TECHNICAL REPORT (NI 43-101) - ENGLISH - InfoMine

FEASIBILITY STUDY – ESTRADES PROJECT
TECHNICAL REPORT 43-101
PROJECT N o AV105365
PREPARED FOR
COGITORE RESOURCES INC
GENIVAR Limited Partnership
152, Murdoch Avenue
Rouyn-Noranda (Quebec)
J9X 1E1
Telephone : (819) 797-3222
Fax : (819) 762-6640
Rouyn-Noranda, October 24 th , 2008

Feasibility Report
1. COVER PAGE
2. TABLE OF CONTENT
TABLE OF CONTENTS
Page
3. EXECUTIVE SUMMARY..................................................................................... 1
3.1 Property Description, Location and Access .............................................. 1
3.2 Regional and Property Geology, Deposit Types and Mineralization......... 1
3.3 Mineralization ........................................................................................... 2
3.4 Exploration................................................................................................ 4
3.5 Metallurgical Testwork.............................................................................. 5
3.6 Mineral Resources and Reserves............................................................. 6
3.7 Mining....................................................................................................... 8
3.8 Mineral Processing ................................................................................. 10
3.9 Site Infrastructure ................................................................................... 10
3.10 Environmental Considerations................................................................ 11
3.11 Conclusions and Recommendations ...................................................... 11
4. INTRODUCTION............................................................................................... 13
5. RELIANCE ON OTHER EXPERTS................................................................... 14
6. PROPERTY LOCATION AND DESCRIPTION ................................................. 15
6.1 Property Location ................................................................................... 15
6.2 Property Description and Ownership ...................................................... 17
7. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE
AND PHYSIOGRAPHY ..................................................................................... 19
7.1 Accessibility ............................................................................................ 19
7.2 Climate ................................................................................................... 21
7.3 Local Resources..................................................................................... 21
7.4 Infrastructure .......................................................................................... 21
7.5 Physiography.......................................................................................... 22
8. HISTORY .......................................................................................................... 23
8.1 Exploration.............................................................................................. 23
8.2 Mineral Resource and Mineral Reserve Estimates................................. 27
8.3 Production .............................................................................................. 28
9. GEOLOGICAL SETTING .................................................................................. 30
9.1 Regional Geology ................................................................................... 30
9.2 Property Geology.................................................................................... 32
10. DEPOSIT TYPES.............................................................................................. 39
11. MINERALIZATION ............................................................................................ 40
11.1 Main Zone............................................................................................... 40
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11.2 Central Zone........................................................................................... 44
11.3 East Zone ............................................................................................... 44
11.4 West Zone .............................................................................................. 45
12. EXPLORATION................................................................................................. 46
12.1 Year 2001 ............................................................................................... 46
12.2 Year 2002 ............................................................................................... 47
12.3 Year 2003 ............................................................................................... 48
12.4 Year 2004 ............................................................................................... 49
12.5 Year 2005 ............................................................................................... 49
12.6 Year 2006 ............................................................................................... 51
12.7 Year 2007 ............................................................................................... 53
13. DRILLING.......................................................................................................... 55
14. SAMPLING METHOD AND APPROACH.......................................................... 57
15. SAMPLE PREPARATION, ANALYSES AND SECURITY................................. 58
16. DATA VERIFICATION....................................................................................... 59
17. ADJACENT PROPERTIES ............................................................................... 60
18. MINERAL PROCESSING AND METALLURGICAL TESTING.......................... 61
18.1 Laboratory Testwork Prior to Ore Milling in Matagami............................ 61
18.2 Ore Milling in Matagami.......................................................................... 62
18.3 2007 Laboratory Testwork by SGS Lakefield Research Ltd ................... 64
19. MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES .................... 70
19.1 Introduction............................................................................................. 70
19.2 Resource Estimate for the Old Mine, the Main and Main East
Zones...................................................................................................... 71
19.3 Resource Estimate in the Central Zone.................................................. 72
19.4 Reserve Estimate ................................................................................... 76
20. OTHER RELEVANT DATA AND INFORMATION............................................. 82
21. INTERPRETATION AND CONCLUSIONS ....................................................... 83
22. RECOMMENDATIONS ..................................................................................... 85
23. REFERENCES.................................................................................................. 86
24. DATE AND SIGNATURES PAGE WITH CERTIFICATES OF CONSENTS ..... 88
24.1 Date and Signatures Page...................................................................... 88
24.2 Certificate and Consent of Michel Garon, Eng........................................ 89
24.3 Certificate and Consent of Nicole Rioux, Geo. ....................................... 90
24.4 Certificate and Consent of Mario Blanchette, Eng. ................................. 91
24.5 Certificate and Consent of Marc Lafontaine, Eng. .................................. 92
24.6 Certificate and Consent of Serge Ouellet, Eng. Ph.D. ............................ 93
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25 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON
DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES.............. 94
25.1 Mining..................................................................................................... 94
25.2 Mineral Processing ............................................................................... 110
25.3 Site Infrastructure ................................................................................. 111
25.4 Manpower............................................................................................. 126
25.5 Environmental Considerations.............................................................. 127
25.6 Markets and Contracts.......................................................................... 128
25.7 Financial Analysis................................................................................. 129
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LIST OF TABLES
Page
Table 3-1 Results Obtained with LCT-2 and Calculated by GENIVAR .................. 5
Table 3-2 Resource Estimate ................................................................................ 7
Table 3-3 Estimate of Probable Reserves ............................................................. 7
Table 3-4 Ore Production Over the Mine Life ...................................................... 10
Table 12-1 Exploration – Best Intersections .......................................................... 54
Table 18-1 Matagami Results with the Copper Concentrate of the
Estrades Ore........................................................................................ 63
Table 18-2 Matagami Results with the Zinc Concentrate of the Estrades Ore....... 63
Table 18-3 Ore Mineral Modal Abundance ............................................................ 65
Table 18-4 Results Obtained for the Locked Cycle Test LCT-2............................. 68
Calculations by SGS Lakefield............................................................. 68
Table 18-5 Results Obtained for the Locked Cycle Test LCT-2............................. 68
Calculations by GENIVAR ................................................................... 68
Table 19-1 Resource Estimate for the Old Mine, Main and Main East
Zones, by Scott Wilson Roscoe Postle ................................................ 71
Table 19-2 Resource Estimate in the Central Zone by GENIVAR ......................... 73
Table 19-3 Estimate of the Probable Reserves in the Old Mine ............................ 76
Table 19-4 Remaining Indicated Resources in the Old Mine ................................. 76
Table 19-5 Estimate of the Probable Reserves in the Main Zone.......................... 78
Table 19-6 Remaining Indicated Resources in the Main Zone .............................. 78
Table 19-7 Estimate of the Probable Reserves in the Central Zone ...................... 81
Table 19-8 Remaining Indicated Resources in the Central Zone........................... 81
Table 25-1-1 Diesel Mobile Equipment Used Underground .................................... 104
Table 25-1-2 Underground Manpower for Development and Production................ 105
Table 25-1-3 Total Ore Production Over the Mine Life (Tonnes Milled).................. 106
Table 25-1-4 Ore Production from Development and Stoping (Tonnes Milled)....... 107
Table 25-1-5 Waste from Development (Tonnes) ................................................... 107
Table 25-1-6 Unit Costs for Development............................................................... 108
Table 25-1-7 Unit Costs of Various Development Supplies .................................... 108
Table 25-1-8 Direct Mining Costs............................................................................ 109
Table 25-3-1 Capital Costs for the Site Infrastructure to Start Operations .............. 125
Table 25-4-1 Total Number of Employees Required on Site................................... 126
Table 25-7-1 Pre-Tax Cashflow of the Estrades Project ......................................... 132
Table 25-7-2 Sensitivity Analysis for the Estrades Project...................................... 134
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LIST OF FIGURES
Page
Figure 6-1 Location Map.................................................................................... 16
Figure 6-2 Property Blocks ................................................................................ 18
Figure 7-1 Acces Road to the Property.............................................................. 20
Figure 7-2 Current Mine Site Infrastructure ....................................................... 22
Figure 9-1 Regional Geology............................................................................. 31
Figure 9-2 Property Geology ............................................................................. 36
Figure 9-3 Stratigraphy – Plan View .................................................................. 37
Figure 9-4 Stratigraphic Column........................................................................ 38
Figure 11-1 Longitudinal Section ......................................................................... 42
Figure 11-2 Vertical Section 2400 W ................................................................... 43
Figure 18-1 Flowsheet for the Locked Cycle Test LCT-2..................................... 69
Figure 19-1 Section 1750W of the Central Zone ................................................. 74
Figure 19-2
Outline of Resources for the South, Center and North
Areas of the Central Zone ................................................................ 75
Figure 19-3 Outline of Probable Reserves for the Old Mine ................................ 77
Figure 19-4 Outline of Probable Reserves for the Main Zone.............................. 79
Figure 19-5 Outline of Probable Reserves for the Central Zone.......................... 80
Figure 25-1-1 Main Zone Longitudinal 3D View...................................................... 97
Figure 25-1-2 Main Zone Elevation 4 749 m........................................................... 98
Figure 25-1-3 Main Zone Elevation 4 637 m........................................................... 99
Figure 25-1-4 The Main and Central Zone............................................................ 102
Figure 25-1-5 The Central Zone and Access Ramp.............................................. 103
Figure 25-3-1 General Arrangement of the Site Layout ........................................ 113
Figure 25-7-1 Cashflow Sensitivity........................................................................ 135
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LIST OF APPENDIX
Appendix 1 :
Appendix 2 :
Detailed Calculations by GENIVAR for Locked Cycle
test LCT-2
CIM DEFINITION STANDARDS - For Mineral Resource
and Mineral Reserve Mineral Resource
Appendix 3 :
Abbreviations
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3. EXECUTIVE SUMMARY
3.1 Property Description, Location and Access
The property consists of three (3) distinct blocks: the Estrades Block, the
Newiska Block, and Estrades Mining Lease #795 (ML 795). The Estrades and
Newiska Blocks comprise a total of 138 claims and cover an area of 3,167.35
ha.. The Estrades Block comprises 101 claims covering 1,595.20 ha in Orvilliers,
Estrades and Estrées townships. Most of the claims of the Estrades Block are
contiguous with Mining Lease #795.
The Estrades-Newiska property is located in Northwestern Quebec
approximately 95 km NE of the town of La Sarre and 35 km WNW for the former
village of Joutel (figure 6.1). The property is situated in the northern part of the
southwestern corner of NTS quadrangles 32E10 and is bounded between
longitudes 78°41’ W and 78°58’W, and latitudes 49°33’N and 49°37’N. The
property comprises a total area of 3,319.14 ha..
The property is accessed from the village of Authier North via an all season
public gravel road which runs from the villages of Authier Nord to the former
village of Joutel (Authier Nord-Joutel Road). The mine site is 35 km northwest of
the public road. Since mine closure, the road to the mine site is not open during
winter.
3.2 Regional and Property Geology, Deposit Types and Mineralization
The rocks of the area are constituted of meta-volcanic and sedimentary rocks of
the Harricana-Turgeon Belt (HTB) which is located in the NW part of the Abitibi
Subprovince. The regional metamorphism is of greenschist facies. Rocks are
east-west striking and vertically dipping. Four regional lithostratigraphic domains
are recognized in the area and they are bounded to the north by the Orvilliers
Pluton and to the south by the Mistaouac pluton. A major regional deformation
zone, the Casa-Berardi Break, is located approximately two kilometres north of
the property limit within the Taïbi sediments.
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The Estrades deposits area is underlain by a succession of east-west striking,
steeply dipping, Archean meta-volcanic and meta-sedimentary rocks. Most
stratigraphic units are intruded by later felsic and mafic dykes and sills.
Stratigraphy is interpreted to face south, based solely on the occurrence of the
alteration zone and stringer mineralization on the north side of the massive
sulphide deposit.
The Estrades deposit shows similarities with other Archean volcanogenic
massive sulphide (VMS) deposits and is considered as such (Welch, 1995).
The Estrades Unit (EU), a polymetallic (Zn, Cu, Pb, Ag, Ag, Au) VMS deposit
unit, is hosted in altered felsic rocks referred to as the “Main Felsic Unit (MFU)”
(Welch, 1995). This mineralization has been identified as classic syngenetic
exhalative type. The EU and the MFU units strike almost east-west, over several
kilometres, with a nearly vertical dip. Base metal and gold mineralization is found
on the property in four VMS lenses, namely the Main, Central, East and West
Zones.
3.3 Mineralization
Owners and Work Done
In 1984, Golden Hope Mines Ltd of the Noramco Group of companies (Noramco)
acquired a large block of claims covering a west-northwest trending swarm of
long airborne electromagnetic (EM) conductors. A deal was signed with Teck
Exploration Ltd. (Teck), whereby Golden Hope provided the property and the
exploration funds, and Teck provided its technical and managerial expertise.
In 1985, line cutting, several geophysical surveys and drilling of reverse
circulation (RC) drill holes as well as of 31,966 m of diamond drill holes were
completed. Diamond drill targets were selected entirely on the basis of
geophysical anomalies. Eight such anomalies were selected and the third to be
drilled was the one corresponding to the Estrades deposit (hole H-003: 15% Zn,
3% Cu, 0.2 oz/t Au, 9 oz/t Ag over 35 ft).
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Between 1985 and 1988, following the discovery hole, Teck completed 77,000
metres of diamond drilling, conducted numerous geophysical programs on the
deposit and this led to the detailed drilling of the Main Zone and the discovery of
the West, Central and East Zones.
Breakwater Resources Inc. (Breakwater) became involved in the project in 1988
when it gained control of Noramco. At that time, the Estrades property was
owned by Noramco affiliates Golden Hope Resources Inc (40%) and Golden
Group Exploration Inc (60%). Breakwater earned a 20% undivided interest in
the Estrades deposit by completing a feasibility study on the Estrades deposit
and incurring expenditures of no less than $3 million. The agreement also
granted Breakwater the option to earn an additional 50% interest by making a
cash payment of $0.5 million to Golden Hope-Golden Group and bringing the
property into production. A feasibility report was completed by Wright Engineers
Limited for Breakwater. That report addresses the “Ore Reserve Estimate –
Phase 1” of the Estrades deposit.
In February 1990, Breakwater exercised its right of earning a full 70% interest in
the mine and formed a JV with Golden Hope-Golden Group (Breakwater 70%,
Golden Group 18%, Golden Hope 12%). The mine was then operated as a JV
until suspension of operations in June 1991. No work other than technical and
engineering studies has since been carried out at the mine.
Due to low metal prices and exchange rates, the mine was subsequently allowed
to flood and kept on a care and maintenance basis.
In 1994, Arimetco International Inc. (Arimetco), a United States copper producer,
assumed full management of Breakwater with the ultimate intent of merging with
Breakwater. But in 1995, it notified Breakwater that it no longer intended to
complete the merge. Arimetco and Breakwater negociated for the Breakwater
debt. Breakwater’s interest in the Estrades Mine and Mining Lease #795 was
thus transferred to Arimetco as full settlement of that debt. In 1996, Arimetco
declared bankruptcy.
After a series of transactions, Atlas Precious Metals Inc. (Atlas), became in 2003
the owner of Breakwater’s original 70% interest in Estrades which Cogitore (then
Woodruff Capital Management Inc.) purchased in 2005. Cogitore then also
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purchased from Orvilliers Resources Ltd. the remaining 30% interest (June 30,
2005 agreement). Through these two agreements, Cogitore now owns a 100%
interest in Mining Lease #795.
Mineral Resource and Mineral Reserve Estimates
Mineral resource and reserve estimates were carried out by Teck (1989),
Noramco and Breakwater (1989), Wright Engineers Ltd. (1989), Breakwater
(1992), Derry Michener Booth and Wahl (DMBW, 1997), SRK (2002), Western
Range (2003) and Scott Wilson Roscoe Postle (2006).
In 2006, Scott Wilson Roscoe Postle estimated for the Main and Main-East
Zones total indicated mineral resources of 592,000 tonnes grading 9.82 %Zn,
0.81 %Cu, 5.21 g/t Au, 168.0 g/t Ag and 0.90 %Pb.
Production
From August 1990 to June 1991, the ore was milled, on a custom-milling basis,
at the Matagami mill which is located 128 km from the mine. At that time, the
Matagami mill was operated by Noranda Minerals Inc. A total of 174,946 tonnes
at an average grade of 12.93% Zn, 1.14% Cu, 6.35 g/t Au and 172.30 g/t Ag are
reported to have been milled.
3.4 Exploration
Between 2002 and 2006, various exploration programs were carried out on the
property consisting of geophysical surveys, diamond drilling and geological
compilation, as reported by Cloutier (2005).
Thirty-five (36) diamond drill holes for a total of 16,146 m were drilled in 2001,
2005, 2006 and 2007. All holes were drilled for gold and base metal exploration.
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3.5 Metallurgical Testwork
Prior to ore milling in Matagami, laboratory testwork started out in early 1987 and
was undertaken by Lakefield Research and Bacon Donaldson and Associates.
From August 1990 to June 1991, the Estrades ore was milled at the concentrator
of the Matagami Division of Noranda Minerals Inc. Concentrates of copper and
zinc were produced. The copper concentrate contained high amounts of zinc
and lead and the copper grade was low as the strategy was to reduce the copper
grade in order to increase gold recovery.
In early 2007, on behalf of Cogitore Resources Inc., GENIVAR requested SGS
Lakefield to initiate a metallurgical test program intended to apply previous
findings to the development of a new flowsheet. The program aimed to improve
the level of understanding of the Cu, Pb and Zn circuits, and optimization focused
on improvements in selectivity between all minerals and improving grade and
recovery. A total of 28 batch flotation and 2 locked cycle tests were completed,
exploring both sequential and bulk Cu-Pb flowsheets. The results obtained
suggested that the fully sequential approach (Cu/Pb/Zn) offered the most promise
for the Estrades ore compared to the bulk flotation (Cu+Pb/Zn).
The results obtained are presented in the table below:
Table 3-1
Results Obtained with LCT-2 and Calculated by GENIVAR
Product Weight Assays, %, g/t % Distribution
% Cu Pb Zn Au Ag Cu Pb Zn Au Ag
Cu Conc 1.7 20.6 4.7 13.7 82.6 2522 60.2 6.7 1.9 32.6 21.7
Pb Conc 1.8 2.1 42.1 15.6 74.8 2227 6.6 62.1 2.2 30.9 20.1
Zn Conc 20.7 0.5 0.9 56.0 3.2 215 16.0 15.7 89.8 15.0 22.1
Zn 1st Cl Tail 7.4 0.7 0.8 6.8 3.9 318 8.1 4.9 3.9 6.5 11.6
Zn Rghr Tail 68.3 0.1 0.2 0.4 1.0 73 9.1 10.6 2.3 15.0 24.5
Head 100.0 0.6 1.2 12.9 4.4 202 100 100 100 100 100
Head-calc 0.7 1.3 12.7 4.8 209
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3.6 Mineral Resources and Reserves
For the purpose of this study, the 43-101 technical report done by Scott Wilson
Roscoe Postle Associates Inc.(SWRPA), on the resource estimate for the Old
Mine, the Main and Main East Zones, was used. This report was completed in
November 2006.
For the SWRPA evaluation, original assays were used for interpretation of the
mineralized envelope. Zn, Cu, Au and Ag grades were converted into dollar
values and a cut-off grade of $120/t NSR was used. No value was given to the
Pb in the ore. A minimum width of 2 metres was used for the interpretation of the
envelope.
For the resource estimate of the Central Zone done by GENIVAR, the
mineralization outline and the faults position are based upon the geological
interpretation provided by Cogitore and reviewed by GENIVAR. The minimum
horizontal width was established at 2 m and the vertical extension of the mining
blocks was limited to 25 m. The distance between sections was set at 25 m. But
because the drilling pattern is not regular on these sections, a weighting factor
was applied to determine the tonnage of blocks which were too far apart (over 25
m of distance). The estimate was done within the volume determined by the
geological interpretation.
Table 3-2 indicates the results obtained by Scott Wilson Roscoe Postle and
GENIVAR.
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Table 3-2
Resource Estimate
Tonnes Zn % Cu % Pb % Au g/t Ag g/t
INDICATED
Old Mine* 54,000 13.06 0.88 1.35 7.64 196.00
Main Zone,190-415* 337,000 10.71 0.79 1.02 6.01 186.00
Main Zone, below 415* 170,000 8.43 0.54 0.64 2.87 144.00
Main East Zone* 31,000 1.97 2.38 0.31 5.03 52.00
Central Zone** 148,500 8.13 1.13 0.75 4.10 122.38
Total Indicated 740,000 9.48 0.87 0.87 4.99 158.88
INFERRED
Main Zone* 4,000 8.69 0.54 0.52 1.88 150.00
Main East Zone* 28,000 6.88 0.31 0.40 2.50 60.00
Central Zone** 18,700 5.40 0.75 0.89 1.61 60.26
Total Inferred 50,700 6.49 0.49 0.59 2.11 68.30
* by SWRP ** by GENIVAR
GENIVAR established a mine plan for the Old Mine, the Main Zone between
levels 190 and 415 and the Central Zone. For both the Main and Central Zones,
shrinkage stoping is used, following the vein width, with no dilution. Long-hole
stoping was selected for the Old Mine and a 10% dilution factor was considered.
Table 3-3 shows the results obtained for the ‘’probable’’ reserve estimate.
Table 3-3
Estimate of Probable Reserves
Tonnes Zn % Cu % Pb % Au g/t Ag g/t
PROBABLE
Old Mine 51,900 11.75 0.80 1.22 6.90 177.60
Main Zone,190-415 275,050 10.43 0.81 1.01 5.60 186.94
Central Zone 124,950 7.54 1.31 0.71 3.89 127.05
Total 451,900 9.78 0.95 0.95 5.28 169.31
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3.7 Mining
Mine Dewatering and Rehabilitation
A period of about 4 months of work has been estimated for dewatering at a cost
of $850,000. Mine rehabilitation will be carried out as the ramp becomes
accessible and therefore will be undertaken during dewatering. It is assumed
that rehabilitation will start about a month after the beginning of dewatering and
will be completed a month after dewatering is finished. Rehabilitation costs have
been established at $2,250,000 for 1,500 m of ramp.
Ramp Development
A ramp with a 4.0 m height and 5.0 m width, will be driven at a grade of 15% and
will go from the 190 level down to the 415. A total of 14 months have been
planned to complete its excavation.
Mining Methods
The mining method selected for both the Main and Central Zones is shrinkage
stoping. The ore outline with a minimum width of 2.0 m, a dip nearly vertical and
a good linearity along strike make it appropriate for shrinkage stoping.
For the Main Zone, every 56.0 m, an access will be driven from the ramp to the
Main Zone. From this access, drifts will be excavated in the mineralization at
vein width and a height of 2.4 m. Each stope will stop 5.0 m from the level
above. Vertical pillars of 4.0 m will also be left between them when the quantity
of ore is too voluminous for a single stope. No waste backfilling will be required
since the stopes’ walls will be supported as mining progresses. But backfilling
could be undertaken and would be more economical than hauling waste to
surface. This issue could be examined at the outset of the project. Mucking is
planned to be carried out with 5 yard 3 scooptrams and 26 ton trucks. About 82%
of the total indicated resources will be extracted with no dilution. A total of
275,050 tonnes of ore will be mined from the Main Zone.
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For the Central Zone, only 4 stopes will be mined in that area and mining will be
similar to what will be done in the Main Zone, following the width of the vein, no
dilution, drifts with a height of 2.4 m, etc… A total of 124,950 tonnes of ore will
be mined which represents 84% of the total indicated resources with no dilution.
In order to reduce the ventilation requirements, ore handling and hauling in this
area will be carried out with equipment on track. Two drifts with a height of 2.4 m
and a width of 2.8 m will be excavated from the Main Zone area to the Central
Zone and will allow for the access as well as ventilation of that part of the
underground operation.
The long-hole method has been selected to recover the ore left in the Old Mine
as most of the development is apparently complete to suit that mining method. It
will be mined in 30.0 to 35.0 m blocks separated by 5.0 m pillars. A total of
51,900 tonnes of ore will be mined and this includes a 10% dilution.
Ventilation
The air volume necessary for this underground operation has been evaluated in
compliance with Quebec regulations and particularly, the minimum requirements
for diesel engines. Therefore, an existing raise will be extended to the last level
and this will also provide an emergency exit in compliance with the current
regulations. A new raise will also be excavated nearby the ramp (circular, 11.5’
dia., 250 m long). These raises will supply fresh air to the underground workings.
A total of $1.725 M has been estimated for extending the existing ventilation raise
and drive the new one. Air will be exhausted through the ramp. The total air
requirement amounts to 310,000 cfm @ 13.5 ‘’ w.g. and 880 bhp.
Production and Costs
Development in mineralization will start about 12 months after the beginning of
the project, after completion of U/G dewatering and rehabilitation, and stoping will
start producing ore in about the 14 th month. Ore production will gradually ramp
up and the production phase will be reached in the 21 th month of the project with
more than 77% of the planned maximum rate of 20,000 tonnes of ore per month.
The planned maximum rate will be achieved a month later and maintained until
nearly the end of operations. The production schedule is shown in Table 3-4
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below. Direct mining costs will be $76.65/tonne of ore over a nearly three year
period.
Table 3-4
Ore Production Over the Mine Life
Ore Production
Months 7 - 18 19 - 30 31 - 42 Total
Main Zone 27,150 143,300 104,600 275,050
Central Zone 3,150 36,500 85,300 124,950
Old Mine 13,400 38,500 51,900
Total 43,700 218,300 189,900 451,900
3.8 Mineral Processing
On the mine site there is no milling facility and therefore, 6 operations were
approached in order to find one which could treat the Estrades ore on a custom
basis. All of them are located within a 150 km radius but in some cases, require a
travelling distance beyond 300 km.
Finally, two operations demonstrated interest in treating the ore and some
discussions were held with each of them. This allowed to determine preliminary
costs that were used for this study. Both options would require modifications to
the current installations with addition of new equipment. GENIVAR estimated the
costs involved in these modifications.
However, no agreement has been reached with these operations yet and this will
take place when Cogitore Resources will decide to go ahead with the project.
3.9 Site Infrastructure
A site visit was made by GENIVAR during the week of April 23 rd , 2007, to
evaluate the conditions of the existing infrastructure. Following the visit, a list and
description of the new surface infrastructure required to ensure proper operations
were carried out and a cost evaluation has been produced for different options:
overhaul, replacement through purchasing or rental. A total of $7.0 M will be
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required in replacement and overhaul of infrastructure in order to start operations.
The main expenses will be for the electrical infrastructure ($1.9 M) and the
replacement of the current generators ($1.5 M).
3.10 Environmental Considerations
The only major environmental structures remaining on site are the ponds that are
constructed with clay material. After a clean up, ponds will be functional. Ore and
waste rocks are partly potentially acid generating and metals leachable. The
planned waste management approach and infrastructures are designed to handle
these risks safely and according to Quebec’s provincial environmental laws and
regulations.
The project is certified being in compliance with the James Bay municipality
regulations. According to a map provided in 2007 by the Abitibiwinni First Nation
Council, the mine area is not in a zone where traditional hunting and fishing
activities are taking place. Quebec’s Ministry of Environment issued in 2007 a
certificate of authorization for underground mine de-flooding and extraction of
approximately 422,000 tonnes of ore at a rate of 500 tonnes per day. Requests
for other environmental permits (not limited to but such as: wastewater, potable
water, solid waste disposal area, and closure plan) is not done yet.
3.11 Conclusions and Recommendations
The financial analysis shows that the project is marginal for the base case
scenario selected, with a rate of return of 7.4% and a negative net present value
of $1.4 M at a discount rate of 10%. But the sensitivity analysis indicates it has
an interesting economic potential and demonstrates a high degree of sensitivity
to the exchange rate and the zinc price in particular.
The study has been spread over an 18 month period and the cost of many items
has been estimated some time ago. Since then, some of these costs have
changed significantly, including the cost of the diesel fuel which has dramatically
increased recently.
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The recommendations are as follows:
Some further drilling should be done in the Central Zone to increase the
confidence level and optimize stope development;
Laboratory test work should be done to adapt the flowsheet to the
concentrator which would be eventually selected and run the tests with a
head sample having grades closer to the grades of the mineral reserve to
confirm the metallurgical results;
In light of the current zinc and lead prices, it is recommended to delay the
production decision and wait for better prices, particularly for zinc;
An overall cost update should be undertaken before taking the final
decision to go ahead with the project;
The property demonstrates a very good potential for the discovery of other
volcanogenic massive sulphide deposits and therefore, exploration should
continue in that area. The discovery of additional ore zones could
dramatically improve the economics of the Estrades project.
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4. INTRODUCTION
In June 2006, Cogitore Resources Inc. gave GENIVAR the mandate to undertake
a complete feasibility study for the Estrades project in order to determine its
economic viability. This was to be done in the form of a 43-101 technical report.
The data and information required to carry out the study were received from
Cogitore Resources. They provided the geological information with a 3D model
built by Scott Wilson Roscoe Postle as well as info on the past operations, site
infrastructures, metallurgical test work and results from their own archives. A
preliminary evaluation performed by Met-Chem in 2006 was also obtained from
Cogitore Resources.
A site visit was conducted on September 21, 2006, by Mr. Michel Garon, Eng.,
Senior Mining Engineer of GENIVAR and project manager of the current
feasibility study, with Mr. Gérald Riverin, President & CEO of Cogitore
Resources, Mr. Dave Comba, member of the Cogitore Board of Directors, Mr.
Glenn O’Gorman, Consulting Mine Engineer and former manager of the Estrades
Mine and Mr. Bernard Salmon, Consulting Geological Engineer of Scott Wilson
Roscoe Postle.
Other GENIVAR representatives visited the site. Mr. Serge Ouellet, Eng., Ph.D.
Environmental Sciences, made a visit on November 10, 2006, for the
environmental aspect of the project. In April 2007, Mr. Stephan Dupuis, jr.Eng.
and Mr. Sébastien Quévillon, Tech. went to the site to evaluate the conditions of
the access road and in June 2007, Mr. Stephan Dupuis, jr.Eng. returned with Mr.
Carl Jolin, Eng. for the same purpose. During the week of April 23, 2007, Mr.
Jean-François Brouillette, Eng. and Mr. Sébastien Morin, Eng. spent some time
on site to assess the conditions of the mechanical and electrical infrastructures
as well as the buildings.
In early 2007, SGS Lakefield Research Limited was mandated by GENIVAR to
conduct metallurgical test work on drill core samples selected by Cogitore
Resources, in order to optimize the flowsheet while focusing on improvements in
selectivity between minerals and improving grade and recovery.
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5. RELIANCE ON OTHER EXPERTS
Mr. Serge Ouellet, Eng. Ph.D. Environmental Sciences, from GENIVAR, worked
on the environmental aspect of the project and visited the site on November 10,
2006. Mr. Ouellet was responsible for writing the section 25.5 Environmental
Considerations.
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6. PROPERTY LOCATION AND DESCRIPTION
6.1 Property Location
The Estrades-Newiska property is located in Northwestern Quebec
approximately 95 km NE of the town of La Sarre and 35 km WNW of the former
village of Joutel in the southwestern corner of NTS sheets 32E10 and is bounded
between longitudes 78°41’ W and 78°58’W, and latitudes 49°33’N and 49°37’N.
The property comprises a total area of 3,319.14 ha.
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Figure 6-1
Location Map
6.2 Property Description and Ownership
The property consists of three (3) distinct blocks: the Estrades Block, the
Newiska Block, and Estrades Mining Lease #795 (ML 795). The Estrades and
Newiska Blocks comprise a total of 138 claims and cover an area of 3,167.35
ha.. The Estrades Block comprises 101 claims covering 1,595.20 ha in Orvilliers,
Estrades and Estrées townships, and the Newiska Block comprises 37 claims
covering 1,572.15 ha, all located in Estrades Township. Mining Lease #795
covers an area of 151.79 ha. Most of the claims of the Estrades Block are
contiguous with Mining Lease #795 (Figure 4-2). Claims of the Estrades and
Newiska Blocks are not surveyed (unpatented); however, the mining lease was
surveyed by a registered land surveyor. At the time of writing this report, all
claims are in good standing until 2007 and 2008, and credits to keep claims in
good standing are sufficient (Table 24-1, Appendix 1).
Of the 138 claims forming the Estrades and Newiska Blocks, 130 claims are
currently recorded under the name of Inmet Mining Corporation (Inmet), the
others being recorded under the name of Woodruff Capital Management
(Woodruff). The 130 claims owned by Inmet are subject to the April 23, 2004
agreement (the "Inmet Agreement") between Woodruff and Inmet, granting
Woodruff the right to earn a 50% interest in the claims. All of the Inmet claims
are in good standing until 2007 and 2008. Mining Lease #795 is recorded under
the name of Woodruff. At the time of writing this report, the name change from
Woodruff to Cogitore was not yet done in the Ministère des Ressources
Naturelles, de la Faune et des Parcs register (MRNFP – Ministry of Natural
Resources, Wildlife and Parks).
The mining lease covers 151.79 ha and straddles the Orvilliers, Estrades and
Estrées townships, and is adjacent in part to the south boundary of Puisseaux
Township. In the Québec Government records, Mining Lease #795 is shown as
being "active" with an expiry date of January 20, 2012, with no lien attached, and
recorded under the name of Woodruff Capital Management Inc. Cogitore is in
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6.2 Property Description and Ownership
The property owned by Cogitore Resources in the Estrades project consists of
three (3) distinct blocks:
• the Estrades Block: 110 claims totalling 1,747.2 hectares located in
Estrées, Estrades and Orvilliers townships; these claims are subject to a
50% earn-in agreement with Inmet Mining Corporation;
• the Newiska Block, 37 map staked “cells” totalling 1,564 hectares
located in Estrades township; these claims are subject to a 50% earn-in
agreement with Inmet Mining Corporation; and
• the Estrades Mining Lease (Bail #795) located in Estrées, Estrades and
Orvilliers townships and covering 151.79 hectares; Quebec government
records show Lease #795 to be owned 100% by Cogitore Resources Inc.
The location of the Estrades claims and mining lease is shown on Figure 6-2.
This report deals exclusively with the deposit located within Mining Lease #795
and therefore does not cover the Newiska block which has been discussed at
length in the Technical Report prepared by Cloutier (2005).
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Figure 6-2
Property Blocks
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7. ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND
PHYSIOGRAPHY
7.1 Accessibility
The property is accessed from the village of Authier North via an all season
public gravel road which runs from the village of Authier Nord to the former
village of Joutel (Authier Nord-Joutel Road). The mine site is 35 km northwest of
the public road as illustrated in Figure 7-1. Since mine closure, the road to the
mine site has not been opened during winter except during exploration drilling
operations.
Access in winter is also possible along an 8 km long winter bush road that follows
the Puisseaux-Estrées township line and crosses the Wawagosic River before
joining the La Sarre-Selbaie all weather gravel road, 60 km north of Villebois
village.
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7.2 Climate
Climate is characterized by short mild summers and long cold winters with mean
temperatures ranging from –15°C in January to 16°C in July. Peak temperatures
can reach -40°C in the winter and 30°C in the summer.
7.3 Local Resources
The Abitibi region has a long history of mining activity and therefore mining
suppliers and contractors are locally available. The closest town, La Sarre, a
municipality of 7,728 inhabitants (2001 census), supplies most of the workforce
for the Casa Berardi Gold mine currently operated by Aurizon Mines Ltd. (as it
did for former producing mines Selbaie). The Estrades Project enjoys the
support of local communities.
7.4 Infrastructure
At the time of acquisition by Cogitore, the surface and underground infrastructure
at the Estrades Mine included the following:
Underground infrastructure including a ramp and ventilation raise;
Surface infrastructure including a garage, various sheds and trailers
used as workshop, warehouse, administration buildings, and dry
facilities;
Facilities providing basic services to the mine, including electric power,
heat, water treatment and supply, as well as sewage treatment;
Some mining equipment left at mine closure.
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Presently, all building infrastructure is still on site but Cogitore has re-located
some of the equipment to a safe storage place near Rouyn-Noranda. A picture
of the current infrastructure is show below.
Figure 7-2
Current Mine Site Infrastructure
7.5 Physiography
The vertical relief in the area is almost flat with a mean altitude of 275 m above
sea level. Rare hills (10 m to 20 m) are dispersed on the property. The property
area is primarily covered by swamps and black spruce. Overburden consists
essentially of a thick layer (>20 m) of fluvio-glacial till, locally topped by organic
soils. No outcrops have been reported on the property.
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8. HISTORY
8.1 Exploration
History of exploration works and mining is summarized below, largely
taken from Cloutier (2005) and Salmon (2006);
1960: The Federal Government released aeromagnetic maps covering
the general area;
1976: Geophysical coverage (Input Mark VI airborne geophysical
survey) of the area including the present property. The survey was
sponsored by the MRNFP;
1977: Noranda Exploration drilled one hole (77-1) at the eastern end of
the Estrades Block to test an isolated Input short conductor. Hole 77-1
intersected approximately five metres of thin pyrite-pyrrhotite bands at a
vertical depth of 90 m (GM 33109);
1984: Golden Hope Mines Ltd of the Noramco Group of companies
(Noramco) acquired a large block of claims covering a west-northwest
trending swarm of long airborne electromagnetic (EM) conductors. A
deal was signed with Teck Exploration Ltd. (Teck), whereby Golden
Hope provided the property and the exploration funds, and Teck
provided its technical and managerial expertise. During the summer,
various exploration work was conducted by Teck including line cutting
and 112 km of EM and magnetic surveys (Mag);
1985: Line cutting (140 km), several geophysical surveys (66 km of EM,
104 km of Mag, 50 km of IP and EM-37), and drilling of 300 reverse
circulation (RC) drill holes and 120 diamond drill holes for 31,966 m
were completed. Results of the overburden drilling were disappointing.
Diamond drill targets were selected entirely on the basis of geophysical
anomalies. Eight such anomalies were selected, none of which was
rated very high. The third anomaly to be drilled was the one
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corresponding to the Estrades deposit (hole H-003: 15% Zn, 3% Cu, 0.2
oz/t Au, 9 oz/t Ag over 35 ft);
1985-1988: Following the discovery hole, Teck completed 77,000
metres of diamond drilling and conducted numerous geophysical
programs on the deposit. This discovery led to the detailed drilling of the
Main Zone and the discovery of the West, Central and East Zones;
1986-1987: Additional geophysical surveys (Mag 148 km, IP 196 km)
and diamond drill holes (66 holes for 23,621 m);
1987-1988: Pulse EM (10 holes), overburden drilling (47 holes for 906
m) and diamond drilling (107 holes for 33,345 m) were completed.
Subsequent lithogeochemical sampling, metallurgical testing,
preliminary engineering studies and research investigations were
completed in-house;
1985-1990: Noramco-Teck joint-venture drilled 24 holes covering the
actual Estrades Block;
1988-1989: Breakwater Resources Inc. (Breakwater) became involved
in the project in 1988, when it gained control of Noramco. At that time,
the Estrades property was owned by Noramco affiliates Golden Hope
Resources Inc (40%) and Golden Group Exploration Inc (60%).
Breakwater earned a 20% undivided interest in the Estrades deposit by
completing a feasibility study on the Estrades deposit and incurring
expenditures of no less than $3 million. The agreement also granted
Breakwater the option to earn an additional 50% interest by making a
cash payment of $0.5 million to Golden Hope-Golden Group and
bringing the property into production. A feasibility report was completed
by Wright Engineers Limited for Breakwater. That report addresses the
“Ore Reserve Estimate – Phase 1” of the Estrades deposit;
1990-1991: In February 1990, Breakwater exercised its right of earning
a full 70% interest in the mine and formed a JV with Golden Hope-
Golden Group (Breakwater 70%, Golden Group 18%, Golden Hope
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12%). The mine was then operated as a JV until suspension of
operations in June 1991. No work other than technical and engineering
studies has since been carried out at the mine;
1992: During the third quarter, the mine was dewatered to confirm the
geological structure on which the new Breakwater reserves were based
and assess the condition of the underground workings. Due to low
metal prices and exchange rates, Breakwater delayed the re-opening of
the mine. The mine was subsequently allowed to flood and kept on a
care and maintenance basis;
1994 Arimetco International Inc. (Arimetco), a United States copper
producer, assumed full management of Breakwater with the ultimate
intent of merging with Breakwater;
1995: Arimetco notified Breakwater that it no longer intended to
complete the merge with Breakwater. Arimetco and Breakwater
negociated for the Breakwater debt. Breakwater’s interest in the
Estrades Mine and Mining Lease #795 was thus transferred to Arimetco
as full settlement of that debt. A production royalty of 2.5% NSR on the
first 450,000 tonnes of ore produced and 3% on tonnes in excess of
450,000 tonnes was retained by Breakwater;
1996: Arimetco declared bankruptcy;
1999: Arimetco transferred its interest in the Estrades Mine (and in
Mining Lease #795) to Western Gold Resources Inc. (Western Gold). In
the mean time, the Québec Government had initiated legal proceedings
against Arimetco to revoke Mining Lease #795 for lack of compliance.
Western Gold was able to re-activate Mining Lease #795 by making, on
behalf of Arimetco all lease payments that were in arrears. Western
Gold merged with Atlas Minerals Inc. to form Atlas Precious Metals Inc.
(Atlas), thus becoming the owner of Breakwater’s original 70% interest
in Estrades;
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2002: SRK Consulting (SRK) carried out a diligence evaluation of the
Estrades Project for Atlas in November. A mineral reserve estimate,
based on extensive review of all available data, drill core, mine records
and maps, consultants’ reports and discussions with a former engineer
at the mine while in operation, was carried out;
2003: Atlas acquired 100% of the outstanding shares of Western Gold.
Atlas commissioned Western Range Services Inc. (Western Range) to
do a resource evaluation;
2005: Woodruff Capital Management Inc. (now Cogitore Resources Inc.)
purchased a 70% interest in Mining Lease #795 from Atlas Precious
Metals Inc. and then purchased from Orvilliers Resources Ltd. the
remaining 30% interest. Through these two agreements, Cogitore now
owns a 100% interest in Mining Lease #795.
2006: Cogitore carried out 8,140 m of diamond drilling in the Mining
Lease #795, in and around the known “Main Zone” deposit, followed by
borehole PEM surveys. Cogitore also completed 3,233 m of diamond
drilling in claims adjacent to Mining Lease #795.
2007 : A feasibility study initiated in 2006 was pursued in 2007, along
with permitting work which resulted in a Certificate of Authorization for
the mining project delivered in August 2007. A total of 4,259 metres
were drilled and the program included re-entering three representative
holes through the Main Lens to get a 200 kg composite metallurgical
sample for the feasibility study. In addition, exploration holes were
drilled to test a felsic unit similar to the Estrades mine unit and located
600 metres to the north of the mine horizon, and finally to follow-up the
East Zone sector at depth. Drilling in the East Zone sector was
successful in discovering a new copper zone returning 0.84% copper
over 23 metres, including a higher grade zone of 3.1% copper and 21
g/t silver over 4.4 metres. This new copper zone is a stringer zone of
the type that is commonly found below massive sulphide deposits.
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8.2 Mineral Resource and Mineral Reserve Estimates
Previous mineral resource and reserve estimates were carried out by Teck
(1989), Noramco and Breakwater (1989), Wright Engineers Ltd. (1989),
Breakwater (1992), Derry Michener Booth and Wahl (DMBW, 1997), SRK (2002),
Western Range (2003) and Scott Wilson Roscoe Postle (2006). All mineral
resource and reserve estimates used different parameters (cut-off grades,
minimum mining widths, dilution factors, specific gravity, combinations of metal
prices and mill recoveries) or different estimation methodologies (polygons on
vertical longitudinal section or block modeling).
In 1990, prior to going underground, Breakwater estimated the Main Zone of the
Estrades deposit, down to a vertical depth of 600 m, to contain 941,000 tonnes of
ore at an average grade of 10.68% Zn, 0.94% Cu, 0.92% Pb, 5.59 g/t Au, and
182 g/t Ag at an NSR cut-off of $80. Dilution was estimated at 24%.
Breakwater reported also a diluted “ore reserve” for the Central Zone, which is
located to the east of the Main Zone, of 400,000 tonnes at 6.30% Zn, 0.68% Cu,
0.63% Pb, 3.97 g/t Au, and 84.55 g/t Ag. The average minimum mining width
was 2.2 m. At that time, due to mining costs at $91 (Taylor, 1990), the Central
Zone was considered marginal with an NSR value per tonne of ore at $85.
On May 15, 1991, Breakwater revised the "reserves" of the Main Zone above
Elevation 4,600 m to 259,303 tonnes at a grade of 12.59% Zn, 0.79% Cu, 7.35
g/t Au and 210 g/t Ag. The crown pillar was excluded from these reserves. A
minimum mining width of 2.0 m was used. In 1992, Breakwater estimated the
mineral reserves to contain 271,415 tonnes at an average grade of 13.07% Zn,
0.88% Cu, 7.52 g/t Au and 214.31 g/t Ag, above Elevation 4,600 m.
In 2002, at the request of Atlas Minerals Inc., SRK Consulting prepared a due
diligence evaluation for the Estrades Project. Mineral resources considered by
SRK were primarily a reflection of work performed by Teck, Breakwater and
others. Based upon this past work, resources were reported as “measured”,
“indicated” and “inferred” in the Main Zone (Breakwater) and as geologic mineral
inventory (Teck) for the Central and the East Zones. SRK adjusted the 1991
Breakwater reserve estimate to achieve a mineral reserve estimate of 324,715
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tonnes at an average grade of 9.66% Zn, 0.57% Cu, 5.28 g/t Au, and 157 g/t Ag
above the 4,600 m elevation.
In 2003, Western Range was retained by Atlas to provide a new “ore reserve
statement” for the project. The mineral reserve estimate was carried out through
a kriged block model. A two-metre minimum mining width was used. At a
US$65 cut-off grade, the mineral resources for the Main, Central and the East
zones were estimated to contain 1,068,271 tonnes at an average grade of 8.65%
Zn, 05% Cu, 0.88% Pb, 4.29 g/t Au, and 143.86 g/t Ag, down to a vertical depth
of 600 m. In April 2005, Robert Sim of Sim Geological was asked by Woodruff to
open the digital files containing the Western Range block model and to further
detail Western Range numbers for the Main, Central, and East zones but without
checking or redoing the calculations. Sim concluded that the Western Range
numbers looked too high when comparison is made between ore thickness as
modeled by Western Range and the actual thickness based on drill data. Sim
concluded that Western Range overestimated the mineral resource.
In November 2006, Scott Wilson Roscoe Postle undertook a mineral resource
estimate. Original assays were used for interpretation of the mineralized
envelope. Zn, Cu, Au and Ag grades of each sample were converted into dollar
values based on NSR. No value was given to the Pb in the ore. A cut-off grade
of $120/t was used and this was calculated with the parameters described in
chapter 19 of the current report. A minimum width of 2 metres was used for the
interpretation of the envelope. For the Main and Main-East Zones, total
‘’indicated’’ mineral resources of 592,000 tonnes were estimated at a grade of
9.82 %Zn, 0.81 %Cu, 5.21 g/t Au, 168.0 g/t Ag and 0.90 %Pb.
8.3 Production
The Main Zone was mined between July 1990 and May 1991. A total of 166,928
tonnes at an average grade of 13.06% Zn, 1.30% Cu, 6.11 g/t Au and 169.16 g/t
Ag are reported to have been mined. Mining was done on a contract basis.
From August 1990 to June 1991, the ore was milled, on a custom-milling basis,
at the Matagami mill which is located 128 km from the mine. At that time, the
Matagami mill was operated by Noranda Minerals Inc. A total of 174,946 tonnes
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at an average grade of 12.93% Zn, 1.14% Cu, 6.35 g/t Au and 172.30 g/t Ag are
reported to have been milled. There was no explanation for the discrepancy
between the mined and milled tonnage (+4.8%) and grades. The milled tonnage
and grades are normally the base for payments to the mine and therefore, they
probably represent the official production figures.
Operations were suspended in June 1991 due to low metal prices and excessive
contract mining and processing costs. Monthly production of mined and milled
ore is presented in Item 16 of this report.
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9. GEOLOGICAL SETTING
9.1 Regional Geology
The rocks of the area are constituted of meta-volcanic and sedimentary rocks of
the Harricana-Turgeon Belt (HTB) which is located in the NW part of the Abitibi
Subprovince. The regional metamorphism is of greenschist facies. Rocks are
east-west striking and vertically dipping.
Four regional lithostratigraphic domains are recognized in the area (Figure 9-1):
the Orvilliers-Desmazures Basaltic Domain (5 km wide), the Taïbi Sediments
Domain (1.5 km wide), the Joutel-Raymond Basaltic-Rhyolitic Domain (>5 km
wide), and the Cartwright Hills Basaltic to Komatiitic Basaltic Domain (< 2 km
wide).
Those lithostratigraphic domains are bounded to the north by the Orvilliers Pluton
which is of quartz granodiorite to monzodiorite composition, and to the south by
the Mistaouac pluton which is of a tonalite to diorite composition.
A major regional deformation zone, the Casa-Berardi Break, is located
approximately two kilometres north of the property limit within the Taïbi
sediments. The Casa Berardi Break is a graphitic fault with injections of quartzcarbonate
veining. Iron formations which are well defined on magnetic maps
occur in the southern portion of the Taïbi sediments.
Rocks are cut by two major east-northeast to northeast trending diabase dykes.
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Figure 9-1
Regional Geology
CIM DEFINITION STANDARDS - For Mineral Resource and Mineral Reserve
Mineral Resource
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9.2 Property Geology
The Estrades deposit area is underlain by a succession of east-west striking,
steeply dipping, Archean meta-volcanic and meta-sedimentary rocks (Figure 9-
2). Most stratigraphic units are intruded by later felsic and mafic dykes and sills.
Stratigraphy is interpreted to face south, but this is based solely on the
occurrence of the alteration zone and stringer mineralization on the north side of
the massive sulphide deposit. The rhyolitic host to the deposit has recently been
dated as 2,719.8 ±2.8 Ma.
Stratigraphy is described below from youngest to oldest rocks (Unit 13 to Unit 1).
These rock units were summarized by O’Dowd et al. (1989) and by Welch
(1995). Sequence numbering is, however, from Welch (Figures 9-3 and 9-4).
PROTEROZOIC
Unit XIII: Late Intrusions
The most prominent intrusion is a northeast trending diabase dyke that runs
through the middle of the Main Zone, but not through the “ore horizon”. It is a
fine to medium-grained, magnetic diabase with 50% mafic minerals and 50%
plagioclase minerals often with a well developed typical “diabase texture”.
ARCHEAN - JOUTEL RAYMOND DOMAIN (JRD)
Unit XII: Mafic to Intermediate Volcanics
This unit is the southernmost and at the top of the stratigraphic sequence. Rocks
consist of fine-grained to medium-grained, pillowed flows and flow breccias
locally containing 5% to 10% quartz and quartz-carbonate filled amygdules, and
feldspar phenocrysts (1% to 2%). Alteration consists principally of chlorite and
carbonate. Pyrite occurs as fine-grained disseminations.
Unit XI: Intermediate Volcanics
Sheared intermediate volcanic rocks are weakly altered but strongly deformed.
Clasts within the volcaniclastic unit may be weakly sericitized and the finegrained
matrix typically shows various degrees of chlorite alteration. There is no
significant mineralization in this unit.
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Unit X: Sedimentary Unit
A thin, sedimentary unit occurs within mafic to intermediate volcanics. The metasediments
consist of siltstone, argillite, and minor greywacke. This unit is locally
brecciated, with graphite-rich contacts.
Unit IX: Felsic Hanging Wall Unit
Two minor felsic tuff horizons occur within 100 m of the Main Felsic Unit (Unit 7).
They constitute the Hanging wall Felsic Unit (HFU). These horizons were
described as containing lapilli to block size fragments in an often darker
(chlorite), felsic ash-sized matrix. Both horizons are moderately to strongly
sericitized but contain more chlorite towards their lower contact. Disseminated,
fine to coarse grained pyrite (

Feasibility Report
however sphalerite is common, as well as chalcopyrite and galena. Elevated
values of both silver and gold occur in the hanging wall and footwall.
Unit VI: Felsic Rocks
This unit is medium grey, siliceous, massive felsic rock containing up to 5%
quartz crystals, and is moderately foliated but with an overall uniform massive
appearance. This unit is not significantly altered or mineralized.
Unit V: Mafic to Intermediate Volcanics
This unit consists of mafic to intermediate volcanics. The rock is fine-grained,
light grey to dark green with flows that are massive to moderately foliated, often
amygdaloidal, quartz and quartz-carbonate filled, locally feldspar porphyritic and
contain patchy secondary carbonate. Trace, disseminated, fine-grained pyrite is
the dominant sulphide.
Unit IV: Felsic Tuff
The Footwall Felsic Unit (FFU) is a sheared, monolithologic (felsic fragments)
lapilli tuff to tuff breccia unit that is usually strongly sericitized. This unit is
depleted in CaO and Na2O, and enriched in K2O, with elevated base metal
values (Clark, 1986).
Unit III: Mafic Volcanic Flows
A succession of mafic flows and tuffs occurs south of the Casa Berardi
sediments. The flows are fine-grained, dark green, foliated, and locally
amygdaloïdal. Thin interflow units of monomictic fragments and matrix
supported, mafic tuff, lapilli tuff and minor crystal tuff are intercalated with these
flows. Both flows and interflow units are weakly chloritized, carbonatized, and
contain trace-disseminated pyrite.
ARCHEAN - TAÏBI-SEDIMENTS DOMAIN (TSD)
Unit II: Taïbi Sedimentary Roks (TSD)
This unit, which ranges from 700 m to 1500 m in thickness, is composed of
sandstone, siltstone, greywacke and argillite. The unit hosts the Casa Berardi
Fault (CBF), a four metres wide graphitic fault with quartz-carbonate veining.
Iron formation occurs in the southern portion of the sedimentary package and is
evident on magnetic maps as a series of magnetic highs traversing the centre of
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the property block. This iron formation consists of fine-grained alternating
laminae and beds of magnetite and chert. The Casa Berardi sediments are
variably sericitized and carbonatized. The alteration increases towards the CBF
where the sediments are strongly sericitized and contain up to 20% ankerite
(Clark, 1986) and, locally, pyrite and arsenopyrite-bearing, smoky to dark quartz
veins containing pyrite and arsenopyrite. Anomalous gold occurs locally.
ARCHEAN - ORVILLIERS DESMAZURES DOMAIN (ODD)
Unit I: Mafic Volcanics (ODD)
This unit is the northernmost and forms the base of the stratigraphic sequence.
Rocks consist predominantly of massive flows, though pillowed and porphyritic
flows are recognized. Interflow sediments or tuffs with siliceous chert-like
laminations separate some flows. Most interflow breccias are probably flow
breccias. The rocks are typical greenschist facies rocks and contain chlorite,
calcite, epidote, and quartz. Base metal mineralization (Cu, Zn) is not common,
however, pyrite is ubiquitous as fine disseminated grains.
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Figure 9-2
Property Geology
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Figure 9-3
Stratigraphy – Plan View
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Figure 9-4
Stratigraphic Column
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10. DEPOSIT TYPES
The Estrades deposit shows similarities with other Archean volcanogenic
massive sulphide (VMS) deposits and is considered as such (Welch, 1995).
Exploration for such type of deposits is essentially guided by the recognition of:
Volcanic flows and volcanic cycles;
Presence of a hiatus (quiescence) period between volcanic flows;
Presence of sericite-chlorite alteration;
Background base metal mineralization, either disseminated or stringer;
Geophysical anomalies which could indicate the presence of conductive
mineralization.
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11. MINERALIZATION
The Estrades Unit (EU), a polymetallic (Zn, Cu, Pb, Ag, Ag, Au) VMS deposit
unit, is hosted in altered felsic rocks referred to as the “Main Felsic Unit (MFU)”
(Welch, 1995). This mineralization is considered as classic Archean age of the
syngenetic exhalative type. The EU and the MFU units strike almost east-west,
over several kilometres, with a nearly vertical dip. Base metal and gold
mineralization is found on the property in several zones, namely the Main,
Central, East and West Zones (Figure 11-1).
The deepest drill hole (hole H-279), targeting the Estrades Unit, has intersected
sulphide mineralization at 930 m below surface (O’Dowd et al., 1989).
The deposit is covered by glacial sediments consisting of silt, clays and sandy
gravel with variable thickness.
11.1 Main Zone
The Main Zone is mineralized over a strike length of 600 m, between sections
2100W and 2600W, and extends at depth to at least 650 m below surface with
an average width of 3.8 m. Economic mineralization is essentially concentrated
between sections 2275W and 2500W, and between sections 2125W and 2225W.
All past production originated in this zone. The Main Zone is subdivided into the
following four components, from north to south:
Footwall stringer zone;
Banded massive sulphides;
Fragmental massive sulphides;
Hanging wall massive sulphides.
Pyrite is the predominant sulphide mineral followed by, in decreasing abundance,
sphalerite, chalcopyrite, galena, and pyrrhotite. The precious metal content is
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represented by a silver-gold amalgam, ranging in composition from silver-rich
electrum to the gold-rich kustelite.
The pinch and swell nature of the Main Zone is clearly shown on section 2400W
(Figure 11-2).
There are two main faults which are associated with the Main Zone deposit:
Main Fault: the dominant structure within the deposit, strikes at 338° and
dip 65°SW, separating the Main Zone from the Central Zone;
Fault A: strikes at 303° to 325, dips 50° SW, and cuts the East Zone.
On a vertical longitudinal section, the massive sulphide body has a funnel shape
plunging vertically with its eastern margin apparently bounded by the Main Fault.
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Figure 11-1 Longitudinal Section
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Figure 11-2 Vertical Section 2400 W
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11.2 Central Zone
The Central Zone occurs between sections 1600W and 2100W. The zone has a
strike length of 500 m and an average width of two metres. Drilling to a depth of
400 m indicates a massive sulphide body which appears to be cut by several
faults. The western limit of the Central Zone is interpreted to terminate against
the Main Fault.
In 1990, Breakwater reported a diluted “ore reserve” for the Central Zone of
400,000 tonnes 6.30% Zn, 0.68% Cu, 0.63% Pb, 3.97 g/t Au, and 84.55 g/t Ag
over an average minimum width of 2.2 m. At that time, due to mining costs at
$91 (Taylor, 1990), the Central Zone was considered marginal with an NSR
value per tonne of ore at $85. Resources calculated by GENIVAR are discussed
in section 19.3 of this report.
11.3 East Zone
The East Zone occurs between sections 1500W and 1000W and consists of a
thin discontinuous body of mostly pyrite. The zone has been identified to a depth
of approximately 600 m. Massive sulphide thicknesses range from one metre to
2.5 m. Three of the best base metal intersections are reported in the following
holes (Welch, 1995):
Hole 221 - 2.1 m at 10.1% Cu, 4.8% Zn, 1.7 g/t Au
Hole 230 - 3.3 m at 5.9% Cu, 9.4% Zn, 2.8 g/t Au
Hole 082 - 0.9 m at 9.3% Cu, 2.7% Zn, 1.5 g/t Au
Narrower, gold rich intersections are also present in the East Zone.
Hole 227 - 0.6 m at 1.2% Cu, 3.4% Zn, 50.1 g/t Au
Hole 087 - 0.6 m at 0.02% Cu, 5.8% Zn, 20.95 g/t Au
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11.4 West Zone
The West Zone extends from section 2900W to section 3000W. It has been
intersected in only two drill holes. The maximum thickness intersected is three
metres. Mineralization is essentially constituted of barren pyrite.
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12. EXPLORATION
Inmet (2001-2003) and Cogitore (2005-2006) carried out various exploration
programs on the property consisting of geophysical surveys, diamond drilling
and geological compilation.
Thirty-six (36) diamond drill holes for a total of 16,146 m were drilled in 2001,
2005, 2006 and 2007 by Inmet and Woodruff/Cogitore. All holes were drilled for
gold and base metal exploration. Cogitore carried out exploration in the
Estrades Block, the Estrades Lease and in the Newiska Block.
This report focused on the Estrades Mining Lease and details on the work done
in other blocks can be found in Cloutier (2005) and Salmon (2006).
12.1 Year 2001
From early January to the end of February 2001, line cutting grids, followed by
surface geophysical surveys, diamond drilling and down hole geophysical
surveys were completed in a selected area of the property.
LINE CUTTING GRIDS
Early in January, three distinct grid lines were cut for a total of 40.2 km, two on
the Estrades claim block (27.4 km) and one on the Newiska Block (12.8 km).
SURFACE GEOPHYSICAL SURVEYS – ESTRADES BLOCK
From January 16 to 29, surface TDEM (Time Domain Electromagnetic) surveys,
for a total of 23.2 line-kilometres, were completed on the two above grids of the
Estrades Block. The work was completed by Abitibi Géophysique of Val d’Or for
Inmet. Abitibi Géophysique concluded that no definite response that could be
caused by a large sulphide mineralized body was detected. However, close
review of the profiles by Cloutier (2005) suggested that a good conductor was in
fact located at depth, on L4900E, in the east end of the block.
DIAMOND DRILLING
After the completion of the above surface TDEM surveys, three holes were drilled
for a total of 1,592 m. Drilling occurred from January 16 to February 14, 2001.
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The first two holes (NK-01 and NK-02) were drilled on the Newiska Block and the
last one (EME-01) on the Estrades Block. The work was contracted out to
Forage Rouillier of Amos, Québec, under the direction of Mathieu Guay,
Geologist and Marco Gagnon, Project Manager, both working for Inmet. The
core diameter is BQ. The core is stored at the former Norbec Mine in the Rouyn-
Noranda district. Borehole pulse EM was completed in all the three holes.
ESTRADES BLOCK
Hole EME-01 was drilled to test the eastern extension of the Estrades Mine felsic
unit. It was drilled at an azimuth of 348° and an inclination of -66° and reached a
depth of 810 m. The hole encountered 35 m of overburden. From 35 m to the
end of the hole at 810 m, rocks mostly consisted of weakly altered volcanic rocks
of the Joutel Raymond Domain (JRD). A wide sedimentary unit was intersected
from 132 m to 319 m. Faulting was encountered from 249 m to 271 m. Two
felsic units were intersected from 319 m to 383 m and from 676 m to the end of
the hole at 810 m. The latter unit shows a strong typical VMS alteration. Two
minor metallic values were intersected in the later felsic unit which returned 13
ppm Cu and 1,060 ppm Zn over 1.5 m, and 17 ppm Cu, and 887 ppm Zn over
1.5 m.
DOWN HOLE GEOPHYSICAL SURVEYS
All holes drilled by Cogitore on the Estrades property have been systematically
surveyed with borehole Pulse EM surveys.
12.2 Year 2002
SURFACE GEOPHYSICAL SURVEY
A surface geophysical survey over a selected area of the property and a
compilation of geotechnical information were completed during the year. In
March, a TEM (Transient Electromagnetic) survey was done on the Estrades
Block. The survey covered the area not covered with the 2001 survey between
L6E and L34E. The work was contracted out by Inmet to Quantec Geoscience
Inc. of Porcupine, Ontario.
The TEM survey was successful in delineating several weak axial conductors
generally trending east west. All conductors are weak and reflect probable
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overburden related features. However, the axes appear to be concurrent with
the stratigraphy and could also be explained by mineralized contacts or shears.
LITHOGEOCHEMICAL STUDY
An extensive lithogeochemical study was completed by Inmet on the Estrades
and Newiska Blocks. The study was based on all published geochemical data
from core analysis of previous drill holes. Results were presented on drill hole
cross-sections, projected on a vertical longitudinal section traced along the
Estrades deposit and traced on surface compilation maps for both the Estrades
and the Newiska Blocks (unpublished data – Cogitore files).
From this study, alteration zones were identified in several areas based on
lithogeochemical results. Such alteration zones compare with similar patterns
common in VMS deposits such as Noranda and Matagami VMS mining camps.
12.3 Year 2003
AIRBORNE GEOPHYSICAL SURVEY
In October, a limited airborne geophysical survey was done over three distinct
areas of the property. The work consisted in an EM and magnetic survey. The
work was completed by Géophysique GPR International Inc. of Longueuil,
Québec, for Inmet (Arsenault and Mouge, 2003).
A total of 140 km of lines were surveyed over the three blocks labelled: “Bloc
Ouest, Bloc Nord-Est et Bloc Sud-Est”. Lines were spaced at 100 m and flown in
an NS direction. The helicopter speed was 75 km/h with an altitude of 65 m
above ground whereas the transmitter - receiver was towed behind the helicopter
at 30 m above ground.
The system used the Geotech EM in the Time Domain (VTEM) with a highresolution
magnetic instrument.
Numerous poor EM conductors were detected, possibly caused by subsurface
formations.Three possibly interesting conductors were identified on L2020 of
claim 5259484, on L1450 of claim 5262674 and on L3240 of claim 5259429.
Locally moderately magnetic formations were detected.
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The EM conductor on L2020 coincides with the deep conductor identified in the
2001 TDEM survey in the east end of the Estrades Block.
12.4 Year 2004
No work was done in 2004.
12.5 Year 2005
All work was carried out by Woodruff/Cogitore. The work consisted in line cutting
(19 km), a surface geophysical survey and drilling of six holes for a total of 2,992
m.
LINE CUTTING GRID
Approximately 17 km of lines were cut in the western part of the Estrades Block
in preparation for a deep EM survey and two kilometres of lines were cut in the
western part of the Newiska Block in preparation for a DeepEM survey on Line
90W.
SURFACE GEOPHYSICAL SURVEYS
Approximately 11 km of lines were surveyed with a Crone system DeepEM
survey (a fixed loop time domain EM system) on the Estrades Block. No
significant anomalies were detected.
DIAMOND DRILLING
Three holes totalling 1,880 m were drilled, with three holes in the Estrades
claims. Only anomalous values were intersected.
The work was contracted out to Forage Rouillier of Amos, Québec, under the
direction of Tony Brisson, Exploration Manager of Woodruff/Cogitore. The core
diameter is BQ. The core is stored at Inmet’s former Norbec Mine in the Rouyn-
Noranda district.
ESTRADES BLOCK
Holes EME-02, EME-03, and EME-04 were drilled on the Estrades Block. Hole
EME-02 was drilled to test the Estrades Mine horizon to a depth of about 600 m,
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approximately 200 m down-dip from hole H-82 which had returned 9.3% Cu and
2.7% Zn over 0.9 m of core length. It was drilled with an azimuth of 168° and an
inclination of 69° to a depth of 781 m. The hole intersected a sulphide
(chalcopyrite, pyrite, sphalerite) stringer zone that seemed to line up with the
zone in H-82 but the hole was stopped short of the top contact of the Main Felsic
Unit (MFU) due to logistical problems related to an early break-up. Very strong
sericite alteration is associated with the stringer zone which extends intermittently
from 760 m to 773 m. Significant assays in copper, zinc, gold, and silver are
presented in Table 10-1.
In all likelihood, these stringers occur at the same stratigraphic interval as the
Estrades deposit and the high copper to zinc ratio suggests proximity to a vent
zone for VMS.
Hole EME-03 was drilled to test a conductor identified from the 2003 VTEM
survey and from the 2001 surface TDEM survey. It was drilled with an azimuth of
168° at an inclination of 60° to a depth of 501 m. The hole successfully
explained the conductor as being a 10 m zone of pyrrhotite stringers intersected
from 439 m to 451 m. Assays from the sulphide zone did not show any
anomalous metal values. It should be noted, however, that the MFU hosting the
Estrades deposit was not intersected in the drill hole (contrary to what was
expected from a compilation map).
Hole EME-04 was drilled to test the Estrades Mine horizon to a depth of about
500 m, approximately 500 m east of Inmet's hole EME-01. It was drilled with an
azimuth of 168° at an inclination of 69° to a depth of 598 m. The hole intersected
altered felsic volcanics cut by numerous gabbro dykes or sills. Alteration is
locally very strong and consists mostly of sericite with local talc. The top of the
MFU seems to have been properly tested with this hole but no sulphides were
present.
DOWN-HOLE GEOPHYSICAL SURVEYS
Hole EME-03 was surveyed with a Crone borehole PEM (pulse EM) system.
Holes EME-02 and EME-04 could not be surveyed due to spring break-up
conditions.
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12.6 Year 2006
Work in 2006 consisted in:
Line cutting in preparation for an InfiniTEM survey to be done in late
2006 to the east of Mining Lease #795;
Diamond drilling in both Mining Lease #795 and adjacent claims;
Borehole PEM surveys in most of the 2006 holes and in a few 2005
holes that could not be surveyed in 2005.
A preliminary evaluation of capital and operating costs for resuming production at
the Estrades deposit was completed by Met-Chem Canada Inc., the results of
which led to the resource evaluation by Scott Wilson RPA.
LINE CUTTING GRID
Approximately 52 km of lines were cut partly within Mining Lease #795 (20 km)
and east of the lease in the Lac Rouillard area (32 km) in preparation for a deep
penetrating surface InfiniTEM survey targeting the east extension of the Estrades
Mine horizon and a second potential horizon located about 400 m to the north of
the Estrades horizon. The InfiniTEMsurvey is scheduled to begin in November
2006.
DIAMOND DRILLING
From January to April 2006, Cogitore drilled 20 holes for a total of 8,708 m
(including wedge-cuts and holes aborted due to deviation). The holes were
drilled in Mining Lease #795, in the surrounding Estrades claims.
The work was completed by Forage Rouillier of Amos, Québec, under the
direction of Tony Brisson, Exploration Manager with Cogitore. The core diameter
was BQ and the core is stored at the Estrades Mining Lease property, except for
high grade mineralized sections which are stored at the Cogitore office in Rouyn-
Noranda. Borehole PEM surveys were done in most of the holes.
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MINING LEASE #795
Seventeen holes were drilled for a total of 7,040 m, including six holes
abandoned at depths less than 50 m due to unacceptable deviation in either
azimuth or dip. Drilling inside the Mining Lease was grouped with three different
objectives:
Fill gaps and confirm continuity within the limits of the known resources;
Find possible extensions of the massive sulphides outside of the known
resources but in the immediate vicinity of the deposit;
Wide space testing of the Estrades Mine horizon to the west of the
deposit.
One new hole (EST-03) and two wedge-cuts of old holes (H-165W and H-193W),
were drilled inside the limits of the known resources in order to fill gaps that were
thought to be too large. These three holes intersected high grade massive
sulphides which confirmed the high grade nature of the deposit and the continuity
of the massive sulphides. The most significant results from these holes are
reported in Table 12-1.
Seven holes (EST-01, EST-02, EST-04, EST-05, EST-06, EST-07, EST-08) were
drilled to find extensions of massive sulphides in the immediate vicinity of the
deposit. One of the holes (EST-01) was stopped at 730 m due to an
unacceptable deviation, whereas the other six holes reached their target which
was the Estrades Mine horizon. Results of these holes are reported in Table
10.0
Although holes EST-04, EST-05, EST-06 and EST-07 intersected massive
sulphides, only EST-05 had results which were considered of potential economic
interest. Hole EST-08 was dyked out and borehole PEM failed to indicate
conductive material in the immediate proximity of the hole, therefore it must be
concluded that no massive sulphides are located within 100 m to 150 m from the
hole. One hole (EST-09) was drilled inside the Mining Lease approximately 800
m west of the Main Zone as part of wide space testing of the Estrades Mine
horizon. This hole did intersect the Estrades Mine horizon but with no
encouraging mineralization. In conclusion, Cogitore’s drilling has confirmed the
quality and continuity of the deposit inside the previously known limits but has
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also shown that the deposit (the Main Zone) is well closed off by drilling in all
directions.
ESTRADES ADJACENT CLAIMS (ESTRADES BLOCK)
In 2006, two holes (EME-05 and EME-06) were drilled outside of Mining Lease
#795 but along the Estrades Mine horizon. EME-06 was drilled about 1,200 m
west of the Main Zone to test the mine horizon. Like EST-09, this hole did
intersect the Estrades Mine horizon but with no encouraging mineralization. Hole
EME-05 was drilled in the “East Zone” area to test an “off-hole” anomaly detected
in hole H-113A. The hole intersected the mine horizon at a vertical depth of
about 800 m which was below the geophysical target. The hole did not intersect
significant mineralization, with the best assay being 0.72% Cu and 0.92% Zn
over 0.48 m. However, the Estrades horizon may be dyked out by a gabbro at
the target depth since borehole PEM showed a 12-channel “off-hole” conductor
exactly in the centre of the gabbro interval. It is interesting that the Estrades
deposit (the Main Zone) is a moderate to poor conductor, typically with
responses that show up only in the first 10 to 12 channels. Some follow-up is
required in this sector as the anomaly in H-113A remains unexplained and EME-
05 also shows an unexplained “off-hole” conductor in its vicinity.
12.7 Year 2007
Only one drill hole (length of 189.5 m) was done. This hole, EST-10, was drilled
in the Main-East sector, on a mining lease. It intersected a narrow mineralization
over 2.6 m at 2.67 %Zn, 0.62 %Cu, 0.15 %Pb, 44.57 gpt Ag and 3.26 gpt Au,
which reduces the potential of mining this area. This hole has also intersected
another horizon north of the Main-East over 1 m at 11.00 %Zn, 1.22 %Cu, 0.24
%Pb, 64.1 gpt Ag and 3.1 gpt Au.
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Table 12-1
Exploration – Best Intersections
Hole From To Length Cu % Zn % Pb% Au g/t Ag g/t
Year 2005
EME-02 763.40 772.40 9.00 0.64 0.12 - 0.09 5.30
including 767.75 771.60 3.85 1.30 0.23 - 0.33 9.80
including 771.30 771.60 0.30 4.05 0.79 - 0.70 27.50
Year 2006
EST-03 343.39 344.69 1.30 0.32 14.21 1.46 8.02 125.60
EST-03 346.37 347.08 0.71 0.71 23.90 3.43 3.32 230.00
H-165W 387.65 391.55 3.90 1.27 28.76 3.90 29.83 370.00
H-193W 467.70 469.90 2.20 1.04 20.57 0.78 3.59 267.80
EST-02 745.90 767.56 21.66 Disseminated pyrite, no significant
values
EST-04 892.65 895.16 2.51 0.93 0.51 0.08 2.76 138.03
Including 893.72 895.16 1.44 1.54 0.81 0.14 1.76 232.80
EST-05 595.46 600.18 4.72 0.27 6.29 0.28 2.32 47.47
including 595.46 598.40 2.94 0.28 9.47 0.43 3.42 58.07
EST-06 511.56 513.10 1.54 0.69 12.62 0.81 3.22 61.63
EST-07 600.62 601.52 0.90 0.02 1.23 0.23 0.22 20.60
EST-08
Dyked out (diabase)
Year 2007
EST-10 150.60 151.60 1.00 1.22 11.00 0.24 3.10 64.10
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13. DRILLING
Diamond drill holes were planned (azimuth, dip, length) by Cogitore geologists on
vertical cross sections and on surface plan views in order to hit geological units
relatively perpendicular to their strike and dip. This way, mineralized intersections
are relatively close to their true thickness.
Drill collars were spotted on the field lines with the use of modern surveying
equipment. Front sight and back sight are identified with pickets.
Hole deviations (azimuth and dip) were measured with survey instruments
(tropari, multishot, or reflex) approximately every 30 m. In addition, dip angles
were measured at intervals varying from six metres to 25 m using Microsync or
Easy Dip instruments. All of these instruments provide accuracy better than +1
degree.
The core is placed in sequential order in core boxes labelled with a hole number.
Each run is identified by a wood block on which the depth of the hole is marked.
Missing (not recovered) core is identified by a wood stick indicating the length of
the missing section. At the end of each shift, core boxes are transported by
drillers to the core shack.
Upon receipt, core boxes are placed on tables and opened. Core is washed and
verified for accuracy. Then a geologist describes the core geology and enters
geological and structural data into a digital logging package. Drill hole logs show
hole parameters, core description, and sampling intervals. Core logging is
carried out in French.
The core recovery is generally very good, nearly 100%, with the exception of
short intervals within fault zones. Such intervals are generally marked during
drilling and checked later by the geology personnel for depth accuracy and
missing sections. RQD measurements have been made on a few holes but not
systematically in every hole done at Estrades.
All core boxes from drilling by Cogitore and Inmet Mining are stored at the mine
site. Only a few core boxes with mineralization intersections are currently kept at
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Cogitore’s office in Rouyn-Noranda. Core boxes from pre-2001 drilling have all
disappeared or deteriorated behond any potential use.
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14. SAMPLING METHOD AND APPROACH
In 2007, only one drill hole was done, EST-10 on the Main East Zone. A total of
31 samples were shipped to the analytical laboratory LabExpert in Rouyn-
Noranda to be assayed for Cu, Zn, Pb, Au and Ag.
7 samples were sent to the ALS Chemex laboratory for litho geochemistry.
3 drill holes were also carried out in order to undertake metallurgical test work in
laboratory. A total of 162 kg of mineralized material was sent to SGS Lakefield
for that purpose.
No statistical analysis was done in the course of the current study.
Further information about the sampling method and approach used for the
samples of the previous drilling campaigns can be found in Salmon (2006).
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15. SAMPLE PREPARATION, ANALYSES AND SECURITY
We refer the reader to the Scott Wilson Roscoe Postle technical report of
November 20, 2006, Mineral Resource Estimate for the Main Zone of the
Estrades Project, North Western Québec, Canada, for information related to this
chapter.
GENIVAR cannot comment on the samples handling, preparation as well as on
the security procedures used by Cogitore Resources.
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16. DATA VERIFICATION
A database, built for Gemcom, was transferred from Roscoe Postle to GENIVAR.
The drawings were examined, including plans and sections. The data for hole
EST-10, done on the Main East Zone, were provided from a Prolog data base.
Particular attention was paid to the block model done by Roscoe Postle in order
to validate the resource.
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17. ADJACENT PROPERTIES
There are no adjacent properties that should have a material impact on the
Estrades deposit.
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18. MINERAL PROCESSING AND METALLURGICAL TESTING
18.1 Laboratory Testwork Prior to Ore Milling in Matagami
Laboratory testwork started out in early 1987 and was undertaken by Lakefield
Research and Bacon Donaldson and Associates (Racine, 1991). The following
conclusions were drawn:
The ore demonstrates a complex metallurgy and the presence of
secondary copper minerals like chalcocite, covellite and digenite is
observed;
5% of the gold content is in the form of coarse electrum (particles > 100
µm), an alloy of gold and silver, soluble in a cyanide solution.
Cyanidation tests recovered 80 to 85% of the gold within 48 hours. 15%
of the gold cannot be recovered by cyanidation;
Sphalerite, galena and chalcopyrite are mainly contained in fine
inclusions within pyrite, requiring a relatively fine grind to allow a good
liberation;
A high percentage of zinc is recovered in the copper concentrate even
though zinc depressants are used in the flotation circuit. Two reasons
can explain this situation:
The presence of covellite in the ore, which in the course of the
process dissolves and as a result, copper ions activate sphalerite
and enhance its flotation with copper;
The presence of fine particles of chalcopyrite which stick to the
sphalerite grains and bring them into the copper concentrate.
The combination of zinc sulphate and cyanide gives the best results in
laboratory to depress zinc during copper flotation;
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Laboratory testwork produces a zinc concentrate of about 52 %Zinc with
a recovery of approximately 85%;
It is difficult to produce a saleable lead concentrate, the highest grade
achieved being 19 %Pb.
18.2 Ore Milling in Matagami
The Estrades ore was milled from August 1990 to June 1991 at the concentrator
of the Matagami Division of Noranda Minerals Inc. Concentrates of copper and
zinc were produced, the Estrades copper concentrate was blended with the
Matagami concentrate and the zinc concentrate was kept separate as reported
by Racine (1991).
The copper concentrate contained high amounts of zinc and lead and the copper
grade was low as the strategy was to pull hard in order to increase gold recovery.
The reagents used as zinc depressants in the copper circuit were zinc sulphate
(ZnSO 4 ) and sulphur dioxide (SO 2 ).
The following table shows the metallurgical performance obtained for the copper
concentrate during milling in Matagami.
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Table 18-1
Matagami Results with the Copper Concentrate of the Estrades Ore
The table below illustrates the metallurgical performance achieved for the zinc
concentrate.
Table 18-2
Matagami Results with the Zinc Concentrate of the Estrades Ore
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18.3 2007 Laboratory Testwork by SGS Lakefield Research Ltd
In early 2007, on behalf of Cogitore Resources Inc., GENIVAR requested SGS
Lakefield to initiate a metallurgical test program intended to apply previous
findings to the development of a new flowsheet. The program aimed to improve
the level of understanding of the Cu, Pb and Zn circuits with optimization focused
on improvements in selectivity between all minerals and improving grade and
recovery.
Here’s a summary of the major findings and results (SGS Lakefield, An
Investigation into the Recovery of Cu/Pb/Zn from Estrades Ore’’, December 14,
2007).
18.3.1 Sample Preparation
Approximately 162 kilograms of sulphide samples were sent to SGS Lakefield,
representing material from 3 drill holes (H-264, H-172 and H-155) including some
wall-rock samples. The material used for the test work was obtained in January
2007 by doing “wedge cuts” from holes that were initially drilled in 1987 and
1989. A single Master Composite was prepared by combining all of the material
received, in proportions calculated to obtain a weighted average of about 11 %
Zn and 5 g/t Au. Then it was stage crushed to –10 mesh, rotary split into 2 x 10-
kg and the remainder was split into 2-kg test charges and head sample.
The Master composite graded 0.75 %Cu, 1.60 %Pb, 14.0 %Zn, 5.66 g/t Au and
233 g/t Ag, which was a bit higher grade than estimated when the sample was
prepared. A complete chemical head analyses was performed on this sample.
18.3.2 Ore Mineralogy
A detailed mineralogical study was completed on the Master Composite to
identify mineral associations and develop grade limiting/recovery relationships.
Ore mineral modal abundance was determined with the QEMSCAN technology
and is outlined below for the major minerals.
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Table 18-3
Ore Mineral Modal Abundance
Minerals Abundance %
Pyrite 23.7
Quartz 21.4
Sphalerite 18.6
Micas 11.7
Chlorites 10.7
Fe-Oxides 5.4
Chalcopyrite 1.9
Clays 1.8
Galena 1.8
In the sample examined, 78% of the sphalerite, 76% of the chalcopyrite and 59%
of the galena by mass occur as free particles.
18.3.3 Gravity Recovery
Two tests were completed on the Master Composite to determine the potential
for producing a high grade coarse gold/silver concentrate by gravity
concentration. The feed samples were prepared by grinding 10-kg of material in
a laboratory ball mill to 135 µm and 245 µm. The ground material was fed to a
Knelson centrifugal concentrator and further upgraded through Mozley tabling.
Overall gravity gold and silver recovery was approximately 23% and 2%
respectively for both of the tested feed sizes, indicating that the samples as
tested are not suitable for gold and silver recovery by this method alone. In both
cases a high grade gold concentrate was produced with a high lead (Pb) content
making it suitable to be combined into a final Pb concentrate.
18.3.4 Flotation Testwork
A total of 28 batch flotation and 2 locked cycle tests were completed, exploring
both sequential and bulk Cu-Pb flowsheets. The batch flotation tests evaluated
rougher kinetics by investigating various depressant dosages, pH levels and the
effect of reagent type and dosage. Of the 28 flotation tests performed, 12 of
them were completed as bulk Cu-Pb with the remaining 16 following a sequential
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Cu-Pb-Zn style of flowsheet. Rougher kinetics were investigated with 13 tests (5
bulk and 8 sequential).
Cleaner flotation testwork investigated the sequential grade–recovery
relationships for copper, lead and zinc by observing the effect of regrind, pulp
chemistry and reagent/dosages. Bulk cleaner flotation explored various
separation methods to produce saleable copper and lead concentrates.
18.3.4.1 Effect of Primary Grind
Batch flotation also determined the effect of primary grinding in the target P 80
range of 52 to 65 microns. Finer grinding demonstrated minimal effect in terms
of Cu/Zn selectivity and only minor increases in Zn recovery to the Cu
concentrate was observed. Increased liberation, at finer grinding, improved
selectivity within the Pb circuit, lowering the recovery of Zn to the Pb concentrate.
18.3.4.2 Rougher Flotation
In the copper circuit, detrimental effects on copper recovery were observed when
using cyanide alone or in a complex as a zinc depressant. MBS (sodium
metabisulfite) and ZnSO 4 (zinc sulphate) were tested as well and as a result,
ZnSO 4 used alone was found to be effective enough in improving selectivity
between Cu and Zn in the copper circuit.
The same conclusion was drawn for the lead circuit.
In the zinc circuit, sphalerite flotation operated well with high rougher recoveries
and no particular problems were encountered.
18.3.4.3 Cleaner Flotation
In the copper cleaning circuit, the use of SO 2 to pH of 4.5 yielded better
selectivity than a high pH circuit in terms of Zn and Pb rejections. Several
depressant combinations and dosages were tested with MBS, NaCN (sodium
cyanide) and ZnSO 4 to improve selectivity, but no significant improvement was
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observed. NaCN, however, had adverse effect upon flotation, generating a dull
dark froth.
In the lead cleaning circuit, MBS and ZnSO 4 were tested as depressants and it
produced unfavourable results. It did not prevent zinc from floating and the
concentrate was high in zinc. The best results were obtained with a 3:1
ZnSO 4 :NaCN complex.
In the zinc circuit, sphalerite responded again well to flotation with high cleaner
recoveries. No regrinding was necessary.
18.3.4.4 Bulk Cu-Pb Rougher Flotation
Rougher recoveries for Cu and Pb to the bulk Cu-Pb rougher concentrates were
consistent throughout all tests involving different depressants attempting to
improve Zn rejection. MBS, NaCN and ZnSO 4 were tried as depressants.
Different separation methods were investigated over the course of the testing
program to produce separate Cu and Pb products from flotation of a bulk Cu-Pb
concentrate. However, the results obtained suggested that the fully sequential
approach (Cu/Pb/Zn) offered the most promise for the Estrades ore compared to
the bulk flotation (Cu+Pb/Zn) followed by separation of the bulk.
18.3.4.5 Locked Cycle Tests
The fully sequential Cu-Pb-Zn flowsheet was studied further through confirmatory
locked cycle testing. Two closed loop, six cycle, 2-kg tests were completed. The
second locked cycle test (LCT-2) was completed under similar conditions to the
first one (LCT-1) excluding a 4 th Cu cleaner stage. Slightly more reagent was
added into the initial stage of Cu cleaning in comparison to LCT-1. A regrinding
stage was used in both Cu and Pb cleaning circuits for the rougher concentrates.
The first zinc cleaner tail was not recycled in the circuit but rather discarded.
The results obtained for LCT-2 are indicated below and the flowsheet is
presented in Figure 18-1. The primary grind was performed at a P 80 of 54 µm.
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Table 18-4
Results Obtained for the Locked Cycle Test LCT-2
Calculations by SGS Lakefield
Product Weight Assays, %, g/t % Distribution
% Cu Pb Zn Au Ag Cu Pb Zn Au Ag
Cu Conc 1.8 20.6 4.8 13.7 83.0 2526 54.7 6.4 1.9 31.0 21.5
Pb Conc 1.9 2.1 42.1 15.6 74.4 2211 5.9 59.3 2.3 29.1 19.7
Zn Conc 21.2 0.5 0.9 56.0 3.2 215 14.5 15.0 94.0 14.2 21.8
Zn 1st Cl Tail 7.6 0.7 0.8 6.8 3.9 318 7.5 4.7 4.1 6.2 11.5
Zn Rghr Tail 70.0 0.1 0.2 0.4 1.0 73 8.5 10.0 2.4 14.2 24.3
Head (calc.) 102.0 0.7 1.3 12.7 4.8 209 91.1 95.5 105 94.8 98.7
GENIVAR did its own calculations from the results obtained with LCT-2 by SGS
Lakefield and used the average of the last two cycles for the weights and grades
of the products. This calculation shows that a steady state was nearly reached in
the last two cycles as demonstrated by the grades calculated from all of the
products generated by the entire locked cycle test (Head-calc) which compare
very well to the averages of the last two cycles (Head) in Table 18-5 below. This
is also confirmed by the weight obtained per cycle from the average of the last
two and the one obtained from all the products generated during the cycle test
which are fairly close. The detailed calculations are shown in Appendix 1.
Table 18-5
Results Obtained for the Locked Cycle Test LCT-2
Calculations by GENIVAR
Product Weight Assays, %, g/t % Distribution
% Cu Pb Zn Au Ag Cu Pb Zn Au Ag
Cu Conc 1.7 20.6 4.7 13.7 82.6 2522 60.2 6.7 1.9 32.6 21.7
Pb Conc 1.8 2.1 42.1 15.6 74.8 2227 6.6 62.1 2.2 30.9 20.1
Zn Conc 20.7 0.5 0.9 56.0 3.2 215 16.0 15.7 89.8 15.0 22.1
Zn 1st Cl Tail 7.4 0.7 0.8 6.8 3.9 318 8.1 4.9 3.9 6.5 11.6
Zn Rghr Tail 68.3 0.1 0.2 0.4 1.0 73 9.1 10.6 2.3 15.0 24.5
Head 100.0 0.6 1.2 12.9 4.4 202 100 100 100 100 100
Head-calc 0.7 1.3 12.7 4.8 209
The results of Table-18-5 were used to do the financial analysis of the project.
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Figure 18-1
Flowsheet for the Locked Cycle Test LCT-2
Figure 18-1
Flowsheet for the Locked Cycle Test LCT-2
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19. MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES
19.1 Introduction
To delineate the Estrades mineral deposit, drill holes were drilled from surface as
well as from underground workings.
A database for the Old Mine, Main and Main East Zones, was provided by Scott
Wilson Roscoe Postle Associates Inc. (SWRPA). This is the same database that
SWRPA used for the resource evaluation done and detailed in a 43-101 technical
report issued in November 2006. This database was developed from a Prolog
database supplied by Cogitore Resources.
This database was created with the Gemcom software and includes all the
coordinates of the drill holes’ collars in the UTM NAD83 system and in a local
coordinates system. The point of origin of this local grid, 2,400 m W and 0 m N
at a 5,000 m elevation, corresponds to the following UTM point: 654276 m E and
5494554 m N at a 285.3 m elevation. The North of this grid points towards the
348.18° azimuth. All of the work done on the resource estimation and the
localization of the planned underground workings are based on this grid system.
The drill hole database from SWRPA also includes the drill hole deviation tests,
the geological information observed from the drill holes and the Zn, Cu, Pb, Ag
and Au assays with a NSR (Net Smelter Return) calculation as well as an
estimation of the density.
Reserve calculations were done by GENIVAR. For the Old Mine and the Main
Zone, they were done from a 3D block model done by SWRPA. For the Central
Zone, the resource and reserve calculations were done by GENIVAR using a
sections method. The Main East Zone was not deemed economic.
According to the CIM standards, the spacing between drill holes and the
confidence level of both the geological and mineralization continuity allow the
resources to be categorized as ‘’indicated’’ and the reserves as ‘’probable’’. No
chip samples taken from the underground drifts were used to carry out the
calculations.
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19.2 Resource Estimate for the Old Mine, the Main and Main East Zones
For the purpose of this study, the NI 43-101 compliant resource estimate of Scott
Wilson Roscoe Postle completed in November 2006 for the Old Mine, the Main
and Main East Zones, was used.
For the SWRP resource estimate, original assays were used for interpretation of
the mineralized envelope. Zn, Cu, Au and Ag grades were converted into dollar
values with no value given to the Pb in the ore. A cut-off grade of 120$/t NSR
and a minimum width of 2 metres were used for the interpretation of the
envelope.
Table 19-1 indicates the results obtained by SWRP.
Table 19-1
Resource Estimate for the Old Mine, Main and Main East Zones, by
Scott Wilson Roscoe Postle
Tonnes Zn % Cu % Pb % Au g/t Ag g/t
INDICATED
Old Mine 54,000 13.06 0.88 1.35 7.64 196.00
Main Zone,190-415 337,000 10.71 0.79 1.02 6.01 186.00
Main Zone, below 415 170,000 8.43 0.54 0.64 2.87 144.00
Main East Zone 31,000 1.97 2.38 0.31 5.03 52.00
Total Indicated 592,000 9.82 0.81 0.90 5.21 168.00
INFE RRED
Main 4,000 8.69 0.54 0.52 1.88 150.00
Main East 28,000 6.88 0.31 0.40 2.50 60.00
Total Inferred 32,000 7.12 0.34 0.42 2.41 73.00
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19.3 Resource Estimate in the Central Zone
The resource evaluation for the Central Zone was done by GENIVAR using
Promine software with a method by sections. The potential zone includes the
area from section 1800 W to 1675 W. 17 drill holes yielded values allowing to
define the potentially economic area. The evaluation of this sector is complicated
by the occurrence of faults splitting the zone. Some holes intersected the same
zone twice as a result of these faults as shown in Figure 19-1. As illustrated in
Figure 19-2, the Central Zone has been divided in three areas, the South, Center
and North.
Even though the drilling density is relatively limited, the confidence level in the
continuity of the mineralization is acceptable for the Central Zone. Because of
the difficulty resulting from the occurrence of the faults limiting the zones, the
resource calculation was undertaken through blocks on transverse sections. The
mineralization outline and the faults position are based upon the geological
interpretation provided by Cogitore and reviewed by GENIVAR. The horizontal
thickness of the zone was established at a minimum of 2 m and the blocks
vertical extension was limited to 25 m. The distance between sections was set at
25 m. But because the drilling pattern is not regular on these sections, a
weighting factor has been applied to determine the tonnage of blocks which were
too far apart (over 25 m of distance). The estimate was done within the volume
determined by the geological interpretation.
Beyond the limits used for the resource estimate calculation, other drill holes
have intersected significant mineralization over small widths. Considering the
lack of continuity and information due to the presence of faults, these
intersections have not been included in the resource calculation.
The results of the ‘’indicated’’ resource estimate by section are presented in
Table 19-2 and illustrated in Figure 19-2.
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Table 19-2
Resource Estimate in the Central Zone by GENIVAR
Section Area Tonnes Zn % Cu % Pb % Au g/t Ag g/t
Indicated Resources
1700W Center 1,200 3.39 1.40 0.50 2.10 82.60
1750W Center 5,800 18.40 0.69 1.40 8.60 184.20
1800W Center 4,200 6.81 0.61 0.70 3.10 65.50
Sub-total 11,200 12.43 0.74 1.04 5.84 128.88
1700W North 11,000 7.94 0.91 1.00 6.60 96.80
1750W North 19,000 15.34 0.59 0.90 8.00 117.20
1800W North 21,000 5.44 0.86 0.30 2.60 100.60
Sub-total 51,000 9.62 0.77 0.67 5.45 105.88
1675W South 16,500 4.27 2.57 0.50 2.90 155.80
1700W South 12,800 3.47 1.87 0.30 1.80 108.00
1725W South 8,200 6.73 0.98 0.80 0.80 75.80
1750W South 25,700 8.60 0.47 0.70 5.30 107.20
1750W inf South 7,200 13.34 0.79 1.80 1.60 204.80
1775W South 6,700 6.78 0.78 1.10 0.80 128.40
1800W South 9,200 4.86 2.41 0.90 3.90 180.60
Sub-total 86,300 6.69 1.38 0.76 3.09 131.21
Total 148,500 8.13 1.13 0.75 4.10 122.38
Inferred Resources (lower area with an influence of 25 m)
Total 18,700 5.40 0.75 0.89 1.61 60.26
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At proximity, other drill holes intersected interesting values but on very small
widths. The extension of these intersections, taking into account the fault
system, is not yet defined.
19.4 Reserve Estimate
19.4.1 Reserve Estimate in the Old Mine
The ‘’probable’’ reserves in the Old Mine, with a 10% dilution, were estimated at
51,900 tonnes grading 11.75 %Zn, 0.80%Cu, 1.22 %Pb, 6.90 g/t Au and 177.60
g/t Ag, with about 87% of the total ‘’indicated’’ resources extracted with long-hole
stoping. The mined reserves outline is illustrated in Figure 19-3 and the detailed
results are presented in Table 19-3.
Table 19-3
Estimate of the Probable Reserves in the Old Mine
Ore Panel Tonnes Zn % Cu % Pb % Au g/t Ag g/t
D 6,300 9.75 0.58 0.82 6.16 126.38
E 9,700 16.51 0.88 1.64 8.57 229.77
F 13,900 10.90 1.08 1.19 6.60 196.84
G 22,000 10.76 0.65 1.16 6.56 157.06
Total 51,900 11.75 0.80 1.22 6.90 177.60
The remaining resources in the Old Mine are indicated in Table 19-4.
Table 19-4
Remaining Indicated Resources in the Old Mine
Tonnes Zn % Cu % Pb % Au g/t Ag g/t
Total 6,820 13.85 0.88 1.42 7.97 200.44
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19.4.2 Reserve Estimate in the Main Zone
The mine plan has identified ‘’probable’’ reserves between levels 190 and 415.
About 82% of the total ‘’indicated’’ resources will be mined without dilution,
following the vein width, using the shrinkage stoping mining method. Table 19-5
presents the results and Figure 19-4 illustrates the ‘’probable’’ reserve outline.
Table 19-5
Estimate of the Probable Reserves in the Main Zone
Stope Tonnes Zn % Cu % Pb % Au g/t Ag g/t
1 West 22,500 8.38 1.85 1.23 4.45 215.17
1 Center 30,150 7.38 0.33 0.90 6.52 107.29
1 East 22,700 6.08 0.83 0.40 11.60 160.74
2 West 27,400 9.72 0.74 1.08 4.70 227.27
2 Center 23,050 14.18 0.49 1.98 6.33 201.51
2 Center-East 18,100 8.19 1.14 0.74 4.22 225.08
2 East 15,900 6.62 0.78 0.36 3.99 86.86
3 West 25,000 8.61 0.70 0.84 3.77 156.07
3 Center 28,900 16.90 0.55 1.33 8.50 235.24
3 East 24,300 10.87 0.81 0.70 3.63 184.67
4 Center-East 16,050 14.21 0.61 1.79 5.14 216.98
4 East 21,000 13.20 1.24 0.69 2.69 223.93
Total 275,050 10.43 0.81 1.01 5.60 186.94
The remaining resources in the Main Zone are shown in Table 19-6 below.
Table 19-6
Remaining Indicated Resources in the Main Zone
Tonnes Zn % Cu % Pb % Au g/t Ag g/t
Total 61,950 11.95 0.70 1.06 7.83 181.01
19.4.3 Reserve Estimate in the Central Zone
Three stopes are planned in the Central Zone. Taking into account the available
information and the overall configuration of the mineralization, the middle portion
of the Central Zone cannot be mined. Table 19-7 and Figure 19-5 present a
summary of the reserve calculations.
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Table 19-7
Estimate of the Probable Reserves in the Central Zone
Area Tonnes Zn % Cu % Pb % Au g/t Ag g/t
North 39,000 10.11 0.81 0.62 5.80 110.06
South 2 47,050 5.58 1.54 0.58 4.10 127.09
South 1 38,900 7.33 1.52 0.96 1.73 144.03
Total 124,950 7.54 1.31 0.71 3.89 127.05
About 84% of the ‘’indicated’’ resources will be mined, without dilution, following the vein
width, using the shrinkage stoping mining method.
The remaining resources are shown in Table 19-8 below.
Table 19-8
Remaining Indicated Resources in the Central Zone
Section Tonnes Zn % Cu % Pb % Au g/t Ag g/t
Total 23,550 11.25 0.18 0.96 5.21 97.66
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20. OTHER RELEVANT DATA AND INFORMATION
There is no additional information to be included in this section. All of the other
chapters provide the reader with enough data and information in order to have a
proper understanding of the project.
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21. INTERPRETATION AND CONCLUSIONS
Resource and Reserve Estimate
According to the CIM standards, the spacing between drill holes and the
confidence level of both the geological and mineralization continuity allow the
resources to be categorized as ‘’indicated’’ and the reserves as ‘’probable’’. The
total reserves were estimated at 451,900 tonnes grading 9.78 %Zn, 0.95 %Cu,
0.95 %Pb, 5.28 g/t Au and 169.31 g/t Ag (including no dilution for both the Main
and Central Zones and a 10% dilution for the Old Mine).
For the Central Zone, even though the drilling density is relatively limited, the
confidence level in the continuity of the mineralization is acceptable.
Although density measurements were made for many of the holes drilled through
the Main zone, they were not made on a systematic basis for all holes drilled in
the late 80’s by Teck Exploration and Golden Hope. For the holes drilled by
Cogitore in the Main Zone, density measurements were made systematically
both for the ore and for the immediate host rocks. Because of the lack of
systematic density measurements, especially in the Central Zone, the specific
gravity of the ore represents the least accurate parameter in the resource and
reserve estimates.
Environment
The Quebec Ministry of Environment has issued in 2007 a certificate of
authorization for underground mine dewatering and ore extraction at a rate of
500 tonnes per day. However, the operating scenario of the current feasibility
study uses a production rate of 1,000 tonnes per day and therefore, the
certificate of authorization will have to be modified prior to the development and
production start-up, when the go ahead will be given.
Metallurgical Test Work
The metallurgical recoveries achieved in the laboratory with the locked cycle test
LCT-2 were used to calculate the revenues of the financial analysis. However,
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because of differences in grades between the head sample used for LCT-2 and
the average grades of the mineral reserve, the quantities of concentrates
produced were adjusted in proportion to the grades of the reserve, assuming that
recoveries would stay the same, which is not necessarily the case for each
metal.
Financial Analysis
Indications of market treatment charges were obtained directly from smelters
based in Canada for the three concentrates. The custom milling costs and the
capital expenses for mill modifications were estimated by GENIVAR from
detailed reviews of potential mill sites for the Estrades ore. However, no formal
agreement has been negotiated with any smelter and concentrator. Negotiations
will have to take place when Cogitore Resources is ready to go ahead with the
project in order to determine more precisely these costs.
Mining costs as well as pre-production capital expenditures were established by
GENIVAR and are shown in section 25.
The financial analysis shows that the project is marginal for the base case
scenario selected, with a rate of return of 7.4% and a negative net present value
of $1.4 M at a discount rate of 10%. But the sensitivity analysis indicates it has
an interesting economic potential and demonstrates a high degree of sensitivity
to the exchange rate and the zinc price in particular.
The study has been spread over an 18 month period and the cost of many items
has been estimated some time ago. Since then, some of these costs have
changed significantly, including the cost of the diesel fuel which has dramatically
increased recently.
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22. RECOMMENDATIONS
Some further drilling should be done in the Central Zone to increase the
confidence level and optimize stope development;
Laboratory test work should be done to adapt the flowsheet to the
concentrator which would be eventually selected and run the tests with a
head sample having grades closer to the grades of the mineral reserve to
confirm the metallurgical results;
In light of the current zinc and lead prices, it is recommended to delay the
production decision and wait for better prices, particularly for zinc;
An overall cost update should be undertaken before taking the final
decision to go ahead with the project;
The property demonstrates a very good potential for the discovery of other
volcanogenic massive sulphide deposits and therefore, exploration should
continue in that area. The discovery of additional ore zones could
dramatically improve the economics of the Estrades project.
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23. REFERENCES
Arsenault, J.L.;Mouge, P.; 2003: Levé magnétique et électromagnétique héliporté
sur le projet Estrades (70-428), Géophysique GPR International Inc., présenté à
Corporation Inmet.
Brouillette, Jean-François; Morin, Sébastien; November 2007: Estrades Project –
Mechanical, Electrical and Building Infrastructures Evaluation, GENIVAR.
Clark, L.A.; 1986 : Geochemical Exploration of the Estrades Deposit Alteration
Zones, Estrades Township, Quebec, for Golden Hope Resources Inc. and Teck
Exploration Ltd., Unpublished, Private Company Report.
Cloutier, J.-P.; 2005: Technical Report, Form 43-101F1 (Volume 1 of 3),
Estrades-Newiska Property, Estrées, Orvilliers and Estrades Townships,
Quebec, NTS 32E10, Prepared for Woodruff Capital Management Inc..
Goffaux, Daniel; Jauron, Richard; Duchesne, Pierre; Gagnon, Garant; May 2006:
Preliminary Evaluation of Capital and Operating Costs for the Estrades Project,
Met-Chem Canada Inc..
Lang, Jake; Imeson, Dan; December 14, 2007: An Investigation into the
Recovery of Cu/Zn/Pb from Estrades Ore, SGS Lakefield Research Limited.
Nathalie Gauthier; Août 2007: Simulation du réseau de ventilation – Mine
Estrades, GENIVAR.
O’Dowd, P.; Welch, M.; Landry, A.C.; 1989: Noramco Exploration – Estrades
Property, Estrades Deposit ‘’Main Zone’’, 1989 Definition Drilling Program and
Mineral Inventory, Golden Hope Resources Inc., Golden Group Exploration Inc.,
Breakwater Resources Inc., MRN GM 53422.
Racine, Louis; Juin 1991: Traitement du minerai d’Estrades au concentrateur de
la Division Matagami, Division Matagami, Minéraux Noranda inc..
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Salmon, Bernard; November 20, 2006: Mineral Resource Estimate for the Main
Zone of the Estrades Project, Northwestern Québec, Canada, Scott Wilson
Roscoe Postle Associates Inc..
Taylor, J.R.; February 1990: Breakwater Resources Ltd., Estrades Project,
Custom Milling Option, Final Report.
Welch, M.J.; 1995: Metal Zoning, Geochemistry and Alteration of the Archean,
Estrades Zn-Au Massive Sulphide Deposit, Northwestern Quebec, Canada;
Department of Earth Sciences, Laurentian University, Sudbury, Ontario; Thesis
Presented in Partial Fulfillment of the Requirements for the Degree of Master of
Science.
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24. DATE AND SIGNATURES PAGE WITH CERTIFICATES OF CONSENTS
24.1 Date and Signatures Page
This report entitled ‘’FEASIBILITY STUDY – ESTRADES PROJECT -
TECHNICAL REPORT 43-101’’ and dated October 24 th 2008, has been prepared
by the following persons:
Dated in Rouyn-Noranda, Quebec
October 24 th , 2008
Signed and Sealed by:
GENIVAR
Michel Garon, Eng.
Signed and Sealed by:
GENIVAR
Nicole Rioux, Geo.
Signed and Sealed by:
GENIVAR
Mario Blanchette, Eng.
Signed and Sealed by:
GENIVAR
Marc Lafontaine, Eng.
Signed and Sealed by:
GENIVAR
Serge Ouellet, Eng. Ph. D.
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24.2 Certificate and Consent of Michel Garon, Eng.
I, Michel Garon, do hereby certify that:
1) I am an Engineer employed as Senior Mining Engineer by GENIVAR at 152 Murdoch
Avenue, Rouyn-Noranda, Quebec.
2) I received a Bachelor’s Degree in Applied Sciences from the Université de Montréal
(Montréal, Québec) in 1975 and a Master’s Degree from the same university in 1976.
3) I am a registered member of the Ordre des Ingénieurs du Québec (OIQ member no.
28151).
4) I have over 30 years of experience as an engineer in the mining industry. My experience
has been acquired mostly with Noranda. I have been working for GENIVAR since June
2006 as Senior Mining Engineer.
5) I have read the definition of “qualified person” set out in Regulation 43-101 (“R 43-101”)
standards for disclosure for mineral projects and certify that by reason of my education,
affiliation with a professional association (as defined in R 43-101) and past relevant work
experience, I fulfill the requirements to be a “qualified person” for the purposes of
R 43-101.
6) I am responsible for the preparation of the technical report entitled “Feasibility Study-
Estrades Project-Technical Report 43-101’’ dated October 24 th , 2008. I coordinated the
preparation of the document and made a visit to the Estrades site.
7) I am “independent” (as such term is defined in Section 1.4 of R 43-101) of Cogitore
Resources Inc..
8) I have never had any prior involvement with the property that is the subject of the
Technical Report.
9) As of the date of this certificate, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be
disclosed to make the Technical Report not misleading.
10) I have read R 43-101, Appendix 43-101A1 and the Technical Report which has been
prepared in compliance with that instrument and form.
11) I consent to the filing of the Technical Report with any stock exchange and other
regulatory authority and any publication by them for regulatory purposes, or presentation
of excerpts or a summary, including electronic publication in the public company files on
their websites accessible by the public, of the Technical Report.
Dated: October 24 th , 2008
Signed and Sealed by: Michel Garon, Eng.
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24.3 Certificate and Consent of Nicole Rioux, Geo.
I, Nicole Rioux, do hereby certify that:
1) I am a Professional Geologist employed as Senior Geologist by GENIVAR at 152
Murdoch Avenue, Rouyn-Noranda, Quebec.
2) I received a Bachelor’s Degree in Geology from the Université du Québec à Montréal
(Montréal, Québec) in 1981.
3) I am a registered member of the Ordre des Géologues du Québec (OGQ licence no.
326).
4) I have over 25 years of experience as a geologist. My experience has been acquired
mostly in exploration, mining geology and teaching in geology since 1982. I worked for
Promine Consultant and later joined GENIVAR in February 2005 as Senior Geologist.
5) I have read the definition of “qualified person” set out in Regulation 43-101 (“R 43-101”)
standards for disclosure for mineral projects and certify that by reason of my education,
affiliation with a professional association (as defined in R 43-101) and past relevant work
experience, I fulfill the requirements to be a “qualified person” for the purposes of
R 43-101.
6) I am responsible for the preparation of the Technical Report entitled “Feasibility Study-
Estrades Project-Technical Report 43-101’’ dated October 24 th , 2008 and I was
particularly responsible for the resource and reserve estimates of the Central Zone. I did
not make a visit to the Estrades site.
7) I am “independent” (as such term is defined in Section 1.4 of R 43-101) of Cogitore
Resources Inc..
8) I have never had any prior involvement with the property that is the subject of the
Technical Report.
9) As of the date of this certificate, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be
disclosed to make the Technical Report not misleading.
10) I have read R 43-101, Appendix 43-101A1 and the Technical Report which has been
prepared in compliance with that instrument and form.
11) I consent to the filing of the Technical Report with any stock exchange and other
regulatory authority and any publication by them for regulatory purposes, or presentation
of excerpts or a summary, including electronic publication in the public company files on
their websites accessible by the public, of the Technical Report.
Dated: October 24 th , 2008
Signed and Sealed by: Nicole Rioux, Geo.
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24.4 Certificate and Consent of Mario Blanchette, Eng.
I, Mario Blanchette, do hereby certify that:
1) I am an Engineer employed as Senior Mining Engineer by GENIVAR at 1075, 3 rd Avenue
East, Val-d’Or, Quebec.
2) I received a Bachelor’s Degree in Applied Sciences from the Université Laval (Québec
city, Québec) in 1990.
3) I am a registered member of the Ordre des Ingénieurs du Québec (OIQ member no.
105651).
4) I have about 18 years of experience as an engineer in the mining industry. I have been
working for GENIVAR since October 2006 as Senior Mining Engineer.
5) I have read the definition of “qualified person” set out in Regulation 43-101 (“R 43-101”)
standards for disclosure for mineral projects and certify that by reason of my education,
affiliation with a professional association (as defined in R 43-101) and past relevant work
experience, I fulfill the requirements to be a “qualified person” for the purposes of
R 43-101.
6) I am responsible for the preparation of the section 25.1 Mining of the Technical Report
entitled “Feasibility Study-Estrades Project-Technical Report 43-101’’ dated October 24 th ,
2008. I did not make a visit to the Estrades site.
7) I am “independent” (as such term is defined in Section 1.4 of R 43-101) of Cogitore
Resources Inc..
8) I have never had any prior involvement with the property that is the subject of the
Technical Report.
9) As of the date of this certificate, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be
disclosed to make the Technical Report not misleading.
10) I have read R 43-101, Appendix 43-101A1 and the Technical Report which has been
prepared in compliance with that instrument and form.
11) I consent to the filing of the Technical Report with any stock exchange and other
regulatory authority and any publication by them for regulatory purposes, or presentation
of excerpts or a summary, including electronic publication in the public company files on
their websites accessible by the public, of the Technical Report.
Dated: October 24 th , 2008
Signed and Sealed by: Mario Blanchette, Eng.
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24.5 Certificate and Consent of Marc Lafontaine, Eng.
I, Marc Lafontaine, do hereby certify that:
1) I am an Engineer employed as Senior Mineral Processing Engineer by GENIVAR at
1462, rue de la Québéçoise, Val-d’Or, Quebec.
2) I received a Bachelor’s Degree in Applied Sciences from the Université Laval (Québec
city, Québec) in 1994.
3) I am a registered member of the Ordre des Ingénieurs du Québec (OIQ member no.
114548).
4) I have over 10 years of experience as an engineer in the mining industry. I have been
working for GENIVAR since June 2005 as Senior Mineral Processing Engineer and
Manager – Mineral Processing, since January 2008.
5) I have read the definition of “qualified person” set out in Regulation 43-101 (“R 43-101”)
standards for disclosure for mineral projects and certify that by reason of my education,
affiliation with a professional association (as defined in R 43-101) and past relevant work
experience, I fulfill the requirements to be a “qualified person” for the purposes of
R 43-101.
6) I am responsible for the coordination of the work done for sections 18.0 and 25.2, related
to mineral processing, of the Technical Report entitled “Feasibility Study-Estrades
Project-Technical Report 43-101’’ dated October 24 th , 2008. I did not make a visit to the
Estrades site.
7) I am “independent” (as such term is defined in Section 1.4 of R 43-101) of Cogitore
Resources Inc..
8) I have never had any prior involvement with the property that is the subject of the
Technical Report.
9) As of the date of this certificate, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be
disclosed to make the Technical Report not misleading.
10) I have read R 43-101, Appendix 43-101A1 and the Technical Report which has been
prepared in compliance with that instrument and form.
11) I consent to the filing of the Technical Report with any stock exchange and other
regulatory authority and any publication by them for regulatory purposes, or presentation
of excerpts or a summary, including electronic publication in the public company files on
their websites accessible by the public, of the Technical Report.
Dated: October 24 th , 2008
Signed and Sealed by: Marc Lafontaine, Eng.
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24.6 Certificate and Consent of Serge Ouellet, Eng. Ph.D.
I, Serge Ouellet, do hereby certify that:
1) I am a Professional Engineer employed as Senior Engineer and Director of Geotechnical
Services by GENIVAR at 1075, 3 e Avenue Est, Val-d’Or, Quebec.
2) I received a Bachelor’s Degree in Engineering Geology from the Université Laval
(Québec city, Québec) in 1992 and a Master’s Degree from the same university in 1995.
I received a Philosophiae Doctor Degree in Environmental Sciences from the Université
du Québec (Rouyn-Noranda, Quebec) in 2006.
3) I am a registered member of the Ordre des Ingénieurs du Québec (OIQ member no.
109358).
4) I have 12 years of experience as an engineer and 7 years as engineer in mining
environment. I have been working for GENIVAR since August 2006 as Senior Engineer.
5) I have read the definition of “qualified person” set out in Regulation 43-101 (“R 43-101”)
standards for disclosure for mineral projects and certify that by reason of my education,
affiliation with a professional association (as defined in R 43-101) and past relevant work
experience, I fulfill the requirements to be a “qualified person” for the purposes of
R 43-101.
6) I am responsible for the preparation of section 25.5 Environmental Considerations of the
Technical Report entitled “Feasibility Study-Estrades Project-Technical Report 43-101’’
dated October 24 th , 2008. I made a visit to the Estrades site in November 2006.
7) I am “independent” (as such term is defined in Section 1.4 of R 43-101) of Cogitore
Resources Inc..
8) I have never had any prior involvement with the property that is the subject of the
Technical Report.
9) As of the date of this certificate, to the best of my knowledge, information and belief, the
Technical Report contains all scientific and technical information that is required to be
disclosed to make the Technical Report not misleading.
10) I have read R 43-101, Appendix 43-101A1 and the Technical Report which has been
prepared in compliance with that instrument and form.
11) I consent to the filing of the Technical Report with any stock exchange and other
regulatory authority and any publication by them for regulatory purposes, or presentation
of excerpts or a summary, including electronic publication in the public company files on
their websites accessible by the public, of the Technical Report.
Dated: October 24 th , 2008
Signed and Sealed by: Serge Ouellet, Eng. Ph.D.
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25 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON
DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES
25.1 Mining
25.1.1 Mine Dewatering and Rehabilitation
It is difficult to assess precisely the time and work which will be necessary to
dewater and rehabilitate the mine workings. A period of about 4 months of work
has been estimated with 5 employees, 24 hours per day over 7 days a week, for
the dewatering activity. This amounts to a total of $850,000 and a 15%
contingency has been added to it.
Mine rehabilitation will be carried out as the ramp becomes accessible and
therefore will be undertaken during dewatering. It is assumed that rehabilitation
will start about a month after the beginning of dewatering and will be completed a
month after dewatering is finished. A cost of $1,500 per metre has been
evaluated for that activity and will amount to a total of $2,250,000 for 1,500 m of
ramp. A 15% contingency has also been added to that amount.
Dewatering and rehabilitation are then planned over a total period of about 5
months and another 40 days will be necessary afterwards to put in place the
powder magazines, lunch rooms, emergency exits and other required
underground infrastructures.
25.1.2 Ramp Development
A ramp with a 4.0 m height and 5.0 m width, will be driven at a grade of 15%. At
the moment, this ramp has been located about 30 m from the Main Zone, but this
could likely be reduced because the shrinkage method will be used for that zone
and as a result, blasting in the stopes won’t be severe and won’t have much
impact upon the rock integrity. The ramp will go from the 190 level down to the
415.
Productivity for the ramp development has been evaluated at 6.0 m per day. A
total of 14 months have been planned to complete its excavation.
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25.1.3 Shrinkage Stoping for the Main Zone
The mining method selected for that zone is shrinkage stoping. Although it
appears to be an old and an obsolete method when considering all the new
equipment on the market aiming at more productive methods, it fits in properly
with this type of narrow and high value mineralization where grade control is of
paramount importance. Furthermore, it does also provide good ground control as
mining progresses.
The long-hole concept developed by MET-CHEM in their preliminary evaluation
was also examined. It calls for much development but stopes are developed and
mined gradually as the ramp progresses towards the 415 level and this allows
production to start quickly. However, taking into account the small size of the
stopes, this method involves a great deal of logistics for drilling, blasting and
backfilling. This is a constraint to the numerous moves of equipment required.
Furthermore, management of the mining cycle could become extremely difficult if
ever ground problems arise, which is another risk resulting from that mining
method. Ground problems were experienced in the past and this problem could
likely resurface with a long-hole method.
The ore outline with a minimum width of 2.0 m, a dip nearly vertical and a good
linearity along strike make it appropriate for shrinkage stoping.
Every 56.0 m, an access will be driven from the ramp to the Main Zone. From
this access, drifts will be excavated in the mineralization at vein width and a
height of 2.4 m. Each stope will have a raise done with an Alimak machine which
will follow mineralization. We assume in the current study that mineralization
offers a fairly constant dip which, if it’s not the case, will require the use of
conventional raising methods which are lightly less costly, a little better for grade
control but much slower. Raises will be timbered and will be used for both
miners and equipment haulage.
Each stope will stop 5.0 m from the level above. Vertical pillars of 4.0 m will also
be left between them when the quantity of ore is too voluminous for a single
stope.
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In this study, development in the ore is done with slushers and Cavo machines
but small scooptrams could also be used where the width allows it. The advance
rate is planned at 6.0 m per day.
Mucking is planned to be carried out with 5 yard 3 scooptrams and 26 ton trucks.
With this method no waste backfilling will be required. The stopes’ walls will be
supported as mining progresses. But backfilling could be undertaken and would
be more economical than hauling waste to surface. This issue could be
examined at the outset of the project.
Productivity in the stopes varies mainly with the ore width as well as with the ore
specific gravity. The average rate of advance during mining has been estimated
at 1.9 m per worker per day.
About 82% of the total indicated resources will be mined with no dilution. A total
of 275,050 tonnes of ore will be mined from the Main Zone.
As ramp development will progress toward the 415 level, access, stope
development as well production will be undertaken as soon as possible in the
upper levels while making sure it doesn’t interfere with other activities.
Figure 25-1-1 illustrates a longitudinal as well as a 3D view of the Main Zone and
Figures 25-1-2 and 25-1-3 show the plan view of two levels.
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Figure 25-1-1 Main Zone Longitudinal 3D View
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Figure 25-1-2 Main Zone Elevation 4 749 m
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Figure 25-1-3 Main Zone Elevation 4 637 m
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25.1.4 Shrinkage Stoping for the Central Zone
Shrinkage stoping is also recommended for the Central Zone. This zone is
disturbed by a number of fault planes which makes it very difficult to be mined
with the long-hole method. This would require very extensive development since
these faults cut the zone in three distinct sectors. Only 4 stopes will be mined in
that area and mining will be similar to what will be done in the Main Zone,
following the width of the vein, no dilution, drifts with a height of 2.4 m, etc….
A total of 124,950 tonnes of ore will be mined and as a result 84% of the total
indicated resources will be mined with no dilution.
However, in order to reduce the ventilation requirements, ore handling and
hauling in this area will be carried out with equipment on track. Two drifts with a
height of 2.4 m and a width of 2.8 m will be excavated from the Main Zone area
to the Central Zone and will allow for the access as well as ventilation of that part
of the underground operation. These drifts will have a length of about 450.0 m.
Their excavation will require a total of 5 months.
Figure 25-1-4 shows a view of both the Main and Central Zones and Figure 25-1-
5 illustrates a plan view of one level of the Central Zone.
25.1.5 Long-Hole Method for the Old Mine
Some rehabilitation work will have to be undertaken in order to access and mine
safely the ore panels left in that area. A provision of $1,200,000 was estimated
for that work. The long-hole method is planned to recover these panels as most
of the development is apparently complete to suit that mining method.
Each panel will be mined one after the other in 30.0 to 35.0 m blocks separated
by 5.0 m pillars. A total of 51,900 tonnes of ore will be mined and this includes a
10% dilution. About 87% of the total indicated resources will be extracted.
25.1.6 Underground Ventilation
GENIVAR did some simulation on the ventilation network of the Estrades
underground project and determined the power required to supply the
necessary volume (Nathalie Gauthier, GENIVAR, Simulation du réseau de
ventilation-Mine Estrades, Août 2007).
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The air volume necessary for this underground operation has been evaluated in
compliance with Quebec regulations and particularly, the minimum requirements
for diesel engines.
The available information indicates there is in place a ventilation raise (8’ x 8’)
with a man way in the old mine section and for the purpose of this study we
assume this information is valid.
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Figure 25-1-4 The Main and Central Zone
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Figure 25-1-5 The Central Zone and Access Ramp
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The mine life being short, the use of more than one raise was preferred rather
than enlarging the existing one, which will allow the raises to be operational more
rapidly. As a result, the existing raise will be extended to the last level in order to
have an emergency exit in compliance with the current regulations. A new raise
will also be excavated near the ramp (circular, 11.5’ dia., 250 m long). These
raises will supply fresh air to the underground workings. Air will be exhausted
through the ramp. A total of $1.725 M has been estimated for extending the
existing ventilation raise and drive the new one.
The total air requirement amounts to 310,000 cfm @ 13.5 ‘’ w.g. and 880 bhp.
Air velocity in the ramp will be in the order of 1,450 feet per minute.
The air requirement was determined for the diesel equipment indicated in the
table below.
Table 25-1-1 Diesel Mobile Equipment Used Underground
Vehicles Number HP per Vehicle
Scoop, 5 y 3 4 231
Scoop, 3.5 y 3 2 185
Truck, 26 tons 4 400
Tractor 8 57
Scissor Lift 2 82
Jumbo, 2 booms 2 75
Because the Central Zone will be mined with equipment on track, fresh air will be
provided to that area from the Main Zone with secondary ventilation as the
necessary volume will be minimal.
25.1.7 Underground Production Manpower
After about 6 months of underground dewatering and rehabilitation, development
will start and production will follow. This will be spread over a period of 3 years
and the production manpower required during that period is shown in the table
below for the underground sector. This doesn’t include maintenance labour.
The current evaluation was carried out as if manpower was hired by Cogitore. All
the jobs required were identified and the number of people determined.
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However, most of the work will more likely be contracted out. The costs were
therefore calculated in order to be valid for both options, considering the labour
shortage and the high costs of the current labour market.
The assumed working schedule uses 10 hr shifts, 5 days a week.
Table 25-1-2 Underground Manpower for Development and Production
Nb of Employees
Months 7 - 18 19 - 30 31 - 42
Development 25 19 0
Stoping 7 37 12
Mucking 8 8 8
Total 40 64 20
25.1.8 Ore Production
Development in mineralization will start about 12 months after the beginning of
the project and stoping will start producing ore in about the 14 th month.
Ore production will gradually ramp up and the production phase will be reached
in the 21 th month of the project with more than 77% of the planned maximum rate
of 20,000 tonnes of ore per month. The planned maximum rate will be achieved
a month later and maintained until nearly the end of operations.
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Table 25-1-3 Total Ore Production Over the Mine Life (Tonnes Milled)
Months 7 -18 19 - 30 31 - 42 Total
Main Zone
Tonnes 27,150 143,300 104,600 275,050
%Zn 7.28 9.30 12.79 10.43
%C u 0.88 0.82 0.78 0.81
%Pb 0.85 1.03 1.03 1.01
Au, g/t 7.42 5.78 4.89 5.60
Ag, g/t 151.51 181.75 203.25 186.94
Central Zone
Tonnes 3,150 36,500 85,300 124,950
%Zn 7.33 6.79 7.87 7.54
%C u 1.52 1.53 1.20 1.31
%Pb 0.96 0.84 0.65 0.71
Au, g/t 1.73 2.47 4.58 3.89
Ag, g/t 144.03 138.77 121.41 127.05
Old Mine
Tonnes 13,400 38,500 51,900
%Zn 13.33 11.20 11.75
%C u 0.74 0.82 0.80
%Pb 1.25 1.21 1.22
Au, g/t 7.44 6.71 6.90
Ag, g/t 181.18 176.36 177.60
Total
Tonnes 43,700 218,300 189,900 451,900
%Zn 9.14 9.22 10.58 9.78
%Cu 0.88 0.94 0.97 0.95
%Pb 0.98 1.03 0.85 0.95
Au, g/t 7.02 5,39 4.75 5.28
Ag, g/t 160.07 173.61 166.48 169.31
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Table 25-1-4 Ore Production from Development and Stoping (Tonnes Milled)
Months 7 - 18 19 - 30 31 - 42 Total
Main Zone
Development 15,490 17,550 33,040
Stoping 11,660 125,750 104,600 242,010
Total 27,150 143,300 104,600 275,050
Central Zone
Development 3,150 2,430 5,580
Stoping 34,070 85,300 119,370
Total 3,150 36,500 85,300 124,950
Old Mine
Development
Stoping 13,400 38,500 51,900
Total 13,400 38,500 51,900
Total
Development 18,640 19,980 38,620
Stoping 25,060 198,320 189,900 413,280
Total 43,700 218,300 189,900 451,900
Waste will be produced during development and a part of this material could
likely be dumped in empty stopes underground in order to avoid hauling it to
surface. This alternative has not been examined yet but should be considered at
the outset of the project.
Waste will come mainly from excavation of the ramp as well as from the two drifts
which will be driven from the Main Zone to the Central. Furthermore, some 76 m
of waste development per level will be carried out around the Main Zone area for
lunch rooms, electrical rooms, pumping stations, powder magazines, etc…
Table 25-1-5 Waste from Development (Tonnes)
Months 7 - 18 19 - 30 31 - 42 Total
Main Zone 103,566 52,813 156,379
Central Zone 32,579 3,064 35,643
Old Mine
Total 136,145 55,877 192,022
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25.1.9 Mining Costs
In the current market there is no significant difference between purchasing, rental
with or without a purchasing option or having the job done by a contractor.
Therefore, for a part of the mobile and fixed equipment, we’ve assumed that it will
be rented. This increases the operating costs but reduces significantly the
amount of capital outlay required at the beginning of the project.
The maintenance labour is not included in these costs.
Table 25-1-6 shows the unit costs for development.
Table 25-1-6 Unit Costs for Development
Development
Dimensions
(Width x Height)
$ Per Metre
Ramp 5.0m x 4.0m 2,802.85
Drift 4.0m x 3.5m 2,574.35
Draw Point 3.0m x 3.0m 1,649.26
Other Development, Waste 3.0m x 3.0m 1,649.26
Muck Bay 4.0m x 3.5m 2,574.35
Ore Drift 2.78m x 2.4m 1,438.75
Alimak Raise 2.78m x 2.4m 2,000.00
The following table illustrates the unit costs of various development supplies.
Table 25-1-7 Unit Costs of Various Development Supplies
Supplies
$ Per Metre
Ventilation (Temporary) 32.00
Piping 84.30
Rock Bolting 143.67
Muck, Ramp 33.00
Explosives 193.32
Telephone, Blasting Line 13.00
Drill Material (Maintenance Included) 141.40
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Table 25-1-8 presents the direct mining costs over the three year period,
including pre-production and production.
Table 25-1-8 Direct Mining Costs
Costs Per Year, $
Months 7 - 18 19 - 30 31 - 42 Total
Development
Main Zone 7,308,070 4,198,232 11,506,302
Old Mine 1,380,000 1,380,000
Central Zone 3,466,695 543,510 4,010,205
Sub-Total 12,154,765 4,741,742 16,896,507
Stoping –
Mucking
Main Zone 1,219,182 7,013,117 2,691,000 10,923,299
Old Mine 459,932 1,847,605 2,307,537
Central Zone 2,607,020 1,897,500 4,504,520
Sub-Total 1,679,114 11,467,742 4,588,500 17,735,353
Total 13,833,879 16,209,484 4,588,500 34,631,860
$/Tonne Milled 76.65
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25.2 Mineral Processing
On the mine site there is no milling facility and therefore, 6 operations were
approached in order to find one which could treat the Estrades ore on a custom
basis. All of them are located within a 175 km radius but in some cases, require a
travelling distance beyond 300 km.
Finally, two operations demonstrated interest in treating the ore and some
discussions were held with each of them. This allowed to determine preliminary
costs that were used for this study. Both options would require modifications to
the current installations with addition of new equipment. GENIVAR estimated the
costs involved in these modifications.
However, no agreement has been reached with these operations yet and this will
take place when Cogitore Resources decides to go ahead with the project.
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25.3 Site Infrastructure
25.3.1 Introduction
A site visit was made by GENIVAR during the week of April 23 rd 2007, to
evaluate the conditions of the existing infrastructure. Following the visit, a list and
description of the new surface infrastructure required to ensure proper operations
were carried out and a cost evaluation (±10%) has been produced for different
options, overhaul, replacement through purchasing or rental (Brouillette, Morin,
2007). Figure 25-3-1 shows the general arrangement of the site layout.
25.3.2 Mechanical Infrastructure
25.3.2.1 Compressed Air System
The underground requirements for compressed air have been evaluated at
3,000 CFM at 100 PSI. Two Sullair variable capacity screw type compressors,
Series 25, 300 HP, are actually on site and located in a separate room beside
the main mechanical shop. A compressed air receiver of 220 ft 3 is located
outside next to the compressor room. If compressed air requirements increase
during the operation of the mine, a diesel compressor could be added to the
system.
The compressors and the receiver appear in good condition and could likely be
used again after being overhauled. Purchasing new compressors similar to the
existing ones would cost approximately twice the price of the overhaul.
Consequently, there is no advantage to buy new ones.
25.3.2.2 Main Underground Ventilation
The ventilation requirements for the operation of the mine have been evaluated
at 310,000 CFM at 13.6 inches of water as reported by Gauthier (2007). The air
shall be heated with propane burners to ensure a minimum air temperature of
5 o C into the vent raise. At the site, infrastructure for the main ventilation is
currently non existent. During the site visit, the vent raise collar was flooded,
therefore it was impossible to assess the condition of the existing infrastructure.
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A concrete base used for supporting the propane tank of the ventilation heater is
still on site. It is actually impossible to estimate if the collar of the vent raise can
be used or not. Consequently, we assume that the entire infrastructure for the
ventilation system will have to be completely rebuilt.
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Figure 25.3.1 General Arrangement of the Site Layout
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The ventilation system will include the following items:
Four (4) 84’’ diameter axial fans of 250 HP for a total of 310,000 CFM;
Two (2) burners of 16,000,000 BTUH;
One 96,000-litre propane tank and propane piping;
Vent duct distribution into the vent raise;
Concrete base for burner;
Enlargement of the existing vent raise which reaches the collar, from
2.44 m x 2.44 m to 3.66 m x 3.66 m and raise collar rehab.
Two options have been studied for the existing vent raise: the first option is to
keep the raise as is, at 2.44 m x 2.44 m x 25 m. The second option is to enlarge
the size of the vent raise at 3.66 m x 3.66 m x 25 m. With the first option, the
motors of the four fans would be 350 HP compared to 250 HP for the second
option as the total pressure is more important with a smaller raise (17.0 inches of
water compared to 13.6 inches of water). Considering the cost of enlarging the
raise, the cost of the fans and the cost of diesel consumption, the second option
costs approximately $400,000 less than the first one based on a three year
operation. As a result, the enlargement of the existing raise of 2.44 m x 2.44 m x
25 m has been selected.
25.3.2.3 Lime and Acid Plants (Water Treatment)
The amount of water to be treated during operation of the mine has been
evaluated at 1,200 m 3 /day. The amount of water to be treated during dewatering
of the mine before normal operation has been evaluated at 4,150 m 3 /day. The
lime plant and acid plant will be designed to treat water for both of these periods.
At the mine site, there’s currently no installation for either a lime or acid plant.
The proposed system will consist of two prefabricated containers, one for the
lime plant and one for the acid plant. All the water will be pumped into the lime
plant container to be treated. After passing through a series of two static mixers
to add lime and flocculant, the treated water will be discharged into the second
sedimentation pond. Lime preparation will be performed by a system of screw
conveyor feeding dehydrated lime (1,000 Kg bags) into a mixing tank. Milk of
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lime will be pumped into a static mixer to mix lime and water. The flocculant
produced into another mixing tank will be pumped into the second static mixer.
The acid plant container will consist of one acid tank fed by a manual pump.
Then, the acid will be pumped by a peristaltic pump into the overflow pipe
between the first sedimentation pond and the polishing pond.
25.3.2.4 Process Water System
The process water consumption has been estimated at 250 USGPM.
On site, a 4’’ diameter steel pipe (process water distribution pipe) and a 6’’
diameter steel pipe (process water return pipe) have been installed from the
decantation pond to the vent raise (the last 100 m of pipes are flooded).
There are no advantages in reusing the existing steel pipes for this application,
because the pipes need to be insulated on site which will cost more than buying
pre-insulated pipes.
However, the cheapest and recommended option is to install buried pipes in the
ground.
The process water from underground (return line) will be pumped to surface
through the vent raise directly to the sedimentation pond. The process water
(distribution line) will be pumped from the polishing pond to the vent raise and
underground with a submersible pump installed into a 36’’ diameter water intake.
A heated cabin will be built over the process water intake to ensure accessibility
to the pump all year long.
25.3.2.5 Diesel and Gasoline Distribution System
The diesel distribution system has two functions; provide diesel for the genset
and provide diesel for the underground and surface equipments. The diesel
consumption for each of these applications has been established as follows:
Diesel consumption for underground and surface equipments has been
estimated at 500,000 litres/year or 9,615 litres/week.
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Diesel consumption for the genset based on a 5-day operation has been
estimated as follows:
o First year : 2,500,000 litres
o Second year : 3,750,000 litres
o Third year : 2,500,000 litres
The current diesel distribution system is not in compliance with the prevailing
regulations and therefore new equipment will have to be put in place.
Therefore, a new diesel storage tank size will be selected to provide a storage
capacity of at least a week of production. The diesel and gasoline distribution
systems will include:
One diesel 50,000 litres ULC-S653 dike tank for the genset;
One diesel 37,750 litres ULC-S653 dike tank for the genset;
One diesel pump to supply diesel daily to tanks of each generator;
One diesel 22,500 litres ULC-S653 dike tank complete with a
distribution pump station to fill vehicles;
One gasoline 10,000 litres ULC-S653 dike tank complete with a remote
dispenser unit to fill vehicles.
25.3.2.6 Distribution Systems of New and Used Oil
The new oil distribution system which is already installed on site could be used
again after upgrading it to be in compliance with regulations and make it more
convenient.
The used oil distribution system could also be used again after doing some
modifications in order to comply with regulations.
25.3.3 Buildings
25.3.3.1 Main Office
The surface area required for the main office trailers has been estimated at 2,900
ft 2 for 22 persons. The current building is in very bad condition and a major
overhaul would be required in order to use it again. Therefore, it is
recommended to rent new office trailers.
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25.3.3.2 Dry Trailers
The peak for workers at the mine site has been evaluated at about 100
employees. The requirements have then been evaluated as follows:
1,440 ft 2 for the men’s dry with 100 lockers, 100 hooks to dry
clothes and 240 ft 2 for the women’s dry (6 persons) with
showers and water-closets;
816 ft 2 for the men’s showers, water-closets and hand-sinks;
200 ft 2 for the infirmary.
The existing facility of 1,600 ft 2 is not large enough and would require a major
overhaul. Therefore, it is recommended to rent a new dry of 2,700 ft 2 .
25.3.3.3 Mechanical Shop and Cold Storage
The existing mechanical shop and cold storage are in good condition and could
be used again. The main doors require repairs for both buildings. The electrical
distribution equipment is reusable. The electrical cabling and lighting inside the
buildings have to be replaced. For the mechanical shop, a ventilation system to
provide adequate air change is non existent. The propane unit heater and
electrical unit heater will also have to be replaced. The compressed air and water
service pipes are reusable but require some repairs.
After repair and a number of modifications these buildings will be adequate for
the future operation.
25.3.3.4 Potable Water System
The potable water for the main office, the dry, the mechanical shop and the
underground lunch room comes from an existing 8’’ diameter well. The existing
pumping system with pumps, pressure tank and reserve are not reusable and will
have to be replaced by a completely new system. The wood cabin used to
protect the pumping system is reusable but requires minor repairs (doors and
access). The distribution piping to buildings has been damaged in a few places
but most of the piping was not accessible during the visit. The water piping
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distribution has to be replaced. The costs assessment, assumes that the water
from the well is potable, the flow is sufficient and does not require any treatment.
A pumping test and sampling will be required to evaluate the capacity of the well
and the water quality.
25.3.3.5 Waste Water Treatment
The existing installation must be replaced. The proposed system is an advanced
secondary treatment of biologic reactor type. The effluent of this system would be
discharged into the environment.
25.3.3.6 Camp Site
At the camp site, two dormitory trailers of 20 person capacity and one dormitory
trailer of 8 person capacity are still on site. All trailers are in bad condition.
Overhaul of these trailers is possible but not recommended.
The capacity of the dormitory trailers will have to be for 75 persons (average) and
100 persons at the peak of operation (second year only). The suggested
arrangement for the dormitory trailers is as follows:
Assembly of modular trailers with 6,900 ft 2 of service area :
o Two (2) modular trailers with 36 rooms, simple occupancy (72
persons);
o One (1) modular trailer with 28 rooms, simple occupancy for
workers (28 persons). This complex will be necessary in the
second year of operation for the peak of 100 persons at the mine
site;
o One (1) modular trailer with sanitary services of 100 person
capacity.
Due to the important capital cost required to buy a new facility or overhaul the
existing one, it is recommended to rent a new dormitory.
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The kitchen and dining trailers are non existent. New modular buildings of 2,120
ft 2 will have to be erected. The proposed kitchen and dining arrangements offer
75 sitting places and it is recommended to go for rental.
There’s currently no recreation trailers. New modular buildings of 1,600 ft 2 will
have to be erected or rented. The proposed recreation trailers will include a living
room with a television, a pool and ping-pong tables.
Another possibility is to restore one of the existing dormitory trailers and
transform it into a recreation room. Due to the important capital cost required to
buy a new recreation trailer or overhaul the existing dormitory, it is recommended
to rent a new recreation trailer.
The potable water for the camp comes from an existing 8’’ diameter well. The
existing pumping system with pumps, pressure tank and reserve are not reusable
and will have to be replaced as well as the distribution piping. We assume that
the water from the well is potable and does not require any treatment. A pumping
test and sampling will be required to evaluate the capacity of the well and the
water quality.
For the waste water treatment, the existing system is not planned to be used
again as the costs of upgrading it would be too high. It is then proposed to
replace it with a new system.
25.3.4 Electrical Infrastructure
25.3.4.1 Main Power Plant
All the power supply equipment actually available on site is in poor to bad
condition. Generator sets are obsolete and the cost of repair, operation and
maintenance would be much greater than the cost of buying and operating new
sets. It’s then recommended to replace the current ones with new equipment.
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The original generator configuration of "n+1" and the available power of two
times 900 kW prime will not be enough for all of the site requirement. The new
system shall have three 900 kW prime models plus one spare 900 kW prime.
Power cabling has been ripped off. New power cabling will have to be installed.
We don’t know in what conditions the actual generator power distribution
breakers are. The internal mechanisms have to be tested in a specialized shop
to ensure good working conditions. However, Federal Pioneer electronic relays
of that age tend to fail to operate, especially after being stored the way the
current ones are.
An entirely new local distribution system will also be needed (cabling and main
breaker) for each generator set.
Even if the generator sets’ shelters could be repaired at a fair price, the field fit
cost of installing the new generators, makes the option of buying a new shelter
more economical. The new shelters also provide for daily fuel tanks installation.
Control equipments for generator synchronism and load sharing are totally
obsolete. New ones will have to be included with the new generators.
The option of supplying the mine site through Hydro-Quebec power grid, with an
overhead power line, was also examined. Unfortunately, after studying the
Estrades project geographical location and estimating power demand, Hydro-
Quebec representatives stated that no economical power source was available to
supply the site this far. The costs of infrastructure estimated by Hydro-Quebec to
effectively bring power to the site, i.e. building a new sub-station, would be far
greater than the savings obtained on kilowatt-hour rates.
25.3.4.2 Main Distribution Switch Gear
The current switch gear equipment has been totally ripped apart by copper
thieves and can no longer be used. A complete 600 V main distribution will have
to be purchased.
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25.3.4.3 Medium Voltage Distribution
The site medium voltage distribution system consists of a set of two 1,500 kVA
oil type step-up transformers. These transformers are used to raise the voltage
from 600 V to 4,160 V. The transformers seem to be in relatively good condition,
though they will need to be electrically tested. A minor oil leakage indicates that
one of the transformers may need re-gasketing. Finally, oil tests shall be done to
confirm if the transformers are ready to be operated. New transformers are not
necessary unless pre-operational tests show otherwise.
Both transformers are actually tied to medium voltage fused load break switches.
The switches are in fair condition and, with shop repairs, could be used again
The two main medium voltage transformers feed all remote installations through
an overhead line. The line will supply the main ventilation, the underground
distribution, the pond pumping stations, the lime plant, the guard house and the
camp installations.
.
The actual overhead line is in really poor condition. Almost all wood poles are
rotten and some are already broken. New poles and horizontal wooden struts
will have to be installed to repair the overhead line. All guys and anchors will
also have to be added or replaced.
The power cable and power cable support equipment will be reusable to wire the
line between the camp installation and the underground power take-off.
However, for the first part of the overhead line, between the step-up transformers
and the underground take-off, bigger 477 kcmil ACSR cables will be required.
A transformer is actually installed on the ground and a cable is buried up to the
old camp site. This transformer installation is not environmentally compliant and
the actual cable is too short (burned with kitchen modules).
In order to supply the camp, the overhead line will be extended to the camp site
and pole mounted transformers will be used to feed all trailers as needed.
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25.3.4.4 Main Mine Ventilation System
Nothing is currently available for the main ventilation power supply. A new
system must be purchased and installed. In order to supply the ventilation
system, a derivation will be taken from the overhead line, through one of the two
refurbished load-break fuse switches, which will be used to supply a new 1,750
kVA step-down transformer. This will be used to supply the main fan VFD
(variable frequency drive) and building through a 2,000 A power distribution
panel.
Four 250 HP fans are required for the main mine ventilation system. To optimize
overall power consumption and lower start-up current on the power house,
variable speed drives will have to be installed to control the fan motors.
The air heating system will use a local specialized controller for all propane gas
and burner operation.
25.3.4.5 Underground Power Distribution
The existing underground distribution system could not be evaluated due to the
mine being currently flooded. The power medium voltage cable will eventually be
reusable once dried. This study considers that most of the medium voltage cable
already installed underground can be reused.
To feed the various mine levels, a total of eight 500 kVA new portable electrical
sub-stations are planned. All sub-stations will have the same feeder
configuration as well as the flexibility to feed the constantly moving load. The
load shall be moving from one sub-station to the other depending on mine
development.
25.3.4.6 Compressed Air System
Existing compressor starters are obsolete and will be replaced with electronic
soft starters. Power cabling will have to be redone since it has been ripped off.
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25.3.4.7 Mine and Process Water Distribution and Treatment
For both the fresh and waste water systems, new pump stations will be installed
and equipped with electrical distribution. A pole transformer will be installed on
the overhead line to supply each station.
New lime and acid plants will be installed and equipped with electrical
distribution. A pole transformer will be installed on the overhead line to supply
them.
25.3.4.8 Mechanical Shop and Cold Storage
The main distribution trough will be reusable with all its switches and starters.
However, a good overhaul shall be done on all of the distribution equipment prior
to put load on them. The mechanical shop’s main trough will be supplied from
the main 600 V switch gear.
25.3.4.9 Service Buildings
The service buildings, when delivered, will have their own internal distribution
system already in place.
All service buildings near the main 600 V distribution, will be supplied directly
through buried cable. The following buildings will be wired that way:
Dry building;
Office building;
Potable water station;
Mechanical shop.
25.3.4.10 Camp Site
The power to all installations will be provided by pole transformer. Every building
or building group, depending on the voltage and power needs, will be supplied by
its own transformer.
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25.3.5 Site Preparation and Access Road
Site preparation will require grading at both mine and camp sites, surface
preparation for diesel and propane storage, dismantling of old reservoirs and
clean-up, particularly at the mine site. For the access road which is 36 km long,
minimal resurfacing will be done with natural gravel, only 5 culverts will be
replaced and some deforestation will be performed over about 20 km on each
side of the road.
For the costs evaluation purposes, it is assumed that the same contractor will
take care of both the site and the road.
25.3.6 Ore Stockpile Concrete Pad
A concrete pad of 25 m x 25 m will be built to stockpile ore hauled from the
underground operation. It will be designed in order to allow for water drainage.
25.3.7 Waste Pad
A waste pad for 200,000 tonnes of material will be put in place. It will be
surrounded by a drainage ditch directing water into a settling pond.
25.3.8 Capital Costs
The table below shows the capital costs which will be incurred at the outset of the
project for renting, purchasing, installing or overhauling the various
infrastructures required to start mining operations. They all include a 10%
contingency.
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Table 25-3-1 Capital Costs for the Site Infrastructure to Start Operations
Infrastructure Costs, $
Mobilization Costs for Rentals
Main Office 54,000
Dry 58,300
Camp 552,300
Sub-Total 664,600
Purchases, Repairs and Installation
Main U/G Ventilation System 665,600
Generators 1,543,300
Mechanical Infrastructure 736,200
Buildings 346,500
Electrical Infrastructure 1,976,700
Access Road and Site Preparation 300,000
Ore Stockpile Concrete Pad 194,700
Waste Pad 110,000
Sub-Total 5,873,000
Total 6,537,600
EPCM @ 8% of total 523,008
Total with EPCM 7,060,608
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25.4 Manpower
The planned work schedule is of 5 days a week at 10 hours per day. In the
current study it is assumed that all of the workers are Cogitore Resources’s
employees. However, it is more than likely that mining (development and
production) will be contracted out. Our costs evaluation does apply to both
situations. It reflects the labour shortage that prevails on the market at the
moment and consequently, the high wages offered.
The following table illustrates the total number of employees required to run the
site from the time dewatering and rehabilitation are completed.
Table 25-4-1 Total Number of Employees Required on Site
Months 7 - 18 19 - 30 31 - 42
Development 25 19 0
Stoping 7 37 12
Mucking 8 8 8
Staff 22 22 22
Services 11 11 11
Total 73 97 53
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25.5 Environmental Considerations
25.5.1 Physical Environmental Aspects
The mine area is included in the hydrographic network of the Harricana river.
Locally, this network is mainly constituted of small creeks flowing in the north
direction. Estrades mine site covers an area of approximately 139,000 m² where
the main infrastructures are surrounded by a ditch allowing segregation of
runoffs. Clayey soils, having a thickness between 10 and 45 meters, cover the
property. The only major environmental structures remaining on site are the
ponds that are constructed with this clay material; after clean up, ponds will be
functional. Ore and waste rocks are partly potentially acid generating and metal
leachable; the planned waste management approach and infrastructure are
designed to handle these risks safely and according to Quebec’s province
environmental laws and regulations.
25.5.2 Permitting
The project is certified being in compliance with the James Bay municipality
regulations. According to a map provided in 2007 by the Abitibiwinni First Nation
Council, the mine area is not in a zone where traditional hunting and fishing
activities are taking place. Quebec’s Ministry of Environment issued in 2007 a
certificate of authorization for underground mine de-flooding and extraction of
approximately 422,000 tonnes of ore at a rate of 500 tonnes per day. Requests
for other environmental permits (not limited to but such as: wastewater, potable
water, solid waste disposal area, and closure plan) is not done already.
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25.6 Markets and Contracts
All of the concentrates produced from the Estrades project will be sold to
smelters at terms and conditions prevailing in the industry.
At the moment, Cogitore Resources has just obtained quotations for the purpose
of this feasibility. No contracts have been signed yet with any smelting operation.
In a further stage, Cogitore Resources will look for the best terms available on
the market at that time.
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25.7 Financial Analysis
A pre-tax financial model has been developed for the Estrades Project. All costs
are in current Canadian dollars with no allowance for inflation or escalation.
The economic valuation of the project has been conducted with the Net Present
Value method. The NPV method converts all cash flows of investments and
revenues occurring throughout the planning horizon of a project to an equivalent
single sum at present time using a discount rate. The discount rate used in the
analysis is 10%.
25.7.1 Assumptions
Production
For the current feasibility study, after about 6 months of underground dewatering
and rehabilitation, it is assumed that development will start and production will
follow. Development in mineralization will begin approximately 12 months after
the beginning of the project and stoping will start producing ore approximately in
the 14 th month.
Ore production will gradually ramp up and the production phase will be reached
in the 21 th month of the project with more than 77% of the planned maximum rate
of 20,000 tonnes of ore per month. The planned maximum rate will be achieved
a month later and maintained until nearly the end of operations towards the end
of the third year.
Revenues
For the purpose of revenue calculations, metallurgical recoveries achieved in the
laboratory with the locked cycle test LCT-2 were used. However, because of
differences in grades between the head sample used for LCT-2 and the average
grades of the mineral reserve, the quantities of concentrates produced were
adjusted in proportion to the grades of the reserve, assuming that recoveries
would stay the same.
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The following metal prices used for the base case are shown below as well as
the exchange rate. Prices are in US$.
Zinc 0.90
Copper 3.25
Lead 0.85
Gold 850.00
Silver 15.00
Exchange Rate 1.05
Quotations for market treatment charges were obtained from smelters based in
Canada for the three concentrates. However, no smelting and refining
agreements have been reached yet and this will take place when Cogitore
Resources will decide to go ahead with the project.
Operating and Capital Costs
Most of the costs were estimated with a 15% contingency included, except for the
mechanical and electrical infrastructures where a 10% contingency was applied
to the quotations obtained from suppliers for purchases and rentals as well as to
the costs of overhauling.
The custom milling costs were estimated for the purpose of this study, based on
the costs incurred by similar operations, but no agreement has been reached
with the owner of the concentrator. This has to be negotiated when Cogitore
Resources is ready to go ahead.
The costs of ore hauling from the mine site to the concentrator, are estimates
obtained from a transport company of the Abitibi area. For the concentrates
freight rates, these are estimates provided by CN Railways.
A total of about $51 M will be spent in capital expenditures until the 21 th month of
the project is reached when more than 77% of the planned maximum rate of
20,000 tonnes of ore per month will be reached and this will mark the outset of
the production phase. After dewatering and rehabilitation, development and
production will be spread over a total of nearly 3 years.
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25.7.2 Cash Flow Analysis
The pre-tax cash flow is presented in Table 25-7-1 for the best of the two milling
options examined.
The net project cash flow before taxes amounts to about $5.4 M with an internal
rate of return of 7.4% and a negative net present value of $1.4 M at a discount
rate of 10%.
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Table 25-7-1 Pre-Tax Cashflow of the Estrades Project
Dew-
Rehab
Pre-Production Production Cl osure Total
Years Year 1 Year 2 Year 3 Year 4 Year 5
Months 0-6 7-18 19-20 21-30 31-42 43-48
PRODUCTION (‘000 DMT)
Ore Milled 43,700 22,750 195,550 189,900 451,900
Concentrates:
Copper 1,121 623 5,358 5,370 12,472
Zinc 6,405 3,364 28,911 32,218 70,898
Lead 632 346 2,970 2,381 6,329
REVENUES (‘000$)
NSR ** 12,089 5,789 49,757 48,432 116,067
COSTS (‘000$)
Dewatering/ Rehab. 3,565 3,565
U/G Vent. Raises 1,725 1,725
Infrastructure 6,538 3,404 791 3,955 2,965 (912) 16,741
E.P.C.M. 523 523
Mining /Ore Hauling
Development 12,154 1,569 3,173 16,896
Stoping - Mucking 1,678 1,249 10,219 4,589 17,735
Ore Hauling 2,010 1,047 8,995 8,735 20,787
Total 15,842 3,865 22,387 13,324 55,418
Ore Treatment
Modifications 6,480 6,480
Milling, Operations 656 341 2,933 2,849 6,779
Gen. Admin.
Services-Labour 3,557 593 2,964 3,254 10,368
Misc. Equipment 2,444 407 2,036 2,200 7,087
Water Treatment 350 60 290 350 175 1,225
Closure 750 750
TOTAL COSTS 17,106 27,978 6,057 34,565 24,942 13 110,661
CASH FLOW (‘000$)
CASH FLOW (17,106) (15,889) (268) 15,192 23,490 (13) 5,406
Cash Flow, cum. (17,106) (32,995) (33,263) (18,071) 5,419 5,406
R.O.R. (%) 7.4
NPV (‘000$) @ 10% (1,434)
**NSR: Net Smelter Return
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25.7.2 Sensitivity Analysis
The following parameters were used to produce the sensitivity analysis upon the
cash flow and are presented in decreasing order of importance:
1- Exchange Rate
2- Zinc Price
3- Gold Price
4- Zn Recovery in the Zinc Concentrate
5- Zinc Concentrate Treatment Charge
6- Silver Price
7- Copper Price
8- Lead Price
Both the exchange rate and the zinc price have a tremendous impact upon the
project economics. A ±10 % variation of the exchange rate has a ±$12 M impact
upon the cash flow and a similar variation of the zinc price results in a ±$6 M.
The results are presented in Table 25-7-2 and in Figure 25-7-1.
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Table 25-7-2 Sensitivity Analysis for the Estrades Project
Cash Flow Resulting from the Variation of Parameters
Price of Metals -50% -40% -30% -20% -10% 0% 10% 20% 30% 40% 50%
Zn -25,252 -19,121 -1,989 -6,858 -725 5,406 11,538 17,669 23,539 29,081 34,624
Cu -3,895 -2,058 -164 1,674 3,568 5,406 7,300 9,138 11,033 12,870 14,765
Pb 2,950 3,418 3,944 4,412 4,938 5,406 5,932 6,400 6,926 7,394 7,920
Ag -2,325 -780 767 2,313 3,860 5,406 6,952 8,498 10,045 11,591 13,138
Au -15,405 -11,242 -7,080 -2,918 1,244 5,406 9,568 13,730 17,892 22,054 26,217
Exch Rate -54,245 -42,774 -30,155 -18,684 -6,065 5,406 18,024 29,495 42,115 53,585 66,205
TC-Zn 16,572 14,339 12,105 9,872 7,639 5,406 3,172 940 -1,294 -3,527 -5,760
Zn Recovery 1,300 5,406 9,466
Variation of Parameters
Price of Metals, US$
Zn 0.45 0.54 0.63 0.72 0.81 0.90 0.99 1.08 1.17 1.26 1.35
Cu 1.63 1.95 2.28 2.60 2.93 3.25 3.58 3.90 4.23 4.55 4.88
Pb 0.43 0.51 0.60 0.68 0.77 0.85 0.94 1.02 1.11 1.19 1.28
Ag 7.50 9.00 10.50 12.00 13.50 15.00 16.50 18.00 19.50 21.00 22.50
Au 425.00 510.00 595.00 680.00 765.00 850.00 935.00 1,020.00 1,105.00 1,190.00 1,275.00
Exch Rate 0.53 0.63 0.74 0.84 0.95 1.05 1.16 1.26 1.37 1.47 1.58
TC-Zn 150.00 180.00 210.00 240.00 270.00 300.00 330.00 360.00 390.00 420.00 450.00
Zn Recovery 80.80 89.80 98.70
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Figure 25-7-1 Cashflow Sensitivity
Cashflow Sensitivity
80 000
60 000
40 000
Cashflow (,000$)
20 000
0
-20 000
-40 000
-50% -40% -30% -20% -10% 0% 10% 20% 30% 40% 50%
-60 000
-80 000
Variation of Parameters, %
Zn Cu Pb Ag Au Exch Rate TC-Zn Zn Recovery
GENIVAR 135
AV105365

APPENDIX 1

DETAILED CALCULATIONS BY GENIVAR FOR LOCKED CYCLE TEST LCT-2
Average of the last 2 cycles
Weight Assays, %, g/t Units % Distribution
g % Cu Pb Zn Au Ag Cu Pb Zn Au Ag Cu Pb Zn Au Ag
Cu Conc 35.40 1.74 20.65 4.75 13.75 82.65 2 522.00 36.0 8.3 24.0 144.0 4 393.1 60.2 6.7 1.9 32.6 21.7
Pb Conc 37.00 1.82 2.15 42.10 15.60 74.85 2 227.00 3.9 76.6 28.4 136.3 4 054.6 6.6 62.1 2.2 30.9 20.1
Zn Conc 421.45 20.74 0.46 0.94 56.00 3.19 215.00 9.5 19.4 1 161.3 66.2 4 458.7 16.0 15.7 89.8 15.0 22.1
Zn Cl T 149.90 7.38 0.66 0.82 6.79 3.90 318.00 4.9 6.0 50.1 28.8 2 345.6 8.1 4.9 3.9 6.5 11.6
Tails 1 388.50 68.32 0.08 0.19 0.43 0.97 72.70 5.5 13.0 29.4 66.3 4 967.1 9.1 10.5 2.3 15.0 24.6
Heads 2 032.25 100.00 0.60 1.23 12.93 4.41 202.19 59.8 123.3 1 293.1 441.4 20 219.1 100.0 100.0 100.0 100.0 100.0
Head calculated 0.67 1.32 12.7 4.76 209.5
Av weight/cycle: 1 984.8 g/cycle

APPENDIX 2

CIM DEFINITION STANDARDS - For Mineral Resource and Mineral Reserve
Mineral Resource
Mineral Resources are sub-divided, in order of increasing geological confidence, into
Inferred, Indicated and Measured categories. An Inferred Mineral Resource has a lower
level of confidence than that applied to an Indicated Mineral Resource. An Indicated
Mineral Resource has a higher level of confidence than an Inferred Mineral Resource
but has a lower level of confidence than a Measured Mineral Resource.
A Mineral Resource is a concentration or occurrence of diamonds, natural solid
inorganic material, or natural solid fossilized organic material including base and
precious metals, coal, and industrial minerals in or on the Earth’s crust in such
form and quantity and of such a grade or quality that it has reasonable prospects
for economic extraction. The location, quantity, grade, geological characteristics
and continuity of a Mineral Resource are known, estimated or interpreted from
specific geological evidence and knowledge.
The term Mineral Resource covers mineralization and natural material of intrinsic
economic interest which has been identified and estimated through exploration and
sampling and within which Mineral Reserves may subsequently be defined by the
consideration and application of technical, economic, legal, environmental, socioeconomic
and governmental factors. The phrase ‘reasonable prospects for economic
extraction’ implies a judgement by the Qualified Person in respect of the technical and
economic factors likely to influence the prospect of economic extraction. A Mineral
Resource is an inventory of mineralization that under realistically assumed and justifiable
technical and economic conditions might become economically extractable. These
assumptions must be presented explicitly in both public and technical reports.
Inferred Mineral Resource
An ‘Inferred Mineral Resource’ is that part of a Mineral Resource for which
quantity and grade or quality can be estimated on the basis of geological evidence
and limited sampling and reasonably assumed, but not verified, geological and
grade continuity. The estimate is based on limited information and sampling
gathered through appropriate techniques from locations such as outcrops,
trenches, pits, workings and drill holes.
Due to the uncertainty that may be attached to Inferred Mineral Resources, it cannot be
assumed that all or any part of an Inferred Mineral Resource will be upgraded to an
Indicated or Measured Mineral Resource as a result of continued exploration.
Confidence in the estimate is insufficient to allow the meaningful application of technical
and economic parameters or to enable an evaluation of economic viability worthy of
public disclosure. Inferred Mineral Resources must be excluded from estimates forming
the basis of feasibility or other economic studies.
Indicated Mineral Resource
An ‘Indicated Mineral Resource’ is that part of a Mineral Resource for which
quantity, grade or quality, densities, shape and physical characteristics, can be
estimated with a level of confidence sufficient to allow the appropriate application

of technical and economic parameters, to support mine planning and evaluation
of the economic viability of the deposit. The estimate is based on detailed and
reliable exploration and testing information gathered through appropriate
techniques from locations such as outcrops, trenches, pits, workings and drill
holes that are spaced closely enough for geological and grade continuity to be
reasonably assumed.
Mineralization may be classified as an Indicated Mineral Resource by the Qualified
Person when the nature, quality, quantity and distribution of data are such as to allow
confident interpretation of the geological framework and to reasonably assume the
continuity of mineralization. The Qualified Person must recognize the importance of the
Indicated Mineral Resource category to the advancement of the feasibility of the project.
An Indicated Mineral Resource estimate is of sufficient quality to support a Preliminary
Feasibility Study which can serve as the basis for major development decisions.
Measured Mineral Resource
A ‘Measured Mineral Resource’ is that part of a Mineral Resource for which
quantity, grade or quality, densities, shape, and physical characteristics are so
well established that they can be estimated with confidence sufficient to allow the
appropriate application of technical and economic parameters, to support
production planning and evaluation of the economic viability of the deposit. The
estimate is based on detailed and reliable exploration, sampling and testing
information gathered through appropriate techniques from locations such as
outcrops, trenches, pits, workings and drill holes that are spaced closely enough
to confirm both geological and grade continuity.
Mineralization or other natural material of economic interest may be classified as a
Measured Mineral Resource by the Qualified Person when the nature, quality, quantity
and distribution of data are such that the tonnage and grade of the mineralization can be
estimated to within close limits and that variation from the estimate would not
significantly affect potential economic viability. This category requires a high level of
confidence in, and understanding of, the geology and controls of the mineral deposit.
Mineral Reserve
Mineral Reserves are sub-divided in order of increasing confidence into Probable
Mineral Reserves and Proven Mineral Reserves. A Probable Mineral Reserve has a
lower level of confidence than a Proven Mineral Reserve.
A Mineral Reserve is the economically mineable part of a Measured or Indicated
Mineral Resource demonstrated by at least a Preliminary Feasibility Study. This
Study must include adequate information on mining, processing, metallurgical,
economic and other relevant factors that demonstrate, at the time of reporting,
that economic extraction can be justified. A Mineral Reserve includes diluting
materials and allowances for losses that may occur when the material is mined.
Mineral Reserves are those parts of Mineral Resources which, after the application of all
mining factors, result in an estimated tonnage and grade which, in the opinion of the
Qualified Person(s) making the estimates, is the basis of an economically viable project
after taking account of all relevant processing, metallurgical, economic, marketing, legal,

environment, socio-economic and government factors. Mineral Reserves are inclusive of
diluting material that will be mined in conjunction with the Mineral Reserves and
delivered to the treatment plant or equivalent facility. The term ‘Mineral Reserve’ need
not necessarily signify that extraction facilities are in place or operative or that all
governmental approvals have been received. It does signify that there are reasonable
expectations of such approvals.
Probable Mineral Reserve
A ‘Probable Mineral Reserve’ is the economically mineable part of an Indicated
and, in some circumstances, a Measured Mineral Resource demonstrated by at
least a Preliminary Feasibility Study. This Study must include adequate
information on mining, processing, metallurgical, economic, and other relevant
factors that demonstrate, at the time of reporting, that economic extraction can be
justified.
Proven Mineral Reserve
A ‘Proven Mineral Reserve’ is the economically mineable part of a Measured
Mineral Resource demonstrated by at least a Preliminary Feasibility Study. This
Study must include adequate information on mining, processing, metallurgical,
economic, and other relevant factors that demonstrate, at the time of reporting,
that economic extraction is justified.
Application of the Proven Mineral Reserve category implies that the Qualified Person
has the highest degree of confidence in the estimate with the consequent expectation in
the minds of the readers of the report. The term should be restricted to that part of the
deposit where production planning is taking place and for which any variation in the
estimate would not significantly affect potential economic viability.

APPENDIX 3

ABBREVIATIONS
Unless otherwise noted, all units of measurement used in this report are expressed
according to the metric system. The following conversion factors and their respective
abbreviations are used in this report:
1 ounce troy (oz) = 31.1035 grams (g)
1 ton: 2,000 pounds
1 tonne (t) = 1 metric tonne
1 metre (m) = 3.28 feet (ft)
Other abbreviations used in this report are as follows:
Au: gold
Ag: silver
bhp: break horsepower
cfm: cubic feet per minute
cm: centimetre
Cu: copper
ddh: diamond drill hole
ft: foot
ft 2 : square foot
g: gram
g/t: grams per tonne
ha: hectare
HP: horsepower
kcmil: 1,000’s of circular mil
kg: kilogramme
km: kilometre
kV: kilovolt
kVA: kilovolt-ampere
kW: kilowatt
m: metre
m 2 : square metre
M: million
Ma: million years

Mil: 1/1,000 inch
mm: millimetre
MW : megawatt
NSR: net smelter return
oz: ounce
Pb: lead
PSI: pounds per square inch
P 80 : dimension for which 80% of the particles of a material are smaller
T: tonne
t/d: tonnes/day
V: volt
VG: visible gold
w.g: water gauge
Zn: zinc
µm: micrometer