various surveys with different sets of control that have been converted to NAD83 and, in particular, a

topography surface produced from orthographic data.

16.4 Production Rate

A number of factors are considered when establishing an appropriate mining and processing rate. Key

factors in relation to Kwanika include:

Resource Size: A typical mine life is 12.5 to 20 years; beyond this, time- value discounting

shows an insignificant contribution to the NPV of the project, and capital investment typically is

targeted at projects with a payback period of 3 to 5 years.

Operational Constraints: Power, water, critical supplies, or limited infrastructure for operations

support can limit production.

Construction Constraints: Physical size and weight of equipment and shipping limits can

determine the maximum size of available units. For this evaluation, the largest proven units are

assumed.

Project Financial Performance: Generally, economies of scale can be realized at higher

production rates, and lead to reduced unit operating costs. These are tempered to the above

mentioned physical and operational constraints and generally higher capital requirements for

higher tonnage throughputs.

Higher production rates generally pay back fixed capital at a faster rate, thereby improving project NPV.

A throughput of 15,000t/d sets the Kwanika mine life at 13.5 years for the Updated PEA mining resource.

With the current mining resource, it is unlikely that a project NPV improvement will be achieved by

significantly increasing the mill throughput. If the mineable resource base is significantly increased in

future studies, the NPV advantage of a higher throughput rate should be investigated.

16.5 Mine Planning 3D Block Model

Two resource models are used in this study, based on the Resource Statements in the Technical Report

dated March 2011 by RPA.

The resource models contain whole block Cu (%), Au (g/t), Ag (g/t) and Mo (%) grades.

The resource models also contain an SG (density) item and a topography (TOPO) item representing the

proportion of a block below the topographic surface. The RPA 3DBMs, have been imported into

MineSight® (CKWAN15.DAT, SKWAN15.DAT) and extra items added for mine planning tasks.

The PEA model dimensions are provided in Figure 16-1 and Figure 16-2. A short list of the mine

planning 3DBM items is given in Table 16-3. The total Central model area is illustrated for orientation in

plan view in Figure 16-3. The total South model area is illustrated for orientation in plan view in Figure

16-4.



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Note: X = Easting, Y = Northing, Z = Elevation.

Figure 16-1 Kwanika Central Mine Planning Model Limits

Note: X = Easting, Y = Northing, Z = Elevation.

Figure 16-2 Kwanika South Mine Planning Model Limits

Table 16-3 3DBM Items

Item Item By Description

TOPO RPA Percent of block below surface

ROCK RPA Rock Type Code

AU RPA Gold Grade (g/t)

CU RPA Copper Grade (%)

CUEQ RPA Calculated Copper Equivalent %, based on March 2010 model

CLASS RPA Resource category (1=Measured; 2=Indicated; 3=Inferred)

DOL$ RPA Dollar Value of block

AG RPA Silver grade (g/t)

MO RPA Molybdenum grade (%)

SG MMTS Bulk density (tonnes/m3)

CUEQR RPA Calculated Copper Equivalent %, based on March 2011 Model

NSR MMTS Net Smelter Return ($/t), based on Sept 2011 NSP

CUEQ2 MMTS Calculated Copper Equivalent %, based on Sept 2011 NSP



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Figure 16-3 Kwanika Central Model Area with NSR >$11.90/t Grade Shell – Plan View with

5m Topography Contours



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Figure 16-4 Kwanika South Model Area with NSR >$11.90/t Grade Shell – Plan View with 5m

Topography Contours

16.5.1 Net Smelter Return

Mill feed cut-offs are determined using the NSR in CDN$/t, which is calculated for each block in the

3DBM using the Net Smelter Price (NSP) as calculated in Appendix C. The NSR (net of offsite charges

and onsite mill recovery) is used as a cut-off item for break-even mill feed/mine rock selection and for

the grade bins for cash flow optimization. The NSP is based on base case metal prices, US dollar

exchange rate, and offsite transportation, smelting and refining charges, from earlier work estimated in

late 2011 (Appendix C). The metal prices and resultant NSPs used are shown in Table 16-4.



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Table 16-4 Metal Prices and NSP from September 2011



Gold US$1,178.78/oz $36.30/gm



Metallurgical recoveries used for the NSR calculations are based on test work done by SGS Metallurgical

Services Ltd and are detailed in Section 13.0.

The method of calculating NSR is as follows:

NSR (CDN$/t recovered) = dolval_Cu + dolval_Au + dolval_Ag + dolval_Mo

Where:

Dolval_Cu = Cu% / 100 x (NSPCu) x 2204.62 lb/tonne x (RecCu)

Dolval_Au = Au g/tonne x (NSPAu) x (RecAu)

Dolval_Ag = Ag g/tonne x (NSPAg) x (RecAg)

Dolval_Mo = Mo% / 100 x (NSPMo) x 2204.62 lb/tonne x (RecMo)

NSP = Net Smelter Price

Rec = Recovery %

16.5.2 Mining Loss and Dilution

The Kwanika Project is planned as an open pit and underground operation; suitable mining parameters

have been used to match the deposit characteristics with the proposed mining method and equipment size.

The 3D block model (3DBM) block dimensions in plan view are 10m Northing x 10m Easting with block

heights of 10m.

The pits are planned to be selectively mined with large shovels and trucks to achieve a mill feed rate of

15,000t/d. The majority of the mine rock mining will be carried out with large equipment at high mining

rates. Blasthole sampling will be used to determine the mine rock/resource boundaries for identifying

material designations on the pit bench for daily operations in the Ore Control System (OCS). Blasthole

cuttings will be assayed across each bench giving a higher resolution of resource and mine rock than the

3DBM used in this study, which has been built from wider spaced exploration drillholes. From the

blasthole assays, resource and mine rock boundaries will be defined for the production shovels and the

OCS. This methodology is typical of ore control in operations in porphyry deposits.

Mining dilution for the open pit assumes whole block dilution from the grade interpolation. Additional

dilution has been added to account for material on ore/waste boundaries. At this time the grade of the

dilution is set at zero even though it will close to the cut-off grade at the cut-off boundary.

Mining loss is an allowance for material lost at the ore/waste digging line in the pit and through

misdirected loads, spillage, etc. during mining.

The mining reserves used for scheduling are estimated from grades in the 3DBM within the detailed pit

designs with the appropriate mining loss and dilution applied as described above. The mining recoveries

and dilution convert the in situ resource material tonnages into a ROM mill feed.



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For the underground there is no opportunity for selectivity or grade control. Global loss and dilution is

included in the mining resource by including all material within the slope limits regardless of grade. A

further allowance is included to account for difference in the designed stope limit and the eventual actual

caved stope.

For this PEA, a global factor has been estimated for mining loss and dilution for both the Open Pit and

Underground. The estimated overall values are shown with Table 16-1.

16.5.3 Detailed Pit Designs

MMTS has completed PEA-level pit designs demonstrating the viability of accessing and mining

economical resources at the Kwanika site. The designs are developed using MineSight® software,

estimated geotechnical parameters, suitable road widths for the equipment size, and minimum mining

widths based on efficient operation for the size of mining equipment chosen for the Project.

16.5.4 Haul Road Widths

Haul road widths are designed to provide safe, efficient haulage, and to comply with the following BC

mines regulations:

for dual lane traffic, a travel width of not less than three times the width of the widest haulage

vehicle used on the road, plus an allowance for shoulder barrier(s)

where single lane traffic exists, a travel width of not less than two times the width of the widest

haulage vehicle used on the road, plus an allowance for shoulder barrier(s)

shoulder barriers of at least 3/4 of the height of the largest tire on any vehicle hauling on the road,

placed along the edge of the haulage road wherever a drop-off greater than 3m exists; the

shoulder barriers are designed at 34° slope (slightly less than the angle of repose); the total road

width equals the barrier width plus the travel width.

Figure 16-5 to Figure 16-8 show typical road cross sections for haul roads.

Ditches are included in the travel width allowance. Ditches are not added to the in-pit high wall roads;

there is adequate water drainage at the edge of the road between the crowned surface and lateral

embankments, such as high walls or lateral impact berms. During run-off, when water is flowing, this

ditch allowance is still part of the running surface, and can be used as lateral clearance for haul trucks. It

can also be driven on, if required, to avoid obstructions. In practice, specifically designed excavated

ditches in haul roads tend to be filled in by road grading; and when maintained as open ditches, can create

a hazard if the wheel of a haul truck or light vehicle should happen to get caught in them. Avoiding the

addition of a ditch width to the three-truck travel width on the in-pit high wall roads can significantly

reduce the pit mine rock stripping.

Based on a 136t truck, the haul road design basis is as follows:

largest vehicle overall width: 3.5m

double lane high wall haul road allowance: 17.3m

double lane external haul road allowance: 24.1m

single lane high wall haul road allowance: 13.8m

single lane external haul road allowance: 20.6m


NI43‐101 Technical Report for the Kwanika Project Prepared for Serengeti Resources Inc.

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Figure 16-5 Dual Lane High Wall Haul Road Cross Section

Figure 16-6 Dual Lane External Haul Road Cross Section


NI43‐101 Technical Report for the Kwanika Project Prepared for Serengeti Resources Inc.

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Figure 16-7 Single Lane High Wall Haul Road Cross Section

Figure 16-8 Single Lane External Haul Road Cross Section



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16.5.5 Design Standards

Design parameters for the pit phases and the mine rock storage facility and tailings storage facility (RSF

& TSF) are estimated based on similar projects in the project area.

Minimum Mining Width

A minimum mining width between pit phases is reserved to maintain a suitable mining platform for

efficient mining operations. This width is established based on equipment size and operating

characteristics. For this Kwanika PEA, the minimum mining width generally conforms to 30m, which

provides sufficient room for two-sided truck loading but, due to the configuration of merging pits, it is

sometimes less.

Access Considerations

For this PEA an access ramp is not designed for the very last bench of each pit bottom, on the assumption

that the ramp is mill feed grade and will be removed upon retreat.

Road grades are designed at a maximum grade of 8%. Steeper roads (10%) can be considered after more

weather data has been accumulated. Switchbacks are designed flat, with ramps entering and exiting at

design grade. In practice however, grades are transitioned such that visibility and haul speeds are

optimized going around the switchback. Where possible, switchbacks are located such that they tie into

future phase access development.

Access up from the lowest pit benches requires a spiral ramp that exits at the eastern side of the pit rim

and joins with infrastructure features. Switchbacks and flat grade segments are minimized.

Bench Height

The Kwanika pit designs are based on the digging reach of the large shovels (10m operating bench) with

double benching between high wall berms; therefore, there are berms every 20m vertical. Single

benching is employed, if necessary, to maximize recovery and maintain the safety berm sequence as

warranted.

Berm Width

A minimum safety berm width of 8m is assumed. Where haul roads intersect designed safety berms, the

haul road width is counted towards the safety berm width for the purpose of calculating the maximum

overall pit slope angle.

16.5.6 LG Phase Selection

The LG pits discussed previously are used to evaluate alternatives for determining the economic pit limit

and the optimal pushbacks or phases before commencing detailed design work. LG pits provide a

geometric guide to detailed pit designs. Details considered are the addition of roads and bench access,

removal of impractical mining areas with a width less than the minimum working width, and ensuring the

pit slopes meet the detailed geotechnical recommendations.

The LG pit cases selected as the economic pit limits for the Kwanika mine area discussed above are Pit

Case A06 and B06 (detailed in Section 14.3).

Smaller pit shells exist within the ultimate pit limits that have higher economic margins due to lower strip

ratios or better grades. Mining these pits as phases from higher to lower margins, provides the



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opportunity to maximize revenues and minimize mining costs at the start of mining operations and

thereby shortens the project capital payback. Where a higher number of smaller push-backs increase this

effect, it needs to be balanced with the higher efficiencies and resultant lower unit mining costs of big

mining areas from bigger pushbacks. The first phases (starter pits) have the greatest effect on capital pre-

stripping requirements.

The lower LG price case pits are used as guides to place the initial starter pit with the following

constraints.

The starter pit must be:

large enough to accommodate the multiple unit mining operations of drilling, blasting, loading,

and hauling

have bench sizes large enough so the number of benches mined per year is reasonable (sinking

rate)

wide enough so the shovels can load the trucks efficiently

The description of the detailed pit design phases uses the following naming conventions:

The prefix “CK” indicates Central Zone,

The prefix “SK” indicates South Zone,

The first digit signifies the original LG Pit Case used,

The middle digit signifies the pit phase number,

The last digit signifies the revision number,

A suffix of ‘i’ indicates that the resource tonnage is for that phase, incremental from the previous

phase. If there is no ‘i’ specified, it is cumulative up to the phase indicated. (E.g. CK625i means

the incremental resource tonnage from CK615 to CK625.)

16.5.7 Kwanika Detailed Pit Phase Designs

The Kwanika pit design includes two phases for each zone (Central and South). Access to each bench is

continuously provided by ramps built into the high walls.

Kwanika Central Phase CK615

Mining of Phase CK615 begins during pre-production. The design intention of the first phase is to

expose mineralized material for the mill start-up with a minimum pre-strip and provide continuous feed

to the mill once the phase is completed. This phase begins mining at a bench elevation of 1012m on the

south end of the mining area and is mined down to a bottom pit elevation of 830m. The haul road from

the pit bottom reaches the surface at an elevation of 988m on the east side of the pit where it connects to

external haul roads. The placement of ramps in the pit should be further optimized in future studies to

reduce haul cycle times. An illustration of pit CK615 is provided in Figure 16-9.



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Figure 16-9 Plan View of Kwanika Central Pit CK615



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Kwanika Central Phase CK625

Pit Phase CK625 expands beyond CK625 on all but the Eastern-most side. It originates at a bench

elevation of 1012m and is mined down to a bottom pit elevation of 780m. The haul road from the pit

bottom exits the pit at an elevation of 988m. CK625 is incremental to CK615. An illustration of the Pit

CK625 is provided in Figure 16-10.

Figure 16-10 Plan View of Kwanika Central Pit CK625



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Kwanika Phase SK612

Mining of Phase SK612 begins after CK625 has been mined out. The design intention of the first phase

is to expose mineralized material for the mill with a minimum pre-strip and provide continuous feed to

the mill once the phase is completed. This phase begins mining at a bench elevation of 1030m on the

south end of the mining area and is mined down to a bottom pit elevation of 840m. The haul road from

the pit bottom reaches the surface at an elevation of 980m on the east side of the pit where it connects to

external haul roads. The placement of ramps in the pit should be further optimized in future studies to

reduce haul cycle times. An illustration of Pit SK612 is provided in Figure 16-11.

Figure 16-11 Plan View of Kwanika South Pit P612



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Kwanika Phase SK622

Pit Phase SK622 is the final expansion of the South Pit, starting at a bench elevation of 1056m and is

mined down to a bottom pit elevation of 760m. SK622 is incremental to SK612. The haul road from the

pit bottom reaches the surface at an elevation of 980m on the east side of the pit. An illustration of the pit

is provided in Figure 16-12.

Figure 16-12 Plan View of Kwanika South Pit SK622



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A north-south section view of all the Kwanika Central Pit phases is shown in Figure 16-13. A north-

south section view of all the Kwanika South Pit phases is shown in Figure 16-14.

Figure 16-13 North-South Section View of all Central Pits at East 351500 – Looking East

Figure 16-14 North-South Section View of all South Pits at East 352400 – Looking East



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16.5.8 Kwanika Central Underground Mining

The central underground block cave mine is directly below CK625 and is split into 2stopes – “Tall” and

“West”. The tall stope originates at elevation 780m and is mined down to a bottom elevation of 540m.

The West stope branches off the Tall stope from elevation 560m and is mined down to elevation 480m.

The access ramp to both underground stopes originates from surface at the South East corner of Central

Pit at an elevation of 990m. An illustration of the underground stopes is provided in Figure 16-15.

Figure 16-15 Side View of Kwanika Underground Stopes

16.5.9 Pit Delineated Resources

Pit delineated resources are estimated using the previously discussed parameters. The pit delineated

resources are listed above in Table 16-1.

16.6 Mine Plan

16.6.1 LOM Production Schedule

The Kwanika mine production schedule is developed with MS-SP, a comprehensive long-range

scheduling tool for open pit mines. It is typically used to produce a LOM schedule that will maximize

the NPV of a property, subject to user specified conditions and constraints. Annual production

requirements, mine operating considerations, product prices, recoveries, destination capacities, equipment

performance and operating costs are used to determine the optimal production schedule. The

underground resource has not been scheduled using MS-SP but is added into the schedule separately.



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Scheduling results are presented by period as well as cumulatively and include:

tonnes and grade mined by period broken down by material type, bench, and mining phase

tonnes transported by period to different destinations (mill, stockpiles, and RSF)

The mine schedule considers Time 0 the time that the mill starts; the full capacity production of mill feed

is expected in Year 1. The production schedule specifies:

Pre-production: Year -1

Mine Load and Haul Fleet Selection

The mine load and haul fleet has been selected prior to production scheduling. Similar projects in the

area have shown that the lowest cost-per-tonne fleet of shovels and haul trucks for large hard rock open

pit mines that are currently being used at BC mines are the 15m3 bucket, diesel-hydraulic shovel matched

with the 136t truck.

Diesel hydraulic drills with 150mm hole size are chosen to match this size of truck/shovel fleet. The

following performance and costs are estimated based on this fleet.

Productivities of the selected equipment are derived from truck/shovel matching studies, and include

detailed truck haul cycle estimates, for multiple pit-to-destination combinations.

Load times for the shovels include operator efficiency. Details for the load times are included in

Appendix C.

In order to optimize the project NPV, grade bins have been specified based on NSR block values. The

MS-SP optimizer develops a COG strategy to increase the project NPV by stockpiling lower grade

material for processing later in the LOM schedule, increasing mill head grades and, therefore, revenues

early in the production schedule. Blast hole assays will be used for mill feed COGs and for a COG

strategy. The material types specified in Table 16-5 are COGs used for selectivity within the MS-SP

optimized scheduler. Mining operations will not use this many grade bins in actual operations.

Table 16-5 Material Types Defined for MS-SP

NSR Grade Bins

(CDN$/t)

Sub Grade <11.90

Low Grade 15.50

Mid-Low Grade 19.50

Mid Grade 25.50

Mid-High Grade 34.50

High Grade >34.50



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The MS-SP schedule utilizes these main criteria in each period to maximize the NPV (the design basis is

provided in Appendix C):

Mining precedence (Phase 2 after Phase 1)

the haul cycle time (including 0.5 min dump and maneuver), and resultant variable unit cost

shovel productivity (including 2.75 min load & exchange time per hauler, 70% overall

efficiency)

estimated operating and capital costs, process recoveries, and metal prices

360 mine operating days scheduled per year and 24h/day

annual mill feed of 5,475kt/a is targeted based on an average throughput of 15,000t/d

Cut-off Grade Optimization

Typically, the mill feed grade can be increased by sending low and mid-grade classes to stockpiles in

early periods of the production schedule. The mill feed grade is maximized and this effectively increases

the revenue per tonne milled early in the schedule. A small amount of sub-grade stockpiling has been

employed however; further optimization of stockpile usage will be performed in future studies.

Schedule Results

The summarized production schedule results are shown in Table 16-6. Tonnes and grades are ROM from

the resources reported in Table 16-1. Full results are in Appendix C.



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Table 16-6 Life of Mine Production Summary



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Figure 16-16 shows the LOM mill feed production schedule.

0.00

0.20

0.40

0.60

0.80

1.00

1.20

1.40

1.60

1.80

0

1000

2000

3000

4000

5000

6000

AU

g/t

A

g g/

t

C

U %

kTo

nn

es

Mill Feed UG to Mill Open Pit to Mill

CU AU

Ag

Figure 16-16 ROM Mill Feed Sources and Mill Head Grades for Feed Cu, Au, Ag, and Mo

The cumulative strip ratio (open pit waste divided by [open pit mill feed + open pit to stockpile]) by

period is shown in Figure 16-17.

0.0

0.5

1.0

1.5

2.0

2.5

3.0

3.5

4.0

4.5

5.0

Stri

p R

atio

t

:t

Period

Strip Ratio Cum Strip Ratio

Figure 16-17 Cumulative Open Pit Strip Ratio



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16.6.2 Rock Storage Facilities

Design Parameters

All RSFs are designed with a natural angle of repose of 37°, with exterior edges re- sloped at the time of

placement to 3:1. A 20% swell factor is applied to in situ volumes to calculate the volumes that need to

be placed.

Further details are available in Section 18.9.

Construction Methods

Mine rock placement is done using primarily bottom-up construction methods. Bottom-up placement

involves the truck placing the material in lifts up to 30-60m high and constructing the RSF to final limits

from the bottom working upwards. Some top-down placement, involving trucks dumping the material

from the top bench crest down to the platform or topography below may be required.

Tailing Storage

A portion of the mined waste rock will be used to construct the confinement for the tailings storage

facility (TSF). This construction must provide sufficient tailings storage volume within its footprint. As

a minimum requirement, the TSF surrounds the flotation tailings with at least with an elevation

approximately 10m higher than the edge of the tailings surface. The geometry of the tailings containment

dam will be explored in more detail in the next level of study.

Further details are available in Section 18.9.

Foundation Preparation

Necessary stability studies will be conducted for foundation preparation requirements and will be

followed to allow for stable RSF and TSF construction.

Sub Grade ROM Stockpile

There is a sub-grade stockpile throughout the mining schedule. A long term sub-grade stockpile is placed

close to the primary crusher in order to maximize the grade of the plant feed and smooth strip ratio and

fleet requirements. In the mine production schedule, the sub-grade stockpile reaches a maximum size of

0.6Mt.

Annual Mine Waste Rock Volumes

Annual mine waste rock volumes produced from the 15,000t/d schedule are shown by pit phase by year

in Table 16-7.



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Table 16-7 Mine Waste Rock Tonnages by Area and Year (kt)

Year Pit Phase (kt)

Total CK615 CK625i SK612 SK622i

-1 11,999 - - - 11,999

1 12,877 122 - - 12,999

2 1,504 11,779 - - 13,283

3 2 15,994 - - 15,997

4 - 2,582 14,249 - 16,830

5 - - 8,669 8,080 16,749

6 - - 2,314 14,435 16,748

7 - - 471 16,278 16,749

8 - - - 16,568 16,568

9 - - - 8,196 8,196

10 - - - 4,407 4,407

11 - - - 3,526 3,526

12 - - - 2,339 2,339

13 - - - 1,698 1,698

14 - - - 24 24

Total 26,383 30,477 25,703 75,549 158,112

RSF Access Roads

Access to the RSF throughout the life of mine is via a double-lane external mine rock haul road (24.1m),

from the south end. This road network also connects the mining areas with the stockpile and plant areas.

Final RSF Configuration

The final RSF for the Kwanika Project has overall slope angles of 26° (2:1). The final post closure

configuration is adapted in accordance with the closure plan. Costs for this work are included by keeping

the ancillary equipment in use during the later years of the operation after the mine rock strip ratio drops

to low levels. This allows for reclamation in the latter part of the LOM schedule. Future studies will

look for earlier reclamation opportunities.

16.6.3 Mine Pre-Production Detail

Pre-Production Description

There are three primary objectives for mine pre-production development:

expose sufficient mill feed for start-up

establish mining areas that will support the equipment required to achieve mill feed requirements

on a sustainable basis

act as a source of material that is required for construction of the mine, mill, initial TSF berm and

site infrastructure



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Mine pre-production site development activities are currently scheduled to start in Year -1 in order to

meet the timeline for overall site development. Site development for the mine area will consist of:

topsoil salvage and foundation preparation

drainage control and water management structures

access to construction and initial mining areas

initial pit bench development

o haul road construction

o infrastructure construction

Topsoil Salvage and Foundation Prep

Topsoil salvage and foundation preparation will likely be required. An estimate of the extent of

foundation preparation and topsoil salvage will be performed in future studies.

Mine Drainage

The primary purpose of the diversion ditch network is to prevent contact surface water from impacting

certain areas. These diversion ditches are primarily located around the perimeter of the pit, the RSF, the

TSF, the sub-grade stockpile and all mine haul roads. A water diversion channel is also required for

Kwanika Creek which currently runs eastward through the future tailings storage facility and southward

through the future South Pit mining area.

The contact water will be directed into a water collection pond downstream of the TSF, where the water

will be reclaimed to the process plant.

Details on mine drainage are available in Section 18.9.

Mine Power

The mine requires electric power for mine offices, mine maintenance facilities, explosive manufacturing

and storage facilities. All equipment used in the open pit and underground are diesel powered and do not

require an electric power supply.

Mine Infrastructure Construction

Site preparation allowance has been made for the mine equipment assembly area, explosives

manufacturing plant, ammonium nitrate prill storage, and explosives magazines. Facilities such as the

mine offices, maintenance shops, and fuel tanks will be available at the mine site before pre-production

mining commences.

Initial Pit Development

Once the access roads are in place, the larger mine equipment will have access to the working areas and

will commence mining. The Phase 1 pit is small in area, but due to the relatively flat terrain, pre-

stripping operations can be carried out in a typical drill, blast, load and haul method. The pre-production

mining is designed to ensure that sufficient working area is developed to provide sustainable mill feed

after start-up. Any mill feed grade material encountered will be stockpiled for processes after start-up.



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16.6.4 Mine Production Detail

End-of-period (EoP) mine status maps have been developed and are shown in Appendix C. Each period

is described below with illustrative EoP figures.

Pre-Production

During the pre-production period, CK615 is mined to an elevation of 940m. Mine-rock material is stored

in the RSF, up to an elevation of 1020m, and in the tailings containment dam to an elevation of 1045m

for the flotation tailings in later periods. Topsoil is stockpiled to the northwest of the mining area. A

berm for the leach residue impoundment is established east of the pit. The North Kwanika Creek

diversion is also constructed north-east of the TSF to divert Kwanika Creek around the TSF. Mill feed

material from CK615 will be brought to the stockpile.

A plan view of the Pit/RSF/TSF during the pre-production stage is shown in Figure 16-18.

Year 1

By the end of Year 1, CK615 is mined to 910m and CK625i is mined to 1010m elevation. Mine rock

from the pits is placed on the North RSF to an elevation of 1030m and expands the tailings containment

dam at an elevation of 1045m. The majority of mill feed material from CK615 and CK625i will be taken

directly to the mill. A small portion will however be sent to the sub-grade stockpile. The material sent to

the stockpile in the pre-production phase is reclaimed as mill feed in Year 1. Flotation tailings from the

plant will be contained within the crescent shaped TSF constructed during the pre-production phase.

Year 2

By the end of Year 2, CK615 is mined to 850m and CK625i is mined to 960m elevation. A portion of the

mine rock from the pits extends to the tailings containment dam at an elevation of 1045m. All other mine

rock is transported along the haul road to the North RSF to an elevation of 1040m. Underground

development continues. All mill feed material from the pit is taken directly to the mill and a portion of

the sub-grade stockpile is reclaimed as mill feed.

Year 3

By the end of Year 3, CK615 is completely mined out to 830m and CK625i is mined to 860m elevation.

A portion of the mine rock from the pits extends to the tailings containment dam at an elevation of

1045m. All other mine rock is transported along the haul road to the North RSF to an elevation of 1050m.

Underground development continues. All mill feed material from the pit is taken directly to the mill and

none of the sub-grade stockpile is reclaimed as mill feed.

Year 4

By the end of Year 4, CK625i is mined out to 780m elevation, UGTALL has started production and

SK612 is mined to 950m elevation. A portion of the mine rock from the pits is transported to the tailings

containment dam to an elevation of 1050m. All other mine rock is transported along the haul road to the

North RSF to an elevation of 1060m. Underground development continues. All mill feed material from

the pit is taken directly to the mill and none of the sub-grade stockpile is reclaimed as mill feed.

The Pit/RSF/TSF plan view at the end of Year 4 is shown in Figure 16-19.



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Year 5

By the end of Year 5, UGTALL continues production, SK612 is mined to 880m elevation, and SK622i is

mined to 1010m elevation. A portion of the mine rock from the pits is transported to the tailings

containment dam to an elevation of 1060m. All other mine rock is transported along the haul road to the

North RSF to an elevation of 1070m. Underground development continues. All mill feed material from


Year 10

Backfill of Phase SK612 commences in Year 8. At the end of Year 10, UGTALL continues production,

SK612 is completely mined out to 840m elevation, and SK622i is mined to 880m elevation. The tailings

containment dam is complete, the North RSF is no longer in use and the South RSF is raised to 1000m.

Mill feed material from the pit is sent to both the mill and the stockpile and some of the low-grade

stockpile is reclaimed to the mill. Net inventory in the stockpile decreases. All mill feed material from


Year 14/Life of Mine

At the end of Year 14, all pits and all underground stopes are mined to completion. Mine-rock from the

pit brings the final elevation of the North RSF to 1100m, the tailings containment dam to 1083m and in-

pit waste storage in SK612 to an elevation of 1040m. All sub-grade stockpile inventory is reclaimed as

mill feed.

The Pit/RSF/TSF plan view at the end of Year 14 is shown in Figure 16-20.



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Figure 16-18 EOP – Time 00 End of Pre-production – Start of Production



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Figure 16-19 EOP – Year 4



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Figure 16-20 EOP – LOM



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16.7 Mine Operations

The mining operations are typical of open-pit operations in northern British Columbia and employ

accepted bulk mining methods and equipment. There is considerable operating and technical expertise,

services, and support in northern British Columbia. A large capacity operation is designed and large-

scale equipment is specified for the major operating functions in the mine to generate high productivities,

which reduce unit mining costs. Large-scale equipment also reduces the on-site labour requirements, and

dilutes the fixed overhead costs for mine operations.

16.8 General Organization

The Kwanika operations are organized as illustrated in Figure 16-21. Mine operations is organized into

three areas: direct mining, mine maintenance, and general mine expense (GME). Other areas of the

organization are dealt with elsewhere in the report.

The direct mining area accounts for drilling, blasting, loading, hauling, and pit maintenance activities in

the mine. Costs collected for this area include the mine operating labour, mine operating supplies,

equipment operating hours and supplies, and distributed mine maintenance costs. The distributed mine

maintenance costs include items such as maintenance labour, repair parts, and energy (fuel or electricity)

which contribute to the operating cost of the equipment and are distributed as an hourly operating cost

which is applied to the scheduled equipment operating hours.

The mine maintenance area accounts for the overheads of supervision, planning, and implementation of

all activities within the mine maintenance function. Costs collected for this area include salaried

personnel (supervisors, technical planners, and clerical), operating supplies for the various services

provided by this area and general shop costs. The costs of these items are not included in the distributed

mine maintenance costs.

The GME area accounts for the supervision, safety, and training of all personnel required for the direct

mining activities as well as technical support from mine engineering, environmental and geology

functions. Costs collected for this area include the salaries of personnel and operating supplies for the

various services provided by this function.

In this study, direct mining and mine maintenance are planned as an owner operated fleet with the

equipment ownership and labour being directly under operations. It may be possible to contract out some

of the direct mining activities under typical mine stripping contracts and maintenance and repair contracts

(MARC) as has been done at other operations. The viability and cost effectiveness of contracting can be

determined in future detailed planning and commercial negotiations. The exception for this study

involves blasting, where similar to other northern British Columbian mining operations, the mine

employs the blasting crew(s) but, due to the specialty expertise required, the supply and onsite

manufacturing of blasting materials is assumed to be contracted out. Similar to other operations,

infrastructure required for blasting will be provided by the operations.



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Figure 16-21 General Organization Chart

Details of the mine operation organization will be updated in future studies.

16.8.1 Direct Mining Area

The direct mining area accounts for the drilling, blasting, loading, hauling, and pit maintenance activities

in the mine.

In situ rock requires drilling and blasting to create suitable fragmentation for efficient loading and hauling

of both mineralized and mine rock material. Mill feed material and mine-rock limits are defined in the

blasted muck pile through blasthole assays and grade control technicians. A fleet management system

assists in optimizing deployment and utilization of the loading and haulage fleet to meet the production

plan. Support personnel and equipment are required to maintain the mining area, ensuring the operation

runs safely and efficiently. General descriptions of the direct mining unit operations follow.

Drilling

Areas are prepared on the bench floor blast patterns in the in situ rock. The spacing and burden between

blastholes is estimated to be 5m. The blasthole drills are fitted with GPS navigation and drill control

systems to optimize drilling. The GPS navigation enables stake-less drilling and is recommended for

efficiency in aligning with hole collar locations and accuracy of set-up. The drills are also fitted with

automatic samplers to provide grade control samples from the drill cutting in the mineralized material

zones. These samples are used for blasthole kriging to define the mineralized material/mine rock

boundaries on the bench as well as stockpile grade bins for the grade control system to the mill OCS.

Diesel hydraulic drills (150mm bit size) are used for production drilling; both in mill feed material and

mine rock.

Blasting

Powder Factor

A powder factor of 0.25 kg/t is assumed to provide adequate fragmentation and digging conditions for the

shovels. A detailed blasting study should be conducted in the next study phase.



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Explosives

A contracted explosives supplier provides the blasting materials and technology for the mine. Because of

the remote nature of the operation, an explosives plant is built on site. The nature of the business

relationship between the explosives supplier and the mining operator will determine who is responsible

for obtaining the various manufacture, storage and transportation permits, as well as any necessary

licenses for blasting operations. This will be established during commercial negotiations. For this study,

the costs are derived from the assumption that the explosives contractor delivers the prescribed

explosives to the blastholes and supplies all blasting accessories. Costs are reported on a per kilogram

basis for explosives.

Until the extent of groundwater and surface water in the blastholes is determined, it is assumed that the

holes will use a 70/30 emulsion/ammonium nitrate-fuel oil (ANFO) mix explosive (“wet” product).

Blasting accessories for initiation and detonation of the explosives in the blast holes are stored in

magazines in an appropriate location as described below.

Specifications for blasting plant and explosives storage magazines and the locations of these facilities

must adhere to the Explosives Act of Canada, regulations as published by the Explosives Regulatory

Division of Natural Resources Canada, and regulations as published by the MEMPR in BC, in particular

the Health, Safety and Reclamation Codes for Mines in BC. The explosives manufacturing plant and the

magazines are located east of the RSF area, as determined by the table of distances that govern the

manufacturing and storage of explosives and blasting agents.

Explosives Loading

Loading of the explosives is done with bulk explosives loading trucks provided by the explosives

supplier. The trucks should be equipped with GPS guidance and be able to receive automatic loading

instructions for each hole from the engineering office. This practice is common now in Canada and the

explosives supplier’s trucks have this capability already installed. The GPS guidance is a necessity for

compatibility with stake-less drilling. The explosives product used is a mix of ANFO and emulsion.

The holes are also stemmed to avoid fly-rock and excessive air blasts. Crushed rock is provided for

stemming material and is dumped adjacent to the blast pattern. A loader with a side dump bucket is

included in the mine fleet to tram and dump the crush into the hole. The crushed rock is provided by the

on-site rock crusher specified for mine roads.

From time to time during inclement weather (i.e. high snowfall, rainfall, etc.), it will be necessary to

square-off the pattern, tie in the loaded holes, and blast before the snow accumulation gets too high or the

explosives product gets wet. In these instances, it is necessary to square-off the pattern by loading

specific holes to complete certain rows in the pattern.

Blasting Operations

The blasting crew comprises day-shift-only mine employees. Based on existing mines of similar size and

previous experience, the estimated crew size is two people. The blasting crew coordinates the drilling

and blasting activities to ensure a minimum two weeks of broken material inventory is maintained for

each shovel. Also, the blast patterns are not staked and therefore the blasting activities must also have

GPS control. The blasters require handheld GPS units to identify the holes for the pattern tie-in. A



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detonation system is used consisting of electric cap initiation, detonating cord, surface delay connectors,

non-electric single-delay caps and boosters.

The explosives contractor supplies and manufactures bulk explosives on site. The explosives contractor’s

employees deliver explosives to the blasthole using a digital controlled "Smart" truck as is common in

Canadian surface mines.

Based on the desired powder factor, the blasting specifications for the Kwanika operations have been

evaluated for a small diameter hole size. The blasting assumptions are summarized in Table 16-8. It is

assumed that all rock requires drilling and blasting. These parameters will be re-evaluated in the future

with a detailed blasting study, using site specific rock strength parameters.

Table 16-8 Blasting Assumptions

Blasting Pattern –

Mineralized Material & Mine Rock

Specifications

Spacing 5m

Burden 5m

Hole Size 5.9″

150mm

Explosive In-Hole Density 1.25 g/cc

Explosive Average Downhole Loading 22.1 kg/m

Bench Height 10m

Collar 3.5m

Loaded Column 7.5m

Sub-drill 1m

Charge per Hole 166kg/hole

Rock SG 2.70t/m3

Yield per Hole 675t/hole

Powder factor 0.25kg/t

Loading

Mill feed material and mine rock boundaries are defined in the blasted muck pile using the OCS.

The design basis assumes one loader type is purchased in order to simplify the maintenance function and

reduce capital equipment and maintenance spares. This applies to other mine fleet equipment as well.

Three 15.3m3 diesel hydraulic shovels have been selected as the primary digging units. All mining

equipment has therefore been selected as diesel powered. The excavators are selected to mine mill feed

material in triple split bench areas while the shovels mine full bench mine rock.

There are years when there is a component of mill feed material being reclaimed from the stockpile to

feed the mill. In these years, it is intended to relocate the necessary loading equipment to the stockpile

area for the required length of time.

Bench widths are designed to ensure maximum operating widths to enable double-sided loading of trucks

by the shovels. In some areas (such as pit bottoms or where ramps run across narrow pushback sections),



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single-side loading is necessary and the productivity for the shovels is reduced. For this study, it is

assumed that this represents a small percentage of the total material mined.

Optimization of the shovel fleet will be conducted in future studies.

Hauling

Mill feed material and mine rock haulage is handled by large haul trucks with a 136t payload. Haulage

profiles have been estimated from pit centroids at each bench to designated dumping points for each time

period. These haul profiles are inputs to the truck simulation program and the resulting cycle times are

used in MS-SP, which is set to maximize project NPV by using the shortest haul to a feasible destination.

The payload, loading time, and haul cycle then determine the truck productivity.

Pit Maintenance

Pit maintenance services include haul road maintenance, mine dewatering, transporting operating

supplies, relocating equipment, and snow removal.

A rock crusher for road grading material is included to improve truck travel speeds, reduce mechanical

fatigue to the haul trucks, increase traction (improving safety), and to enhance tire life, which is a major

mine operating cost.

During winter, the snow fleet is operated by mine operations staff that will not be required for activities

such as dust control and summer field programs. During severe winter storms, additional crew members

are drawn from truck and shovel operations to operate the snow fleet. This ensures any fleets deemed to

be priority will remain operating.

16.8.2 Mine Maintenance Area

The mine maintenance area accounts for the supervision and planning of the mine maintenance activities.

Mine maintenance activities are directed under the mine general foreman who assumes overall

responsibility for mine maintenance and reports to the mine superintendent (in an alternate organization,

this position may be filled at a superintendent-level reporting to the general manager). Maintenance

planners coordinate planned maintenance schedules. The daily maintenance shift coordination is carried

out by mechanical and electrical foremen.

The mine maintenance department performs break-down field maintenance and repairs, regular

preventive maintenance, component change-outs, in-field fuel/lube servicing, and tire change-outs. Due

to the remoteness of the site, some larger critical equipment components will need to be kept in onsite

inventory, and broken down components will be sent out during the next shipping season.

16.8.3 General Mine Expense Area

This section describes the mine GME as estimated in the mine cost model in Appendix C.

The GME area accounts for the supervision, and training for the direct mining activities as well as

technical support from mine engineering and geology functions. Mine operation supervision extends

down to the shift foreman level.



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A mine general foreman is responsible for overall supervision for the mining operation, including overall

open pit supervision and equipment coordination. Supervision is also required for drilling and blasting,

training, and dewatering. A mine shift foreman is required on each 12-hour shift, with overall

responsibility for the shift operation. Security/first-aid staff and mine clerks report to the mine

superintendent.

Initial training and equipment operation are provided by experienced operators. As performance reaches

adequate levels, the number of trainers is decreased to a sustaining level.

A chief mine engineer directs the mine engineering department. The senior mining engineer coordinates

the mining engineers, mine planning groups, surveyors, and geotechnical monitoring personnel.

Surveyors assume responsibility for surveying for the entire property. Surveying uses GPS-based

systems.

The geology department includes pit geologists, and grade control technicians. This department is

responsible for local step-out and infill drill programs for onsite exploration activities and updates the

long range mine mineralized material models. The geology department also provides grade control

support to mine operations, manages and executes the blasthole sampling and blasthole kriging of the

short range blasthole models for operations planning and mill feed material grade definition.

Geotechnical considerations are covered by the mine engineering department.

The Environmental department is normally comprised of a department head and environmental

professionals including biologists, environmental engineers and technicians. Major responsibilities

include regulatory administration, environmental permitting and monitoring, and operational waste

management and reclamation.

16.9 Mine Closure and Reclamation

At the end of the mine life, an approved closure and reclamation plan is implemented that meets the

approved end land use objectives and satisfies regulatory commitments. The mining costs provide an

allowance for general reclamation activities. Details on mine closure and reclamation are available in

Section 18.11.

Open Pit Reclamation



potential as quickly as possible.

RSF Reclamation

The RSFs are constructed at the natural angle of repose of 37° for the dump lifts, and with benches the

overall slope is configured to the final designed slope angle. This reduces the cost of RSF reclamation at

the end of the mine life when RSFs are re-sloped to a closure slope of 26°. Re-sloped RSFs are designed

to blend with naturally occurring topographic features. At the end of the mine life, topsoil from the

topsoil stockpile is placed on the RSF where possible. The surfaces can then be planted/seeded as

required required to control erosion and meet end land use objectives.



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TSF Reclamation

The flotation tailings will progressively capped and the outer slopes of the tailings storage dam will be re-

sloped to blend with the natural landscape and to provide access for wildlife. Natural seepage water

collected within the water storage pond will be pumped and discharged to the open pit until the water

quality meets discharge criteria.

Mine Roads and Dykes

At the end of the mine life, decommissioned mine roads are scarified and capped with available surficial

soils. Dykes and dams that are exposed above the water line will also be scarified and capped with

suitable soils. The surfaces can then be planted/seeded as required.

Creek Diversions

The Kwanika Creek diversion channels will either continue to operate or will be decommissioned, as per

the conditions required in the environmental authorizations.

16.10 Mine Equipment

The mining equipment descriptions in this section provide general specifications so that dimensions and

capacities can be determined from the manufactures specification documents.

Major Mine Equipment

The production requirements for the major mining equipment over the LOM are summarized in Table

16-9. According to the current production schedule and the haulage assumptions, a maximum of 13

trucks are required over the LOM. All other equipment requirements are detailed in Appendix C.

Table 16-9 Major Mine Equipment Requirements

PP Y5 Y10 Y13 Max

Drilling

Primary Drill – 150mm Diesel Hydraulic Drill 2 3 3 1 3

Loading

Primary Shovel –15.3 m3 Diesel Hydraulic Shovel 2 3 3 1 3

Hauling

Haul Truck – 136t 6 13 5 5 13

The haul truck fleet size schedule is shown in Figure 16-22.



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0

2

4

6

8

10

12

14

Nu

mb

er

Of

Tru

cks

Period

Haul Truck Fleet Size

Figure 16-22 Haul Truck Fleet Size

Ancillary and Support Equipment

A suitable complement of ancillary and support equipment has been included in the cost model. Details

can be found in Appendix C.

16.10.1 Drilling Equipment

The primary production drilling is done in both mineralized material and mine rock with diesel hydraulic

rotary drills using a 150mm bit size. The production drills are fitted with GPS navigation and drill

control systems to optimize drilling. Production drilling assumptions are listed in Table 16-10.

Table 16-10 Production Drilling Assumptions

Production Drill – Mineralized

Material & Mine Rock

Diesel Hydraulic

Rotary

Bench Height 10m

Subgrade 1.0m

Hole Size 150mm

Penetration Rate 30.0m/h

Over Drill 0.5m

Hole Depth 11.5m

Setup Time 4.0 min

Drill Time 23.0 min

Move Time 4.0 min

Total Cycle Time 31.0 min

Holes per Hour 1.94

Re-drills 10%



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Development drilling requirements have not been detailed in this study, but it is assumed to be minimal

and an allowance has been made for costing purposes.

A detailed drill study is recommended for more advanced project studies. This will help determine the

penetration rate that can be expected for the selected drills and the specific rock types that exist within the

pit area.

16.10.2 Blasting Equipment Facilities

Blasting activities are detailed in Section 16.8.1.

A blasthole stemming unit will be required to load cuttings into the hole and stem the unloaded portion of

the hole. This unit will be provided by the Kwanika operation.

The selected explosive plant location is designed at the required clearance distance from infrastructure

and working areas.

16.10.3 Loading and Hauling Equipment

The shovel-truck fleet selected for Kwanika is the 15.3m3 class of diesel hydraulic shovel, and the 136t

payload class of truck.

16.10.4 Dewatering Equipment

It is important to control water that is in active mining areas. In-pit water generally increases the mining

cost especially blasting, where explosives loading, explosives costs (ANFO vs. emulsions), and blast

performance is affected by water. Flooded pit working areas need to be drained, and rock cuts to tires are

increased in wet conditions. The presence of water in the shovel digging area greatly decreases the

average tire life of the trucks. Rocks are easily hidden in puddles that the haul trucks have to drive

through and this can lead to instantaneous tire failure. Wet muck that the shovels dig freezes to the sides

of the truck boxes in the wintertime and this “carry back” results in less material being hauled per truck

load (i.e. lower productivities). Water also affects the stability of walls and dumps. All of these effects

need to be addressed by an effective pit dewatering program.

No hydrology work has been performed to this point. At this stage of planning, an allowance for pit

dewatering activities will include the following:

sloped pit floors as required

in-pit sumps

water collection system(s)

Pit water will be collected and treated prior to discharge to the environment.



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16.10.5 Mine Ancillary Facilities

Shops and Offices

In addition to providing an area for maintenance bays, tire shops, and a wash bay, the maintenance shop

will also house:

a welding bay

an electrical shop

an ambulance

a first aid room

a first aid office

a machine shop area

a mine dry

a warehouse

offices for administration, mine supervision, and engineering/geology staff

a lunch room and foreman’s office

The recommended final maintenance facilities have service bays, a welding bay, and wash bays. The

mine maintenance facility also includes a machine shop area, tool storage area, warehouse, and office

complex. A separate tire bay facility is required with an exterior heated pad to accommodate at least two

trucks and a tire manipulator; the pad area should be 30m x 30m. A layout of the truck shop has yet to be

detailed.



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17.0 Recovery Methods A conventional copper-gold flotation process is assumed for the Kwanika project including crushing,

grinding, and multi-stage froth flotation to produce a copper-molybdenum concentrate. There will be

gold and silver credits. Future work will need to evaluate the economics of producing a separate,

molybdenum concentrate and operating a separate a gold and silver circuit.

The process is supported by test work discussed in Section 13.



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18.0 Project Infrastructure The Kwanika Property in north central British Columbia, is situated in the Omineca Mining Division,



round by four-wheel-drive vehicle; provided there is active snow removal in winter.

18.1 General Site Geotechnical Investigation

To date, there has not been a general site geotechnical investigation. Pre-feasibility level geotechnical

and hydrology studies will be needed, in order to optimize the locations of site infrastructure.

18.2 Site Layout

The Project site will have open pit and underground mining-related facilities, process-related facilities, and

a permanent camp. The site layout is compact in order to reduce the impact of disturbance and to

increase efficiency. The site will require a small amount of backfill. The site plan will be further

optimized in future studies.

The general site layout is shown in Figure 18-1.



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Figure 18-1 General Site Layout



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18.3 Site Roads

18.3.1 Access Road

Currently there is an existing forest service road between Fort St. James and the Tsayta Lake Road

providing surface access to the site. For the purposes of this study, it is assumed that the ~30km long

Tsayta Lake Road will be upgraded to meet the needs of the operation.

18.3.2 On Site Transport Road

On-site service roads will be constructed connecting to the Project access road, the explosives storage and

manufacturing facilities, tailings storage facility, processing plant, rock storage facility, and open pits.

The haul roads between the pit, the primary crusher at the plant site, the RSF, and the TSF will be

constructed with a top course of crushed mine rock.

18.4 Process Plant and Process Related Facilities

The plant site is located south of the RSF.

The plant site pad will be gravel surfaced. The site will be graded so as to control water flow and capture

all contact water before treatment or disposal.

The main process-related facilities for the Project will include:

Primary crushing facility

Main process plant including; primary grinding, flotation, regrinding, and water treatment

Mine rock storage facility

Tailings storage facility

18.5 Ancillary Buildings

The main ancillary facilities for the Project will be located adjacent to the plant site and will include:

Mine maintenance facilities

Truck shop; including warehouse

Permanent camp; including general offices

Explosive storage facilities

Sewage treatment facility

Potable water system

Electrical sub-station

18.6 Communications System

A satellite-based system will be needed for external voice and data communications services. An on-site

network will be established that will connect buildings and radio transceivers will be used for remote

monitoring and control. An ultra-high frequency (UHF) radio system will be used for mobile

communication.

18.7 Power Supply

Electrical power that will be supplied to the Project includes extending a connection from the Kemess

Power Line 75km to the Kwanika Project site.



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18.8 Power Distribution

18.8.1 Plant Site

Power will be distributed in the plant at 13.8kV. More work is required to determine details of the power

distribution system.

18.8.2 Remote Loads

Remote loads will be fed via overhead 13.8kV lines. These include:

primary crushing

tailings/water management

explosive manufacturing

permanent camp

truck shop

pit

18.8.3 Ancillary Systems

Ancillary systems to be provided under electrical include:

emergency power

general power and lighting (indoor and outdoor)

electrical heating

heat trace

fire alarm

communications

closed circuit TV

18.9 Mine Rock, Tailing, and Water Management

This section describes the projected LOM tailings, mine rock, and water management requirements for

the Project.

The overall topography of the property area is mountainous with an elevation range between 950 and

1415masl.

18.9.1 Solids Management

Solids management essentially comprises the containment and long term management of the waste

products derived from the mining and milling process (e.g. mine rock and tailings). The tailings stream

will be pumped to the Tailing Storage Facility (TSF), located directly west of the Plant Site.

Information Supplied and Design Basis

According to the mining schedule, the tailings production will be approximately 7,100kt/a. In general,

the mine rock production ranges between 2,000 and 17,000kt/a.

A summary of the key parameters used in this study for the design of the tailings management are listed as

follows:



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Tailings Impoundment

The following parameters have been used for the design of the tailings impoundment:

Swell factor used to calculate tailings tonnage = 30%

Total LOM volume of tailings to be accommodated by the tailings impoundment: approximately

35Mm3.

The tailings will not require any special chemical suppression or management measures.

Mine Rock Production

Approximately 12,000,000t (approximately 4,450,000m3) of mine rock will be produced during pre-

production. Mine rock produced during this period will be utilized for civil earthworks construction (e.g.

tailings starter dams and water diversion structures) and the foundation for the Rock Storage Facility

(RSF).

During mining operations, mine rock production will range between 2,000,000t/a and 17,000,000t/a,

averaging 11,750,000t/a. The following parameters have been used for designing the mine rock

containment facility:

direct haul from mine pit to RSF

SG of mine rock: 2.7

average overall side slopes of mine rock storage facility during mining operations: 26°

maximum mine rock bench height: 30-60m

swell factor of 20%

total LOM volume of mine rock to be accommodated by the mine rock pile: approximately

70Mm3

It is assumed that the majority of mine rock will be non-acid generating (NAG); the remaining small

portion of potentially acid generating (PAG) mine rock, if present, will be placed inside of the RSF where

water infiltration can be limited. The PAG mine rock could, if required, be surrounded with till

overburden fill (or flotation tailings) to limit the top infiltration and potential leachate generation. .

Detailed studies on the ARD/ML characteristics of the mine rock will need to be conducted to support a

final mine rock disposal plan.

Tailing and Mine Rock Management

Some of the considerations that helped develop the proposed tailings storage facility design and location

were:

Due to the large mass of mine rock, the mine rock storage facility should be located as close as

possible to the pit. The tailings/residue storage locations are less cost sensitive to distance as

these materials are hydraulically transported by pipeline to their point of deposition, but it is still

advantageous to have them stored relatively close to the plant.

Items that need to be considered in future studies are:

Surface water from snow melt and rainfall events as well as bleed water from deposited

tailings/residue has to be effectively captured and managed across the disturbed areas of the mine

site. This requires an understanding of the local catchments, topography, climatic conditions,

and near surface geotechnical conditions.

A prefeasibility level geotechnical investigation of pertinent structures and surficial material

types should be undertaken to confirm the locations of the TSF and RSF are optimal.



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Chemical characterization of the mine rock and tailings/residue. Physical characteristics of mine

rock and tailings/residue, including the stability of the RSF and TSF under both static and

seismic loading conditions.

Ongoing design work will need to mitigate potential environmental impacts and use current accepted

standards as a means of reducing risk.

18.9.2 Water Management

Water management describes and compiles processes that enable the effective management of water from

the disturbed mining areas in order to reduce the risk of uncontrolled discharge into the surrounding

environment. This requires the review and interpretation of existing climatic data to estimate the

rainfall, snowfall, evaporation, and sublimation expected at the site. These parameters are used to

develop a water balance flow sheet that takes into account all water entering, being contained within, or

leaving the area of mine disturbance within a defined time period. This flow sheet feeds into a LOM

water balance, which estimates on a monthly basis the volume and capacity requirements for containment

facilities, pumps, pipelines, and other supporting infrastructure, as well as any operational requirements.

To date no significant water management work has been completed for the Kwanika Project. In future

studies these water management items should be considered:

Preliminary estimated annual pit inflow quantities and quality to the proposed open pit should be

performed once the open pit and underground stopes have been optimized during the next level

of study,

An estimate of the fresh water makeup for the process plant,

A site water schematic water balance flow sheet,

A life of mine site water schematic water balance flow sheet. This model will be used to size

civil and mechanical infrastructure to support the water requirements of the Project, and

Background hydrological and hydro-geological studies in order to develop an overall water

management plan. This will also include details for the proposed diversion of Kwanika Creek.

18.9.3 Kwanika Creek Diversion

This section summarizes Knight Piésold’s conceptual design for a Diversion Channel for Kwanika Creek.

The length of diversion channel required and the costs have been re-calculated by MMTS based on

updated cost assumptions and the revised mine plan.

Kwanika Creek is a fish-bearing stream that runs through the Kwanika leases. The proposed TSF and South

Pit locations intersect the existing route of Kwanika Creek and, as a result, requires two diversions to

maintain hydraulic and ecological continuity between these two lakes until the mine is closed. The proposed

Kwanika Creek Division Channels are shown in Figure 18-1. The location of the diversion channel was

chosen to maintain flow along as much of the existing Kwanika Creek alignment as possible and maintain

the hydraulic boundaries of the area as a whole.

The consequence of exceeding the capacity of the diversion system would be the flooding of the South

Zone Pit and the subsequent dewatering efforts in the pit. As this consequence is considered to be severe,

the diversion system was sized with consideration of the 1 in 10,000 year flood, which has a very low

likelihood of occurring over the life of the mine. Upon further study it may be concluded that a less

severe event may be appropriate for the design, however the difference between a 500 year event and a



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10,000 year event is comparatively small, so the effect on the size and cost of the diversion would be

relatively minor, and is thus considered to be a prudent planning criteria.

Conceptual Diversion Designs

To divert Kwanika Creek around the tailings storage facility and the South Zone Pit, a diversion channel

is required along the north side of the tailings containment dam and the east side of the pit on the valley

floor (see Figure 18-2). For this design, it is assumed that the channel has a trapezoidal shape and an

average gradient of 0.5% and will be lined with sufficient rip-rap. Floods up to and including the 1 in

10,000 year event could be successfully diverted with north and east channel dimensions of 2m depth and

80m width. This design provides 0.5m of freeboard.

MMTS estimates the diversion would be comprised of approximately 1600m of valley bottom channel

(800m for the South diversion, 800m for the North diversion). The cost for the diversion is estimated to

be in the order of $3 to $4 million. Measures to control seepage from both sections of the channel may

be required because of stability considerations for the pit walls. These measures could include

bentonite lining of the valley bottom channel and shotcrete or slush grouting of the rock excavated

channel.

Figure 18-2 Proposed Kwanika Creek Diversion Channel – Valley Bottom Diversion

18.9.4 Pre-Production Earthworks for Mine Rock and Water Management

Prior to mining and processing operations, it will be necessary to prepare for and construct the major

structures listed below:

Upgrade access road,

Construct power line,

Kwanika Creek Diversion Channel – North diversion,

Water management ditches,

Starter dams (see details below),

Plant construction,

Other infrastructure construction,

Pre-Stripping open pit mining area, and

Start underground development.



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Water Retention Dam

The construction of a water retention dam is required during the pre-production construction stage to

collect surface runoff water from the Rock Storage Facility (RSF). This will later be used to collect

seepage water from the flotation tailings during the mine operation and the early stage after closure until

the water quality meets discharge criteria.

It is planned that the construction fill materials for these dams will be mainly sourced from the pre-

production open pit excavation and diversion channel excavations.

Containment Berm for Tailings Management

Prior to process plant operation, it will be necessary to construct a starter containment berm to an elevation

of 1045m to act as an initial containment for the flotation tailings. The proposed location and extent of

the containment berm are shown in Figure 18-1.

Pre-production mine rock will form the primary containment structure and will be internally lined with

lower permeability material.

18.9.5 Post-Production Earthworks for Mine Rock and Water Management

At the cessation of mining operations an approved closure plan will be implemented to return the

operating area to a condition that will meet the end land use objectives.

The flotation tailings will be progressively capped and the outer slopes of the RSF will be re-sloped to

blend with the natural landscape and to enable access for wildlife. Natural seepage water collected within

the water storage pond will be pumped and discharged to the open pit until the water quality meets

discharge criteria.



potential as quickly as possible. The Kwanika Creek diversion channels will either continue to operate or

will be decommissioned, as required by approval conditions.

This plan will require more detailed work based on the findings and recommendations from the ongoing

studies being conducted for the Project.



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19.0 Market Studies and Contracts There have been no market studies conducted and no contracts reached between Serengeti Resources Inc.

and refiners at the time of this PEA.

Table 19-1 shows the metal prices and resultant Net Smelter Prices for the base case in the financial

analysis using the three-year trailing average (as of October 15, 2012). The exchange rate used is 0.95

US$/CDN$.




Gold US$1,427.00/oz $44.57/gm




Concentrates will be sold into the general market. This will either be to North American, European, or

Asian smelters and refineries.



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20.0 Environmental Studies, Permitting and Social or Community

Impact

20.1 Regulatory Framework

Serengeti’s Kwanika Project falls within the category of a “reviewable project” of the British Columbia

Environmental Assessment Act (BCEAA), administered by the BC Environmental Assessment Office

(BCEAO), and will also likely trigger the Canadian Environmental Assessment Act (CEAA).

The project is deemed to have a production capacity of 5.48Mtonnes/yr.; well over the threshold of

75,000 tonnes/year which triggers BCEA requirement. A proposed diversion of Kwanika Creek, a fish-

bearing stream, will require a Federal Fisheries Act approval, and will likely trigger the requirements of

the CEAA. Other requirements of Provincial and Federal Acts and Regulations may also apply,

depending upon final design components. The cost of a CEAA process could add $2.0M or more to

costs, as well as at least one to two years of approval time. Attempts will be made in future studies to

parallel ongoing engineering work with the approval process to have less impact on the project schedule.

Additional costs of monitoring, mitigation and decommissioning may also be required.

The current mine plan only requires limited sections of steam diversions. Future work should be diligent

in considering mine plan alternatives and details that do not require federal approvals triggering the full

CEAA process.

20.2 Regional Land Use Processes

The Project is located within lands that have been dedicated in the Fort St. James Land and Resource

Management Plan, approved by government in 1999. The Project area is within the Multi-Value

Resource Management Zone Land Use designation, where lands are managed to integrate a wide range of

resource values, including mining.

20.3 Environmental and Corporate Social Responsibility Programs Already in Progress

Project related work completed to date on the environmental and corporate Social Responsibility/First

Nations/Community Impact aspects of the Kwanika project is described by Serengeti below. This work

has been undertaken mainly in support of the past exploration programs and will positively contribute to

ongoing background studies and relationships with the other stakeholders in the area.

A project specific Valued Ecosystem Component (VEC) study was completed in collaboration with the

Takla Lake First Nation (TLFN) in 2008. This study identified the principal issues that should be

addressed in further environmental assessment studies on the project.

A baseline water quality report was conducted over a ten month period in 2008. Annually, since that

time selected drainages have been sampled and analyzed pre and post drilling activity on the property.

A project area, archaeological overview assessment (AOA) was completed in 2008. Since that time,

proposed drilling plans are reviewed annually with TLFN within whose traditional territory the project

lays. Proposed drilling sites are evaluated via Preliminary Field Reconnaissance (PFR) conducted by a



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registered archaeologist, commonly accompanied by a TLFN elder to evaluate and minimize impact on

archaeological values.

An Exploration Agreement covering activities up to and including mine construction on the project was

signed with TLFN in 2010. This agreement provided for communication protocols, project related

employment and training opportunities, community capacity building and community improvements,

principally around education, as well as providing for TLFN participation in any project environmental

assessment process. Since that time, annual project impact mitigation strategies are implemented

collaboratively with TLFN including the employment of a locally hired environmental monitor and

monitoring program.

In terms of employment and contracting opportunities, local First Nations members have generally

composed greater than 50% of the onsite project employment and significant contracting opportunities

have been provided since 2007.

To date, no formal permitting activities have been initiated on the project other than through the Ministry

of Energy, Mines and Natural Gas, Notice of Work (NOW) process where a multiyear permit is in place

covering exploration activities.

20.4 Fisheries Resources and Permitting Issues

British Columbia government fisheries inventory data indicate that the following fish species are present

in Kwanika Creek:

Rainbow trout,

Dolly Varden char,

Burbot,

Mountain whitefish, and

Peamouth chub.

There are no fish obstructions listed in Kwanika Creek and therefore migratory fish species residing in

the upper Nation River system including Tsayata Lake could potentially use Kwanika Creek for

spawning.

Re-routing Kwanika Creek through a channel in order to develop an open pit in the Kwanika Valley

would require a Section 35(2) approval from Fisheries and Oceans Canada (DFO) under the federal

Fisheries authorizations required for both diversion of the watercourse and development of the open pit.

In order for DFO to issue a Fisheries Act approval the proponent would need to demonstrate the

following:

That the impact is reasonably unavoidable

That the level of impact poses an acceptable risk to fish in consideration of DFO risk assessment

criteria, and

That the impact can be compensated for and achieve the no net loss policy of DFO.

The project has the potential to compromise a large amount of fish habitat. Providing a diversion channel

around the open pit would enable migratory fish access to upper Kwanika Creek and its tributaries,



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reducing the potential impact to fish. It is expected that DFO will review this stream diversion carefully

and will likely require special provisions to mitigate potential impacts to fisheries resources.

At this scoping study level, the order of magnitude cost of the replacement of the lost fish habitat in

Kwanika Creek through the construction of fish compensation measures in the diversion channel is

estimated to be $3 Million. Detailed studies are recommended to fully identify the level of impact to fish

and fish habitat, fish habitat compensation options, and potential fisheries related permitting risks

associated with the Project.



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21.0 Capital and Operating Cost Estimates

21.1 Capital Cost Estimate

All currencies in this section are expressed in Canadian dollars. Costs in this report have been converted

using a fixed currency exchange rate of US$0.95 to CDN$1.00. The expected accuracy range of the

capital cost estimate is +/- 40%.

An initial capital of $363.6M is estimated for the Project.

This PEA estimate is prepared with a base date of Q4 2012 and does not include any escalation past this

date.

Initial capital has been designated as all capital expenditures required prior to mill start-up for producing

copper concentrates for shipment to contract smelters. Sustaining Capital includes replacement

equipment purchases and continued underground development. A summary of the major capital costs is

shown in Table 21-1 and Table 21-2.

Table 21-1 Initial Capital Cost Summary


($000)

Open Pit Mining – Equipment 56,920

Open Pit Mining – Pre-Production 31,575



Processing Plant 129,000

Site Infrastructure 41,170

Access & Power 24,900

Contingency 58,520

Total Initial Capital Cost 363,565

Table 21-2 Sustaining Capital Cost Summary


($000)

Open Pit Mining – Sustaining 20,790



Reclamation 1,810

Total Sustaining Capital Cost 143,870

The detailed breakdown of this capital cost estimate is included in Appendix C.

21.1.1 Exclusions

Owner’s costs for past exploration and studies and future costs for future studies have not been

included as well as certain royalties, taxes and interests. At this stage of study the future financial



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capacity of equity investors or partners are unknown. These details and options will need to be

assessed in future more advanced studies.

21.1.2 Open Pit Mine Capital Costs

Open Pit mine capital costs are a factored estimate using information from other projects and adjusted to

this case. The open pit mine equipment capital costs include costs of delivery and assembly. The open pit

mine initial capital costs are shown in Table 21-3.

Table 21-3 Open Pit Mine Initial Capital Costs

CDN$ (000)

Pre-production (Operating costs during Capital period) 31,575

Mobile Equipment 53,219

Communication 200

Safety 1,000

Other Mining Costs 2,500

Total Open Pit Mine Initial Capital 88,494

21.1.3 Underground Mine Capital Costs

Underground mine capital costs are a factored estimate using information from other projects and

adjusted to this case.The underground mine initial capital costs are shown in Table 21-4.

Table 21-4 Underground Mine Initial Capital Costs

CDN$ (000)

Mobile Equipment 10,080

Development Costs 11,400

Total UG Mine Initial Capital 21,480

21.1.4 Mining Basis of Estimate

The MMTS estimate of start-up capital costs includes the following:

mine mobile equipment fleet

support and auxiliary equipment

fleet and drill monitoring and management software and systems

Open Pit mine mobile equipment capital costs are shown in Table 21-5. The mobile equipment capital

schedule includes all equipment purchases to the end of Year 1. See Appendix C for details.

Table 21-5 Open Pit Mine Mobile Equipment Initial Capital Schedule

Fleet Capital Cost Capital (CDN$ M)

Diesel Drill –150mm 3.3

Diesel Hydraulic Shovel – 15m3

10.6

Haul Truck – 136t 22.7

Other 16.6

Total Open Pit Initial Capital Cost $53.2M



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Underground mine mobile equipment capital costs are shown in Table 21-6. The mobile equipment

capital schedule includes all equipment purchases to the end of Year 1. See Appendix C for details.

Table 21-6 Underground Mine Mobile Equipment Initial Capital Schedule

Fleet Capital Cost Capital (CDN$ M)

Drill Jumbo 1.0

Haul Truck 2.8

LHD 2.3

Other 4.0

Total Underground Initial Capital Cost $10.1M

21.1.5 Assumptions

The Capital estimate is based on the Base Case as described and includes scoping level assumptions.

Future engineering and cost studies are required to quantify the geotechnical, hydrological, and

environmental conditions for the site which can significantly impact the design requirements for the

project and subsequently the capital costs.

21.1.6 Contingency

Contingency allowances have been built-in to cover additional costs, which will be incurred as a result of

more detailed design and investigations for the different aspects of the project. An additional

contingency of approximately 20% ($58.52 Million) has also been added to cover changes not anticipated

in this scoping level study.

21.2 Operating Cost Estimate

The operating costs for the Project, as shown in

Table 21-7, are estimated at an overall cost of $21.20/t of mill feed. The estimate is based on an average

annual process rate of 5,475,000t milled at an average grade of 0.38% copper, 0.295g/t gold, 1.378g/t

silver, and 0.016% molybdenum (molybdenum is only present in the South Zone) including mine

dilution.

The cost estimates in this section are based upon budget prices in Q4 2012 or based on the information from the

database of MMTS. When required, costs in this report have been converted using an average currency

exchange rate of US$0.95 to CDN$1.00. All costs are reflected in 2012 Canadian dollars. The expected

accuracy range of the operating cost estimate is +/-35%.



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Table 21-7 Operating Cost Summary

$/tonne Mined $/tonne Milled

Mine (OP) $2.42

Mine (UG) $7.50*

Mine Total $9.14**

TSF $0.18

Mill $10.69

G & A $1.20

Tailing Treatment Included in G&A

Water Treatment Included in G&A

Total $21.20

*Note: UG cost/tonne mined based on total tonnes mined by UG only.

**Note: Total mining cost (OP + UG) divided by total tonnes milled.

The operating costs are defined as the direct operating costs including mining, processing, tailings

handling, water treatment, and G&A. The power is estimated to be $0.05/kWh. The power cost is based

on other studies in the general area of the Project.

The water treatment and tailings storage costs are included in the G&A costs.

21.2.1 Open Pit Mine Operating Costs

All Open Pit mining operating costs are shown in Canadian dollars. Open Pit mine operating costs are

derived from a combination of vendor estimates and historical data collected by MMTS. This includes the

labour, maintenance, major component repairs, fuel, and consumables costs. The 2012 fleet hourly

operating costs are used as a constant basis over the scheduled periods and the estimates are input for

sustaining and replacement capital.

Unit costs for consumable and labour rates are estimated from the sources listed below while the

magnitude of consumables and labour requirements are determined for each specific activity from

experience and first principles.

The unit costs are based on the following data:

Salaries for the supervisory and administrative job categories are based on MMTS’s experience

of similar functions in Canadian mines. An average burden rate of 20% was applied to base

salaries to include all statutory Canadian holidays, social insurance, medical and insurance costs,

pension, and vacation costs.

For hourly employees, general labour rates expected in Canadian mines are used. An average

burden rate of 25% is applied to base wages to include all statutory Canadian holidays, social

insurance, medical and insurance costs, pension, and vacation costs.

Unit costs are based on mine designs with optimized size and mine fleet makeup. Factors

considered include distance to pit and mine rock storage piles, maintenance facilities, and current

and future topography.

Unit costs freight for all consumables, tires, and fuel, are based on budgetary quotations. The

long term fuel price is estimated at a delivered cost to site of CDN$1.00/L.

All mine equipment is assumed to be diesel-hydraulic.



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Blasting costs are based on studies from similar projects and historical blasting costs.

Geotechnical costs for high wall control blasting, horizontal drains, etc. are based on other study

data collected by MMTS.

Mining equipment consumables, major equipment replacements, sustaining capital, labour

loading factors, equipment life, and costs are based on MMTS’s database from similar mining

operations.

LOM unit open pit operating costs are listed in Table 21-8. Complete mine cost tables, including mine

capital and operating cost schedules, are available in Appendix C.

Table 21-8 Open Pit Mining Costs per Tonne Material Mined LOM Cost

($/t Material Mined)

Drilling 0.22

Blasting 0.31

Loading 0.36

Hauling 1.01

Pit Maintenance 0.25

GME 0.25

Other Mining Costs 0.03

Total Mining Cost 2.42

21.2.2 Mine Fuel Consumption

Fuel consumption rates are estimated in the mine schedule for each equipment type. These consumption

rates are applied to the operating hours of the equipment to estimate the total fuel consumption. Fuel

costs have been included in the unit operating costs estimated above. Explosive factory fuel consumption

is estimated to be 44L diesel fuel per tonne of explosives based on the quantity of explosives used.

21.2.3 Underground Operating Costs

Underground mining operating costs have been developed by AMEC as $7.50/tonne (including rockmass

conditioning, delivery of mill feed to an on-site mill, assuming truck haulage and inclusive of

underground mine management and supervision but exclusive of capital development, processing and

administration). See Appendix A for more details.

21.2.4 Process Operating Costs

The operating cost estimate for Kwanika has been based on a benchmark with comparable industrial size

operations. The primary grinding P80 of 75µm and the concentrate regrinding P80 of 26µm

recommended by the testwork, along with the estimated power rate for Kwanika have been used to

estimate an operating cost of $10.70/tonne milled.



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21.2.5 General and Administrative

G&A costs are the costs that do not relate directly to the mining or processing operating costs. The costs

include:

personnel – general manager and staffing in accounting, purchasing, and environmental

departments, and other G&A departments

G&A expenses – insurance, administrative supplies, medical services, legal services, human

resources related expenses, travelling, accommodation/camp costs, air/bus crew transportation,

and external assay/testing.

The G&A cost is estimated at $1.20/t milled, including approximately $0.05/t milled for water treatment

and $0.05/t milled and is based on similar projects in the area.



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22.0 Economic Analysis

22.1 Introduction

An economic evaluation of the Project incorporating all the relevant capital, operating, working,

sustaining costs, and royalties has been performed. The evaluation is based on a pre-tax financial model.

For the 13.5 year mine life and 73.6Mt resources inventory, the following pre-tax financial parameters

have been calculated using the base case metal prices:

13.4% internal rate of return (IRR)

7.3-year payback on US$507.5M capital

US$263M net present value (NPV) at 5% discount value.

Table 22-1 shows the metal prices and resultant Net Smelter Prices for the base case in the financial

analysis using the three-year trailing average (as of October 15, 2012). The exchange rate used is 0.95

US$/CDN$.




Gold US$1,427.00/oz $44.57/gm




Sensitivity analyses are used to evaluate the project economics.

The detailed financial model is provided in Appendix D.

It should be noted that the data in the financial analysis incorporates engineering and cost estimates at a

scoping level of study which is not deem suitable for capital investment or production decisions. As well

Inferred Mineral Resources have been included in the production schedule and cash flow model. These

are considered too geologically speculative to have the economic considerations applied to them that

would enable them to be categorized as Mineral Reserves. Therefore, there can be no certainty that the

estimates contained in the PEA will be realized.

22.2 Pre-Tax Model

22.2.1 Metal Production Financial Model

The metal production values indicated in Table 22-2 are a summary of the results of the production

schedule, which is used in the cash flow to determine projected revenues. The pre-tax economic results in

Canadian dollars are listed in Table 22-3.



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Table 22-2 Metal Production from Kwanika Project

Years 1 to 5 LOM

Total Tonnes to Mill (000s) 27,375 73,663

Annual Tonnes to Mill (000s) 5,475 5,475

Average Grade

Copper (%) 0.35 0.377

Gold (g/t) 0.285 0.295

Silver (g/t) 1.057 1.378

Moly (%)* 0.001 0.016

Total Production (after Recovery)

Copper (000s lb) 187,993 544,892

Gold (000s oz) 175.6 489.0

Silver (000s oz) 697.7 2,447.7

Moly (000s lb)* 344.6 5,251

Average Annual Production

Copper (000s lb) 37,599 40,512

Gold (000s oz) 35.1 36.4

Silver (000s oz) 139.5 182.0

Moly (000s lb)* 68.9 390.4

* South Pit only

Table 22-3 Summary of the Economic Evaluation

Unit Base Case

Metal Price

Copper US$/lb 3.63

Gold US$/oz 1,427.00

Silver US$/oz 27.50

Molybdenum US$/lb 14.45

Exchange Rate US$:CDN$ $0.95

Economic Results (Pre-Tax)

Undiscounted cash flow $ M 567.1

NPV (at 5%) $ M 262.6

NPV (at 8%) $ M 143.3

NPV (at 10%) $ M 81.2

IRR % 13.4

Payback years 7.3



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22.2.2 Financial Evaluations of NPV and IRR

The production schedule has been incorporated into the 100% equity pre-tax financial model to develop

annual recovered metal production from the relationships of tonnage processed, head grades, and

recoveries.

Metal revenues are calculated based on market prices. Unit operating costs for underground mining,

processing, site services, G&A, and off-site charges (smelting, transportation, and royalties) areas are

applied to annual milled tonnages. Open pit mining operating costs are derived from first principles of

productivities and hourly operating costs. All operating costs are then added together to determine the

overall operating cost, which is deducted from the revenues to derive annual operating cash-flow (Net

Revenue).

Initial and sustaining capital costs have been incorporated on a year-by-year basis over the mine life and

deducted from the net revenue to determine the net cash flow before taxes. Initial capital expenditures

include costs accumulated prior to first production of copper, gold, silver and molybdenum; sustaining

capital includes expenditures for mining and processing additions, replacement of equipment, tailings

embankment construction, water treatment, and environmental/closure costs.

The financial analysis uses twelve months of the first year on-site operating cost as working capital cost.

The undiscounted annual cash flows are illustrated in Figure 22-1.

($600)

($400)

($200)

$0

$200

$400

$600

$800

-1 1 2 3 4 5 6 7 8 9 10 11 12 13 14

Cas

hFl

ow

(C

DN

$M

)

Pre-tax cash flow CAD$M Cumulative pre-tax cash flow CAD$M

Figure 22-1 Undiscounted Annual and Cumulative Cash Flow



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22.2.3 Sensitivity Analysis

Sensitivity analyses have been carried out on the following parameters:

copper and gold metal price

exchange rate

initial capital expenditure

on-site operating costs

copper and gold recovery

The analyses are presented graphically as financial outcomes in terms of NPV and IRR. Both the Project

NPV and IRR are most sensitive to Copper Price followed closely by Copper & Gold Recovery, with

Capital Cost having the least impact. The NPV and IRR sensitivities can be seen in Figure 22-2 and

Figure 22-3. These results are presented graphically only to show trends for future evaluation. At a

scoping level of engineering and costing the absolute values are not deemed relevant for economic

evaluation.



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0%

50%

100%

150%

200%

250%

-15% -10% -5% 0% 5% 10% 15% 20%

Sen

siti

vity

to

Bas

e C

ase

% ofBase Case

Input Sensitivities at 5% discount rate








Figure 22-2 Base Case Sensitivity to Pre-Tax NPV @ 5%



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0%

20%

40%

60%

80%

100%

120%

140%

160%

180%

-15% -10% -5% 0% 5% 10% 15% 20%

Sen

siti

vity

to

Bas

e C

ase

% of Base Case

Input Sensitivities to IRR








Figure 22-3 Base Case Sensitivity to Pre-Tax IRR

22.2.4 Royalties

The Kwanika mining leases are un-encumbered by Royalty Agreements.



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23.0 Adjacent Properties

23.1 Regional

The Quesnel Trough is the host to several other porphyry copper ± gold mines and significant deposits.

These deposits include: the Mount Polley Mine, the former Kemess Mine and its related infrastructure

located north of Kwanika, and the Mount Milligan Mine development project located approximately

85km south of Kwanika.

23.2 Local District

The adjacent Lustdust claims, owned by Alpha Gold Corporation, are located immediately to the north of

the Kwanika property. The Lustdust property has been the subject of exploration for fifteen years on

various precious and base metal vein and skarn occurrences and contains a small Indicated and Inferred

copper-gold Mineral Resource known as the Canyon Creek Zone. The other significant prospect in the

general vicinity of Kwanika is the Lorraine porphyry copper-gold property jointly controlled by Teck

Corporation and Lorraine Copper Corp. which contains a modest, Indicated and Inferred Mineral

Resource in two deposits. Both of these properties are the subject of current NI-43-101 reports.



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24.0 Other Relevant Data and Information MMTS has relied upon the previous Technical Reports by RPA on the Kwanika Property to define the

geological resources, as well as Serengeti to provide information regarding the existence and extent of

any environmental, legal, regulatory or First Nations liabilities to which the project is subject. Based on

the scoping level project defined by this PEA study, the knowledge of land use expectations, and the

regulatory process in British Columbia there is nothing that has come to light that is not a normal part of

the proposed mining operation. At this time, MMTS is not aware of any constraints in this regard that

may prevent the Mineral Resources at Kwanika from being exploited.



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25.0 Interpretations and Conclusions The Kwanika deposit represents a copper-gold-silver deposit that is amenable to open pit and

underground mining through the use of block caving and conventional milling consisting of flotation

concentration.

25.1 Geology and Resource Modeling

RPA carried out estimates of Mineral Resources for the Central Zone in 2008 and South Zones on the

Kwanika Property in 2009 and 2010. The current Mineral Resources estimates are summarized in Table

14-1 and Table 14-2 above.

RPA drew the following conclusions:

Drilling, core handling and sampling and security protocols were appropriate and samples should

be representative of the mineralization.

Conventional assaying techniques were used, sample QA/QC protocols were adequate and

checks at a secondary laboratory were consistent with the primary laboratory results.

Validation by RPA showed the sample database was reasonably free of errors, representative and

appropriate for use in Mineral Resource Estimates.

Due to a lower density of drillholes, geological interpretation, wireframe modelling and

geostatistical analysis in the South Zone are preliminary in nature and all of the resource was

categorized as Inferred.

Due to a sufficient density of drilling in the Central Zone, a block model was constrained by 3D

wire-framed geological domains constructed with grade interpolation by ordinary kriging for the

Central Zone and much of the resource was characterized as Indicated category.

Additional drilling would be required to increase the confidence level Inferred Mineral Resources

to Indicated category particularly in the South Zone.

25.2 Metallurgy

The metallurgical test work indicated that the mineralization responds well to the process

consisting of conventional flotation. Copper recovery of 88.5% in the locked cycle test to a

copper concentrate grading 27.7% Cu confirmed earlier results obtained in batch cleaner tests.

The concentrate contained 20.9g/t Au at a gold recovery of 65.2%. The test material responded

well to various collectors, which will allow the least costly reagents to be used in a commercial

operation. A follow-up test designed to investigate improved gold recovery from tailings

indicates that up to 10.5% additional gold can be recovered by very fine grinding and flotation of

rougher tailings.

The test program is only preliminary in scope, and as such, the composite sample taken from the

Central Zone may not be representative of the entire deposit.

With the project now focused on the Central and South pits, and Underground stopes in this PEA

mine plan, Metallurgical samples need to be collected that represent this part of the resource area

and appropriate lab test run for the anticipated feed grade range.

A mill throughput in the order of 15,000 tonnes per day is indicated as a basis of estimate.

Further evaluation of throughput is required in future studies. Variability Sampling and testing is

needed representing the rock type, mineralization, alteration, hardness, and grade variability

throughout the deposit. These include additional bond work index to characterize the hardness



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variability throughput the deposit, and better define the association of copper with precious

metals in the whole range of the expected mill feed grade, and their response to flotation.

The presence of impurities in final concentrate requires quantification.

Flotation products require the following evaluations:

o Settling and filtration of copper-gold concentrate.

o Sedimentation of flotation tailings.

o Potential acid generation from tailings.

Future flotation test work should assess the molybdenum recovery to the bulk copper concentrate

(Cu-Au-Mo), and evaluate molybdenum separation to a separate molybdenum concentrate.

Process design and costing in future studies will be required after the through put optimization is

completed.

25.3 Underground Mine Plan

The mineral zone evaluated for underground mining and the rock characteristics are suitable for

Block Caving. This mining method will provide a low operating cost.

Additional economic underground resources are possible both below and adjacent to the block

cave stope outline used in this report. Optimization of the underground mine plan should add to

these potential economic resources. This should include the use of optimization software to

define the economic footprint and also evaluation of the preferred development/extraction levels.

Further evaluation of the extraction levels could increase the amount of mineable material, and

can also improve project cashflow. Designing an extraction level closer to the open pit bottom

will allow higher grade underground mill feed to be produced earlier in the productions schedule

and with less up front capital development costs which will be positive to the project financials.

After the conceptual economic stope limits and preferred extraction levels are optimized,

planning and design can proceed at a Pre-Feasibility level. This will require geotechnical studies

and Engineering to evaluate caving parameters, caving rate, pre-conditioning requirements, and

draw point design details.

25.4 Open Pit Mine Plan

The Central and South pits are designed to their break-even economic limit and provide low cost

mining for the near surface material. The lowest benches have the highest incremental strip ratio

and are the higher cost on both a “per tonne mined” and “per tonne milled” basis. Optimization

of the mining method at the open pit/underground interface has the potential to improve the

overall economics of the property. Material on the bottom benches of the pit may produce a

better net return if mined as part of the underground interface. An evaluation is required to verify

this.

Geotechnical field investigations, including drilling, are required and the open pit design needs to

be taken to a Prefeasibility study level. This redesign work will also be based on the

Underground mining limit optimization, equipment optimization studies, and water and waste

rock management details.

25.5 General

The financial results from the base case plan used in this report are based on the parameters and

assumptions as described which are preliminary in nature. The financial sensitivity to these parameters

and assumptions has been indicated and as noted there are other areas in the mine plan that need



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optimization that could increase or decrease the economic mining limits. To advance the project to a Pre-

Feasibility level will require significant site investigation work and testing, including exploration drilling,

Geotechnical drilling and testing, metallurgical sampling and testing, and base line Environmental

studies. Before this Pre-Feasibility work is done, several optimization studies should be undertaken to

test the project’s sensitivity to key parameters and to guide the scope of a future Pre-Feasibility Study.

25.6 Opportunities

This PEA study provides a scoping level basis for a viable operation with the opportunity to add

more economic resources both on site and in the local area.

There are other properties in the local area that have the potential of using the Kwanika facilities

on a contract or joint venture basis.

The expanded resource can use the facilities and infrastructure from this study. A significant

mineralized resource in the Kwanika deposit surround the resources used in this mine plan. The

potential exists for some of this marginal material to be brought into an economic resource base

after this proposed operation has met capital payback, or if future expansions can provide a lower

operating cost due to economies of scale.



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26.0 Recommendations MMTS recommends advancing the project in two steps. First is a series of optimization studies to refine

the engineering and costing in the areas with higher economic sensitivity. This work will be based on

more engineering but will require minimal additional field data. The results of these studies will be a

revised PEA. If still justified after the Optimization studies, the project can then advance to a Pre-

Feasibility Study level, which will involve more field data for upgrading the inferred resources,

geotechnical evaluations, metallurgical test work, and field data collection to start the environmental

impact assessment and reclamation planning.

26.1 Optimization Studies

Based on the project’s economic sensitivity to various parameters, further engineering assessment work

should be done at a scoping level using existing data to reduce the down side risk and to focus the

planning and design work before extensive expenditures are incurred for a Pre-Feasibility study. These

optimizations should include:

Optimize the underground conceptual plan and add material not included in the current plan.

Investigate the ‘best’ economic mining method for the interface between open pit and

underground.

Refine equipment and planning parameters, for open pit, underground, and processing.

Estimate mining loss and dilution and dilution grades.

Update the capital cost estimates including a project timeline for development expenditure and

more detailed indirect costs.

With revisions to the underground plan, optimize the production schedule to bring higher grade

underground mill feed forward in time and with less up front development capital.

26.2 Pre-Feasibility Study

The project should advance to a Pre-Feasibility Study based on the results from the Optimization Studies.

The results of the Optimization studies will not only allow a “go/no go” decision but will also establish

areas requiring more detailed data and analysis for the Pre-Feasibility work.

26.3 Future Engineering Study Costs

To advance the project the following approximate costs will be incurred:

Optimization Studies $50,000

Pre-Feasibility $3,500,000



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27.0 References AMEC Americas Ltd. (2013): Preliminary Caveability Assessment – Central Zone - Kwanika Deposit –

Amended January 2013

Antill., T. (2011): Independent Review of Serengeti’s Kwanika Water Quality Program: EDI

Environmental Dynamics Inc.

Casselman, J., and Wilson, K. (2009): Kwanika Project 2008 Water Sampling Program: Ecofor

Clarke, H. (2011): Kwanika Project 2011 Water Sampling Program: Serengeti Resources

Craig, F. (2008): Archaeological Overview Assessment - Serengeti Resources Inc, Kwanika Creek

Claim, BC: Archer CRM Partnership

Davis., D. (2009): Preliminary Field Reconnaissance (PFR): Archer

Knight Piésold Consulting (2010): Fatal Flaw Overview Assessment for Creek Diversion at the Kwanika

Project – Channel Option

McNay, S. (2010): Environmental Monitoring on the Kwanika Mine Property: Wildlife Informetrics

Rennie, D.W., and Scott, K. (2009): Technical Report on the Kwanika Project, Fort St. James, British

Columbia: Scott Wilson, RPA.

Rennie, D.W. (2010): Technical Report on the Kwanika Project, Fort St. James, British Columbia: Scott

Wilson, RPA.

Rennie, D.W. (2011): Technical Report on the Kwanika Project, Fort St. James, British Columbia: RPA.

Samson, H. (2009): Kwanika Project 2009 Water Sampling Program: Serengeti Resources

Samson, H. (2010): Kwanika Project 2010 Water Sampling Program: Serengeti Resources

SGS Canada Inc. (2009): A Report on the Recovery of Copper and Gold from Kwanika Deposit.

SGS Canada Inc. (2009): Memo: Kwanika Follow-up Testwork Summary.

Wilson, K. (2008): Kwanika Creek Claim, Archaeological Overview Assessment Review, Technical

Memorandum: Ecofor

Wilson, K., (2009): Serengeti Resources Kwanika Property 2008 VEC Scoping Report: Ecofor

Casselman, J., and Wilson, K. (2008): Hydro and Stream Temperatures Report Near and Within the

Kwanika Property: Ecofor

Wilson, K. (2010): Kwanika Property- Caribou Overview for the 2010 Drill Program, Technical

Memorandum: Ecofor



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28.0 Certificates of Qualified Persons

Certificates are included in this section for the following Qualified Persons:

David W. Rennie (P.Eng.), Roscoe Postle Associates Inc.

James H. Gray (P.Eng.), Moose Mountain Technical Services

Tracey D. Meintjes (P.Eng.), Moose Mountain Technical Services



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DAVID W. RENNIE

I, David W. Rennie, of Vancouver, British Columbia, do hereby certify:

1. I am a Principal Geologist with Roscoe Postle Associates Inc. My office address is Suite 388,

1130 West Pender Street, Vancouver, British Columbia, Canada V6E 4A4.

2. This certificate applies to the technical report entitled “NI 43-101 Technical Report for the

Kwanika Property Preliminary Economic Assessment 2013”, dated March 4, 2013 (the

“Technical Report”).

3. I am a graduate of the University of British Columbia in 1979 with a Bachelor of Applied Science

degree in Geological Engineering.

4. I am a member in good standing of the Association of Professional Engineers and Geoscientists of

British Columbia (# 13572).

5. My relevant experience includes:

Review and report as a consultant on numerous exploration and mining projects around the

world for due diligence and regulatory requirements.

Consultant Geologist to a number of major international mining companies providing

expertise in conventional and geostatistical resource estimation for properties in North and

South Americas, and Africa.

Chief Geologist and Chief Engineer at a gold-silver mine in southern B.C.

Exploration geologist in charge of exploration work and claim staking with two mining

companies in British Columbia.

6. I am a “Qualified Person” for purposes of National Instrument 43-101 (the “Instrument”).

7. My most recent personal inspection of the Property was on October 18, 2011.

8. I am responsible for Subsections of Sections 1, 25, and 26 that pertain to the mineral resource

estimate, as well as Sections 4 through 12, and 14 (excluding 14.3) of the Technical Report.

9. I am independent of Serengeti Resources Inc. as defined by Section 1.5 of the Instrument.

10. I have no prior involvement with the Property that is the subject of the Technical Report.

11. I have read the Instrument and the Technical Report has been prepared in compliance with the

Instrument.

12. As of the date of this certificate, to the best of my knowledge, information, and belief, the

Technical Report, within my sections of responsibility referred to above, contains all scientific

and technical information that is required to be disclosed to make the Technical Report not

misleading.

Signed and dated this 4th

day of March 2013 at Vancouver, British Columbia

“Original document signed and sealed by

David W. Rennie, P.Eng.”

David W. Rennie, P.Eng.

Principal Geologist

Roscoe Postle Associates Inc.



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JAMES H. GRAY

I, James H. Gray, of Calgary, Alberta, do hereby certify:

1. I am a Mining Engineer with Moose Mountain Technical Services with a business address at

1975 1st Avenue South, Cranbrook, BC, V1C 6Y3.

2. This certificate applies to the Technical Report entitled “NI43-101 Technical Report for the



3. I am a graduate of the University of British Columbia (Bachelor of Applied Science – Mineral

Engineering, 1975).

4. I am a member in good standing of the Association of Professional Engineers and Geoscientists

of British Columbia (#11919), and the Association of Professional Engineers, and Geoscientists

of Alberta (Member #M47177).

5. My relevant experience includes operation, supervision, and engineering in North America, South

America, Australia, Eastern Europe, and Greenland.


7. My most recent personal inspection of the Property was on October 18, 2011.

8. I am responsible for Subsections of Sections 1, 21, 22, 25, and 26 that pertain to mining, and

Sections 2, 3, 14.3, 15, 16, 18 through 24 of the Technical Report, as well as the general

compilation of the report.




Instrument.




misleading.


day of March 2013 at Calgary, Alberta


James H. Gray, P.Eng.”

James H. Gray, P.Eng.

Principal Mining Engineer




of 139

TRACEY D. MEINTJES

I, Tracey Meintjes, of Vancouver, British Columbia, do hereby certify:

1. I am a Mining Engineer with Moose Mountain Technical Services with a business address at

1975 1st Avenue South, Cranbrook, BC, V1C 6Y3.

2. This certificate applies to the technical report entitled “NI43-101 Technical Report for the



3. I am a graduate of the Technikon Witwatersrand, (NHD Extraction Metallurgy – 1996).

4. I am a member in good standing of the Association of Professional Engineers and Geoscientists

of British Columbia (#37018).

5. My relevant experience includes process engineering, operation, supervision, and mine

engineering in South Africa and North America. I have been working in my profession

continuously since 1996.


7. I have not personally inspected the Property.

8. I am responsible for Subsections of Sections 1, 21, 22, 25, and 26 for issues pertaining to

metallurgy and mineral processing as well as Section 13 and 17 of the Technical Report.




Instrument.




misleading.


day of March 2013 at Calgary, Alberta


Tracey D. Meintjes, P.Eng.”

Tracey D. Meintjes, P.Eng.

Principal Mining Engineer



APPENDIX A Block Caveability Assessment (Amended Jan2013)

Kwanika DepositPreliminary Caveability Assessmenty y(Amended January 2013)

Company Name: Serengeti Resources Inc.Date: April 5, 2012From: Stephen Godden

SERENGETI RESOURCES INC.KWANIKA DEPOSIT

PRELIMINARY CAVEABILITY ASSESSMENT

Project No.: 170432 Page 2 14 January 2013

C O N T E N T S

1.0 INTRODUCTION ......................................................................................................... 1

2.0 SUMMARY .................................................................................................................. 3 2.1 Caveability and Operational Success .............................................................. 4 2.2 AMEC’s Preliminary Caveability Assessment ................................................. 6 2.3 Preliminary Cost Estimates ............................................................................. 6 2.4 Conclusions and Recommendations ............................................................... 7

3.0 GEOLOGY ................................................................................................................ 10 3.1 Regional Geology .......................................................................................... 10 3.2 Property Geology ........................................................................................... 10

4.0 OPTIONS AND COSTS ............................................................................................ 14 4.1 Underground Stoping .................................................................................... 15 4.2 AMEC’s Preliminary Cost Estimates ............................................................. 18 4.3 Comments and Conclusions .......................................................................... 19

5.0 BACKGROUND ......................................................................................................... 21 5.1 Natural Caving ............................................................................................... 21 5.2 The Importance of Scale and Fragmentation ................................................ 22 5.3 Limitations and Benefits ................................................................................ 23

5.3.1 Rockmass Rating Systems ................................................................ 24 5.3.2 The Roles of Veinlets, Faults and Shears ......................................... 24 5.3.3 Current Practice ................................................................................. 25

5.4 Rockmass Preconditioning ............................................................................ 25

6.0 THE CENTRAL ZONE ROCKMASS ......................................................................... 28 6.1 Caveability of the Mineralized Mass .............................................................. 31

6.1.1 An Empirical Assessment .................................................................. 32 6.1.2 Discussion ......................................................................................... 33 6.1.3 Comments ......................................................................................... 34

6.2 Fracturation of the Mineralized Mass ............................................................ 36 6.3 Caveability of the Overburden Rockmass ..................................................... 38

7.0 REFERENCES .......................................................................................................... 40

T A B L E S

Table 1: A Summary of Variables Used In MMTS’s Preliminary Financial Models ....................... 18 Table 2: A Summary of AMEC’s Preliminary and Provisional Unit Operating Cost Estimates for

Block Caving at 10 to 20 ktpd, Central Zone, Kwanika Property ..................................... 18 Table 3: A Summary of AMEC’s Caveability Inputs and Outcomes, Central Zone, Kwanika

Property ............................................................................................................................ 33




F I G U R E S

Figure 1: A Google Earth Image Showing the Location of the Kwanika ........................................... 1 Figure 2: A Conceptual Layout of the Kwanika Open Pit and Underground Mine Workings (looking

approximately north) ........................................................................................................... 3 Figure 3: A Geology Plan of the Kwanika Property, Showing to the Locations of the Central Zone

and South Zone Mineralization ........................................................................................ 11 Figure 4: Geology Section 200 N through Central Zone of the Kwanika Property .......................... 12 Figure 5: Geology Section 150N through Central Zone of the Kwanika Property, Showing to the

Distribution of the Supergene Enrichment Zone and the Location of the Adjacent Sedimentary Basin ........................................................................................................... 13

Figure 6: An East-West Section through Central Zone, Showing the Distribution of Higher-Grade Cu-Au Mineralization ........................................................................................................ 14

Figure 7: A Conceptual Layout of the Kwanika Open Pit and Underground Mine Workings (looking approximately north) ......................................................................................................... 17

Figure 8: A General View of Frozen Drillcore from Drillhole K-08-91, Located in a Wire Mesh Gabion at the Kwanika Project Site .................................................................................. 28

Figure 9: An Example of the Post-Drilling/Pre-Logging Condition of the Core (top photolog) and the Split and Generally Degraded Condition of the Same Core, as Seen During AMEC’s February 2012 Site Visit (bottom photolog) ..................................................................... 30

Figure 10: An Example of the Post-Drilling/Pre-Logging Condition of the Core (top photolog) and the Much-Broken, Post Hand-Split Condition of the Same Core, as Seen During AMEC’s February 2012 Site Visit (bottom photolog) ..................................................................... 31

Figure 11: Laubscher’s Caving Chart (from Bartlett, 1998), with the Estimated Central Zone MRMR Values Highlighted in Orange .......................................................................................... 33

Figure 12: An Example of the Widely and Occasionally Moderately to Closely Jointed Host Rockmass of the Main Mineralized Zone, Central Zone, Kwanika Property (note the 1.0 m long steel rules) ............................................................................................................ 37




1.0 INTRODUCTION

A meeting was held in February 2012 between Serengeti Resources Inc. (“Serengeti”), Moose Mountain Technical Services (“MMTS”) and AMEC to discuss mining options for exploiting the underground portion of the Central Zone deposit on Serengeti’s Kwanika property. It was concluded that given the available, average Mineral Resource grades of the target mineralization (29.9 Mt at 0.52% Cu and 0.54 g/tonne Au) and due to unit mining costs, only the block caving method could realistically be employed. In consequence of this, Serengeti asked AMEC to carry out a preliminary caveability assessment to test the technical viability of block caving in Central Zone. Serengeti also asked AMEC to compile preliminary estimates of unit operating costs for block caving, for inclusion in preliminary Project cashflow models to be compiled by Moose Mountain Technical Services (“MMTS”), for Serengeti. The results of AMEC’s investigations are presented in this report.

Central Zone is located in the northern part of Serengeti’s Kwanika property, in central British Columbia, some 150 kilometres north-northwest of Fort St. James (Figure 1). The basis for the caveability assessment was in detailed examinations of selected (by AMEC) drillcores during a site visit, made between February 27 and February 29, 2012 (“AMEC’s February 2012 site visit”). The cores were examined by Stephen Godden, Principal Mining Consultant, assisted by Patrick Lee, Mining Engineer in Training.

Figure 1: A Google Earth Image Showing the Location of the Kwanika Property in Northern BC, Canada




The preliminary nature of AMEC’s investigations is emphasized. More detailed geotechnical investigations and characterizations are required before firm and final conclusions can be made concerning appropriate cave stoping strategies, layouts and unit costs. The key areas that require consideration are identified in the main report text; a summary is provided in the following Section 2.




2.0 SUMMARY

AMEC’s preliminary caveability assessment was based on detailed, on-site examinations of selected (by AMEC) drillcore intersections of the mineralization targeted for underground mining. Reference was also made to drillcore photologs and to geology core logs that were supplied by Serengeti Resources Inc. (“Serengeti”). The geology logs included core recoveries and RQDs, by 3.05 m (10 ft) drillcore run.

For purposes of the investigations reported here, the target mineralization was divided into two zones: the “main mineralized zone” located vertically below the proposed open pit; and the “western extension zone” that forms an approximately horizontal mass that extends to the west from the base of the main mineralized zone (Figure 2). The conceptual mine plan and schedule for cave mining, which were prepared by Moose Mountain Technical Services (“MMTS”) for Serengeti, were assumed for purposes of analysis. The plan envisions block cave mining taking place in two lifts: from 680 m level to the bottom of the proposed Central Zone open pit at 780 m elevation; and to 680 m level from 480 m level (Figure 2, the level intervals are defined by their elevations).

Figure 2: A Conceptual Layout of the Kwanika Open Pit and Underground Mine Workings (looking approximately north) (from a report by MMTS entitled ‘Serengeti – Kwanika Potential Block Caving Cost Estimate and Cashflow with the Small O/P Case’ and dated February 03, 2012)

The “main mineralized zone”

The “western extension zone”

The upper production level at 680 m elevation

The lower production level at 480 m elevation

The access spiral ramp and

ventilation raise The final planned openpit (in RED)




For purposes of the preliminary investigations reported here, stress was discounted as a dominant control on rockmass caveability, because both mineralized zones are:

• located at moderate depths (up to 300 m below the proposed pit floor, or approximately 600 m below the current surface); and

• hosted in moderately competent to competent rocks (the estimated uniaxial compressive strengths of the dominant rock types approximate to 100 MPa in the main mineralized zone and 150 MPa in the western extension zone, although minor intersections of clay-rich material with estimated compressive strengths of less than 40 MPa exist [which preliminary strengths estimates need to be confirmed by means of structured laboratory testing]).

It should, however, be emphasized that stress will inevitably play a potentially beneficial role, which impact cannot be assessed without the benefit of orientated discontinuity data and detailed computer-based rockmass simulations. Such data is not currently available and such simulations fall outside the scope of the current investigations.

2.1 Caveability and Operational Success

All rockmasses can be made to cave naturally, as long as the mined spans over which they are allowed to deflect are sufficiently large to allow the caving process to start (the spans as defined, for example, by the dimensions of an undercut). However, it is the manner of caving and the resultant fragmentation size distribution that in large measure determine the potential for success of a cave mining operation. For example, depending on the prevailing rockmass characteristics, the caving process can be difficult and potentially dangerous to initiate, oversize blocks can either clog individual drawpoints and/or require extensive secondary blasting to ensure a block size distribution that is consistent with productive, cost efficient mining. Difficulties can also arise if the footprint of a target mineralized zone approximates to, or is less than, the spans required to either initiate a cave or to ensure that the caving process continues smoothly and progressively to a target elevation.

When stress does not play a dominant role in the caving process (as is assumed to be the case for the preliminary investigations reported here), rockmass caveability for the most part depends on the frequency and interaction of contained flaws such as joints. However, optimum results are achieved only if:

• the cave front migrates upwards through the target rockmass, in a smooth and progressive manner;




• the ultimate mined dimensions of the undercut were sufficiently large to promote continued upwards migration of the cave front, despite the span-limiting effects of the natural caving angle of the rockmass (which typically extends over the undercut area and thereby progressively reduces the effective span over which the caving rockmass can deflect); and

• the blocks of rock within a caved mass are sufficiently small to allow for both the controlled / controllable drawdown of the caved mass and optimal mucking cycle times at the production drawpoints.

With the preceding points in mind, it may be concluded that operational success could only be achieved over a narrow range of theoretically ideal rockmass conditions. For example, rockmasses with high fracture frequencies per metre, hence weak to moderately strong rockmasses with MRMR values of less than approximately 50, as defined by Laubscher’s modified rockmass rating system (which has, until fairly recently, been widely accepted as a means of directly assessing rockmass caveabilities).

In contrast, the majority of modern, mechanized cave mining operations successfully extract competent rock containing few open discontinuities that have MRMR values in excess of 50, and sometimes as much as 65. Portions of the western extension zone in particular fall into this latter category.

The difference between what in theory is possible and what in practice can be (and has been) achieved depends in part on the presence of large-scale structures such as faults and shear zones. Their impact varies with their frequency and distribution compared with the mined spans over which a target rockmass is able to deflect. Their effect is to promote bulk destabilization of an otherwise competent mass, thereby releasing large volumes of loosened blocks bounded by joints.

Weakly bonded veinlets in particular also play an important role, insofar as they can promote good fragmentation of a caved mass, due to comminution during drawdown of joint-bounded blocks. If such veinlets are not present in sufficient density and/or their bonding strengths are too high, rockmass preconditioning might be required. Two main preconditioning methods are used when block caving takes place in high quality rockmasses at moderate depths: bulk in situ (confined) blasting and hydraulic fracturing. In either case, the objective is to improve the caveability and/or fragmentation characteristics of a target mineralized rockmass by extending, shearing or opening existing fractures and, most importantly, by creating new fractures within the target mass.




2.2 AMEC’s Preliminary Caveability Assessment

The purpose of a caveability assessment is, in the first place, to assess the required undercut dimensions to achieve a natural caving condition. If these dimensions are compared with the footprint of a target mass, an assessment can be made as to whether it would cave smoothly and progressively without excessive dilution being incurred and/or without the need for rockmass preconditioning.

Laubscher’s empirical method for assessing rockmass caveability was employed for purposes of AMEC’s preliminary study. The preliminary and provisional results show that while the footprints of the main mineralized zone above both 680 m level and 480 m level are probably sufficient to initiate natural caves, they might be insufficient to ensure natural caving to the target elevations (the base of the proposed open pit from 680 m level and 680 m level from 480 m level).

Scrutiny of drillcore K-08-62 in particular shows that the main mineralized zone contains numerous structures that are recorded, in the drillhole logs, as shear zones that are invariably associated with strong sericite alteration and that are characterized by well-developed clay alteration. Their presence elevates confidence in the natural caveability of the main mineralized zone. However, an element of risk nevertheless remains for cave heights in excess of approximately 150 m, due to the limited footprint of mineralized mass. This suggests that rockmass preconditioning might be required; an alternative would be to increase the footprint and accept elevated levels of planned dilution as a result.

In the case of the western extension zone, provisional and preliminary analysis suggests that sufficiently large undercuts could be developed to induce a natural cave. However, the timing and nature of the initial cave might reasonably be questioned, due to the generally competent and moderately- to widely-jointed nature of the rockmass (there appears to be a trend towards an increasingly competent rockmass to the west,, over the western extension zone, where only a few fault- and shear-zones appear to be developed). This suggests that the mineralized mass is likely to cave dynamically and intermittently and, because the examined veinlets appear to be uniformly competent and strong, preconditioning would almost certainly be required before block cave mining could successfully be employed.

2.3 Preliminary Cost Estimates

AMEC’s provisional and preliminary cost analysis suggests a unit mining cost approximating to C$7.00 per tonne mined for a block caving operation located in British Columbia, exclusive of capital development and rockmass preconditioning. Unit costs for rockmass preconditioning vary widely between locations, the method used




and the amount of rockmass preconditioning required. For purposes of the preliminary investigations reported here, it was assumed that the unit cost would be approximately C$1.00 per tonne mined. As regards these and other operating costs it is emphasized that:

• the estimated costs are for owner-operated mines (higher rates apply in contractor-mined operations);

• for purposes of preliminary analysis, AMEC recommends a unit cost of C$3.00 per tonne milled for G&A;

• according to information available to AMEC, unit operating costs have escalated by approximately 25% per year since 2008 (when they dropped slightly following five straight years of progressive increase), due mainly to inflationary pressure resulting from increasing metal prices;

• to the best of AMEC’s knowledge, the capital cost for projects has gone up approximately five-fold since 2002, not least due to the cost of capital items such as mining equipment, the cost of which continues to escalate rapidly due to strong demand; and

• according to information available to AMEC, the lead time for key mining equipment such as jumbos and LHDs can currently be as much as eighteen months, sometimes more.

2.4 Conclusions and Recommendations

AMEC’s provisional and preliminary analysis suggests that block cave mining is a technically and economically viable proposition. However, block caving in the main mineralized zone appears to be a better economic proposition, due to:

• the generally more amenable rockmass conditions that appear to be developed in the main mineralized zone (the rockmass appears to be more competent, far fewer faults appear to be developed and the bonding strength of the veinlets appears to be greater in the western extension zone); and

• the taller (200 m) in the main mineralized zone above 480 level that yields a better development metres to ore tonnes ratio compared with the western extension zone, especially as regards production drawpoint development.

It is for the reasons outlined that AMEC recommends that a cost benefit analysis of cave mining in the western extension zone is carried out, assuming the need for preconditioning.




A key difference between the main and western extension zones appears to be the severity of alteration, which is generally more developed in the main mineralized zone. In this regard, AMEC recommends that that the final volume/target mass of above-average grade mineralization is zoned according to the dominant alteration type or types, inclusive of potassic, sericitic and sodic alteration. Various other key data gaps exist, the closure of which would be required before the project could be advanced to the pre-feasibility stage of mine design and planning. In these regards, AMEC recommends that the following studies are completed:

• that the block cave dimensions and footprints of the target mineralized zone or zones are accurately determined;

• that carefully located and orientated, HQ diameter and (triple tube) cored geotechnical holes are drilled to enable discontinuity orientations and frequencies to accurately be defined (inclusive of fault and shear zones);

• that a three-dimensional fault- and shear-zone model is compiled, using data derived from previous and new drillcore logs;

• that a targeted laboratory testing program is carried out to determine both the average compressive strengths and indirect tensile strengths of the dominant rock types, as well as the average shear parameters and tensile strengths of the dominant discontinuity types; and

• that a program of comprehensive geotechnical rockmass characterizations is carried out to establish the number, distribution and average characteristics of geotechnical domains (i.e. zones or areas in which geotechnically similar rockmass conditions apply), inclusive of both the mineralized zones and sedimentary overburden.

Once the studies outlined are complete, detailed rockmass simulations should be compiled to assess in detail the caveability of the target mineralized zone or zones, inclusive of preconditioning requirements. The results would provide details of the parameters required for mine design and planning, within the scope of which consideration should be given to the cost benefits of placing, to the extent possible, mine development in ore (i.e. in above cut-off mineralized material, to enhance potential financial returns).

Risk assessments should also be carried out, inclusive of subsidence modeling, to assess the extent and potential surface impacts of the caving operations (inclusive of the siting of surface infrastructure, dumps and dams). Although no evidence of groundwater movement was seen in the core examined during AMEC’s February site




visit, an investigation and characterization of the geohydrological conditions before and after caving should be carried out.




3.0 GEOLOGY

For purposes of the investigations reported here, AMEC relied on the geological information contained in a technical report by Roscoe Postle Associates Inc. entitled ‘Technical Report on the Kwanika Project, Fort St. James, British Columbia, Canada’ and dated March 03, 2011. The text presented in the following Sections 3.1 and 3.2 is a modified version of the text contained in the cited technical report that is available on www.sedar.com.

3.1 Regional Geology

The Kwanika property is situated in the northern part of the Upper Triassic to Lower Jurassic Quesnel Trough, which comprises a belt of Lower Mesozoic volcanic rocks and intrusions lying between highly deformed Proterozoic and Paleozoic strata to the east and deformed Upper Paleozoic strata to the west (Garnett, 1978). Quesnel Trough hosts numerous alkalic and calc-alkalic porphyry copper-gold deposits, the dominant host rocks for which include andesitic volcanics intruded by monzodiorites and monzonites.

3.2 Property Geology

Two porphyry-type deposits, called Central Zone and South Zone, are located on the Kwanika property. The former contains copper-gold mineralization, the latter contains copper-molybdenum-gold-silver mineralization. It is Central Zone that is the subject of the investigations reported here. South Zone is a separate deposit that is located some two kilometres to the south of Central Zone (Figure 3).

The Central Zone deposit is characterized by the presence of two major and several minor intrusive bodies of the multi-phase Hogem Batholith that intrudes a succession of andesitic rocks of the Takla Volcanic Group. Hypogene mineralization consists of disseminated chalcopyrite, bornite, and pyrite in and around a potassic-altered monzonite stock. Where strongly mineralized, the unit commonly displays quartz stock work and hydrothermal brecciation. The highest mineral grades occur within zones of strong to intense, texture destructive albite-hematite alteration that commonly occurs at the top of the hypogene mineralized zone.

The intrusive and volcanic units of the Central Zone have been locally rotated and unconformably overlain to the west by a clastic sediment-filled, ‘U’-shaped basin that probably represents a half-graben. Both the Central Zone mineralization and the sedimentary sequence are truncated to the west by the regionally dominant Pinchi fault (Figure 3). The principal sedimentary rock types include well-bedded, weakly pyritic siltstones, fine- to coarse-grained and massive to thickly bedded sandstones, and polymictic conglomerates containing clasts of sandstone, siltstone and unidentified volcanic and intrusive rocks. Copper-gold mineralization in the Central Zone occurs to




the east of the sedimentary basin and below the basin from Geology Section 225 S to Geology Section 350 N (for example, Figure 4).

Figure 3: A Geology Plan of the Kwanika Property, Showing to the Locations of the Central Zone and South Zone Mineralization (from a NI43-101 Technical Report by Roscoe Postle Associates Inc. entitled ‘Technical Report on the Kwanika Project, Fort St. James, British Columbia, Canada’ and dated March 03, 2011)




Figure 4: Geology Section 200 N through Central Zone of the Kwanika Property (supplied by Serengeti Resources Inc. and in a NI43-101 Technical Report by Roscoe Postle Associates Inc. entitled ‘Technical Report on the Kwanika Project, Fort St. James, British Columbia, Canada’ and dated March 03, 2011)

A supergene enrichment zone exists below the unconformable contact with the sedimentary sequence (for example, Figure 5). Its thickness varies between five and 70 metres; it extends laterally for up to 500 m. Two distinct assemblages of supergene mineralization are observed in the Central Zone: minor, supergene oxide (native copper) and prevalent, supergene sulphide (chalcocite, covellite).




Figure 5: Geology Section 150N through Central Zone of the Kwanika Property, Showing to the Distribution of the Supergene Enrichment Zone and the Location of the Adjacent Sedimentary Basin (from Serengeti’s website www.serengetiresources.com)




4.0 OPTIONS AND COSTS

A potential, 300 m deep Central Zone open pit has been defined by Moose Mountain Technical Services (“MMTS”), at the request of Serengeti. It targets higher-grade, near-surface mineralization, to an elevation of approximately 780 m.

A meeting was held in February 2012 between representatives of Serengeti (David Moore and Hugh Samson), MMTS (Jim Gray) and AMEC (Stephen Godden), at Serengeti’s Vancouver offices (the “February meeting”). During the February meeting it was demonstrated by MMTS that, as a stand-alone project, the proposed open pit would be sufficient to cover the capital costs of project development only, assuming MMTS’s cost estimates. Underground mining of the higher-grade mineralized material, located beneath the proposed open pit, would therefore be required to ensure a profit margin sufficient to ensure sustainable economic viability. For purposes of this report, the higher-grade mineralization located vertically below the proposed open pit is called the “main mineralized zone”. The area to the west of the open pit, which too is targeted for underground mining, is called the “western extension zone” (Figure 6).

Figure 6: An East-West Section through Central Zone, Showing the Distribution of Higher-Grade Cu-Au Mineralization (from a hard copy of a computer screen image supplied by Serengeti)

The boundary of the higher-grade

mineralized envelope

The cross-section profile of MMTS’s proposed open pit

The “western extension zone”

The “main mineralized zone”




4.1 Underground Stoping

Three options for underground stoping were considered during the February meeting: open stoping, sub-level caving and block caving. MMTS had previously compiled a preliminary underground mining model, inclusive of cost estimates, for sub-level caving with production levels at 30 m intervals that extended off a main access spiral ramp. Assuming an all-in cost of C$35.00 per tonne for mining, processing and site overheads, MMTS demonstrated that the project would be marginal, or at least that the operating margin would quickly be eroded if operating costs increased or lower metal prices prevailed (MMTS assumed copper at C$2.84/lb NSP and gold at C$1,128/oz NSP).

If sub-level caving results in marginal economics, the option of open-stoping could not be used because, according to information available to AMEC, the method is usually at least C$15.00 per tonne more expensive than sub-level caving. However, the cheaper option of block caving could be employed, if it yielded a robust financial return.

In consequence of the findings outlined, MMTS carried out a preliminary analysis of the economic viability of block caving the higher-grade mineralization targeted for underground mining (Gray, 2012). The following design assumptions were applied:

• the main mineralized zone would be cave-mined on two production levels at 680 m elevation (upper) and 480 m elevation (lower) (see Figure 7) –

− the height of cave is limited to 100 m for the upper level, thereby bringing it into production sooner than might otherwise be the case, and

− the cave for the lower level takes the remaining 200 m of mineralized material in the main mineralized zone; and

• level development extends to the west, as appropriate, to exploit the higher-grade mineralization distributed across the western extension zone, where the thickness of the higher-grade mineralization is limited to approximately 100 m.

MMTS’s preliminary and provisional planning for underground mining further assumed that the access ramp would start from surface, rather than off a pit wall, due to the potential for destabilization as a result of cave mining activity. Its development to 680 m level, as well as that of the provisionally planned service raise, was scheduled to start during Phase 1 of the planned open pit (the pit profile highlighted in PURPLE on Figure 7). Production on 680 m level started when Phase 2 of the planned open pit was complete (the pit profile highlighted in RED on Figure 7), at which time




development to 480 m level also commenced. Finally, production on 480 m level was scheduled to start when production on 680 m level was complete.




Figure 7: A Conceptual Layout of the Kwanika Open Pit and Underground Mine Workings (looking approximately north) (from a report by MMTS entitled ‘Serengeti – Kwanika Potential Block Caving Cost Estimate and Cashflow with the Small O/P Case’ and dated February 03, 2012)

MMTS demonstrated that for the input variables summarized on Table 1 and assuming full utilization of both the Central Zone and South Zone resources (the latter by open pit mining only, for an estimated combined mine life of 13.8 years):

• for a unit cost for block cave mining of C$5.00 per tonne, the combined open pit and underground mining project would yield an estimated cumulative net cashflow of $463.1 million, an NPV of $61.6 million at a discount rate of 8%, an NPV of $9 million at a discount rate of 10% and an IRR of 10.4%; whereas




• a unit cost for block cave mining of C$10.00 per tonne would yield an estimated cumulative net cashflow of $313.1 million, an NPV of -$6.9 million at a discount rate of 8%, an NPV of -$48.3 million at a discount rate of 10% and an IRR of 7.7%.

Table 1: A Summary of Variables Used In MMTS’s Preliminary Financial Models (from a report by MMTS entitled ‘Kwanika Potential Block Caving Cost Estimate and Cashflow with the Small O/P Case’, dated February 03, 2012) Parameter Value/Cost Capital CostsCapital Development

- to and on 680 m level - to and on 480 m level

Mining Equipment

C$30.77 million C$68.29 million C$22.12 million

Unit Operating CostsOpen Pit Mining Block Cave Mining Processing G&A*

C$ 2.33/tonne

C$5.00 or 10.00/tonne C$10.70/tonne C$ 1.20/tonne

Stoping** Planned Recovery Planned Dilution

95% 2%

Process RecoveriesCopper Gold Silver

89% 70% 75%

Metal Prices (NSP)Copper Gold Silver

C$2.84/lb

C$1,127/oz C$13.79/oz

Notes: * - For the reasons later described, AMEC recommends that a unit cost of C$3.00/tonne is assumed for purposes of preliminary analysis

** - In AMEC’s experience, planned recovery rates typically average 90% in block cave operations, along with an average planned dilution rate of 10%

4.2 AMEC’s Preliminary Cost Estimates

Table 2 summarizes AMEC’s provisional and preliminary unit operating cost estimates for block caving in Central Zone at the Kwanika property. The subsequent comments and discussions apply.

Table 2: A Summary of AMEC’s Preliminary and Provisional Unit Operating Cost Estimates for Block Caving at 10 to 20 ktpd, Central Zone, Kwanika Property Parameter Value/Cost Unit Operating CostsBlock Cave Mining Pre-Conditioning

C$7.00/tonne mined C$1.00/tonne mined

G&A C$3.00/tonne milled




The cost of block cave mining generally ranges between C$5.00 and C$7.00 per tonne of ore mined, for ramp-accessed stopes (for all mining activities, up to and including delivery of broken ore to an on-site mill, assuming truck haulage and inclusive of underground mine management and supervision but exclusive of capital development, processing and administration). The range of costs is based on AMEC’s experience and knowledge, reviews of previous studies and reviews of year-end reports for currently operating mines.

For an operation producing between 10ktpd and 20ktpd, data available to AMEC from a variety of mines indicates a unit mining cost of approximately C$6.50 per tonne. This suggests a unit cost approximating to C$7.00 per tonne for a block caving operation located in British Columbia, exclusive of capital development and rockmass preconditioning that, for the reasons later outlined, will probably be required to successfully block cave at least portions of the Central Zone orebody.

Unit costs for rockmass preconditioning vary widely between locations, the method used and the amount of rockmass preconditioning required. For purposes of the preliminary investigations reported here, it was assumed that the unit cost would approximate to C$1.00 per tonne mined.

It is emphasized that the unit costs outlined are for owner-operated mines. Formal camp facilities will almost certainly be required if the Kwanika project develops fully to the mine production and processing phase, due mainly to its semi-remote location. In AMEC’s opinion, the G&A cost of C$1.20 per tonne that has thus far been assumed might be optimistic – AMEC recommends that, for purposes of preliminary analysis, a unit cost for G&A of C$3.00 per tonne milled should be assumed.

4.3 Comments and Conclusions

If the unit operating costs outlined are accepted as reasonable estimates as of March 2012, MMTS’s analysis suggests that the project, as earlier outlined, would probably be sustainably profitable if the higher-grade mineralization, targeted for underground mining in Central Zone, was exploited using block caving methods. The key question however remains: is the Central Zone rockmass amenable to block cave mining? It is this question that is considered in the following sections.

The March 2012 date of AMEC’s preliminary and provisional cost estimates should be emphasized: according to information available to AMEC, unit operating costs have escalated by approximately 25% per year since 2008 (when they dropped slightly following five straight years of progressive increase), due mainly to inflationary pressure resulting from increasing metal prices.




To the best of AMEC’s knowledge, the capital cost for mining projects has gone up approximately five-fold since 2002, not least due to the cost of capital items such as mining equipment, the cost of which continues to increase year-on-year, due to strong demand. The lead time for key items such as jumbos and LHDs can currently be as much as eighteen months, and sometimes more.




5.0 BACKGROUND

5.1 Natural Caving

All rockmasses can be made to cave naturally, as long as the mined spans over which a rockmass is allowed to deflect under the influence of gravity are sufficiently large to allow the caving process to start (the spans as defined, for example, by the dimensions of an undercut). Caving can, however, stop if a stable geometry develops in a cave back. This affects directly the continued ability of a deflecting rockmass to cave, although it should be noted that details of the caving geometry can by influenced strongly by the presence of major discontinuities such as faults.

According to Laubscher (1994), two types of natural caving can occur: stress caving and subsidence caving. Laubscher defines stress caving as that which occurs in virgin cave blocks when the stresses acting in the cave back exceed the strength of the rockmass. He defines subsidence caving as that which occurs when adjacent mining has removed the lateral restraint on the block being caved, which can result in a rapid propagation of the cave. For either case, Laubscher identified the following variables as controlling the natural caveability of deflecting rockmasses:

• rockmass strength, as defined by his modified rockmass rating system (MRMR);

• rockmass structure (i.e. the presence and continuity of flaws such as joints);

• the directions and magnitudes of in situ stress plus the magnitudes of induced stress (hence the magnitudes and directions of application of field stress) versus the orientation of jointing and/or the strength of the bulk in situ rock targeted for caving;

• the hydraulic radius of the orebody of interest, versus the footprint of the same orebody (i.e. the dimensions over which the rockmass is able to deflect); and

• the presence or lack of groundwater.

In the case of the main mineralized zone and the west extension zone, it may be assumed that little or no groundwater exists and that stress is not a dominant controlling factor, because:

• no evidence for groundwater movement, in the form of wallrock alteration / weathering, joint plane staining or limonite accumulation along joint planes was found in, or reported for, the examined drillcores, below the main surface




weathered zone that extends to between approximately 25 m and 35 m below surface;

• the target mineralization is located at moderate depths (up to 300 m below the proposed pit floor, or up to approximately 600 m below the current surface); and

• qualitative assessments of rock strengths, carried out during AMEC’s February 2012 site visit, showed that the target mineralization is hosted in moderately competent to competent rocks (the estimated uniaxial compressive strengths of the dominant rock types approximate to 100 MPa in the main mineralized zone and 150 MPa in the western extension zone, although minor intersections of clay-rich material with estimated compressive strengths of less than 40 MPa exist).

In other words, according to Laubscher’s classification and assessment method, natural caving of the target mineralized mass will depend on, and vary according to, the density and orientation of jointing and the mined spans over which the target mass is able to deflect.

5.2 The Importance of Scale and Fragmentation

Apart from the fundamental aspects of rockmass caveability, it is the manner of caving and the resultant fragmentation size distribution that in large measure determine the potential for success of a cave mining operation. For example, for the type of rockmass conditions outlined above for the target mineralized mass:

• if the host rockmass was comprised of massive to widely jointed and competent rock, large undercut spans might be required before a natural cave would initiate (thereafter intermittent, large slab-type caving activity only might develop because caving would in large measure require dynamic, snap-through type failures along the under-surface of the deflecting mass, much in the manner suggested by conventional beam theory [Brady and Brown, 2005]); whereas

• if the host rockmass was moderately jointed and the joints persistently interacted to form blocks, progressive caving due to gravity effects could occur once the critical spans for the rockmass were achieved (i.e. once the dimensions of the undercut matched the magnitudes required to induce the onset of failure).

Sudden, dynamic or large initial caves can result in shock-loading, hence accelerated damage to drawpoints. In some circumstances, dangerous airblasts can also develop, with further airblasts created if a significant air gap / void existed on top of a previously caved mass and large, slab-type caves occurred at intermittent intervals. If large block




sizes only were generated during a cave, an erratic drawdown profile could develop, which should be avoided for the reasons described in Brown, 2007. Individual drawpoints could also become blocked and/or extensive secondary blasting might be required to reduce the blocks to dimensions suitable for handling using production LHDs (which would inevitably increase mucking cycle times to the detriment of operational efficiencies, hence costs). If groundwater is present it would tend to promote the natural caving process, due to pore pressure effects along discontinuities. However, if bulk failures along water-bearing structures occurred, large volume water inflows might occur to the detriment of safety.

In other words, optimum results could be achieved only if a cave front migrated smoothly and progressively through the deflecting host rockmass, and the blocks of rock within a caved mass were sufficiently small to allow for both its controlled / controllable drawdown and optimum mucking cycle times at the production drawpoints. The wealth of accumulated industry experience serves only to emphasize these points (for example, Brown, 2007 and Chitombo, 2010).

5.3 Limitations and Benefits

Given the constraints outlined, it may reasonably be concluded that in theory, natural caving could only successfully be achieved over a narrow range of rockmass conditions. Such rockmasses comprise what are sometimes termed “secondary rocks”, or rocks that cave and fragment readily due to a high frequency of open discontinuities, hence rockmasses with high fracture frequencies per metre which may or may not include faults and shears (Chitombo, 2010). In other words, in weak to moderately strong rockmasses at moderate depths with MRMR values of less than approximately 50, as defined by Laubscher’s modified rockmass rating system (Laubscher, 1994, Laubscher and Jakubec, 2001, which system has until recently been widely accepted as a means of assessing rockmass caveability, based on MRMR determinations and his caving chart – see Section 6).

In contrast, the majority of modern, mechanized cave mining operations extract what is sometimes termed “primary rock”, or competent rock containing few open discontinuities but which, when examined closely, contain high frequencies of small-scale veins (“veinlets”) coupled with widely-spaced faults (Chitombo, 2010). Although such rockmasses can have MRMR values in excess of 50 (and sometimes as much as 65), block cave mining has become accepted and widely successful international mining practice, albeit that some challenges need still to be resolved (Brown, 2007).




5.3.1 Rockmass Rating Systems

Rockmass rating systems were originally developed with the purpose of defining average rockmass conditions for assessing ground behaviour and support requirements for small span excavations such as tunnels. Bieniawski’s RMR system was subsequently extended by a number of authors, ostensibly to render it more suitable for particular applications (e.g. Kendorski et al 1983 and Romana 1985). However, only Laubscher has attempted to extend Bieniawski’s RMR system to cater for diverse mining situations.

The fundamental difference between Bieniawski’s RMR and Laubchser’s MRMR systems was the recognition that in situ rockmass ratings (RMR) had to be adjusted according to the mining environment so that the final ratings (MRMR) could be used for mine design. Adjustments were introduced for weathering, mining-induced stresses, joint orientation and blasting effects (Laubscher 1990, 1994). Some revisions were published in 2001 (Laubscher and Jakubec, 2001), but the fundamentals of the original MRMR system remained largely unchanged.

5.3.2 The Roles of Veinlets, Faults and Shears

The successes of current block cave operations mining primary rock highlight the limitations of the variables that contribute to MRMR estimates: they do not include the potential benefits of either weak to moderate-strength veinlets or, specifically at least, large-scale discontinuities such as faults and shears (although consideration is given to ‘major structures’ within the scope of the amended MRMR system [Laubscher and Jakubec, 2001]).

Veinlets are not specifically considered within the scope of MRMR estimations, or indeed any of the more established rockmass rating systems (the RMR system [Bieniawski 1974, 1989] and the Q system [Barton et al 1974]). This limitation applies because they do not qualify as discontinuities: veinlets may be described as healed structures; whereas discontinuities are naturally occurring, in situ flaws across which the structural continuity of a rockmass is broken and along which local rockmass parting can occur. Despite this, industry cave mining experience shows that weakly bonded veinlets in particular can promote good fragmentation, as a consequence of comminution during the draw-down of a caved mass of joint-bounded blocks (Brown 2007 and Chitombo 2010).

In the case of faults and shears, they tend to act as large-scale and often dominant discontinuities within an otherwise generally jointed and/or layered/bedded rockmass. As such, they can promote bulk destabilization of an otherwise competent mass, to thereby release large volumes of loosened blocks bounded by joints. As the caved




mass is drawn down block rotation can occur, causing mechanical attrition and the progressive reduction in block size, to the further benefit of the bulk caving and fragmentation process.

Although faults and shears are indirectly considered within the scope of all rockmass rating systems (usually within RQD and/or fracture frequency determinations), their impact on bulk rockmass behaviour/caveability can only realistically be assessed by means of computer-based simulations. This is because their behaviour is similar to that of joints, but on a much larger scale: their selective failure will occur if they are free to move by virtue of their interaction with other major discontinuities or a combination of smaller-scale discontinuities, over the mined spans across which movement can occur. In other words, if they are free to move then they will, irrespective of the condition or quality of the general rockmass in which they are developed.

The limitations outlined do not render inappropriate Laubscher’s (empirical) method for widespread application. The method should, however, be used with caution, as a preliminary tool for preliminary estimations of rockmass caveability, especially when the assessed MRMR value exceeds 50 (Brown, 2007). In other words, his method provides a guideline only - more detailed, analytic methods for assessing rockmass caveability should be applied before firm and final conclusions can be made and both design and planning can be advanced with confidence.

5.3.3 Current Practice

Current, developing practice for assessing rockmass caveability includes numerical modelling of closely (geotechnically) defined rockmasses, thereby to better understand the complexities of the physics of caving that varies between places and rockmasses. Mining block heights, production rates, undercutting strategies, extraction level layouts and mining rates also have to be considered within the scope of analysis and design for least risk mining, especially when primary rocks are targeted for block cave mining. For purposes of the preliminary investigations reported here, caveability and fragmentation potential only are considered, with the objective of establishing the efficacy of the block caving method for exploiting higher-grade mineralization contained within Central Zone of the Kwanika property.

5.4 Rockmass Preconditioning

Circumstances can arise where the potential for natural caving is limited by virtue of a high average MRMR for the target mass, coupled with a lack of bulk destabilizing structures such faults and shears and/or a lack of weakly to moderately bonded veinlets. As earlier outlined, it is in such circumstances that the caving process can be




difficult and potentially dangerous to initiate, oversize blocks can either clog individual drawpoints and/or extensive secondary blasting can be required to ensure a block size distribution that is consistent with productive, cost efficient mining. Difficulties can also arise if the footprint of a target mineralized zone approximates to, or is less than, the spans required to either initiate a cave or to ensure that the caving process continues smoothly and progressively to a target elevation.

It is in the types of circumstances outlined that rockmass caveability, as well as potential block size distributions, can be enhanced by means of rockmass preconditioning. Industry experience encompasses two main preconditioning methods when block caving takes place at moderate depths: bulk in situ / confined blasting; and hydraulic fracturing (Brown 2007). In either case, the objectives are to improve the caveability and/or fragmentation characteristics of a target mineralized rockmass by extending, shearing or opening existing fractures and, most importantly, by creating new fractures within the target mass.

If suitable access exists, an obvious option would be to drill and blast long holes across a cave back to promote smooth and progressive caving. Blastholes can also be drilled in the lateral direction from adjacent excavations such as a spiral ramp, which might be a viable option for the main mineralized zone. Whatever the case, industry experience to date suggests that the method can yield robust results (Brown, 2007), not least because with careful blast design, dilution from the wallrocks surrounding a target mineralized zone can be kept to a practicable minimum, especially if perimeter blasting is employed. Very steep caving angles can also be induced, to the further benefit of project economics.

There is an inevitable additional cost for preconditioning using blasting methods, which varies with the number and length of holes to be blasted. However, the cost is not necessarily onerous. For example, industry case histories include preconditioning of a rockmass with a mean MRMR of 65, by drilling and blasting 19, 100 m to 112 m long, 140 mm diameter vertical drillholes across a 7,000 m² area, using an emulsion-type blasting agent (Brown, 2007).

Hydraulic fracturing is cheaper, but it is a technically more challenging method to employ, not least as the fluids (which are introduced under pressure) can selectively move along pre-existing fractures and/or selectively induce factures parallel to the direction of maximum principal stress (the latter for the reasons apparent from consideration of the elastic prototype ‘hole in an infinite elastic solid’ – Coates, 1981). In either case, the effect is to limit the extent and frequency of new fracture development.




The introduction of water-based fluids into an otherwise dry rockmass can also be problematic, if swelling-type clays are present. This is because mud or slurry reporting to drawpoints can be problematic and unsafe, depending on the volume and viscosity of the material. In contrast, if gypsum infillings along discontinuity planes are present, they can partially be dissolved and eroded by the introduced fluids, to the benefit of rockmass caveability (Brown, 2007).




6.0 THE CENTRAL ZONE ROCKMASS

The caveability of the Central Zone rockmass was initially assessed through the close examination of selected (by AMEC) drillcores that were stored in sealed containers at site (i.e. not in wire mesh gabions, which core was difficult to access and generally frozen in the core trays during AMEC’s February 2012 site visit – for example, Figure 8). Drillcores K-07-14, K-08-58, K-08-62, K-08-78 and K-08-93 were selected because they covered all principal parts of the target mineralized mass, inclusive of the central portion of the main mineralized zone.

Figure 8: A General View of Frozen Drillcore from Drillhole K-08-91, Located in a Wire Mesh Gabion at the Kwanika Project Site (the approximately 5 ft of accumulated snow was removed immediately prior to the photograph being taken) (taken during AMEC’s February 2012 site visit)

Cross-references were made to the drillcore-relevant photologs and the geology logs during the drillcore examination process; reference was made to the photologs and geology logs of other drillcores during the process of post-visit analysis leading to AMEC’s findings that are detailed in this report. The photologs and geology logs were




supplied by Serengeti. The geology logs included drillcore recoveries rates and RQDs by 3.05 m long drillcore run, which values were determined by Serengeti’s responsible geologists, during the initial logging process.

Although the focus of AMEC’s investigations was on the host monzonite, dioritic-monzonite, diorite and andesite rockmass for the copper-gold mineralization, intersections of the overburden sedimentary sequence across the western extension zone were examined to the extent possible (most of the drillcore intersections are stored in wire mesh gabions). Cross-references to the relevant drillcore photologs were again made, to facilitate an understanding of the in situ geotechnical characteristics of the sedimentary sequence.

The drillcore photologs were reported by Serengeti to have been taken shortly after each core-relevant hole was completed, and prior to commencement of the drillcore logging process. As such, the condition of the drillcore seen in the photologs may reasonably be construed to be close to that found when the core was removed from the core barrels.

The importance of Serengeti’s photologs to the investigations reported here was emphasized by the condition of the drillcore seen during AMEC’s February 2012 site visit: all the host rocks had been split for purposes of sampling, sometimes using a hand-splitter but more usually using a diamond saw. Mechanical damage during core tray loading, logging, transport and handling was evident, as was the much degraded nature of clay-rich intervals that were logged as shear zones. Figures 9 & 10 are examples of this – they are photolog pairs showing the original condition of the stated core intervals and the condition of the same core intervals, as seen during AMEC’s February 2012 site visit.

The opening up of numerous, otherwise tightly infilled discontinuities was also seen, especially where the infilling comprised albite, calcite or gypsum. This may in part be attributed to in-box mechanical degradation during transport and handling, but the breakdown of the mineral-rock contacts caused by thermal cycling, over several winter and summer seasons since the holes were drilled, might also be a contributing cause. Whatever the case, the condition of the core seen by AMEC precluded conventional geotechnical drillcore logging. The drillcores were instead examined in detail, at least to the extent possible, with Serengeti’s photologs being used to estimate in situ discontinuity densities.




Figure 9: An Example of the Post-Drilling/Pre-Logging Condition of the Core (top photolog) and the Split and Generally Degraded Condition of the Same Core, as Seen During AMEC’s February 2012 Site Visit (bottom photolog) (Drillhole K-08-62, Box 40, 238.87 m to 244.21 m)

A weak, clay-rich and heavily altered interval logged as a shear zone

The same drillcore interval, but now represented as a loose, clay-

rich mass, as seen by AMEC during its February 2012 site visit




Figure 10: An Example of the Post-Drilling/Pre-Logging Condition of the Core (top photolog) and the Much-Broken, Post Hand-Split Condition of the Same Core, as Seen During AMEC’s February 2012 Site Visit (bottom photolog) (Drillhole K-01-14, Box 21, 137.25 m to 142.69 m)

6.1 Caveability of the Mineralized Mass

For the reasons earlier outlined, the natural caveability of a deposit defines the ability of an orebody to cave freely and spontaneously, once undercut to sufficient dimensions and without preconditioning. Because all rockmasses can cave naturally if the free/undercut dimensions are sufficient for the prevailing rockmass conditions, the purpose of a caveability assessment is, in the first place, to assess the required undercut dimensions to achieve a natural caving condition, and to compare these with the footprint of the target orebody to assess whether the rockmass will cave smoothly and progressively, without excessive dilution being incurred and/or without the need for rockmass preconditioning.




6.1.1 An Empirical Assessment

Given the moderate depth of the target mineralized mass, Laubscher’s empirical approach was used to assess its natural caveability and the likely undercut dimensions required to ensure caving to the target elevation:

• estimated MRMR values were compiled by AMEC from a combination of –

− qualitative compressive strength assessments of the various different dominant rock types (principally, variously altered quartz monzonite with occasional andesite and dioritic intervals) that were carried out during AMEC’s examination of the previously listed drillcores,

− assessments of average drillcore recoveries, based on the logged values, as detailed in the drillcore-relevant logs and confirmed by cross-references to drillcore-relevant photologs (they are very rarely less than 95% and typically approximate to 99% in the volcanic sequence and its intrusives),

− assessments of average RQDs, based in part on the logged values, as detailed in the drillcore-relevant logs and confirmed by cross-references to drillcore-relevant photologs (they are rarely less than 80% and typically average 82% to 85% in the volcanic sequence and its intrusives), and

− estimated average fracture frequencies for the same drillcores, as suggested by the drillcore-relevant photologs supplied by Serengeti;

• adjustment factors of 1.0 were applied for weathering (no slaking expected in time period which will promote caving), joint orientation (no orientated data available), mining induced stress (due to the moderate depth of the mineralization and a lack of orientated discontinuity data to assess whether stress might drive shear failure of blocks) and blasting (which is not applicable in naturally caving rockmasses); and

• the adjusted MRMR values were plotted on Laubscher’s caving chart (from Bartlett, 1998 and in Brown, 2007) to estimate the hydraulic radius at which natural caving might occur (Figure 11).

Table 3 summarizes the assessed range of adjusted MRMR values for the Central Zone mineralized mass. The range of hydraulic radii determined from Laubscher’s caving chart are also presented, along with the upper-bound equivalent square dimensions that Laubscher’s empirical method suggests are required for natural caving to occur.




Table 3: A Summary of AMEC’s Caveability Inputs and Outcomes, Central Zone, Kwanika Property

Adjusted

MRMR Range

Forecast Dimensions for Natural Caving Estimated

Hydraulic Radius Equivalent Maximum

Square Dimension

30 to 65

15 to 39

156 m by 156 m

Figure 11: Laubscher’s Caving Chart (from Bartlett, 1998), with the Estimated Central

Zone MRMR Values Highlighted in Orange

6.1.2 Discussion

It may be seen from Table 3 that the maximum estimated hydraulic radius (39) yields an equivalent maximum square dimension for the onset of caving of 156 m by 156 m for the mineralized mass.




Preliminary and provisional analysis suggests that the average, best cut footprint of the target mineralized zone above the provisionally planned 680 m level approximates to 255 m in the north-south direction (i.e. between geological sections) and approximately 125 m in the east-west direction. This yields a hydraulic radius of approximately 42, which exceeds the worst case minimum requirements of 39 (Table 3). However, depending on the geometry of the caving front, a stable rockmass configuration could develop. For example, if a 70º caving angle is assumed the footprint would be insufficient to ensure continued natural caving of the mineralized mass to the target elevation, insofar as a stable rockmass condition could develop at around 150 m above the undercut.

Preliminary and provisional analysis suggests that the average, best cut footprint of the target mineralized zone between the provisionally planned 480 m and 680 m levels approximates to 255 m in the north-south direction and approximately 150 m in the east-west direction. If this is the case, the same general findings as for natural caveability of the mineralized mass above 680 m level apply.

6.1.3 Comments

The results presented above are based on Laubscher’s empirical method that, as earlier outlined, should be used with caution and as a preliminary tool only for estimating rockmass caveability, especially when the assessed MRMR value exceeds 50. This key point is emphasized because scrutiny of drillcore K-08-62 in particular shows that the main mineralized zone contains numerous structures that in the drillhole log are recorded as shear zones and are invariably associated with strong sericite alteration, as well as being characterized by well-developed clay alteration. The clay appears to be a swelling type, as suggested by the much degraded condition of the shear zone intersections (for example, see Figures 10 & 11 above) and the behaviour of the clay when mixed with water.

The balance of available information suggests that the shear zones are steeply inclined or even vertical. Despite this, general industry experience suggests that their presence will strongly influence rockmass caveability, thereby reducing the undercut dimensions and effective spans required to initially induce and then maintain smooth and progressive caving (Brown 2007, Chitombo 2010). This potential benefit cannot, however, be assessed at the current level of investigation, due in part to a lack of detailed geotechnical information on which such analyses depend. For example, the required information includes the frequency, orientation and distribution of dominant discontinuity sets, inclusive of faults, from which detailed computer simulations of rockmass caveability can be compiled (which approach reflects current industry practice, as earlier outlined). Such considerations fall outside the scope of the preliminary investigations reported here.




It should also be emphasized that there appears to be a general increase in rockmass strength, both in terms of a lessening discontinuity frequency and an increase in compressive strength towards and within the western extension zone. Insufficient information is currently available to define geotechnical domains where different bulk rockmass properties might apply, as well as to establish the reasons for the differences between each domain. However, it can be stated that the higher, estimated MRMR values apply in the western extension zone where the mineralized mass appears to be less severely altered and where MRMR values in excess of 50 most often apply.

If the assumption that a maximum MRMR value of 50 applies in the main mineralized zone, the maximum hydraulic radius reduces to approximately 27 and the equivalent maximum square dimension to ensure natural caving reduces to 104 m by 104 m, which is well within the projected footprint of 125 m by 255 m.

It may be concluded that, according to Laubscher’s empirical method, natural rockmass caving is possible to cave heights of approximately 100 m, or at least extensive rockmass preconditioning would probably not be required. However, the potential for natural caveability decreases as the cave height increases, to the point where a risk of a frozen cave might exist for cave heights in excess of approximately 150 m. It is in such circumstances that rockmass preconditioning would be required.

If the influence of the shear zones on cave propagation is added to the mix, then the confidence in the natural caveability of the main mineralized zone increases. However, an element of insinuated risk nevertheless remains for cave heights much in excess of approximately 150 m, due to the limited footprint of the main mineralized zone. This gap would need to be closed going forward, by means of detailed geotechnical investigations and rockmass modelling. A start could, however, be made by identifying and zoning those areas influenced by the dominant alteration types, inclusive of potassic, sericitic and sodic alteration.

If the high-end values of the estimated MRMR range for the Central Zone rockmass (30 to 65) apply across the western extension zone, it would not necessarily be problematic because its footprint is not constrained in the east-west direction. Provisional and preliminary analysis suggests that sufficiently large undercuts could be developed to induce a natural cave. However, their timing and nature might reasonably be questioned, due to the generally competent and widely jointed nature of the rockmass. A few minor fault- and shear-zones were noted in the core, but their number is probably too small to significantly affect the western extension’s natural caveability (dynamic and intermittent caves may reasonably be expected). Rockmass preconditioning would, therefore, almost certainly be required before block cave mining could successfully be employed.




6.2 Fracturation of the Mineralized Mass

In general, the higher-grade mineralized mass is widely jointed (average joint spacings >2.0 m); only occasionally is it moderately jointed (average joint spacings 0.6 m to 2.0 m) and only infrequently is it closely jointed (average joint spacings <0.6 m) (for example, see Figure 12). However, the same rockmass contains irregular micro-veining that is generally but variously developed. Four veinlet infilling types were identified:

• as may reasonably be expected (given the fact that the deposit is of the porphyry copper-gold type), the most common type is a strongly developed, grey quartz vein stockwork and quartz+sulphide microveining and fracture network - densities typically vary up to 40% by volume of affected rock, with an average of approximately 25%, as defined in the drillhole geology logs supplied by Serengeti;

• the second most common infilling type, or at least what appears to be a commonly developed vein infilling type (that, according to Serengeti’s drillhole geology logs, can comprise up to 20% of an veined drillcore intersection, sometimes more), is albite (a felsic plagioclase feldspar) that is probably associated with a sodic alteration phase, either as a (conditioning) precursor to, or as a contemporaneous event of, the main sulphide mineralizing phase (sodic alteration, mainly as secondary albite, is associated with potassic alteration in some porphyry Cu-Au deposits [Sinclair, 2007]);

• calcite infillings are evident along strongly irregular and very thin microveinlets that are locally very well developed in what appears to be closely developed (brittle) fracture networks; and

• gypsum is present as a locally well-developed infilling along planar to undulating joints, which infilling type can be three to four millimetres thick.




Figure 12: An Example of the Widely and Occasionally Moderately to Closely Jointed

Host Rockmass of the Main Mineralized Zone, Central Zone, Kwanika Property (note the 1.0 m long steel rules) (drillhole K-08-62, 238.37 m to 273.14 – top of section at top left, photologs supplied by Serengeti)




The presence of micro-veinlets of the types outlined were found to strongly influence the fracturation characteristics of the drillcore intersections of mineralized material that were examined during AMEC’s February 2012 site visit. For example:

• where the host rocks were observed to be, and were logged as, extensively and deeply affected by potassic alteration, such as in the main mineralized zone, the core tends to break down readily into small blocks where albite, calcite and especially gypsum infillings are present; however

• the same cannot be said of mineralized material from the western extension zone - not only is the rockmass much more competent and less disturbed, but the bonding strengths of especially albite infillings with the wallrocks is much higher too (much more energy is required to break the drillcore that can remain intact, even when hit hard with a geology hammer).

In some respects the difference in bonding strength between the main mineralized zone and the western extension zone may reasonably be expected: depending on the sequence of alteration events, mineral bondings can be comparatively much lower in severely potassic/sodic altered mineralized core. However, there is some evidence to suggest that the reduction in bonding strength might be the result of a sericitic alteration overprint that selectively affected parts of the mineralized mass.

Whatever the case, knowledge of the distribution and sequencing of the main and secondary alteration events is central to the understanding of the fracturation characteristics of the mineralized mass. This is emphasized because the balance of available evidence suggests that the main mineralized zone might consistently have good fracturation characteristics, whereas the western extension zone might consistently have poor fracturation characteristics. As such, and irrespective of any considerations of natural caveability and the undercut dimensions required to promote a natural cave, rockmass preconditioning in the western extension zone in particular might be required to ensure both a smooth and progressive cave and a block size distribution that is consistent with productive and cost efficient block cave mining.

6.3 Caveability of the Overburden Rockmass

It is established in Section 3 that the western extension zone is overlain by:

• either variable thicknesses of unmineralized to weakly mineralized monzonites, dioritic monzonites and andesites; or




• an unmineralized and layered sequence of competent sandstones, fissile shales and macro-bedded to massive conglomerates (the contact between the target mineralization’s host rockmass and the sedimentary sequence is marked by an erosional surface/unconformity that also marks the truncated western margin of the main mineralized zone).

Planning and design for least risk mining requires that in either case, the overburden rocks should be allowed to, or made to, smoothly and progressively cave, thereby to preclude the possibility of a significant air gap / void developing above the caved mineralized mass, as it is progressively drawn down. If a void was to develop then the possibility of an airblast could not be ruled out in the event of a dynamic failure of the overburden mass. An exception could, however, be made if it could conclusively be shown that the overburden mass was inherently stable over the effective span at the final position of the cave front.

In AMEC’s experience, the fissile shales are likely to cave readily and to degrade quickly to a mass of small-sized blocks. Depending on the thickness of the sandstone beds, they too might cave readily. However, the macro-bedded to massive conglomerates could hang up, or at least dynamically and intermittently fail as large slabs.

The impact of the sedimentary sequence on design and planning considerations depends heavily on the details of its stratigraphy above the target mineralized mass and/or its position compared to the top contact of the host rocks for the target mineralization. Such details were not available at the time of AMEC’s preliminary investigations reported here. This gap would have to be closed during the pre-feasibility level of project development, when such details will become important within the scope of design and planning for least risk mining.




7.0 REFERENCES

Bartlett, P.J., 1998. Planning, implementation, operation and monitoring of a cave mining method with coarse fragmentation with reference to cave mining at Premier Diamond Mine. PhD thesis (unpublished), University of Pretoria, South Africa.

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