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FORM 6-K
SECURITIES AND EXCHANGE COMMISSION
Washington, D.C. 20549
Report of Foreign Private Issuer
Pursuant to Rule 13a-16 or 15d-16
of the Securities Exchange Act of 1934
For the month of June, 2003
Commission File Number 000 - 13979
Breakwater Resources Ltd.
(Translation of registrant's name into English)
95 Wellington Street West, Suite 2000, Toronto, Ontario M5J 2N7
(Address of principal executive offices)
Indicate by check mark whether the registrant files or will file annual reports
under cover Form 20-F or Form 40-F.
Form 20-F X Form 40-F.....
Indicate by check mark if the registrant is submitting the Form 6-K in paper as
permitted by Regulation S-T Rule 101(b)(1):
Indicate by check mark if the registrant is submitting the Form 6-K in paper as
permitted by Regulation S-T Rule 101(b)(7):
Indicate by check mark whether the registrant by furnishing the information
contained in this Form is also thereby furnishing the information to the
Commission pursuant to Rule 12g3-2(b) under the Securities Exchange Act of 1934.
Yes ..... No X
If "Yes" is marked, indicate below the file number assigned to the registrant in
connection with Rule 12g3-2(b): 82- ________
SIGNATURES
Pursuant to the requirements of the Securities Exchange Act of 1934, the
registrant has duly caused this report to be signed on its behalf by the
undersigned, thereunto duly authorized.
BREAKWATER RESOURCES LTD.
(Registrant)
By: (signed) E. Ann Wilkinson
---------------------------------
E. Ann Wilkinson
Corporate Secretary
Date: June 12, 2003
INDEX TO EXHIBITS
The following documents are being filed with the Commission as exhibits to,
and are incorporated by reference into and form part of, the report on Form 6-K.
Number
- ------
1. Press Release Dated June 12, 2003
2. Langlois Mine Feasibility Study prepared by SRK Consulting - August 2001
Revised June 2003
EXHIBIT 1
[LOGO] b
- --------------------------------------------------------------------------------
BREAKWATER RESOURCES LTD.
95 WELLINGTON STREET WEST, SUITE 2000
TORONTO, ONT., M5J 2N7
Tel: (416) 363-4798
Fax: (416) 363-1315
- --------------------------------------------------------------------------------
MEDIA RELEASE
BREAKWATER ANNOUNCES RESULTS OF
UPDATED LANGLOIS FEASIBILITY STUDY
JUNE 12, 2003... (BWR - TSX)
Breakwater Resources Ltd. announces the results of the updated SRK Consulting
(SRK) feasibility study for the Langlois Mine, located in northwestern Quebec.
The 2003 feasibility study includes the new mineral reserves in Zone 97 which
were proven up following a 28 hole, 11,511 metre in-fill drill program which was
completed in April 2003.
On May 22, 2003, Breakwater announced a 25 percent increase in mineral reserves
in Zone 97, which adds a further year's production to the expected life of the
mine. In total, the mineral reserves for Zone 97 have increased by 419,600
tonnes, at a grade of 8.1 percent zinc, 1.7 percent copper, 46.9 grams of silver
per tonne and 0.1 grams of gold per tonne. The resources and reserves for Zones
3 and 4 have not changed.
Breakwater purchased the Langlois Mine in May 2000 and operated the mine until
November 2000. Operations were temporarily suspended when problems associated
with the main ore pass system combined with low metal prices made it uneconomic
to operate. The Langlois Mine was placed on care-and-maintenance in 2001 pending
the resolution of the ore pass problem and an improvement in the price of zinc.
A feasibility study to reopen the Langlois Mine including a technical resolution
for the ore pass problem was conducted by SRK in August 2001, following an
extensive drilling program. The original SRK feasibility study indicated a net
pre-tax cash flow of $60.9 million based on a zinc price of US$0.50 per pound, a
copper price of US$0.80 per pound, a silver price of US$5.00 per ounce and an
exchange rate of US$0.66 per Can$. The internal rate of return was 24.0 percent
and the NPV at 8.0 percent was $26.4 million.
The 2003 SRK feasibility study now indicates that the total net pre-tax cash
flow is estimated to be $71.1 million, based on the metal prices shown in Table
I. The internal rate of return has increased to 25.3 percent and the NPV at 8.0
percent has increased to $30.9 million. Due to the recent decrease in the
US/Canadian dollar exchange rate, an
exchange rate of US$0.70 per Can$ has been used. As well, due to the tight zinc
concentrate market, zinc smelter treatment charges have been reduced. Based on
these price assumptions, the operating cost per pound of payable zinc including
smelting, shipping and by-product credits for copper and precious metals is
US$0.38. The 2003 SRK feasibility study can be accessed at www.sedar.com by
advancing to Breakwater's public documents.
Table I - Metal Prices and Exchange Rate
----------------------------------------
Zinc US$0.50/lb
Copper US$0.80/lb
Silver US$5.00/oz
Gold US$343/oz
Exchange Rate US$0.70/Cdn$
----------------------------------------
An estimated $38.2 million in capital is required over the life of the mine of
which approximately $16.4 million must be expended prior to the start of
production. The majority of the capital requirements are related to the
underground mine. The total operating cost to mine gate is estimated at
Cdn$55.61 per tonne milled over the life of the mine. This cost is in the range
of the historical 2000 operating cost.
The project economics are sensitive to the price of zinc which is currently
depressed and is trading at a 16 year low of US$0.36 per pound. As well, the
project is sensitive to the US/Canadian dollar exchange rate. The 2003 SRK
feasibility study was also run based on a long-term zinc price of US$0.45 per
pound. The total net pre-tax cash flow decreases to $41.5 million, the internal
rate of return decreases to 15.2 percent and the NPV at 8.0 percent decreases to
$12.6 million. Based on this price assumption, the operating cost per pound of
payable zinc including smelting, shipping and by-product credits for copper and
precious metals is estimated to be US$0.36. The Company is investigating several
strategies, which would improve the economics of the mine, including
rationalizing concentrate freight rates. A decision to reopen the Langlois mine
awaits an improvement in the price of zinc and financing.
The 2003 SRK feasibility mine plan provides for milling 3,322,760 tonnes over a
period of eight years based on mineral reserves as of May 1, 2003 shown in Table
II.
Table II - Summary of Mineral Resources and Mineral Reserves for the Langlois Mine
- ----------------------------------------------------------------------------------
Tonnes Zn Cu Au Ag
(000s) (%) (%) (g/t) (g/t)
- ----------------------------------------------------------------------------------
Proven and Probable Reserves 3,323 10.8 0.8 0.1 52
Measured and Indicated Resources (1) 4,981 11.2 0.8 0.1 54
Inferred Resources 1,254 9.7 0.5 0.1 40
- ----------------------------------------------------------------------------------
(1) Measured and Indicated Resources include Proven and Probable Reserves
The 2003 feasibility operating plan incorporates several improvements to ensure
reliability of production and to control costs. These improvements include the
elimination of ore passes for Zone 97, a steel lined storage bin,
pre-development of several sublevels in Zone 97 so that it can produce higher
grade ore continuously when mine operations resume and improvements for the
underground mobile equipment fleet.
Table III summarizes the estimated life of mine production statistics.
Table III - Life of Mine Production Statistics
- ----------------------------------------------------------------------
TONNES MILLED 3,323,000
- ----------------------------------------------------------------------
HEAD GRADE Zn (%) 10.8
Cu (%) 0.8
Au (g/t) 0.1
Ag (g/t) 52.1
RECOVERIES Zn (%) 93.7
Cu (%) 82.3
Au (%) 29.0
Ag (%) 35.8
CONCENTRATE GRADE Zn (%) 54.8
Cu (%) 23.9
TONNES CONCENTRATE Zn (t) 612,000
Cu (t) 94,000
CONTAINED METAL Zn (t) 335,000
Cu (t) 22,700
Au (oz) 2,600
Ag (oz) 1,991,000
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This press release contains certain forward-looking statements that involve a
number of risks and uncertainties. There can be no assurance that such
statements will prove to be accurate; actual results and future events could
differ materially from those anticipated in such statements. Risk and
uncertainties are disclosed under the heading "Risk Factors" in the
Corporation's Annual Report on Form 20-F filed with certain Canadian securities
regulators and with the United States Securities and Exchange Commission.
Breakwater Resources Ltd. is a mineral resource company engaged in the
acquisition, exploration, development and mining of base metal and precious
metal deposits in the Americas and North Africa. Breakwater has four producing
zinc mines: the Bouchard-Hebert mine in Quebec, Canada; the Bougrine mine in
Tunisia; the El Mochito mine in Honduras; and the El Toqui mine in Chile.
For further information please contact:
- ----------------------------- ----------------------------------
COLIN K. BENNER TORBEN JENSEN
- ----------------------------- ----------------------------------
President and Manager of Engineering and
- ----------------------------- ----------------------------------
Chief Executive Officer North American Exploration
- ----------------------------- ----------------------------------
(416) 363-4798 Ext. 269 (416) 363-4798 Ext. 232
- ----------------------------- ----------------------------------
Email: investorinfo@breakwater.ca Website: www.breakwater.ca
LANGLOIS MINE
LEBEL-SUR-QUEVILLON, QUEBEC
2CB003.02
FEASIBILITY STUDY
PREPARED FOR:
BREAKWATER RESOURCES LTD.
Suite 2000, 95 Wellington Street West
Toronto, Ontario
M5J 2N7
AUGUST 2001 - REVISED JUNE 2003
[LOGO] SRK CONSULTING
Engineers and Scientists
Authors: _______________________ _________________________
SRK Consulting Breakwater Resources Ltd.
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
TABLE OF CONTENTS
EXECUTIVE SUMMARY .........................................................................1
1.0 INTRODUCTION .............................................................7
1.1 TERMS OF REFERENCE .......................................................7
1.2 BACKGROUND ...............................................................8
1.3 FEASIBILITY STUDY OBJECTIVES AND APPROACH ...............................12
2.0 GEOLOGY .................................................................13
2.1 REGIONAL AND PROPERTY GEOLOGY ...........................................13
2.2 MINERALIZATION ..........................................................15
2.3 MINERAL RESOURCES AND RESERVES ..........................................19
2.4 EXPLORATION POTENTIAL ...................................................27
3.0 HISTORICAL OPERATING CHALLENGES .........................................28
3.1 ORE PASSES ..............................................................28
3.2 MINE CAPITAL DEVELOPMENT ................................................29
3.3 MOBILE EQUIPMENT MAINTENANCE ............................................29
3.4 ORE STORAGE .............................................................29
3.5 STOPE DILUTION AND SUBLEVEL REHABILITATION ..............................29
3.6 SRK MINE REVIEW .......................................................30
4.0 FEASIBILITY STUDY - MINING ..............................................31
4.1 MINING ROCK MECHANICS ...................................................31
4.2 MINING METHODS ..........................................................37
4.3 STOPE SEQUENCING ........................................................50
4.4 PRODUCTION RATE .........................................................53
4.5 MOBILE EQUIPMENT ........................................................56
4.6 PLANNED PRODUCTIVITIES ..................................................58
4.7 VENTILATION .............................................................59
4.8 ORE AND WASTE HANDLING SYSTEMS ..........................................61
5.0 MINERAL PROCESSING ......................................................68
5.1 PROCESSING ..............................................................68
5.2 METALLURGICAL RESULTS ...................................................73
5.3 OPERATING PLAN ..........................................................74
5.4 FUTURE METALLURGICAL PROJECTS ...........................................76
6.0 ENVIRONMENTAL CONSIDERATIONS ............................................77
6.1 LICENCES AND CERTIFICATES OF AUTHORIZATION ..............................77
6.2 WASTE MANAGEMENT ........................................................78
6.3 WATER MANAGEMENT ........................................................78
7.0 MANPOWER ................................................................80
7.1 UNDERGROUND .............................................................81
7.2 PROCESSING ..............................................................83
7.3 ADMINISTRATION AND TECHNICAL SERVICES ...................................83
7.4 ORGANIZATION CHART ......................................................84
7.5 SALARIES, WAGES AND BENEFITS ............................................85
8.0 PROJECT SCHEDULE ........................................................86
8.1 CONSTRUCTIO PERIOD ......................................................86
8.2 LIFE OF MINE PRODUCTION AND DEVELOPMENT .................................88
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
9.0 CONCENTRATE SHIPMENT AND MARKETING ......................................90
9.1 CONCENTRATE SPECIFICATIONS ..............................................90
9.2 CONCENTRATE SHIPMENT ....................................................90
9.3 CONCENTRATE TERMS .......................................................90
10.0 OPERATING COST ..........................................................92
10.1 MINING COST .............................................................93
10.2 MILLING COST ............................................................98
10.3 ADMINISTRATION COST .....................................................99
11.0 CAPITAL COST ...........................................................100
11.1 DEFERRED DEVELOPMENT ...................................................100
11.2 EQUIPMENT ..............................................................102
11.3 SALVAGE VALUE ..........................................................103
11.4 ENVIRONMENTAL AND ABANDONMENT ..........................................104
12.0 ECONOMICS ..............................................................105
12.1 CASHFLOW ...............................................................105
12.2 PROJECT SENSITIVITY ....................................................108
13.0 OPPORTUNITIES AND RISKS ................................................110
13.1 OPPORTUNITIES ..........................................................110
13.2 RISKS ..................................................................113
TABLES
TABLE I - LANGLOIS MINE RESOURCES AND RESERVES ............................................3
TABLE II - FEASIBILITY OPERATING PLAN .....................................................3
TABLE III - OPERATING COST ................................................................5
TABLE IV CAPITAL ..........................................................................5
TABLE V - METAL PRICES AND EXCHANGE RATE ..................................................5
TABLE VI - PRE-TAX CASH FLOW SUMMARY ......................................................6
TABLE 1-1 - HISTORICAL PRODUCTION .........................................................9
TABLE 2-1 - RESOURCES FOR THE LANGLOIS MINE AS OF JANUARY 31, 2001 .......................21
TABLE 2-2 - RESERVES FOR THE LANGLOIS MINE AS OF JANUARY 31, 2001.........................21
TABLE 2-3 - RESOURCES FOR THE LANGLOIS MINE AS OF MAY 1, 2003 ............................26
TABLE 2-4 - RESERVES FOR THE LANGLOIS MINE AS OF MAY 1, 2003 .............................26
TABLE 2-5 - HISTORICAL RECONCILIATION ....................................................27
TABLE 4-1 - ROCK MASS CLASSIFICATION VALUES FOR THE MAIN GEOTECHNICAL DOMAINS ............33
TABLE 4-2 - STABILITY NUMBER VALUES FOR THE MAIN GEOTECHNICAL DOMAINS ....................33
TABLE 4-3 - PASTE BACKFILL LOGISTICS .....................................................49
TABLE 4-4 - PRODUCTION STATISTICS ........................................................55
TABLE 4-5 - TONNES MINED BY ZONE .........................................................56
TABLE 4-6 - EXISTING UNDERGROUND MOBILE EQUIPMENT ........................................57
TABLE 4-7 - NEW UNDERGROUND MOBILE EQUIPMENT PURCHASES ...................................57
TABLE 4-8 - PLANNED MINING PRODUCTIVITIES ................................................58
TABLE 5-1 - ORE COMPOSITION ..............................................................68
TABLE 5-2 - GRINDING CIRCUIT DATA ........................................................69
TABLE 5-3 - SCREEN ANALYSIS WEIGHT DISTRIBUTION SUMMARY ..................................69
TABLE 5-4 - COPPER FLOTATION REAGENTS.....................................................70
TABLE 5-5 - ZINC FLOTATION REAGENTS ......................................................70
TABLE 5-6 - METALLURGY-HISTORICAL PRODUCTION 1997 - 2000 .................................73
TABLE 5-7 - METALLURGY - OPERATING PLAN ..................................................74
TABLE 5-8 - METALLURGY SUMMARY ...........................................................75
TABLE 5-9 - IRON IN SPHALERITE ...........................................................75
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
TABLE 5-10 - METALLURGICAL RESULTS .......................................................75
TABLE 7-1 - TOTAL MINE SITE MANPOWER .....................................................80
TABLE 7-2 - MINE DEPARTMENT PERSONNEL ....................................................81
TABLE 7-3 - CONTRACTOR PERSONNEL .........................................................82
TABLE 7-4 - MAINTENANCE DEPARTMENT PERSONNEL .............................................82
TABLE 7-5 - ELECTRICAL DEPARTMENT PERSONNEL ..............................................83
TABLE 7-6 - MILL & ASSAY LABORATORY PERSONNEL ............................................83
TABLE 7-7 - ADMINISTRATION AND TECHNICAL SERVICES ........................................83
TABLE 7-8 - HOURLY RATE EMPLOYEES WAGE SCALE .............................................85
TABLE 8-1 - PRODUCTION SCHEDULE ..........................................................88
TABLE 8-2 - DEVELOPMENT SCHEDULE .........................................................89
TABLE 9-1 - SHIPPING COST TO SMELTER .....................................................90
TABLE 9-2 - CONCENTRATE SPECIFICATIONS ...................................................91
TABLE 10-1 - OPERATING COST ..............................................................92
TABLE 10-2 - OPERATING COST COMPARISON ...................................................92
TABLE 10-3 - MINING COST .................................................................93
TABLE 10-4 - STOPE PREPARATION ...........................................................94
TABLE 10-5 - EXTRACTION ..................................................................95
TABLE 10-6 - SURFACE SERVICES ............................................................97
TABLE 10-7 - WORKING CAPITAL .............................................................98
TABLE 10-8 - MILLING & ENVIRONMENTAL OPERATING COST ......................................98
TABLE 10-9 - ADMINISTRATION COST .........................................................99
TABLE 11-1 CAPITAL ......................................................................100
TABLE 11-2 - DEFERRED DEVELOPMENT SUMMARY ...............................................100
TABLE 11-3 - DEFERRED DEVELOPMENT - EXCAVATION ..........................................101
TABLE 11-4 - CAPITAL EQUIPMENT ..........................................................102
TABLE 11-5 - SALVAGE VALUE ..............................................................104
TABLE 12-1 - PRE-TAX CASH FLOW SUMMARY ..................................................105
TABLE 12-2 - METAL PRICES AND EXCHANGE RATE .............................................106
TABLE 12-3 - CONCENTRATE TREATMENT CHARGES ..............................................106
TABLE 12-4 - PROJECT SENSITIVITY ........................................................108
FIGURES
FIGURE 1-1 - LOCATION MAP .................................................................8
FIGURE 1-2 - PROPERTY MAP ................................................................10
FIGURE 1-3 - SURFACE INSTALLATIONS .......................................................11
FIGURE 2-1-REGIONAL GEOLOGY MAP OF LANGLOIS MINE .........................................14
FIGURE 2-2 - GEOLOGIC MAP OF THE LANGLOIS MINE AREA ......................................14
FIGURE 2-3 - LONGITUDINAL SECTION (LOOKING NORTH) SHOWING LOCATION OF ZONES 3, 4
AND 97 AND THE CURRENT MINE DEVELOPMENT ..............................................16
FIGURE 2-4 - TYPICAL CROSS-SECTION (LOOKING WEST) THROUGH ZONE 97 SHOWING GEOMETRY
OF THE SULPHIDE MINERALIZATION .......................................................17
FIGURE 2-5 - PHOTOGRAPHS SHOWING .........................................................18
FIGURE 4-1 - MATHEW'S STABILITY GRAPH ....................................................34
FIGURE 4-2 - TYPICAL DRIFT GROUND SUPPORT ................................................35
FIGURE 4-3 - LANGLOIS MINE LONGITUDINAL PROJECTION .......................................39
FIGURE 4-4 - ORE ZONES AND MAIN LEVELS ...................................................41
FIGURE 4-5 - TYPICAL ACCESS RAMP AND SUBLEVEL LAYOUT OF ZONE 4 ...........................41
FIGURE 4-6 - TYPICAL LEVEL IN ZONE 3 .....................................................42
FIGURE 4-7 - ZONE 97 SUBLEVEL ACCESS TO ONE MINING DOMAIN ................................43
FIGURE 4-8 - ZONE 4 STYROFOAM COLUMN SLOT ................................................43
FIGURE 4-9 - ZONE 3 REAMED RAISE IN PASTE BACKFILL .......................................44
FIGURE 4-10 - PASTE BACKFILL DISTRIBUTION NETWORK ........................................47
FIGURE 4-11 - PASTE BACKFILL PLACEMENT ...................................................48
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
FIGURE 4-12 - LONGITUDINAL SECTION - ZONES 3 & 4 .........................................51
FIGURE 4-13 - LONGITUDINAL SECTION - ZONE 97 .............................................52
FIGURE 4-14 - MAXIMUM ANNUAL STOPE SEQUENCING IN ZONE 3 OR ZONE 4 ........................53
FIGURE 4-15 - MAXIMUM ANNUAL STOPE SEQUENCING IN ZONE 97 .................................53
FIGURE 4-16 - VENTILATION NETWORK - YEAR 6 ...............................................61
FIGURE 4-17 - DRIFT CROSS SECTION ........................................................62
FIGURE 4-18 - SLASHING OF ACCESS DRIFT 9 LEVEL, ZONE 97 ..................................63
FIGURE 4-19 - TYPICAL STOPE MUCKING ARRANGEMENT ..........................................64
FIGURE 4-20 - ACTUAL AND FUTURE ORE AND WASTE PASSES .....................................66
FIGURE 5-1 - MILL FLOWSHEET ..............................................................68
FIGURE 5-2 - TAILINGS DEPOSITION PLAN ....................................................72
FIGURE 7-1 - ORGANIZATION CHART ..........................................................84
FIGURE 12-1 MINE SITE COST DISTRIBUTION .................................................107
APPENDICES
APPENDIX A - CERTIFICATES OF QUALIFIED PERSONS
APPENDIX B - BREAKWATER LANGLOIS MINE FEASIBILITY PROCESS, DECEMBER 2000.
APPENDIX C - SRK - LANGLOIS MINE RESERVE AUDIT, 2001
APPENDIX D - BREAKWATER - RESOURCE AND RESERVE SUMMARY, DECEMBER 2000
APPENDIX E - ASSESSMENT OF EFFECTS OF EXTRACTING SHAFT PILLAR
APPENDIX F - CIM STANDARDS ON MINERAL RESOURCES AND RESERVES: DEFINITIONS AND GUIDELINES" (AUGUST 2000)
APPENDIX G - C. PAGE LETTER REPORT, JULY 2000
APPENDIX H - K. REIPAS OBSERVATIONS AND DISCUSSIONS
APPENDIX J - LANGLOIS ROCK MECHANICS REPORT
APPENDIX K - SERVICES TECHNIQUES, APRIL 2000
APPENDIX L - BLASTING SIMULATION
APPENDIX M - BLASTING SKETCH
APPENDIX N - PASTE FILL REPORT
APPENDIX O - LONG TERM
APPENDIX P - MAINTENANCE PROGRAM AND MINE VISIT SUMMARY
APPENDIX Q - TRUCKS
APPENDIX R - EOLAVAL
APPENDIX S - ZONE 97 INTAKE ARRANGEMENT
APPENDIX T - MCINTOSH/REDPATH ENGINEERING REPORT CONCERNING LANGLOIS MINE CONCEPTUAL ORE/WASTE PASS LINING OPTIONS REVIEW
APPENDIX U - LICENCES AND CERTIFICATES OF AUTHORIZATION
APPENDIX V - SAMPLING STATIONS
APPENDIX W - STAFF BUDGET SALARY
APPENDIX X - PRODUCTION SCHEDULE DETAILS
APPENDIX Y - UNIT COSTS
APPENDIX Z - HISTORICAL COSTS
APPENDIX AA - UNDERGROUND DEPARTMENT
APPENDIX BB - MAINTENANCE DEPARTMENT
APPENDIX CC - ELECTRICAL DEPARTMENT
APPENDIX DD - FIXED COSTS
APPENDIX EE - ELECTRIC POWER COST
APPENDIX FF - PROPANE COST
APPENDIX GG - DIESEL FUEL COST
APPENDIX HH - BUSSING CONTRACT
APPENDIX JJ - ENGINEERING DEPARTMENT
APPENDIX KK - GEOLOGY DEPARTMENT
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
APPENDIX LL - VIREMENT
APPENDIX MM - MILL BUDGET
APPENDIX NN - ENVIRONMENTAL BUDGET
APPENDIX OO - ADMINISTRATION DEPARTMENT
APPENDIX PP - OFF SITE ADMINISTRATION
APPENDIX QQ - RAISE BORING ESTIMATES
APPENDIX RR - EQUIPMENT COST
APPENDIX SS - SALVAGE VALUE
APPENDIX TT - CLOSURE PLAN
APPENDIX UU - PRE-TAX CASH FLOW
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
EXECUTIVE SUMMARY
PROPERTY DESCRIPTION
The Langlois mine is situated in northwestern Quebec, approximately
213km north of Val-d'Or. It is an 1,800tpd trackless underground
zinc/copper mine, employing longhole open stoping methods and paste
backfill. The mine facilities include a head frame, shaft, hoisting
plant, a paste backfill plant, mechanical and electrical shops, a
service building, a 2,500tpd zinc/copper concentrator and a tailings
pond.
The Langlois mine has produced zinc (along with lesser values of
copper, silver and gold) from narrow, tabular volcanogenic massive
sulphide ("VMS") bodies with near vertical dips. These massive
sulphide bodies, referred to as zones, are relatively thin (1 to 8m),
but exhibit considerable vertical and lateral extent.
Production has come from two zones to-date, namely Zones 3 and 4,
while a third VMS body, Zone 97, is located approximately 1km to the
east and has only been developed along one level. Zone 97 hosts the
majority of the mining reserves at higher than mine average grades.
HISTORY
The Langlois mine commenced production in 1995, when it was owned by
Cambior Inc. The initial mining method was inappropriate to the
context of the reserves. Large blasthole stopes were attempted in the
narrow undulating ore zone, hosted within highly foliated rock of weak
to moderate strength.
Excessive dilution was experienced, compounded by problems with very
sticky muck and rapidly deteriorating ore passes. Production delays
and shortfalls often occurred.
Some improvements in controlling dilution were made as mining
progressed, but the mine was never able to fully recover from this
situation. When the mine did not deliver according to plan, attempts
were made to lower costs, including starving the mine of capital. This
delayed the capital development required to access the best part of
the reserves in Zone 97.
Breakwater Resources Ltd. purchased 100% of the Langlois mine in May
2000 and operated the mine until November 2000 when it announced that
operations were being temporarily suspended because it had become
uneconomic to operate the mine. The operating cost per pound of
payable zinc for the period May-November 2000 was US$0.52/lb.
OPERATING CHALLENGES
Breakwater realized that any plan to resume production on a profitable
basis would have to address the historic operating challenges at the
Langlois mine that had adversely affected production surety and costs.
Historically, ore passes had proven to be an unreliable method of muck
handling. The problem has been excessive wear rates, followed by
plugging or collapse, resulting in production shortfalls.
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
The Langlois mine had been starved of capital funding for a number of
years, delaying the development of Zone 97. Bringing Zone 97 into
production will directly impact the grade to the mill, and additional
work places will become available, improving the mine's production
potential.
Mechanical availability on mobile equipment was low and equipment
operating costs were high at the time the mine was shut down. Frequent
mechanical breakdowns and equipment shortages hampered production
work.
With ore passes operated at low levels out of necessity, and no
underground ore storage bins, there was insufficient ore storage
between mine and mill to smooth out the peaks and valleys in
underground production.
Stope dilution had been mainly brought under control over the years
through changes in operating practices, but there was a desire to
achieve further improvements.
TERMS OF REFERENCE
SRK Consulting (SRK) was hired by Breakwater in November 2000 to
prepare a feasibility study with the primary objective of bringing the
Langlois mine back to profitability.
With the Langlois mine closed, the experienced mine staff were
available to carry out engineering studies for the feasibility. The
feasibility mining plan has been prepared mainly by the mining
engineers and mine operators experienced in managing the past
production at the Langlois mine. SRK and Breakwater agreed that
production surety was critical, and that the plan must be developed on
an "achievable" basis. SRK's scope of involvement included auditing
the feasibility process and engineering work, and providing an
independent evaluation of the reserves and resources.
FEASIBILITY STUDY PROCESS
To achieve the objectives of the study, a "feasibility process" was
followed. Background information was collected and organized, and a
brainstorming session was held followed by a period of developing the
concepts and scoping options that would address historic mining
problems. The scoping options were assembled into three mining
alternatives (complete mining systems from stope to mill) that were
studied at the pre-feasibility stage.
The pre-feasibility mining alternatives were selected to evaluate the
variables of production rate and cut-off grade.
A decision was made to select a production rate of 450,000 tonnes/year
and to increase the cut-off grade, creating a high-grade alternative
with a mine life of eight years. The operating schedule for the
underground mine will be five days per week, two 8-hour shifts per
day.
Throughout the planning process, the approach taken has been to
introduce only those changes required to address the most serious mine
operating problems. Uncertainty has been reduced by incorporating past
practices where possible and by relying on associated historical
records for productivities and costs.
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FEASIBILITY MINE PLAN
The feasibility mine plan does not include mining of all of the
January 31, 2001 mineral reserves, since it is a high-grade
alternative.
The feasibility mine plan is based on the resources and reserves as of
May 1, 2003 shown in Table I. August 2000 CIM Standards have been
followed.
Table I - Langlois Mine Resources and Reserves
- --------------------------------------------------------------------------------
Tonnage Zinc (%) Copper (%) Silver (g/t) Gold (g/t)
- --------------------------------------------------------------------------------
Mineral Resources
Measured & Indicated 4,980,668 11.15 0.79 53.96 0.09
Inferred 1,254,508 9.71 0.52 40.02 0.14
Mineral Reserves
- --------------------------------------------------------------------------------
Proven & Probable 3,322,760 10.78 0.82 52.13 0.09
- --------------------------------------------------------------------------------
The high-grade mining in the feasibility study does not isolate the
lower grade reserves, which are currently excluded from the mining
schedule. These reserves can be brought into production if metals
prices increase sufficiently.
The feasibility operating plan is shown in Table II.
Table II - Feasibility Operating Plan
- -------------------------------------------------------------------------------------------------------------------------
YEAR 2 YEAR 3 YEAR 4 YEAR 5 YEAR 6 YEAR 7 YEAR 8 YEAR 9 YEAR 10 TOTAL
- -------------------------------------------------------------------------------------------------------------------------
PRODUCTION
Tonnes Milled 73,878 381,045 450,000 450,000 450,000 450,000 450,000 450,000 167,837 3,322,760
HEAD GRADE
Zn (%) 11.29 10.51 10.78 10.93 10.97 10.96 10.73 10.51 10.59 10.78
Cu (%) 0.63 0.63 0.69 0.73 0.87 0.97 0.96 0.89 0.92 0.82
Au (g/t) 0.09 0.08 0.08 0.08 0.09 0.09 0.09 0.08 0.08 0.09
Ag (g/t) 54.35 50.10 51.66 52.54 53.79 53.89 52.16 50.22 51.77 52.13
RECOVERIES
Zn (%) 93.89 93.76 93.81 93.64 93.53 93.54 93.68 93.57 93.56 93.65
Cu (%) 77.46 76.56 78.99 80.48 83.79 85.48 84.98 84.81 85.27 82.33
Au (%) 29.22 30.00 29.22 29.22 28.44 28.44 29.22 28.44 28.44 28.95
Ag (%) 33.87 33.92 35.11 36.50 36.42 36.32 36.03 35.89 35.89 35.75
CONC. GRADE
Zn (%) 52.80 52.80 53.75 54.25 55.00 55.50 56.00 56.00 56.00 54.82
Cu (%) 22.00 22.00 22.50 23.00 24.00 24.50 25.00 25.50 25.50 23.87
TONNES CONC.
Zn (t) 14,832 71,115 84,665 84,897 83,947 82,124 80,774 79,025 29,695 612,075
Cu (t) 1,639 8,354 10,901 11,495 13,668 15,229 14,685 13,320 5,163 94,454
CONTAINED METAL
Zn (t) 7,831 37,549 45,507 46,057 46,171 46,134 45,233 44,254 16,649 335,366
Cu (t) 361 1,838 2,453 2,644 3,280 3,731 3,671 3,397 1,317 22,691
Au (oz) 62 294 338 338 370 370 381 329 123 2,606
Ag (oz) 43,724 208,177 262,424 277,456 283,442 283,192 271,893 260,775 100,363 1,991,345
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The planned average zinc head grade of 10.78% is significantly higher
than the head grades achieved during the last four years of
production, which ranged from 6.4 to 7.9%. The planned zinc grade is
higher because:
o Zone 97 high-grade tonnes will become part of the production
stream for the first time.
o The feasibility plan incorporates a higher cut-off grade
than past mining.
o The feasibility plan includes mining of the higher grade
shaft pillar which was previously considered sterilized.
This report describes the rock mechanics work done to
support this decision. Refer to section 4.1.5
The feasibility operating plan incorporates the following improvements
to ensure reliability of production and to control costs:
o There are no ore passes planned for Zone 97 due to their
unreliable nature. A fleet of new 20 tonne trucks will haul ore
from stoping areas to the shaft area, and the past problem of
collapsing ore passes will be avoided.
o A new steel lined storage bin is planned from level 10 to level
11 to provide ore storage while overcoming the past hang up
problems.
o The feasibility study mining plan provides for pre-development of
several sublevels in Zone 97 so that it can produce higher grade
ore continuously when mine operations resume.
o Several improvements are planned for the underground mobile
equipment fleet:
> Refurbishing of the existing mobile equipment prior to going
back into service.
> Several new units will be purchased to meet mine plan
requirements.
> Many improvements are planned to the mobile maintenance
program.
> Two graders will be purchased, along with a small crusher
for road material.
> A new underground garage is planned on level 13.
o Zone 97 employs an overhand benching method with a reduced
stoping height to control dilution and avoid the associated
delays.
Before the Langlois mine can resume full production, approximately 18
months of construction and development work is required. The major
construction projects include:
o Construction of the level 10 - 11 steel-lined ore bin.
o Steel lining critical ore pass sections.
o Construction work on ore and waste pass dumping points.
o Rehabilitation of the existing underground mobile equipment
fleet.
o Construction of the upgraded ventilation system.
o Backfill systems including paste backfill pumping and
delivery on level 6 and cement slurry delivery to Zone 4.
o Development and construction of a new mobile maintenance
shop on level 13 and shop improvements on level 9.
OPERATING COSTS
The total operating cost to mine gate is estimated at Cdn$55.61 per
tonne milled over the life of the mine. The average operating cost
from Year 4 to Year 9 is $51.73/tonne. This cost is in the range of
the historical 2000 operating cost. Refer to Table III. The operating
cost per pound of payable zinc including smelting, shipping and
by-product credits for copper, silver and gold is US$0.376.
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Table III - Operating Cost
($000's)
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Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total Cost/t
Mining 2,358 7,760 14,467 15,363 15,720 16,143 16,064 15,951 14,954 5,483 124,263 $37.40
Milling 264 858 4,605 5,160 5,061 5,116 5,098 5,029 4,794 1,925 37,910 $11.41
Administration 950 2,352 2,534 2,534 2,534 2,534 2,534 2,534 2,534 1,546 22,586 $6.80
Total Operating 3,572 10,970 21,606 23,057 23,315 23,793 23,696 23,514 22,282 8,954 184,759 $55.61
Total Operating/tonne
Milled - $148.49 $56.70 $51.24 $51.81 $52.87 $52.66 $52.25 $49.52 $53.34 $55.61
Cost/lb Payable Zinc US$ - - $0.408 $0.373 $0.368 $0.359 $0.350 $0.350 $0.345 $0.357 $0.376
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* 6 months
CAPITAL COSTS
An estimated $38.2 million in capital is required over the life of the
mine, equivalent to US$0.043 per pound of payable zinc. This is
comprised of $47.1 million in expenditures on development,
construction and equipment, offset by a salvage value of $8.9 million
at the end of the project.
Table IV Capital
($ 000)
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Item Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total
Deferred Development 4,268 6,717 6,302 3,445 2,845 2,124 2,065 2,293 1,261 125 31,445
Infrastructure & Equipment 340 5,044 4,152 2,387 1,100 1,260 725 500 150 - 15,658
Salvage Value - - - - - - - - - (8,935) (8,935)
Total 4,608 11,761 10,454 5,832 3,945 3,384 2,790 2,793 1,411 (8,810) 38,168
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Approximately $16.4 million must be expended prior to the start of
production at the beginning of Year 3. The majority of the capital
requirements are related to the underground mine, with the mill
accounting for only $0.8 million of the total.
FINANCIAL RESULTS
Total mine site costs (operating and capital) amount to US$0.419 per
pound of payable zinc including smelting, shipping and by-product
credits for copper, silver and gold.
The cashflow is based on mining and milling 3.3 million tonnes at the
grades and recoveries shown in Table II. Metal price and exchange rate
assumptions are shown in Table V.
Table V - Metal Prices and Exchange Rate
-----------------------------------------------
Zinc US$0.50/lb
Copper US$0.80/lb
Silver US$5.00/oz
Gold US$343.00/oz
Exchange Rate Cdn$0.70/$US
-----------------------------------------------
The total net pre-tax cashflow is Cdn$71.1 million. At a zinc price of
US$0.50/lb. the mine produces a positive cashflow starting in Year 4. On a
cumulative basis , the mine is cash positive starting in Year 6. The internal
rate of return is 25.3% and the NPV at 8.0% is $30.9 million.
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Table VI - Pre-Tax Cash Flow Summary
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YEAR * 6 months YEAR 1* YEAR 2 YEAR 3 YEAR 4 YEAR 5 YEAR 6 YEAR 7 YEAR 8 YEAR 9 YEAR 10* TOTAL
- ------------------------------------------------------------------------------------------------------------------------------------
PAYABLE METAL
Zinc (`000 lb.) - 14,649 70,238 85,277 86,307 86,521 86,452 84,764 82,929 31,162 628,299
Copper (`000 lb.) - 759 3,868 5,167 5,575 6,930 7,890 7,770 7,194 2,789 47,942
Silver (`000 oz) - 40 192 241 255 261 261 250 240 92 1,832
Gold (oz) - 12 40 11 - - - - - - 63
SMELTER REVENUE (CDN $000) - 11,624 55,978 68,547 69,843 71,584 72,629 71,212 69,171 26,105 516,693
Smelter, Capital and Operating
Cost (8,180) (29,242) (62,141) (60,549) (57,801) (57,977) (57,221) (55,805) (52,143) (4,559) (445,618)
CASHFLOW (8,180) (17,618) (6,163) 7,998 12,042 13,607 15,408 15,407 17,028 21,546 71,075
Cumulative (8,180) (25,798) (31,961) (23,963) (11,921) 1,686 17,094 32,501 49,529 71,075
NPV @ 8.0% 30,890
10.0% 24,616
12.0% 19,344
IRR 25.3%
Cost/lb. Payable Zinc US$/lb.
Zn Treatment, Shipping Costs - $0.198 $0.198 $0.193 $0.191 $0.188 $0.186 $0.183 $0.183 $0.183 $0.189
Credit for Byproducts - ($0.038) ($0.039) ($0.042) ($0.045) ($0.054) ($0.060) ($0.060) ($0.058) ($0.059) ($0.052)
Operating Cost excl. Deprec. - $0.558 $0.249 $0.222 $0.222 $0.225 $0.224 $0.226 $0.220 $0.233 $0.239
Total Operating Cost - $0.718 $0.408 $0.373 $0.368 $0.359 $0.350 $0.350 $0.345 $0.357 $0.376
Capital - $0.562 $0.104 $0.048 $0.032 $0.039 $0.027 $0.023 $0.012 ($0.198) $0.043
Total - $1.280 $0.512 $0.421 $0.400 $0.398 $0.377 $0.373 $0.357 $0.159 $0.419
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RISKS
The project economics are sensitive to the price of zinc. A +/-10%
change in the zinc price causes a change of +/-$29.4 million in the
net pre-tax cashflow.
The current zinc price (May, 2003) is near the bottom of a long-term
cycle, and the project is presently uneconomic.
OPPORTUNITIES
The feasibility study has been prepared on a conservative basis,
introducing only those changes required to address the most
significant operating challenges. Many of the productivities and costs
used are based on past mine practices, some of them not fully
efficient.
During the development of the study, several opportunities were
identified that have the potential of improving the project economics.
These are fully described in section 13.0. Some of the more promising
opportunities are summarized below.
o Shotcrete could be used to support ore drifts, reducing
rehabilitation costs and improving stope cycle times.
o Electric-hydraulic drilling equipment could be used to
replace some of the production drills.
o There is potential to increase mineral reserves through
ongoing exploration, or as a result of higher metal prices.
o Management effectiveness. There is an opportunity during the
capital period to review the management systems and
personnel.
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1.0 INTRODUCTION
1.1 TERMS OF REFERENCE
SRK Consulting (SRK) met with Breakwater Resources Ltd. (Breakwater)
on 24 November 2000 to discuss SRK's involvement with a study team to
return the Langlois mine to profitability. SRK responded with a
discussion document that outlined a feasibility process. Breakwater
agreed to the proposed study process and framework, and requested SRK
to prepare a proposal that would outline the full feasibility work
plan and schedule. It was agreed that the Langlois, Breakwater and SRK
personnel would work as a team.
With the Langlois mine closed, the experienced mine staff were
available to carry out engineering studies for the feasibility. SRK's
scope of involvement in 2001 included:
o Review and audit the feasibility process and engineering
work
o Independent evaluation of the reserves and resources
o Rock mechanics for stope design
o Final sign-off on the feasibility document
The objective of the feasibility study was to design an economic and
practical mine based on current reserves, including zone 97. It was
anticipated that capital would be needed to reopen the mine, and
therefore the study would be prepared to a bankable standard. Work
began in December 2000.
The sources of information used in preparing the 2001 feasibility
report included the following:
o Langlois mine reserves, audited by SRK, June, 2001
o Rock mechanics parameters developed by SRK and Langlois mine
personnel
o Existing mine engineering plans, schedules and drawings
o Historical operating and cost data from mining and milling
operations
o Previous consultants reports
o Information and estimates prepared by SRK, equipment
suppliers and contractors
The key project personnel are listed below:
Breakwater Resources Ltd. Langlois Mine Personnel
------------------------- -----------------------
C. K. Benner, President and CEO Daniel Vallieres, Acting Manager
Torben Jensen, Manager of Engineering Marc Bernard, Mine Superintendent
and North American Exploration Tony Brisson, Chief Geologist
Denis Vaillancourt, Senior Geologist
Ron Durham, Mine Planning Engineer
SRK Consulting Alain Cossette, Rock Mechanics Engineer
-------------- Andre Dessureault, Senior Mine Technician
Chris Page, Corporate Consultant Martine Deshaies, Senior Metallurgist
Ken Reipas, Principal Mining Engineer
Mike Michaud, Senior Project Geologist
Jean-Francois Couture, Associate Geologist
Jarek Jakubec, Principal Rock Mechanics Engineer
Each of the SRK project team members have spent several days on site
and in the underground mine.
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In May 2003, Breakwater updated the 2001feasibility study to
incorporate the results of additional underground diamond drilling in
Zone 97. SRK was retained to provide an independent review of the
updated feasibility study.
1.2 BACKGROUND
1.2.1 PROPERTY DESCRIPTION, LOCATION AND ACCESS
The Langlois mine is situated in northwestern Quebec, approximately
48km northeast of the town of Lebel-sur-Que villon and 213km north of
Val-d'Or. Lebel-sur-Que villon has a population of 3,500. The mine is
accessed via a gravel road maintained by Breakwater and Domtar, a
forest products company with operations in the area. The mine
facilities include a head frame, shaft, hoisting plant, a paste
backfill plant, mechanical and electrical shops, a service building, a
zinc/copper concentrator and a tailings pond. The mine produces zinc
and copper concentrates that are sold and forwarded to smelters for
further processing. Gold and silver by-products are also produced.
Figure 1-1 - Location Map
[PICTURE]
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1.2.2 LEGAL TITLE
As of June 2003, Breakwater Resources Ltd owns 100% of the Langlois
mine. The mill area, Zones 3, 4 and 97 are held through two mining
leases totalling 133 hectares, granted until the year 2015. The
tailings pond is held under a separate lease totalling 188 hectares,
granted until 2003. The property also contains mineral claims in
Grevet, Ruette and Mountain Townships, some of which are in
partnership with Metco Resources Inc. or BP-Norex. Cambior Inc. was
the sole owner of the mine from September 1993 to May 2000.
All mining and environmental permits have been issued and are up to
date. Refer to section 6.0.
There are no royalties payable on mineral production at Langlois.
1.2.3 PROPERTY HISTORY
The deposit, originally known as the Grevet Project, was discovered in
1989 by Serem-Quebec Inc. (50 %) and VSM Exploration Inc. (50 %).
Cambior acquired its initial 50 % interest in the Grevet Project in
July 1992 with the acquisition of VSM. In September 1993, Cambior
purchased the remaining 50 % interest in the project from Serem to
obtain 100 % ownership.
In 1994, Cambior commenced an underground exploration program designed
to delineate mineral reserves. Due to the success of the underground
exploration program, development work on the property commenced in the
third quarter of 1994 and was completed in December 1995. Commercial
production began at the mine in February 1996. The production goal was
900,000 tonnes per year, however, this was never achieved.
Table 1-1 - Historical Production
-----------------------------------------------------------------------------
1996 1997 1998 1999 2000
-----------------------------------------------------------------------------
PRODUCTION
Tonnes Milled 537,234 261,068 414,742 402,224 310,466
-----------------------------------------------------------------------------
Production at Langlois was halted in December 1996 due to high
dilution problems in the mine caused by the high stope heights. These
problems were rectified by modifications to the mining method and
production was resumed in July 1997. The production rate was decreased
slightly to under 500,000 tonnes per year on a five day per week
basis. The mine operated during 2000 at the budgeted production level,
however it experienced ore pass problems that, in combination with
weak metal prices, forced the company to temporarily suspend
operations and place the mine on a care and maintenance basis.
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Figure 1-2 - Property Map
[PICTURE]
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Figure 1-3 - Surface Installations
[PICTURE]
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1.3 FEASIBILITY STUDY OBJECTIVES AND APPROACH
The primary objective of the study was to bring the Langlois mine back
to profitability. This was to be done by using the experience of
personnel at the Langlois mine in partnership with SRK's input and
study management. The specific objectives to be met were to develop:
o an extraction strategy that maximises utilisation of the
resources, combining the developed resources in Zone 3 and 4 with
those of Zone 97
o mining methods that are appropriate to the context (variable
geometry, high value mineralization, weak wall rocks)
o a mining method that minimises dilution
o a reliable production system for delivering metal to the mill
o maximum use of the existing infrastructure
To achieve these objectives, a "feasibility process" was followed that
ensured all project team members understood the available background
information. Such background included; mineralization and reserves,
rock mass characteristics, previous and current stope designs,
excavation behaviour, stoping efficiency, ore pass problems,
production achievements and historic costs.
A series of technical sessions were held to build on the work
conducted by the Breakwater staff done prior to the initiation of the
feasibility study. This was followed by a period of developing
concepts and scoping options for mining.
The scoping options were assembled into three mining alternatives that
were studied and ranked at the pre-feasibility stage. The best
alternative was advanced to feasibility level and was fully detailed
in the 2001 feasibility report, which has since been updated in 2003
due to the additional drilling in Zone 97.
Throughout the planning process, the approach taken has been to
introduce only those changes required to address the difficult
operating challenges described in Section 3.0 of this report. This
approach reduces uncertainty, by continuing past practices where
applicable and relying on historical records for productivities and
costs.
During the process, several opportunities were identified that hold
potential to further improve productivity and operating costs. These
opportunities will be evaluated and tested during operations at
Langlois. They are individually discussed in Section 13.0 of this
report.
A complete description of the 2001 feasibility process is included in
the reference document Appendix B - Breakwater Langlois Mine
Feasibility Process, December 2000.
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2.0 GEOLOGY
2.1 REGIONAL AND PROPERTY GEOLOGY
The Langlois mine produces zinc (along with lesser values of copper,
silver and gold) from narrow, tabular volcanogenic massive sulphide
("VMS") bodies. They are hosted within mafic to intermediate volcanic
and volcaniclastic units in the central-east portion of the northern
Archean volcanic belt of the Abitibi Sub-province, or more precisely,
within the Miquelon segment. The volcanic complex is bounded to the
south by the Mountain South pluton, a large synkinematic felsic
intrusion whose thermal aureole has metamorphosed the volcanic rocks
(Figure 2-1).
The lithologies in the area consist predominately of a succession of
mafic to intermediate lava flows and volcaniclastics with less
abundant felsic volcanic and sedimentary units. The predominant
structure in the area is the Cameron shear zone, which trends 120(0)
and extends for more than 80km along strike and is up to 5km thick.
The massive sulphide horizons at the Langlois mine are hosted by the
strongly schistosed rocks of the Cameron shear zone. The intense
ductile deformation has obliterated most of the primary stratigraphic
relationships and textures and has imparted a strong easterly
penetrative foliation. The rock sequence has been affected by a
regional deformation, which has formed sub-vertical isoclinal folds.
In addition, the regional metamorphism altered many of the rocks to
green schist facies.
Numerous sulphide bodies in the region have been identified within
volcanogenic corridors oriented easterly, oblique to the strike of the
volcanic complex (Figure 2-2). These corridors may represent a primary
fracture pattern that allowed volcanogenic fluids to escape. Three
corridors are host to several significant zinc-rich massive sulphide
bodies, including Langlois mine, Grevet B and Orphee. A fourth
corridor may exist in the area between Grevet B and Orphee properties
where a polymetallic stockwork zone was intersected by recent
drilling. These volcanogenic corridors are grossly spaced at regular
intervals across the volcanic complex at approximately 800 to 1,000m.
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Figure 2-1-Regional Geology Map of Langlois Mine.
[PICTURE]
Figure 2-2 - Geologic Map of the Langlois Mine Area.
[PICTURE]
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2.2 MINERALIZATION
Ore production at the Langlois mine has come exclusively from two
zones, namely Zone 3, since 1995, and Zone 4, since 1997. A third
lens, Zone 97, which is located approximately 1km to the east, has
been developed along one level. Zones 3 and 4 are located between 100
and 700m below the topographic surface, while Zone 97 is located
between 300 and 900m below surface (Figure 2-3). Zone 97 consists of
several distinct massive sulphide bodies inscribed within a
predominantly felsic package of rocks approximately 100m thick. In
this package, massive sulphide occurrences are closely distributed and
many have limited lateral or vertical extent. However, three zones
display continuity across several drill sections, namely Zone 97
South, Zone 97 Main and Zone 97 North (Appendix C - SRK - Langlois
Mine Reserve Audit, 2001).
Each massive sulphide body is relatively thin (1 to 8m), but with
considerable vertical and lateral extension (> 500m in either
direction). The massive sulphide zones trend southeast with a near
vertical dip, sub-parallel to the regional structural fabric. The
zones are stacked across the felsic sequence along a narrow corridor
slightly oblique to the main structural trend (Figure 2-4). From
southwest to northeast the zones are: Zone 5 (small uneconomic lens
near surface), Zone 4, Zone 3 and Zone 97. In longitudinal section,
each massive sulphide zone portrays an elongated lensoid shape, whose
long axis plunges moderately towards the southeast, parallel to the
plunge of the regional stretching lineation. In addition, the centre
of gravity of each lens becomes progressively deeper moving along the
stacking corridor toward the northeast.
The sulphide zones consist of massive and semi-massive (stringer)
sulphide mineralization of primarily pyrite, sphalerite (predominantly
marmatite) and to a lesser degree, chalcopyrite and pyrrhotite (Figure
2-5). Sulphide mineralization varies from fine to coarse grained and
with or without layering. The sulphides have experienced significant
remobilisation, deformation and a recrystallization. Locally, sulphide
mineralization can grade up to 60% zinc. Mafic dikes cut the
mineralized zones in many areas, and have historically been a major
contributor to dilution. In addition, the well-foliated, chloritic
volcanic host rocks have caused ground-control problems and excessive
dilution.
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Figure 2-3 - Longitudinal Section (looking north)
Showing Location of Zones 3, 4 and 97 and the Current Mine Development.
[PICTURE]
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Figure 2-4 - Typical Cross-Section (looking west) through Zone 97
Showing Geometry of the Sulphide Mineralization.
[PICTURE]
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Figure 2-5 - Photographs showing
[PICTURE]
a)
[PICTURE]
b)
[PICTURE]
c)
a) banded pyrite-sphalerite-chalcopyrite in Zone 3,
b) highly deformed banded sphalerite-pyrite mineralization in Zone
97, and,
c) typical drill core from Zone 97 showing highly foliated volcanics
adjacent to the massive sulphide mineralization.
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2.3 MINERAL RESOURCES AND RESERVES
The resources and reserves for the Langlois mine occur within three
separate zones, namely Zones 97 and 3, which host the majority of the
resources and reserves, and Zone 4. The current mineral resources in
this feasibility study for Zones 3 and 4 are essentially the same as
the January 31, 2001 resources calculated by Breakwater, while the
resources for Zone 97 have been re-estimated by SRK first in 2001
(Appendix C - SRK - Langlois Mine Reserve Audit, 2001) and
subsequently in May, 2003, incorporating the latest drill results. The
updated mineral resources and reserves for Zone 97 are considerably
different than the reserves estimated by Breakwater in 2001 due
primarily to the fact that the reserves are now based on a mining plan
designed to optimize the economics of the resources after examining
various mining and processing scenarios.
2.3.1 DATABASE
The database for the Langlois mine consists of 1,992 surface and
underground core holes, in excess of 2,000 underground chip samples
and several thousand muck samples. The database includes the survey,
assay, density and geological data for each drill hole and channel
sample. The database is maintained at the mine site utilizing the
Prolog computer software program, and has since been imported into the
Gemcom software package by SRK. The current resources and reserves
have been estimated based on 248 core hole intersections and 62
channel sample strings for Zone 97 (which includes 27 drill hole
intersections from May, 2003), 389 core holes and 113 chip samples for
Zone 3 (Main), and 45 and 49 core holes and numerous chip samples for
Zones 3A and 4, respectively.
The majority of the drilling for Zones 3 and 4 was completed prior to
Breakwater acquiring the mine in 2000. However, since that time
Breakwater has continued to drill Zone 97 in order to increase data
density and better define the extents of the mineralization. The
majority of drilling is either BQ or AQ diameter core and is typically
drilled along sections aligned north-south, intersecting the
mineralized zones at as close to a right angle as possible. The
section lines are spaced at 20m intervals in Zone 97 and 10m in Zones
3 and 4. Both the drilling samples and chip samples were designed to
cross the entire width of the massive sulphide mineralization. Drill
core recovery is typically greater than 95%. The drill core is stored
on-site in racks, with sample numbers marked for easy review.
After a review of the data, it is SRK's opinion that the drill logs
provide sufficient description and recognition of the lithology,
alteration, geological structures and mineralization to correlate
geological boundaries between drill holes. This indicates that
previous exploration personnel and current mine operating personnel
have exercised great care and attention to detail in the collection,
verification and storage of the data.
2.3.2 SAMPLING METHOD AND APPROACH
Sampling and assaying methodologies and check assay procedures for
drill core have been well documented by Breakwater. The procedure for
the geologist is to describe the lithology, alteration, structure and
other details and mark out and label the sample intervals and numbers
on the core boxes. The sample intervals are designed not to cross
lithologic boundaries or massive sulphide facies (Appendix D -
Breakwater - Resource and Reserve Summary, December 2000).
Typically, drilling data is spaced at approximately 40m in Zone 97
except between level 8 and 9 (i.e. 60m levels), where the spacing is
20m, 10-15m in Zone 3, except in the area between level 6
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and 7, where the majority of resources remain, and 20m for Zone 4.
Mineralization intersected in drill holes is sampled according to
geology with samples typically ranging in core length from 0.2 to
1.0m. For each sample, a measure of density is taken as well as an
assay for silver, copper, lead and zinc and for gold whenever
chalcopyrite is visible. The core is cut in half with a hydraulic
splitter, with half forming the sample and the other half left in the
box. All of the sampling was carried out by geologists or qualified
technicians under the direct supervision of the project geologist.
Due to the visible nature of the mineralization, estimates of the zinc
grade are made during underground face mapping and core logging and
correlated with the assay results, thereby providing an opportunity to
identify any assay discrepancies.
Several historic underground holes (<1%) intersected the massive
sulphide zone at an angle of less than 15 degrees, and as such, these
results are considered to be less reliable. Although the assay data
from these drill intersections is considered by SRK to be reliable,
and therefore used in the resource estimate, their spatial
distribution was not used to define the geometry of the sulphide
bodies. A number of intersections from surface drilling were handled
in a similar manner.
2.3.3 SAMPLE PREPARATION, ANALYSES AND SECURITY
SRK had an opportunity to visit the on-site assay laboratory and to
review the QA/QC program at the mine. All of the assays for the
operation, including exploration and definition drilling, chip and
muck sampling and mill samples are completed at the mine. Breakwater
uses a relatively common method of analysis of digestion followed by
atomic absorption for zinc and copper assays, and a fire assay
followed by atomic absorption for silver and gold assays. Internal
quality control measures established by Breakwater ensure the quality
of sample preparation and accuracy for zinc, copper, silver and gold
assays and density measurements. The assay laboratory has an extensive
QA/QC program consisting of sample blanks, standards and duplicates.
In addition, at least 5% of the samples are sent to an independent
commercial assay laboratory within the Abitibi region.
An extensive quality control program at the Langlois mine, including
sample preparation, analysis and security, has ensured that the assay
data used in the resource and reserve estimates is very reliable.
2.3.4 DATA VERIFICATION
After reviewing the sampling and assaying procedures and the extensive
QA/QC program implemented by Breakwater, SRK is confident of the
quality of the data. Although verification sampling was not completed
by SRK, numerous drill cores were examined to compare the recorded
geology and mineralization with the assay grades. In addition, SRK had
an opportunity to review the results of ore reconciliation that has
been completed at the mine since 1995. This reconciliation compared
the resources/reserves with the mill production, providing an
opportunity to verify the quality of the data.
2.3.5 RESOURCES FOR ZONES 3 AND 4
The resources for Zones 3 and 4 have been estimated by Breakwater as
of January 31, 2001, utilizing 2-dimensional polygons on
cross-section, using the information from core drilling and
underground chip sampling across development faces. Two-dimensional
kriging and polygons on
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longitudinal section have also been used to a lesser degree, typically
in areas where there is less available data. The resources have been
estimated at several NSR cut-off values based on the width of the
mineralized zone, with a minimum mining width of 3.0m. The January 31,
2001 resources and reserves calculated by Breakwater and audited by
SRK (Appendix C - SRK -Langlois Mine Reserve Audit, 2001) are
summarized in the following tables.
Table 2-1 - Resources for the Langlois Mine as of January 31, 2001
- ------------------------------------------------------------------------------------------------------------------------
Category Zone Tonnes Zinc Copper (%) Silver Gold NSR
(%) (g/t) (g/t) ($Cdn)
- ------------------------------------------------------------------------------------------------------------------------
Measured Zone 3 567,505 9.56 0.55 31.84 0.11 86.93
Zone 4 49,943 15.24 0.50 40.49 0.16 132.90
Zone 97 254,742 12.77 0.78 62.86 0.12 118.09
Total 872,190 10.82 0.61 41.50 0.12 98.66
Indicated Zone 3 929,384 9.59 0.51 34.45 0.05 86.74
Zone 4 232,393 11.47 0.58 42.36 0.13 103.60
Zone 97 1,936,315 13.22 0.90 55.09 0.08 122.70
Total 3,098,092 12.00 0.76 47.94 0.07 110.48
Total Measured and Indicated All Zones 3,970,282 11.74 0.73 46.53 0.1 107.88
- ------------------------------------------------------------------------------------------------------------------------
Inferred Zone 3 543,716 8.46 0.29 23.40 0.11 74.09
Zone 4 206,842 9.26 0.29 32.98 0.13 81.17
Zone 97 451,490 12.24 1.32 48.51 0.03 119.86
Total 1,202,048 10.02 0.68 34.48 0.08 92.50
- ------------------------------------------------------------------------------------------------------------------------
Note: Mineral resources include reserves.
Table 2-2 - Reserves for the Langlois Mine as of January 31, 2001
-------------------------------------------------------------------------------------------
Category Zone Tonnes Zinc Copper (%) Silver Gold NSR
(%) (g/t) (g/t) ($Cdn)
-------------------------------------------------------------------------------------------
Proven Zone 3 448,377 8.76 0.50 33.45 0,11 79.96
Zone 4 60,071 12.04 0.39 40.47 0.16 105.60
Zone 97 256,109 11.13 0.70 66.62 0,11 104.07
Total 764,557 9.81 0,56 45.12 0,11 90.05
Probable Zone 3 799,800 8.10 0.43 34,28 0,05 73,90
Zone 4 166,854 11.42 0.58 55,29 0,13 104.03
Zone 97 2,183,013 11.04 0.75 55,34 0,08 103.22
Total 3,127,893 10.34 0.66 50.31 0.07 96.04
-------------------------------------------------------------------------------------------
Total Zone 3 1,248,177 8.34 0,45 33,99 0,07 75,90
Zone 4 226,926 11.58 0,53 51,37 0,14 104,44
Zone 97 2,439,121 11.05 0,75 56.52 0,08 103,31
Total 3,914,224 10.22 0.64 49.04 0.08 94.63
-------------------------------------------------------------------------------------------
Note: The NSR values are based on US$0.55/lb zinc, US$0.90/lb copper, US$5.00/oz silver,
US$275.00/oz gold and an exchange rate of Cdn$1.55=US$1.00
Mine geology personnel extrapolate the zones of "potentially economic"
sulphide mineralization from hole to hole only when sufficiently
confident. Generally, the limit of the "ore grade" higher-grade zinc
mineralization is congruent with the boundaries of the massive
sulphide mineralization. Once the outlines of the sulphide
mineralization are defined, the assays are composited across the
entire width of the mineralized envelope, or 3.0m horizontal
thickness, whichever is greater, in order to accommodate the
2-dimensional, cross-section model. In general, only regular ore
shapes are delineated. Isolated zones of mineralization, areas of
erratic grade, or areas of geologic
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uncertainty are not extrapolated, and are therefore not included in
the resources/reserves until additional confirmation data is acquired.
The mafic dikes cannot be mined separately, as they are too thin to
selectively mine, and therefore, are included in the mineralized zone
for resource estimation. SRK considers this to be an appropriate
approach.
There are two exceptions, Zone 3 between level 6 and 7 where ordinary
kriging was used to estimate the resource based solely on the results
of the surface exploration core drilling, and a portion of Zone 4
where two-dimensional polygons on longitudinal sections were used,
particularly along the perimeter of the mineralized zone. These
methods were used due to the somewhat larger drill spacing. Any areas
estimated using polygons on longitudinal section have been classified
as Inferred Resources (CIM 2000 definitions). The resources from Zones
3 and 4 are essentially unchanged from the January 31, 2001 resources
prepared by Breakwater.
In order to evaluate the polygonal model constructed by Breakwater,
SRK constructed a parallel inverse distance resource model. The two
models correlate very well within the same volume, confirming the
validity of the resources for Zone 3. In addition, SRK constructed a
second grade model using ordinary kriging to interpolate the zinc
grades into the model. There is essentially no difference between
these interpolation methods at a zero grade cut-off; however, the
difference between these two interpolation techniques is considerably
more pronounced at higher cut-off grades. This would only become a
significant concern if the cut-off grade were increased to a point
where the mine planning selectively targets higher-grade stopes
estimated by polygonal methods.
2.3.6 RESOURCES FOR ZONE 97
Although Breakwater had estimated the resources for Zone 97 as of
January 31, 2001 using a polygonal methodology (summarized in Tables
2-1 and 2-2), SRK re-estimated the resources in 2001 and again in 2003
primarily because:
o The resource for Zone 97 was initially estimated over a 3.0m
minimum mining width, which masked the opportunity to
determine the in-situ resource, and therefore, evaluate
different mining methods such as reduced mining widths,
different production rates, etc.
o Based on the relatively high variability of grade in Zone
97, SRK utilized ordinary kriging to estimate the resource,
which provided a more appropriate amount of averaging of
grades during interpolation, particularly with the larger
drill spacing of approximately 40m.
o Kriging was also considered more suitable to Zone 97 than a
polygonal method because of the pronounced trend of the zinc
mineralization moderately down plunge to the east, which was
evident on the grade contour plots on longitudinal section
and confirmed by the very well defined curves on the
semi-variogram (tool to measure the spatial continuity of
grade) and;
o Incorporation of 2001 and 2003 drill program results.
SRK then constructed a new resource model for Zone 97 using ordinary
kriging to interpolate grades. In order to construct this model, SRK
completed a preliminary review of the zinc grade distribution by
completing a variographic analysis. Although the variograms differ
somewhat by metal, they generally show the direction of maximum
continuity to be moderately plunging to the east. The most
characteristic feature of the mineralization is the moderately high
local variability in grade, with the nugget value (measure of local
variability) for zinc approximately 3/10 of the sill value (measure of
variance), confirming that local predictions of grade would not be
reliable and grade interpolation using kriging was an appropriate
technique.
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In order to complete the resource estimate for the three zones, the
assay values for each metal (Zn, Cu, Ag, Au) and density were
composited using density and length weighting over the true thickness
of the zone. Variography and kriging were completed on the product of
the grade (including Zn, Cu, Ag, Au and the density) by the true
thickness. This value per model block was subsequently divided by the
true thickness to obtain the grade. In this way the true thickness and
density are considered during grade interpolation, something not
possible with cross-sectional interpolation. Capping of the zinc
grades was not deemed to be necessary and is appropriately accounted
for during kriging.
A preliminary analysis of the composites completed by SRK indicates
that the copper and zinc mineralization are slightly skewed to the
left (i.e. mean is to the left of the mode), while the gold and silver
composites approach a log-normal distribution. It also indicates that
the coefficient of variation (a measure of grade variability) is
relatively high compared to other more typical VMS deposits that have
experienced a lesser degree of metamorphism.
SRK had an opportunity to compare the results of the exploration and
production core drilling and the underground chip samples for the same
area/volume of the deposit in order to ensure that the chip samples
and drill hole data have similar sample support, and could therefore,
be combined and used for grade interpolation. The mean grades and the
coefficient of variation, a measure of grade variability, compare very
well indicating congruent sample support. In addition, any chip
samples that were isolated and did not form a continuous string of
samples across the mineralized zone were not used in the resource
estimate.
There were also a number of historic holes, which have a number of
missing gold assays. In these areas, Breakwater has typically used the
average gold grades based on mill head grades, since gold does not
have a definable correlation with any of the other metals.
Fortunately, the majority of these holes lacking adequate gold assays
occur in Zones 3 and 4 where there is production data to estimate the
gold grade. Zone 97 has an adequate number of gold assays from
drilling and underground sampling to estimate a reliable resource.
2.3.7 CONVERSION OF RESOURCES TO RESERVES
The reserves were calculated by converting indicated or measured
resources based on a minimum mining width of 3.0m for Zones 3 and 4
and 2.2m for Zone 97. The reserves consist of contiguous zones of
mineralization delineated in the geological model, while isolated
areas are not included.
2.3.8 RESERVES FOR ZONES 3 AND 4
The reserves for Zones 3 and 4 are based on the mining method selected
for developed and undeveloped areas. Where stope development already
exists, stope dimensions will remain 30m in height. For the portions
of Zones 3 and 4 that are not yet developed, sublevel spacing will be
reduced to 20m. All stopes will be mined in a retreating sequence and
will be accessible by ramp.
Ore recovery, excluding pillars, is 95% for Zones 3 and 4. Dilution,
defined as the volume of waste recovered as a percentage of the ore
removed, has a density of 2.75 t/m3 and a grade of zero and estimated
to be 25% by volume (approximately 18-20% by tonnage) for Zone 3 and
35% for Zone 4. The higher dilution rate for Zone 4 is related to
ground conditions, where sheared,
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chloritic host rocks often account for excessive dilution. Internal
dilution, and to some degree, external dilution, are accounted for
during the initial construction of the two-dimensional sectional
polygon, when regular shapes are drawn and the dilution is included in
the composited grades.
Dilution and recovery for the planned mining method and stope
dimensions for Zones 3 and 4 are based on the historic mining
information. It is SRK's opinion that the planned dilution and mining
recovery factors are appropriate given the characteristics of the rock
strength, continuity of grade and geometry, and the mining method.
The current (2003 feasibility study) reserves for Zone 4 remain
essentially the same as estimated January 31, 2001, with the exception
of a pillar now being left behind. The reserves for Zone 3 are
substantially different for the feasibility study, which consists of
fewer tonnes but at a higher grade. This is primarily due to the fact
that in 2001 a decision was made to include the mining of the shaft
pillar (higher grade portion of the Zone 3 resource) in the mine plan
(Appendix E -Assessment of Effects of Extracting Shaft Pillar, 2001).
In addition, several of the relatively lower grade areas comprising
Zone 3 have been excluded from the mine plan.
2.3.9 RESERVES FOR ZONE 97
Zone 97 will be mined using an overhand benching method of stoping.
Stope dimensions are planned at 11m high and, on average, 94m in
strike length. The average ore width is approximately 3m. Stope
sequencing will be in a retreat fashion to the central access cross
cut provided at each sublevel. The reduced stope height is designed to
control dilution while allowing long, yet stable stope dimensions. The
longer strike length provides for more continuous mining operations,
such as drilling with less moving between work places. Ramp access is
planned, as is backfilling with cemented paste fill.
The ore recovery for Zone 97, excluding pillars is 95%, which is
comparable to the historical recovery rates realized in Zones 3 and 4.
Because of the highly variable width of mineralization in Zone 97, SRK
constructed a "skin or layer" adjacent to the mineralized zone to
calculate dilution, rather than a global dilution factor that ignores
geometry and assumes a constant dilution percentage over mineralized
zones of variable thickness. A thickness of over-break equivalent to
35cm off of each wall has been assumed with a minimum mining thickness
of 2.2m. This constitutes an average dilution factor of approximately
30% by volume, which correlates well with the historical dilution for
Zones 3 and 4.
In total, there are three mining pillars in Zone 97. The pillars,
located along Levels 9, 10 and 13, are 9.5 metres high. It is assumed
that 50% of the pillars will be recovered once the reserves in these
areas are exhausted.
Typically, drilling data is spaced at approximately 40m in Zone 97
except between level 8 and 9 (i.e. 60m level spacing), where the
spacing is 20m. This is considered to be an adequate drilling density
for delineating the zones of sulphide mineralization; however, this
drilling is not sufficient to predict the occurrence of numerous mafic
dikes that occur within the ore zones and late stage faults that
transect the mineralization. As such, drilling will be required for
Zone 97, and for a portion of Zones 3 and 4 to obtain sufficient data
density before final stope design can be completed.
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2.3.10 RESOURCE AND RESERVE CLASSIFICATION
The resource classification is based essentially on the density of
drill hole and chip sample data and the continuity of zinc grade
(since zinc accounts for the majority of the value of the deposit).
Isolated areas of mineralization or areas that have currently no
indication that they can be mined at a profit have not been included
in the reserves.
The resources and reserves for the Langlois Mine as of May 1, 2003
have been prepared in accordance with "CIM STANDARDS ON MINERAL
RESOURCES AND RESERVES: DEFINITIONS AND GUIDELINES" (August 2000)
(Appendix F). Accordingly, the resources have been classified as
measured, indicated or inferred and the reserves have been classified
as proven or probable based on the measured or indicated resources.
Measured resources, and hence proven reserves, occur where the grade
and geometry of the deposit is known with a high degree of confidence,
allowing detailed mine design and mine planning to proceed. Measured
resources are typically located in Zones 3 and 4, adjacent to
underground workings where the data spacing is typically less than
15m. The resources adjacent to the development on level 9 in Zone 97
were classified as proven reserves as of January 31, 2001; however,
they have now been classified as probable by SRK.
In the opinion of SRK, the density of assay data and the knowledge of
the geometry of the majority of Zones 97, 3 and 4, provide sufficient
confidence to determine the grade and tonnage of the deposit. In
addition, the continuity of grade and the geometry of the deposits are
known with sufficient confidence to complete a mine design and mine
plan. Although the actual location of planned stopes may vary somewhat
during mining, this is not expected to significantly change the
economics of the project. The indicated resources, and hence probable
reserves, have been defined primarily within that portion of the
deposit having a drill spacing of approximately 40m or less.
Inferred resources are located primarily along the perimeter of the
deposit, along the strike and down dip extensions of the deposit.
2.3.11 TABULATION OF RESOURCES AND RESERVES
The resources and reserves for the Langlois mine occur within three
separate zones, namely Zones 97 and 3, which host the majority of the
resources and reserves, and Zone 4, as summarised in the following
tables.
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Table 2-3 - Resources for the Langlois Mine as of May 1, 2003
- -------------------------------------------------------------------------------------------------
Category Tonnage Zinc (%) Copper (%) Silver (g/t) Gold (g/t)
- -------------------------------------------------------------------------------------------------
MEASURED & INDICATED
ZONE 3
Measured 567,505 9.56 0.55 31.84 0.11
Indicated 929,384 9.59 0.51 34.45 0.05
ZONE 4
Measured 49,943 15.24 0.50 40.49 0.16
Indicated 232,393 11.47 0.58 42.36 0.13
ZONE 97 - Indicated only
(no measured)
Zone 97 Main 2,203,958 13.71 1.18 72.17 0.12
Zone 97 South 165,187 8.79 0.35 46.61 0.01
Zone 97 North 832,298 7.33 0.38 48.12 0.03
TOTAL MEASURED & INDICATED 4,980,668 11.15 0.79 53.96 0.09
INFERRED RESOURCES
Zone 3 543,716 8.46 0.29 23.40 0.11
Zone 4 206,842 9.26 0.29 32.98 0.13
Zone 97 503,950 11.24 0.86 60.83 0.17
- -------------------------------------------------------------------------------------------------
TOTAL INFERRED 1,254,508 9.71 0.52 40.02 0.14
- -------------------------------------------------------------------------------------------------
Note: Mineral resources include reserves.
Table 2-4 - Reserves for the Langlois Mine as of May 1, 2003
- -------------------------------------------------------------------------------------------------
Category Tonnage Zinc (%) Copper (%) Silver (g/t) Gold (g/t)
- -------------------------------------------------------------------------------------------------
PROVEN & PROBABLE
ZONE 3
Proven 459,513 8.92 0.52 35.16 0.11
Probable 642,591 9.19 0.45 38.65 0.02
Subtotal 1,102,104 9.08 0.48 37.19 0.06
95% Recovery 1,046,999 9.08 0.48 37.19 0.06
Pillar Zone 3 (48,833) 9.50 0.48 39.58 0.04
TOTAL ZONE 3 998,166 9.06 0.48 37.08 0.06
ZONE 4
Proven 63,233 12.04 0.39 40.47 0.16
Probable 175,637 11.42 0.58 55.29 0.13
Subtotal 238,870 11.58 0.53 51.37 0.14
95% Recovery 226,926 11.58 0.53 51.37 0.14
Pillar Zone 4 (21,595) 11.58 0.53 51.37 0.14
TOTAL ZONE 4 205,331 11.58 0.53 51.37 0.14
ZONE 97
Probable 2,313,709 11.52 1.02 59.46 0.09
95% Recovery 2,198,023 11.52 1.02 59.46 0.09
Pillar Zone 97 (78,760) 11.71 1.02 64.50 0.08
TOTAL ZONE 97 2,119,263 11.51 1.02 59.28 0.09
- -------------------------------------------------------------------------------------------------
TOTAL PROVEN & PROBABLE 3,322,760 10.78 0.82 52.13 0.09
- -------------------------------------------------------------------------------------------------
Note: The NSR values used to outline mineral reserves are based on US$0.50/lb
zinc US$0.80/lb copper, US$5.00/oz silver, US$343.00/oz gold and an
exchange rate of Cdn$1.43=US$1.00.for Zone 97 and US$0.55/lb
zinc,US$0.90/lb copper, US$5.00/oz silver, US$275.00/oz gold and an
exchange rate of Cdn$1.55=US$1.00 for Zones 3 and 4. Since the reserves
for the various zones have been calculated at different times, slightly
different parameters have been used to determine the NSR value of the
resources; however, SRK does not believe that these differences would
materially impact the current reserves.
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2.3.12 RECONCILIATION
Since 1995, a detailed reconciliation of the ore reserves and mill
production has been compiled on the mined-out stopes. The
reconciliation of the mining reserves with the mill shows an ore
dilution of 30% by volume and a mining ore recovery of 98.7% for 2000
(based on a combination of various stope heights in various zones).
The results of the historical reconciliation as summarized in Table
2-5, illustrates the variable dilution and mining recovery rates by
year as a result of the relative proportions of the different zones
that were mined during any given year and the variable stope heights.
Table 2-5 - Historical Reconciliation
----------------------------------------------------------
Year Dilution by volume Mining Recovery
----------------------------------------------------------
2000 30.0% 98.7%
1999 22.2% 91.0%
1998 22.1% 87.8%
1997 24.4% 82.5%
----------------------------------------------------------
2.4 EXPLORATION POTENTIAL
SRK believes there is excellent potential to add to the current
resource base, either in areas adjacent to the known resources, or
elsewhere on the property. These areas include:
o Between level 6 and 8, a small satellite lens located 200m
west of Zone 3 requires further drilling to justify
development. It currently contains an estimated resource of
100,000 tonnes grading 6.3% Zn and is open to the west.
o In 1998, exploration drilling intersected sub-economic
massive sulphides 300m west of Zone 97, between level 8 and
10. This sector is open down-plunge and should be drill
tested when underground development on level 13 is
completed.
o Langlois mine corridor along the western extensions of Zones
3, 4 and 5 that remains untested below a northeast trending
fault.
o Langlois mine corridor located east of the mine; a drill
target has already been proposed to test the depth extension
of Zone 97, as suggested by an incomplete off-hole Pulse-EM
response detected at the end of hole 91-GRM-134. Also, the
surface extension of the Zone 97 felsic horizon has received
limited drilling.
o Grevet B corridor - On the Grevet B property, a second
massive sulphide lens was intersected in 1997 during
delineation drilling on the main sulphide lens. This modest
discovery attests to the stacked nature of massive sulphide
lenses. Although lateral and depth extensions of both lenses
have been more or less closed out, the northeastern
extension of the Grevet B corridor remain relatively poorly
defined and thus warrants additional drilling.
o Orphee corridor (option to Metco) - New resource may be
identified along the Orphee corridor, mostly to the north
and east of the main Orphee lens where three distinct
sulphide-bearing felsic horizons have been intersected by
previous drilling.
o A large region located on the BP-Norex claim block, between
Grevet B and Orphee corridors.
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3.0 HISTORICAL OPERATING CHALLENGES
At the feasibility stage, the previous owners of Langlois appeared to
have had an optimistic view of orebody size, geometry and stability.
Consequently the initial mining method, design, procedures and
equipment were not appropriate to the context.
The ore is generally narrow and sinuous with weak, foliated walls.
Mining was initiated on the basis of 60m level intervals, 114mm
diameter blast holes, and primary/secondary stope sequencing. The mine
was unable to control the stability of the stopes, resulting in major
dilution and severe stope production delays.
The mine was never able to fully recover from this situation. As soon
as the mine did not deliver according to the plan, attempts were made
to lower costs, including starving the mine of capital. At the same
time, the previous owners were also experiencing financial
difficulties as well.
Various operating challenges have been experienced at the Langlois
mine since the initial start-up and over the years the mine operators
have made changes to solve some of them.
The following sections describe the operating challenges that
continued to adversely affect the production of metal to the mill
during the last few months of production, ending December 2000. These
are the challenges that have been addressed in the mine plan to allow
the operation to successfully resume production.
3.1 ORE PASSES
Historically, the ore passes at the Langlois mine have proven to be
unreliable with excessive wear rates, followed by plugging or
collapse. Many approaches have been taken to try to overcome these
matters including various designs, ground support methods and ore pass
operating strategies.
Operating the ore passes full is a common way to reduce wear, but at
the Langlois mine the broken ore is very sticky with poor flow
characteristics and the approach only lead to hang-ups. For this
reason, ore passes were operated at low to empty levels, which
increases the wear.
The ground support measures that have been tested in ore passes
include:
o Standard resin rebar bolts
o Flexible cable type resin bolts
o Very tight bolting patterns
o Cable bolts
o Regula shotcrete
o High resistance (Fondag) shotcrete
Regardless of the ground support, the ore passes have only been
reliable for approximately the first 100,000 tonnes of muck passed
through them following which their remaining life is variable and
uncertain.
The loss of use of ore passes has had the following effects on the
mining operation:
o Loss of production
o Resources assigned to non-productive work attempting to
clear blockages
o Excessive rehandling of ore by scoop and by truck. At times,
all or most of the daily production has been re-handled at
least once.
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o Significant dilution has been experienced from excessive ore
pass wear, thus compounding the dilution factor.
One of the existing ore passes that handles a large portion of mine
production is particularly important. It is a shallow angle ore pass
from level 10 to level 11 and it feeds one of the mine's two existing
loading pockets. It has experienced many hang up problems resulting in
production delays. This is a critical ore pass.
3.2 MINE CAPITAL DEVELOPMENT
The Langlois mine has been starved of capital funding for a number of
years. To a large degree, the result has been the curtailment of
capitalized waste development.
Development of the planned access drifts to Zone 97 were delayed, and
this high-grade zone was not readied for production on a timely basis.
To put this into perspective, consider that as of December 31, 2000:
o Zone 97 contains 60% of Langlois mineral reserves
o The Zone 97 diluted zinc grade was 25% higher than the rest
of the mineral reserves
Bringing Zone 97 into production will directly impact the grade to the
mill, and additional work places will become available thus improving
the mine's production potential.
3.3 MOBILE EQUIPMENT MAINTENANCE
Mechanical availabilities were low and equipment operating costs were
high at the time the Langlois mine was shut down. Frequent mechanical
breakdowns and equipment shortages hampered production work.
The underlying cause has principally been due to difficulty in
attracting and retaining well-qualified and experienced mechanics at
Langlois. This has been caused in part by the tight market for
mechanics, the uncertain past performance of the mine thus the lack of
job security and by wage competition from other resource sectors.
Also, a full time maintenance planner was not utilised in the past.
3.4 ORE STORAGE
With ore passes being operated at low levels and no underground ore
storage bins, the only significant storage has been on surface. The
capacity of the surface bin (2,000 tonnes live) has not been
sufficient to assist in smoothing out the highs and lows in
underground production.
3.5 STOPE DILUTION AND SUBLEVEL REHABILITATION
Stope dilution has been brought under control over the years through
changes in operating practices. Further improvements are possible by
designing more closely spaced sublevels with reductions in the
vertical exposure of open stope hangingwall (this approach has been
taken in zone 97).
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With the overhand stoping sequence at Langlois, stope wall control and
sublevel rehabilitation are at least partly related. Each successive
stope works off the backfill floor of the stope below. If waste layers
have peeled off the walls of the lower stope, then there will be more
rehabilitation to make the mucking drift of the stope above safe.
Stope cycle times typically have a significant component of
rehabilitation time. Improvements in stope ground control have the
potential to reduce dilution and rehabilitation, and speed up the
stope cycle time.
3.6 SRK MINE REVIEW
The feasibility study addresses the most significant issues causing
the uneconomic situation at the Langlois mine. SRK mining engineers
visited the mine in July 2000 and also during the start of the
feasibility process in January 2001 to evaluate the mining and
production systems and achievements. SRK needed to verify that the
current achievements were as stated and that the feasibility study
would properly address the issues, to produce a mine plan with a
better, more certain outcome.
Chris Page, Corporate Consultant, visited the mine from July 7 to 14,
2000 and prepared and issued a report documenting conditions at the
mine and the state of mine planning at the time. This report is
included on the reference list to the feasibility study (Appendix G -
C. Page Letter Report, July 2000.).
Ken Reipas, Principal Mining Engineer, visited the mine from January
10 to 18, 2001 and audited the underground production environment,
reviewed mine engineering procedures, and reviewed the background
information being prepared by the Langlois mine personnel.
Observations and discussions were recorded and filed for reference
(Appendix H - K. Reipas Observations and Discussions).
Also, during the period January to April 2001, SRK project personnel
made a number of visits to the mine for various purposes related to
the feasibility study.
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4.0 FEASIBILITY STUDY - MINING
4.1 MINING ROCK MECHANICS
4.1.1 ROCK MECHANICS REVIEW
Since the Langlois mine has already been in operation, there is less
uncertainty in the rock mass competency and behaviour compared to mine
designs that are based only on drill core data. Operational history
provides valuable information which overrides empirical analysis.
Based on the past mine operating experience and SRK's underground
geotechnical assessment, the following can be concluded.
4.1.1.1 ROCK MASS STRENGTH AND COMPETENCY
Intact rock strength (IRS) for both the host rock and ore is in excess
of 100 Mpa, except chloritic schist, locally developed at the contact
in Zones 4. IRS of the massive sulphide is in excess of 150 MPa.
Intact rock strength (not competency) is similar for the various rock
types in all three Zones (3, 4 and 97).
Two strongly developed open joint sets (dipping approximately 60(0))
together with foliation (sub-parallel to the ore body) are the main
"defects" influencing the rock mass competency. Well-developed
foliation is the main cause of strong anisotropy but unless disturbed
by blasting or induced stress, the foliation is not a "weak link".
Undisturbed foliation exhibits strong cohesion.
Ore (often with included layers of waste) is usually reasonably
competent with rock mass rating (RMR) values varying from 40 - 70 and
rock mass strength (RMS) of approximately 35 - 70 MPa.
The host rock is typically highly foliated with RMR values from 45 -
60 (Zone 3 and 4) and RMS of approximately 35 - 60 MPa. There are
locally developed areas (especially in Zone 4) where the contact wall
rock is very weak with RMR values as low as 30. This weaker contact
material cannot be kept in place and will always be taken as
over-break.
In general, Zone 97 is more competent than Zones 3 and 4. The rock
mass could be described as "fair to good" and suitable for open
stoping. The overriding property which will constrain the stope
dimensions is the foliation, rather than intact rock strength.
4.1.1.2 LARGE-SCALE WEAKENING STRUCTURES
There is only one major identified structure (Eastern Fault). Local
structures (outside the joint pattern) do exist and these will
influence individual stopes but cannot be allowed for in an overall
design. The dykes intersecting the rock mass do not appear to cause
major stability problems. Large-scale structures will be an
operational issue, not a design constrain.
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4.1.1.3 DISTURBANCE (STRESS)
There are no signs of unusual stresses. It is assumed that the
horizontal stress is approximately twice the overburden stress. The
overburden stress could vary from 10 - 20 MPa from the top of the
current ore zones to the bottom of Zone 97 respectively. These
stresses, together with a factor for induced stress, are high enough
to fail the rock mass in the vicinity of the stopes but they are not
unusually high. The estimated stresses are sufficient to account for
the deterioration seen in stope development and justify full
structural reinforcement of the in-ore development. Very little intact
rock failure was seen during underground visits.
4.1.2 ROCK MASS CLASSIFICATION
In order to design a mining operation, it is necessary to work with
numbers, therefore, all mining investigations require that the rock
mass be classified. The objective is to assess the rock mass
numerically by applying numbers to the various geological features,
which impact on the rock mass ability to resist the disturbing forces
of stress and gravity.
The Laubscher's mining rock mass (MRMR) classification is one of the
most versatile and widely used in the mining industry as it takes into
consideration both "nature-made" and "man-made" disturbance. For
example, blasting damage observed underground at the Langlois mine is
believed to be one of the key weakening "man-made" disturbances
resulting in excessive overbreak.
Although the MRMR classification is not very commonly used among open
stoping mines, it does provide reliable guidelines in terms of
critical dimensions related to large-scale rock mass stability. The
other method widely applied to open stope mining operations is the
modified "Q'" index. Both of these systems were used to assess the
rock mass competency, which was one of the primary parameter
determining the stope sizes.
In the context of the mining method used at the Langlois mine, the
rock mass in each zone (3, 4 and 97) was divided into 3 basic domains;
footwall, ore body and hanging wall. Based on the underground
observations, in most cases the footwall and hanging wall rock mass
competency is similar and the two zones were therefore grouped under
the term waste. In Zone 4 there are, however, locally developed zones
of weaker contacts (north wall) where decreased cohesion on the
foliation planes can cause separation and local instability,
especially if the blasting design is too aggressive. The detailed
location of such zones is very difficult to predict, so the mining
design had to take such constraints into consideration. Stope
dimensions were influenced by the weaker rock mass.
The details of the rock mass classification data are discussed in
Appendix J - Langlois Rock Mechanics Report and the summary results
used for the stope design are tabulated in Table 4-1 below.
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Table 4-1 - Rock Mass Classification Values for the Main Geotechnical Domains
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Geotechnical Minimum Maximum
Zone Domain ----------------------------------------
RMR Q RMR Q
Waste contact 30 0.2 41 0.7
Zones 3 & 4
Waste strong 43 0.9 57 4.2
Zone 97 Waste strong 49 1.7 66 11.5
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It has to be stressed that rock masses unfortunately do not conform to
an ideal pattern and therefore a certain amount of
judgment/interpretation based on experience was required. The
classification systems are used as guidelines rather than as a precise
engineering tool.
4.1.3 STOPE DIMENSIONING
The most widely used design method for the determination of open stope
dimensions in Canada is Mathew's Method. The method is based on the
modified "Q'" number and empirical stability chart. The details of the
assessment are included Appendix J - Langlois Rock Mechanics Report.
The range of "wall" Stability Numbers for the Langlois mine used in
the analysis is summarised in Table 4-2 below.
Table 4-2 - Stability Number Values for the Main Geotechnical Domains
- --------------------------------------------------------------------------------
Q' Factors Stability number N'
Domain ---------------------------------------------------------------
Min Max A B C Min Max
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Zones 3 and 4 0.9 4.2 0.7 0.8 7 3.5 16.5
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Zone 4 contact 0.2 0.7 0.7 0.8 7 0.8 3.0
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Zone 97 -
shallow 1.7 11.5 1.0 0.8 7 9.5 64.0
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Zone 97 -
deep 1.7 11.5 0.4 0.8 7 3.8 26.0
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The values derived depend on a large number of assumptions, which can
vary significantly. For example, the variability for a single rock
type can be 3.5 to 49 depending on whether Q' direct is used, whether
Q' is estimated from RMR values (which SRK prefers) or whether minimum
or maximum values are used.
Figure 4-1 is a stability chart illustrating various stope dimensions
for zones 3, 4 and 97. It will be noted that in the Mathew's chart,
the current stope dimensions plot at least in the "unsupported
transition" zone for most of the rock type ranges. But historical
results indicate that the stope walls have not performed as well as
would have been predicted by the Mathew's method (and if we had used
the conventional method for estimating stability, the values would
have been even higher). The initial stope size of 60m high by 20m on
strike had a "wall" HR = 7.5 and this plots
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mostly within the zones of "transition" or "support necessary" but
significant problems were experienced.
It must be recognised that empirical design charts are very
approximate. The "wall" Hydraulic Radius (HR) is the same for a 30m
high by 20m long stope (current dimension) compared to 14m high by
100m long stope (planned for Zone 97) but it is concluded that the
lower stope height (even though much longer) is significantly more
stable.
Figure 4-1 - Mathew's Stability Graph
[PICTURE]
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4.1.4 GROUND SUPPORT
4.1.4.1 ORE AND WASTE DRIFT SUPPORT
The backs of the ore and waste drifts are supported with 1.5m rebars
on a 1.2m x 1.2m pattern. Chain link (50mm x 50mm) screening is
installed on the back of the ore drifts, while 100mm x 100mm welded
mesh screen is installed on the back of the waste drifts. The chain
link and mesh are held up with 0.9m mechanical rockbolts. The walls
are supported with two rows of 1.83m straps held up with 1.5m
splitsets on a 1.2m x 1.2m pattern. This support standard was
established by empirical methods and has been improved over the years
with operating experience. No modification of the design is planned
for Zone 97.
Figure 4-2 - Typical Drift Ground Support
[PICTURE]
4.1.4.2 STOPE SUPPORT
In addition to the standard ore drift support, cable bolts were used
in Zones 3 and 4 consisting of six 5m long cables in a fan pattern
with a spacing of 2.4m between each row. In all zones, the
installation of 1.5m rebars will replace cable bolts. It is a less
expensive way to achieve support. It is also a stiffer support which
will prevent early movements which usually resulted in dilution and
extra rehabilitation work. The stope support design is mainly governed
by experience and some improvements can be made in the future.
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4.1.5 SHAFT PILLAR MINING
The shaft at Langlois mine is located in the country rock
approximately 100m away from the steeply dipping orebody. Since the
shaft does not intersect the orebody, a shaft pillar may only be
required if mining related movements or induced stresses are expected
to result in instability in the shaft.
Based on a previous geotechnical assessment by C. Doucet, (Appendix K
- Services Techniques, April 2000), a 100m wide pillar of ore opposite
the shaft had been designated as a shaft pillar, to protect the shaft
from anticipated adverse effects associated with stoping. The shaft
pillar was located in an area where the ore grade was above the mine
average.
As part of the feasibility study, the necessity of leaving a shaft
pillar was re-assessed by SRK Consulting. Due to the higher grade of
the shaft pillar ore, and its distance from the shaft, it was decided
that further analysis work was justified.
The objectives of the analyses were to determine whether the shaft
pillar was required and whether any instability would occur in the
shaft if the designed 100m wide shaft pillar, and other portions of
the orebody were mined out.
The potential stress changes and deformations in the shaft resulting
from the extraction of the shaft pillar and the surrounding ground
were assessed by conducting a series of numerical stress analyses
using a tabular excavation modeling program (Minsim-D, COMRO, 1993).
To assess the potential for local damage in the shaft, back analysis
methods as well as empirical relationships between induced
stresses/strains and excavation damage were considered.
SRK's analysis shows that the shaft will not be adversely affected by
the proposed mining. All criteria indicate that stresses and strain
values will be well below those that have caused instability problems
in mine shafts in the past. The only concern is that some loosening of
the rock in the shaft walls may occur as the stresses decrease. If
loosening has not been a problem in the past, it is unlikely to become
a problem with further mining. The details of this analysis are
included in Appendix E-Assessment of Effects of Extracting Shaft
Pillar.
4.1.6 CURRENT GROUND CONDITIONS
In May 2003, Breakwater mine site staff conducted an underground
inspection of all main levels and sublevels from 9 level to 7 level.
In general, the mine workings have shown little sign of deterioration
since operations were suspended at the end of 2000. Specific
observations reported included:
o No rock falls observed.
o No major deterioration of any access ramps or levels.
o An open stope on 9 level, left open since late 2000, had
experienced some wall caving. (This is not unexpected for an
open stope left so long without backfilling.)
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4.2 MINING METHODS
4.2.1 GENERAL DESCRIPTION
The original method of mining applied in Zone 3 in 1996 (Refer to
Figure 4-3) consisted of mining long hole stopes 60m high by 20m in
strike length by the full ore width which averaged about 4 to 5 m. A
haulage drift was driven in waste paralleling the ore, which provided
a drawpoint for each stope. With this independent access to each
stope, primary and secondary stope sequencing was used. Production
drill holes of 114mm (4 1/2") diameter were used, averaging 60m in
length.
During 1997, this method was abandoned due to excessive dilution
brought on by deviation of the blast holes in the narrow ore zone, the
undulating nature of the ore zone, and the weak to moderate strength
of the foliated host rock. After this experience, the stope heights
were reduced to 30m by driving additional sublevels, and a retreating
stope sequence was employed to save on waste development and to avoid
the problems of mining the pillars created by stope and pillar type
sequencing.
Further improvements in controlling dilution in Zone 3 were made as
mining progressed. In the narrower parts of the ore zone, the drill
hole diameter was reduced to 64mm (2 1/2") and the stope height was
reduced to 20m.
In 1998, stoping began in Zone 4, where the hanging wall rock has a
more intense layer of foliation next to the contact. Here, longhole
stopes were mined with dimensions of 15m high by 20m strike length,
using a retreating sequence.
All of the historic mining described above included trackless
development and ramp access to stoping areas, production mucking with
3.5yd and 4yd scooptrams to ore passes, and cemented paste backfilling
of the stopes.
Ore pass failure was experienced throughout the mining history.
The planned mining areas and methods incorporated in the feasibility
study are introduced in the following paragraphs, while the full
details are presented in the following sections of this report: 4.2.2
Definition Drilling, 4.2.3 Development, 4.2.4 Slot Raising, 4.2.5
Production Drilling, 4.2.6 Blasting, 4.2.7 Production Mucking, and
4.2.8 Backfilling. Within most of these sections, details are provided
for each ore zone.
4.2.1.1 ZONE 3
Where stope development already exists, longhole stope dimensions of
30m in height, and 164mm (4 1/2") diameter production drilling will be
maintained. For the portions of Zone 3 that are not yet developed,
improvements are planned to control dilution.
o Sublevel spacing will be reduced to 20m.
o Production drilling will employ 64mm (2 1/2") diameter
holes.
All stopes will be mined in a retreating sequence and will be
accessible by ramp. Production mucking will be to ore passes. Back up
ore passes will be provided and critical sections will be steel lined.
Stopes will be filled with cemented paste backfill.
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4.2.1.2 ZONE 4
Zone 4 longhole stope dimensions will remain the same as the previous
method. Ramp access will be provided to the sublevels. The stopes will
be filled with cemented rockfill as they are located in an area of the
mine that cannot be serviced by the paste backfill system.
4.2.1.3 ZONE 97
Major improvements in future mine performance (tonnage rate and grade)
will be achieved by bringing Zone 97 on line once production resumes.
64% of future production will come from this zone. Zone 97 is a
higher-grade area that was not previously brought into production due
to the lack of funding by the previous owners for the capital
development required to access it. Refer to Figure 4-3. The
feasibility study mining plan provides for pre-development of several
sublevels in Zone 97 so that it can contribute continuously to
production when mine operations resume. Significantly, Zone 97 will
yield higher-grade tonnes than the past mine average, and will provide
additional working places that make the overall mine production rate
achievable.
An overhand benching method of stoping will be applied in Zone 97.
Stope dimensions are planned at 14m open vertical span and, on
average, 45 to 50m in strike length. The average ore width is
approximately 3m. Stope sequencing will be in a retreating fashion
towards the central access cross cut provided at each sublevel. Where
the length of a sublevel is more than about 70m from its extremity to
the central access cross cut, it will be divided into two stopes, with
the second stope mined after the backfilling of the first.
The reduced stope height is planned in order to control dilution while
allowing long, yet stable, stope dimensions. The long strike length
means that the unit mining operations such as drilling can be more
continuous, with less moving between work places. Ramp access is
planned, as is backfilling with cemented paste fill that will be
pumped to the zone along 6 level. The use of a pump is required due to
the large distance from the shaft to Zone 97.
There are no ore passes planned for Zone 97 due to their unreliable
nature. A fleet of new 20 tonne trucks will haul ore from stoping
areas to the shaft area, and the past problem of collapsing ore passes
will be avoided. (Truck selection is discussed in section 4.8.2) Ore
and waste will be hauled on 9 level and 13 level and will be dumped
into steel lined ore/waste passes near the shaft.
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Figure 4-3 - Langlois Mine Longitudinal Projection
[PICTURE]
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4.2.2 DEFINITION DRILLING
During previous operations, definition diamond drilling has been
carried out on a relatively dense pattern to determine the final
mining reserve. In order to determine the density of required
drilling, a statistical study was carried out in 1994 during Cambior's
original preliminary feasibility study.
For areas of Zone 3 with 30m sublevels, the density of drilling was
10m horizontally by 15m vertically. For the areas of Zones 3 and 4
with 15m sublevels, the density of drilling was spaced wider at 20m
horizontally by approximately 18m vertically. In Zone 97, the
definition drilling density was spaced at 20m horizontally by 20m
vertically.
Pneumatic diamond drills drilling AQ-ATW or 27mm diameter core were
used for hole lengths less than 120m. Electric diamond drills, (75 to
100 HP) drilling BQ or 36.5mm diameter core, were required for holes
in excess of 120m.
All holes were surveyed once complete and acid dip tests were carried
out every 30m. For BQ holes in excess of 120m, strike and dip
measurements were taken with a Pajari type instrument.
The core was logged for RQD and rate of fracturing measurements were
taken on the mineral-bearing zones and their immediate strata. Drill
core was preserved as reference on an approximate pattern of 100m by
50m.
The feasibility study provides for the continuation of these past
definition drilling practices.
4.2.3 DEVELOPMENT
Future mining at the Langlois mine is planned in three areas that are
individually provided with ramp access. These three areas are not,
however, interconnected by ramp but are essentially separate for
mobile equipment planning purposes. Each of these areas is accessed by
main levels from the shaft. The three areas are:
o Levels 4, 5, and 6 including mining of Zone 4 and upper Zone
3.
o Level 9 including mining of lower Zone 3 and upper Zone 97.
o Level 13 including mining of lower Zone 97.
Refer to Figures 4-4, 4-10, 4-12, 4-13.
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Figure 4-4 - Ore Zones and Main Levels
[PICTURE]
4.2.3.1 ZONE 4
Zone 4, is currently accessed by a 3.2m high x 3.7m wide 15% gradient
access ramp in waste. Crosscuts from the ramp to the ore zone average
20m in length. Existing lower level drifts in ore are sized at 4.0m
high by ore width (2.2m minimum). Existing higher-level ore drifts are
dimensioned smaller at 2.7m high by ore width because they were driven
as part of a captive mining plan, which has since been abandoned in
favour of ramp access.
The ramp access will be advanced to service the planned stoping
levels, and future ore drifts will be driven at 3.2m high by ore width
with a minimum width of 2.2m. Levels include access to the waste fill
raise since cemented rock fill is planned in this zone. Planned
development equipment includes a single boom drill jumbo and a 3.5yd
scooptram. The layout of a typical access ramp sublevel is shown in
Figure 4-5.
Figure 4-5 - Typical Access Ramp and Sublevel Layout of Zone 4
[PICTURE]
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4.2.3.2 ZONE 3
The access concept for Zone 3 is similar to Zone 4. A ramp in waste
provides access to the sublevels. Refer to Figure 4-6. Ore drifts will
be driven by similar equipment, silled out to ore width at 3.2m high.
The minimum mining (stoping) width in Zone 3 is 3.0m. The retreating
mining sequence does not require a waste drive parallel to the ore
zone.
Figure 4-6 - Typical Level in Zone 3
[PICTURE]
4.2.3.3 ZONE 97
An existing drift on level 9 currently accesses Zone 97. This drift
has nominal dimensions of 2.9m high x 2.9m wide and will be slashed
out to 3.5m high x 3.7m wide to accommodate 20 tonne haulage trucks. A
second ore haulage drift will be driven on level 13 from the shaft
area to the ore zone. A third drift will be driven to Zone 97 on level
6. This drift will provide access for backfill distribution, assist
with ventilation, and provide a drilling platform for exploration and
definition drilling.
Within the overall Zone 97 mining area, six 14% gradient ramps will
provide access to six mining domains. Refer to Figures 4-7 and 4-13.
The ramps will be driven 3.5m high by 3.7m wide. This will provide
1.5m of width clearance for a 20 tonne truck. Safety bays are planned,
and the resulting configuration is in compliance with the "Regulation
respecting Occupational Health and Safety in Mines", (Quebec),
Regulations 43 and 44. The ramps are located approximately 40m away
from the ore zone for long term stability.
In the ore, sublevels at a vertical spacing of 11m (floor to floor)
will be accessed from the ramp by cross cuts. Each cross cut includes
a remuck bay suitable for loading trucks. The retreat mining sequence
does not require a waste drive parallel to the ore zone.
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Figure 4-7 - Zone 97 Sublevel Access to One Mining Domain
[PICTURE]
4.2.4 SLOT RAISE
The stope slot opening methods that are planned for the Langlois mine
are based on proven methods, and they are described below for each ore
zone.
4.2.4.1 ZONE 4 SLOT RAISES
The main method of opening stope slots in Zone 4 will be to blast
against 15m long styrofoam columns placed into the cemented waste
rockfill of the previously mined stope. Refer to Figure 4-8. This is a
method that has been used successfully in the past at the Langlois
mine in areas of retreating stope sequence. The styrofoam blocks are
anchored in place on cables prior to backfilling with cemented rock
fill. After backfilling, the production long hole drill is used to
drill two reamed holes in ore immediately in front of the styrofoam
column. These holes are blasted against the soft styrofoam to open the
slot.
This method cannot be used for the primary opening on a sublevel. In
these cases a conventional raise or drop raise will be used.
Figure 4-8 - Zone 4 Styrofoam Column Slot
[PICTURE]
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4.2.4.2 ZONE 3 SLOT RAISES
The main method of opening stope slots in Zone 3 will be to blast into
20m or 30m long raises reamed in the paste backfill of the previously
mined stope. Refer to Figure 4-9. This method has been used
successfully in the past at the Langlois mine in areas of retreating
stope sequence. Once backfilling is completed, the production ITH
drill places a 114mm (4 1/2") diameter hole through the backfill to
the sublevel below. A simple 762mm (30") diameter reaming head is
attached and pulled back up in a manner similar to a raise boring
operation. The production long hole drill is then used to drill two
114mm (4 1/2") diameter holes in ore immediately in front of the small
raise in the backfill. These holes are blasted into the raise to open
the slot
This method cannot be used for the primary opening on a sublevel. In
these cases a conventional raise or drop raise will be used.
Figure 4-9 - Zone 3 Reamed Raise in Paste Backfill
[PICTURE]
4.2.4.3 ZONE 97 SLOT RAISES
All of the initial stope slots on Zone 97 sublevels will be opened by
using 1.8m x 1.8m drop raises. These raises will be very short,
averaging 8m in length. They will be drilled with 54mm (2 1/8") holes
by the production drills. Zone 97 employs a benching method that opens
45 to 50m long stopes prior to backfilling.
If a sublevel is more than 70m in strike length (from its center cross
cut to its extremity), it will be divided into two retreating stopes.
The slot raise for the first stope will be a drop raise positioned at
the extremity of the sublevel. After backfilling the first stope, the
second stope will be re-slotted by blasting against the paste backfill
using styrofoam slots (similar to Zone 4 practice previously
described).
4.2.5 PRODUCTION DRILLING
Drill patterns in Zones 3 and 4 are based on previous experience. Zone
97 will employ a similar type of pattern. Although it is recognized
that drill spacing is best maintained greater than burden,
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compromises have been made to accommodate variable, narrow ore widths.
In the three hole patterns, the centre hole is shifted towards the
free face to reduce the effective burden. The planned drilling
patterns are described for each ore zone in the following paragraphs.
Drill patterns are included in Appendix M - Blasting Sketch.
4.2.5.1 ZONE 4 PRODUCTION DRILLING
In Zone 4, tire-mounted, top hammer pneumatic long hole drills
(existing equipment) will drill 54mm (2 1/8") diameter holes
approximately 15m in length. The burden is fixed at 1.2m and the
nominal spacing of 1.0m varies according to the width of the ore.
There are three holes on each ring of drilling. The drilling factor
averages about 6.0 tonnes (diluted) per metre of drilling.
4.2.5.2 ZONE 3 PRODUCTION DRILLING
In Zone 3, crawler-mounted, ITH drills will drill 114mm (4 1/2")
diameter holes of approximately 30m in length. The burden and spacing
are variable according to the width of the ore. The burden ranges from
1.4m to a 2.1m maximum and the spacing is a nominal 1.0m. In areas of
wider ore, 4 holes are placed on each drill line. The centre holes are
shifted by 30% of the burden, towards the free face. Casings of 75mm
(3") diameter are inserted in all the holes to prevent hole blockages
and to reduce stope wall blast damage. The drilling factor averages
about 9.5 tonnes (diluted) per metre of drilling.
4.2.5.3 ZONE 97 PRODUCTION DRILLING
In Zone 97, tire mounted, top hammer pneumatic long hole drills will
drill 54mm (2 1/8") diameter holes of approximately 8m in length. The
burden is fixed at 1.2m and the nominal spacing of 0.7m varies
according to the width of the ore. There are three or more holes on
each line of drilling, depending on the ore width. The centre holes
are shifted by 20% of the burden, towards the free face. The drilling
factor averages about 5.0 tonnes (diluted) per metre of drilling.
Standard practice in the past has been to survey all drill hole
breakthrough points and assess the deviation against established
criteria. Any holes with excessive deviation were re-drilled before
the production drill left the area. This practice will be continued.
4.2.6 BLASTING
The blasting practices described below are largely based on the
practices developed at the Langlois mine while it was in production.
Anfo has been the most commonly used blasting agent in the past and
this is also planned for future blasting. Cartridges of emulsion will
be used when wet conditions are encountered, or when it is necessary
to pre-load rings in the "next" blast. Anfo will be pneumatically
loaded into the 54mm (2 1/8") diameter holes and gravity loaded into
the 75mm (3") casings in the 114mm (4 1/2") diameter holes. Design
powder factors vary depending on the ore width, but when loading with
anfo, the powder factor will be in the range of 0.4 to 0.5 kg per
tonne. Low-density anfo will be used at times in the outside holes to
reduce blast damage.
All holes will be double primed to reduce the risk of blast failure in
this narrow blasting situation, with normally only three holes per
row. Two non-electric detonators with suitable primers are planned in
each hole. Tie in of holes will use detonating cord, initiated by
electric caps.
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When blasting 54mm (2 1/8") diameter holes in Zone 4, 4 rows will be
blasted at one time, while 10 rows in Zone 97 will form one blast.
These blasts in Zone 97 will have adequate relief considering the
short 8m holes with free faces above and below (drifts) in addition to
stope void.
In Zone 97 the stopes will be approximately 45 to 50m long, and the
blasting will be done in stages of 10 rows as noted above. Blast
simulations have been performed to estimate the throw distance of the
muckpile. The simulations indicate that normal remote mucking
techniques will recover the blasted ore. This is not expected to be an
operational concern. However, if there are excessive mucking
distances, teleremote mucking can be applied. It is being used
successfully at other mines to improve productivity and safety.
Results of the blast simulation are included in Appendix L - Blasting
Simulation
During past operations, the Langlois engineering department developed
standards for blast hole loading. The blasters were trained to follow
these standards, and copies were available at the blast site. Blasting
instructions were issued for each blast, providing detailed
instructions and referring to the standards. These practices will be
continued to support future blasting operations. Typical examples of
these hole-loading standards are included in Appendix M - Blasting
Sketch.
4.2.7 PRODUCTION MUCKING
All production mucking will be performed by 3.5 and 4.0 yd scooptrams
equipped with remote controls. Retreat mining is planned in all areas
so each stope will be mucked from one draw point oriented along
strike. The scooptram operator will use remote control mucking when
the bucket of the scooptram passes the brow of the draw point. To
provide protection for the operator in these situations, a remote
operator bay, or an elevated concrete stand will be provided.
In Zones 3 and 4, mucking will be to an ore pass with a grizzly. The
estimated average one way mucking distances are 170m in Zone 3 and 90m
in Zone 4. The upper portion of Zone 3 Inf. East and Inf. Centre will
be trucked to 8 Level.
In Zone 97, mucking will be to a remuck bay located in the central
crosscut from the access ramp. The average one way mucking distance
from drawpoint to the remuck is estimated at 100m. The production
scooptram will load trucks with ore at the remuck bay. The extra time
required for truck loading has been accounted for in the planned
mucking productivity.
4.2.8 BACKFILLING
Three types of backfill are planned for the Langlois mine. Paste
backfill, which will contribute 83% of the total backfill, has been
well proven at the Langlois mine during past operations. Cemented
rock-fill (5% of total) will be used in Zone 4 because paste backfill
cannot be delivered by gravity flow to this zone, and a pump
installation cannot be economically justified in this case. Finally,
un-cemented waste rock fill (12% of total) will be use in conjunction
with paste backfill for fence construction and for creating a suitable
mucking floor on top of the paste fill. Stopes mined early in the
schedule are filled with waste in Zone 97.
Paste backfill will be delivered underground through two existing
operational lines connecting the surface backfill plant with
underground levels. The placement rate of the paste is approximately
70 tonnes/hr. See Figure 4-10 for the paste backfill distribution
network.
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Figure 4-10 - Paste Backfill Distribution Network.
[PICTURE]
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All of Zone 3 can be serviced by gravity flow, however a backfill pump
is required for paste delivery to Zone 97. This pump will be installed
on level 6, in the Zone 97 access drift. The backfill will be
distributed horizontally in a 203mm pipe on level 6, and then
vertically through drill holes to Zone 97. The Langlois mine
engineering department has previously prepared a separate report on
this subject. See Appendix N - Paste Fill Report.
Figure 4-11 schematically illustrates the planned placement of paste
backfill in Zone 3 and Zone 97.
Figure 4-11 - Paste Backfill Placement
[PICTURE]
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Table 4-3 provides information on the paste backfill logistics for
Zone 3 and Zone 97.
Table 4-3 - Paste Backfill Logistics
----------------------------------------------------------------------------------------------
ZONE 3 ZONE 97
----------------------------------------------------------------------------------------------
Pipe distribution 150mm pipe in drift Same
----------------------------------------------------------------------------------------------
Fence 250t of waste placed by scooptram from the upper
drift and covered with a 75mm thick shotcrete layer.
----------------------------------------------------------------------------------------------
Floor A minimum of 600mm layer of rock-fill to allow
scooptram travelling on paste backfill. The waste is
placed by scooptram.
----------------------------------------------------------------------------------------------
Cement content
Plug 6.5% cement N/a
Residual 3.0% cement 3.0% cement
Average 4.3% cement 3.0% cement
----------------------------------------------------------------------------------------------
Placement rate:
(for same tonnage stope)
----------------------------------------------------------------------------------------------
Fence and curing 5 days 5 days
----------------------------------------------------------------------------------------------
Plug (with curing) 5 days Not required.
----------------------------------------------------------------------------------------------
Residual (with curing) 7 days 8 days
----------------------------------------------------------------------------------------------
Total 17 days 13 days
----------------------------------------------------------------------------------------------
----------------------------------------------------------------------------------------------
On a 5 day schedule 4 weeks 3 weeks
----------------------------------------------------------------------------------------------
Cemented rock fill is planned for Zone 4. An existing surface
stockpile of previously hoisted mine development waste will be used to
supply the underground requirements. The waste stockpile is very close
to the backfill delivery raise, and waste will be trucked and dumped
as required. Underground, the waste rock will be transported by 3.5 yd
scooptram to the stope.
Cement slurry batches of 3.0 cubic metres will be prepared on surface
using the existing paste backfill plant facilities. The cement slurry
will be pumped a short distance on surface through a new delivery
system, to a drill hole. The drill hole will deliver the slurry to an
underground slurry plant consisting mainly of an agitated holding tank
and a sump for flush/washdown water. A piping system will deliver the
slurry to the stope being filled. Every time a bucket of waste is
dumped into the stope, a metered amount of cement slurry will also be
sprayed into the stope. Other mines, such as Barrick's Holt McDermott,
have been very successful with this placement method. Five percent
cement by weight is planned to achieve the necessary backfill
strength.
Un-cemented waste rock fill generated by development will be used
underground as much as possible to avoid the cost of hoisting and
future disposal. This practice also serves to reduce the required
quantity of higher cost paste backfill. The use of un-cemented waste
fill is constrained by two factors. Firstly, this fill has no strength
and can only be used in certain parts of the stope. Secondly, in some
mining domains there is little extra time available in the stope cycle
to wait for waste rock filling that is often slower than paste
filling.
Waste can easily be mucked from the development face to a remuck bay
near the stope to be filled by truck and/or scooptram and then it can
be placed by scooptram into the stope. Allowances have been made in
the estimated equipment operating hours for this function.
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4.3 STOPE SEQUENCING
Historically at the Langlois mine production planning has been done on
the basis of "domains", and this same approach has been used in the
feasibility study.
Development and production tonnes are planned and reported on the
basis of domains. Each domain is an area, or sub-block of the mine,
encompassing many individual stopes and the corresponding portion of
the minable reserves. Refer to Figure 4-12.
Within each domain the individual stopes are mined in an overhand
retreating sequence, retreating from the domain extremities towards
the access point, and advancing up-dip, working on top of backfill,
not under it. In general, no pillars are planned along strike between
stopes, as each successive stope is mined up against the backfill of
the previously mined stope. This method of sequencing has been
extensively and successfully used at the Langlois mine in the past.
Retreat sequencing allows development in waste to be minimized, and
avoids the problem of mining the pillars created by stope and pillar
type sequencing.
Some sill pillars are planned between domains as retreat mining
progresses. They are located at levels 9, 10, and 13 as shown in
Figure 4-13. These pillars are the necessary result of opening up
multiple, independent mining domains to support the planned production
rate. These sill pillars have been accounted for in the reserving
process by the application of a reduced mining recovery.
The timing of the mining of any individual stope depends on the mining
activities in the stopes above, beside and below it. Individual stopes
are not independent, but rather, dependent on the completion of mining
in the neighbouring stopes. Thus, the stope mining cycle (days to mine
and fill) becomes a key factor in how fast stopes can be mined when
they are inter-related in a retreating sequence.
Typical stope cycle for Zone 3 or Zone 4:
Production drilling :12 working days
Blasting and mucking :13 working days
Backfilling :17 working days
Rehabilitation :20 working days
------------------------------------------------
Total cycle time :62 working days
Each year has 250 working days. In these domains, with an access
located centrally to the domain, a maximum of 12 stopes per year can
be mined with three sublevels developed in advance. This is possible
because there can be overlaps in stope cycles for stopes not
immediately adjacent to each other. Figure 4-14 shows the sequencing
to achieve the maximum of 12 stopes per year in one domain with
central access.
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[PICTURE]
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[PICTURE]
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Figure 4-14 - Maximum Annual Stope Sequencing in Zone 3 or Zone 4
[PICTURE]
Generally, in Zone 97, 8 to 10 stopes per year (maximum 12) were
scheduled in each domain. This is possible assuming that development
and production drilling are carried out in advance. An allowance has
been kept in the schedule in case re-slotting a stope is required for
ground considerations.
Typical stope cycle for Zone 97 (Typical Size = 5,000t):
Blasting and mucking :15 days
Back fill :15 days
Rehabilitation : 8 days
----------------------------------------
Total cycle time :38 days
The rehabilitation time has been reduced as it is anticipated that
there will be less damage to the stope walls and backs due to the
reduced stope height.
Figure 4-15 shows the sequencing to achieve the maximum of 12 stopes
per year in one Zone 97 domain with central access.
Figure 4-15 - Maximum Annual Stope Sequencing in Zone 97
[PICTURE]
4.4 Production Rate
Three mining alternatives were studied at the prefeasibility stage,
including two different production rates. All three alternatives were
based on operating the mine five days per week, with two 8-hour shifts
per day. The alternatives were:
o (1) 450,000 tonnes/year using Dec. 31/01 reserves
o (2) 600,000 tonnes/year using Dec. 31/01 reserves
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o (3) 450,000 tonnes/year using the higher grade portion of
the reserves
Pre-feasibility level cash flow models were prepared for these
Alternatives. The 450,000 tonnes/year high-grade alternative (3) was
selected. In economic terms, Alternative (3) was slightly better than
Alternative (2).
Other factors that favoured the selection of Alternative (3) were:
o The high-grade mining plan does not isolate the lower grade
reserves, which are currently excluded from the mining
schedule. They can be brought into production if metals
prices increase sufficiently.
o Alternative (3) has lower risk capital than Alternative (2).
o The certainty of achieving planned production would be much
less if planning for 600,000 tonnes/year. A higher
production rate is more difficult to manage, particularly
with small stopes cycling quickly.
The selection of the 5-2 operating schedule was based partly on
previous Langlois mine operating experience and on other factors:
o When operating with three 8-hour shifts per day in the past,
significant delays were encountered waiting for blasting
gases to clear.
o When working three 8-hour shifts per day, it is not possible
to have all men below collar for the full 8-hour shift, due
to shift change information exchange and multiple cage runs.
o Shift-to-shift communications are far more effective without
a third crew.
o More hours worked per week can only be achieved by adding a
third crew. This means three relatively smaller crews, and
supervision costs increase.
o An analysis of more work hours per week theoretically does
allow an increase in equipment utilisation and possible
reductions in fleet size. At the Langlois mine there are
essentially three captive areas with three relatively small
fleets. Only very modest reductions are possible in the
equipment fleets.
o Past experience was that development crews did not benefit
from 10-hour shifts.
The selected production rate of 450,000 tonnes per year is achieved by
the sum of the contributions from each mining domain. In this
feasibility study mining plan, the production scheduled from any one
domain depends on the following factors:
o The maximum possible rate of cycling stopes as described in
section 4.3 "Stope Sequencing".
o The timing of the start of domain production depends on when
sufficient pre-development is completed to sustain
production.
o The grade of the mineral reserves in the domain. Reserves in
the lower grade domains (such as Zone 3 Inferior West, and
Zone 3 Inferior East Extension) are only partially extracted
at a lower mining rate in order to achieve the higher
overall grade of alternative
(3). Refer to Figure 4-13.
The detailed production schedule, broken down by domain and by year,
with all metal grades, is included in Appendix O - Long Term.
The mining plan to support 450,000 tonnes/year of higher-grade
reserves has been prepared with input from the mining engineers and
mine operators experienced in managing the past production at the
Langlois mine. SRK has audited the mining plan during its development,
and it is SRK's opinion the plan has been prepared on an achievable
basis. The Langlois mine has detailed historical mining records
available that have been used to advantage during this process.
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Table 4-4 shows how the feasibility study compares to the historical
achievements at the Langlois mine.
Table 4-4 - Production Statistics
HISTORICAL DATA
Year Working Production Development Total Grade Total
Days Tonnes Tonnes Tonnes % Zn Tonnes
Per Day
1996 304 423,826 113,408 537,234 6.77 1,767
1997 140* 167,387 93,681 261,068 6.36 1,865
1998 250 385,241 29,501 414,741 6.53 1,659
1999 257 331,684 70,540 402,224 7.51 1,565
2000 212* 261,152 43,517 304,669 7.92 1,437
Total 1,163 1,569,290 350,647 1,919,937 7.00 1,651
Average per Year 233 313,858 70,129 383,987
FEASIBILITY STUDY
Year 2 - - 73,879 73,879 11.29 -
Year 3 250 288,321 92,724 381,045 10.51 1,524
Year 4 250 376,963 73,037 450,000 10.78 1,800
Year 5 250 359,719 90,281 450,000 10.93 1,800
Year 6 250 340,976 109,024 450,000 10.97 1,800
Year 7 250 341,174 108,826 450,000 10.96 1,800
Year 8 250 341,041 108,959 450,000 10.73 1,800
Year 9 250 401,245 48,755 450,000 10.51 1,800
Year 10 104 167,727 110 167,837 10.59 1,614
Total 1,854 2,617,165 705,595 3,322,760 10.78 1,792
Average per Year (9 years) 290,796 78,400 369,196
- -------------------------------------------------------------------------------------------
* Shutdown part of year
The planned average zinc head grade of 10.78% is significantly higher
than the head grades achieved during the last four years of production
ranging from 6.4 to 7.9%. The planned zinc grade is higher because:
o Zone 97 high grade tonnes will become part of the production
stream for the first time.
o The feasibility plan incorporates a higher cut-off grade
than past mining.
o The feasibility plan includes mining of the higher grade
shaft pillar which was previously considered sterilized.
This report describes the rock mechanics work done to
support this decision. Refer to section 4.1.5
The mining plan incorporates the following improvements to ensure
reliability of production:
o Zone 97 will be brought into production providing higher
grade ore as well as additional work places. Refer to Table
4-3.
o Zone 97 will use truck haulage to avoid past problems with
ore pass wear and failure.
o In Zone 3 and Zone 4, alternate back up ore passes are
provided. Critical sections of ore passes will be steel
lined.
o A new steel lined storage bin is provided to overcome the
previous difficulties at the level 10 to level 11 ore pass
and feeder.
o Improvements are planned to the mine maintenance program.
o Zone 97 employs a reduced stoping height to control dilution
and avoid the associated delays.
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These improvements are detailed in other sections of this report.
Table 4-5 - Tonnes Mined By Zone
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DEVELOPMENT & PRODUCTION TONNES
- --------------------------------------------------------------------------------
Category Tonnes Zn % Cu % Ag g/t Au g/t
Zone 3 Development 51,868 9.31 0.46 37.89 0.05
Production 946,298 9.05 0.48 37.04 0.06
SUBTOTAL 998,166 9.06 0.48 37.08 0.06
Zone 4 Development 19,503 11.58 0.53 51.37 0.14
Production 185,828 11.58 0.53 51.37 0.14
SUBTOTAL 205,331 11.58 0.53 51.37 0.14
Zone 97 Development 634,223 11.56 1.01 59.40 0.09
Production 1,485,040 11.49 1.03 59.24 0.09
SUBTOTAL 2,119,263 11.51 1.02 59.29 0.09
All Zones Development 705,595 11.40 0.96 57.60 0.09
Production 2,617,165 10.61 0.80 50.65 0.08
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TOTAL 3,322,760 10.78 0.83 52.13 0.08
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4.5 MOBILE EQUIPMENT
Breakwater personnel have visited other operating mines in Quebec,
Ontario and Manitoba in order to assess the possible improvements to
the Langlois mine maintenance program with regards to mobile
equipment. The findings of the visits can be found in Appendix P -
Maintenance Program and Mine Visit Summary
Several improvements are planned for the Langlois mine underground
mobile equipment fleet to increase mechanical availability and reduce
equipment operating costs. Improvements include:
o Refurbishing of the existing mobile equipment prior to going
back into service.
o Several new units will be purchased to meet mine plan
requirements.
o Many improvements are planned to the mobile maintenance
program.
o Two graders will be purchased, and a small crusher for road
material.
o A new underground garage on level 13 and improvements to the
existing level 9 garage.
Table 4-6 provides a list of the existing underground mobile fleet and
shows the allowances for refurbishing this equipment, which is
currently parked in or near the existing garage on level 9. The
equipment was properly prepared for long term storage at the time of
mine shutdown in December 2000. One of the production scooptrams is
scheduled for early replacement and is therefore not being
refurbished.
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Table 4-6 - Existing Underground Mobile Equipment
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Refurbishing Total
Existing Allowance Refurbishing
Equipment Description Units Per Unit ($) Allowance ($)
- ----------------------------------------------------------------------------------------------------
Underground Truck 13 tonne 1 100,000 100,000
Development Jumbo Single boom, electric/hydraulic 2 75,000 150,000
Scissor Lift Dux 3 50,000 150,000
Production Scoop* 4.0 cuyd capacity 4 75,000 225,000
Production Scoop 3.5 cuyd capacity 1 75,000 75,000
Development Scoop 3.0 cuyd capacity 4 75,000 300,000
Services Scoop 2.0 cuyd capacity 5 50,000 250,000
ITH Drill, Crawler CMS, 114mm, diesel 1 75,000 75,000
ITH Drill, Rubber Tire McClean, 114mm, electric 2 75,000 150,000
Pneumatic Top Hammer Drill BCI, 54mm 2 50,000 100,000
Service Vehicle 2-Tractor, 2-Forklift 4 25,000 100,000
- ----------------------------------------------------------------------------------------------------
TOTAL 29 1,675,000
- ----------------------------------------------------------------------------------------------------
* Only three units will be refurbished
To meet the requirements of the mine plan, several new units are
required, and these purchases are shown in Table 4-7. Mostly new units
will be purchased; with the only exceptions being rebuilt units for
the service scooptrams and the two graders. Two graders are required
because of the captive areas in the mine.
The equipment purchases are planned to meet the maximum requirements
in the mine plan. Approximately half of the maximum fleet will be new
equipment. Most of the mobile equipment purchases will be made in the
pre-production period (prior to Year 3).
Table 4-7 - New Underground Mobile Equipment Purchases
- ----------------------------------------------------------------------------------------------------
Equipment Description Existing New Maximum
Units Equipment Planned
Purchases Fleet
- ----------------------------------------------------------------------------------------------------
Underground Truck 13 tonne 1 - 1
Underground Truck 20 tonne - 4 4
Development Jumbo Single boom, electric/hydraulic 2 2 4
Scissor Lift Dux 3 5 8
Production Scoop* 4.0 cuyd capacity 4 3 5
Production Scoop 3.5 cuyd capacity 1 - 1
Development Scoop 3.0 cuyd capacity 4 - 4
Services Scoop 2.0 cuyd capacity 5 3 8
ITH Drill, Crawler CMS, 114mm, diesel 1 - 1
ITH Drill, Rubber Tire McClean, 114mm, electric 2 - 2
Pneumatic Top Hammer Drill BCI, 54mm 2 3 5
Road Grader - 2 2
Service Vehicle 2-Tractor, 2-Forklift 4 6 10
- ----------------------------------------------------------------------------------------------------
TOTAL 29 28 55
- ----------------------------------------------------------------------------------------------------
* 2 Scoops to be retired
Planned improvements to the Langlois mine mobile maintenance program
include the following:
o Underground garages: A new bay will be added to the level 9
shop and a new maintenance facility will be constructed on
level 13. Wash bays will be constructed for both shops.
o There will be underground warehousing of high turnover items
near the two main shops, including access to parts manuals.
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o A full time maintenance planner will be hired in addition to
the clerk used in the past. The planner will order parts and
organize deliveries underground. PM kits will be used.
o Of the maintenance workforce, 6 must be Class 1 mechanics.
If they cannot be hired, a maintenance contractor will be
used. In the past there were as few as 2 Class 1 mechanics.
o A maintenance computer program will be purchased to manage
work orders and the PM program. Compliance to the scheduled
PM's will be tracked. Equipment cost per hour will be
tracked.
o An oil analysis program will be initiated.
o Training services will be purchased with new equipment
units.
4.6 PLANNED PRODUCTIVITIES
The mining productivities used in the feasibility study are shown in
Table 4-8. They are based on scheduled shift durations of 8-hours.
Table 4-8 - Planned Mining Productivities
-------------------------------------------------------------------------------------
Item Description Productivity
-------------------------------------------------------------------------------------
Trackless Development Advance Ore & Waste Drifts & Ramps 0.92m/manshift
Production Drilling 54mm diameter 65m/manshift
(Excludes Moves) 114mm diameter 50m/manshift
Blasting 54mm & 114mm diameter 350t/manshift
Production Mucking (3.5 & 4.0 cuyd)
Tramming to Remuck (Zone 3) 330t/manshift
Loading Trucks (Zone 97) 343t/manshift
Trucking - Ore & Waste 20t truck (Zone 97 Ore & Waste) 293t/manshift
Paste Backfilling - 70t/hr
Note: 8 hr manshifts
-------------------------------------------------------------------------------------
It is difficult to directly compare each of these productivities to
historical performances since they were not all rigorously tracked.
More effort was historically placed on tracking unit costs. Some of
the planned productivities are based on estimates of past performance
at the Langlois mine, while others have been calculated starting from
basics (cycle times and payloads, etc.). The productivities are
individually described below.
Trackless drifting performance has been tracked, and the actual
results for 1999 and 2000 are 0.84 and 1.04m/ms respectively. The
planned drifting productivity of 0.92m/ms falls well in line with past
achievements and is also reasonable by industry standards.
The drilling productivities shown include time spent moving, which is
allowed for in the study. Estimates of past drilling productivities
provided verbally by mine operators are as follows:
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o For 114mm diameter drilling, 45 to 60m/ms
o For 54mm diameter drilling, approximately 80m/ms
The blasting productivity has not been examined in detail, however the
figure used in the feasibility study is appropriate for the conditions
based on experience at other mines. Also, it is relatively easy to
pro-rate the size of blasting crews required in the future based on
the number of blasters used in the past.
Production mucking has been considered for two conditions. Firstly,
for Zone 3 mucking from the draw point to a remuck bay or ore pass,
and secondly for Zone 97 mucking from the drawpoint to a remuck and/or
truck. In the second case the scooptram must load the truck with about
4 buckets of ore (3 from the remuck), before returning to the draw
point.
The scooptram mucking productivities have been checked from first
principles, including travel distances, average speeds, and remote
control bucket filling. The historical productivity for 3.5 and 4.0yd
scooptrams is approximately 300t/ms.
Trucks of 20t capacity have not been used at the Langlois mine in the
past, therefore the trucking productivity has been estimated from
first principles. Productivities were developed individually for all
domains and varies from 600 to 200t/ms based on the average haulage
distance from the shaft to the production area that year. See Appendix
Q - Trucks.
The paste backfilling rate is the historical rate.
4.7 VENTILATION
The Langlois mine ventilation system will be upgraded from its current
220,000cfm to 360,000cfm to provide for activities in Zone 97 and to
provide additional fresh air for the increase in the diesel equipment
fleet planned. The peak year for ventilation requirements is Year 6
when the maximum amount of diesel equipment will be in use.
The total quantity of ventilation required for Year 6 was determined
from the estimated diesel equipment list for each domain. Each diesel
unit has a certificate issued by CANMET specifying the required
ventilation volume for the unit. For each mining domain the total
ventilation required is determined by the maximum number of diesel
units working in the domain, but it is not a straightforward sum (it
is less). For several diesel units operating in one ventilation
circuit, the total air requirement must be determined by following
Quebec mining regulations. (Regulation 102 (2)(a) of Regulation
respecting occupational health and safety in mines, September 12,
2000)
For Year 6, this process was used for each active domain, and the
total was just over 300,000cfm. This amount was then increased to
provide ventilation for other parts of the mine with no diesel
equipment operating, and to allow for some contingency. The design
fresh airflow was estimated at 360,000cfm.
The future ventilation network was modelled using Eolaval software.
See Appendix R -EOLAVAL. Figure 4-16 shows the ventilation network and
the design air flows on the main levels, ramps and in the ventilation
raises. The following description highlights the features of the
future ventilation system, and the elements of work required to
achieve it.
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4.7.1 ZONE 97 VENTILATION
A new 3.66m diameter fresh air raise is required for Zone 97 from
surface to level 9. It must also be extended from level 9 to level 13.
Domain 14a will be provided with a 1.83m diameter bored raise from 13
level to 15 level.
On the new 3.66m raise, two surface mounted Joy 84-30-1180 fans in
series operating at 8.3 inches water gauge will deliver 220,000cfm. A
building at the raise collar will house the fans and a propane heating
system. See Appendix S - Zone 97 Intake Arrangement.
Exhaust air from Zone 97 will travel through levels 6, 9, and 13 and
through the existing 1.83m diameter raise bore hole from surface to
level 9 in Zone 97. A surface mounted fan on this raise will exhaust
70,000cfm. This smaller bored raise will eventually be extended from
level 9 to level 13.
At present, Zone 97 is ventilated by level 9 fresh air that is
exhausted to surface through the existing 1.83m raisebore hole. This
system can supply 60,000 cfm to support the initial development work
and it does not have to be interrupted during construction of the
permanent ventilation network.
4.7.2 ZONE 3 VENTILATION
There are no significant changes planned for Zone 3 ventilation.
4.7.3 ZONE 4 VENTILATION
In Zone 4, exhaust air will travel to surface through the existing
backfill raise. This will be a dual-purpose raise, allowing waste
backfill to be sent underground at shift change. While the lower
section of the raise stores waste backfill, ventilation will be
maintained through upper level break throughs to the raise.
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Figure 4-16 - Ventilation Network - Year 6
[PICTURE]
4.8 ORE AND WASTE HANDLING SYSTEMS
4.8.1 ORE AND WASTE HANDLING SYSTEM SELECTION
Three ore handling systems were evaluated for Zone 97 to determine
which would be the most appropriate.
The first option was to excavate conventionally supported disposable
ore passes, each with an expected life of approximately 100,000t. A
new disposable ore pass would be excavated just prior to the existing
ore pass becoming unusable.
The second option was to install long-lasting support in each ore pass
at an estimated cost of $8,500/m.
The third option was to use trucks to haul the ore and waste from
remucks to the ore handling system at the shaft.
The long-lasting support option was not chosen, as it required the
highest capital cost and was not entirely risk free. The other two
options were both comparable in the cost analysis, however the truck
option was found to be more efficient and more certain and was thus
chosen. The
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disposable ore pass option requires an extra raise development crew
that could slow down other activities. As well, lost production,
dilution and re-handling cannot be avoided and therefore the approach
is less certain.
4.8.2 TRUCK SELECTION
Several sizes of truck were evaluated for Zone 97 muck handling
systems. The truck size analysis considered the total system costs
including capital, truck operating cost, drift excavation and
ventilation costs. The 20 tonne trucks were found to be the most
efficient and economic. The cost of over-size excavations required by
larger trucks is greater than the cost savings associated with better
performance. Zone 97 is divided into two main sectors: the 9 level
access and the 13 level access. These sectors will need at least two
trucks each in order to have a versatile and operational fleet
(maintenance and mechanical downtime). See Appendix Q - Trucks.
4.8.3 EXCAVATION REQUIREMENT
The excavation requirement for the 20 tonne capacity truck is 3.5m
high x 3.7m wide, which is similar to the historic ramp size of 3.2m
high x 3.7m wide. Refer to Figure 4-17.
Figure 4-17 - Drift Cross Section
[PICTURE]
The 9 level access drift, which is currently 2.9m high x 2.9m wide,
will have to be slashed in order to accommodate the trucks. A total of
1,140m of drift will be slashed. Refer to Figure 4-18.
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Figure 4-18 - Slashing of Access Drift 9 Level, Zone 97
[PICTURE]
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4.8.4 TRUCK PERFORMANCE
The trucks will be used on level 9 and level 13 to haul ore and waste
from Zone 97 sub levels to the ore/waste passes located near the
shaft. Refer to Figure 4-19. Haulage profiles were detailed for
several stages of mine life. The cycle time ranged from 11 to 27
minutes per trip with a maximum speed of 17km/h in the haulage drift
and 10km/h in the ramp. Trucks will be loaded by 3.5 or 4.0 yd
capacity scooptrams and they will dump on a grizzly at the main ore
pass.
Figure 4-19 - Typical Stope Mucking Arrangement
[PICTURE]
4.8.5 ORE PASSES
During 2000, a new ore handling system was commissioned on level 11
which included a 570 raise from 11 level to 10 level. The ore pass
handled ore from level 9 and above. A rockbreaker/grizzly is situated
on 10 level and there is a vibrating pan feeder situated on 11 level
which feeds a conveyor, that transfers the ore to the shaft and a new
loading pocket.
The ore pass could handle waste, however, the zinc ore tended to arch
at the outlet of the raise and failed to slide on the footwall of the
raise, thus creating blockages. In order to prevent the blockages, the
raise would be left empty. This resulted in rapid deterioration of the
Fondag lining system in the raise. The Fondag further affected zinc
recovery in the mill.
Based on an investigation detailed in an external report prepared by
Jenike Johanson (project report 9098-2) concerning the flow
characteristics of the ore at the Langlois mine, a new ore pass system
will have to be developed.
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The findings of the report indicated that no acceptable modification
to the existing ore pass system can be made as the slope angle of the
current ore pass is too shallow (57 degrees and 62 degrees). According
to the Jenike report, the new ore pass system has to be vertical or at
a very steep angle and should have a minimal internal diameter of 2.7m
for flow-through purposes. It was also recommended that the new ore
pass system be lined.
McIntosh Redpath Engineering Limited was commissioned to evaluate
different lining options for the ore pass system. This study reviewed
alternate types of liner and their related installations, comparative
capital costs, durability, ground support and ongoing maintenance
issues. The new ore pass system will be lined with a mild steel
long-lasting liner that will be equipped with control chains to
control the flow.
The main ore pass connecting the loading system (11 level) and the
rock breaker (10 level) will have an internal diameter of 4.27m and a
storage capacity of 2,000 tonnes. This vertical silo has been
estimated to cost $1,500,000. The remainder of the ore pass system
from 10 level to 8 level will be a lined vertical raise with an
internal diameter of 2.7m. The total cost for the ore pass between
each level (approx. 60m) is estimated to be $884,000. This type of
steel lined vertical ore pass is expected to have very low maintenance
requirements. The construction of the ore pass system between 11 level
and 8 level will take 6 months to complete. After an optimization was
completed, the option to extend the orepass from 8 level to 6 level
was replaced by trucking the ore instead. See Appendix T -
McIntosh/Redpath Engineering Report Concerning Langlois Mine
Conceptual Ore/Waste Pass Lining Options Review.
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Figure 4-20 - Actual and Future Ore and Waste Passes
[PICTURE]
4.8.6 MATERIAL SIZING
Three material sizing systems will be in operation. The current system
on level 10 will handle all of the production from Zones 3 and 4, in
addition to production from Zone 97 above 9 level, for a total of 2.1
million tonnes of ore. The system includes a Teledyne 710 rock breaker
and a grizzly with 300mm openings. The system is fed by the main ore
pass above level 10 and delivers ore to the level 11 conveyor/loading
pocket.
The 13 level will handle a total of 1.2 million tonnes of ore from
Zone 97 below 9 level. The ore will be dumped in a short ore pass,
which feeds a 125-Hp jaw crusher located on level 14. The ore will be
crushed to minus 150mm and then passed through a raise to the loading
pocket at 16 level.
An additional system will be made on the 5th level for Zone 3 and 4
material above 6th level. This system will consist of a grizzly with
300mm openings as well as a rockbreaker. Due to the small tonnage
involved (up to 260t/d), this installation is not as critical as the
others.
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4.8.7 HOISTING
Ore is skipped to surface by the production hoist from levels 11 and
16 and will also be skipped from level 5. Two loading pockets are
already in place and operational. An additional pocket, similar to the
11th level facility will be installed on 5 level however it will not
include a transfer conveyor. The maximum continuous skipping rate is
3,500 tonnes per day. The hoist is equipped with two seven tonne
capacity skips and is exclusively reserved for ore and waste handling
purposes. Men and materials are transported by the service hoist.
The conversion of the existing ore pass from 11th to 10th level to a
vertical silo having a storage capacity of 2,000 tonnes will permit a
steady feed at the loading pocket. It is estimated that the ore
storage capacity for the 11th level system will be 3,000 tonnes (new
ore bin plus the raises) and 1,000 tonnes for the 16th level.
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5.0 MINERAL PROCESSING
5.1 PROCESSING
The Langlois mill processes approximately 2,100 tonnes per day, 4 days
per week with one day per week left for maintenance. The nominal
capacity of the mill is 2,500 tonnes per day. Zinc and copper
concentrates are produced by selective flotation, with some gold and
silver recovered in the copper concentrate.
Table 5-1 - Ore Composition
--------------------------------------------
Specific gravity of ore 3.6
Specific gravity of zinc concentrate 4.0
Specific gravity of copper concentrate 4.4
Specific gravity of tailing 3.4
--------------------------------------------
Figure 5-1 - Mill Flowsheet
[PICTURE]
5.1.1 PRIMARY CRUSHING
There will be three underground facilities for primary crushing. On 14
level the ore is crushed to minus 165mm (6.5") in a 940mm x 1245mm
(37"x 49") jaw crusher. On 5 and 10 level, ore will be sized to 300mm
(12") with a grizzly/rockbreaker. The ore is then skipped to surface
where it is stored in two fine ore storage bins that have capacities
of 600 and 2,400 tonnes.
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5.1.2 GRINDING
Four vibrating feeders control the feed to the grinding circuit at 90
tonnes/hour. The grinding circuit consists of a 6.1m (20-ft) diameter
x 3.4m (11-ft) long semi autogenous grinding ("SAG") mill in open
circuit and a 4.0m (13-ft) x 6.1m (20-ft) long ball mill in closed
circuit with five cyclones (380mm diameter). There is also a 3.1m
(10-ft) diameter x 3.2m (10.5-ft) copper regrind mill however it is
currently not in use as it doesn't provide any metallurgical
advantage. Tests will be carried out in the future to see if the
regrind mill can be used in the zinc flotation circuit. The regrind
mill is in closed circuit with four cyclones (150mm diameter).
Table 5-2 - Grinding Circuit Data
- -------------------------------------------------------------------------------------------
SAG Mill Ball Mill Cu Regrind
- -------------------------------------------------------------------------------------------
Avg. Slurry Density % Solids 80 72
Diameter Mm Ball of Slug of 50.8 Ball of 12.7
101.6
Consumption Kg/t 0.11 1.60
Mill Work index of ore KWh/t 9.3 11.8 11.8
Mill speed Rpm 9.41 16.03 12.84
Mill speed % of critical 54.92 76.90 54.55
Mill lining Material Rubber Rubber Rubber
Mill motor Hp 1,500 2,500 800
- -------------------------------------------------------------------------------------------
The grinding fineness is approximately 88 % -200 mesh and the
circulating load is 200 %. SO2 is added as a depressing agent for
sphalerite and pyrite. The grinding circuit overflow is conditioned by
agitation with SO2, Aerophine 3418-A and Aerofloat 208, as collecting
agents.
Table 5-3 - Screen Analysis Weight Distribution Summary
- ------------------------------------------------------------------------------------------------------
Mesh
--------------------------------------------------------------------------------------
65 100 150 170 200 250 325 400 500
Cyclone Feed 68.9 64.4 57.9 53.4 48.4 42.8 30.1 25.8 24.9
Cyclone O/F 99.9 99.4 96.3 93.4 90.1 86.4 76.4 70.5 69.0
Zn conc. 100.0 100.0 99.7 98.7 96.7 92.8 80.0 73.0 61.7
Cu conc. 100.0 99.9 99.7 99.5 99.0 97.9 90.6 83.9 73.8
Tailing 99.4 97.4 93.4 90.2 87.2 84.3 73.2 67.1 54.8
- ------------------------------------------------------------------------------------------------------
5.1.3 FLOTATION
The grinding circuit overflow feeds the copper flotation circuit,
which contains a conditioner and conventional rougher, scavenger and
cleaner stages using Outokumpu cells. At each stage, water, air,
3418-A, 208 (to recover silver and gold) and MIBC frother are added.
The tail of the scavenging and secondary scavenging stages constitutes
the feed to the zinc flotation circuit. The re-cleaner concentrate is
sent to the copper thickener (4.0m, 13-ft diameter).
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Table 5-4 - Copper Flotation Reagents
- ----------------------------------------------------------------------------------------------
Reagents Conditioner Rougher Scavenger Cleaner Re-cleaner Thickener
- ----------------------------------------------------------------------------------------------
SO2 400 g/t 100 g/t
208 6.7 g/t 3.8 g/t 1 g/t
3418-A 13.2 g/t 8.8 g/t
MIBC 8 g/t 1.3 g/t 1.2 g/t 0.5 g/t
PH 7.0 10.5
% solids 41 6 18
- ----------------------------------------------------------------------------------------------
Pulp sent to the zinc flotation circuit is conditioned in two tanks
with copper sulphate to reactivate sphalerite in the first tank and
with 507-B in the second tank as a collector. Lime is added to the
discharge of the second tank in order to increase pH and depress
pyrite. The circuit includes rougher, scavenger and cleaner stages.
MIBC is added as a frother. The tail from the primary and secondary
scavengers constitutes the final mill tailings, and the concentrate
from the third cleaning is sent to the zinc thickener (10.1m, 33-ft
diameter).
Table 5-5 - Zinc Flotation Reagents
- --------------------------------------------------------------------------------------------------------------------
Reagents Conditioner Conditioner Rougher Scavenger Cleaner 2nd Cleaner Re-cleaner Thickener
1 2
- --------------------------------------------------------------------------------------------------------------------
CuSO4 1100 g/t
507-b 54 g/t 8.8 g/t 1.2 g/t
MIBC 4 g/t 1.1 0.7 g/t 0.7 g/t 0.5 g/t
Lime 1990 g/t
PH 9.9 9.9 10.3 10.3 10.3 10.5
% solids 33 12 20
- --------------------------------------------------------------------------------------------------------------------
5.1.4 DE-WATERING
Underflows from the zinc and copper thickeners are pumped separately
from the mill building to the filtration and loading stations and
stored in tanks (Copper - 110 tonnes = 42 days capacity and Zinc - 375
tonnes = 21 hour capacity). A storage building is connected to the
building for the storage of final concentrates (6,000 tonne capacity).
The storage building is located beside the railway to facilitate
concentrate loading, and can also be used for trucks if required.
Filtration is done by an Eimco filter press to obtain concentrates
with 7.0% moisture.
5.1.5 TAILINGS DISPOSAL
Approximately 50% of the tailings are used in the paste backfill and
the remaining tailings are sent to the 1.88 km2 tailings pond, located
3 km from the mine site. The tailings are discharged under water to
prevent acid generation.
Most of the tailings pond overflow is recycled as process water, with
a portion of tailings pond overflow released to the Wedding River
after being treated with caustic lime to neutralize pH levels.
The pond has approximately 3,750m of dams, which have a maximum height
of 12m, and it is accessed by a 1.2km road. The dams are composed of a
mix of gravel and till with a geotextile membrane to ensure its
impermeability.
The tailings line is 250mm in diameter and the recycled water line is
280mm in diameter. The lines average 5.5km in length. The pumping of
fresh and recycled water to the mill is ensured by
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two 300HP pumping stations. The first one is located at the edge of
the Wedding River near the discharge of the pond. The second station
is located near the principal dam (extreme west) of the tailings pond.
Figure 5-2 presents the tailings deposition plan for the life of the
mine. The total remaining volume available for tailings disposal is
1,867,864 m(3) at a density of 1.3 t/m(3), which equates to 2,428,000
tonnes available without raising the dam heights.
o Tonnes available in tailings pond - 2,428,000t
o Tonnes in the feasibility study - 2,139,000t
o Tons available at the end of mine life - 289,000t
This tonnage represents the minimum amount available. Past practice
has shown that after adjustments following bathymetry, more tonnage
(available space) remains than calculated. Only minor amounts of work
such as maintenance and inspections are necessary for the life of the
project.
5.1.6 PASTE BACKFILL PLANT
The paste backfill plant uses a backfill preparation process that
involves thickening and filtering of mill tailings which are mixed
with cement and water to make a paste that is sent underground by
gravity. The capacity of the paste backfill plant is 70 tonnes per
hour using one disc filter. The paste backfill circuit operates on a
batch of 3.5t due to being limited by the milling rate, which is
maintained between 90 and 95 t/h. A second filter is maintained on
standby.
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Figure 5-2 - Tailings Deposition Plan
[PICTURE]
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5.2 METALLURGICAL RESULTS
5.2.1 HISTORICAL METALLURGICAL RESULTS
The Langlois mill has historically produced a zinc concentrate in the
range of 52.0 to 52.9% (during four previous years) from head grades
ranging from 6.4 - 7.9% zinc. The grade of the copper concentrate has
been in the range of 21.3 - 24.0%. Recovery of zinc has been in the
range of 92.6 to 93.6% and copper recovery has been in the range of
69.9 - 75.1%.
Table 5-6 lists the operating parameters from 1997 - 2000 along with
the metallurgical results.
Table 5-6 - Metallurgy-Historical Production 1997 - 2000
- -----------------------------------------------------------------------------------------
1997 1998 1999 2000
- -----------------------------------------------------------------------------------------
PRODUCTION
Tonnes Milled 261,068 414,742 402,224 310,466
HEAD GRADE
Zn (%) 6.36 6.53 7.51 7.92
Cu (%) 0.36 0.33 0.36 0.36
Au (g/t) 0.14 0.14 0.16 0.19
Ag (g/t) 29.08 26.74 28.78 31.44
RECOVERIES
Zn (%) 92.6 93.3 93.6 92.8
Cu (%) 75.1 73.6 69.9 70.6
Au (%) 37.3 39.8 31.7 32.1
Ag (%) 43.7 35.4 30.1 30.9
CONC. GRADE
Zn (%) 52.7 52.9 52.0 52.7
Cu (%) 21.3 24.0 21.6 22.3
IMPURITIES
Fe (%) in conc. Zn 10.9 10.6 10.3 10.0
Pb (%) in conc. Cu 5.7 5.4 3.4 3.5
Zn (%) in conc. Cu 7.57 7.24 7.02 6.99
TONNES CONC.
Zn (t) 29,181 47,770 54,397 43,287
Cu (t) 3,322 4,198 4,739 3,589
TONNES METAL
Zn (t) 15,382 25,281 28,280 22,792
Cu (T) 709 1,009 1,023 800
Au (0z) 515 736 674 613
Ag (Oz) 90,914 126,224 112,063 96,976
- -----------------------------------------------------------------------------------------
The zinc recovery in 2000 was affected by contamination from the early
deterioration of a new shotcrete lining system in the underground ore
passes. A new ore pass was put into operation towards the end of June
2000. The shotcrete (Fondag) lining deteriorated and became mixed into
the ore, adversely affecting zinc recovery during the summer of 2000.
Usually, zinc recovery maintains a constant winter/summer cycle with
recovery at its highest during the summer months. This was not the
case during the summer of 2000.
The Fondag shotcrete contains high levels of calcium and aluminium.
This contamination resulted in a lack of selectivity in the zinc
flotation circuit and compromises were carried out to maintain the
best relationship between the recovery and the grade of concentrate.
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5.3 OPERATING PLAN
The operating plan for the feasibility study includes metallurgical
values which are believed to be achievable. Recoveries of copper, gold
and silver are based on projected head grades applied against formulas
determined by graphs for each metal following historic head grades.
From year 4 onwards, the copper recovery has been significantly
increased, as it is believed that with higher copper head grades, the
circuit will work more efficiently.
The zinc head grade does not seem to affect the zinc recovery. The
mill water temperature is a more influential factor for determining
zinc recovery. Recoveries improve in the summer when the water is
warmer.
The zinc recovery has been gradually increased over the years in the
operating plan. This is justified, as there are plans to recover heat
and to use this to control the water temperature during the winter
months and to advance certain research tasks as of the first year of
operation. A study has been initiated to find an economical system to
heat the water required by the operation.
Additionally, feeding the mill on a regular basis, with no
interruptions will be very beneficial to advancing metallurgical work
and optimising the circuit. Work is also planned on improving the
start-up procedure for the mill in order to decrease the time
necessary to stabilize the circuits.
Table 5-7 details the operating plan for the feasibility study.
Table 5-7 - Metallurgy - Operating Plan
- --------------------------------------------------------------------------------------------------------------------------
YEAR 2 YEAR 3 YEAR 4 YEAR 5 YEAR 6 YEAR 7 YEAR 8 YEAR 9 YEAR 10 TOTAL
- --------------------------------------------------------------------------------------------------------------------------
PRODUCTION
Tonnes Milled 73,878 381,045 450,000 450,000 450,000 450,000 450,000 450,000 167,837 3,322,760
HEAD GRADE
Zn (%) 11.29 10.51 10.78 10.93 10.97 10.96 10.73 10.51 10.59 10.78
Cu (%) 0.63 0.63 0.69 0.73 0.87 0.97 0.96 0.89 0.92 0.82
Au (g/t) 0.09 0.08 0.08 0.08 0.09 0.09 0.09 0.08 0.08 0.09
Ag (g/t) 54.35 50.10 51.66 52.54 53.79 53.89 52.16 50.22 51.77 52.13
RECOVERIES
Zn (%) 93.89 93.76 93.81 93.64 93.53 93.54 93.68 93.57 93.56 93.65
Cu (%) 77.46 76.56 78.99 80.48 83.79 85.48 84.98 84.81 85.27 82.33
Au (%) 29.22 30.00 29.22 29.22 28.44 28.44 29.22 28.44 28.44 28.95
Ag (%) 33.87 33.92 35.11 36.50 36.42 36.32 36.03 35.89 35.89 35.75
CONC. GRADE
Zn (%) 52.80 52.80 53.75 54.25 55.00 55.50 56.00 56.00 56.00 54.82
Cu (%) 22.00 22.00 22.50 23.00 24.00 24.50 25.00 25.50 25.50 23.87
IMPURITIES
Fe (%) in conc. Zn 10.0 10.0 9.5 9.3 9.0 8.7 8.4 8.2 8.2 9.04
Pb (%) in conc. Cu 3.0 3.0 3.0 3.0 3.0 3.0 3.0 3.0 3.0 3.0
Zn (%) in conc. Cu 7.0 7.0 7.0 7.0 7.0 7.0 7.0 7.0 7.0 7.0
TONNES CONC.
Zn (t) 14,832 71,115 84,665 84,897 83,947 82,124 80,774 79,025 29,695 612,075
Cu (t) 1,639 8,354 10,901 11,495 13,668 15,229 14,685 13,320 5,163 94,454
CONTAINED METAL
Zn (t) 7,831 37,549 45,507 46,057 46,171 46,134 45,233 44,254 16,649 335,366
Cu (t) 361 1,838 2,453 2,644 3,280 3,731 3,671 3,397 1,317 22,691
Au (oz) 62 294 338 338 370 370 381 329 123 2,606
Ag (oz) 43,724 208,177 262,424 277,456 283,442 283,192 271,893 260,775 100,363 1,991,345
- --------------------------------------------------------------------------------------------------------------------------
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Table 5-8 - Metallurgy Summary
-------------------------------------------------------------------------
Assays
-------------------------------------------------------------------------
Zn % Cu % Au g/t Ag g/t
-------------------------------------------------------------------------
Heads 10.78 0.82 0.09 52.13
Zn concentrate 54.82 0.42 0.46 76.0
Cu concentrate 7.0 24.0 0.86 659
Tails 0.62 0.086 0.025 12
-------------------------------------------------------------------------
Distribution
-------------------------------------------------------------------------
Zn % Cu % Au g/t Ag g/t
-------------------------------------------------------------------------
Heads 100 100 100 100
Zn concentrate 93.65 9.43 39.8 41.5
Cu concentrate 1.84 82.33 29.0 35.8
Tails 4.51 8.24 31.2 22.7
-------------------------------------------------------------------------
5.3.1 ZONE 97 TESTWORK
Laboratory testwork on mineralized material from Zone 97 has been
completed. Several tonnes of Zone 97 material were mixed with Zone 3
and 4 material with good metallurgical results. The circuit operated
well with the help of some minor adjustments.
An independent laboratory analysis was made. A microsounder was used
for determining sphalerite particles in the sample in order to compare
it with the historical values from Zones 3 and 4.
The results are found in Table 5-9.
Table 5-9 - Iron in Sphalerite
- ---------------------------------------------------------------------------------------
Period % Fe % Zn % S Comment
- ---------------------------------------------------------------------------------------
Sept 00 6.74 59.31 33.44 Unstable circuit - Fondag
June 00 6.79 60.62 32.94 Stable Circuit- contain Zone 97 ore
Jan 99 6.64 59.92 33.20 Unstable circuit- contain some paste backfill
April 98 6.74 60.47 33.23 typical day
- ---------------------------------------------------------------------------------------
Table 5-10 - Metallurgical Results
---------------------------------------------
Period % Zn in % Fe in % Zn Rec.
Zn Conc. Zn Conc.
---------------------------------------------
Sept 00 50.8 11.05 92.0
June 00 55.3 8.03 95.2
Jan 99 51.0 10.8 91.4
April 98 52.5 10.1 93.3
---------------------------------------------
The iron in sphalerite is similar in both cases. The tests which were
carried out in the Langlois mine mineralogy laboratory were designed
to check compatibility of the existing circuit by using the standard
test (even conditioning, even pH, even calculation of proportioning,
even time of
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flotation, etc). The results did not show anything in particular other
than a need for some operating adjustments to the copper circuit due
to the expected higher grades
5.4 FUTURE METALLURGICAL PROJECTS
As part of the ongoing metallurgical work at the Langlois mine, the
following studies and projects have been or will be initiated.
5.4.1 OPTIMIZATION AND MAINTENANCE OF CONTINUOUS ANALYSER (BOXRAY)
The continuous analyser is a principal tool in the mill. It is
necessary to increase the precision and reliability of the unit in
order to use to its full capacity.
It will be necessary to do a thorough maintenance of the analyser unit
prior to start-up of operations in order to be able to establish
proper calibration curves for regular sampling. The estimated cost for
this work is approximately $20,000.
5.4.2 METALLURGICAL STUDY OF ZONE 97
Two studies are planned with the following objectives:
o To determine the metallurgical performance possible for Zone
97 material and to identify the reagents which will be most
adequate in order to better optimize the flotation circuit.
o To perform an external study by an independent laboratory
(Lakefield) during the shutdown period in order to be well
prepared for the re-opening of mill.
5.4.3 SODA ASH (NA2CO3) STUDY
The objective of this study is to develop means of preventing calcium,
magnesium, and/or aluminium from binding to the surface of sphalerite,
to make it possible to control the circuits when there is paste
backfill present in the mill feed. Potential benefits include reduced
zinc losses and increased quality of the zinc concentrate.
Some work has already been done on this project. Work remaining on the
project is the determination of how to measure the soda ash and how to
control the addition without the addition becoming a controller of pH.
5.4.4 WATER HEATING SYSTEM
A database has been assembled and it is clear that water temperature
has an impact on the zinc recovery. Work is underway to find an
economic method to control the water temperature used in the flotation
circuit and maintain a temperature of 20 degrees C or more. Achieving
such will prevent loss in recovery of approximately 2-3% for 6 months
of the year.
5.4.5 OTHERS
Other projects that are planned include:
o Test collectors and foaming agents.
o Campaigns on the zinc circuit to determine whether a regrind
mill is required.
o Staff training (at all levels).
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6.0 ENVIRONMENTAL CONSIDERATIONS
6.1 LICENCES AND CERTIFICATES OF AUTHORIZATION
The Langlois operation is covered under various environmental permits
which are all in good standing. No new permits will be required in
order to resume operations. The mine has been in compliance with the
existing regulations.
The following is a list of the licenses and certificates of
authorization. The actual licenses can be found in Appendix U -
Licences and Certificates of Authorization.
o Exploitation miniere du projet Grevet (cession) - 19 juin
2000
o Construction et exploitation du parc a residus Projet Grevet
(cession) - 16 juin 2000
o Modification - Systeme de traitement de l'effluent - 24
fevrier 2000
o Modification - Evacuation du parc a residus, frequences des
peches scientifiques et comite de liaison - 23 decembre 1998
o Fosse de drainage - 22 decembre 1998
o Modification - Maintien d'un effluent au parc (hiver 97-98)
22 decembre 1997
o Modification - Amenagement d'un nouveau entrepot de dechets
dangereux - 13 aout 1997
o Traitement des eaux usees - 16 juillet 1997
o Modification - Redemarrage de la Mine Langlois - 23 juin
1997
o Modification - Maintien d'un effluent au parc (hiver 96-97)
- 24 mars 1997
o Alimentation eau potable - 5 juin 1996
o Exploitation d'une sabliere - 20 decembre 1995
o Entreposage dechets dangereux - 13 septembre 1995
o Modification - Construction et exploitation du parc a
residus Projet Grevet - 12 juillet 1995
o Construction et exploitation du parc a residus - 19 mai 1995
o Approbation MRN - Emplacement destinee a recevoir les
residus miniers - 28 fevrier 1995
o Exploitation miniere - 8 fevrier 1995
o Construction de la ligne de transport d'energie a 120 kV -
12 decembre 1994
o Exploitation miniere Grevet - 18 novembre 1994
o Modification - Programme souterrain d'exploration - 25 mai
1994
o Modification - Entreposage des huiles usees - 13 decembre
1993
o Decapage de mort terrain et amenagement d'une halde de
mort-terrain - 10 decembre 1993
o Entreposage d'huiles usees - 28 juillet 1993 Puits d'eau
potable - 19 avril 1993
o Systeme de traitement des eaux usees domestiques du bureau
temporaire - 25 fevrier 1993
o Traitement des eaux usees domestiques - 30 novembre 1992
o Programme souterrain d'exploration - 10 novembre 1992
o The Langlois mine environmental policy.
Langlois mine received its ISO 14001 environmental accreditation in
December 1998 but has since allowed it to lapse. This program included
detailed procedures and action plans designed to
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protect and improve the environment at the mine. This accreditation
was a corporate accreditation of Cambior, so the Langlois mine is no
longer accredited. However the system and standards are maintained in
place.
6.2 WASTE MANAGEMENT
6.2.1 WASTE ROCK
There are two surface dumps for the storage of waste rock. One dump is
for the storage of non-acid generating rock and the other is for
material with a low potential for acid generation.
At the end of the life of the mine, the majority of the non-acid
generating waste will be disposed of underground and the
acid-generating waste will be placed under 1m of water in the tailings
pond to avoid acid generation.
6.2.2 TAILINGS
As discussed in section 5.1.5, all of the tailings not used to make
paste backfill will be placed in the tailings pond. The pond has the
capacity to store all of the tonnes produced in the feasibility study.
At the end of mine life, there will be approximately 289,000t of
tailings volume available. The tailings dams will not require raising
for the tonnes mined in the feasibility study.
6.2.3 DOMESTIC/INDUSTRIAL WASTE
All domestic waste is sent to the municipal dump and the other
industrial waste is sent to the following facilities:
o Scrap metals: Legault Metal
o Paper/cardboard: Service Sani-Tri
o Copper/brass/lead batteries/aerosols/roll-off/etc: Recyclex
o Used oil: Greenhouses of Guyenne
o Old grease: Heist
6.3 WATER MANAGEMENT
Run-off water from the ore stockpile is collected underground by a
borehole in the bottom of the dump. This water is combined with the
tailings and deposited in the tailings pond. The water is then
recycled back to the mill for use as process water.
Discharge water is treated with lime in the pumpbox in the mill to
maintain a pH of approximately 7.5 to 8 to prevent metals from
dissolving. In addition, there is a caustic soda system which is
installed on the tailings pipeline which can also adjust the pH if
required. It has been installed as a precaution.
Water samples are taken on a regular basis. In Appendix V - Sampling
Stations, there is a plan of the sampling stations as well as the
sampling schedule.
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6.3.1 TAILINGS POND WATER DISCHARGE
Following a modification of the Exploitation Certificate issued by the
Ministry of the Environment on December 23, 1998, the discharge from
the tailings pond is allowed according to the following flows, namely:
o Jan. - Mar. 100 m(3)/h
o Apr 245 m(3)/h
o May, Jun. 300 m(3)/h
o Jul. - Dec. 245 m(3)/h
After the closure of the tailings pond, the surplus water will
discharge through the emergency discharge and an alternate discharge
point will be built on the south-eastern side of the pond.
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7.0 MANPOWER
Table 7-1 shows the projected mine site manpower requirements for the
Langlois mine by year for the life of the project. The approximate
average manning level for year 2000 is shown for comparison.
In January 2000 the actual work force was 125 employees plus 19
contractors for a mine site total of 144. As no capital expenditure
was approved during the first quarter of 2000, a portion of the work
force was temporarily laid off. Manpower was increased during the
year. In September 2000 the actual manning level was 157 employees and
39 contractors for a total of 196. Plans called for a continued
increase to the end of the year to reach a total of 248. This level
was not achieved because mine operations were suspended in December
2000. Table 7-1 shows a rough average for the partial year of
production in 2000.
The manpower estimates for future operations are based on the
feasibility mine plan at a higher production rate than was achieved in
2000. For mine production activities such as mucking and trucking,
manpower estimates were generated from the equipment calculations that
determined the required fleet sizes and number of operating units. For
mine service functions, manpower estimates were based on past
experience adjusted to reflect changes in the feasibility mine plan.
The Langlois mine workforce will include some contractors to handle
peak demands and to perform certain functions described in section
7.1.2. In the mine, these are typically areas requiring specialized
skills and/or very expensive equipment that are only needed for
periodic project work.
Most mine personnel are scheduled to work five days per week, with
weekends off. They will work two 8-hour shifts per day. The mill will
operate four days per week, utilizing two 12-hours shifts per day
until the production rate reaches 400,000 tonnes per year. For the
higher production rates they will operate five days per week on three
8-hour shifts per day.
Table 7-1 - Total Mine Site Manpower
- ------------------------------------------------------------------------------------------------------------------
2000 Year Year Year Year Year Year Year Year Year Year
1 2 3 4 4 6 7 8 9 10
- ------------------------------------------------------------------------------------------------------------------
Administration 3 2 2 3 3 3 3 3 3 3 3
Purchasing 4 1 2 4 4 4 4 4 4 4 4
Human Resources 2 1 2 2 2 2 2 2 2 2 2
Safety 2 1 3 4 4 4 4 4 4 4 4
Engineering 11 5 7 11 11 11 11 10 10 10 10
Geology 6 3 4 6 6 6 6 6 6 6 6
Maintenance 23 13 22 33 35 35 35 35 35 35 35
Electrical 6 3 6 7 7 7 7 7 7 7 7
Mine 75 11 50 85 87 87 82 76 76 76 76
Mill & Assay Lab. 23 4 21 25 30 30 30 29 29 29 29
Contractors 39 21 16 19 17 12 10 8 8 8 8
- ------------------------------------------------------------------------------------------------------------------
Total 194 65 135 199 206 201 194 184 184 184 184
- ------------------------------------------------------------------------------------------------------------------
Before the Langlois mine can resume full production, approximately 1.5
years of construction and development work is required. This work is
detailed in Section 8.0. Hiring to build up the work force will take
place over a period of time beginning in Year 1 and extending into
early Year 3.
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Previously, most of the workforce lived in the nearby town of
Lebel-sur-Quevillon. Breakwater will provide transportation from the
town to the mine site.
The underground employees, the mill employees and the maintenance
employees unionized themselves as part of the United Steel Workers of
America in August 1999. Negotiations between the company and the union
started in December 1999.
A conciliator was appointed to assist with negotiations in October
2000, however the first collective agreement has not yet been
finalized.
7.1 UNDERGROUND
Mine department personnel are shown in the following tables for mining
personnel, contractors, and mine maintenance personnel. Mill
maintenance personnel are included with the mill in section 7.2. Note
that engineering and geology are reported under 7.3 "Administration
and Technical Services."
7.1.1 MINING PERSONNEL
The planned manpower for the mine is shown in Table 7-2, compared to
the average for 2000.
Table 7-2 - Mine Department Personnel
- ---------------------------------------------------------------------------------------------------------------------------------
Area Description 2000 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year10
- ---------------------------------------------------------------------------------------------------------------------------------
Production Long hole-114mm 4 0 0 2 2 2 2 2 2 2 2
Long hole-54mm 2 0 0 4 4 4 4 5 5 5 5
Blasting 2 0 0 4 4 4 4 4 4 4 4
Mucking 12 0 4 10 12 12 13 13 13 13 13
Fill 4 0 0 4 4 4 4 4 4 4 4
Development Jumbo 8 0 24 24 24 24 18 12 12 12 12
Jack leg crew 10 0 0 0 0 0 0 0 0 0 0
Services Supervision 9 4 6 11 11 11 11 10 10 10 10
Services 1 0 1 3 3 3 3 3 3 3 3
Hoist man 4 2 4 4 4 4 4 4 4 4 4
Cage/Skip Tender 6 4 3 6 6 6 6 6 6 6 6
Shaft / Deck 2 1 2 2 2 2 2 2 2 2 2
Crusher Operator 2 0 2 3 3 3 3 3 3 3 3
Construction 2 0 2 2 2 2 2 2 2 2 2
Rehabilitation 7 0 2 6 6 6 6 6 6 6 6
Total 75 11 50 85 87 87 82 76 76 76 76
- ---------------------------------------------------------------------------------------------------------------------------------
The mucking category in Table 7-2 includes scoop operators and truck
operators. There will be two backfill operators on each shift working
on backfill piping, fill fences, and monitoring filling. There will be
four development jumbos operating two shifts per day, with three man
crews, for a total of 24 development miners. The supervision category
includes a superintendent, a captain, eight foremen, and one spare
foreman, for a total of 11. Services personnel will be engaged moving
supplies into the mine.
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7.1.2 MINING CONTRACTORS
The planned requirements for contractors are shown in Table 7-3,
compared to the average for 2000.
Contractors will be used to handle the peak in construction activity
at the beginning of the mine plan, to provide security services, and
to perform the following specialized functions:
o Installing steel linings in ore passes
o Driving a track drift on level 6 to Zone 97
o Raise boring large and medium diameter raises
o Diamond drilling to meet definition drilling needs
Table 7-3 - Contractor Personnel
- ---------------------------------------------------------------------------------------------------------------------------
2000 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10
- ---------------------------------------------------------------------------------------------------------------------------
Supervision 1 1 1 1 2 - - - - - -
Drifts 6 - 6 - - - - - - - -
Ore pass Construction - 12 - - - - - - - - -
Raisebore - - - 4 4 - - - - - -
Conventional raise - - - 2 - - - - - - -
Alimak raise 8 - - - - - - - - -
Construction - 4 3 4 3 4 2 - - - -
D.D.H. drilling 15 - 2 4 4 4 4 4 4 4 4
Services 2 - - - - - - - - - -
Mucking 3 - - - - - - - - - -
Security guards 4 4 4 4 4 4 4 4 4 4 4
- ---------------------------------------------------------------------------------------------------------------------------
Total 39 21 16 19 17 12 10 8 8 8 8
- ---------------------------------------------------------------------------------------------------------------------------
Raise bore contractors will be required in Years 3 and 4 to drive the
3.66m (12 ft.) ventilation raise.
7.1.3 UNDERGROUND MAINTENANCE PERSONNEL
The mine maintenance department requirements are shown in Tables 7-4
and 7-5, compared to the average for 2000. If there are difficulties
in hiring the mechanics on a timely basis, then a maintenance
contractor will be used to provide short-term coverage. In the area of
underground mobile maintenance, the target for qualifications is for a
minimum of six of the mechanics to be qualified to the level of a
Class 1 heavy-duty mechanic.
Table 7-4 - Maintenance Department Personnel
- --------------------------------------------------------------------------------------------------------------------------
2000 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10
- --------------------------------------------------------------------------------------------------------------------------
Superintendent 1 1 1 1 1 1 1 1 1 1 1
Supervision 2 2 2 3 3 3 3 3 3 3 3
Planner + clerk 1 1 2 2 2 2 2 2 2 2 2
Mechanic fixed 5 3 6 7 7 7 7 7 7 7 7
Mechanic mobile 13 5 10 19 21 21 21 21 21 21 21
Dry men 1 1 1 1 1 1 1 1 1 1 1
- --------------------------------------------------------------------------------------------------------------------------
Total 23 13 22 33 35 35 35 35 35 35 35
- --------------------------------------------------------------------------------------------------------------------------
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Table 7-5 - Electrical Department Personnel
- -----------------------------------------------------------------------------------------------------------------
2000 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10
- -----------------------------------------------------------------------------------------------------------------
Supervision 1 1 1 1 1 1 1 1 1 1 1
Electrician 5 2 5 6 6 6 6 6 6 6 6
- -----------------------------------------------------------------------------------------------------------------
Total 6 3 6 7 7 7 7 7 7 7 7
- -----------------------------------------------------------------------------------------------------------------
7.2 PROCESSING
The mill personnel are shown in Tables 7-6, and are compared to the
average for 2000. The milling of development ore will begin late in
year 2.
Table 7-6 - Mill & Assay Laboratory Personnel
- --------------------------------------------------------------------------------------------------------------------------
2000 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10
- --------------------------------------------------------------------------------------------------------------------------
Superintendent 1 0 1 1 1 1 1 1 1 1 1
General Foreman 1 0 1 1 1 1 1 1 1 1 1
Metallurgist 2 1 1 2 3 3 3 2 2 2 2
Technician 1 0 1 1 1 1 1 1 1 1 1
Environment 1 1 1 1 1 1 1 1 1 1 1
Electrical 1 1 2 2 2 2 2 2 2 2 2
Maintenance 5 1 4 5 5 5 5 5 5 5 5
Operators 8 0 8 9 13 13 13 13 13 13 13
Assay Lab. 3 0 2 3 3 3 3 3 3 3 3
- --------------------------------------------------------------------------------------------------------------------------
Total 23 4 21 25 30 30 30 29 29 29 29
- --------------------------------------------------------------------------------------------------------------------------
7.3 ADMINISTRATION AND TECHNICAL SERVICES
The administration and technical personnel are shown in Tables 7-7,
compared to the average for 2000. Two of the four positions in safety
are for certified trainers that will teach underground mining skills.
Table 7-7 - Administration and Technical Services
- ---------------------------------------------------------------------------------------------------------------------------
2000 Year 1 Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10
- ---------------------------------------------------------------------------------------------------------------------------
Administration 3 2 2 3 3 3 3 3 3 3 3
Purchasing 4 1 2 4 4 4 4 4 4 4 4
Human resources 2 1 2 2 2 2 2 2 2 2 2
Safety 2 1 3 4 4 4 4 4 4 4 4
Engineering 11 5 7 11 11 11 11 10 10 10 10
Geology 6 3 4 6 6 6 6 6 6 6 6
- ---------------------------------------------------------------------------------------------------------------------------
Total 28 13 20 30 30 30 30 29 29 29 29
- ---------------------------------------------------------------------------------------------------------------------------
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7.4 ORGANIZATION CHART
Figure 7-1 - Organization Chart
[PICTURE]
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7.5 SALARIES, WAGES AND BENEFITS
The wage scale in effect during the mining in year 2000 is shown in
Table 7-8. Salary information is included in Appendix W - Staff Budget
Salary.
Table 7-8 - Hourly Rate Employees Wage Scale
- -------------------------------------------------------------------------------
Classification Hourly rate
- -------------------------------------------------------------------------------
Unskilled labourer $16.10
Miner helper, Mill operator helper, 3rd class mechanic,
Timber man helper. $18.95
Heavy machinery operator, 3rd class electrician, Cage tender
$20.20
Timber man, 2nd class miner
1st miner, 2nd class mechanic, Mill operator $21.30
2nd class electrician, 1st class mechanic, Hoist man $22.15
1st class electrician, Mill operator leader $22.90
- -------------------------------------------------------------------------------
A provision has been included in the feasibility plan to pay hourly
mine employees a production bonus based on productivity, similar to
past practices.
The pay roll burden for social benefits averages 47% of the direct
wages and 45% for staff salaries. About half of the pay roll burden is
comprised of non-negotiable benefits (CSST, RAMQ, social insurance,
Quebec retirement plan, working norms). The other half includes all
the group insurance policies, vacation pay, sick leave, floating
holidays, stock option plan, pension plan, social club, transportation
and meal fees ($50/week- house or apt., $25/week - room housing
allowance).
The decision has been made to use the wage rates and burden levels
that were in place for the original (2001) feasibility study. Prior to
the operation returning to commercial production, however, the portion
of the collective agreement governing two key items (i.e. wages and
fringe benefits) must first be finalized and agreed with the USWA.
Regional employment market conditions at the time that these
negotiations are conducted will partially determine the outcome, but
for the purpose of the 2003 feasibility update it is believed to be
reasonable that these wage rates and burden levels are realistic and
attainable.
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8.0 PROJECT SCHEDULE
8.1 CONSTRUCTION PERIOD
There are construction projects that must be completed during the
pre-production period and during production start-up.
The major construction projects included in the schedule are:
o Modifications to the level 10 - 11 ore pass. This will be
converted into a 4.27m diameter steel lined bin.
o Ore pass lining work. A new steel-lined ore pass will be
built in Zone 3 up to level 8.
o Construction work on ore and waste pass dumping points.
o Rehabilitation of the existing underground mobile equipment
fleet.
o Construction of the upgraded ventilation system.
o Backfill systems including paste backfill pumping and
delivery on level 6 and cement slurry delivery to Zone 4.
o Development and construction of a new mobile maintenance
shop on level 13 and shop improvements on level 9.
o Installation of loading and material sizing facilities on
level 5.
The key aspects of the construction schedule for the mine start up
activities are:
o Completion of the new, steel lined ore storage bin between levels
10 and 11. This includes modifications to the ore feeder and
conveyor system below the bin on level 11.
o Steel lining a new ore passe that feeds the new ore bin up to
level 8.
o On level 9, the ore dump and waste dump are set up for rail cars.
The dumps must be modified to be suitable for truck dumping. This
must be done prior to development work on level 9.
o The level 13 waste dump must be set up for truck dumping prior to
the start of development on level 13.
o On level 5 the loading facilities must be commissioned prior to
resumption of production.
o Equipment refurbishing will start early in the schedule (mid-Year
1) to prepare the development fleet for use beginning early Year
2. The production equipment must be readied for January 2003.
There are 28 units to refurbish and this will take many weeks.
o Four jumbo crews will begin work in Year 2. Track must be
stripped on level 9 ahead of time for jumbo 4. Level 9 drift
slashing is followed by Zone 97 development. Jumbo 3 begins the
level 13 access drift toward Zone 97 in mid-Year 2.
o A track drift crew will drive the level 6 access to Zone 97
during Year 2. The Zone 97 paste backfill system requires a
pumping station on level 6. This will be installed once level 6
development is complete. Following the pump installation, holes
will be drilled from 6 level to Zone 97 below to distribute the
paste backfill. This can all be done before backfill is needed in
Zone 97.
o Start dates for stope production were modeled for stope
production in some of the ore zones so that the relative timing
of development advance and backfill system readiness could be
assessed.
o The Zone 4 cement slurry delivery/distribution system will be
installed on an upper level in the zone which is accessed near
the end of Zone 4 development. The schedule indicates there is
sufficient time to install the system prior to cemented rockfill
being needed in Zone 4.
o The timing of driving the new 3.66m diameter ventilation raise
depends on activities on level 9. Only after level 9 slashing is
complete will there be access to the pilot hole location on level
9. This dictates the timing of drilling the pilot hole.
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o The second leg of the new ventilation raise from level 9 to 13
can only be driven after level 13 development is completed.
o A small diameter fresh air raise will be driven from level 11 to
13 at the start of level 13 development. This raise will provide
the fresh air to the development face on level 13. It will also
serve as a future back up ore pass.
o The excavation for the new mobile maintenance shop on level 13
will be driven before development is started on the level.
Construction work to set up the shop follows after completion of
the excavation.
o Improvements are planned for the existing level 9 mobile
maintenance shop. These may be best implemented after the
development equipment has been refurbished.
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8.2 LIFE OF MINE PRODUCTION AND DEVELOPMENT
Table 8-1 is the life of mine production schedule showing the total
ore production which includes development ore and stope production
ore. All ore shown in Year 2 is from development in ore. Stope
production begins in Year 3.
Production schedule details are included in Appendix X - Production
Schedule Details:
o Development ore tonnes by year
o Stope production tonnes by year
o Ore grade information by domain.
Table 8-1 - Production Schedule
- ---------------------------------------------------------------------------------------------------------------------------------
Zone Domain Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Total
- ---------------------------------------------------------------------------------------------------------------------------------
Zone 3 Sup. West - 21,833 21,834 21,834 14,556 - - - - 80,057
Inf. West - 37,388 37,388 6,231 - - - - - 81,007
Inf. East - 58,821 58,821 58,821 58,821 58,821 58,821 58,821 14,705 426,449
Inf. East Ext. - - - - - - 21,945 21,945 - 43,890
Inf. Centre - 53,982 54,499 68,411 29,219 792 4,644 64,846 38,501 314,894
Development 24,295 21,735 1,652 550 771 771 1,102 881 110 51,870
SUBTOTAL (T) 24,295 193,759 174,194 155,847 103,367 60,384 86,512 146,493 53,316 998,167
Zone 4 Centre - 43,216 43,216 43,216 30,251 25,929 - - - 185,828
Development 18,115 - 397 397 397 198 - - - 19,504
SUBTOTAL (T) 18,115 43,216 43,613 43,613 30,648 26,127 - - - 205,332
Zone 97 9A - 37,222 37,222 37,222 37,222 - - - - 148,888
9B - 35,859 35,859 35,859 44,823 44,823 44,823 44,823 17,929 304,798
10A - - - - - 46,765 46,765 46,765 23,382 163,677
13A - - 37,633 37,633 37,633 37,633 37,633 37,633 15,053 240,851
13B - - 50,491 50,491 50,491 50,491 50,491 50,491 20,197 323,143
14A - - - - 37,960 75,920 75,920 75,920 37,960 303,680
Development 31,469 70,989 70,988 89,334 107,856 107,857 107,857 47,874 - 634,224
SUBTOTAL (T) 31,469 144,070 232,193 250,539 315,985 363,489 363,489 303,506 114,521 2,119,261
TOTAL TONNES 73,879 381,045 450,000 450,000 450,000 450,000 450,000 450,000 167,837 3,322,760
Grade Zinc (%) 11.29% 10.51% 10.78% 10.93% 10.97% 10.96% 10.73% 10.51% 10.59% 10.78%
Copper (%) 0.63% 0.63% 0.69% 0.73% 0.87% 0.97% 0.96% 0.89% 0.92% 0.82%
Gold (g/t) 0.09 0.08 0.08 0.08 0.09 0.09 0.09 0.08 0.08 0.09
Silver (g/t) 54.35 50.10 51.66 52.54 53.79 53.89 52.16 50.22 51.77 52.13
- ---------------------------------------------------------------------------------------------------------------------------------
Table 8-2 is the life of mine development schedule showing the total
lateral jumbo development requirements including ore and waste
advance. Total development is comprised of 52% ore advance and 48%
waste advance. Ore and waste development details are included in
Appendix X - Production Schedule Details
Development has been planned on the basis of four crews, each with a
set of equipment including an electric-hydraulic jumbo, scooptram,
truck and scissor lift. Each crew is capable of an average of 110m
advance per month. All four crews are scheduled to begin development
work in Year 2. A fifth crew is required during Year 2 to advance the
track drift on 6 Level. This drift will be advanced with the existing
track equipment
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Table 8-2 - Development Schedule
- ----------------------------------------------------------------------------------------------------------------------------------
Zone Domain Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Total
- ----------------------------------------------------------------------------------------------------------------------------------
Waste Development (m) - Jumbo
Zone 3 Sup. West 80 - - - - - - - - 80
Inf. West - - - - - - - - - -
Inf. East 437 255 - - - - - - - 692
Inf. East Ext. - - - - - - - - - -
Inf. Centre 280 167 - - - - - - - 447
SUBTOTAL (M) 797 422 - - - - - - - 1,219
Zone 4 Centre 817 - - - - - - - - 817
SUBTOTAL (M) 817 - - - - - - - - 817
Zone 97 6 Lvl. Track Dr. 1,283 - - - - - - - - 1,283
9A 306 274 267 224 - - - - - 1,071
9B 390 271 264 264 178 278 347 535 - 2,527
10A - 152 152 152 110 210 190 - - 966
13A 770 1,200 408 451 274 267 336 118 - 3,824
13B - 310 358 267 447 329 453 - - 2,164
14A - - 507 507 275 165 110 110 - 1,674
SUBTOTAL (M) 2,749 2,207 1,956 1,865 1,284 1,249 1,436 763 - 13,509
TOTAL METRES 4,363 2,629 1,956 1,865 1,284 1,249 1,436 763 - 15,545
Ore Development (m) - Jumbo
Zone 3 Sup. West - - - - - - - - - -
Inf. West - - - - - - - - - -
Inf. East 467 224 - - - - - - - 691
Inf. East Ext. - - - - - - - - - -
Inf. Centre 137 274 - - - - - - - 411
SUBTOTAL (M) 604 498 - - - - - - - 1,102
Tonnes 24,295 21,735 1,652 550 771 771 1,102 881 110 51,870
Zone 4 Centre 450 - - - - - - - - 450
SUBTOTAL (M) 450 - - - - - - - - 450
Tonnes 18,115 - 397 397 397 198 - - - 19,504
Zone 97 9A 355 355 355 355 - - - - - 1,420
9B 409 409 409 511 511 511 511 204 - 3,475
10A - - - - 464 464 464 232 - 1,624
13A - 492 492 492 492 492 492 197 - 3,149
13B - 468 468 468 468 468 468 187 - 2,995
14A - - - 345 690 690 690 345 - 2,760
SUBTOTAL (M) 764 1,724 1,724 2,171 2,625 2,625 2,625 1,165 - 15,423
Tonnes 31,469 70,989 70,988 89,334 107,856 107,857 107,857 47,874 - 634,224
Total Metres 1,818 2,222 1,724 2,171 2,625 2,625 2,625 1,165 - 16,975
Tonnes Tonnes 73,879 92,724 73,037 90,281 109,024 108,826 108,959 48,755 110 705,595
Total Waste & Ore Metres 6,640 4,851 3,680 4,036 3,909 3,874 4,061 1,928 - 32,520
- ----------------------------------------------------------------------------------------------------------------------------------
Note: Ore Development Includes Slot Raises
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9.0 CONCENTRATE SHIPMENT AND MARKETING
9.1 CONCENTRATE SPECIFICATIONS
9.1.1 CONCENTRATE QUALITY
The zinc concentrate has an average grade of 54.8% Zn and the copper
concentrate has an average grade of 24.0% Cu. The zinc concentrate has
9% iron which is penalised (US$2.00 per tonne of zinc concentrate) and
the copper concentrate pays a penalty on lead (US$0.36 per tonne of
copper concentrate). Moisture content is 7.50% for both the zinc and
copper concentrate.
9.2 CONCENTRATE SHIPMENT
Zinc concentrate would be loaded on CN railcars directly at the mine
site and transported to Noranda's CEZ smelter in Valleyfield or to the
port of Montreal. The distance to Valleyfield is 850km and the cost is
$41.20 per wet tonne.
Copper concentrate would be sent by rail to Noranda's Horne smelter in
Rouyn-Noranda. The distance to the Horne smelter is 300km and the cost
is $17.00 per dry tonne.
Table 9-1 - Shipping Cost to Smelter
($000's)
- -----------------------------------------------------------------------------------------------------------------
Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total
- -----------------------------------------------------------------------------------------------------------------
Zinc Concentrate 672 3,220 3,833 3,844 3,801 3,763 3,657 3,578 1,344 27,712
- -----------------------------------------------------------------------------------------------------------------
Copper Concentrate 28 142 185 195 232 259 250 226 88 1,605
- -----------------------------------------------------------------------------------------------------------------
Total 700 3,362 4,018 4,039 4,033 4,022 3,907 3,804 1,432 29,317
- -----------------------------------------------------------------------------------------------------------------
* 6 months
9.3 CONCENTRATE TERMS
Zinc and copper concentrates are sold to Noranda on terms and
conditions, which mirror the annual benchmarks in the industry. In
addition Noranda is paid a freight capture premium which represents a
sharing of Breakwater's savings on transportation incurred by selling
to Noranda rather than to Europe.
Breakwater conservatively forecasts that long-term smelter treatment
charges are US$160 per dry metric tonne of concentrate on a basis
US$1,000 per tonne zinc. The price participation is plus 16 cents /
minus 14 cents per dollar change in zinc metal price on a per tonne
basis. The freight capture is estimated to be between US$10.00 -
$20.00 per dmt.
Breakwater conservatively forecasts the copper treatment charges at
US$80 per dry metric tonne of concentrate on a basis of US$0.90 per
pound copper and the copper refining charge at US$0.08 per pound of
payable copper. The freight capture is estimated to be US$20.00 to
US$30.00 per dmt.
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Table 9-2 - Concentrate Specifications
- ---------------------------------------------------------------------------------------------------
Zinc Concentrate Copper Concentrate
- ---------------------------------------------------------------------------------------------------
Zn % 52.00 Cu % 24-28.8
Pb % 0.32 As % 0.003-0.004
Cu % 0.63 Au g/t Up to 8.2
Fe % 9-10 Ag g/t 450 - 900
As % 0.01 Sb % 0.007
Sb % 0.0009 Bi % 0.02 - 0.04
Au g/t 0.30 S % 34.3 - 36.2
Ag g/t 70 Hg ppm 2 - 5
F ppm 10 Fe % 27.3 - 29.1
Cl ppm 100 Cd ppm 92 - 110
S % 35.00 Pb % 1.19 - 4.87
Ge % 0.001 Al2O3 % 0.04 - 0.06
Se % 0.02 Zn % 3.91 - 4.96
Sn % 0.02 F ppm 0 - 56
Te % - SiO2 % 1.07 - 1.30
Hg % 0.0022 Co ppm 6.3 - 15
Bi % 0.005 Se % <0.044
Ni % 0.003 Te ppm < 3
Co % 0.002 Ni ppm 9 - 480
Cd % 0.19 Mn % 26 - 31
MgO % 0.08 CaO % 0.03
Mn % 0.019 Sn ppm 30 - 320
Al2O3 % 0.08 MgO % 0.12 - 0.45
CaO % 0.13 Cl ppm < 100
SiO2 % 0.25 Mo % -
Insols % -
In % -
U % -
Ge % -
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10.0 OPERATING COST
The total operating cost to mine gate is estimated at Cdn$55.61 per
tonne milled over the life of the mine. The average operating cost
from Year 4 to Year 9 is $51.73/tonne. This cost is in the range of
the historical 1999 operating cost. Table 10-1 summarizes the total
operating cost. Note that the cost per lb payable zinc is based on the
operating costs tabulated, plus the smelter treatment costs, not show
in the table. The following sections explain the operating costs,
while further details are in Appendix Y - Unit Costs.
As part of the current feasibility study update, suppliers were
contacted for updated material costs for 2003. A second version of the
project economic model was prepared using current 2003 prices, and it
was determined that there had not been a significant change, so the
original 2001 material costs have been used in this current study.
Table 10-1 - Operating Cost
($000's)
- ------------------------------------------------------------------------------------------------------------------------------------
Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total Cost/t
- ------------------------------------------------------------------------------------------------------------------------------------
Mining 2,358 7,760 14,467 15,363 15,720 16,143 16,064 15,951 14,954 5,483 124,263 $37.40
Milling 264 858 4,605 5,160 5,061 5,116 5,098 5,029 4,794 1,925 37,910 $11.41
Administration 950 2,352 2,534 2,534 2,534 2,534 2,534 2,534 2,534 1,546 22,586 $6.80
Total Operating 3,572 10,970 21,606 23,057 23,315 23,793 23,696 23,514 22,282 8,954 184,759 $55.61
Total Operating/tonne Milled - $148.49 $56.70 $51.24 $51.81 $52.87 $52.66 $52.25 $49.52 $53.34 $55.61
Cost/lb Payable Zinc US$ - - $0.408 $0.373 $0.368 $0.359 $0.350 $0.350 $0.345 $0.357 $0.376
- ------------------------------------------------------------------------------------------------------------------------------------
* 6 months
Table 10-2 provides a more detailed breakdown of the mining cost, and
makes a comparison to the average operating costs in the feasibility
plan to the actual costs from 1999 and 2000. In addition, the
feasibility operating cost for Year 6 has been included, representing
the typical costs at full production.
Table 10-2 - Operating Cost Comparison
($/tonne)
--------------------------------------------------------------------------------------------
HISTORIC HISTORIC HISTORIC FEASIBILITY LIFE OF MINE
ACTUAL ACTUAL AVERAGE OF PLAN FEASIBILITY
1999 2000 1999 & 2000 YEAR 6 PLAN
--------------------------------------------------------------------------------------------
Definition Drilling 1.31 0.57 0.94 0.67 0.61
Stope Preparation 7.82 7.12 7.47 6.15 5.44
Extraction 14.06 16.66 15.36 13.16 13.50
Mine Services 6.28 6.40 6.34 5.84 6.52
Mechanical Services 2.56 2.45 2.51 2.87 3.41
Electrical Services 1.19 1.08 1.14 1.05 1.22
Surface Services 3.69 4.09 3.89 4.26 5.20
Engineering 1.43 1.74 1.59 1.74 1.90
Geology 1.11 1.18 1.15 1.11 1.21
Allocation to Capital (2.60) (1.72) (2.16) (0.98) (1.61)
Total Mining 36.85 39.57 38.21 35.87 37.40
Milling 9.69 10.52 10.11 11.37 11.41
Administration 5.20 6.24 5.72 5.63 6.80
--------------------------------------------------------------------------------------------
TOTAL $51.74 $56.33 $54.04 $52.87 $55.61
--------------------------------------------------------------------------------------------
COST/LB PAYABLE ZINC $0.490 $0.520 $0.500 $0.359 $0.376
US$
--------------------------------------------------------------------------------------------
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The cost per tonne milled in the feasibility plan is higher than the
average 1999-2000 cost, due partly to the effect of averaging in costs
from Year 1, 2 and 10, when the production rate is well below design.
Many of the cost categories shown above compare much more closely in
Year 6 during full production. On a cost per pound payable zinc basis,
the feasibility plan has a reduced cost due to mining higher grade
material.
Stope preparation includes development and slot raising. The
feasibility cost of stope preparation is lower than historical costs
as the practice of cable bolting the ore drifts will be replaced with
lower cost rebar bolting. Also, the feasibility unit cost for drifting
is lower than historical costs as accounting practices have changed
regarding the service equipment. In the past it was charged to
drifting under stope preparation, while the current practice is to
include these charges under mechanical services resulting in a higher
cost in this category.
Extraction includes drilling, blasting, mucking, backfilling and drift
rehabilitation.
Mine services include supervision, shaft operations, crushing and
construction.
Surface services includes electric power, propane, diesel fuel,
bussing, road and yard maintenance and waste handling. Energy costs
have increased since 2000 and more surface waste handling is included
in the feasibility plan. The feasibility plan also includes an
increase in the ventilation requirement for the mine which results in
additional electricity consumption and propane consumption for mine
air heating.
Feasibility plan geology and engineering costs are slightly higher
than historical costs as additional personnel will be required due to
the increased amount of development faces related to Zone 97 as well
as longer distances to cover.
Feasibility plan milling costs are higher than historical costs due to
additional reagent consumption due to milling higher grade ore.
Feasibility plan administration costs are the average of historical
costs. An amount of $0.5 million has been included for Langlois
severance costs in Year 10.
10.1 MINING COST
The following table is a summary by year of each major area of the
underground mine operating cost.
Table 10-3 - Mining Cost
($000's)
- ----------------------------------------------------------------------------------------------------------------------------------
Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total Cost/t
- ----------------------------------------------------------------------------------------------------------------------------------
Definition Drilling - 50 321 379 303 303 303 152 152 57 2,020 $0.61
Stope Preparation - 1,685 2,449 2,051 2,407 2,766 2,725 2,709 1,263 20 18,075 $5.44
Extraction - 843 5,064 6,037 5,960 5,923 6,002 6,094 6,376 2,553 44,852 $13.50
Mine Services 625 1,824 2,630 2,630 2,630 2,630 2,541 2,541 2,541 1,085 21,677 $6.52
Mechanical Services 400 1,203 1,279 1,247 1,312 1,291 1,337 1,319 1,375 573 11,336 $3.41
Electrical Services 115 473 473 473 473 473 473 473 473 145 4,044 $1.22
Surface Services 878 2,062 2,090 1,956 1,936 1,919 1,894 1,917 1,864 777 17,293 $5.20
Engineering 215 621 781 781 781 781 712 712 712 201 6,297 $1.90
Geology 125 336 499 499 499 499 499 499 499 72 4,026 $1.21
Allocation to Capital - (1,337) (1,119) (690) (581) (442) (422) (465) (301) - (5,357) ($1.61)
- ----------------------------------------------------------------------------------------------------------------------------------
Total 2,358 7,760 14,467 15,363 15,720 16,143 16,064 15,951 14,954 5,483 124,263 $37.40
- ----------------------------------------------------------------------------------------------------------------------------------
* 6 months
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10.1.1 DEFINITION DRILLING
The feasibility study includes definition drilling during Year 3 and
Year 4 in order to better define the ore in Zone 97. Other areas will
also be drilled, but the drilling quantity drops off in Year 8, as all
areas will be sufficiently defined. The lower interval development
planned in Zone 97 will allow substantial savings in definition
drilling at $0.61/t compared to historical costs of $1.00/t hoisted.
Since 1993, 123,512m of definition and exploration drilling have been
carried out at the Langlois mine at an average cost of $37.31/m,
including assaying and cementing. For the year 2000, the average cost
was $34.94/m.
10.1.2 STOPE PREPARATION
Table 10-4 shows the quantities of work and the associated costs for
stope preparation.
Table 10-4 - Stope Preparation
($ 000)
- -----------------------------------------------------------------------------------------------------------------------------
Cost/m Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total
- -----------------------------------------------------------------------------------------------------------------------------
Quantity
Drifts in Ore (m) - 1,818 2,222 1,724 2,171 2,625 2,625 2,625 1,165 - 16,975
Open Raises (m) - - 135 90 - - - - - - 225
Drop Raises (m) - 64 164 240 262 306 288 300 168 10 1,802
Raclette Slots (m) - - 180 300 400 235 170 180 160 30 1,655
Styrofoam Slots (m) - 64 324 288 320 290 290 200 88 - 1,864
Cost
Drifts in Ore $887/m - 1,613 1,971 1,529 1,926 2,328 2,329 2,327 1,034 - 15,057
Open Raises $1,225/m - - 165 110 - - - - - - 275
Drop Raises $800/m - 51 131 192 210 245 230 241 134 8 1,442
Raclette Slots $413/m - - 74 124 165 97 70 75 66 12 683
Styrofoam Slots $332/m - 21 108 96 106 96 96 66 29 - 618
Total - 1,685 2,449 2,051 2,407 2,766 2,725 2,709 1,263 20 18,075
- -----------------------------------------------------------------------------------------------------------------------------
* 6 months
Drifts in ore cost $887/m. This can be further broken down to $423/m
for labour, $253/m for materials, and $211/m for equipment. (As
previously noted, material costs are based on 2001 prices.)
The labour cost is based on two 3 man crews working 3 headings at the
same time on a 5 day per week, 2 8-hour shifts per day basis.
Historically, labour costs were $410/m. Material cost includes ground
support, piping, ventilation, etc. Historically it was $245/m.
Historically, equipment cost was $480/m. The difference between
historical and planned equipment cost reflects a change in accounting
practice. In the past the cost for 2 yd. service scoop and service
vehicles were redistributed in the development account, from now on
they will be redistributed in mech. dept. account.
Open raises cost $1,225/m, drop raises $800/m, raclette slots $413/m,
and styrofoam slots $332/m. These costs include manpower and
materials. Refer to Appendix Z - Historical Costs.
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10.1.3 EXTRACTION
Table 10-5 shows the quantities of work and the associated costs for
extraction.
Table 10-5 - Extraction ($ 000)
- ------------------------------------------------------------------------------------------------------------------------------------
Cost/t Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Total
- ------------------------------------------------------------------------------------------------------------------------------------
UNITS - TONNES
Longhole Drilling - 114mm - - 134,636 135,153 149,065 102,596 59,612 63,465 123,667 53,206 821,400
Longhole Drilling -54mm Z3,4 - - 80,604 80,604 49,447 30,251 25,929 21,945 21,945 - 310,726
Longhole Drilling - 54mm Z97 - - 73,080 161,205 161,205 208,130 255,633 255,633 255,633 114,522 1,485,040
Blasting 114mm - - 134,636 135,153 149,065 102,596 59,612 63,465 123,667 53,206 821,400
Blasting 54mm Z3, 4 - - 80,604 80,604 49,447 30,251 25,929 21,945 21,945 - 310,726
Blasting 54mm Z97 - - 73,080 161,205 161,205 208,130 255,633 255,633 255,633 114,522 1,485,040
Mucking - - 288,321 376,962 359,717 340,976 341,175 341,043 401,245 167,727 2,617,165
Remucking - Truck Z4, 97 - 31,468 144,069 232,194 250,540 315,985 363,489 363,489 303,507 114,522 2,119,263
Remucking - Truck Z3 Est 18,221 99,788 85,485 95,837 66,608 45,288 48,094 93,081 - 552,401
Remucking - Scooptram - 1,573 7,203 11,610 12,527 15,799 18,174 18,174 15,175 5,726 105,963
Load Waste to Trucks 76,021 91,919 55,120 47,482 18,772 11,272 18,862 5,860 2,133 327,441
Backfill - Paste - 30,440 187,396 221,657 220,483 223,840 223,430 238,637 242,476 90,352 1,678,711
Backfill - Rock - - 24,510 33,375 31,650 31,073 31,525 34,104 40,124 16,773 243,134
Backfill - Cemented Rock - - 20,467 20,655 20,655 14,515 12,374 - - - 88,665
COST ($ 000) Cost/t
Longhole Drilling - 114mm $2.16 - 291 291 322 222 129 137 267 115 1,774
Longhole Drilling -54mm Z3,4 $1.83 - 148 148 90 55 47 40 40 - 568
Longhole Drilling -54mm Z97 $2.08 - 152 335 335 433 532 532 532 238 3,089
Blasting 114mm $1.18 - 159 160 176 121 70 75 146 63 970
Blasting 54mm Z3,4 $1.86 - 150 150 92 56 48 41 41 - 578
Blasting 54mm Z97 $1.98 - 145 319 319 412 506 506 506 227 2,940
Mucking $2.15 - 620 811 773 733 734 733 863 361 5,628
Remucking - Truck Z4, 97 $2.21 71 318 513 554 698 803 803 671 253 4,684
Remucking - Truck Z3 Est $1.61 29 161 138 154 107 73 77 150 - 889
Remucking - Scooptram $215 3 15 25 28 34 39 39 33 12 228
Load Waste to Trucks $2.15 163 197 119 102 40 24 41 13 5 704
Backfill - Paste $8.78 267 1,645 1,946 1,936 1,966 1,962 2,095 2,129 793 14,739
Backfill - Rock $2.15 - 53 72 68 67 68 73 86 36 523
Backfill - Cemented Rock $5.32 - 109 110 110 78 66 - - - 473
Drift Rehabilitation - 282 845 844 845 845 845 846 844 422 6,618
Ore Pass - 28 56 56 56 56 56 56 55 28 447
- ------------------------------------------------------------------------------------------------------------------------------------
TOTAL 843 5,064 6,037 5,960 5,923 6,002 6,094 6,376 2,553 44,852
- ------------------------------------------------------------------------------------------------------------------------------------
Long hole drilling (114mm diameter) is estimated to cost $2.16/t or
$20.54/m, including manpower at $8.25/m, materials at $7.79/m, and
equipment at $4.50/m. This assumes a long hole driller will average
50m/shift working five days a week, two shifts per day.
Long hole drilling (54mm diameter) in Zones 3 and 4 is estimated to
cost $1.83/t or $10.99/m, including manpower at $6.17/m, materials at
$1.99/m, and equipment at $2.83/m. Long hole drilling (54mm) in Zone
97 is estimated to cost more due to the narrower ore at $2.08/t or
$10.40/m, including manpower at $5.57/m, materials at $1.99/m, and
equipment at $2.83/m. This assumes that a long hole driller will
average 65m/shift working five days a week, two shifts per day.
Blasting 114mm diameter holes is estimated to cost $1.18/t , while
blasting 54mm diameter holes in Zones 3, 4 is estimated to cost
$1.85/t and in Zone 97 it is estimated to cost $1.98/t.
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95 SRK CONSULTING
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
All longhole blasting will be carried out by 2 crews of 2 men working
five days a week, two shifts per day (160 man shifts/week). Refer to
Appendix Y - Unit Costs for further details.
Mucking and trucking operating costs per hour have been estimated from
basics, starting with tires, fuel, maintenance labour and parts, etc.
Haulage distances have been evaluated to estimate cycle times and
productivities. Costs per tonne have been derived from these
productivities.
Mucking and scooptram rehandling will cost $2.15/t, while trucking
will cost $2.21/t. Trucking costs for Zone 3 material have been
estimated at $1.61/t.
Paste backfilling will cost $8.78/t, un-cemented rock backfilling will
cost $2.08/t, and cemented rock backfilling will cost $5.32/t.
Rehabilitation costs are based on 3 men per shift scaling, spot
bolting, and screening drifts where needed.
Historical ore pass operating costs have been $10,000/month. Since
there are no ore/waste passes in Zone 97 (truck haulage), the future
projection for this account has been reduced by about 50% to
$4,700/month to include maintenance of ore passes in Zones 3 and 4.
10.1.4 MINE SERVICES
Mine services costs include all aspects of shaft operation,
maintenance of some underground infrastructure (lunch rooms, shaft
stations, tuggers, etc.), crusher and rock breaker operation, service
men, training, supervision and operating costs for four pick-up
trucks. Further details are in Appendix AA - Underground Department.
10.1.5 MECHANICAL SERVICES
Mechanical services costs include manpower and materials for
maintaining the mine buildings, underground garages and mobile
equipment, compressors, generator, hoisting plant, shaft loading
system, surface ventilation equipment, and surface mobile equipment.
Also included are supervision salaries, one pick-up truck and various
expenses (small tools and tools replacement, safety equipment).
Further details are in Appendix BB - Maintenance Department.
10.1.6 ELECTRICAL SERVICES
Electrical services costs include manpower and materials for
maintaining the mine buildings, underground mobile equipment,
communication and automation systems, heating system, compressors,
generator, hoisting plant, miner's lamps, shaft loading system,
ventilation equipment, etc. Also included are supervision salaries,
one pick-up truck and various expenses (small tools and tools
replacement, electrical equipment, safety equipment). Further details
are in Appendix CC - Electrical Department.
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96 SRK CONSULTING
JUNE 2003
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
10.1.7 SURFACE SERVICES
Table 10-6 - Surface Services
($000's)
- ------------------------------------------------------------------------------------------------------------------------------------
Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total Cost/t
- ------------------------------------------------------------------------------------------------------------------------------------
FIXED COSTS 80 249 248 248 249 249 249 249 249 104 2,174 $0.65
VARIABLE COSTS
Surface Waste Haulage - 442 218 81 65 19 19 50 - - 894 $0.27
Electricity 540 796 936 936 936 960 936 936 936 390 8,302 $2.50
Propane 60 159 254 254 253 254 254 254 253 106 2,101 $0.63
Diesel Fuel - 19 38 41 38 41 40 32 30 12 291 $0.09
Transport (Bussing) 198 397 396 396 395 396 396 396 396 165 3,531 $1.06
- ------------------------------------------------------------------------------------------------------------------------------------
Subtotal 798 1,813 1,842 1,708 1,687 1,670 1,645 1,668 1,615 673 15,119 $4.55
- ------------------------------------------------------------------------------------------------------------------------------------
TOTAL 878 2,062 2,090 1,956 1,936 1,919 1,894 1,917 1,864 777 17,293 $5.20
- ------------------------------------------------------------------------------------------------------------------------------------
* 6 MONTHS
Fixed costs in surface services include manpower and material for the
dry, the yard and the waste handling. It also includes the contract to
maintain the mine site access road. Further details are in Appendix DD
- Fixed Costs.
The Langlois mine is a significant consumer of electricity. A study
was conducted which analyzed the annual electrical power cost by
examining the utilization hours for all of the electrical equipment.
Using this data, projections were made of future electrical power
consumption. In accordance with the Hydro-Quebec pricing rate
effective in May 2001, the unit cost per kilowatt-hour is estimated to
be $0.0462. The underground cost of electricity has been estimated to
be $8,302,000 for the life of the mine. See Appendix EE - Electric
Power Cost.
Propane is used to heat the fresh air during cold periods. To evaluate
the annual propane cost, an analysis of the historical propane
consumption in relation to the flow of heated air was conducted.
Utilizing data from the winters of 1998-1999 and 1999-2000, the
average consumption of propane was estimated to be 3.8 litres per
cubic foot per minute (cfm) of fresh air. See Appendix FF -Propane
Cost.
The total propane cost for the project is $2,101,000, considering the
planned increase in ventilation requirements, and a fixed price of
$0.192 per litre of propane (1999-2000 price).
A diesel fuel price of $0.4838 per litre (the average price of 1999
and 2000) is included in the estimation of the unit costs (mucking,
trucking, etc.). To account for the recent rise in the diesel fuel
price, an extra $0.0484 per litre has been accrued and added to the
surface service cost account. This addition represents an increase of
10% and brings the estimation of the diesel fuel price to $0.5322 per
litre for the whole life of the project. See Appendix GG - Diesel Fuel
Cost.
Mine employees are transported to the mine from the town of
Lebel-sur-Que Villon by bus and this cost is included in surface
services. Further details are in Appendix HH - Bussing Contract.
10.1.8 ENGINEERING
Engineering costs include salaries for (1 Chief Engineer, 1 Senior
Engineer, 2 Engineers, 1 Senior Technician, and 6 Technicians),
expenses for one pick-up truck and various expenses (small tools,
computer, safety equipment). Further details are in Appendix JJ -
Engineering Department.
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
10.1.9 GEOLOGY
Geology costs include salaries for (1 Chief Geologist, 1 Senior
Geologist, 1 Geologist, and 3 Technicians), and various expenses (lab.
analyses, small tools, safety equipment). Further details are in
Appendix KK - Geology Department.
10.1.10 ALLOCATION TO CAPITAL
Allocation to capital is a transfer of operating costs (portions of
the supervision and engineering salaries of people attributed to
capital expenditures) to the capital cost. See Appendix LL -Virement
10.4.11 WORKING CAPITAL
Working capital requirements have been estimated and are shown in the
table below.
Table 10-7 - Working Capital
($000's)
- ---------------------------------------------------------------------------------------------------------------------
Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10 Total
- ---------------------------------------------------------------------------------------------------------------------
Working Capital 1,278 4,876 1,529 221 315 199 (95) (224) (8,099) -
- ---------------------------------------------------------------------------------------------------------------------
* 6 months
Working capital takes into account the time difference between
concentrate sales and revenue. It has been calculated by taking the
annual minesite revenue divided by 5.
10.2 MILLING COST
The following table is a summary by major area of the milling and
environmental operating costs.
Table 10-8 - Milling & Environmental Operating Cost
($000's)
- ------------------------------------------------------------------------------------------------------------------------------
Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total Cost/t
- ------------------------------------------------------------------------------------------------------------------------------
Milling 204 731 4,371 4,926 4,827 4,882 4,864 4,795 4,560 1,827 35,987 $10.83
Environmental 60 127 234 234 234 234 234 234 234 98 1,923 $0.58
- ------------------------------------------------------------------------------------------------------------------------------
Total 264 858 4,605 5,160 5,061 5,116 5,098 5,029 4,794 1,925 37,910 $11.41
- ------------------------------------------------------------------------------------------------------------------------------
* 6 months
Milling cost includes manpower, reagents, maintenance, and
electricity. See Appendix MM - Mill Budget.
Environment costs include manpower, reagents, and maintenance. See
Appendix NN -Environmental Budget.
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98 SRK CONSULTING
JUNE 2003
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
10.3 ADMINISTRATION COST
Table 10-9 - Administration Cost
($000's)
- -----------------------------------------------------------------------------------------------------------------------------------
Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total Cost/t
- -----------------------------------------------------------------------------------------------------------------------------------
Site Administration 818 1,887 2,069 2,069 2,069 2,069 2,069 2,069 2,069 1,352 18,540 $5.58
- -----------------------------------------------------------------------------------------------------------------------------------
Administrative Services 132 465 465 465 465 465 465 465 465 194 4,046 $1.22
- -----------------------------------------------------------------------------------------------------------------------------------
Total 950 2,352 2,534 2,534 2,534 2,534 2,534 2,534 2,534 1,546 22,586 $6.80
- -----------------------------------------------------------------------------------------------------------------------------------
* 6 months
10.3.1 SITE ADMINISTRATION
Site administration costs include salaries for (1 General Manager, 1
General Superintendent, 1 Human Resources Superintendent, 1 Safety
Co-ordinator, 1 Human Resources Co-ordinator, 2 Trainers, 1 Nurse, 1
Secretary, 1 Purchasing Agent, 3 Warehouse Clerks), expenses for two
pickup trucks, housing expenses and apartments revenue, expenses for
security guards, association fees, professional fees, recruiting
costs, employees help program, mining rights permit, general
insurance, municipal taxes, school taxes, phone fees, postal service
fees and various expenses (small tools, computer, safety equipment).
See Appendix OO - Administration Department.
10.3.2 ADMINISTRATIVE SERVICES
Administrative services cost includes salaries for (1 Administration
Director, 1 Accountant, 2 Clerks, and 1 Purchasing Agent). Also
included are the costs for the computer system. See Appendix PP - Off
Site Administration.
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99 SRK CONSULTING
JUNE 2003
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
11.0 CAPITAL COST
An estimated $38.2 million in capital is required over the life of the
mine. This is comprised of $47.1 million in expenditures offset by a
salvage value of $8.9 million at the end of the project. Approximately
$16.4 million must be expended prior to the start of production at the
beginning of Year 3. The majority of the capital requirements are
related to the underground mine, with the mill accounting for only
$0.8million of the total.
The project capital requirements are presented by year in Table 11-1
in these categories:
o Deferred development
o Equipment
o Salvage value
Deferred development includes development of the underground
infrastructure and construction activities. Equipment costs are mainly
for the purchase of new underground and surface equipment, including
stationary equipment and mobile equipment. A small portion of
equipment cost is for the repair of existing equipment. Salvage value
is credited to the project in year 2009 and includes the residual
value of all buildings and equipment
There is no capital provision for exploration activities aimed at
defining new resources and reserves. The feasibility study is based
only on the currently defined reserves
Table 11-1 Capital
($ 000)
- -----------------------------------------------------------------------------------------------------------------------------
Item Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total
- -----------------------------------------------------------------------------------------------------------------------------
Deferred Development 4,268 6,717 6,302 3,445 2,845 2,124 2,065 2,293 1,261 125 31,445
Infrastructure & Equipment 340 5,044 4,152 2,387 1,100 1,260 725 500 150 - 15,658
Salvage Value - - - - - - - - - (8,935) (8,935)
- -----------------------------------------------------------------------------------------------------------------------------
Total 4,608 11,761 10,454 5,832 3,945 3,384 2,790 2,793 1,411 (8,810) 38,168
- -----------------------------------------------------------------------------------------------------------------------------
* 6 months
11.1 DEFERRED DEVELOPMENT
The deferred development capital estimates are shown in Table 11-2.
The allocation from operating is a transfer of some operating costs to
capital. These transferred costs include portions salaries of
personnel supporting project work such as supervision, engineering,
etc. The development and construction components are described below.
Table 11-2 - Deferred Development Summary
($ 000)
- ------------------------------------------------------------------------------------------------------------------------------
Item Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total
- ------------------------------------------------------------------------------------------------------------------------------
Development - 4,767 4,588 2,255 1,763 1,181 1,143 1,328 710 - 17,735
Construction 4,268 613 596 500 500 500 500 500 250 125 8,352
Allocation from Operating - 1,337 1,118 690 582 443 422 465 301 - 5,358
- ------------------------------------------------------------------------------------------------------------------------------
Subtotal 4,268 6,717 6,302 3,445 2,845 2,124 2,065 2,293 1,261 125 31,445
- ------------------------------------------------------------------------------------------------------------------------------
* 6 months
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100 SRK CONSULTING
JUNE 2003
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
11.1.1 DEVELOPMENT
Development requirements have been conceptually planned for the life
of the mine. Capitalized underground development generally includes
most of the excavations in waste. These excavations include access
ramps, main access/haulage drifts, waste cross cuts, ore and waste
passes, and ventilation raises. Quantities and cost estimates for
capital development are shown in Table 11-3.
Most of this development is related to accessing Zone 97 and
supporting the ongoing production there. Work is required on levels 6,
9 and 13 to create suitable access drifts to this new zone. Also,
considerable waste development is needed to drive the ramp systems and
crosscuts to the sublevels that are closely spaced at 11m.
Table 11-3 - Deferred Development - Excavation
- ----------------------------------------------------------------------------------------------------------------------------------
Cost/m Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total
- ----------------------------------------------------------------------------------------------------------------------------------
EXCAVATION
Drifts (m) - 1,535 835 - - - - - - - 2,370
Crosscuts (m) - 1,184 774 600 495 715 760 715 330 - 5,572
Ramps (m) - 1,126 1,020 1,356 1,370 569 489 721 433 - 7,084
6 Lvl Track Drift - 1,283 - - - - - - - - 1,283
Open Raises (m) - - 80 - - - - - - - 80
1.8m Raisebore (m) - - 140 280 - - - - - - 420
3.7m Raisebore (m) - - 780 - - - - - - - 780
COST ($ 000)
Drifts $880/m - 1,351 735 - - - - - - - 2,086
Crosscuts $880/m - 1,042 681 529 436 630 669 630 291 - 4,908
Ramps $968/m - 1,091 988 1,313 1,327 551 474 698 419 - 6,861
6 Lvl Track Drift $1,000/m - 1,283 - - - - - - - - 1,283
Open Raises $1,225/m - - 98 - - - - - - - 98
1.8m Raisebore $1,476/m - - 207 413 - - - - - - 620
3.7m Raisebore $2,409/m - - 1,879 - - - - - - - 1,879
- ----------------------------------------------------------------------------------------------------------------------------------
Total - 4,767 4,588 2,255 1,763 1,181 1,143 1,328 710 - 17,735
- ----------------------------------------------------------------------------------------------------------------------------------
* 6 months
At the beginning of Year 2, crews are scheduled to begin the lateral
development work. An additional jumbo crew commences in mid Year 2.
Each crew is capable of advancing 110m per month. The largest item
under raising is the 3.7m diameter ventilation raise for zone 97,
scheduled the start in mid Year 2.
The track drift on level 6 will also start at the beginning of Year 2.
The heading will have to be driven on three shifts per day assuming an
advance rate of 110m per month.
The drifts and crosscuts sized at 3.5 x 3.7m will cost $880/m. This
breaks down to $423 for labour, $246 materials, and $211 equipment
costs. Ramping costs are estimated at $968/m, 10% more than level
development.
Quotations from a raise boring contractor are $1,476/m for 1.8m
diameter completed raise, and $2,409/m for 3.7m diameter. See Appendix
QQ - Raise Boring Estimates.
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101 SRK CONSULTING
JUNE 2003
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
11.1.2 CONSTRUCTION
The construction cost estimate has been detailed for Years 1 and 2.
Allowances for construction have been made in ensuing years. Refer to
Table 11-2.
The major construction projects included in the capital cost estimate
are:
o Modifications to the level 10 - 11 ore pass and
feeder/conveyor. This will be converted into a 4.3m diameter
steel lined bin. ($1,895,000)
o The ore pass system will consist of 3.0m diameter vertical
raise lined with steel from level 10 to level 8.
($1,933,000)
o Construction work on ore and waste pass dumping points. o
Installation of Zone 97 paste backfill underground
distribution system.
o Expansion of the electrical distribution system.
o Fuel line from surface to underground.
o Development and construction of a new mobile maintenance
shop on level 13 and shop improvements on level 9.
11.2 EQUIPMENT
The capital cost estimate for equipment is shown in Table 11-4. During
Year 1, 5 pick up trucks are needed as replacement units ($200,000),
the computer system will be updated ($60,000) and roof repairs will be
made to solve the leaking problem ($100,000). A total of $200,000 has
been included in Years 1 and 2 under a general category for
miscellaneous items. This annual allowance is increased to $500,000
beginning Year 3.
Table 11-4 - Capital Equipment
- ------------------------------------------------------------------------------------------------------------------------
Item ($ 000) Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total
- ------------------------------------------------------------------------------------------------------------------------
Jack leg/Stopers - 50 25 25 25 - - - - - 125
Portable Bumpers - - 10 10 - - - - - - 20
Zone 97 Backfill System - 220 70 - - - - - - - 290
Backfill Pump - 360 - - - - - - - - 360
Forklift 25 - - - - - - - - - 25
Haulage Truck - 994 497 497 - - - - - - 1,988
Jumbo - 800 - - - - - - - - 800
Production Scooptram - - 440 880 - - - - - - 1,320
Service Scooptram - 360 180 - - - - - - - 540
Scissor Lift - 250 750 - - 250 - - - - 1,250
Service Vehicle - 105 210 - - - - - - - 315
Bar & Arm - - - 100 - - - - - - 100
BCI Long hole Drill - - 270 - - 135 - - - - 405
Grader - 240 240 - - - - - - - 480
Equipment Replacement - - 450 450 450 450 450 225 - - 2,475
Equipment Rehabilitation - 1,255 100 - - - - - - - 1,355
General - - 350 350 350 350 200 200 100 - 1,900
Pickup Trucks 200 - - - - - - - - - 200
Jaw Crusher - 125 - - - - - - - - 125
Computers - 60 - - - - - - - - 60
Cavity Monitor - 60 - - - - - - - - 60
Ventilation 97 Zone - 90 485 - - - - - - - 575
Oil Separator 15 - - - - - - - - - 15
Repair Roof 100 - - - - - - - - - 100
Mill General - 75 75 75 275 75 75 75 50 - 775
- ------------------------------------------------------------------------------------------------------------------------
Subtotal 340 5,044 4,152 2,387 1,100 1,260 725 500 150 - 15,658
- ------------------------------------------------------------------------------------------------------------------------
* 6 months
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
During Year 2, a paste backfill pump ($360,000) will be purchased and
installation of the zone 97 paste backfill delivery system will begin
on level 6 ($290,000). Two 20 tonne truck ($994,000), two jumbos
($800,000), two service scooptrams ($360,000), one scissor lift
($250,000), two service vehicles ($105,000), one grader ($240,000) and
a small jaw crusher ($125,000) will be purchased. An allowance of
($1,355,000) is included for the rehabilitation of existing equipment
which is carried out in 2002 and 2003.
An annual allowance of $450,000 is included for equipment replacement,
beginning in Year 3 and continuing until Year 7 and dropping to
$225,000 in Year 8.
During Year 3, an additional 20 tonne truck ($497,000) will be
purchased for the new mining areas in Zone 97. Additional equipment
needed for these areas include one 4.0 yd scooptram ($440,000) for
production, one 2.0 yd scooptram ($180,000) for service, three scissor
lifts ($750,000) for rehabilitation and four service vehicles
($210,000). In addition, one grader ($240,000) for zone 97 on 13 level
and two more BCI long hole drills ($270,000) will be purchased.
During Year 4, one additional 20 tonne truck ($497,000), two 4.0 yd
diesel scooptrams ($880,000) and one bar and arm drill ($100,000) will
be purchased for zone 97.
During Year 6, one additional scissor lift ($250,000) will be
purchased for Zone 97 as well as one BCI long hole drill ($135,000).
See Appendix RR - Equipment Cost.
Mill capital cost for years 2 to 8 have been included at $75,000 per
year except for Year 5 where $275,000 is required. The capital cost in
Year 5 includes enlarging the piping distribution, replacement of
pumps, replacement of the copper thickener (too small for future
copper grades), incorporation of a soda ash system, installing a water
heating system and an improved lime system.
11.3 SALVAGE VALUE
As shown in Table 11-5 a salvage value of $8.9 million has been
estimated for the plant, equipment and buildings. This value is
comprised of $0.6 million for buildings, $1.8 million for 9 houses and
3 apartment buildings in town, $2.7 million for the mill equipment and
$3.8 million for the underground equipment.
11.3.1 SALVAGE VALUE OVERVIEW
The buildings include the office, head frame, ore and waste bins, and
mechanical shop, etc. The mill equipment includes the compressors,
generator, SAG mill, ball mill, hoists (2), front-end loader,
mechanical tools, paste backfill plant, grinding circuit, flotation
circuit, thickeners and filtration circuit, pickups (4). The
underground equipment includes the electrical substations,
compressors, crusher, cage/skip, pumps, ventilation system, 20 tonne
trucks, service vehicles, scissor lifts, 2.0 yd scooptrams, 3.0 yd
scooptrams, 4.0 yd scooptrams, development jumbos, long hole drills
and various other small gear. See Appendix SS - Salvage Value.
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103 SRK CONSULTING
JUNE 2003
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
Table 11-5 - Salvage Value
($ 000)
- -----------------------------------------------------------------------------------------------------------------------------
Item Year 1* Year 2 Year 3 Year 4 Year 5 Year 6 Year 7 Year 8 Year 9 Year 10* Total
- -----------------------------------------------------------------------------------------------------------------------------
Salvage Value
Surface Installations - - - - - - - - - (1,417) (1,417)
Surface Technique - - - - - - - - - (21) (21)
Stationary Equipment - - - - - - - - - (567) (567)
Mobile Equipment - - - - - - - - - (1,793) (1,793)
Mill - - - - - - - - - (2,730) (2,730)
Buildings - - - - - - - - - (603) (603)
Houses & Apartments - - - - - - - - - (1,804) (1,804)
- -----------------------------------------------------------------------------------------------------------------------------
Total - - - - - - - - - (8,935) (8,935)
- -----------------------------------------------------------------------------------------------------------------------------
* 6 months
11.4 ENVIRONMENTAL AND ABANDONMENT
The environmental and abandonment cost is evaluated at $1,701,000 in
the Langlois mine closure plan. This sum includes all the work to
restore the mine site (dismantling all the buildings and
infrastructures, rearrangement of the tailings pond). See Appendix TT
- Closure Plan.
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104 SRK CONSULTING
JUNE 2003
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2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
12.0 ECONOMICS
12.1 CASHFLOW
Table 12-1 - Pre-Tax Cash Flow Summary
- -----------------------------------------------------------------------------------------------------------------------------------
YEAR * 6 months YEAR 1* YEAR 2 YEAR 3 YEAR 4 YEAR 5 YEAR 6 YEAR 7 YEAR 8 YEAR 9 YEAR 10* TOTAL
- -----------------------------------------------------------------------------------------------------------------------------------
Exchange Rate - 0.70 0.70 0.70 0.70 0.70 0.70 0.70 0.70 0.70
METAL PRICES
Zinc Price US$/lb - $0.50 $0.50 $0.50 $0.50 $0.50 $0.50 $0.50 $0.50 $0.50
Copper US$/lb - $0.80 $0.80 $0.80 $0.80 $0.80 $0.80 $0.80 $0.80 $0.80
Silver US$/oz - $5.00 $5.00 $5.00 $5.00 $5.00 $5.00 $5.00 $5.00 $5.00
ANNUAL ORE PROD. (`000) - 73.9 381.1 450.0 450.0 450.0 450.0 450.0 450.0 167.8 3,322.8
Zinc (%) - 11.29 10.51 10.78 10.93 10.97 10.96 10.73 10.51 10.59 10.78
Copper (%) - 0.63 0.63 0.69 0.73 0.87 0.97 0.96 0.89 0.92 0.82
Silver (g/t) - 54.35 50.10 51.66 52.54 53.79 53.89 52.16 50.22 51.77 52.13
Gold (g/t) - 0.09 0.08 0.08 0.08 0.09 0.09 0.09 0.08 0.08 0.09
METAL RECOVERY
Zinc in Zinc Con. (%) - 93.9 93.8 93.8 93.6 93.5 93.5 93.7 93.6 93.6 93.6
Copper in Copper Con. (%) - 77.5 76.6 79.0 80.5 83.8 85.5 85.0 84.8 85.3 82.3
Silver in Copper Con. (%) - 33.9 33.9 35.1 36.5 36.4 36.3 36.0 35.9 35.9 35.8
Gold in Copper Con. (%) - 29.2 30.0 29.2 29.2 28.4 28.4 29.2 28.4 28.4 29.0
CONCENTRATE GRADE
Zinc (%) - 52.8 52.8 53.8 54.3 55.0 55.5 56.0 56.0 56.0 54.8
Copper (%) - 22.0 22.0 22.5 23.0 24.0 24.5 25.0 25.5 25.5 24.0
CONCENTRATE PRODUCTION (`000 T)
Zinc - 14.8 71.1 84.7 84.9 84.0 83.1 80.8 79.0 29.7 612.1
Copper - 1.6 8.4 10.9 11.5 13.7 15.2 14.7 13.3 5.2 94.5
CONTAINED METAL
Zinc (`000 t) - 7.8 37.5 45.5 46.1 46.2 46.1 45.2 44.3 16.7 335.4
Copper (`000 t) - 0.4 1.8 2.5 2.6 3.3 3.7 3.7 3.4 1.3 22.7
Silver (`000 oz) - 43.7 208.1 262.4 277.5 283.4 283.2 271.9 260.8 100.3 1,991.3
Gold (oz) - 62 294 338 338 370 370 381 329 124 2,606
PAYABLE METAL
Zinc (`000 lb.) - 14,649 70,238 85,277 86,307 86,521 86,452 84,764 82,929 31,162 628,299
Copper (`000 lb.) - 759 3,868 5,167 5,575 6,930 7,890 7,770 7,194 2,789 47,942
Silver (`000 oz) - 40 192 241 255 261 261 250 240 92 1,832
Gold (oz) - 12 40 11 - - - - - - 63
SMELTER REVENUE (CDN $000) - 11,624 55,978 68,547 69,843 71,584 72,629 71,212 69,171 26,105 516,693
Less Treatment Charges - (4,125) (19,862) (23,709) (23,845) (23,938) (23,946) (23,196) (22,479) (8,476) (173,576)
Less Losses - (20) (99) (122) (125) (130) (134) (131) (127) (48) (936)
Less Transportation - (388) (1,882) (2,281) (2,312) (2,385) (2,434) (2,360) (2,264) (858) (17,164)
Less Operating
Mining (2,358) (7,760) (14,467) (15,363) (15,720) (16,143) (16,064) (15,951) (14,954) (5,483) (124,263)
Milling (264) (858) (4,605) (5,160) (5,061) (5,116) (5,098) (5,029) (4,794) (1,925) (37,910)
Administration (950) (2,352) (2,534) (2,534) (2,534) (2,534) (2,534) (2,534) (2,534) (1,546) (22,586)
Total Operating (3,572) (10,970) (21,606) (23,057) (23,315) (23,793) (23,696) (23,514) (22,282) (8,954) (184,759)
Less Shipping Cost to Smelter - (699) (3,362) (4,018) (4,039) (4,033) (4,022) (3,907) (3,804) (1,432) (29,317)
Less Working Capital Change - (1,278) (4,876) (1,529) (221) (315) (199) 95 224 8,099 -
Less Capital (4,608) (11,761) (10,454) (5,832) (3,945) (3,384) (2,790) (2,793) (1,411) 8,810 (38,166)
Less Environ. & Abandonment - - - - - - - - - (1,701) (1,701)
CASHFLOW (8,180) (17,618) (6,163) 7,998 12,042 13,607 15,408 15,407 17,028 21,546 71,075
Cumulative (8,180) (25,798) (31,961) (23,963) (11,921) 1,686 17,094 32,501 49,529 71,075
NPV @ 8.0% 30,890
10.0% 24,616
12.0% 19,344
IRR 25.3%
Cost/lb. Payable Zinc US$/lb.
Zn Treatment, Shipping Costs - $0.198 $0.198 $0.193 $0.191 $0.188 $0.186 $0.183 $0.183 $0.183 $0.189
Credit for Byproducts - ($0.038) ($0.039) ($0.042) ($0.045) ($0.054) ($0.060) ($0.060) ($0.058)($0.059) ($0.052)
Operating Cost excl. Deprec. - $0.558 $0.249 $0.222 $0.222 $0.225 $0.224 $0.226 $0.220 $0.233 $0.239
Total Operating Cost - $0.718 $0.408 $0.373 $0.368 $0.359 $0.350 $0.350 $0.345 $0.357 $0.376
Capital - $0.562 $0.104 $0.048 $0.032 $0.039 $0.027 $0.023 $0.012 ($0.198) $0.043
Total - $1.280 $0.512 $0.421 $0.400 $0.398 $0.377 $0.373 $0.357 $0.159 $0.419
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The cashflow is based on mining and milling 3.3 million tonnes at an
average grade of 10.78% zinc, 0.82% copper, 52.1 g/t silver and 0.09
g/t gold. Average metal recoveries over the life of the mine are 93.7%
zinc recovery, 82.3 copper recovery, 35.8% silver recovery and 29.0%
gold recovery. The average zinc concentrate grade is 54.8% zinc and
the average copper concentrate grade is 24.0% copper.
In total, 612,075 tonnes of zinc concentrate and 94,454 tonnes of
copper concentrate are produced. Contained metal amounts to 335,366
tonnes of zinc, 22,691 tonnes of copper, 1,991,345 ounces of silver
and 2,606 ounces of gold.
The total smelter revenue is $516.7 million and has been calculated
based on the following metal prices and exchange rate:
Table 12-2 - Metal Prices and Exchange Rate
-----------------------------------
Zinc US$0.50/lb
Copper US$0.80/lb
Silver US$5.00/oz
Gold US$343/oz
Exchange Rate Cdn$0.70/$US
-----------------------------------
Smelter penalties are applied to the zinc concentrate for iron and to
the copper concentrate for lead.
Payable metal is calculated according to Breakwater's existing smelter
contracts. The total payable zinc is 628,299,332 lbs. Payable copper
is 47,942,312 lbs. Payable silver is 1,832,038 oz and payable gold
amounts to 63 oz.
Total treatment charges for zinc and copper concentrates are included
according to Table 12-3. For more information see Section 9-3
Concentrate Terms. The total treatment charges are $173.6 million.
Table 12-3 - Concentrate Treatment Charges
($US)
-----------------------------------------------
TC Price Level
-----------------------------------------------
Zinc $ 160/t $1,000
Copper $ 80/t $1,984
-----------------------------------------------
Losses during transportation and assay exchange have been included and
amount to a total of $0.9 million.
Based on the above revenue and costs, the Net Smelter Return is
Cdn$97.81/tonne milled. (excluding concentrate transportation costs)
Operating and capital costs are deducted as per Sections 9.0 and 10.0.
The total operating cost is $184.8 million. Shipping charges of $29.3
million include rail freight for zinc concentrates to Noranda's CEZ
smelter at Valleyfield, Quebec and rail freight for the copper
concentrate to Noranda's Horne smelter in Rouyn-Noranda, Quebec. For
more information see Section 9-2 Concentrate Shipment.
Capital cost including salvage value totals $38.2 million. In addition
the environmental and rehabilitation cost is $1.7 million.
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The total site operating cost per tonne milled before shipping is
Cdn$55.61 per tonne. The total operating cost per pound of payable
zinc including shipping is US$0.376 and the capital cost per pound of
payable zinc is US$0.043 (Cdn$11.49/tonne milled) for a total
operating and capital cost of $0.419 per pound of payable zinc
including by-product credits for copper, silver and gold ($67.09/tonne
milled).
Figure 12-1 Mine Site Cost Distribution
[PICTURE]
The total net pre-tax cashflow is Cdn$71.1 million. At a zinc price of
US$0.50/lb. the mine produces a positive cashflow starting in Year 4.
On a cumulative basis , the mine is cash positive starting in Year 6.
The internal rate of return is 25.3% and the NPV at 8.0% is $30.9
million. Details of the cash flows can be found in Appendix UU -
Pre-Tax Cash Flows.
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12.2 PROJECT SENSITIVITY
Table 12-4 - Project Sensitivity
- --------------------------------------------------------------------------------------------------------------------
Cash Cost Cash Flow IRR NPV 8% NPV 10% NPV 12%
US$/lb Zn ($ `000) (%) ($ `000) ($ `000) ($ `000)
- --------------------------------------------------------------------------------------------------------------------
Base Case Zn $0.50/lb
Cu $0.80/lb
Exchange 0.70 $0.376 $71,075 25.3% $30,890 $24,616 $19,344
A. Metal Price
Zn Price + $0.05/lb $0.389 $100,465 34.9% $49,136 $40,969 $34,052
- $0.05/lb $0.355 $41,548 15.2% $12,559 $8,186 $4,567
Cu Price + $0.05/lb $0.368 $74,488 26.3% $32,952 $26,451 $20,984
- $0.05/lb $0.375 $67,661 24.2% $28,828 $22,780 $17,704
Zn+Cu Price + $0.05/lb $0.385 $103,874 35.9% $51,195 $42,802 $35,690
- $0.05/lb $0.358 $38,134 14.1% $10,497 $6,350 $2,926
Prices May 30, 2003 Zn $0.357/lb $0.329 ($18,952) -7.6% ($24,974) ($25,448) ($25,680)
B. Exchange Rate + $0.05 US/Cdn $0.388 $49,356 18.0% $17,461 $12,592 $8,540
- $0.05 US/Cdn $0.355 $96,135 33.4% $46,385 $38,490 $31,811
C. Metallurgy
Improved Zn Conc. Grade + 1% $0.369 $73,130 26.0% $32,192 $25,789 $20,404
Improved Zn Recovery + 1% $0.370 $73,586 26.1% $32,447 $26,010 $20,598
Improved Zn Grade & Recovery + 1% $0.368 $75,662 26.8% $33,762 $27,195 $21,669
Improved Cu Conc. Grade + 1% $0.372 $71,143 25.3% $30,930 $24,651 $19,376
Improved Cu Recovery + 1% $0.371 $71,519 25.4% $31,160 $24,856 $19,559
Improved Cu Grade & Recovery + 1% $0.371 $71,587 25.4% $31,200 $24,892 $19,590
D. Cost
Operating Cost + 10% $0.392 $52,599 18.4% $18,941 $13,793 $9,506
- 10% $0.351 $89,551 32.5% $42,840 $35,439 $29,183
Capital Cost + 10% $0.372 $66,365 22.6% $27,325 $21,268 $16,192
- 10% $0.372 $75,785 28.2% $34,456 $27,965 $22,497
Transportation Cost + 10% $0.375 $68,143 24.3% $29,063 $22,977 $17,869
- 10% $0.368 $74,007 26.3% $32,718 $26,255 $20,820
All Costs + 10% $0.395 $44,957 15.2% $13,548 $8,805 $4,877
E. Production
Grade Change - 10% $0.389 $45,974 16.8% $15,331 $10,676 $6,811
Throughput - 10% $0.390 $48,245 18.3% $17,253 $12,520 $8,581
Grade plus Throughput - 10% $0.408 $25,656 10.0% $3,251 ($24) ($2,697)
F. Smelter Charges
Zn Charges + $10/t con $0.381 $62,333 22.3% $25,435 $19,720 $14,935
Cu Charges + $10/t con $0.373 $69,728 24.9% $30,070 $23,884 $18,689
All smelter charges + $10/t con $0.383 $60,986 21.9% $24,614 $18,988 $14,280
- --------------------------------------------------------------------------------------------------------------------
The project sensitivity table indicates that the Langlois mine
feasibility plan is most sensitive to zinc metal prices and exchange
rates, which are items beyond control. Improvements in metallurgy
result in gains with zinc grade and recovery contributing the most.
Reductions in operating, capital and transportation costs also result
in improved cash flows. Reductions in head grade and throughput can be
adverse to the pre-tax cashflow. As well increases in smelter
treatment charges result in additional cost.
A US$0.05/lb of payable zinc rise increases the pre-tax cashflow by
almost 40% to $100.5 million and a US$0.05/lb decrease results in a
net pre-tax cashflow of $41.5 million.
Fluctuations in the price of copper have less effect on the net
pre-tax cashflow. Current prices as of May 30, 2001 (US$0.357/lb zinc)
result in a negative net pre-tax cashflow of -$18.9 million.
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A $0.05 US/Cdn increase in the exchange rate increases the net pre-tax
cashflow to $96.1 million, whereas a $0.05 US/Cdn decrease results in
a net pre-tax cashflow of $49.4 million.
A 1% improvement in zinc concentrate grade and recovery results in a
net pre-tax cashflow of $73.1 million. Improvements in copper
metallurgy have less of an effect.
A 10% increase in operating cost reduces the net pre-tax cashflow by
about 26%, whereas a 10% decrease in operating cost increases net
pre-tax cashflow to $89.6 million from $71.1 million.
Capital cost fluctuations do not affect the net pre-tax cashflow as
much as increases or decreases in operating cost. A 10% increase in
capital returns $66.4 million, while a 10% decrease returns $75.8
million. The feasibility plan is also not as sensitive in the
transportation cost area.
A 10% reduction in head grade reduces net pre-tax cashflow to $46.0
million while a 10% reduction in throughput results in a cashflow
reduction to $48.2 million.
An increase of $10/t in zinc treatment charges reduces net pre-tax
cashflow to $62.3 million, while a $10/t increase in copper treatment
charges reduces net pre-tax cashflow slightly to $69.7 million.
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13.0 OPPORTUNITIES AND RISKS
13.1 OPPORTUNITIES
This feasibility study relies on historical mine operating data, and
has been developed on the basis of introducing only those changes
required to address the most challenging operating problems and
deficiencies described in Section 3.0.
During the course of the development of the feasibility study, several
opportunities were identified that hold the potential to further
improve productivity and operating costs. These opportunities were not
incorporated into the feasibility plan as there was not enough
information or studies had not yet been carried out to ensure that
they would be successful. These opportunities should be evaluated and
appropriately tested at the Langlois mine before final engineering
design or during the pre-production period. They are individually
discussed below:
13.1.1 MINING
13.1.1.1 UNDERHAND BENCH AND FILL
The feasibility study is based on overhand stope sequencing in all
areas. In Zone 97 there is a possibility of employing an underhand
bench and fill method. The potential benefits are a reduction in
up-front capital development to support production, and less
rehabilitation of sublevels. Issues to investigate include a survey of
other mines working under paste backfill (equipment only, not
personnel), the cement content requirements in the backfill, and the
effect on production scheduling of blasting (upholes) and mucking from
the same level.
13.1.1.2 TELE-REMOTE MUCKING
Tele-remote mucking could be employed in Zone 97. The scooptram
operator would be located at the remuck bay in the crosscut and would
not get on and off his machine. This practice is frequently employed
in Australia with much higher productivities, less damage to
equipment, and improved safety for the operator.
This practice would eliminate the need for safety bays in the mucking
drifts. Safety bay cut outs will weaken the stope walls to some extent
by undercutting the foliation.
Tele-remote could also prove advantageous during remote mucking if the
ore is thrown farther than expected during blasting.
13.1.1.3 SHOTCRETE
The use of shotcrete for support in the Zone 97 ore drifts would be
more expensive initially, but would greatly reduce future
rehabilitation time and costs. One possibility to keep the shotcrete
costs down would be to use a wet mix with a slick line from surface,
and small delivery trucks.
Time saved in sublevel rehabilitation would speed up the stope cycle,
contributing to production surety.
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13.1.2 EQUIPMENT
13.1.2.1 ZONE 97 ELECTRIC-HYDRAULIC EQUIPMENT
Electric-hydraulic development jumbos are included in the feasibility
plan, but Zone 97 production drilling equipment is planned as
pneumatic. Electric-hydraulic production drills are more expensive to
purchase, but are much more energy efficient. Issues to investigate
include capital versus operating costs, available hole sizes, and
machine size and manoeuvrability in the narrow ore drifts.
13.1.2.2 ZONE 3 DRILLING EQUIPMENT
The feasibility plan includes 114mm diameter ITH production drilling
in Zone 3 on 30m and 20m sublevels. It may prove economically
attractive to drill the 20m sublevels with 54mm diameter holes. Issues
to investigate include 114 mm versus 54mm drilling costs, hole
deviation, and improvements in stope wall control.
This opportunity can be assessed independently of the
electric-hydraulic versus pneumatic question.
13.1.2.3 TELE-TRAM TRUCKS
There is a possibility of using tele-tram trucks in Zone 97. The
advantage would be less back height requirements at dumping points.
This has the potential to save on some development, but perhaps more
significantly, to improve the flexibility of how the trucks can be
used for hauling development waste to stopes or remuck areas. Issues
to investigate include capital costs, models available (capacity and
dimensions), and reliability and maintenance cost estimates.
13.1.3 MANAGEMENT
13.1.3.1 MANAGEMENT AND OPERATIONAL EFFECTIVENESS
Mine-wide improvements in operational effectiveness as measured by
productivities achieved, are certainly possible, but may only be
implemented through a management philosophy change. There is an
opportunity during the capital period to make a fresh start by
initiating changes in management systems and personnel. This is a
difficult process as it affects people, but at least for the
management positions that are currently vacant there is an opportunity
for very selective recruitment.
13.1.3.2 MINE PLANNING AND SCHEDULING SYSTEMS
Integrated systems for efficient resource modelling and mine planning
are widely available and should be considered for the Langlois mine.
Scheduling software is also available that can be of great benefit in
planning and scheduling development and production activities. The
Langlois mine has small stopes cycling quickly, and good scheduling
and communication will help ensure production targets are achieved.
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13.1.3.3 SHIFT SCHEDULE
The mine plan shift schedule is for 80 hours per week for most
underground activities. There is an opportunity to work more hours per
week. This can be viewed as an opportunity to catch up should there be
any periods of production shortfall.
The cage and hoist are scheduled to be operational three shifts per
day, or 120 hours per week, to facilitate the movement of materials
underground. The use of limited amounts of overtime work on production
activities, such as stope mucking, is an opportunity that can be used
for short periods.
13.1.4 RESERVES
13.1.1.1 EXPLORATION
There is potential for mineable reserves to be increased through
ongoing exploration at the Langlois mine, and this has been described
in Section 2.6 of this report.
13.1.1.2 CUT-OFF GRADE
Mineral reserves may also be increased through an increase in the NSR
value of the ore, and the corresponding decrease in cut-off grade. The
feasibility mine plan does not include mining of all of the December
31, 2000 mineral reserves, since it is a high-grade alternative. If
metal prices and/or the exchange rate improve, the mineral reserves
will increase. The mine plan is compatible with accessing most of the
lower grade areas.
13.1.1.3 ZONE 97 NORTH AND SOUTH LENSES
The Langlois mine resources in Zone 97 includes tonnes from three
parallel lenses; main, north, and south. Refer to section (2.3) of
this report. Only main lens resources have been converted to mineral
reserves. The north and south lenses were not brought into reserves
because they are narrow, and their diluted grades are too low.
However, at higher metal prices, it is anticipated that some of these
resources would become mineable.
13.1.5 CONSUMABLES
13.1.5.1 CEMENT BINDER REPLACEMENT
It may be possible to reduce the cost of cemented paste backfill by
replacing some of the cement with lower cost fly ash. Issues to
investigate include any capital costs for fly ash handling, fly ash
versus cement cost per tonne, backfill strength testing and strength
gain timing.
13.1.5.2 CORPORATE LEVEL COST NEGOTIATION
At the corporate level, negotiations with suppliers may result in
savings on some of the high cost consumables. There has been some
success in the past, and this is an area that will be pursued in the
future.
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13.2 RISKS
The following risks to the financial success of the project have been
identified.
13.2.1 METAL PRICES
The financial results of the project are directly related to metals
prices, particularly the price of zinc. Refer to Section 12.2 -
Project Sensitivity.
13.2.2 PRODUCTION RATE
There is always a risk of not achieving the planned production rate.
The Langlois mine stope sizes are small and they cycle quickly.
Several stopes must be in the production cycle at any one time,
leading to many individual activities to manage.
A key factor in managing this risk is the effective scheduling of
development and production activities, and good communication and
co-operation between the engineering and mine operations groups.
13.2.3 GRADE AND DILUTION
The most likely cause of not achieving planned grades will be
excessive dilution. The risk is perhaps the greatest in Zone 97 where
there is no stoping experience.
It is possible that Zone 97 is more distorted than we have assumed so
the dilution could be greater than estimated.
The planned minimum mining width for Zone 97 is 2.2m. If this is not
achieved in the areas of thin ore, then dilution will tend to be
greater than planned. It will be a challenge for development crews to
follow the ore zone without undercutting the wall rock. Good geology
control will be required for success.
The minimum blast size (number of rows) in Zone 97 may be dictated by
production requirements, and may be larger than ideal from a ground
control point of view. This could have an adverse affect on dilution.
13.2.4 CAPITAL OVERRUNS
Capital overruns are a risk on any project. There is no capital
contingency.
If there are any delays in the start up of production, the financing
requirements of the project could be increased.
13.2.5 UNDERGROUND BACKFILL PUMPING
There is some risk associated with operating a backfill pumping
station underground. The perceived risks are in the areas of
reliability and operating cost.
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APPENDIX A - CERTIFICATES OF QUALIFIED PERSONS
Note:
Daniel Vallieres, Ronald Durham, Alain Cossette, Tony Brisson, Marc Bernard,
Andre Dessureault, and Martine Deshaies were all qualified persons who were
involved with the original 2001 Langlois Mine Feasibility Study. These persons
have either left the employ of Breakwater Resources Ltd. or have been
transferred to other Breakwater Divisions. They were not involved in the 2003
update to the Langlois Mine Feasibility Study.
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CERTIFICATE AND CONSENT
TO ACCOMPANY THE BREAKWATER-LANGLOIS MINE FEASIBILITY STUDY,
LEBEL-SUR-QUEVILLON, QUEBEC, CANADA
I, Ken S. Reipas, residing at 43 Deverell Street, Whitby, Ontario do hereby
certify that:
1) I am a Principal Mining Engineer with the firm of Steffen Robertson and
Kirsten (Canada) Inc. (SRK) with an office at Suite 602, 357 Bay Street,
Toronto, CANADA.
2) I am a graduate of Queen's University with a B.Sc. in Mining Engineering
1981, and have practiced my profession continuously since 1981;
3) I am a Professional Engineer registered with the Association of
Professional Engineers of Ontario;
4) I have not received, nor do I expect to receive, any interest, directly
or indirectly, in the Mine Langlois Feasibility Project or securities of
Breakwater Resources Limited.
5) I am not aware of any material fact or material change with respect to
the subject matter of the technical report, which is not reflected in the
technical report, the omission to disclose which makes the technical report
misleading.
6) I, as a qualified person, am independent of the issuer as defined in
Section 1.5 of National Instrument 43-101.
7) I have not had any prior involvement with the property that is subject
to the technical report.
8) I have read National Instrument 43-101 and Form 43-101F1 and the
technical report has been prepared in compliance with this Instrument and
Form 43-101F1.
9) Steffen Robertson and Kirsten (Canada) Inc. was retained by Breakwater
Resources Ltd. to prepare a report on the Langlois Mine Feasibility Project
in accordance with National Instrument 43-101. The following report is
based on SRK's review of project files, SRK's review of engineering studies
carried out by Breakwater Resources Ltd. personnel, and on observations
made by several of SRK's professional staff during numerous visits to site
in 2000 and in 2001.
10) I was a co-author of sections 1, 3, 4, 7, 8, 10, 11, 12, and 13 of the
report and provided supervision and peer review.
11) I hereby consent to use of this report and our name in the preparation
of a prospectus for submission to any Provincial regulatory authority.
"SEAL"
"Ken S. Reipas"
Toronto, Canada Ken S. Reipas, P. Eng.,
June 06, 2003 Principal Mining Engineer
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CERTIFICATE AND CONSENT
TO ACCOMPANY THE BREAKWATER-LANGLOIS MINE FEASIBILITY STUDY,
LEBEL-SUR-QUEVILLON, QUEBEC, CANADA
I, Michael J. Michaud, residing at 43 Eastlawn Street, Oshawa, Ontario do hereby
certify that:
1) I am a Senior Geologist with the firm of Steffen Robertson and Kirsten
(Canada) Inc. (SRK) with an office at Suite 850, 121 King Street West,
Toronto, Canada.
2) I am a graduate of the University of Waterloo with a HBSc. in Earth
Science, MSc. from Lakehead University in 1998, and have practiced my
profession continuously since 1987.
3) I am a a fellow with the Geological Association of Canada and a
Professional Geoscientist registered with the Association of
Professional Engineers and Geoscientists of the province of British
Columbia;
4) I have not received, nor do I expect to receive, any interest, directly
or indirectly, in the Langlois Mine or securities of Breakwater
Resources Ltd.
5) I am not aware of any material fact or material change with respect to
the subject matter of the technical report, which is not reflected in
the technical report, the omission to disclose which makes the
technical report misleading.
6) I, as the qualified person, am independent of the issuer as defined in
Section 1.5 of National Instrument 43-101.
7) I have not had any prior involvement with the property that is subject
to the technical report.
8) I have read National Instrument 43-101 and Form 43-101F1 and the
technical report has been prepared in compliance with this Instrument
and Form 43-101F1.
9) Steffen Robertson and Kirsten (Canada) Inc. was retained by Breakwater
Resources Ltd. to prepare a feasibility study for the Langlois Mine in
accordance with National Instrument 43-101. The following report is
based on our review of project files, discussions with Breakwater
Resources Ltd personnel and observations made during several site
visits between January and April, 2001.
10) I was the co-author of the report.
11) I hereby consent to use of this report for submission to any Provincial
regulatory authority.
Toronto, Canada Michael J. Michaud, P.Geo.,
June 06, 2003 Senior Geologist
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116 SRK CONSULTING
JUNE 2003
- --------------------------------------------------------------------------------
2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
CERTIFICATE AND CONSENT
TO ACCOMPANY THE BREAKWATER-LANGLOIS MINE FEASIBILITY STUDY,
LEBEL-SUR-QUEVILLON, QUEBEC, CANADA
I, Torben Jensen, residing at 10 Asprey Court, Brampton, Ontario do hereby
certify that:
1) I am a Mining Engineer working for Breakwater Resources Ltd., with an
office located at 2000-95 Wellington St. W., Toronto, Ontario.
2) I am a graduate of South Dakota School of Mines with a B.Sc. in Mining
Engineering, 1978, and a graduate of Haileybury School of Mines with a
Mining Engineering Technologist Diploma, 1975 and have practiced my
profession continuously since 1978.
3) I am a Professional Engineer registered with the Association of
Professional Engineers of Ontario;
4) I have reviewed all sections of the report and have provided
supervision and peer review. I was a co-author of section 12 of the
report
5) I, as a co-author, agree with the facts in this report. Everything I
have prepared was done with diligence.
Toronto, Ontario, Canada Torben Jensen, P.Eng
June 06, 2003 Manager of Engineering and
North American Exploration
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117 SRK CONSULTING
JUNE 2003
- --------------------------------------------------------------------------------
2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
CERTIFICATE AND CONSENT
TO ACCOMPANY THE BREAKWATER-LANGLOIS MINE FEASIBILITY STUDY,
LEBEL-SUR-QUEVILLON, QUEBEC, CANADA
I, Daniel Vallieres, residing at 69 Cote du Plateau, Lebel-sur-Quevillon, Quebec
do hereby certify that:
1) I am a Mining Engineer working for Breakwater Resources Ltd., Langlois
Mine located at Lebel-sur-Quevillon, Quebec, Canada.
2) I am a graduate of Laval University with a B.Sc. in Mining Engineering
1991, and have practiced my profession continuously since 1991.
3) I am an Engineer registered with the Ordre des Ingenieurs du Quebec
(O.I.Q).
4) I have reviewed all sections of the report and have provided
supervision and peer review.
5) I, as a co-author, agree with the facts in this report. Everything I
have prepared was done with diligence.
Lebel-sur-Quevillon, Quebec, Canada Daniel Vallieres ing.,
June 14, 2001 Mining Engineer
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118 SRK CONSULTING
JUNE 2003
- --------------------------------------------------------------------------------
2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
CERTIFICATE AND CONSENT
TO ACCOMPANY THE BREAKWATER-LANGLOIS MINE FEASIBILITY STUDY,
LEBEL-SUR-QUEVILLON, QUEBEC, CANADA
I, Ronald Durham, residing at 64 Des Sapins, Lebel-sur-Quevillon, Quebec do
hereby certify that:
1) I am a Mining Engineer working for Breakwater Resources Ltd., Langlois
Mine located at Lebel-sur-Quevillon, Quebec, Canada.
2) I am a graduate of Laval University with a B.Sc. in Mining Engineering
1994, and have practiced my profession continuously since 1994.
3) I am an Engineer registered with the Ordre des Ingenieurs du Quebec
(O.I.Q).
4) I was a co-author of sections 1, 3, 4, 7, 8, 10, 11 of the report.
5) I, as a co-author, agree with the facts in the sections I have written.
Everything I have prepared was done with diligence.
Lebel-sur-Quevillon, Quebec, Canada Ronald Durham ing.,
June 14, 2001 Mining Engineer
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119 SRK CONSULTING
JUNE 2003
- --------------------------------------------------------------------------------
2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
CERTIFICATE AND CONSENT
TO ACCOMPANY THE BREAKWATER-LANGLOIS MINE FEASIBILITY STUDY,
LEBEL-SUR-QUEVILLON, QUEBEC, CANADA
I, Alain Cossette, residing at 53 Rivest place, Lebel-sur-Quevillon, Quebec do
hereby certify that:
1) I am a Mining Engineer working for Breakwater Resources Ltd., Langlois
Mine located at Lebel-sur-Quevillon, Quebec, Canada.
2) I am a graduate of L'Ecole Polytechnique de Montreal with a B.Sc. in
Mining Engineering 1997, and have practiced my profession continuously
since 1997.
3) I am an Engineer registered with the Ordre des Ingenieurs du Quebec
(O.I.Q).
4) I was a co-author of sections 2, 3, 4, 7, 8 and 10 of the report.
5) I, as a co-author, agree with the facts in the sections I have written.
Everything I have prepared was done with diligence.
Lebel-sur-Quevillon, Quebec, Canada Alain Cossette ing.,
June 14, 2001 Mining Engineer
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120 SRK CONSULTING
JUNE 2003
- --------------------------------------------------------------------------------
2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
CERTIFICATE AND CONSENT
TO ACCOMPANY THE BREAKWATER-LANGLOIS MINE FEASIBILITY STUDY,
LEBEL-SUR-QUEVILLON, QUEBEC, CANADA
I, Tony Brisson, residing at 45 des Hetres, Lebel-sur-Quevillon, Quebec do
hereby certify that:
1) I am a Geologist working for Breakwater Resources Ltd., Langlois Mine,
located at Lebel-sur-Quevillon, Quebec, Canada.
2) I am a graduate Universite du Quebec a Chicoutimi (U.Q.A.C.) with a
B.Sc. in geology, 1992 and have practiced my profession continuously
since 1992.
3) I am a Geologist registered with the Association professionnelle des
geologues et geophysiciens du Quebec.
4) I have reviewed section 2 of the report.
5) I, as a co-author, agree with the facts in the sections I have written.
Everything I have prepared was done with diligence.
Lebel-sur-Quevillon, Quebec, Canada Tony Brisson, B.Sc.,
June 29, 2001 Chief Geologist
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121 SRK CONSULTING
JUNE 2003
- --------------------------------------------------------------------------------
2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
CERTIFICATE AND CONSENT
TO ACCOMPANY THE BREAKWATER-LANGLOIS MINE FEASIBILITY STUDY,
LEBEL-SUR-QUEVILLON, QUEBEC, CANADA
I, Denis Vaillancourt, residing at 111 Lesage, Val D'or, Quebec do hereby
certify that:
1) I am a Geologist working for Breakwater Resources Ltd., Langlois Mine,
located at Lebel-sur-Quevillon, Quebec, Canada.
2) I am a graduate of Universite du Quebec a Montreal (U.Q.A.M.) with a
M.Sc. in geology, 1996, and have practiced my profession continuously
since 1996.
3) I am a Geologist registered with the Association professionnelle des
geologues et geophysiciens du Quebec.
4) I was a co-author of sections 2 and 4 of the report.
5) I, as a co-author, agree with the facts in the sections I have written.
Everything I have prepared was done with diligence.
Lebel-sur-Quevillon, Quebec, Canada Denis Vaillancourt M.Sc.,
June 14, 2001 Senior Geologist
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122 SRK CONSULTING
JUNE 2003
- --------------------------------------------------------------------------------
2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
CERTIFICATE AND CONSENT
TO ACCOMPANY THE BREAKWATER-LANGLOIS MINE FEASIBILITY STUDY,
LEBEL-SUR-QUEVILLON, QUEBEC, CANADA
I, Marc Bernard, residing at 49 Des Hetres, Lebel-sur-Quevillon, Quebec do
hereby certify that:
1) I am a Mining Technician working for Breakwater Resources Ltd.,
Langlois Mine located at Technician, Quebec, Canada.
2) I am a graduate of College Region Amiante with a D.E.C. in Mining
Technology 1980, and have practiced my profession continuously since
1980.
3) I was a co-author of section 4 of the report.
4) I, as a co-author, agree with the facts in the section I have written.
Everything I have prepared was done with diligence.
Technician, Quebec, Canada Marc Bernard
June 14, 2001 Mine Superintendent
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123 SRK CONSULTING
JUNE 2003
- --------------------------------------------------------------------------------
2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
CERTIFICATE AND CONSENT
TO ACCOMPANY THE BREAKWATER-LANGLOIS MINE FEASIBILITY STUDY,
LEBEL-SUR-QUEVILLON, QUEBEC, CANADA
I, Andre Dessureault, residing at 125 Principale Sud, Lebel-sur-Quevillon,
Quebec do hereby certify that:
1) I am a Mining Technician working for Breakwater Resources Ltd.,
Langlois Mine located at Quevillon, Quebec, Canada.
2) I am a graduate of College de l'Abitibi-Temiscamingue with a D.E.C. in
Mining Technology 1990, and have practiced my profession continuously
since 1990.
3) I was a co-author of sections 3, 4, 7, 8 and 10 of the report.
4) I, as a co-author, agree with the facts in the sections I have written.
Everything I have prepared was done with diligence.
Lebel-sur-Quevillon, Quebec, Canada Andre Dessureault
June 14, 2001 Senior Mining Technician
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124 SRK CONSULTING
JUNE 2003
- --------------------------------------------------------------------------------
2CB003.02 LANGLOIS MINE FEASIBILITY STUDY, LEBEL-SUR-QUEVILLON, QUEBEC
CERTIFICATE AND CONSENT
TO ACCOMPANY THE BREAKWATER-LANGLOIS MINE FEASIBILITY STUDY,
LEBEL-SUR-QUEVILLON, QUEBEC, CANADA
I, Martine Deshaies, residing at 69 Cote du Plateau, Lebel-sur-Quevillon, Quebec
do hereby certify that:
1) I am a Mining Engineer working for Breakwater Resources Ltd., Langlois
Mine located at Lebel-sur-Quevillon, Quebec, Canada.
2) I am a graduate of Laval University with a B.Sc. in Mining Engineering
1992, and have practiced my profession continuously since 1992.
3) I am an Engineer registered with the Ordre des Ingenieurs du Quebec
(O.I.Q).
4) I was a co-author of sections 5, 6, 7, 9 and 10 of the report.
5) I, as a co-author, agree with the facts in the sections I have written.
Everything I have prepared was done with diligence.
Lebel-sur-Quevillon, Quebec, Canada Martine Deshaies ing.,
June 29, 2001 Mining Engineer
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125 SRK CONSULTING
JUNE 2003