PRELIMINARY MINE FEASABILITY RPT...Limited pyrochlore-columbite deposit located in the James Bay...

YESTERDAY RIVER010

PROJECT TERRANE

INTERIM REPORT

Prepared for

ARGOR EXPLORATIONS LIMITED

CANADIAN BECHTEL LIMITED

JULY 27, 1967

111111111111111111111111111 421,.."._ 83.1-48 YESTERDAY RIVER

010

PROJECT TERRANE

INTERIM REPORT

Prepared for



JULY 27, 1967

42HSNVM*a S3. 1-46 YESTERDAY RIVER O1OC

TABLE OF CONTENTS

1.0 SUMMARY AND PROJECT ECONOMICSl. l SUMMARY1.2 PROJECT ECONOMICS

2.0 INTRODUCTION2.1 SCOPE AND RESPONSIBILITIES

3.0 CAPITAL AND OPERATING COST ESTIMATES3.1 BASIS OF CAPITAL COST ESTIMATE3.2 OPERATING COST ESTIMATE

4.0 ORE RESERVES4.1 GEOLOGICAL RESERVES4.2 RECOVERABLE RESERVES

5.0 MINING METHODS5.1 OPEN PIT CASE I Si III5.2 UNDERGROUND CASE II5.3 UNDERGROUND CASE III

6.0 APPENDIX6.1 LIST OF DRAWINGS

PAGE

l

l 3

66

88

12

141415

18182228

3131

, t~

11111111111 ( 421, .... _ 83.1-48 YESTERDAY RIVER 010C

• TABLE OF CONTENTS

PAGE

1.0 SUMMARY AND PROJECT ECONOMICS 1 1.1 SUMMARY 1 1.2 PROJECT ECONOMICS 3

2.0 INTRODUCTION 6 2.1 SCOPE AND RESPONSIBILITIES 6

3.0 CAPITAL AND OPERATING COST ESTIMATES 8 3.1 BASIS OF CAPITAL COST ESTIMATE 8 3.2 OPERATING COST ESTIMATE 12

4.0 ORE RESERVES 14 4.1 GEOLOGICAL RESERVES 14 4.2 RECOVERABLE RESERVES 15

5.0 MINING METHODS 18

• 5.1 OPEN PIT CASE I & III 18 5.2 UNDERGROUND CASE II 22 5.3 UNDERGROUND CASE III 28

6.0 APPENDIX 31 6.1 LIST OF DRAWINGS 31

•

1.0 SUMMARY AND PROJECT ECONOMICS

1.1 Summary

This report comprises a preliminary mining appraisal together with an order of magnitude assessment of both capital and operating costs of the proposed mining and processing operations to be performed on Argor Explorations' Limited pyrochlore-columbite deposit located in the James Bay lowlands area of Northern Ontario.

The property is located approximately 31 miles south of Moosonee, Ontario. Access to the property is by helicopter from Moosonee in the summertime; in the winter by either winter trail from Moosonee or ski-plane from Moosonee or Timmins.

The mining study has been carried out on the basis of the following three cases:

Case I: Mining the entire orebody by open pit methods, excluding the wedge of low grade material in the north end of the deposit, but including all columbite ore.

Case II: Mining the orebody entirely by underground methods, excluding all columbite ore.

Case III: Mining the south end of the orebody by open pit followed by underground development of the northwest limb, excluding all columbite ore.

The assessment of the three cases is based on an annual tonnage sufficient to produce 7.5 million pounds per year of Cb205 for a period of 20 years

- l -

1.0 SUMMARY AND PROJECT ECONOMICS

1.1 Summary

This report comprises a preliminary mining appraisal together with an order of magnitude assessment of both capital and operating costs of the proposed mining and processing operations to be performed on Argor Explorations' Limited pyrochlore-columbite deposit located in the James Bay lowlands area of Northern Ontario.

The property is located-approximately 31 miles south of Moosonee, Ontario. Access to the property is by helicopter from Moosonee in the summertime; in the winter by either winter trail from Moosonee or ski-plane from Moosonee or Timmins.

The mining study has been carried out on the basis of the following three cases:

Case I: Mining the entire ore body by open pit methods, excluding the wedge of low grade material in the north end of the deposit, but including all columbite ore.

Case II: Mining the ore body entirely by underground methods, excluding all columbite ore.

Case III: Mining the south end of the orebody by open pit followed by underground development of the northwest limb, excluding all columbite ore.

The assessment of the three cases is based on an annual tonnage sufficient to produce 7.5 million pounds per year of Cb205 for a period of 20 years

- 1 -

at an overall mill recovery of 7596.

Recoverable ore reserves have been calculated on the basis of the preliminary mining layouts and are estimated to be as follows:

Total

Tons in ThousandsIII

II Open Pit Underground

Indicated 96Cb20 5

Possible

23,0390.51

23,0390.51

8,7320.57

8,4740.59

17,2060.58

12,7140.59

6,3650.49

12,7140.59

6,3650.49

Additional deep hole drilling is required to prove up the ore tonnages and grades for Case II between the 525 and 825 elevation.

Open pit mining will be carried out using wheel mounted front end loaders loading into 22-ton capacity trucks. Stripping of the overburden will be done with 5*5 yard wheel mounted front end loaders loading into 35-ton capacity trucks.

Pre-production stripping will involve the removal of some 4.9 million cubic yards of overburden consisting of peat, till, unconsolidated silty sand and a siltstone-mudstone-sandstone sequence. Stripping quantities are based on an overall slope of 3 horizontal to l vertical. It has been assumed that stripping will be carried out on a contract basis.

Mining underground will be done using a sublevellonghole method. Ore will be trammed from drawpoints to the ore pass system by low profile, rubber tired frontend loaders. The mining efficiency has been taken at IB'fc of the total available ore.

- 2 -

•

•

•

at an overall mill recovery of 75%.

Recoverable ore reserves have been calculated on the basis of the preliminary mining layouts and are estimated to be as follows:

I

23,039 0.51

23,039 0.51

Tons in Thousands III

II

8,732 0.57

8,474 0.59

17,206 0.58

Open pit

12,714 0.59

12,714 0.59

Underground

6,365 0.49

6,365 0.49

Additional deep hole drilling is required to prove up the ore tonnages and grades for Case II between the 525 and 825 elevation •

Open pit mining will be carried out using wheel mounted front end loaders loading into 22-ton capacity trucks. Stripping of the overburden will be done with 5~ yard wheel mounted front end loaders loading into 35-ton capacity trucks.

Pre-production stripping will involve the removal of some 4.9 million cubic yards of overburden consisting of peat, till, unconsolidated silty sand and a siltstone-mudstone-sandstone sequence. Stripping quantities are based on an overall slope of 3 horizontal to 1 vertical. It has been assumed that stripping will be carried out on a contract basis.

Mining underground will be done using a sublevellonghole method. Ore will be trammed-from drawpoints to the ore pass system by low profile, rubber tired frontend loaders. The miqing efficiency has been taken at 75% of the total available ore •

- 2 -

Total funds are estimated for each of the three cases as 524,212,000, 518,589,000 and 523,291,000 respectively. In addition a further expenditure will be involved in Case III on termination of the open pit in order to develop the underground mine and to purchase equipment suitable for underground extraction. This additional sum is estimated to be 52,621,000 at 1967 costs and has been taken into account in the prepar ation of the financial projections included in this report.

Average mining and processing costs over the 20-yearperiod for the three cases considered are estimatedto be 53.77, 5 5 *05 and 53.56 per ton of ore respectively.

1.2 Project Economics

Cash Flows have been developed over the twenty-year life for each of the three cases. These cash flows have been developed on the following basis:

1. Value of concentrate has been taken at $1.00 Canadian per pound of contained Cb2Q5*

2. Schedules of annual mining tonnage have been prepared for the open pit cases. A constant average annual tonnage has been used for the underground schemes.

3. These schedules have been combined with theoperating costs to arrive at an annual mining and processing cost.

4. Total funds required are estimated as follows:

CASE I II III

Plant Costs 8,650,000 9,922,000 8,650,000Pre-production 12,271,000 6,652,000 12,271,000Other Costs 3.291.000 2,015,000 2.370.000

Total $24,212 I 000 518,589,000 523,291,000

- 3 -•

Total funds are estimated for each of the three cases as $24,212,000, $18,589,000 and $23,291,000 respectively. In addition a further expenditure will be involved in Case IlIon termination of the open pit in order to develop the underground mine and to purchase equipment suitable for underground extraction. This additional sum is estimated to be $2,621,000 at 1967 costs and has been taken into account in the preparation of the financial projections included in this report.

Average mining and processing costs over the 20-year period for 'the three cases considered are estimated to be $3.77, $5.05 and $3.56 per ton of ore respectively.

1.2 Project Economics

Cash Flows have been developed over the twenty-year life for each of the three cases. These cash flows have been developed on the following basis:

1. Value of concentrate has been taken at $1.00 Canadian per pound of contained Cb205.

2. Schedules of annual mining tonnage have been prepared for the open pit cases. A constant average annual tonnage has been used for the underground schemes.

3. These schedules have been combined with the operating costs to arrive at an annual mining and processing cost. '

4. Total funds required are estimated as follows:

CASE I II .!!!

Plant Costs 8,650,000 9,922,000 8,650,000 Pre-production 12,271,000 6,652,000 12,271,000 Other Costs 3,291,000 2,015,000 2,370,000

Total $24,212,000 $18,589,000 $23,291,000

- 3 -

5. A 24 month construction period has beenassumed with 4096 of the total funds spent in the first year.

6. All funds are assumed to have been raisedinternally and no interest has been charged against these funds.

7. Plant replacement costs have been estimated for all mining equipment on the basis of an economic life recommended by the equipment manufacturer.

8. Ontario provincial depreciation rates andtax on profit have been calculated in accord ance with the Mining Tax Act, Chapter 242, Volume 3, the "Revised Statutes of Ontario, 1960".

9. Mining pre-production expenses have been deducted immediately following the 3-year tax free period.

10. Capital cost allowance has been taken according to the varying rates as set forth in the Federal Income Tax Act.

11. A depletion allowance of 33 1/396 of the profits has been taken and federal tax calculated at 5096 of the remainder.

- 4 -

•

~.

5. A 24 month construction period has been assumed with 40% of the total funds spent in the first year.

6. All funds are assumed to have been raised internally and no interest has been charged against these funds.

7. Plant replacement costs have been estimated for all mining equipment on the basis of an economic life recommended by the equipment manufacturer.

8. Ontario provincial depreciation rates and tax on profit have been calculated in accordance with the Mining Tax Act, Chapter 242, Volume 3, the "Revised statutes of Ontario, 1960".

9. Mining pre-production expenses have been deducted immediately following the 3-year tax free period.

10. capital cost allowance has been taken according to the varying rates as set forth in the Federal Income Tax Act.

11. A depletion allowance of 33 1/3% of the profits has been taken and federal tax calculated at 50% of the remainder.

- 4 -

The following table summarizes the total cash flow for each of the three cases:

CASE I II III(Dollars in Thousands)

20 year value of product 150,000 150,000 150,000

20 year operating cost 73,817 86,560 67,971

Plant additions 4,006 2,297 6,237

Taxes 17,763 15.164 18.644

Subtotal Deductions 95,586 104,021 92,852

Cash Flow 54,414 45,979 57,148

Discounted Rate of Return 996

The following three tables illustrate the full calculation of the 20 year cash flow.

- 5 -

•

•

The following table summarizes the ~otal cash flow for each of the three cases:

CASE I II (Dollars in Thousands)

III

20 year value of product 150,000 150,000 150,000

20 year operating cost 73,817 86,560 67,971

Plant additions 4,006 2,297 6,237

Taxes 17,763 15,164 18,644

Subtotal Deductions 95,586 104,021 92,852

Cash Flow 54,414 45,979 57,148

Discounted Rate of Return 9% 12% 12%

The following three tables illustrate the full calculation of the 20 year cash flow •

- 5 -


PRELIMINARY CASH FLOW

CASE I OPEN PIT MINING

j. 2 ^ .i S J3 2 J J'lP.iilSJ.SJ^J.SJLejn.lSJJJ^O TOTAL

Revenue(5l,000) 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 150,000ExpensesMining 1,618 1,618 1,713 1,798 1,794 1,710 1,580 781 693 671 708 840 770 737 737 688 688 638 638 638 21,058Other 2.080 2,080 2.645 3,214 3,185 2,645 2,645 2,970 2,430 2.295 2,160 2,970 2,540 2,700 2,700 2,700 2,700 2.700 2,700 2.700 52,759Total 3.698 3.698 4.358 5,012 4,979 4,355 4,225 3,751 3,123 2,966 2,868 3,810 3.310 3,437 3,437 3,388 3,388 3,338 3-338 3 338 73,817

Gross Oper.Prof. 3 802 3,802 3.142 2.488 2,521 3,145 3,275 3,749 4,377 4,534 4,632 3,690 4,190 4,063 4.063 4,112 4,112 4,162 4.162 4.162 76.183Prov. Tax 191 217 267 106 122 198 226 288 367 394 412 302 364 357 357 354 370 372 378 375 6.017Income 3.611 3.585 2,875 2,382 2,399 2,947 3,049 3,461 4,010 4,140 4,220 3,388 3,826 3,706 3,706 3,758 3,742 3,790 3,784 3,787 70,166

Cap. Cost Allow. - 1,749 3,102 2,274 1,687 1,257 1,210 908 702 643 558 444 502 374 423 15,833Pre-Prod. -__ -__ -__ 2,382 2,399 2,947 1,300 -__ -__ -__ -__ -__ -__ -__ -__ -__ -__ -__ -__ -__ 9,028

Tax Profit 3.611 3 585 2,875 - 359 1,736 2,453 2,963 2,178 2,918 3,004 3,063 3,200 3,298 3,288 3,410 3,364 45,305

Fed. Tax --__ --__ --__ ^-__ ^-__ --__ --__ 120 579 818 988 726 973 1,001 1,021 l, 067 1,099 1,096 1.137 1,121 11,746

Prof.After Tax 3,611 3.585 2.875 - 239 1,157 1,635 1,975 1,452 1,945 2,003 2,042 2,133 2,199 2,192 2 273 2,243 33,559

Plant Addition - 23 20 508 328 483 15 210 20 34 - 870 - 35 340 210 - 480 - 430 4.006

Net Profit 3.611 3.562 2,855 (508) (328) (483) (15) 29 1,137 1,601 1,975 582 1,945 1,968 1,702 1,923 2,199 1,712 2,273 1,813 29,553

Cash Flow 3,611 3.562 2,855 1,874 2,071 2,464 3,034 3,131 3,411 3,288 3,232 1,792 2,853 2,670 2,345 2,481 2,643 2,214 2,647 2,236 54,414

Cum.Cash Flow 3,611 7.173 10.028 11,902 13.973 16,437 19,471 22,602 26,013 29,301 32,533 34,325 37,178 39,848 42,193 44.674 47,317 49,531 52,178 54,414

Payback 8.5 years Return on Investment 9%

•



CASE I OPEN PIT MINING

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 TOTAL

Revenue($l,OOO) 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 150,000 Expenses

Mining Other Total

1,618 2,080 3.698

Gross Oper.Prof. 3.802 Provo Tax 191 Income 3.611

Cap. Cost Allow. Pre-Prod.

1.618 2,080 3.698

3.802 217

3.585

1,713 2.645 4.358

3.142 267

2,875

Tax Profit 3.611 3 585 2,875

Fed. Tax

Prof.After Tax 3,611 3.585 2.875

Plant Addition 23 20

Net Profit 3.611 3,562 2.855

1,798 3.214 5,012

1,794 3,185 4,979

2.488 2,521 106 122

2,382 2,399

1. 710 2.645 4;355

3,145 198

2,947

1,580 2,645 4,225

3,275 226

3,049

781 2,970 3,751

693 2,430 3,123

671 2 1 295 2,966

3,749 4,377 4,534 288 367 394

3,461 4,010 4,140

708 2,160 2,868

4,632 412

4,220

840 2,970 3,810

770 2,540 3,310

3,690 4,190 302 364

3,388 3,826

737 2,700 3,437

4,063 357

3,706

737 2,700 3,437

4.063 357

3,706

688 2,700 3,388

4,112 354

3,758

688 2,700 3,388

4,112 370

3,742

638 2.700 3,338

4,162 372

3,790

638 2,700 3.338

4.162 378

3,784

638 2.700 3 338

4.162 375

3,787

21,058 52.759 73,817

76.183 6.017

70,166

1,749 3,102 2,274 1,687 1,257 1,210 2,382 2,399 2,947 1,300

908 702 643 558 444 502 374 423 15,833 9,028

359 1,736 2,453 2,963 2,178 2,918 3,004 3,063 3,200 3,298 3,288 3,410 3,364 45,305

120 579 818 988 726 973 1,001 1,021 1,067 1,099 1,096 1.137 1,121 11,746

239 1,157 1,635 1,975 1,452 1,945 2,003 2.042 2,133 2,199 2,192 2 273 2,243 33,559

508 328 483 15 210 20 34 870 35 340 210 480 430 4.006

(508) (328) (483) (15) 29 1,137 1,601 1,975 582 1,945 1,968 1,702 1,923 2,199 1,712 2,273 1,813 29,553

Cash Flow 3,611 3.562' 2.855 1,874 2,071 2,464 3,034 3,131 3,411 3,288 3,232 1,792 2,853 2,670 2,345 2,481 2,643 2,214 2,647 2,236 54,414

Cum. Cash Flow 3.611 7.173 10.028 11,902 13.973 16,437 19,471 22,602 26,013 29,301 32,533 34,325 37,178 39,848 42.193 44,674 47,317 49,531 52,178 54.414

Payback Return on Investment

8.5 years 9%



CASE II UNDERGROUND MINING

Revenue (3., 000) ExpensesMiningOther

Total

i l l i .5 l l J J AQ 1]L 12 13 14 15 16 17 18 19 20

7.500 7.500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 150,000

1.800 1,800 1,800 1,800 1,800 1,800 1,800 1,800 1,800 1,800 2,106 2,106 2,106 2,106 2,106 2,106 2,106 2,106 2,106 2,106 39.0602.375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 47,500

4,175 4,175 4,175 4,175 4,175 4,175 4,175 4,175 4,1754,175 4,481 4,481 4,481 4,481 4,481 4,481 4,481 4,481 4,481 4,481 86,560

Gross Oper.Prof. 3,325

Prov.Tax 138

Income 3.187

Cap.Cost Allow. Pre.Prod.

Tax Profit Fed.Tax

3,325 3,325 3,325 3,325 3,325 3,325 3,325

164 186 201 217 230 240 246

3,161 3,139 3,124 3,108 3.095 3,085 3,079

307 3.108 2.330 1,762 1,404- -__ 2,817 - -___ -__ -__

3 187 3 161 3.139 765 1.323 1,675255 441 558

3,325 3,325

256 227

3,069 3,098

1,060 880

2,009 2,218670 739

Prof,After Tax 3.187

Plant Addition __13

Net Profit 3.174

Cash Flow 3,174

Cum.Cash Flow 3.174

3.161 3.139 - - 510

__22 __63 234 ^__57 109

3.139 3.076 (234) (57) 401

882 1,117

147 313

735 804

3.139 3,076 2.890 3,051 2,731 2,497 2,208

6,313 9.389 12279 15330 18061 20558 22766

1,339 1,479

__34 257

1,305 1,222

2,365 2,102

25131 27233

3,019 3,019 3,019 3.019 3,019 3,019 3,019 3,019 3,019 3,019 63,440

232 235 242 246 250 247 253 257 260 264 4.591

2,787 2,784 2,777 2,773 2,769 2.772 2.766 2.762 2,759 2,755 58,849

730 632 495 405 364 393 319 253 211 173 14,826^__ -z__ -Z__ .j;__ -r__ -Z__ ̂ __ -2 -Z__ ̂ __ 2.817

2,057 2,152 2,282 2.368 2,405 2,379 2,447 2,509 2,548 2,582 41,206686 717 761 789 802 793 816 836 849 861 10,573

1,371 1,435 1,521 1,579 1,603 1,586 1,631 1,673 1,699 1,721 30,633

194 234 __23 ___59 144 343 __37 -Z__ __l± ̂ L__ 2,297

1,177 1,201 1,498 1,520 1,459 1,243 1,594 1.673 1,685 1,721 28,336

1,907 1,833 1,993 1,925 1,823 1,636 1,913 1,926 1,896 1.894 45,979

29140 30973 32966 34891 36714 38350 40263 42189 44085 45979

PaybackReturn on Investment

6.3 years 1296

•

•

Revenue ($. , 000) Expenses

Mining Other

Total

Gross Oper.prof.

Prov.Tax

Income

Cap. Cost Allow. pre. Prod.

Tax Profit Fed. Tax

Prof. After Tax

Plant Addition

Net Profit

Cash Flow

Cum. Cash Flow

1 2 3 4 5 6 7 8 9 10 11 12



CASE II UNDERGROUND MINING

13 14 15 16 17 18 19 20

7.500 7.500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 7,500 150,000

1.800 2.375

4,175

1,800 1,800 1,800 1,800 1,800 1,800 1,800 1,800 1,800 2,106 2,106 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375 2,375

4,175 4,175 4,175 4,175 4,175 4,175 4,175 4,175 4,175 4,481 4,481

2,1062,106 2,375 2,375

4,481 4,481

2,106 2,106 2,106 2,106 2,106 2,106 2,375 2,375 2,375 2,375 2,375 2,375

4,481 4,481 4,481 4,481 4,481 4,481

39.060 47,500

86,560

3,325 3,325 3,325 3,325 3,325 3,325 3,325 3,325 3,325 3,325 3,019 3,019 3,019 3,019 3,019 3,019 3,019 3,019 3,019 3,019 63,440

138 164 186 201 217 230 240 246 256 227 232 235 242 246 250 247 253 257 260 264 4.591

3.187 3,161 3,139 3,124 3,108 3,095 3,085 3,079 3,069 3,098 2,787 2,784 2,777 2,773 2,769 2,772 2.766 2,762 2,759 2,755 58,849

3 187 3 161 3.139

3,187 3.161 3,139

13 22 63

307 3,108 2,330 1,762 1,404 1,060 880 2,817

730 632 495 405 364 393 319 253 211 173 14,826 2.817

765 1,323 1,675 2,009 2,218 2,057 2,152 2,282 2,368 2,405 2,379 2,447 2,509 2,548 2,582 41,206 255 441 558 670 739 686 717 761 789 802 793 816 836 849 861 10,573

510 882 1,117 1,339 1,479 1,371 1,435 1,521 1,579 1,603 1,586 1,631 1,673 1,699 1,721 30,633

234 57 109 147 313 _...;;3;.,.;;4 257 194 234 23 59 144 343 37 - 14 ------- -- -....;;;...;; -- 2,297

3.174 3.139 3,076 (234) (57) 401 735 804 1,305 1,222 1,177 1,201 1,498 1,520 1,459 1,243 1,594 1,673 1,685 1,721 28,336

3,174 3.139 3,076 2.890 3.051 2,731 2,497 2,208 2,365 2,102 1.907 1,833 1,993 1.925 1,823 1,636 1.913 1.926 1.896 1.894 45,979

3.174 6.313 9,389 12279 15330 18061 20558 22766 25131 27233 29140 30973 32966 34891 36714 38350 40263 42189 44085 45979


6.3 years 12%



CASE 111 PIT AND UNDERGROUND

i -

Revenue ExpensesMiningOtherTotal

Gross Op. Prof.

i;. 4 Mttm: Tax

income

Cap. Cost All. Pre.Prod.

. j,c....'. liiiji.it.:.:. . - , . ,. i

Tax Prof itTax '

l 2. l i Ji ^ 27,500 7,500 7,500 7,500 7,500 7,500 7,500

343 563 1,471 1,659 1,296 1,394 7082,105 2,105 2,105 2,105 2,160 2,215 2,2152,448 2,668 3,576 3,764 3,456 3,609 2,923

5.052 4,832 3,924 3,736 4,044 3,891 4,577

342 336 254 348 269 284 ; 373

4,710 4,496 3,670 3,388 3,775 3,607 4,204

- 1,742 2,841 - 3,388 3,775 1,865 -'

8 9 10 11

7,500 7,500 7,500 7,500

717 618 649 6502,270 2.240 2,240 2,2402,987 2,858 2,889 2,890

4,513 4,642 4,611 4,610

3 r 5 ' ' "3'9 r "' ' 3^9

4,318 4,245 4,212

2,155 1,604 1,301

4,203' if

979

12

7,500

6452,2152,860

4,640

406

4,234

950

11 li i5. li

7,500 7,500 7,500 7,500

645 641 645 1,1422,215 2,190 2,215 2,5922,860 2,831 2,860 3,734

4,640 4,669 4,640 3,766

Prof, After Tax Post Production 6 1 an t Add.

Net Profit

Cash Plow

CunuCash Plow

4,710 4,496 *,670 i

4,710

4,710

4,710

4,710

4,496

346

4,150

4,150

8,860

3,670

260

3,410

3,410

12,270

1,363454

909

203 328 483 ^-^^

(203) (328) (483) 909

3,185 3,447 3,124 3,750

15,455 18,902 22,026 25,776

1,983 2,641 2,911661 -' ! 880 : i 970

1,322 1,761 1,941

230 __33 340

1,092 1,728 1,601

3,247 3,332 2,902

29,023 32,355 35,257

3,224 3,2841,475 1,095

2,149 2,189

-r__ 678

2,149 1,511

3,128 2,461

38,385 40,845

4,225 4,261 ; 4^276 ^

722 862 1,146 864'l tan ' Ji

3,503 3,399 l, 1,168 1.133

2,335 2,266 1,22^9 4^747i- ' : - ' - ^1,216'."-^':^__ 1,020 1.654 __"65

2,335 1,246 (1,711) 1,682

3,057 2,108 721 2,,?46

43,903 46,011 46,732 49,278

12

7,500

1,7762,5924,368

3,132

' 2J9 :

2,913

730

1,455

289

1,166

1,896

51,174

18

7,500

1,7762,5924,368

3,132

:: 226

2,906

627

.|..:,i;.:L-!U;i.i i .t.. -- i

2,279760

i, 519

268

1,251

1*878

53,052

19

7,500

1,7762,5924,368

3,132

236

2,896

470

2,426809

20

7,500

1,7762,5924,368

3,132

244

2,888

372

2. 516839

1,617 1,677 i ' 4*

^__ __40

1,617 1,637

2,087 2,009

55,139 57,148

TOTAL

150,000

20,89045.79566,685

83,315

6, sal

76,732

17,365rt j 31*,:i.:: ( -..al'iUi:..i..aiikS":

49405112.061

36,9921,2866,237

29,469

57,148

PaybackReturn on Investment

6.5 years 12.3 y*

Revenue Expenses

Mining Other Total

1 2 3

7,500 7,500 7,500

343 563 1,471 2,105 2,105 2,105 2,448 2,668 3,576

4 5

7,500 7,500

1,659 1,296 2,105 2,160 3,764 3,456

6

7,500

1,394 2,215 3,609

7 8 9 10 11 12

7,500 7,500 7,500 7,500 7,500 7,500

708 717 618 649 650 645 2,215 2,270 2,240 2,240 2,240 2,215 2,923 2,987 2,858 2,889 2,890 2,860



CASE III PIT AND UNDERGROUND

13

7,500

645 2,215 2,860

14

7,500

641 2,190 2,831

15 16

7,500 7,500

645 1,142 2,215 2,592 2,860 3,734

17 18

7,500 7,500

1,776 1,776 2,592 2,592 4,368 4,368

19 20 TOTAL

7,500 7,500 150,000

1,776 1,776 2,592 2,592 4,368 4,368

20,890 45,795 66,685

Gro88 Op.prof. 5,052 4,832 3,924 3,736 4,044 3,891 4,577 4,513 4,642 4,611 4,610 4,640 4,640 4,669 4,640 3,166 3,132 3,132 3,132 3,132 83,315

6~'5Ed

76,732

254 269, ' '21M, 313 315 " 4l!;' ,408,t'3fS4' ii lJ!!81 i '~~9 236 244

Income 4,710 4,496 3,670 3,388 ""3,'77'5 ,3,607 4,204 4,318 4,24~ 4,212 4,203 4,234 4,225 4'2~1 ,~. 216 hA'i85 2,913 ~,906 2,896 2,888 I, I ~~l

Cap.Cost All. Pre. Prod.

Tax Profit Fed. Tax

,., I' i I : • I .10..,;: Il~.

4,,710 4,496 , '

Prof.After Tax 4,710 4~496 poat, Production -~laDt Add. 346

1,742 3,388 3,775 1,865

.Ii 'I" 'J. "' Jii";J:: :~ii!~~!-ji' :,: ",;:,~\:'t :~!i!'J i lid"~, ,~ ~I

3..610 ... , ..

411, /

3,670

260 203 328 483

Net Profit 4,710 4,150 3,410 (203) (328) (483)

2,841 2,155 1,604 1,301 ~79 950 722 862 1, 1~ ';~4 730 627 470 372 17,365

-~ '. ""j,i"!'!",i;,i: "~ ,J ..i"i,;"",,,:lii i, .i:;jJli~"LMI:" " lie i"if',,~L:!,;..iiI,: ,tL,lhJ',:iii ",I" :il,~,,~:;:~.r~ ,: ,4f.lUtl"!'::~'~:":Wi;i"l""';:ji,' ,"~:, ' ., " ,. it: ,,~g t~"i~" 1,363

454

909

1.983 I 2.641 2,911 3,224 3,284 3~503 ,3,399 1;,~ 'I~.~, 2,1, a,183 '2,279 2,426,2i~516 49~05~; 661' 880', 970 1,"5 1,095 1,16$ 1,133 ~' 'I ~,i 14 728' 760 809 839 12,061

iU!j ,jj ,iii j 14<'" "

1,322 1,761 1,941 2,149 2,189 2,335 2,2661,22i9'1:"'1~:t"47' 1,45:5 i,519 1,617 1,677 1,286': -!~

1,020 1,6$4 1'165 ...

230 33 340 678 289 268 40 .j .I':b I

36,992 1,2_ 6,237

909 1,092 1,728 1,601 2,149 1,511 2,335 1,246 (1,71i1) 1,.6s2 1,637 29,469 I

C¥b .P1ow 4,710 4,150 3,410 3,185 3,447 3.12. 3,750 3,247 3,3322.,;902 3,128 2,461 ~!' 051· ,2,;108 72~'~'l~

1,166 '1,251 1,617

1, 8~ ·',,1 ,878 2~, 087 2,009 57,148 i\

Cum.Cash Flow 4,710 8,860 12,270 15,455 18,902 22,026 25,776 29,023 32,355 35,257 38,385 40,845 43,903 46,011 46,732 49,278 51,174 53,052 55,139 57,148


6.5 years 12.3 %

,,·1

2.0 INTRODUCTION

2.1 Scope and Responsibilities

Canadian Bechtel Limited was engaged by Argor Explorations Limited on January 6th, 1967 to develop preliminary plans, flowsheets and mining methods together with capital and operating costs for their proposed mine and mill located south of Moosonee, Ontario. As part of the terms of reference Bechtel was required to make a com parison of open pit versus underground mining.

At a meeting held in Toronto on May 17th, 1967 it was agreed that the mining studies would be carried out for the following three cases:

1. All open pit

2. All underground

3. Open pit followed by underground

The responsibilities of Argor and Bechtel in the preparation of this interim report have been as follows:

Argor Explorations Limited

a) Supply of diamond drill hole logs, assay data, geological plan and geological cross-sections with interpreted ore-waste boundaries.

Canadian Bechtel Limited

a) Calculation of geological and recoverable ore reserves.

b) Preliminary design of open pit and underground mining systems.

- 6 -•

2.0 INTRODUCTION

2.1 Scope and Responsibilities

Canadian Bechtel Limited was engaged by Argor Explorations Limited on January 6th, 1967 to develop preliminary plans, flowsheets and mining methods together with capital and operating costs for their proposed mine and mill located south of Moosonee, Ontario. As part of the terms of reference Bechtel was required to make a comparison of open pit versus underground mining.

At a meeting held in Toronto on May 17th, 1967 it was agreed that the mining studies would be carried out for the following three cases:

1. All open pit

2. All underground

3. Open pit followed by underground

The responsibilities of Argor and Bechtel in the preparation of this interim report have been as follows:

Argor Explorations Limited

a) Supply of diamond drill hole logs, assay data, geological plan and geological cross-sections with interpreted ore-waste boundaries.

Canadian Bechtel Limited

a) Calculation of geological and recoverable ore reserves.

b) Preliminary design of open pit and underground mining systems •

- 6 -

c) Estimate of prepreduction costs includingmobilization, access road, camp construction and maintenance, stripping of the overburden, construction of airstrip and demobilization.

d) Development of a factored plant capital costestimate.

e) Estimate of mining operation costs.

The rough estimate of concentrating, general overhead, transportation, sales and head office and contingency expenses were developed jointly by Argor and Bechtel.

- 7 -

•

•

c) Estimate of preproduction costs including mobilization, access road, camp construction and maintenance, stripping of the overburden, construction of airstrip and demobilization.

d) Development of a factored plant capital cost estimate.

e) Estimate of mining operation costs.

The rough estimate of concentrating, general overhead, transportation, sales and head office and contingency expenses were developed jointly by Argor and Bechtel •

- 7 -

3.0 CAPITAL AND OPERATING COST ESTIMATES

3.1 Basis of Capital Cost Estimate

No formal estimate has been prepared of the cost of the three plants. The capital costs have been developed by relating the cost of flotation plants of similar capacity to the situation at Moosonee. It is believed that the figures are sufficiently accurate for comparison purposes, but they should not be used for absolute costs. The plant capital costs are inclusive of engineering, supervision and applicable contractor's fees. The estimate excludes all exploration costs, land purchase and easement costs, legal expenses, the cost of the feasibility study and other expenses incurred to date by the owner.

The pre-production costs for the open pit case involving the removal of 4.9 million cubic yards have been estimated by C. A. Pitts General Contractor Limited and independently by Bechtel. The costs as estimated by Bechtel are used in the estimate of the total cash requirements. The Bechtel estimate includes an additional 346,000 for the rock work required to sink the ramp and a slot into the orebody. The costs of the site access road, the fuel facilities, the airfield and the mine pre-production for Case II have been estimated by Bechtel.

Mine equipment costs, both underground and open pit, shaft excavation and equipment costs, hoisting equipment and the headframe costs have been obtained either directly from the manufacturers involved, or from past quotes previously submitted to Bechtel.

The following table summarizes the order of magnitude total cash requirements for each of the three cases:

- 8 -

•

•

3.0 CAPITAL AND OPERATING COST ESTIMATES

3.1 Basis of Capital Cost Estimate

No formal estimate has been prepared of the cost of the three plants. The capital costs have been developed by relating the cost of flotation plants of similar capacity to the situation at Moosonee. It is believed that the figures are sufficiently accurate for comparison purposes, but they should not be used for absolute costs. The plant capital costs are inclusive of engineering, supervision and applicable contractor's fees. The estimate excludes all exploration costs, land purchase and easement costs, legal expenses, the cost of the feasibility study and other expenses incurred to date by the owner.

The pre-production costs for the open pit case involving the removal of 4.9 million cubic yards have been estimated by C. A. Pitts General Contractor Limited and independently by Bechtel • The costs as estimated by Bechtel are used in the estimate of the total cash requirements. The Bechtel estimate includes an additional $46,000 for the rock work required to sink the ramp and a slot into the orebody. The costs of the site access road, the fuel facilities, the airfield and the mine pre-production for Case II have been estimated by Bechtel.

Mine equipment costs, both underground and open pit, shaft excavation and equipment costs, hoisting equipment and the headframe costs have been obtained either directly from the manufacturers involved, or from past quotes previously submitted to Bechtel.

The following table summarizes the order of magnitude total cash requirements for each of the three cases:

- 8 -

SUMMARY OF ORDER OF MAGNITUDE CAPITAL COST ESTIMATE

ARGOR EXPLORATIONS

PLANT COSTSGeneral Yard FacilitiesWater ServicesTailings DisposalPower PlantOffices and LaboratoriesHeadfrarae and ShopsWarehouse and ShopsPrimary CrushingConcentratorContingency at 15?6

Sub-Total Plant

PRE-PRODUCTION COSTSAccess Road to SiteFuel, Oil Se Grease

FacilitiesStream DiversionMine Pre-ProductionAirfield Se Road

Sub-Total Pre-production

OTHER COSTSMine EquipmentTowns iteInitial InventoryWorking Capital

Sub-Total Other

TOTAL PROJECT COST

IOpen PitMining

X $1000

545350290990250-

445870

3,7901,120

8,650

1,854

773382

8, 646616

12,271

2,016300325650

3,291

$ 24,212

LIMITED

IIUnderground

MiningX S1000

360350235

1,457250

1,939-254

3,7851,292

9,922

2,178

359

2,8171,298

6,652

774300291650

2,015

$ 18,589

IIIO.P. and U.G

MiningX $1000

545350290990250-

445870

3,7901,120

8,650

1,854

773382

8,646616

12,271

1,195300225650

2,370

5 23,291

- 9 -•

SUMMARY OF ORDER OF MAGNITUDE CAPITAL COST ESTIMATE


PLANT COSTS General yard Facilities Water Services Tailings Disposal Power Plant Offices and Laboratories Headframe and Shops Warehouse and Shops Primary Crushing Concentrator Contingency at 15%

Sub-Total Plant

PRE-PRODUCTION COSTS Access Road to Site Fuel, Oil & Grease

Facilities Stream Diversion Mine Pre-Production Airfield & Road

Sub-Total Pre-production

OTHER COSTS Mine Equipment Townsite Initial Inventory Working Capital

sub-Total Other

TOTAL PROJECT COST

I Open pit

Mining X $1000

545 350 290 990 250

445 870

3,790 1,120

8,650

1,854

773 382

8,646 616

12,271

2,016 300 325 650

3,291

$ 24,212

- 9 -

II Underground

Mining X $1000

360 350 235

1,457 250

1,939

254 3,785 1,292

9,922

2,178

359

2,817 1,298

6,652

774 300 291 650

2,015

$ 18,589

III O.P. and U.G.

Mining X $1000

545 350 290 990 250

445 870

3,790 1,120

8,650

1,854

773 382

8,646 616

12,271

1,195 300 225 650

2,370

$ 23,291

3.1.1 Plant Costs

Initial plant costs are estimated for the three cases to be $8 / 650,000, $9 / 922,000 and ?8, 650,000 respectively. Higher capital is required for the underground case due to increased electric power requirements and the cost of the hoisting equipment and shaft house.

3.1.2 Pre-Production Costs

Pre-Production costs are estimated for the three cases to be $12,271,000 f ?6,652,000 and ?12,271,000 respectively.

The following table summarizes the costs of various items which are contained in the pre-production estimate:

Case I S IIIPitts Bechtel Case II

Mobilization St Demobilization 167; 000 160,800 102,700Access Roads 933,000 865,900 865,900Airstrip Construction 83,000 242,200 470,600 Ferry - Build Wharves Se

Operate 204,000 230,000 220,100 Service Vehicles - Supply

S. Operate 218,000 201,200 100,000 Fuel Storage Facilities 60,000 95,300 93,400 Air Freight s Personnel 129,000 41,500 24,500 Camp Construction S Operation 480,000 757,900 456,100 Utilities - Power, Water, Sew

age, etc. 197,000 238,700 94,000 Clear Production Area 32,000 7,500 River Diversion b Dewatering 608,000 256,300 Excavation of Overburden 3,780,000 4,028,500 Indirect Costs 1,134,000 1,077,600 263,700 Plant Write-Off 2,690,000 2,421,600 544,000 Fee b Contingency 1,150,000 1.600,000 600, OOP

Sub-Total 11,865,000 12,225,000 3,835,000

Mine Rockwork -____ 46,000 2.817,000

TOTAL $11,865,000 $12,271,000 $6,652,000

- 10 -

3.1.1 Plant Costs

Initial plant costs are estimated for the three cases to be $8,650,000, $9,922,000 and $8,650,000 respectively. Higher capital is required for the underground case due to increased electric power requirements and the cost of the hoisting equipment and shaft house.

3.1. 2 Pre-Production Costs

Pre-Production costs are estimated for the three cases to be $12,271,000, $6,652,000 and $12,271,000 respectively.

The following table summarizes the costs of various items which are contained in the pre-production estimate:

Mobilization & Demobilization Access Roads Airstrip Construction Ferry - Build Wharves &

Operate Service Vehicles - Supply

& Operate Fuel storage Facilities Air Freight & Personnel Camp Construction & Operation utilities - Power, Water, Sew-

age, etc. Clear Production Area River Diversion & Dewatering Excavation of Overburden Indirect Costs Plant Write-Off Fee & Contingency

Sub-Total

Mine Rockwork

Case pitts 167;000 933,000

83,000

204,000

218,000 60,000

129,000 480,000

197,000 32,000

608,000 3,780,,000 1,134,000 2,690,000 1,150,000

11,865,000

I & III Bechtel 160;800 865,900 242,200

230,000

201,200 95,300 41,500

757,900

238,700 7,500

256,300 4,028,500 1,077,600 2,421,600 1,600,000

12,225,000

46,000

TOTAL $11,865,000 $12,271,000

- 10 -

Case II 102,700 865,900 470,600

220,100

100,000 93,400

'24,500 456,100

94,000

263,700 544,000 600,000

3,835,000

2,817,000

$6,652,000

It should be noted that variations in the individual cost centres between the Pitts and Bechtel estimates can generally be reconciled by different allocation of costs or by differences in some of the basic assumptions. In general the estimate comparisons are considered excellent for this stage of the project development.

3.1.3 Mine Equipment

Mine equipment costs for the three cases are estimated to be ?2,016,000, $774,000 and ?1,195,000 respectively. The higher equipment costs for Case I over Case III are due to the different stripping schedules developed. Case I requires stripping of the overburden to continue in the first year of operation, whereas in Case III overburden stripping commences during the second year. Capital for the equipment of the latter case has been taken from the annual operating profit.

3.1.4 Townsite

An allowance of $300,000 has been included in each case to cover the cost of housing for senior personnel only.

3.1.5 Initial Inventory

An allowance of $325,000, $291,000 and ?225,000 hasbeen included to cover the cost of major equipment spares,

3.1.6 Working Capital

An allowance of ?650,000 has been included in each case to cover the Owner's cost during construction and approximately 30 days of accounts receivable.

- 11 -

3.1.3

3.1.4

3.1.5

3.1.6

It should be noted that variations in the individual cost centres between the Pitts and Bechtel estimates can generally be reconciled by different allocation of costs or by differences in some of the basic assumptions. In general the estimate comparisons are considered excellent for this stage of the project development.

Mine Equipment

Mine equipment costs for the three cases are estimated' to be $2,016,000, $774,000 and $1,195,000 respectively. The higher equipment costs for Case I over Case III are due to the different stripping schedules developed. Case I requires stripping of the overburden to continue in the first year of operation, whereas in Case III overburden stripping commences during the second year. Capital for the equipment of the latter case has been taken from the annual operating profit.

Townsite

An allowance of $300,000 has been included in each case to cover the cost of housing for senior personnel only.

Initial Inventory

An allowance of $325,000, $291,000 and $225,000 has been included to cover the cost of major equipment ~pares.

Working Capital

An allowance of $650,000 has been included in each case to cover the Owner's cost during construction and approximately 30 days of accounts receivable.

- 11 -

3.2 OPERATING COST ESTIMATES

3.2.1 Open Pit

Operating cost estimates for the open pit operations are based on the removal of 3000 tons per day of ore on a two shift basis, approximately 1,530,000 cubic yards per year of overburden and approximately 360,000 cubic yards per year of waste rock.

The following table summarizes the unit mining costs:

Overburden Ore Waste $/C.Y. $/Ton S/Ton

Drilling .013 .035 .035Blasting .015 .075 .075Loading .046 .065 .065Hauling .241 .126 .126Haul Road Maintenance .046 .010 .010Slope Maintenance .227 .010 .010Tractors - .020 .020Drainage - .010 .010General Pit - .020 .020Supervision .013 .035 .035Engineering - .034 .034

Total .601 .440 .440

3.2.2 Underground

Operating cost estimates for the underground operations are based on the removal of 3000 tons per day of ore. In Case III the first 15 years are by open pit mining, the costs of which are included in the previous section,

The following table summarizes the unit underground mining costs:

- 12 -

•

•

3.2

3.2.1

OPERATING COST ESTIMATES

Open pit

Operating cost estimates for the open pit operations are based on the removal of 3000 tons per day of ore on a two shift basis, approximately 1,530,000' cubic yards per year of overburden and approximately 360,000 cubic yards per year of waste rock.

The following table summarizes the unit mining CostSI

Overburden $/c. Y.

Ore $/Ton

Waste $/Ton

Drilling Blasting Loading Hauling

.013

.015

.046

.241

.046

.227

.035

.075

.065

.126

.035

.075

.065

.126 Haul Road Maintenance Slope Maintenance Tractors Drainage General Pit Supervision Engineering

Total

3.2.2 Underground

.010 .010

.010 .010

.020 .020

.010 .010

.020 .020 .013 .035 .035

.:..2l! .034

.601 .440 .440

Operating cost estimates for the underground operations are based on the removal of 3000 tons per day of ore.In Case III the first 15 years are by open pit mining, the costs of which are included in the previous section.

The following table summarizes the unit underground mining costs:

- 12 -

CASE III Year 16-20

Hoist .22 .30Pump .03 .05 .02Ventilation .03 .05 .02Crusher .10 .12 .08Fill .10 .12 .10Development .41 .46 .35Timber .02 .03 .02Drill and Blast .42 .42 .42Haul Se Load .20 .30 .44Supervision Se Eng'g .12 .12 .08Power, Air, Water .10 .12 .06Rock Bolt .03 .04 .03Training .05 .05 .05Miscellaneous .22 .22 .18

Total $ 2.05 5 2.40 $ 1.85

3.2.3 Other Costs

Concentrating, overhead, transportation and miscellaneous expenses are estimated to total ?2.70 per ton of ore milled, as follows:

Concentrating 1.40

General Overhead .70

Transportation .10

Sales St Head Office .25

Contingency

Total

- 13 -

tit

CASE II CASE III Year Year Year 1-10 11-20 16-20

Hoist .22 .30 Pump .03 .05 .02 ventilation .03 .05 .02 Crusher .10 .12 .08 Fill .10 .12 .10 Development .41 .46 .35 Timber .02 .03 .02 Drill and Blast .42 .42 .42 Haul & Load .20 .30 .44 Supervision & Eng'g .12 .12 .08 Power, Air, water .10 .12 .06 Rock Bolt .03 .04 .03 Training .05 .05 .05 Miscellaneous .22 ~ ~

Total $ 2.05 $ 2.40 $ 1.85

3.2.3 other Costs

Concentrating, overhead, transportation and miscellaneous expenses are estimated to total $2.70 per ton of ore milled, as follows:

concentrating 1.40

General Overhead .70

Transportation .10

Sales & Head Office .25

Contingency .25

Total $ 2.70

- 13 -

4.0 ORE RESERVES

4.1 Geological Reserves

Utilizing the drill hole data, assay data and geological sections showing the ore-waste divisions, geological ore reserves have been calculated for the pyrochlore and columbite ore zones between section 6+OOS and 18+OON.

Reinterpretation of the extent of the columbite zone will change the pyrochlore-columbite tonnageratio.

The following table summarizes the total geological ore reserves:

SUMMARY OF GEOLOGICAL RESERVES

SECTION 6+OOS TO SECTION 18+OON

Section

6S-6N

Ore Type

Pyrochlore

6N-18N W.Limb

E.Limb

P + Columbite

Pyrochlore

Pyrochlore

Columbite

Total * Estimated only.

Location

Crown Pillar 175 - 400 400 - 475 Sill Pillar 525 - 825 Crown Pillar 175 - 400 400 - 475 Crown Pillar 175 - 400 400 - 475 Sill Pillar 525 - 825 Crown Pillar 175 - 400 400 - 475 Sill Pillar 525 - 825 Crown Pillar 175 - 400 400 - 475 Sill Pillar 525 - 825

Very little drill

Tons x 10- 6

3.1768.8182.8252.00211.2980.9812.0950.4060.7332.7440.9440.6823.6831.4293.4111.4191.1208.1870.9322.7240.28550.3790.71761.560information.

*Cb2C

0.640.570.570.590.590.410.390.480.400.550.440.460.460.510.460.490.480.480.390.420.480.450.450.52

t

55

*

*

*

*

- 14 -

4.0 ORE RESERVES

4.1 Geological Reserves

Section

6S-6N

6N-18N W.Limb

E.Limb

utilizing the drill hole data, assay data and geological sections showing the ore-waste divisions, geological ore reserves have been calculated for the pyroch1ore and columbite ore zones between section 6+00S and l8+00N.

Reinterpretation of the extent of the columbite zone will change the pyrochlore-columbite tonnage ratio.

The following table summarizes the total geological ore reserves:

SUMMARY OF GEOLOGICAL RESERVES

SECTION 6+005 TO SECTION l8+00N

Ore Type Location Tons %Cb20 5 x 10-6

Pyroch1ore Crown Pillar 3~176 0~64

175 - 400 8.818 0.57 400 - 475 2.825 0.57 Sill pillar 2.002 0.59 525 - 825 11. 298 0.59 *

P + Columbite Crown pillar 0.981 0.41 175 - 400 2.095 0.39 400 - 475 0.406 0.48

pyrochlore Crown pillar 0.733 0.40 175 - 400 2.744 0.55 400 - 475 0.944 0.44 Sill pillar 0.682 0.46 525 - 825 3.683 0.46 *

Pyrochlore Crown Pillar 1.429 0.51 175 - 400 3.411 0.46 400 - 475 1.419 0.49 Sill pillar 1.120 0.48 525 - 825 8.187 0.48 *

P + Co1umbite Crown pillar 0.932 0.39 175 - 400 2.724 0.42 400 - 475 0.2855 0.48 Sill pillar 0.379 0.45 525 - 825 0.717 0.45 *

Total 61. 560 0.52

* Estimated only. Very little drill information.

- 14 -

4.2 RECOVERABLE RESERVES

4.2.1 Case I

Recoverable ore reserves for Case I are based on a pit floor elevation 400 feet below the surface. These reserves are estimated to be 23,039,000 short tons with an average grade of Q.51%

To recover this tonnage an estimated 14,720,000 cubic yards of overburden and 7,257,000 cubic yards of waste rock must be removed.

These reserves are inclusive of the columbite ore zone and are classified as drill indicated.

The following table summarizes the ore reserves and stripping requirements by 50 foot intervals for Case I,

CASE I

RECOVERABLE ORE RESERVES AND STRIPPING QUANTITIES

Ore Waste-6 -6

Elevation Tons x 10 96Cbo05 Cu.Yd. x 10

*f 100 - - 14.720+ 150 4.249 0.54 2.063*t- 200 4.099 0.53 1.744* 250 3.848 0.50 1.388* 300 3.706 0.50 1.208* 350 3.648 0.50 .675t 400 3.489 0.50 .359

Total 23.039 0.51 21.977

4.2.2 Case II

Potential recoverable reserves are estimated to be 17,206,000 short tons at a grade of 0.58* Cb2^5 for an underground mine planned in the area bounded by

- 15 -

4.2

4.2.1

RECOVERABLE RESERVES

Case I

Recoverable ore reserves for Case I are based on a pit floor elevation 400 feet below the surface. These reserves are estimated to be 23,039,000 short tons with an average grade of 0.51% Cb205~

To recover this tonnage an estimated 14,720,000 cubic yards of overburden and 7,257,000 cubic yards of waste rock must be removed.

These reserves are inclusive of the co1umbite ore zone and are classified as drill indicated.

The following table summarizes the ore reserves and stripping requirements by 50 foot intervals for Case I.

CASE I


Ore -6

Waste -6

Elevation Tons x 10 %Cb,OS cu. Yd. x 10

+ 100 + 150 + 200 + 250 + 300 + 350 + 400

4.2.2

14.720 4.249 0.54 2.063 4.099 0.53 1.744 3.848 0.50 1.388 3.706 0.50 1.208 3.648 0.50 .675 3.489 ~ .359

Total 23.039 0.51 21.977

Case II

Potential recoverable reserves are estimated to be 17,206,000 short tons at a grade of 0.58% Cb205 tor an underground mine planned in the area bounded by

- 15 -

sections 6-fOOS to 6+OON and down to the 825 elevation at an extraction recovery of 7596. The following table summarizes these reserves:

CASE II

RECOVERABLE ORB RESERVES

-6 Tons x 10 96Cb205

Drill Indicated (-175'* 475') 8.732 0.57

Possible (-525 1 * 825') 8.474 0.59

Total 17.206 0.58

These reserves are exclusive of the columbite ore zone.

Additional ore, in case of shortages in tonnage or grade, can be obtained from the following sources:

- access provided to east and west limbs- extension beyond 6-fOOS to 8-K)OS at least.

4.2.3 Case III

Recoverable ore reserves are based on a pit floor elevation of 430 followed by a ramp in ore to 475 and an underground operation in the area above the 475 elevation bounded by sections 7-fOON to 16-HDON in the east and west limbs.

Open pit ore reserves are estimated to be 12,714,000 short tons at a grade of Q.59% Cb2O5* They are classified as drill indicated reserves.

To recover this tonnage an estimated 8,859,000 cu. yds. of overburden and 4,489,000 cu. yds. of waste rock must be removed.

- 16 -

sections 6+005 to 6+OON and down to the 825 elevation at an extraction recovery of 75%. The following table summarizes these reserves:

CASE II

RECOVERABLE ORB RESERVES

-6 Tons x 10

Drill Indicated (-175'+ 475') 8.732 0.57

possible (-525' + 825') 8.474 0.59

Total 17.206 0.58

These reserves are exclusive of the columbite ore zone.

Additional ore, in case of shortages in tonnage or grade, can be obtained from the following sources:

4.2.3

- access provided to east and west limbs - extension beyond 6+005 to 8+005 at least.

Case III

Recoverable ore reserves are based on a pit floor elevation of 430 followed by a ramp in ore to 475 and an underground operation in the area above the 475 elevation bounded by sections 7+00N to l6+OON in the east and west limbs.

Open pit ore reserves are estimated to be 12,714,000 short tons at a grade of 0.59% Cb205. They are classified as drill indicated reserves.

To recover this tonnage an estimated 8,859,000 cu. yds. of overburden and 4,489,000 cu. yds. of waste rock must be removed.

- 16 -

Underground ore reserves are estimated to be 6,365,000 tons at a grade of 0.49 96 Cb2o5 and an extraction recovery of 7596 for an underground mine planned in the area bounded by 6N to 16N down to the 475 horizon. The ore reserves are exclusive of the columbite ore zone and are classified as drill indicated.

The following table summarizes the ore reserves for Case III,

CASE III


Ore Waste———————-6———————— —————————^6-

Elevation Tons x 10 96Cb2O5 Cu.Yd. x 10

Open Pit+ 100 - - 8.859* 150 2.160 0.64 1.172* 200 2.064 0.60 .932+ 250 1.956 - .670* 300 1.880 - .464+ 350 1.980 - .246* 400 1.762 0.57 .098* 450 .912 0.57 .017

Sub Total 12.714 0.59 12.458 Columbite .890

Underground+ 475 6.365 0.49 -

TOTAL 19.079 0.56 13.348

- 17 -

Underground ore reserves are estimated to be 6,365,000 tons at a grade of 0.49 % Cb20S and an extraction recovery of 75% for an underground mine planned in the area bounded by 6N to l6N down to the 475 horizon. The ore reserves are exclusive of the columbite ore zone and are classified as drill indicated.

The following table summarizes the ore reserves for Case III.

Elevation

Open pit + 100 + 150 + 200 + 250 + 300 + 350 + 400 + 450

CASE III


Ore =6

Tons x 10 %Cb20 5

2.160 0.64 2.064 0.60 1.956 1.880 1.980 1. 762 0.57

.912 0.57 -Sub Total 12.714 0.59

Columbite

Underg:round + 475 6.365 0.49

TOTAL 19.079 0.56

- 17 -

waste

Cu. Yd. x 10 -6

8.859 1.172

.932

.670

.464

.246

.098

.017

12.458 .890

13.348

5.0 MINING METHODS

5.1 OPEN PIT CASE I Se III

5.1.1 Pit Layout

The following parameters have been used in the design of the open pits:

Bench Height 37*5 feet Bench Slope 70 degrees Overall Pit Slope 55 degrees Berm Width 13 feet Road Width 50 feet in rock

60 feet in dirt Road Grade 10 percent

5 percent on curves Overburden Slope 3H to IV

The ultimate pit road exits from the pits at the south end and is located in the west wall of the pit for the upper 200 feet and in the east wall of the pit for the lower 200 feet. The location of the haul road has been selected for ease of access to the higher grade ore at the south end in the early years of operation. It avoids a potential fault zone along the eastern carbonatite contact, and permits a more logical sequence of overburden removal in the early operating years.

A 50 foot berm has been retained at the contact between the carbonatite and the siltstone-mudstone sequence. This will permit access at all times to clean up sloughed material and maintain drainage ditches.

Two 35 foot berms have been left in the overburden for road access to continue the stripping during operations. These berms are located at the base of the silty sand and midway in the siltstone-mudstone sequence at the approximate elevation of 70 feet. A twenty foot berm has been left at the base of the till.

The peat has been stripped back a distance of 50 feet

- 18 -

5.0

5.1

5.1.1

MINING METHODS

OPEN PIT CASE I & III

pit Layout

The following parameters have been used in the design of the open pits:

Bench Height 37~ feet Bench Slope 70 degrees Overall Pit Slope 55 degrees Berm Width 13 feet Road Width 50 feet in rock

60 feet in dirt Road Grade 10 percent

5 percent on curves Overburden Slope 3H to IV

The ultimate pit road exits from the pits at the south end and is located in the west wall of the pit for the upper 200 feet and in the east wall of the pit for the lower 200 feet. The location of the haul road has been selected for ease of access to the higher grade ore at the south end in the early years of operation. It avoids a potential fault zone along the eastern carbonatite contact, and permits a more logical sequence of overburden removal in the early operating years.

A 50 foot berm has been retained at the contact between the carbonatite and the siltstone-mudstone sequence. This will permit access at all times to clean up sloughed material and maintain drainage ditches.

Two 35 foot berms have been left in the overburden for road access to continue the stripping during operations. These berms are located at the base of the silty sand and midway in the siltstone-mudstone sequence at the approximate elevation of 70 feet. A twenty foot berm has been left at the base of the till.

The peat has been stripped back a distance of 50 feet

- 18 -

from the edge of the pit and an interceptor drainage ditch has been established in the till around the perimeter of the pit.

5.1.2 Pre-Product ion

To conform with the proposed mining schedules and pit designs for Cases I Se III an estimated 4,900,000 and 4,450,000 cubic yards of overburden respectively must be removed in the pre-production phase. Some of the till material will be used in the construction of tailing pond starting dykes and perimeter dykes around the perimeter of the plant site and mining area.

This material will be removed on a contract basis using 5^ cubic yard wheel mounted front end loaders in conjunction with 35 ton haulage trucks.

As soon as the overburden has been stripped to expose the western side of the orebody a ramp is sunk from the 100 elevation at the south end of the orebody to the 150 elevation. A sump is est ablished at the 150 elevation and a slot is driven from the 137 elevation eastwards to the mid-point of the ore zone. This will permit at least three working faces to be established in ore on the 137 elevation.

5.1.3 Pit Development

A preliminary mining schedule showing the annual stripping requirements has been drawn up. One of the prime considerations of this schedule has been to keep stripping requirements to a minimum in the tax free period. However, because of the large volumes of overburden and the rapid rate at which ore must be exposed in the early years, this require ment has been very difficult to meet.

The basic approach used in the development of the mining schedule has been to mine from only 2 levels at any one time in order to keep equipment movements

- 19 -

5.1.2

5.1.3

from the edge of the pit and an interceptor drainage ditch has been established in the till around the perimeter of the pit.

Pre-Production

To conform with the proposed mining schedules and pit designs for Cases I & III an estimated 4,900,000 and 4,450,000 cubic yards of overburden respectively must be removed in the pre-production phase. Some of the till material will be used in the construction of tailing pond starting dykes and perimeter dykes around the perimeter of the plant site and mining area.

This material will be removed on a contract basis using 5~ cubic yard wheel mounted front end loaders in conjunction with 35 ton haulage trucks.

As soon as the overburden has been stripped to expose the western side of the orebody a ramp is sunk from the 100 elevation at the south end of the orebody to the 150 elevation. A sump is established at the 150 elevation and a slot is driven from the 137 elevation eastwards to the mid-point of the ore zone. This will permit at least three working faces to be established in ore on the 137 elevation.

pit Development

A preliminary mining schedule showing the annual stripping requirements has been drawn up. One of the prime considerations of this schedule has been to keep stripping requirements to a minimum in the tax free period. However, because of the large volumes of overburden and the rapid rate at which ore must be exposed in the early years, this requirement has been very difficult to meet.

The basic approach used in the development of the m~n~ng schedule has been to mine from only 2 levels at anyone time in order to keep equipment movements

- 19 -

to a minimum. In the early years it has not been entirely possible to develop a full years ore requirement one year in advance. However, once the overburden removal is completed this require ment is more easily met.

The following tables show the annual ore and waste requirements for Case I and III. Additional studies will be required to optimize these schedules.

CASE I

ANNUAL ORE AND WASTE PRODUCTION

PeriodPre-prod uction

123456789

1011121314151617181920

TOTAL PIT

OreTons x 10

.061

.780

.780

.9801.1901.180.980.980

1.100.900.850.800

1.100.940

1.0001.0001.0001.0001.0001.0001.000

19.621

WasteOverburdenCu.Yd. x lQ6

4.9001.5301.5301.5401.5301.5301.530.630——————————--—

14.720

RockCu.Yd. x 10~ 6

.023

.360

.360

.360

.360

.360

.360

.775

.300

.300

.300

.360

.360

.360

.300

.300

.251

.251

.200

.200

.200

6.340

- 20 -

•

Period

to a minimum. In the early years it has not been entirely possible to develop a full years ore requirement one year in advance. However, once the overburden removal is completed this requirement is more easily met.

The following tables show the annual ore and waste requirements for Case I and III. Additional studies will be required to optimize these schedules.

CASE I


waste Ore Overburden 6 Rock

10-6 -6 Tons x Cu. Yd. x ln Cu~ Yd. x 10

Pre-prod-uction .061 4.900 .023

1 .780 1.530 .360 2 .780 1.530 .360 3 .980 1.540 .360 4 1.190 1.530 .360 5 1.180 1.530 .360 6 .980 1.530 .360 7 .980 .630 .775 8 1.100 .300 9 .900 .300

10 .850 .300 11 .800 .360 12 1.100 .360 13 .940 .360 14 1.000 .300 15 1.000 .300 16 1.000 .251 17 1.000 .251 18 1.000 .200 19 1.000 .200 20 1.000 .200

TOTAL PIT 19.621 14.720 6.340

- 20 -

CASE III


Waste

Period

Pre-prod uction

123456789

10111213141516

Ore Tons x 10'

.052

.780

.780

.780

.780

.800

.820

.820

.840

.830

.830

.830

.820

.820

.810

.810

.512

Overburden Cu.Yds. x 10

4.450

.2001.2981.500.711.700

-6 Rock Cu.Yds. x 10

-6

TOTAL PIT 12.714

5.1.4 Pit Equipment

8.859

.037

.100

.350

.420

.500

.620

.350

.350

.255

.287

.287

.287

.287

.259

.100

4.489

Drilling of the overburden, waste rock and ore will be done with air track machines capable of drilling 4 inch diameter holes.

It is anticipated that light blasting only will be required in approximately 5096 of the overburden, A powder factor of 0.2 pounds per cubic yard has been assumed for the purpose of costing.

Powder costs for ore and waste rock are based on loading the blast holes with 4096 slurry

- 21 -

CASE III


waste Ore

10-6 Overburden_

6 Rock

Period Tons x Cu. Yds. x 10 Cu. Yds. x 10-6

Pre-prod-uction

1 2 3 4 5 6 7 8 9

10 11 12 13 14 15 16

TOTAL

5.1. 4

.052 4.450 .037

.780

.780 .200 .100

.780 1.298 .350

.780 1.500 .420

.800 .711 .500

.820 .700 .620

.820 .350

.840 .350

.830 .255

.830 .287

.830 .287

.820 .287

.820 .287

.810 .259

.810 .100

.512

PIT 12.714 8.859 4.489

pit Equipment

Drilling of the overburden, waste rock and ore will be done with air track machines capable of drilling 4 inch diameter holes.

It is anticipated that light blasting only will be required in approximately 50% of the overburden. A powder factor of 0.2 pounds per cubic yard has been assumed for the purpose of costing.

Powder costs for ore and waste rock are based on loading the blast holes with 40% slurry .

- 21 -

in the toe and 60?6 ANFO in the column. A loading factor of 0.5 pounds per ton of rock.

Loading of ore and waste rock will be done with either the Cat 988 or the Eimco 916, both rubber tired front end loaders equipped with 5*5 yard buckets. The Eimco 916 is four wheeled drive articulated vehicle which can also be used in an underground operation.

Loading of the overburden will be done by a Cat 988 tire mounted front end loader equipped with a 5^ cubic yard bucket, aided by a D-8 equipped with a ripper for dozing and ripping the till. Haulage of the over burden will be done with trucks of the Euclid R-30 series type.

A 3*3 diesel shovel with backhoe and dragline attachments will be used as a utility shovel for ditching and dyke construction.

5.2 UNDERGROUND CASE II

5.2.1 Choice of Mining Method:

The Moosonee area has no previous record of mining. Beyond information yielded by the soils examination and test drilling no practical knowledge is available concerning the behaviour of overburden when exposed nor the ground conditions and amenability to stoping of the underlying carbonatite mass.

Having regard to these uncertainties any proposed mining method must at this stage be considered as tentative and in urgent need of verification at the earliest possible opportunity by direct methods - exploratory shaft, experimental stoping etc.

The required extraction demands a minimum daily production of 2823 tons of ore and the mine falls into the "medium " range. The ore is apparently of reasonably regular grade throughout without any notably

- 22 -

5.2

5.2.1

in the toe and 60% ANFO in the column. A loading factor of 0.5 pounds per ton of rock.

Loading of ore and waste rock will be done with either the Cat 988 or the Eimco 916, both rubber tired front end loaders equipped with 5~ yard buckets. The Eimco 916 is four wheeled drive articulated vehicle which can also be used in an underground operation.

Loading of the overburden will be done by a Cat 988 tire mounted front end loader equipped with a 5~ cubic yard bucket, aided by a D-8 equipped with a ripper for dozing and ripping the till. Haulage of the overburden will be done with trucks of the Euclid R-30 series type.

A 3~ diesel shovel with backhoe and dragline attachments will be used as a utility shovel for ditching and dyke construction.

UNDERGROUND CASE II

Choice of Mining Method;

The Moosonee area has no previous record of mining. Beyond information yielded by the soils examination and test drilling no practical knowledge is available concerning the behaviour of overburden when exposed nor the ground conditions and amenability to stoping of the underlying carbonatite mass.

Having regard to these uncertainties any proposed mining method must at this stage be considered as tentative and in urgent need of verification at the earliest possible opportunity by direct methods -exploratory shaft, experimental stoping etc.

The required extraction demands a minimum daily production of 2823 tons of ore and the mine falls into the "medium" range. The ore is apparently of reasonably regular grade throughout without any notably

- 22 -

high grade areas.

A final mining scheme should be evolved to include at least two separate mining horizons in view of the established thicknesses and this method should be easily adaptable to the lens-like plan of the orebody.

Block caving is recognized as the cheapest form of extraction provided (a) that the ore will cave and break into reasonable sizes due to its own weight and (b) that the subsidence of the surface will not affect the operation adversely. In this case it seems probable that the unsupported spans in this carbonatite are likely to be very large before caving would occur and that the blocks so formed would be unmanageable.

In addition since this area has very poor natural drainage and a potentially heavy water bearing horizon in the overburden it is felt that disturbance of the surface would give rise to a pumping problem of very great magnitude since once the overlying strata deliver water into the mine these entries may never be sealed off. Caving has therefore not been further examined and it is considered that the overlying strata must be held intact by a substantial crown pillar over the whole extent of the orebody.

A tentative thickness of 75' has been assumed,and the loss of ore amounts to 3.176 million tons at Q.64%. Based on the experiences of another producer having a carbonatite orebody it is possible that open stopes may be considered as an extraction method in all parts of the orebody and on successively lower horizons provided adequate pillars are left between working stopes and that working stopes are filled immediately on completion.

Some skeleton pillars must be left between filled working stopes to provide continuous support to the crown pillar although on a final salvage operation these could be removed by another method.

- 23 -

high grade areas.

A final mining scheme should be evolved to include at least two separate mining horizons in view of the established thicknesses and this method should be easily adaptable to the lens-like plan of the orebody.

Block caving is recognized as the cheapest form of extraction provided (a) that the ore will cave and break into reasonable sizes due to its own weight and (b) that the subsidence of the 'surface will not affect the operation adversely. In this case it seems probable that the unsupported spans in this carbonatite are likely to be very large before caving would occur and that the blocks so formed would be unmanageable.

In addition since this area has very poor natural drainage and a potentially heavy water bearing horizon in the overburden it is felt that disturbance of the surface would give rise to a pumping problem of very great magnitude since once the overlying strata deliver water into the mine these entries may never be sealed off. Caving has therefore not been further examined and it is considered that the overlying strata must be held intact by a substantial crown pillar over the whole extent of the orebody.

A tentative thickness of 75 1 has been assumed,and the loss of ore amounts to 3.176 million tons at 0.64%. Based on the experiences of another producer having a carbonatite orebody it is possible that open stopes may be considered as an extraction method in all parts of the orebody and on successively lower horizons provided adequate pillars are left between working stopes and that working stopes are filled immediately on completion.

Some skeleton pillars must be left between filled working stopes to provide continuous support to the crown pillar although on a final salvage operation these could be removed by another method.

- 23 -

Since the supply of skilled labour in the area might be scarce it is believed that sublevel stopes be tried in the opening phase of the mine since this method gives relatively very high output per manshiftand low cost of mining.

A tentative scheme has been laid out using a basic stope having 6 sublevels at 50' intervals, a span of 80' with a final pillar between worked stopes of 20'. Length has been limited to a maximum of 200' for single stopes and 350' for double or back- to-back working. Drilling will be performed by long hole machine from the sublevels and ore will be withdrawn, after breaking, from haulage drives placed in the dividing pillars using front end loaders tramming to a central ore-pass system.

5.2.2 Extraction Plan

Using the basic stope detailed above all ore beneath the crown pillar down to the undercut levels of these stopes, less pillars, will be removed during the first ten years.

Stopes will be arranged transverse i.e. from wall to wall across the width of the mass and stoping will start from the east wall which dips west at about 75O . Mining will be progressive from the north and south extremities retreating towards the shaft situated on 2N at 7W about 250' outside the orebody.

Fill distribution drifts will be cut in the residual pillars just below the crown pillar to enable filling with classified mill tailings to start immediately on completion of mining. Arrangements will be made. to admit a small proportion of cement in the opening and closing stages to give additional strength to floor and backs.

5.2.3 Mining Second Horizon

Provided the stoping method chosen proves satisfactory

- 24 -

Since the supply of skilled labour in the area might be scarce it is believed that sublevel stopes be tried in the opening phase of the mine since this method gives relatively very high output per manshift and low cost of mining.

A tentative scheme has been laid out using a basic stope having 6 sublevels at 50~ intervals, a span of 80' with a final pillar between worked stopes of 20'. Length has been limited to a maximum of 200' for single stopes and 350' for double or backto-back working. Drilling will be performed by long hole machine from the sublevels and ore will be withdrawn, after breaking, from haulage drives placed in the dividing pillars using front end loaders tramming to a central ore-pass system.


Using the basic stope detailed above all ore beneath the crown pillar down to the undercut levels of these stopes, less pillars, will be removed during the first ten years.

stopes will be arranged transverse i.e. from wall to wall across the width of the mass and stoping will start from the east wall which dips west at about 750 •

Mining will be progressive from the north and south extremities retreating towards the shaft situated on 2N at 7W about 250' outside the orebody.

Fill distribution drifts will be cut in the residual pillars just below the crown pillar to enable filling with classified mill tailings to start immediately on completion of mining. Arrangements will be made. to admit a small proportion of cement in the opening and closing stages to give additional strength to floor and backs.

5.2.3 Mining Second Horizon

Provided the stoping method chosen proves satisfactory

- 24 -

and further exploration below shows no radical change in ground conditions the same method is proposed for the second lift leaving a sill or floor pillar to provide a safe roof for the second phase. A distance of 50' is suggested.

It should be stressed that reserve tonnages and grades in the second horizon have a lower order of reliability than the first mining horizon since less diamond drill information is available. Further development and drilling from the upper level will be necessary to prove this latter 10 year reserve adequately.

5.2.4 Details of Mining Sequence

5.2.4.1 Development

i) Shaft

A vertical shaft is proposed to be sunk by contractors from a position on 2N at 7W approximately 250' from the orebody and in a non-mineralized calcite bearing carbonatite. This position is approximate at this stage until further detailed knowledge of precise ground conditions has been obtained by closely spaced diamond drilling.

The location has been chosen for the following reasons.

a) It is situated on the approximate axis of the centre of gravity of the orebody between 6S and 6N from which the extraction for the first 20 years will take place.

b) The orebody at the west side of the ore zone, adjacent to the shaft position, has nearly vertical walls.

c) Between the carbonatite in which the shaft will be sunk and the orebody is a band of uralitized

- 25 -

and further exploration below shows no radical change in ground conditions the same method is proposed for the second lift leaving a sill or floor pillar to provide a safe roof for the second phase. A distance of 50' is suggested.

It should be stressed that reserve tonnages and grades in the second horizon have a lower order of reliability than the first mining horizon since less diamond drill information is available. Further development and drilling from the upper level will be necessary to prove this latter 10 year reserve adequately.

5.2.4 Details of Mining Sequence

5.2.4.1 Development

i) Shaft

A vertical shaft is proposed to be sunk by contractors from a position on 2N at 7W approximately 250' from the orebody and in a non-mineralized calcite bearing carbonatite. This position is approximate at this stage until further detailed knowledge of precise ground conditions has been obtained by closely spaced diamond drilling.

The location has been chosen for the following reasons.

a) It is situated on the approximate axis of the centre of gravity of the orebody between 6S and 6N from which the extraction for the first 20 years will take place.

b) The orebody at the west side of the ore zone, adjacent to the shaft position, has nearly vertical walls.

c) Between the carbonatite in which the shaft will be sunk and the orebody is a band of uralitized

- 25 -

pyroxenite, a harder rock and apparently not extensively faulted. In the corresponding position on the east side of the ore zone the wall rock has been identified as a gneiss and probably contains a fault or shear zone. The orebody here has a westerly dipping east wall at about 70oand even if ground problems are not experienced, all crosscuts to the orebody will increase substantially with depth.

For twenty years life, ore must be recovered down to 825 below ground level and the initial decision should therefore be taken to provide for the shaft to be sunk to its final depth of approximately 1050*.

All ore from the first mining horizon at 475 will be hauled from the stopes to a central ore pass system and conveyed by raises to the crushing station on 825 and thence by a 1000 ton shaft pocket and loading pockets, to a shaft loading station at 1025 for skip hoisting, During the second 10-year period with all ore coming from the 875 haulage level the same crushing, storage and hoisting system will be used.

The shaft will require four compartments two for 5-ton skips in balance driven by a Koepe hoist, two for a large sized cage capable of transporting major pieces of mobile equipment below ground and its counterweight and two more for a manway and service cables,pipes etc. An approximate dimension will be 22' x 12* outside timbers. Shaft stations will be cut at 150' intervals down to 475 then at 525, 675 and 825.

Below this, stations are needed as access to the ore pass system at the crusher discharge level, shaft pocket and loading pocket levels

- 26 -

pyroxenite, a harder rock and apparently not extensively faulted. In the corresponding position on the east side of the ore zone the wall rock has been identified as a gneiss and probably contains a fault or shear zone. The orebody here has a westerly dipping east wall at about 700 and even if ground problems are not experienced, all crosscuts to the orebody will increase substantially with depth.

For twenty years life, ore must be recovered down to 825 below ground level and the initial decision should therefore be taken to provide for the shaft to be sunk to its final depth of approximately 1050'.

All ore from the first mining horizon at 475 will be hauled from the stopes to a central ore pass system and conveyed by raises to the crushing station on 825 and thence by a 1000 ton shaft pocket and loading pockets, to a shaft loading station at 1025 for skip hoisting. During the second 10-year period with all ore coming from the 875 haulage level the same crushing, storage and hoisting system will be used.

The shaft will require·four compartments two for 5~ton skips in balance driven by a Koepe hoist, two for a large sized cage capable of transporting major pieces of mobile equipment below ground and its counterweight and two more for a manway and service cablee,pipes etc. An approximate dimension will be 22' x 12' outside timbers. Shaft stations will be cut at 150' intervals down to 475 then at 525, 675 and 825.

Below this, stations are needed as access to the ore pass system at the crusher discharge level, shaft pocket and loading pocket levels

- 26 -

and for the spill pocket.

From the surface to approximately 80' a concrete collar with water seal at the rock interface is allowed for while in rock the further 970' will be fitted with normal timber sets at 6"O" to 6'6" intervals depending on actual conditions.

The shaft will be designed for 3000 tons per day hoisting in 12 effective hoisting hours or 250 t.p.h. maximum. A winding speed of 1500 feet per minute with the 5 ton skips is indicated to satisfy this condition.

ii) Stope Preparation

From the shaft stations at 175, 325 and 475 crosscuts will be put out to the western ore boundaries and the west wall drives established to serve initially as diamond drilling drifts.

When precise contacts and ore grades have been established cross-cuts will be put through the orebody to establish the east wall drives on these levels from which further diamond drilling will take place to facilitate stope layout. Assuming this close-spaced drilling supports the conclusions drawn from the preparatory drilling from the surface the orebody will be divided transversely into stopes of varying lengths having a height of 300', a width of 80' and a separating pillar between stopes of 20'.

Each stope will be prepared by providing at the 475 level two haulage drives sufficiently sized to permit front end loader operation.

These haulage drives are placed in the separating pillars and are connected to the stopes on either side by boxholes leading to draw slots between 475 and the undercut level at 425.

- 27 -

and for the spill pocket.

From the surface to approximately 80' a concrete collar with water seal at the rock interface is allowed for while in rock the further 970' will be fitted with normal timber sets at 6'0" to 6'6" intervals depending on actual conditions.

The shaft will be designed for 3000 tons per day hoisting in 12 effective hoisting hours or 250 t.p.h. maximum. A winding speed of 1500 feet per minute with the 5 ton skips is indicated to satisfy this condition.

ii) stope Preparation

From the shaft stations at 175, 325 and 475 crosscuts will be put out to the western ore boundaries and the west wall drives established to serve initially as diamond drilling drifts.

When precise contacts and ore grades have been established cross-cuts will be put through the orebody to establish the east wall drives on these levels from which further diamond drilling will take place to facilitate stope layout. Assuming this close-spaced drilling supports the conclusions drawn from the preparatory drilling from the surface the orebody will be divided transversely into stopes of varying lengths having a height of 300', a width of 80' and a separating pillar between stopes of 20' •

Each stope will be prepared by providing at the 475 level two haulage drives sufficiently sized to permit front end loader operation.

These haulage drives are placed in the separating pillars and are connected to the stopes on either side by boxholes leading to draw slots between 475 and the undercut level at 425.

- 27 -

Access to the stopes is made by raise from 475 to 325 and 325 to 175 with connections to the west wall drives at these levels. At intermediate levels of 425, 375, 275 and 225 one raise serves a group of 3 stopes by short longitudinal connecting drives. Sublevels are put through from end to end and a slot raise will be developed at a convenient position depending on stope length. Some stopes - those on the east wall - have the slot raise on the east wall and others, the larger, having a central slot raise to permit working in both directions.

5.3 UNDERGROUND CASE III

5.3.1 Choice of Mining Method

The same mining method has been adopted in this scheme as in the full underground scheme and costs of mining have been taken directly from the earlier underground analysis.


Stope layout however has been freely adapted to meet the requirements dictated by the shape of the orebody and the bifurcation into two limbs between sections 6 and 16N. It has been decided to mine by sub-level stopes, therefore, in a transverse W-E direction and to develop these stopes by means of wall drives, extraction to commence in the northern extremities and to continue by retreating to the open-pit entries.

Two different stope heights are proposed. The majority have six sublevels at 50' intervals as in the full scheme but in the boundary zone between the north face of the open-pit and the underground section stopes having only 150' total height i.e. 3 x 50' sublevels have been laid out. A minimum safety crown pillar of 75' has been left above

- 28 -

•

5.3

5.3.1

5.3.2

Access to the stopes is made by raise from 475 to 325 and 325 to 175 with connections to the west wall drives at these levels. At intermediate levels of 425, 375, 275 and 225 one raise serves a group of 3 stopes by short longitudinal connecting drives. Sublevels are put through from end to end and a slot raise will be developed at a convenient position depending on stope length. Some stopes - those on the east wall - have the slot raise on the east wall and others, the larger, having a central slot raise to permit working in both directions.

UNDERGROUND CASE III

Choice of Mining Method

The same mining method has been adopted in this scheme as in the full underground scheme and costs of mining have been taken directly from the earlier underground analysis.

Extraction Plan

stope layout however has been freely adapted to meet the requirements dictated by the shape of the orebody and the bifurcation into two limbs between sections 6 and l6N. It has been decided to mine by sub-level stopes, therefore, in a transverse W-E direction and to develop these stopes by means of wall drives, extraction to commence in the northern extremities and to continue by retreating to the open-pit entries.

Two different stope heights are proposed. The majority have six sublevels at 50' intervals as in the full scheme but in the boundary zone between the north face of the open-pit and the underground section stopes having only ISO' total height i.e. 3 x 50' sublevels have been laid out. A minimum safety crown pillar of 75' has been left above

- 28 -

these stopes for their protection, this distance being measured between the final opencast face and the top of the 150" stopes and vertically above the tops of the 300' stopes.

5.3.4 Development

Due to the lack of definitive information concerning the distribution of columbite all stopes have been laid out and the development plan relates to system atic preparation regardless of columbite values. Life of the underground section is based on noncolumbite ore only however and should any proposed stopes subsequently prove to have no columbite free ore then development of these need not, in this event, take place. Stopes showing minor proportions of columbite will be worked until columbite appears and is defined and may then be abandoned and filled or, if metallurgical research has by that time solved the problem of treating a pyrochlore-columbite mixture, completely extracted and filled.

5.3.5 Access and Haulage

Access will be independently to both west and east limbs of the underground section with a central access for services and supplies only. From the established open-pit bottom proposed to go to 430' a ramp will be put down to 485' below surface and new sumps established. From 485' the main entries will be driven, after portals are secured and concreted, at -t-2% to cause the underground section to be self-draining. It is anticipated that the northern extremities will reach about 465' and the mining scheme has been based on the average horizon of 475' to coincide with calculated reserve as regards tonnage and estimated grades.

Due to the distances of the stopes from the open pit front end loaders, Eimco 916, will need to be supplemented by a low profile transporter of

- 29 -

5.3.4

5.3.5

these stopes for their protection, this distance being measured between the final opencast face and the top of the 150' stopes and vertically above the tops of the 300' stopes.

Development

Due to the lack of definitive information concerning the distribution of columbite allstopes have been laid out and the development plan relates to systematic preparation regardless of columbite values. Life of the underground section is based on noncolumbite ore only however and should any proposed stopes subsequently prove to have no columbite free ore then development of these need not, in this event, take place. stopes showing minor proportions of columbite will be worked until columbite appears and is defined and may then be abandoned and filled or, if metallurgical research has by that time solved the problem of treating a pyrochlore-colurnbite mixture, completely extracted and filled.

Access and Haulage

Access will be independently to both west and east limbs of the underground section with a central access for services and supplies only. From the established open-pit bottom proposed to go to 430' a ramp will be put down to 485' below surface and new sumps established. From 485' the main entries will be driven, after portals are secured and concreted, at +2% to cause the underground section to be self-draining. It is anticipated that the northern extremities will reach about 465' and the mining scheme has been based on the average horizon of 475 1 to coincide with calculated reserve as regards tonnage and estimated grades.

Due to the distances of the stopes· from the openpit front end loaders, Eimco 916, will need to be supplemented by a low profile transporter of

- 29 -

10 ton capacity, such as the Wagner vehicle designed for this purpose.

The main wall haulages are driven at 20' wide by 9' high to permit passing of these vehicles. Front end loaders in this scheme will transport only to stope boundaries, before transferring their loads to these vehicles. Stope drives on "475" level remain as before at 15' x 9'.

5.3.6 Loading of Open-Pit Trucks

From a point on the main underground haulage roads ramps will be driven to connect with the open-pit at about level 440'. These ramps can be driven up to 1596 and are designed to connect with dumping bins of reinforced concrete built into the walls of the open-pit above the main pit haulage road. These bins will be of 1000 ton capacity, and are controlled by suitable chute gates to permit loading into the regular open-pit haul trucks for transport via the open-pit haul road to the surface crushing installation.

- 30 -

5.3.6

10 ton capacity, such as the Wagner vehicle designed for this purpose.

The main wall haulages are driven at 20' wide by 9' high to permit passing of these vehicles. Front end loaders in this scheme will transport only to stope boundaries, before transferring their loads to these vehicles. stope drives on "475 tI leve 1 remain as before at 15' :x 9 I •

Loading of Open-Pit Trucks

From a point on the main underground haulage roads ramps will be driven to connect with the open-pit at about level 440'. These ramps can be driven up to 15% and are designed to connect with dumping bins of reinforced concrete built into the walls of the open-pit above the main pit haulage road. These bins will be of 1000 ton capacit~and are controlled by suitable chute gates to permit loading into the regular open-pit haul trucks for transport via the open-pit haul road to the surface crushing installation.

- 30 -

1-

1-

1-

1-

I--

-

1-

1-

I--

1. 'f

, I • I • I • I • I • I • I • I • I .. I " I .. I .. I .. I .. I .. I .. I .. I .. I ~ I •. I .. I .. I .. I • I •

---------v--

-- r--- ---.......,

I I I I I I I I I .. I I to I .. I I I ..

-N~

I I .. I

1-

1-

1-

1-

-

•

-

•

•

..

I~ I"" ... ".... I" """100' 1_ ... 4 1_ .... 1:''''

BECHTEL AND COMPANY " MOMTIIAL. '.Q.

ARGOR EXPLORATIONS LTD. 1-

PROJECT TERRANE ~--------------------~.

ULTIMATE PIT CASE I ....... ..

.. I .. I .. I

.... I

r

SHAFT BOTTOM

"—— W i*vi** Lttfl

• E V 11 ION t

KALE l" - 50' IHAWH s. B Y/

BECHTEl AND COMPANYMONTIIAl. P.O.

ARGOR EXPLORATIONS LTD. PROJECT TERRANE

TYPICAL TRANSVERSE STOPING ARRANGEMENT CASE E

;i .. r .tJ **...|i^ ji-ijuL ______ tgjWIJU MB______ II

6396

r • .. .. .. .. .. .. '+00. 4.00w 2+ooW O.L

1 I I

--1-

"

.. .•

,., .. to ..

... •.

.. .. .. ..

If tn-_ .. ::

BECHTEL AND COMPANY MOH"fAt, '.0.

ARGOR EXPLORATIONS L TO. PROJECT TERRANE

TYPICAL TRANSVERSE STOPING ARRANGEMENT CASE n ..... ,.. .. .." .

..

•

•

..

- l " l -

PLAN SUBLEVEL 225

Vft* Tli A r t ON

PLAN LEVEL 475

••VISIONS

BECHTEL AND COMPANYMONTIIAl, r.Q.

ARGOR EXPLORATIONS LTD. PROJECT TERRANE

PRELIMN. MINING PLANS ELEV. 225 a 475

6396

.. "

'HOOS 0.00

----- I!!'J MAIN SHAFT

"------

PLAN SUBLEVEL 225

I

•

,,/

II

..

o

• 4+00S 0+00 4""OON

PLAN LEVEL 475

..

..

/' /'

./

..

..

PYflfOC.HtORE.

.. F",,,ft ¥."",~rIllN SI"fFr

//--

..

J:urlltlt£ VE..-n.lAT,.N SHAFT

.. .. .. .. - .. -

-N~

.. .. at

- .. - -

arran: ......... :

au ... S. J)'llMA

BECHTEL AND COMPANY MONTIIAL. '.G.

ARGOR EXPLORATIONS L TO • PROJECT TERRANE

PRELlMIN. MINING PLANS ELEV. 225 a 475

••• w, •• Mo .IW

•

..

o

V

•? -J \

' ' -V /' \s*

l "its*

1-44\\

\ \\\\

'S

\ \ ,' S*/i\\. VI \X\ \

\ X

\s

i*odS~

r - loo1 T

BECHTEL AND COMPANYMONTIIAL. P.Q*

AR60R EXPLORATIONS LTD. PROJECT TERRANE

ULTIMATE PIT CASE lg

joi H* ____an**iH

6396

• • • - It .. • • .. .. .. ..

/ /

/

- 11_- 400

t. ..

.. ..

-N-h-

.. .. .. ..

It

•

..

It

o

..... " .1"1'10_' aM ... S • .1 I.A

BECHTEL AND COMPANY MONU,"L. '.0.

ARGOR EXPt.CRATIONS LTD. PROJECT TERRANE

~----------------------~ .

..

UlTIMATE PIT CASE m

.. .. I


SCHEDULE

—Interim Report

Dscision Open PitDrilling^

Est- i m a t g Final Reportm. Eng_^

to Proceed

OPEN .PIT__.Access, Road to Site Sunnort A Camp Facilitiesreek Div,ersjiQn ..4....J ...4.

Pi t D rainage Se Excavate PeatExcavate Ti

Silty SandSiltstone - SandstoneFinal Clean UpPit Rock Wo irk

UNDERGROUNDAccess Road to Site

Cam Facilities— SupportShaft SinkingStope Development:

Detailed EngineeringEquipment OrdersEquipment DeliveriesSite WorkConcrete

Building EnclosedEauioinent InstallationPre-operation Testing

•

f.

. "'. '7' • ~lt . .- "~w- -'T"


SCHEDULE

.... --- .. ..D.e.c.i.s.io..n... t-o. Proceed -----------+---.. -~t+0-d .. l+'--+-+·~· ...... -+!-+-+ ~ .~ J ;-~,_. 'T ~~ ). ~ I .~ ,L. ~ .. ~. -l- ~L- ~-+- .+- -f- -J-+-4---+-+---+-t--~f,tj··I-f-~ j u ,·t· ; ! ! ' • 1 . i I I, 1 1 I I .' j

---- .- ----- .-.-- ---.------- - ._~i ~r tt 'r t-+i ~-'"1' 7 7- ,--+.+r---- .+-. >-f-- -1-':~" -t-. "t--K= -'j ;- r- -+-+---........ -..+--I----+- -" - .' f.-f--. ---~~~.?l~----------------------+--+R+~H~~4+4·~ ~. ~--0-- ~ +- 4 ~ +.~- '.-- -t-- +-1-1-. -~~4~~~~~~- -

A R d t S 't '~T i • i I • t, ! ' , I 't· "t ---- '- S~~!~-~t ~a C~~ F ~r ~ 1 it i e" f- :.. -+- --t-o :~ f-, '1 .~. --t . '1 t -.. T~- --~ - . r-r --:--'--+---+---~

f--___ >--. ___ Cr:.~~lL.Qiversion ~ _f. - f---.- .+ __ ·,-t-t- -+ 1-+.I.-~.t~:t-L.r---I-_ .. -t .. W-.. 1-. ___ • __ .~J_t. . .J)rainaqe & Excavate PeatL ;".I-.t=l-H-=- f--! ' -t-: t --'- 1. •••.• : +- . L -. L -i -' ·t'·L +, ,.1. .....--f-4--+---~_+_+___+_+~1_ _

1-- __ ,_._.J;;..~cay.9te _'tL-,,___ R -1=' .-~.--.-- III i • i· "'1-' -;- -+- f- ._, ~ 4-+-' ..... -t--' . f-k,'--: '11 . ~fT. -'~,., ~ ..:~.~ :r,', , L -~!., t'-·.T+1i.·. -'1' +j-.. '.'J.-.. ,J .. - ........ 1 ........... .it. ; 1 ; • I I' Silt" Sand -+. t, ..-r: ·-i t ;':! H- t! 'i....L...I ' I j I j

-,,- - .... ,--------5 ii"~·tone - Sandstone ,----- a- :Li- - ~r I- ;---- ·"r- i .--, 1 ·t- ,'.. -- t--~t-- .. _. fir + r-r--- f - .+'+: t--+ -- . -4---+---I--+-+-+ -- -1-'

Fl' nal clean UT~ ~-t= ~ •. :-+:. ~. ·r.:: I ! -.J. ! : 1-' 1 t: ! I .... . 0 -_. >--"----------'''--''''~'''-''=--.: )~ ~~i-- .. ::t '-r---r- I f ;--+ .. i, .. ti ·_i1_"-' '± __ .. ·i+, .. 11 I,. J. ".J!!= __ ~~f-+,-". +,-+'+._-ll'-'_.+·+--........... ---- f--------- Pi t- P"rlr TAl,... .... lr ~~t-I=1----. ~T . _ . I ~r ..... . . c-,--~~- +-

1---- t-- ....... ··r 1-+ t ·.t- - r-+-'-+--·..j.....4-+--I--~ ... - ~t- -·+-+-+--HH-+-+-t .... - '1+ 1 '- .,. ...+-. t +- -...... l----'I~ .... '-1--1-----11-4---4--+--1- ......

r TTNDE BGROUND ~.... . f-- - +_+ .. !_ H' - ·_·t'· +--~f--+-4--I--4---+-+-t---1---I-+-+-+--+--+-+--+-+-+- +-~ --1----- - t- .. T'''' I-l-+-+---f----f- 1_ ...... 1--'·-+---......I.l.l..ll.l..t.JAl.Ulc~cU.e:IUsJCs""-"'U.Rl..o-a-d-t-o--s-i-t-e---------------1------+1=-+--+.: 1----- .... ~ ... r='j.. lti;

• _. e·

.1- '-PT .ANT f-' - r--- t- f.-

nPTr'I; led Enaineerinq '-~ -...:.r=-= ' __ _ .. .. .. I''''

~.- ___ ._~E:,a~lu~i£ITolm~eunDt~OQKr~d~eUr~sL-~ _________________ ~~_~-=·~:~·-=t-4-=F~Ft~+-~~~4-+~-+~~,~-~-+~~+-~~~ ...... ~-t~~~ . ~~H ~. Egu ipmen t De liver ies f--r= -, - ...::. -:-...... 111ttttDtjjj1IIItutjjj1 •••• ~tjj· ·j-=·ttJtjjj1!tttt1:tjj -'''''

~. Site ...!WC!.:o~r . .!:!k~ ___________________ + ___ +:::-f ...... --+-f'-t-----tf----+---t-_-t--I-+++ .... i- - ~t.-~L~--:.i-+-++++-+-t-~ ... -+-+++ Concrete ::: ,.- '-~-'..:.. .... :::.- i

-' .-- ,,_. _. Steel & Siding ---------------+----......... f-+- = :+~.::::.=:::.:._I__-~__'-.... _1- -->-

~._.-~-- __ Bu j 1 ding Encl~_d._·r ~'~- :':"f--:-':"r-+-~ . --'- - ,Eo1d.i.pmentJlJstallation: .t'i--+'~- 'r- ---I--

1-_-'-___ P..r..e.=..Op.e.r.rit i on 1'e"tina ----f-+---+-+t---t--+.-+-'+ ... -+--++--.+--~-.. _. ___ ,,_ ~....... f- ~-= !.

.... - -- 1- ....

f-.

r--

-~-

.J. ++~+--hI-lH-+-+--+-+-+-+-"f--HH·--+-+-- f-. "c--

4-+--I--l----'-4---I--- - r--' f-- . -- f--.

- A L- . - "- .,,+--~~~-+-+ >--,- e-.. • f- " ,-I-.

-+--+-It-+-t---t-- f..-. 1-1- t~.~"~--4--i-I-----I---+--- t--"- - --i

---f-- f--

>-- T + f-

I-

I i

!

CANADIAN BECHTEL LIMITED 564O PARE STREET MONTREAL 9. P.d,

TELEPHONE ' 342-O1SO

MAILING ADDRESS: P.CX BOX 23-4O ST. LAURENT P.O.

July 21, 1967

Mr. R. M. SmithConsolidated Morrison Explorations Ltd.Suite 170011 King Street WestToronto, Ontario

Dear Mr. Smith:

We are pleased to present this interim report covering our very preliminary assessment of your potential columbium property near Moosonee, Ontario.

It should be stressed that the data presented in this report are of a very preliminary nature and are not intended to be used for appropriation of the capital required to develop the property into production.

The purpose of this interim report is to assist your company in making the decisions necessary to continue with the feasibility study and to proceed with further exploration of the property. To aid you in your decisions, we have presented you with three possible mining schemes together with order of magnitude capital, operating costs and projected earning potentials, assuming a minimum of 20 years reserves.

No consideration has been given to the effect on the cash flows should the recommendations of the Carter Report be implemented, or to savings in capital requirements which might be made in the following areas :

a) Government subsidy in building the access road and operation of the ferry at Moosonee.

b) Elimination of the airfield.

c) Reduced pit stripping quantities when additional soils data becomes available.

, -CANADIAN BECHTEL LIMITED II}

Mr. R. M. Smith Consolidated Morrison Explorations Ltd. Suite 1700 11 King Street West Toronto, Ontario

Dear Mr. Smith:

5640 PARE STREET

MONTREAL 9. P.Q. TBLBPHONE • 342-0160

MArLiNO ADD"'IISS. p.o. BOX 2340 ST. LAU"'BNT P.Q.

July 27, 1967

We are pleased to present this interim report covering our very preliminary assessment of your potential columbium property near Moosonee, Ontario.

It should be stressed that the data presented in this report are of a very preliminary nature and are not intended to be used for appropr iation of the capital required to develop the property into production.

The purpose of this interim report is to assist your company in making the decisions necessary to continue with the feasibility study and to proceed with further exploration of the property. To aid you in your decisions, we have presented you with three possible mining schemes together with order of magnitude capital, operating costs and projected earning potentials, assuming a minimum of 20 year s reserves.

No consideration has been given to the effect on the cash flows should the recommendations of the Carter Report be implemented. or to savings in capital requirements which might be made in the following areas:

a) Government subsidy in building the access road and operation of the ferry at Moosonee.

b) Elimination of the airfield.

c) Reduced pit str ipping quantities when additional soils data becomes available.

•


Mr. R. M. Smith July Z7, 1967 Page 2.

d) Takeover of the contractor's equipment for continuation of the stripping programme.

The preliminary economic assessments given in this report indicate that the discounted rate of return on investment for each of the three schemes, although low, are sufficient to warrant further expend itures to more fully define the total project, to allow a detailed capital cost estimate to be made, and to allow the preparation of more definitive cash flow calculations. Of prime importance is the work required to obtain a bulk sample of the ore for metallurgical testing prior to final plant design. This testing will provide concen trates which will be of value in market studies and will assist in establishing firm sales committments for the product. A further very important benefit from the bulk sample programme is that in sinking an exploration shaft through the overburden, valuable additional information of the overburden strata will be obtained.

A schedule of work, together with key decision points, are presented in the appendix of this report. This schedule will allow additional work to progress in an orderly manner with a minimum of investment at each stage. This work would be divided into three phases, as follows :

Phase I

a) Initiate shaft sinking and underground drifting to produce a bulk sample.

b) Examine, log and record all soil informationdeveloped in sinking the shaft through the overburden.

c) Evaluate the rock encountered during drifting through a preliminary rock mechanics programme.

d) Prepare final comparison of open-pit versus under ground mining methods. Decision required by December 1967.

CANADIAN BECHrEL LIMITED

Mr. R. M. Smith July 27. 1967 Page 2.

d) Takeover of the contractor1s equipment for continuation of the str ipping programme.

The preliminary economic assessments given in this report indicate that the discounted rate of return on investment for each of the three schemes. although low. are sufficient to warrant further expenditures to more fully define the total project. to allow a detailed capital cost estimate to be made. and to allow the preparation of more definitive cash flow calculations. Of prime importance is the work required to obtain a bulk sample of the ore for metallurgical testing prior to final plant design. This testing will provide concentrates which will be of value in market studies and will assist in establishing firm sales committments for the product. A further very important benefit from the bulk sample programme is that in s inking an exploration shaft through the overburden. valuable additional information of the overburden strata will be obtained.

A schedule of work. together with key decision points. are presented in the appendix of this report. This schedule w ill allow additional wqrk to progress in an orderly manner with a minimum of investment at each stage. This work would be divided into three phases. as follows:

Phase I

a) Initiate shaft sinking and underground drifting to produce a bulk sample.

b) Examine. log and record a11 soil information developed in sinking the shaft through the overburden.

c) Evaluate the rock encountered dur ing drifting through a preliminary rock mechanics programme.

d) Prepare final comparison of open-pit versus underground mining methods. Decision required by December 1967.


Mr, R, M. Smith July 27, 1967 Page 3.

Phase II

a)

b)

c)

d)

The bulk sample will be tested by gravity and flotation methods to produce firm criteria for plant design.

If a decision is reached to mine by underground methods, deep drilling will be required to prove up ore reserves between the 525 and 825 horizons. This should be initiated by February l, 1968.

A preliminary design should be made of the plant and a comprehensive report assembled giving full definition of the total project, with a detailed capital cost estimate.

Cash flow calculations prepared at this stage will form the basis of the decision to proceed with the project. Such decision should be made by August 1968, so that some long delivery equipment can be ordered, and concrete and steel design can be initiated.

Phase III

a)

b)

c)

A contract should be let by September 1968 to prepare the access road, so that the mine can be fully developed by February 1971.

Plant construction should commence early 1969 so as to complete all civil and structural work before the 1969/ 1970 winter. The buildings should be enclosed by November 1969.

Mechanical and electrical installations will be carried out during 1970.


Mr. R. M. Smith July 27, 1967 Page 3.

Phase II

a) The bulk sample w ill be tested by gravity and flotation methods to produce firm criter ia for plant design.

b) If a decision is reached to mine by underground methods, deep drilling will be required to prove up ore reserves between the 525 and 825 horizons. This should be initiated by February 1, 1968.

c) A preliminary design should be made of the plant and a comprehensive report assembled giving full definition of the total project. with a detailed capital cost estimate.

d) Cash flow calculations pr epared at this stage will form the basis of the decision to proceed with the project. Such decision should be made by August 1968, so that some long delivery equipment can be ordered, and concrete and steel design can be initiated.

Phase III

a) A contract should be let by September 1968 to prepare the access road. so that the mine can be fully developed by February 1971.

b) Plant construction should commence early 1969 so as to complete all civil and structural work before the 1969/ 1970 winter. The buildings should be enclosed by November 1969.

c) Mechanical and electrical installations will be carried out during 1970.



During our meeting with you on July 17, we presented verbally the contents of this interim report, and reached agreement on the principle concerning the approach to be followed over the next several months. We are in the process of obtaining bids for the shaft and cross-cut work necessary to obtain the bulk sample, and at the same time, we are assembling an estimate of the costs involved in the additional work required to complete the feasibility study.

The effect of these events on our present contract with you is, of course, to extend the schedule so that the feasibility report will be completed approximately nine to ten months from now. Some add itional costs will probably be incurred on our part which will increase our contract limit. We will assess the probable extent of these costs.

It is anticipated that during the period of shaft sinking, cross-cutting, diamond drilling and bulk sample testing, a minimum of activity will be required on our part. We shall continue to use Mr. K. Culver as Project Engineer on a part time basis, so that he may co-ordinate the activities required during this period and arrange for the neces sary application of technical skills for soils evaluation, mining assessments, metallurgical work and other services as required.

Yours very truly


D. RhodesManager of Engineering

DR:cjm



Dur ing our meeting with you on July 17, we presented verbally the contents of this interim report, and reached agreement on the principle concerning the approach to be followed over the next several months. We are in the process of obtaining bids for the shaft and cross-cut work necessary to obtain the bulk sample, and at the same time, we are assembling an estimate of the costs involved in the additional work required to complete the feasibility study.

The effect of these events on our present contract with you is, of course, to f'xtend the schedule so that the feasibility report will be completed approximately nine to ten months from now. Someadditiona1 costs will probably be incurred on our part which will increase our contract limit. We will assess the probable extent of these costs.

It is anticipated that during the period of shaft sinking. cross-cutting, diamond drilling and bulk sample testing, a minimum of activity will be required on our part. We shall continue to use Mr. K. Culver as Project Engineer on a part time basis. so that he may co-ordinate the activities required during this period and arrange for the necessary application of technical skills for soils evaluation, mining assessments, metallurgical work and other services as required.

DR:cjm

Your s very truly


J)~~ D. Rhodes '/ Manager of Engineering

PRELIMINARY MINE FEASABILITY RPT...Limited pyrochlore-columbite deposit located in the James Bay...

Documents

Transcript of PRELIMINARY MINE FEASABILITY RPT...Limited pyrochlore-columbite deposit located in the James Bay...