Mining Dent* COAL MINING COSTS % Qraw-MlBook (S 1m PUBLISHERS OF BOOKS FOB^ Coal Age v Electric Railway Journal Electrical World v Engineering News -Record American Machinist v Ingenieria Internacional Engineering 8 Mining Journal ^ Power Chemical 6 Metallurgical Engineering Electrical Merchandising COAL MINING COSTS BY A. T. SHURICK ? MEMBER AMERICAN INSTITUTE OF MINING ENGINEERS, FORMERLY ASSOCIATE EDITOR "COAL AGE," ETC. FIRST EDITION McGRAW-HILL BOOK COMPANY, INC. NEW YORK: 370 SEVENTH AVENUE LONDON: 6 & 8 BOUVERIE ST., E. C. 4 1922 c *V , r- T ? ;> ; .-^ >r . COPYRIGHT, 1922, BY A. T. SHURICK PREFACE THERE are books on costs in all the important branches of engineering except coal mining. The author has waited patiently the advent of a similar work in his chosen field and none having been forthcoming he has made bold to venture the effort himself. Obviously a work of this character cannot be up-to-date as to costs in dollars and cents because of the wide fluctuations in the purchasing power of the dollar in labor, equipment and material, particularly during the last five years. There has been no hesitancy, therefore, in using data of a number of years back so long as the subject discussed is still in general use as in the case of the comparative costs of wood and masonry brat- tices, etc. The reader will have no difficulty in interpolating the figures given to conform to current standards and to facili- tate this a table giving all the wage scales in the Central Com- petitive District since 1898 has been given on page 158. Thus if a certain piece of work required the services of three men for a certain number of days, say a decade ago, it is a relatively simple matter to estimate the cost in terms of prevailing wage scales. Care has been exercised to give the year during which the different examples cited occurred in order that this inter- polation can readily be effected. A great deal of valuable data of an abstract nature has been obtained from the various State and Federal government reports but in general it has been the endeavor to hold more to specific costs. Thus the cost of haulage or the cost of doing a piece of work under certain well defined conditions has been accepted as of more value than the average cost of mining for a certain district or the capital investment per ton of capacity, etc. In other words it has been the aim to make the work essentially practical. Only a few of the best systems of mining have been discussed but these, it is believed, have been covered in greater detail than in works dealing with this subject alone. The reason for this 48265,; vi PREFACE is that unless full particulars concerning all phases of working under any system are made clear beyond all peradventure the cost figures are worthless. To insure a thorough treatment of all subjects taken up it was deemed advisable to limit the present volume to under- ground costs alone. A great deal of valuable data on outside costs has been assembled in the course of the present work and it is thought the publication of this in a separate volume at a later date will enhance the value of the completed work more than if an attempt were made to straddle the two fields in the present volume. It is thought, if anything, the book errs on the side of con- servatism ; old and tried methods only have been accepted, newer systems and appliances, some of which are very promising at this time, having been used with caution. As an example the use of the underground loading machine or the combined mining and loading machine has not become sufficiently general as yet to justify the amount of space devoted to say, the mine motor which is now an accepted part of nearly every mining operation. In concluding, the author wishes to make sincere acknowledg- ment to the many friends who have furnished him with much valuable material and still more valuable suggestions and advice ; to the various engineering societies, the papers of which have been quoted so freely throughout the book ; to the large engineer- ing companies, several of which in particular have been untiring in their efforts to furnish certain special requests for material; and to the technical press which has been drawn upon liberally. A. T. SHURICK. DEARBORN, MICH. January, 1922. CONTENTS SECTION I MINING COSTS PAGE Government statistical data regarding costs Method of computing tax returns Comparison of costs and distribution of revenue for German and Middle Western Mines Conditions where operations may be conducted at an apparent loss Systems of mining Methods of working in the Pocahontas field Connellsville systems Compara- tive costs of mining different thicknesses of coal Conveyor system Mining machinery Cutting machines Installation and operating costs Comparative cost for alternating and direct current for machines Arc wall cutters Post punchers Repair costs Machine bits Loading machines Mining and loading machines Blasting Dynamite Hydraulic cartridges Shot firing Daymen Miner's wages Losses from idle time Economic aspects of conservation Use of longwall to effect conservation 1 SECTION II SHAFT SINKING Operations Circular or rectangular Equipment Sinking costs Shaft linings Rates of progress Reports and contract forms 176 SECTION III HAULAGE COSTS Tractive effort, drawbar pull and rating of mine motors Number and size of motors required Gathering locomotives Costs Motor losses Power costs Line costs and losses Bonding Computing losses in bonds Track costs Haulage grading estimates Haulage track curves Rails Track frogs Track Mine cars Rope haulage Gravity planes Endless rope haulage Wire rope lubrication Com- parative costs of different systems of haulage Comparison of all systems Animal, compressed air and electric haulage costs Single- and two-stage air motors Compressed-air and animal haul- age costs Electric motor and animal haulage costs Costs and care of mules Gasoline motor vs. animal haulage Storage battery and trolley motors haulage costs Compressed-air and electric haulage costs 219 vii viii CONTENTS PAGB SECTION IV TIMBERING COSTS Ho to buy timber Timber used and costs for the United States Com- puting size of timber Timber framing equipment Timber pre- 61 vatives Steel timbering Concrete timber Cement gun Re- aiming timbers 365 SECTION V MISCELLANEOUS INSIDE COSTS nneling costs American and foreign tunneling records compared Cost of rock tunnel at a coal operation Some examples of tunnel costs Drill steel Explosives Ventilating costs Power required Speci- fication of fans Change in volume of air required Mine lighting Portable electric lamps Oil and acetylene lamps Oil lamps Cost of undergrounds tables Overcasts Stoppings and overcasts Com- parison of doors and overcasts Cost of stone and wood brattices Refuge chambers Mine sprinkling costs 406 GOAL MINING COSTS SECTION I MINING COSTS The U. S. Census report for the year of 1909 shows that the value of the Pennsylvania anthracite produced that year was $148,957,894. The total gross expenses amounted to $139,110,- 444, from which should be deducted $4,864,844 made from charges to miners for explosives, oil and blacksmithing, making the net expenses $134,245,600. The gross expenses are itemized as follows : Services: Salaries $ 4,572,489 Wages 92,169,906 $ 96,742,395 Supplies : Fuel and power 3,189,279 Other supplies 23,472,809 26,662,088 Royalties 7,969,785 Miscellaneous 7,736,176 Total gross expenses $139,110,444 Deductions 4,864,844 Net expenses $134,245,600 The total production in 1909 amounted to 72,215,273 long tons, so that the average value per ton for the output in that year was $2.06 ; the average cost per ton was $1.86 ; and the net returns on the operations for the year were $14,712,294, or an average of 20c per ton. This at first glance looks like a fair return, but attention must be called to the fact that the Census 2 COAL MINING COSTS figures of cost make no allowance for interest on capital invested or borrowed, and no offsetting charges for amortization or depre- ciation. According to the returns to the Bureau of the Census, the entire capital invested in anthracite mining in 1909 was $246,- 700,000, which may appear rather inadequate when one considers the magnitude of the industry, and an annual production of $150,000,000 (in 1911 the output was valued at $175,189,392 and in 1912 it was $177,622,626), but these are the figures reported by the Census Bureau. If on this capitalization an allowance of 4 per cent be made for interest, the net returns for the year amounted in round numbers to $4,844,000. If new breakers and other equipment are charged into operat- ing expenses no allowance need be made for depreciation, but the exhaustion of from 75,000,000 to 80,000,000 tons from the reserves every year should have some amortization charged against it and if 5c. a ton be allowed the margin of $4,800,000 is practically wiped out. The figures covering the cost and value of bituminous coal show even more striking comparisons. There are some slight differences in the statistics of production between the Census figures and those published by the United States Geological Survey for the reason that the Census investigations excluded mines having a production of less than 1000 tons, whereas the Survey includes every small country bank from which it can secure a report. For 1909 the Survey showed a bituminous coal production of 379,744,257 short tons valued at $405,486,777, and the Census Bureau showed a production of 376,865,510 tons valued at $401,577,477, the difference being about 3,000,000 tons in quantity and $4,000,000 in value less than 1 per cent in either case. As the Census figures for cost of mining are the basis of this discussion, the Census figure of production is also used. The total value of the bituminous production, as already stated, was $401,577,477, and the mining expense of producing this value, including salaries of officers, was $378,159,282. As in the case of anthracite, the expenses of production do not include any charges for depreciation, amortization, or interest on capital invested or borrowed. The expenses are divided as follows : MINING COSTS Salaries $ 20,417,392 Wages 282,378,886 Supplies 45,345,932 Royalties . / 12,035,900 Miscellaneous 17,961,172 Total $378,159,282 From this it appears that 75 per cent of the total cost and 70 per cent of the total value was spent in wages. Salaried officials got less than 5.5 per cent. The total capital invested in the bituminous coal mines of the United States in 1909 was, according to the Census bulletin, in round numbers $960,000,000 ($960,289,465), and this does not appear as if there were very much over- valuation, whatever the capitalization may be as represented by stock issue. The dif- ference between the value of the product and the expense of producing it was $23,440,000 in round numbers or a fraction over 2.5 per cent on the capital. According to the figures compiled by the Bureau of the Census, the amount paid in wages was, in 1909, above 80 per cent of the total selling value of coal at the mine mouth. From 1909 to 1913 there were two wage increases granted one in 1910 of 5.55 per cent and another in 1912 of 5.26 per cent. These increases brought the wage cost per ton of coal produced to 92.44c. in 1913. In 1913 the average selling price of coal at the mines in Illinois was $1.14 and in Indiana $1.11 per ton. This leaves a margin of only 21. 6c. in Illinois and 18. 6c. in Indiana. Out of this must be paid the cost of material used at the mines; the cost of making sales; all officers' salaries; general expenses; insurance (liability, fire, storm, etc.) ; taxes (including tax on plant and mineral rights) ; interest on the investment; deprecia- tion of plant; royalties or charges for the exhaustion of coal. The report of the Bureau of the Census for 1909 showed that, without allowing for any interest charge on the investment or for amortization of property, the so-called net returns in Illinois and Indiana were only 3c. per ton in Illinois and less than Ic. per ton in Indiana. The average royalty paid, however, in these two states on coal recovered under lease is 5c. per ton, and the average present valuation of coal land is such as to require a minimum amor- COAL MINING COSTS tization charge of 3c. per ton to recover such, land value within the period of the mine's life. It will therefore be seen that in even so good a year as 1913 an actual profit return was impos- sible. As existing facts show, the industry sustained a sub- stantial deficit in these two states. The average value per ton of all the bituminous coal produced in the United States was $1.07 and the costs averaged a fraction of a cent over $1, so that the margin of profit to cover interest, depreciation and amortization was a little less than 7c. a ton. In some states the expenses exceeded the returns. Take Arkansas, for instance, where the expenses totaled $3,630,526 and the value of the product was $3,508,590. Other instances were : Value of Product Expenses Iowa $12,682,106 $12,816,076 Kentucky 9,940,485 10,127,987 Tennessee 6 548 515 6 691 482 Oklahoma 6,185,078 6,536,441 Virginia 4,336,185 4 392,440 Pennsylvania, by long odds the most important producer, with an output of 137,300,000 tons, showed a total of expenses of $117,440,000 and of value of $129,550,000 a balance on the profit side of a little over $12,000,000, or about 3 l / 3 per cent on the capital invested, $358,600,000. The four competitive states, West Virginia, Illinois, Ohio and Indiana, which rank second, third, four and fifth, respectively, in producing importance, all show such narrow margins between income and outlay that profits are infinitesimal. The figures follow: Value of Product Expenses Difference West Virginia Illinois .... $ 44,344,067 53,030,545 $ 43,024,716 51,697,504 $1,319,351 1,333,041 Ohio 27 353 663 27 153 497 200, 166 Indiana 15,018,123 14,906,831 111,292 $139,746,398 $136,782,548 $2,963,850 og o o ^ PH S S Value Including Minor Products I ai o PH I* I MINING COSTS 1 o co~ IO i-H Oi OOOOO *O CO O T^ CO OI>O CO i 1 O5 CO 00 O "f i-H CO ^^ GO iO ^f ' i T^ O i i O O (N COi i o o oo o cO ^^ O^ ^^ i * ill ^Si> 8 ll 3 Jill o O> O iH O !> CO 00 t>. O~ .coooc^>o OS CO C^l CO CO CO OO CO CO O 1 CO OS CO i-H CM CO 1C CO t-H O 00 OO O IO 1> O CO t^ rH CO .' ft ft ; e a fl - -2 -4-3 03 ^3 -S 5fc fe.Sfe S 6 COAL MINING COSTS These four states with an aggregate production of a little more than the bituminous output of Pennsylvania, showed a total of less than $3,000,000 as the excess of receipts over expenses. The capital invested in the coal-mining industry in these states was something over $310,000,000, so that the returns on the capital were less than 1 per cent. The United States Fuel Administration, organized during the war emergency, was empowered to exact the most intimate details as to operating costs and profits in the coal industry under severe penalities for omissions or incorrect returns. The En- gineers Committee of the Fuel Administration, headed by some of our most prominent engineers and equipped with an excellent organization for assembling and correlating this mass of material compiled a report on general production costs unexcelled in the history of the industry for its authenticity and accuracy. The accompanying table is a summary showing reported and adjusted costs, prices fixed and tonnage for all the principal districts to August 12, 1918. The diagram Fig. 1 shows these, in general, before the labor increase of November, 1917, com- pensated for by the 45c. general advance in coal prices, giving the average costs, "bulk lines," and prices fixed for practically all districts in the country as of August and September, 1917, and covers about 95 per cent of the total output of bituminous coal for the period stated. The costs for each district in the proportion of its output to the total tonnage studied are shown in heavy lines, the * ' bulk lines" are shown by medium lines, and the prices fixed are indicated by light lines. The diagram also shows the weighted average costs, "bulk lines," and prices fixed for the tonnage included, and effectively disposes of the widely circulated asper- sions of profiteering, of which the industry was so freely accused by people having no knowledge of the facts or willfully misrep- resenting them. Diagram Fig. 2 shows the same data for the principal dis- tricts for the full year 1918, giving, however, only reported costs and prices fixed, the prices having been fixed on the August- September, 1917, data, and changed only by the 45c. allowed November 1, 1917, to compensate for the labor increase of that date, reduced May 24, 1918, to 35c. in consideration of equal car distribution ordered at that time. MINING COSTS The weighted average margin between costs and prices for practically the entire bituminous coal production of the country was but 45. 6c., and between the "bulk line," which represents [{Hlp^^ M 1 i_ i :::ts I : ^T :;i;|i=; ; ;j;ii;;: : ' 3 '. j; the higher-cost necessary coal and the price fixed by the Fuel Administrator, was but 26c. As it is known that the capital invested per ton of yearly output in bituminous mines ranges from $2 to nearly $8, interest on which is included with other mines in the "margin" not included in the charted costs, it is evident that taking the COAL MINING COSTS Illllllll rH i-l "~ ^ CO CO *O *O OQO O O CO CO CO TF O5 O^ ^~* i*H ^^ t^ C^ ^^ ^^ ^"^ C^-l CO T~H c^J 05 O O ; 8888 OS TH o cO rH~ 10" TjT OtOiOOCOOr^tX)(NOiOiOO^OOCO t-t Ui t-t ^o^oj^o^aj^di^aiaaJso^cD^o^om' 3^ Illli : 1 I i-ll : t 1 i i^^ III ?^ fe r*> M C C 02 CO || 1 8 ^ A % $ I @ W .3 '3 '& 2 i> d r& ^ I S cf H -a j? t> IISS o o o o w CO t^ 00 OS O TH rH O 00 CO CO CO 10 COAL MINING COSTS S o -T3 I -I I oo t^ d d iO O ^ CQ O CO 99 09 1 s * !> ^2 ^^ I s * !> d o TH TH d d 1 8-8 ISs 88 1> CO CO CO O O O SB O O iO iO >O iO >O CO TH rH Cq (^ CO CO (N (N (N f^s iO IO IO CO rH r4 IvI (N (M (N i I> (N (M - t^cOb-i0^1>t^cO0 CO CO CO CO 1>- 1^ TH TH O O ^ Th< CQ (N (M (M (N (N 8888 o o o o O iO O O O >O co oo o 10 co co CO OO OO OS OO O t^ 00 TH (M OS OS I I ^ I I I I I I J-4 * Cd -4-3 C] -4^ Cj -4-3 O -4^) Cj -4^ GJ -4-3 3 -4-3 C3 ^S^S^m^S^S^S^^^S rlllllllillllllf} CO 1> CO CO CO cO ^ TjH 00 00 O5 O5 I ; I O 'S* PI ^ (N * co co T* ^ (X) 00 00 00 (N (N rH r-l rH ^H ! rH 3 ! C 00 00 Oi O5 O5 O5 i i ; rH III "^ S "03 S "&2 S O S> a S>"& > a 5^^ J^ 00 O2 r 1 O O5 bC rH .2 1 S Q c O l> t^ 00 MINING COSTS 13 industry as a whole no excessive price allowances were given. If prices had been fixed at a point high enough to even cover the highest costs reported in each district, the result would have been to add over a billion dollars to the price paid for coal, with probable labor disturbances, in an effort to obtain some of the abnormal profits which would have gone to the great majority of the tonnage, so serious as to have probably decreased rather than increased the tonnage, which was in fact ample for all the needs of the country. 14 COAL MINING COSTS The prices fixed from this complete investigation of costs have shown in many cases a remarkable compliance with eco- nomic laws. For instance, in Illinois the cost of coal from the different price districts delivered in Chicago was found to be practically identical, showing that the mining of the higher- cost coal is due to its proximity to the principal market and the lower resulting transportation costs. High-grade coal shipped by lake and rail to Minneapolis was found to cost precisely the same per heat unit as a lower-grade coal shipped a much less distance all rail. Anthracite prices as fixed by the President Aug. 23, 1917, with an adjustment for the labor increase of Dec. 1, 1917, were the subject of an intensive study by the committee im- mediately after the first charting of bituminous costs was com- pleted. A technical paper giving the methods adopted and the results of this analysis was presented before the American Institute of Mining Engineers, February, 1919, and the following is abstracted from this paper : The adjustments of cost from a reported to a price-fixing basis, as described for the bituminous methods, were applied but showed only minor adjustments as necessary. The great spread in anthracite prices on the varying sizes, which for the 6 -month period under review ranged in average from $5.244 for nut to $2.074 for barley coal, makes the question of the percentage of sizes produced at the different collieries a vital one. The realization with the same prices for each size must be within very wide limits, when it is considered that the percentage of prepared coal reported from different collieries varied from over 80 per cent to below 30 per cent for fresh-mined coal. Hence, as the spread in prices for the various sizes must be predi- cated on some percentage, it is essential to find some method of adjustment to allow for this variation. The logical method of adjustment is to calculate actual costs to costs as of the standard percentage of sizes, so that the margin between the adjusted costs and the average realization shall be the actual margin for each colliery between its actual costs and actual realization due to its particular percentage of sizes. As a base for realization the actual percentage of sizes for fresh-mined coal for the 6 month period was adopted. This percentage is given in the following table. MINING COSTS 15 PERCENTAGE OF SIZE OF FRESH-MINED COAL Size of Coal MESH, IN INCHES PERCENTAGE OF SIZES Through Over Fresh- mined Washery Fresh- mined and Washery Round Round Broken 4| 3f 31 2^ 21 if 14 f 4 A 1 & A A 1 A A 3| 31 2A 21 if 14 f 1 A 1 A A A A A A 6.8 14.6 19.6 24.7 9.1 11.6 3.2 4.9 3.9 1.6 0.4 1.2 2.3 10.1 10.0 21.4 14.9 27.5 8.8 3.4 6.2 13.5 18.2 23.5 9.2 12.4 4.2 6.8 4.3 1.7 Eee Stove Nut Pea Buckwheat Rice Barley Boiler Screenings For adjustment as a base for fixing a spread of prices the percentages used were, taken at even figures, prepared 65 per cent, pea 9 per cent, buckwheat 12 per cent, and smaller 14 per cent. The adjustment finally arrived at after long study was tested on actual reports from collieries having percentages that varied from over 80 per cent to under 30 per cent prepared coal and was found to be correct within a maximum variation of less than 1% per cent. It was as follows: For Each 1 Per Cent Variation Above Standard, Per Cent Deduction Below Standard, Per Cent Addition Prepared 1.20 1.20 Pea . 0.85 0.85 Buckwheat 75 0.75 Smaller 0.50 0.50 As examples of the working of this adjustment with prices assumed at about the average for the 6 months and taking mines well away from average percentage of sizes, the following may be cited: 16 COAL MINING COSTS Size Base Sizes, Per Cent Base Price Reali- zation Mine A Sizes, Per Cent Correc- tion, Per Cent Actual Reali- zation Mine B Sizes, Per Cent Correc- tion, Per Cent Reali- zation Prepared 65 9 12 14 $5.10 3.70 3.20 2.20 $3.315 0.333 0.384 0.308 73.1 6.4 10.4 10.1 -9.72 + 2.21 + 1.20 + 1.95 $3.730 0.237 0.333 0.222 55.1 15.3 13.7 15.9 + 11.880 - 5.355 - 1.275 - 0.950 $2.810 0.566 0.438 0.350 Pea Buckwheat Smaller Total 100 $4.340 100.0 -4.36 $4.522 100.0 + 4.30 $4.164 Assume cost for each mine $4 . 000 Actual margin . 522 $4.000 0.164 Standard realization $4 . 340 Calculated cost as of standard per cent sizes, $4X0.9564%= 3.826 $4X104.30% Calculated margin $0. 514 $4.340 4.172 $0.168 The correction for mine A is then 4.36 per cent and the adjusted cost $3.826, showing 51.4c. margin on the $4.34 standard realization against 52. 2c. actual margin. Similarly, for mine B, the correction is -{-4.30 per cent, giving an adjusted cost of $4.172 and a margin of 16. 8c. as compared with the actual margin of 16. 4c. Thus the adjusted costs on the chart bear a true relation to the realization received from a scale of prices for the various sizes based on the standard or average percentage of sizes adopted as a base, regardless of the actual percentage of sizes produced by each operation, and prices can be fixed from the chart line of adjusted costs which will result in giving each mine its intended margin. The correction, of course, is an allocation based on realization from the different sizes and could be made more accurately by taking into account each size produced, but at the cost of more time than was available for the work. With a material variation in price, different factors of correction should be calculated. A large percentage of the anthracite coal is owned in fee by operators, who also lease tracts contiguous to their fee holdings. As all report royal- ties on the basis of tonnage produced, the general average 15. 5c. per ton reported is misleading. The actual average royalty reported by operators mining generally from leased lands was 33.25c. and by those generally mining from fee lands, 5.5c. As relatively few operators mine exclusively from either class of lands, no data are available to show the actual average royalties paid, but it is believed that the present average would be approxi- mately 40c. per ton. A few leases, notably those made by the trustees of the Girard estate, owned by the city of Philadelphia, base the royalty payments on a per- centage of the sale price of the coal at the mines instead of requiring fixed royalties. This percentage varies from 15 per cent to as high as 28 per cent of the price. As the labor war bonuses materially add to the sale MINING COSTS 17 price, these have resulted in excessive royalties and serious embarrassment to the operators, who were not allowed to increase the price of coal suffi- ciently to even fully absorb this additional labor cost and by whom the extra royalties must be paid out of already narrow margins. Cost charts were made from averages of the 6 months, showing both the reported and the adjusted costs for standard fresh-mined white ash anthracite, both by collieries and by operating companies. As, in the prices fixed by the President Aug. 23, 1917, a differential of 75c. per ton on pea size and above, equivalent to 52.95c. per ton on all sizes, was established for the independent operators over certain companies with rail- road affiliation, generally known as the l ' companies. ' ' AVERAGE AND BULK LINE COSTS OF WHITE ASH COAL Description Costs, Averages Returned Costs, Adjusted Cost, 90 Per Cent Bulk Line Excluding washery coal: All operations, each colliery separate All company operations, each colliery separate All independent operations, each colliery separate . . . All operations, each company operating two or more $3.85 3.71 4.37 3.85 $3.91 3.79 4.36 3.91 $4.80 4.65 4.97 4 38 Including washery coal: All operations, each company operating two or more collieries consolidated 3.57 3.77 4.36 AVERAGE PRICES RECEIVED FOR WHITE ASH COAL FRESH-MINED BANK TOTAL, INCLUDING COAL COAL BANKS Size Per Average Per Average Per Average Cent Price Cent Price Cent Price 6 8 $4 889 0.4 $4.416 6.2 $4 886 Eee 14 6 5 028 1 2 4.815 13.5 5 027 Stove 19 6 5 161 2.3 5.060 18.2 5.160 Nut 24 7 5 244 10.1 5.246 23.5 5.244 Pea 9 1 3 687 10.0 3.696 9.2 3.698 prepared and pea 74.8 $4.959 24.0 $4.544 70.6 $4.947 Buckwheat 11.6 $3.342 21.4 $3.213 12.4 $3.324 Rice 3.2 2.482 14.9 2.452 4.2 2.473 Barley 4.9 2.231 27.5 1.767 6.8 2.074 Boiler 3.9 2.341 8.8 2.123 4.3 2.304 Screenings 1.6 2.202 3.4 1.555 1.7 2.162 small sizes 25.2 $2 . 795 76.0 $2.339 29.4 $2.697 Grand total 100.0 $4.414 100.0 $2.868 100.0 $4 . 285 18 COAL MINING COSTS The prices received by the companies and independents have not been separately averaged, but, calculating on the differential and assuming the percentages the same for companies and independents which is only approxi- mately the case, the selling price of fresh-mined coal would average for companies $4.287, and for independents, $4.817. Margins over reported costs of companies would be 58c., and for independents 45c., with a general average margin for all fresh-mined coal of 56c., and for all coal, including washery, of 71c. per ton, and under "bulk line" costs, fresh-mined com- panies, 36c.; independents, 15c. ; total, 39c., including washeries consolidated sheets total of 7.5c. These margins include all expenditures for Federal income and excess- MINING COSTS 19 profit taxes, selling expenses, interest charges, expenditures for improve- ments and developments to increase output, excess of capital expenditures over normal cost, and all profit on the investment of about $8 per ton annual output. Effective Dec. 1, 1917, a labor war bonus, ranging from 60c. to $1.10 per day for labor and 25 per cent for contract miners was granted over and above the wage scales effective by agreement Apr. 1, 1916, expiring Apr. 1, 1920, and the prices fixed Aug. 23, 1917, and modified Oct. 1, 1917, by reducing pea coal 60c. per ton, were increased by 35c. per ton to compensate for this labor increase. The actual reported increase in labor costs due to this advance was figured by the Federal Trade Commission from the operators' reports to be 60. 3c. From the actual pay-roll figures later obtained by the United States Fuel Administration, this increase was found to be 76. 3c. per ton. Effective Nov. 1, 1918, a second labor war bonus was granted. The calculated increase in cost due to this is shown in Fig. 3, on which the increases for each operator are found by figuring from the pay rolls for the 6 months the actual increase in pay which would have been given, applying the Nov. 1, 1918, increases, and dividing by the 6 months' tonnage of the colliery. This line, adjusted to per cent of sizes, and compared with the adjusted cost, shows an increase in cost of 74. Ic. As this was necessarily applied to the prepared and pea sizes, 70.6 per cent of the total, the increase on those sizes was $1.05 per ton, which increase was allowed to balance the increased cost of labor. Except for the two increases to compensate for labor increases just noted and the reduction Oct. 1, 1917, of the pea coal price, the anthracite prices are as fixed by the President on Aug. 23, 1917. The present realiza- tion, all companies and all sizes, including washery coal and both the labor increases, is calculated to average $5.13 per ton, while the bulk line of the chart shown in Fig. 3, plus the Nov., 1918, labor increase, would be $5.32. The capital invested per ton output in the larger and better equipped collieries ranges from $5 to $11, with an average investment of from $7.50 to $8. PRICES FIXED BY THE PRESIDENT, AUGUST 23, 1917 WHIT: a ASH RED ASH LYKENS VALLEY Com- pany Inde- pendent Com- pany Inde- pendent Com- pany Inde- pendent Broken $4.55 $5.30 $4.75 $5.50 $5.00 $5.75 Eee 4 45 5.20 4.65 5.40 4.90 5.65 Stove 4 70 5 45 4 90 5.65 5.30 6.05 Chestnut 4 80 5 55 4 90 5.65 5.30 6.05 Pea 4.00 4.75 4.10 4.85 4.35 5.10 20 COAL MINING COSTS FIXED PRICES, DECEMBER 31, 1918 WHITE ASH RED ASH LYKENS VALLEY Com- pany Inde- pendent Com- pany Inde- pendent Com- pany Inde- pendent Broken $5.95 5.85 6.10 6.20 4.80 $6.70 6.60 6.85 6.95 5.55 $6.15 6.05 6.30 6.30 4.90 $6.90 6.80 7.05 7.05 5.75 $6.40 6.30 6.70 6.70 5.15 $7.15 7.05 7.45 7.45 5.90 Eee Stove Chestnut Pea The prices fixed by the President, Aug. 23, 1917, are given in the accompanying table. No price was fixed on sizes smaller than pea, which was decreased 60c. per ton Oct. 1, 1917. There was a general increase of 35c. per ton Dec. 1, 1917, and one of $1.05 per ton Nov. 1, 1918. Sizes smaller than pea were limited to a maximum 50c. per ton below pea coal by order of Nov. 15, 1918. AVERAGE COST PER TON, DECEMBER, 1917, TO MAY, 1918 Fresh- mined coal, 35,256,550 Tons Washery Operations, 3,431,916 Tons Total, In- cluding Washeries, 38,688,466 Tons Labor $2 593 $0 687 $2 423 Supplies 616 260 584 Transportation, mine to breaker Royalty, current 0.004 153 0.007 102 0.004 148 Royalty, advance 0.002 0.002 Depletion 099 077 097 Amortization of cost of leasehold Depreciation . . 0.014 091 0.024 0.86 0.016 0.090 Pro rata suspended cost of stripping .... Contract stripping and loading 0.023 0.009 0.021 0.009 Taxes, local 054 034 0.052 Insurance, current 0.016 0.014 0.016 Insurance, liability . . . 058 0.018 0.055 Officers' salaries and expenses 030 019 029 Office salaries and expenses 0.048 0.024 0.045 Legal expenses 005 0.003 0.005 Miscellaneous 0.026 0.023 0.026 Total $3 841 $1 378 $3.622 Increase over May to November, 1917 .. 0.764 0.365 0.719 MINING COSTS 21 The present fixed prices Dee. 31, 1918, per ton of 2240 Ibs. f .o.b. mines, are given in the accompanying table. Smaller than pea is not to be sold within 50c. of maximum pea-coal price. Thus the selling price of anthracite has been increased but 30.5 per cent over the pre-war price, while the cost of production has gone up 52 per cent, the difference having been absorbed by the operators. The average cost as reported for the six months, Dec., 1917, to May, 1918, inclusive, prior to the increase of Nov. 1, 1918, but including that of Dec. 1, 1917, is given in the accompanying table. Chief among the factors causing fluctuations in mining costs are the changes in wage scales and the tonnage produced. The labor cost per ton forms 70 to 80 per cent of the total f.o.b. mine cost. In 73 mining districts for which detailed statistics for 1918 were published by the Federal Trade Commission the distribution was as follows: Per Cent of f.o.b. Number 1918 Per Cent of Mine Cost that of Production, Total goes to Labor Districts Tons Tonnage 60 to 64 1 247,000 65 to 69 10 175,880,000 35 70 to 74 33 158,877,000 32 75 to 79 18 129,913,000 26 80 to 84 11 32,502,000 7 Totals 73 497 419 000 100 It will be noted that the difference between the labor cost proportions from district to district are much less than between the total costs themselves, which, excluding lignite, ranged from $1.62 per ton in a West Virginia field to $4.45 in an Arkansas field. The principal reasons for the difference in cost of the various fields are the physical conditions under which mining must be carried on, chief of which is the thickness of seam and the extent of the use of modern machinery for mining and transporting coal to the mouth of the mine, as compared with the old fashioned pick mining and mule haulage. It is a mistake, however, to try to measure the advantage one district has over another by a direct comparison of labor or even total f.o.b. mine costs. Allowance must be made for the much heavier investment neces- 22 COAL MINING COSTS sary In the fields where machinery is used to cut down the manual labor for mining and transporting the coal. Nor does it necessarily follow in a given field that the thicker the seam, the lower will be either the labor cost or the total f.o.b. mine cost per ton. The analyses of cost by thickness of seam mined shown in the Federal Trade Commission reports indicate that after a certain thickness of seam is reached in most districts between five and six feet the mining of thicker seams involves higher costs. In other words both labor and total f.o.b. mine costs per ton in a given field are likely to decrease as the thickness of seam increases from two feet up to between five and six feet, and then to rise as the thickness increases still further. Apparently in such cases it is the greater amount of labor and supplies required in timbering the thicker seams which increases the costs. Wage scales for each field are fixed with relation to the par- ticular mining conditions of the field. Common or uniform increases in existing wage scales, however, have widely different results in their effect on the per ton labor costs of the different fields. Thus it was that the wage increase granted in November, 1917, for which a uniform price increase of 45c. per ton was allowed, increased the cost about 28c. per ton in the Illinois mines of F. S. Peabody, according to his testimony before Senator Reed's committee. On the other hand, it has been found that this wage increase in some other fields increased the labor cost as much as 70c. per ton. The second important factor in causing changes in the per ton cost is the fluctuation in the tonnage produced. Obviously the greater the divisor, the less per ton will be the regular upkeep expenses whether in supplies or in general overhead charges. But also the proportion of so-called " non-productive " labor employed in the mine is so large with relation to the labor paid on a per ton basis that an increase in the production will often materially decrease the total labor cost. In fact the increase in production may be so great as to obscure, for a time, the direct effect of an increased wage scale. There is little authentic information available as to costs prior to 1916. There is reason to believe that for the period immediately preceding the war, while costs had been gradually increasing (disregarding the effect of fluctuations in production), there was MINING COSTS 23 no sudden jump, the increase taking place in the wages from time to time having been relatively small as compared to total cost. The 1916 costs, therefore, can be regarded as high water mark for a number of years previous. The experience of the anthracite field, where published labor costs are available as far back as 1913, supports this conclusion. There labor costs on fresh-mined coal of operators who produced about 60,000,000 tons annually were $1.62 per gross ton in 1913, $1.62 in 1914, $1.63 in 1915 and $1.75 in 1916. 3.00- FIG. 4. Production costs in Southwestern Pennsylvania for the years 1916 to 1920. The accompanying charts, Figs. 4 to 9, inclusive, for all of the principal producing fields in the United States these fields produced about 275,000,000 tons in 1918 show the rapid rise of costs since 1916 and also give some measure of the distribution of costs. The figures for 1916-1918 are taken from the Federal Trade Commission reports, those for 1919 and 1920 from the reports made by operators to the National Coal Association, and tabulated by the Senate Committee on Reconstruction ("Calder committee"). The allocation of these costs to labor, supplies and general expense for January -June, 1920, has been compared on the basis of the distribution shown in the Federal Trade Com- mission bulletins which covered the first half of 1920. The fields or districts are those established by the Engineer Committee of the Fuel Administration, and are defined as follows : 24 COAL MINING COSTS (1) Southwest Field, Pennsylvania: The counties of Allegheny, West- moreland, Fayette, Greene and Washington, in the State of Pennsylvania, except (1) that portion of Allegheny County from the lower end of Taren- 8.00 FIG. 5. Production costs in the Indiana district for the years 1916 to 1920. 8.00 FIG. 6. Production costs in the Illinois No. 6 district for the years 1916 to 1920 turn Borough north of the county line; (2) the territory in Westmoreland County from a point opposite the lower end of Tarentum Borough, north along the Allegheny Kiver to the Kiskiminitas Eiver, along the Kiskiminitas Eiver eastward to the Conemaugh Eiver, and continuing along the Cone- MINING COSTS 25 maugh Kiver to the county line of Cambria County; (3) operations on Indian Creek in Westmoreland County; and (4) operations in the Ohio Pyle district of Fayette County. See Fig. 4. 8.00 FIG. 7. Production costs in the Ohio No. 8 district for the years 1916 to 1920. FIG. 8. Production costs in the Pocahontas Field for the years 1916 to 1920. (2) Central Field, Pennsylvania: The counties of Tioga, Lycoming, Clinton, Center, Huntingdon, Bedford, Cameron, Elk, Clearfield, Cambria, 26 COAL MINING COSTS Blair, Somerset, Jefferson, Indiana, Clarion, Armstrong, Butler, Mercer, Lawrence and Beaver, and operations in Allegheny County from the lower end of Tarentum Borough north to the county line, and in Westmoreland County from a point opposite the lower end of Tarentum Borough north along the Allegheny Eiver to the Kiskiminitas Eiver and along the Kiski- minitas Eiver eastward to the county line of Cambria County, operations on the Baltimore & Ohio E.E. from the Somerset County line to and including Indian Creek and the Indian Creek Valley branch of the Baltimore & Ohio E.E. See Fig. 9. (3) Pocahontas Field, West Virginia: Operations on the Norfolk & Western Ey. and branches west of Graham, Va., to Welch, W. Va., including Newhall, Berwind, Canebrake, Hartwell and Beech Fork branches; also operations on the Virginian E.E. and branches, west of Eock to Herndon, W. Va. See Fig. 8. 3.EO annggnnnnnnanoggnDnonnnnnanannnnnnnnnnnannnnnnDcinrannnnnnn FIG. 9. Production costs in the Central Pennsylvania district for the years 1916 to 1920. (4) District No. 8, Ohio: The County of Monroe, the County of Bel- mont, except the township of Warren and operations in the 8-A vein in Flushing and Union Townships, the County of Harrison except the town- ships of Monroe, Franklin, Washington and Freeport, and the County of Jefferson except the townships of Brush Creek, Saline, Eoss, Knox and Springfield. See Fig. 7. (5) Bituminous Field, Indiana: Coal mined in the State of Indiana other than Brazil Block coal. See Fig. 5. (6) District No. 6, Illinois: Including Marion, Jefferson, Franklin, Williamson, Johnson, Hamilton, Saline, White, Gallatin, and mines along the main line of the Illinois Central E.E. between Vandalia and Carbondale in Clinton, Washington, Perry and Jackson Counties. See Fig. 6. MINING COSTS 27 The charts show clearly what has happened to costs since 1916 in some of the principal fields of the United States. The increases in these fields, based on the 1916 cost, are as follows : COSTS PER NET TON (Federal Trade Commission Figures Used Exclusively) PENNSYLVANIA WEST VIRGINIA Southwest Central Pocahontas New River Labor F.o.b. Mine Labor F.o.b. ' Mine Labor F.o.b. Mine Labor F.o.b. Mine 1916 (base) Jan.-March, 1920. . . April- June, 1920. . . $0.82 1.50 1.88 $1.19 2.13 2.62 $0.92 1.97 2.17 $1.32 2.56 2.82 $0.56 1.31 1.51 $0.87 1.89 2.11 $0.74 1.79 1.85 $1.00 2.44 2.39 PER CENT OF INCREASES OVER 1916 Jan.-March, 1920. . . April -June, 1920. . . 84 130 179 120 104 136 94 114 138 175 118 143 142 150 124 120 COSTS PER NET TON OHIO INDIANA ILLINOIS No. 1 No. 8 Bituminous No. 3 No. 6 Labor F.o.b. Mine Labor F.o.b. Mine Labor F.o.b. Mine Labor F.o.b. Mine Labor F.o.b. Mine 1916 (base) .... Jan.-Mar., 1920 Apr.-June, 1920 $0.84 1.65 1.64 $1.17 2.27 2.09 $0.78 1.54 1.79 $1.02 2.14 2.47 $0.87* 1.63 1.94 $1.09 2.00 2.41 $0.89t 1.57 1.77 si.iot 1.96 2.21 $0.86* 1.63 1.85 $1.07* 1.98 2.29 PER CENT OF INCREASES OVER 1916 Jan.-Mar., 1920 Apr.-June, 1920 97 96 94 78 97 130 110 142 87 123 83 121 77 100 78 101 89 115 85 114 * April-December, 1916. t July-December, 1916. In the foregoing table the 1920 figures, while not strictly comparable because not obtained from the same operators as the 1916 figures, are probably representative enough to show in a 28 COAL MINING COSTS general way the change in conditions since 1916. If the Septem- ber, 1920, total cost figures of the Calder committee be compared with the 1916 total f.o.b. mine cost, the increases shown would be yet more marked, as the effect of the wage increase late in the summer of 1920 was to increase costs. Such increase, how- ever, cannot be as exactly measured because the 1916 figures are "revised" costs and exclude selling expense, while the 1920 figures are "reported" costs, and include selling expense, etc. As the events of the past few years have shown, labor in coal mines, just as on railroads, holds the strategic advantage of being able to tie up the whole country through an effective strike, it is not likely that .costs will be materially lessened through any immediate reduction of wages. On the other hand, the necessary writing off of some of the heavy investment charges caused by high cost development during the past few years, in order to get the investment down to present day values, will in many cases increase the overhead charges. Method of computing tax returns. The intent of the law, according to an article on this subject in Black Diamond, is clearly that the cost of the coal in the ground shall be considered as part of the cost of the same coal when removed and sold. The regulations fail to carry out that intent, because of the unwar- ranted assumption that all the tons of coal in the mine cost the same amount per ton. This is what gives the Treasury Depart- ment's method its simplicity, but at the same time robs it of its reasonableness, because the assumption is contrary to fact.' It is self-evident that a ton of coal near the surface or exposed by present workings is worth much more than a ton buried far down in the earth that cannot be removed for many years. If no improvements in mining methods were expected, the value of deeply buried coal would be further reduced by the greater expense that would be required to bring it to the surface. There is a possibility, of course, that improvements in methods may keep pace with the difficulties encountered in the majority of cases. But whether or not inventions may be expected to offset in some measure the increasing difficulties of greater depth, there is nothing to compensate for the time element. And the time element is always an important factor. No extensive deposit, whether coal or ore or other minerals, can be mined in a day MINING COSTS 29 or a year. Almost invariably before a deal in mining properties is consummated the purchaser has, through the aid of specialists, made exhaustive studies as to the extent and quality of the deposit, and the most economical methods of its exploitation. It would not be economical to mine one ton a year, and it would ordinarily be impossible to mine the whole deposit in a year. But a plan is adopted between these extremes usually based on the annual production of a certain definite quantity, and the equipment and machinery to be provided in order to maintain that output is elaborated. Ordinarily the purchaser not only has these plans, but has concrete evidence of them that would convince any fair-minded and disinterested person that in purchasing the property the price he was willing to pay was based upon these engineering reports. And in arriving at that price he always does, either mathematically or intuitively, take into consideration the time that must elapse before his invest- ment can be realized in cash by operations. Let us suppose that the mine is known to contain 1,000,000 tons of coal, and that the plans call for the mining of 20,000 tons per annum, indicating a life of 50 yr. for the mine, and that with these facts in mind the coal lands are purchased for $100,000. This is the cost of the entire deposit. It indicates the average cost per ton is 10c., but although the average is 10c., that figure does not apply to the tons near the mouth nor to those deeply buried. To determine the cost of the several tons let us indicate by V the value of a ton exposed and minable to-day, and assume an interest rate of 6 per cent, and, to avoid unnecessary intrica- cies of computation, let us further assume that each year's production comes at the end of the year. The cost of the coal to be mined the first year would be The cost of the coal to be mined the second year would be The cost of the coal to be mined the third year would be 20,0007(^5}, etc. ; 30 COAL MINING COSTS The sum of this series for 50 terms constitutes the cost of the entire deposit. Therefore: + 1.06 ' (1.06) 2 ' (1.06) 3 ' (1.06)' . It is at once seen that the series in the bracket is equivalent to the present value of an annuity of $1 for 50-yr. at 6 per cent, which is readily computed at $15.761. Our equation is therefore reduced to: $100,000 = 20,0007(15.761). Solving for F, we find that the cost of one ton minable to-day is 31.72c. From this the cost of the 20,000 tons to be mined each year may be computed as follows: Cost Cost of Tonnage per Ton 20,000 Tons 1st 20,000 tons $0 .3172^ 1.06 =$0.2992 $5984 2nd 20,000 tons .3172-i-(1.06) 2 = 0.2823 5646 3rd 20,000 tons .3172-f-(1.06) 3 = 0.2865 5326 4th 20,000 tons .3172 -Ml. 06) 4 = 0.2512 5024 5th 20,000 tons .3172-K1.06) 5 = 0.2370 4740 10th 20,000 tons .3172-h(1.08) 10 = 0.1772 3544 15th 20,000 rons .3172-(1.06) 15 = 0.1324 2648 20th 20,000 tons .3172-h(1.05) 20 = 0.09897 1979 30th 20,000 tons .3172-=-(1.06) 30 = 0.05528 1106 40th 20,000 tons .3172-H1.06) 40 = 0.03088 618 50th 20,000 tons .3172-H1.06) 50 = 0.01725 345 If this table were filled in complete for the 50 yr. the last column would total up to $100,000, which is the cost of the entire deposit. If an interest rate of 8 per cent were adopted the cost of a ton minable to-day would be 40.87c., and the cost of the respect- ive groups of 20,000 tons would range from 37.84 to 0.871c. per ton. Other properties besides mines are acquired for lump sums, and it is conceded in those cases that the purchaser has the right and the duty to analyze his cost and set up in separate accounts a fair apportionment of it. Thus a taxpayer may buy for a lump sum a going store with all the assets and liabilities attached to it. He is expected to apportion this cost and set up separately in his books, the land, buildings, furniture, merchandise, accounts MINING COSTS 31 receivable, good will, accounts payable, etc. The apportionment must be fair, and the net total must agree with the aggregate cost. As another example a merchant buys a shipment of miscel- laneous hides for a lump sum. He then sorts them into numerous grades. The best hides may be worth many times as much as the poorest ones. He is expected to apportion the cost on the basis of quality and value. The apportionment must be fair and the total of the costs thus allocated to the several grades must agree with the aggregate cost. If the hide merchant sold all the poorer grades, but had the best hides on inventory at the end of the year, and priced them at the average cost, there is little doubt but that the Treasury Department would compel him to use a higher figure and would assess an additional tax. With the mining company the situation is reversed. In the nature of its operations it mines and sells first the most accessible coal the coal that really cost it most and is then asked to carry the less accessible coal on its balance sheet at the average cost. The law permits a reasonable allowance for depletion, and any mining company that makes a reasonable and fair appor- tionment of the cost of its mineral deposits should be accorded the same fair treatment that is accorded to the merchant of hides. Comparison of costs and distribution of revenue of German and Middlewestern mines. A study of mining costs and the distribution of revenue at the mines of the Westphalian Syndi- cate in Germany as compared with the mines of Illinois and Indiana throws some interesting light on these questions. The mine operators of the Westphalian district in Germany suffered from severe competition resulting from overproduction, and various efforts were made to find relief as early as 1850. Price agreements, which were forbidden by the German law, were disregarded, notwithstanding the heavy penalties imposed for violations. Finally in 1885 the Westphalian Syndicate was established, and continues to the present date. It is a selling organization without any property and only a nominal working capital. Its affairs are administered by an official who has no finan- cial interest in the mines and acts as chairman of a board made up of one representative from each participating company. The 32 COAL MINING COSTS function of the syndicate is to sell the product of the mines, coke ovens and briquetting plants and to allot to each company the tonnage which it should produce. Twice each year an estimate of the probable requirements is made, and a tonnage is allotted to each company based upon previous production after allowance has been made for the ton- 1075 1050 IOZ5 JTAM la-l / / / 975 ocn i i i 925 'QAfl i I 875 ftCA I *N ^ i 825 ttAA ' EA Ml RNINGS ILLINOIS PICK NERS WOULD HAVE HAL D RUNNING TIME EQUAL ttOF WESTPHALIA MM jf._ ''' HA m y 775 750 / \ / / \ / / \ ' / 675 650 25 / 575 550 525 500 / 4VERA ILLII\ T /W/V EARNINGS OIS PICK Ml NEKS ?/ ! ' / \ / ^ / / \ 1 ^ / 475 X 7 N 450 " 1901 B02 1903 1904 1905 1906 1907 1908 1909 1910 1911 I9IZ 191, YEAR FIG. 10. Wage losses to miners due to intermittent work. nage of companies, such as railroad, furnace and other such corporations, which consume a part of their own production. On May 1 each company is notified how much coal it will be called upon to furnish during the second half of the calendar year, and each mine can make its arrangements for the most MINING COSTS 33 economical production of the tonnage called for. Any company falling short in its supply, if market conditions continue as anticipated, must pay damages for the shortage unless the deficit can be made up by another company. v 125 2.00 21.75 ILK 125 1.00 075 f\Cf\ * ^^ x' x^ ^*^^ *. r X ^* DiV/l r- ^, ?f/J H 7 - ,*^^ _^-* x ^"' , - ^- ^ ^ ... **" nATL "RIAL 5,7> XES //vr/ 1 i ~RSTAND 6EI Vf>?> iX YAtf - 5 / s _ V ^~ I -^ ,, -~ , "* ^ **= ^i^ H^H -^ / ^\ nc I65i 025 1694 1896 1896 1900 I90Z 1904 190^ 1908 YEAR. FIG. 11. Distribution of gross revenue at Westphalia mines. 1594 1596 1900 190Z 1904 1906 1905 YEAR 1910 I9IZ FIG. 12. Production of the Westphalia mines compared with those of Illinois and Indiana. Losses due to inferior preparations are borne by the com- pany responsible for the defect. Prices are agreed upon and fixed in advance semi-annually and take into account the quality of coal produced from each mine, making it immaterial to the purchaser where the coal comes from, because of the adjustment of price to the intrinsic value of the material sold. It has sometimes happened that by some unforeseen con- dition the syndicate was not able to market through its ordinary 34 COAL MINING COSTS trade channels the estimated quantities of coal, and other markets had to be entered in order to permit the mines to operate under the most economical conditions. Losses due to these difficulties are borne alike by all, the syndicate paying to the participants the price agreed upon, having retained a commission, from which all deficits are paid. JC.3 son - - ESTPHAUA ^ '. / ^s^ ,/ >^ 275 250 225 200 175 150 I?R IL .INC IS 2 \ ~~***^. ^-^**" L s ~-^ > / / \ ^ ^ x * , ^- ' '* ^ \ ^ x v f / \ x' \ / "^ ^ / 1 1894 18% 1898 1900 1902 1904 YEAR 1906 1908 1910 1913 FIG. 13. Comparison of the average number of shifts worked per annum at the Westphalia and Middle Western mines. ^.bU 2.25 2.00 e U5 5 1.50 1.25 1.00 1.75 ,.50 *v^^ STP Mil* / f / "'" ^^ - ***" jS '^^ <^ INDIANA. L. S" * i *** ^^ Zi: 'SC ~^** +** ^^ IL L//VC 394 1896 1898 1900 1902 1904 1906 1908 1910 191 Y EAR FIG. 14. Average price per ton of coal at Tipple, Westphalia and Middle Western mines. The advantages of a single seller marketing 50,000,000 tons of coal a year are apparent. Markets are available to the syndi- cate which individual operators could not reach. Its contracts are made for five-year periods, and this assures an income to the operators and enables them to finance their properties and engage in business which, while more remunerative, requires larger investments. Thus they have erected large coking, by- product and briquetting plants. Such financing would be impos- sible with the uncertainties of ordinary competition. The higher returns have made possible an expenditure of MINING COSTS 35 money for improved equipment, safety measures and labor-sav- ing devices quite unknown in this country. Complete extraction of coal is required by the government, and it is estimated that the cost of flushing to sustain overlying strata and to permit of the removal of all coal adds 25c. per ton to the production cost. L-LO 200 C 175 Distribution of i Kfl B Revenue 1909 B-lncfuctes all general f) I7c salaries, but- notdepreci' I.C 7 3 ) vl flfl ~c C -1 ~ ~~~ 8iton t irnerest on invest 1 * ^ ment or sinking fund, C-Represents difference B B < between gross revenue " and operating expense O*75 ~~ including inte reside- predariofysinh'ng fvnd a i- and nei prof its For Jllinnis 2.6$ 1 V) UJ t, For Indiana 0.8t g 5 s ? purpose. s 1 g ^ * d S ft FIG. 15. Comparison of revenue at Illinois, Indiana and Westphalia mines. The coal operators are enabled to provide funds for the pro- tection of the injured employees and for the support of the families of those fatally injured. They also provide pensions for the incapacitated and the aged. The cost of this social insurance in 1909 added 20c. per ton to the cost of production. No protest has been made by the consumer against the higher coal prices which have followed the establishment of the syndi- cate. The increase in price has been generally accepted as the best expedient for solving a most vexatious question. Undoubt- 36 COAL MINING COSTS edly it induced more care and economy in the use of coal and resulted in the adoption of more economical engines and improved boiler settings. The Westphalia production increased from 1,665,000 in 1850 to 81,000,000 tons in 1907; at the same time the number of companies was reduced from 100 to 76, indicating growth of individual companies and concentration of capital. The 17 com- panies in the syndicate the output of which was sold for com- mercial use and which were not allied with the fuel-consuming industries had an aggregate annual production of 28,000,000 tons and a capitalization of $72,450,000 which is an average of $4,200,- 000 each. This indicates an investment for plant and equipment of $2.50 per ton of annual production. The capital account does not include any outpay for coal land, as all the coal belongs to the government. For Illinois the capital invested in 1909 was $1.49 and in Indiana $2.44 per ton of annual production. This latter, however, includes the coal rights, which represent the major portion of the investment. The accompanying diagrams, Figs. 10 to 15, inclusive, show graphically the points developed in this discussion. Conditions where operations may be conducted at an ap- parent loss. Maintenance charges for drainage, timbering, ventilation taxes, etc., are so heavy at mines that it may be more economical to continue operations in the face of an apparent loss than to shut down the mines entirely. Particu- larly is this the case with older mines having long underground hauls and high pumping heads. Some interesting data on this subject will be found in the report of a committee appointed to investigate the receivership of a prominent coal corporation about 1911. This committee found that if the coal properties were shut down, the annual loss will be $420,000. If they were operated at the standardized cost per ton of $0.857 and for an output of 3,000,000 tons and the coal sold at the price realized the previous year, $0.8097, the loss will be $141,900. The standard cost includes a charge for interest of $0.067 and for depreciation of $0.058, a total of $0.125 per ton. The standard costs are 14.8 per cent lower than 1909-10 corresponding costs, 17.4 per cent lower than July and August, 1910, corresponding costs. MINING COSTS 37 This short report was amplified into the following : The coal lands have been injudiciously acquired. Money has been injudiciously spent in equipping the plants. Overhead charges for interest, maintenance and depreciation are therefore high. The current market selling price for coal was so low as to make profitable coal mining difficult, if not impossible, even if the coal lands had been secured without price, and had been equipped with rigid reference to economical operation. To shut down the mines and wait for better prices would entail an annual expense for power, maintenance, supervision, depreciation and interest of $420,000. This does not include an annual charge of $104,494 on book value of coal lands not immediately identified with the plants to be operated. The cost of mining coal if operations are standardized, will be $0.857 per ton for a daily output of 12,000 tons, a monthly output of 250,000 tons and a yearly output of 3,000,000 tons. The loss from continued operation will depend on the price obtained for coal sold as follows: At $0 . 66 loss will amount to $561,000 At $0 . 70 loss will amount to 420,000 At $0.70 loss from operations and loss from suspension of operations will be equal. At $0. 79 loss will amount to 200,000 At $0 . 8097, price netted by coal sales in 1909-10, loss from operation will be 141,900 At $0. 857 there is neither loss nor profit from operation. At $0.921, profit above operation 192.000 This is sufficient to pay interest on obligation. Coal should therefore continue to be mined. At $0.948, profit from operation 272,0000 This pays for operation, for moneys owed and for present administration charges, While waiting for better coal prices, costs of operations were to be standardized as follows : By revaluing all the lands and equipment, thus reducing future operating overhead charges. By putting the management in the hands of a competent and experienced man of reliable character. 38 COAL MINING COSTS By concentrating operations at that plant, or those plants, where coal could be mined most cheaply. By investigating the advantages, if any, to be derived from coking the product of these mines. By investigating the advantages, if any, of establishing a washery at the mines. In making its investigations the committee attempted to determine a standard cost per ton of mined coal for a standard output, which was assumed at 3,000,000 tons each year. The standards adopted for immediate use were, per ton : The existing wage scale for mining labor, $0.485. Current rates of wages for a minimum amount of other efficient working labor, $0.175. Moneys for supervision, supplies and other bills, taxes, insur- ance, etc. ; an efficient minimum, $0.07. Depreciation charges based on revaluations, on experience, and on the present ascertained coal reserve tributary to operat- ing plants, $0.06. Interest at 6 per cent per annum on reappraised values of coal reserves, mining buildings, equipment, etc., actually used for mining operations, $0.067. Other expenses not standard and not directly appertaining to mining operations were : Interest and other charges on investments at present inopera- tive, $0.029. Excessive interest load, due partly to investment in elaborate and unnecessary plants, partly to deficits accumulated from former years, and partly to other causes, $0.035. High costs of administration of the company's business. COSTS FOR 1910 Operation $77,294 ' Maintenance 14,156 General expense, excluding insurance 37,912 $129,362 Less allowance for mining operation 48,000 $81,362 Cost per ton $0. 0271 The output of coal can fluctuate from no tonnage, if the mines are closed, to a maximum daily tonnage of 17,000 tons. MINING COSTS 39 If this maximum, of 17,000 tons daily could be attained it would reduce mining costs about as follows: OUTPUT PER YEAR, 4,250,000 TONS Costs per Ton Mining labor $0 . 455 Other labor 0. 15 Operation . 06 Depreciation . 06 Interest . 045 Total $0.77 TABLE ON BASIS OF 3,000,000 TONS ANNUALLY Daily Output, 12,000 Tons Costs per Ton 1. Mining labor $0.485 2. Other labor 0. 175 3. Total working pay-roll (1 and 2) $0.66 4. Operations $0. 07 5. Depreciation . 06 6. Interest . 067 7. Total overhead charge (4, 5/6) $0. 197 8. Total standard cost per ton of coal (3 and 7) $0 . 857 Systems of mining". A thorough grasp of the economics of the various systems of mining can be obtained best by a brief review of the various stages of development that have lead up to the adoption of the systems now in use. A study of the faults that were found in these older methods and the remedies that were applied to overcome the difficulties is of prime impor- tance. The changes in the systems of mining in the Georges Creek field, one of the oldest bituminous districts in the country, was described in a paper presented before the West Virginia Mining Institute in 1908, from which the following has been excerpted, disregarding the methods that were used prior to 1870 when there was apparently little attempt at systematized effort. Fig. 16 illustrates two methods followed during the years between 1870 and 1880. These workings are inaccessible to surveys at the present time owing to the creeps and squeezes induced by the irregular method of robbing the small pillars. The sketch was constructed from some old projection drawings COAL MINING COSTS MINING COSTS 41 and from information obtained from a number of men actually engaged in the work. The main headings consisting of haulage road and airway were driven on the strike of the coal. In the first method the room headings were driven in pairs from the main entry at intervals of 600 ft. and on the rise of the coal on about 10-per cent grade. From these headings approximately 25 rooms were driven to the right and left with 40-ft. centers on a grade of 4 per cent, giving an average length of about 350 ft. The rooms were 14 ft. wide and pillars 26 ft. These pillars were found to be totally inadequate and extracting them impossible. Cross-cutting the pillars at frequent intervals was then attempted after completion of the rooms, but this was generally accompanied by creeps closing a whole district at a time. The maximum height of the superincumbent strata in this territory is 200 ft. The second method, shown in Fig. 16, was adopted later. The maximum thickness of the overlying strata is 150 ft. By this method headings were driven from the main entry on the rise of the seam, at intervals of 1000 ft., to the level above, and two pairs of cross-headings turned to the right. The rooms were driven from these cross-headings at 50-ft. intervals and 14 ft. wide, leaving a pillar of 36 ft. The length of rooms varied from 300 ft. to 550 ft. These pillars were also of insuffi- cient size; robbing was conducted spasmodically, and, although more coal was recovered than in the adjoining districts, a great deal was lost. In addition to the small pillars, the method of robbing them was calculated to promote squeezes. It appears to have been the method to hold the strata with props until sufficient coal had been removed to enable the weight to break the props. As a general rule, however, before this was attained the weight had induced a creep, which is well known to have no limits within a territory of small pillars. Fig. 17 represents a method in use in 1890. The main entries were driven from the slope on the strike of the seam, sufficient grade being allowed for drainage. Cross-headings were driven on an angle of about 35 degrees to the main entry and headings turned off these parallel to the main entry. Rooms were turned, as shown, from all headings on 100-ft. centers, and pillars split by half-rooms. The length of rooms varied from 300 ft. to 600 ft., and were 15 ft. wide, leaving pillars 42i/ 2 ft. 42 COAL MINING COSTS MINING COSTS 43 wide. These pillars were not strong enough to support the overlying strata, 500 ft. high, and the usual creep or squeeze resulted when pillar drawing commenced. This half room method has the advantage of facilitating gathering of coal and doubling the support of the haulway. The squeeze in this district could have been prevented by turn- ing the rooms from the haulway on 200-ft. centers and, after driving the half-rooms, the resultant pillars would be 85 ft. wide. While this would have avoided a squeeze, the great weight to be supported by this pillar of soft coal would not have per- mitted a very high percentage of recovery. Fig. 18 shows a method adopted in 1900. The maximum dip is 15 per cent, and the greatest thickness of superincumbent strata 425 ft. The slope, together with parallel air-course and man-way, are sunk on the heaviest dip of the coal, and double entries turned off to right and left at intervals of 1000 ft. on a grade of 1% to 2*4 per cent in favor of the loads. From these haulways cross-headings are deflected at intervals of 240 ft. at an angle of about 25 deg. and driven on a grade of 4 per cent to 7 per cent. Rooms varying in length from 100 to 800 ft. are turned on the rise of the coal from these cross-headings. The rooms are driven 15 ft. wide on 65-ft. centers, leaving pillars 50 ft. wide. Twenty-five rooms are driven in each of these diagonal panels. Unusually large protecting pillars are left along the main haulage roads. This system has been found to be especially adapted to rapid gathering of cars thus ensuring a large tonnage. It has been found, however, that a very large recovery from the pillars is impossible, owing to the many sharp angles, which, in a thick seam of soft coal, are always difficult and ofttimes impossible to extract. This sharp-angle method was even resorted to formerly in cross-cutting the pillars preparatory to drawing them, but this has been changed to a rectangular method, thereby increasing the actual percentage of pillar coal recovered from 80 per cent to 83 per cent. The distance of rooms apart has also been increased in the last few years to 100-ft. centers giving pillars 85 ft. thick. It is expected that the extraction of these will show a further increase in the percentage of yield from pillars. The present yield from head- ings, rooms, and pillars under this system is about 90 per cent, 44 COAL MINING COSTS MINING COSTS 45 considering the recovery from headings and rooms at 100 per cent. Fig. 19 illustrates a method instituted in the latter part of 1904. The main haulway is an extension of the slope from the opposite side of the basin. Double entries are turned off from this entry, on 1%-per-cent grade, 400 ft. apart, from which rooms are driven directly on the rise of the coal. Rooms are from 13 ft. to 15 ft. wide and practically no barrier pillar left between the room face and the air-course of the panel above. They are driven at 100-ft. intervals, leaving a pillar 85 ft. wide. The length of a panel is about 2500 ft., containing 22 rooms. There are five such panels in this district and when completed it is proposed to draw the pillars in a retreating fashion with the line of pillar work on an angle of 45 deg. across the whole district. A similar method in another district, but with rooms on a deflection of 35 deg. from a right angle, is yielding SSy 2 per cent from the pillars with a total recovery of 94 per cent from headings, rooms, and pillars, and it is believed that this can at least be duplicated if not exceeded in the case of Fig 19. The maximum dip in this district is 6y 2 per cent with the greatest height of the overlying strata 250 ft. With this resume of systems used at different periods under conditions now known in view, a suggested method of extracting the coal from thick soft seams with a brittle top and a height of superincumbent strata of 400 ft. or less is presented in Fig. 20. The general design of the mine for haulage, drainage, and ventilation is not given, because they are variable quantities, depending on conditions which change with the locality, and the method suggested is therefore limited to the ultimate recov- ery of the coal. By this method a territory under development is divided up into rectangular panels of 10 rooms each. The room headings are driven on easy grades favorable to drainage and haulage and the panels worked in pairs. When the upper heading has been driven to its end the rooms are turned at intervals of 100 ft. with the drawing of pillars following the retreating method. The rooms are 400 ft. long and 13 ft. wide, leaving pillars 87 ft. in width. The rooms in the upper panel are limited by a barrier pillar separating them from the head- ing above, and those on the lower panel are driven through to the gob of the upper panel. The line of pillar work extending 46 COAL MINING COSTS MINING COSTS 47 over the two panels should have an angle of about 45 deg. The length of rooms can be varied to suit the conditions, and, when the height of the overlying measures exceeds 400 ft., the thick- ness of pillars should be increased accordingly. FIG. 20. Suggested method of working thick, soft seams, having a brittle top and 400 ft. cover. Fig. 21 shows the method of drawing the pillars in detail. The rectangular method should be carefully adhered to and all sharp angles avoided. When the work commences a cut is made separating a block of coal 30 ft. wide and 87 ft. long. This piece is further subdivided into blocks varying from 6 X 12 ft - to 15 X 12 ft - in size > which are cut off and extracted successively as shown. In no case should the small blocks cut off exceed in size the distance a man can shovel under average conditions, which is about 15 ft. The largest of the small blocks 48 COAL MINING COSTS should be removed last because it is there that the greatest pres- sure manifests itself in the removal of the original block cut from the pillar. When this block has been removed another cut is made and the process repeated. The line of gob should be approximately on an angle of 45 deg. If it is found impos- sible to take out the small blocks clean, on account of the gob from the preceding work running in on the new cut, a row of props can be set on the upper side of the large block to hold the gob when the small blocks below are removed. This method L. FIG. 21. Method of drawing pillars to be used in the system shown in Fig. 20. of working should result in a total average recovery of fully 95 per cent of the seam from headings, rooms, and pillars and ensure the workings against creeps and squeezes. One of the most interesting studies of costs as applied to the different systems of mining flat seams of coal was included in a paper presented before the February, 1915 meeting of the Am. Inst. of Mining and Metallurgical Engineers, excerpts from which are given herewith. It may be stated that the unit with which we have to deal in studying mining costs is the room ; what takes place at its face is the real productive work of the mine, and all else under- MINING COSTS 49 ground is for the purpose of serving best the worker at the room face. Fig. 22 shows several typical methods of procedure. They are of particular interest in that one may see them in mines following the same plan, working the same seam, under con- ditions which admit of comparison. The features of these methods are given in Table I. In all of these methods variations may be seen, from entries driving with no rooms turned to entries driving with two or more rooms turned and driving as the entries advance ; in respect to the robbing, one may see variations from robbing following --Pillar-. -H 2/K-4ET-H ; I D E F FIG. 22. Typical methods of procedure in working rooms. immediately upon the completion of the first two rooms to the robbing following at an indefinite date after the completion of the first workings of the panel. Where continuous paneling, or advancing robbing, is in effect, robbing is not compelled to wait until the completion of all the entries of the panel. The number of rooms per entry varies from about 12 to an indefinite number, and the depth of the room varies from about 300 to about 800 ft. The amount of timber and the manner and time of placing same depend largely upon the individual miner, and as a rule there are no specific instructions for his guidance ; also, in general, no effort is made to recover the timber in robbing. 50 COAL MINING COSTS TABLE I METHODS OF PROCEDURE IN DRIVING ROOM Sketches A B C D E F Width of room in feet 24 20 20 30 36 36 Width of pillar in feet 36 65 40 45 54 54 Location of track . . In center of room Along robbing rib Along robbing rib Along robbing rib Along robbing rib Along robbing rib Location of gob . . . Along both ribs Opposite robbing rib Opposite robbing rib Between tracks Opposite robbing rib Between tracks Number of men per room Ito2 rooms 1 lor 2 2 6 4 Feet of room face per man Feet of entry per man 48 120 20 85 20 or 10 60 or 30 15 37.5 8.5 15 9 22.5 A method of procedure observed at one mine (but which has not as yet been sufficiently tested out in the matter of recov- ering the pillar to warrant its unreserved adoption), is shown in Fig. 22, E. Here it is intended that rooms shall be driven 36 ft. wide on centers 90 ft. apart, carrying a room face at an angle of 45 deg. and a single track along the robbing rib but curved to parallel and follow the length of the room face. It is intended to work six men to the room, the gathering motor receiving and placing three cars at a time. Immediately upon the completion of the room the pillar is to be withdrawn. By this method of procedure a, high degree of concentration will be effected and the efficiency of the gathering motors, mining machines, and miners will be increased. It is also hoped that by carrying the working face on a diagonal, fewer unexpected falls of top will occur than at present, because the fracture will generally be partly exposed before the entire coal support is removed from beneath it. In most mines the miner at the face is responsible for the safe working conditions of his room. In the above methods of procedure, therefore, one might say that, for the same expendi- ture of time, energy, and watchfulness, the relative degree of MINING COSTS 51 security which the miner may feel as a result of his efforts is inversely proportional to the room space he occupies. It is also true that for cars of the same height the energy expended by the miner, or the work done in loading the coal, is much less where two or more men work per room and the room space per miner is low, than where one man works per room and the room space per miner is high. In the grouping of rooms as outlined in Fig. 22 many arrangements were made from which have matured certain well- defined plans. Probably the consensus of opinion favors the panel system, but even with it there are differences of opinion. Many men think that the entries should be driven to the inside lines of the property and the coal extracted retreating; others think that half of the property should be worked advancing and the remainder retreating ; yet others think that all or nearly all of the coal should be extracted as the entries advance. It is probable that it is best to extract the coal in such a manner that the present worth on the returns from the mining venture will be greatest, both to the property owner and the operator, leaving only such coal during the advance of the entries as will permit of profitable mining until the final exhaustion of the property. Figs. 23 and 24 show typical plans on the panel system. Fig. 23 is the square or rectangular panel, Fig. 24 the continuous panel. For purposes of discussion and demonstration a typical property of 1000 acres of the approximate shape indicated in Fig. 25 will be assumed from which it is desired to produce 2800 tons per day when running at maximum. The dip is 2.5 per cent and the strike lies at an angle of about 10 deg. to the long axis of the property, its general direction being from the upper right hand corner of the property to the lower left hand corner. The coal is fairly clean, 6 ft. thick, and the condition of grades, top and bottom, are fair. It is also assumed that the rate of loading per man per day will be 16 tons. The questions to be decided are what method of procedure and what plan are best, to determine which the following information is desired : 1. What period of time will be required to reach the desired output ? 2. How many day laborers, mining machines, mine cars, mules for gathering, and main-haulage motors will be required? 52 COAL MINING COSTS t. Barrier Pillar Barrier Pillar $- -I ain Errfry FIG. 23. Typical plan of mining on the square or rectangular panel system. ...A i I I l l VI I I I I I \ I I IJ_U_J . 24. Typical plan of mining on the continuous panel system. MINING COSTS 53 3. How much main entry, main entry track and trolley wire ; cross main entry, cross main entry track and trolley wire ; room entry, room entry track, and rooms and room track will be required ? 4. What is the length of the average car haul? 5. What is the relative amount of power for ventilation? 6. What is the acreage of standing pillars, the estimated relative cost of production, and the estimated percentage of recovery ? The methods referred to in Fig. 22, 0, and the plans of mining in Figs. 23 and 24 will be applied to the problem as follows : First Form. Drive the third entry of the panel, turn the last two rooms on this entry first, start removing the pillar immediately upon the completion of the next to the last room, and continue to drive all the rooms in the panel only fast enough to provide for the uninterrupted advance of the robbing. Work two men to the room and in the air-courses and on the pillars. Only this method of procedure will be applied to the square and continuous panels, and the following methods to the square panel. Second Form. Drive the rooms of the panel as they are encountered, turning the first entry of the panel when it is reached, and start robbing immediately upon the completion of the last room on the third entry of the panel. Work one man to each working place. Third Form. Drive the rooms and entries of the panel as they are encountered, start robbing immediately upon the com- pletion of the last room in the third entry, and continue the robbing until the completion of the panel. Work one man to every other room, but advance all rooms and work one man to each pillar. The accompanying table shows the comparison, as well as some other figures to which further reference will be made. From these data it may be concluded that the first form of procedure and the plan of mining, Fig. 24, are best. The period of time required to reach the desired output was determined for the several methods, as shown in Figs. 25, 26, 27, 28 and 29; the location of the working faces from day to day, as determined by the assumed rate of advance of 16 54 COAL MINING COSTS tons per man, was plotted on a map, and the total number of faces at the time the desired output was reached were counted, from which data the tonnage curves were plotted as shown in Figs. 25A, 26A, 27A, 28A and 29A. First Form Continuous Panel Second Form Third Form Ad- vancing Method Output reached months 53 42 62 92 7 Day laborers 60 65 82 102 31 ADVANCING METHOD USES 8 ASSISTANT FOREMEN Mining machines . 9 9 14 18 8 Mine cars 275 310 335 465 155 Mules 18 22 24 32 10 Motors 4 4 5 6 2 7 850 6500 9 300 13 950 600 Main-entry trolley Cross main entry Cross main-entry track Cross main-entry trolley Room entry and room-entry track Room track . 5,550 10,700 15.84C 5,150 12,700 20,500 8,850 33,900 96,800 13,500 50,400 230,300 1,000 7,000 18 100 6,180 5,333 7 420 10 230 3 640 Ventilation power, kilowatt hours 40 62 8 42 65 7 125 168 2 175 277 20 13 8 Relative cost of production 1.33 1.24 1.76 2 1 1 Percentage of recovery 94 95.5 83 80 97 In arriving at the relative number of men and mules required, rates for performing certain tasks, taken from time study observations were used. The amount of rolling stock required is based on the assumption that the equipment will travel at the same rate of mileage per day ; the other items com- pared were taken direct from the maps. In the absence of facts for comparison, opinions of creditable authorities were sought in very instance. These methods of procedure and both plans of mining have been designed to meet certain wants. In some instances certain features of the plan have been prescribed by the land owners in order to safeguard their interests from ' ' squeezes ' ' and losses of coal due to lack of proper supervision. Were the proper supervision supplied and better methods of procedure adopted, the restrictions in the plan of mining might very properly be MINING COSTS 55 COAL MINING COSTS removed. Other details of design have been the result of accept- ing certain " rules of thumb" which have since been proved wrong, and yet other details, although admittedly wrong and expensive, have been introduced rather than combat the wrongs which they are designed to circumvent. In the plan of mining shown in Fig. 23, the frequent inter- position of barrier pillars is for the purpose of confining a squeeze and limiting its range of destructive action. The use of these barriers is imperative under the methods of procedure 3000 2750 2250 ^2000 o o ,.1750 & 750 500 250 Tofal Tonnage "'from all Places (Lfooms and ftobbingl"-? Flat I <-Rooms and fobbing 3& 'Flat Mam Entry and First F/crf. '^and 3^ Flat Entries 20 25 30 60 65 Months of 25 Days per Month FIG. 25A. Tonnage curves under the method of procedure shown in Fig. 25. that involve large areas of long-standing pillars and where the degree of supervision is low. It is to be regretted that their use is so common, for they tend to interfere seriously with the maximum degree of concentration because one is seldom, if ever, able to provide a satisfactory output from a single panel, and then only for a short period of time. Where two or more panels are required to produce the output, the further the work- ings advance the more distantly seperated they become, or other important considerations must be sacrificed. Disadvantages of the unit-panel plan may be seen in the curve in Fig. 30 which shows the great variation in the tonnage obtained daily, vary- MINING COSTS 57 58 COAL MINING COSTS ing from zero at the opening of the panel, augmented by a more or less constant rate of increase, to a certain maximum number of tons, and then a gradual decline to zero again. If a certain number of tons per day gathered from the panel is accepted as 100 per cent efficiency for a gathering motor, as shown in Fig. 30, it will be noticed that the motor is at first working at a very low efficiency, which gradually increases until the maximum is reached, at which time another motor must be added, and the average efficiency of the two motors ig ^-, t* 2250 1500 1250 1000 750 500 250 A f & cf / 4** A* / / f. x> U fc \6 !n\ / S\ ^ u M / i A I \ , A A / \ A \ \ 1 /f* !ah+ / i I Y \ / Y V \ \ 29 // \ 2 / \ /\ / \, binew tflhjn te?w/ A \ ^/- f-f^RiahH 3 _1 A| A / \ 1 ^ \ 5 10 15 20 >5 ' M< >nths 5 40 45 50 55 60 < at 25 Days per Month 5 70 75 f 'O 85 9< FIG. 26A. Tonnage curves under the method of procedure shown in Fig. 26. about 50 per cent; there is a similar drop in efficiency with each motor that is added, until the maximum tonnage from the panel is reached, after which the process of removing motors from the panel is begun. In some measure this degree of effi- ciency may be increased by working the motors over more than one panel, as is often done, and a better efficiency curve might be obtained more nearly in accordance with the full line shown, but in practice a rigid watch must be kept on this detail, or more often than otherwise a lower degree of efficiency than that shown will result. If, in the preceding paragraph, instead of considering the MINING COSTS 60 COAL MINING COSTS efficiency of the gathering motor the efficiency of day laborers or the tons produced per unit of material and equipment in use had been considered, the same general discussion would apply. Where low efficiencies are obtained from day laborers, material, and equipment, low efficiencies are also obtained from the miners at the room face. For these reasons it is difficult, and in practice well-nigh impossible, to establish any constant relation between a given tonnage desired to be uniformly pro- 3000 25 3O 35 4O 45 50 55 6O 65 Months oi 25 DCILJS per Month FIG. 27A. Tonnage curves under method of procedure shown in Fig. 27. duced, and the amount of material, equipment, and day laborers required to produce that tonnage; the efficiency of these quan- tities rises and falls with the rise and fall of the tonnage curve, although in an erratic manner. Thus it would appear that the square panel, while designed to meet certain requirements, does so at the loss of much that is to be desired, and introduces new complications. The barrier pillars are, as the term implies, for the purpose of barricading against some impending danger, such as an unforeseen squeeze. Since no one can predetermine where or when these squeezes MINING COSTS 61 62 COAL MINING COSTS will occur it sometimes happens that the barrier pillars are provided where they are not needed; yet experience has shown the wisdom and necessity of their use under certain conditions. & f* -V V v i V \i Si 1 1 1 1 he>Q js 1 i They would be used less frequently if the square panels were made rectangular, but the same degree of security would not be obtained unless the entries were driven to the limit of the rectangle, with few or no rooms driven as the entries advance. If we accept it as axiomatic that when a room is driven to com- MINING COSTS 63 64 COAL MINING COSTS pletion its pillars should be immediately removed in order to obtain the best results, or that it is equally as fundamental to open up no new entries until ready to mine from them, and 13 ,1. =8 \ I 2 (n?q jdcl suoj. that mining should then be conducted at the maximum rate of production, the rectangular panel that involves either long* standing pillars or long-unproductive entries must be rejected. The continuous panel obviates the necessity for frequently interposing a barrier pillar and it is especially well adapted MINING COSTS 65 to a property where the main entries are driven to the dip. However, the tonnage from a single continuous panel is limited, and where the main entries of a property go to the rise the maximum degree of concentration cannot be obtained or the rooms off the cross entry will go to the dip. Advancing robbing is impracticable because the pockets in the pillars go to the dip. The rate of production from a single room entry rises and falls in the same manner as the rate of production in the room entries of the square panel and the general discussion above in reference to the square panel applies to the continuous panel. l*t>O 1400 1350 1300 1250 1200 1150 1100 1050 1000 950 900 650 <750 u700 g.650 -600 00 1 8 n " K. .\ /! \ f j* ^ . 1 \ \ [J 1 1 \ 1 V 1 \ 1 \ \ / V J500 f> J \ . i 400 350 300 \ J s J \ 250 ... % - 3 S < ^*- -*. ^ N- ^ J - \ 150 100 50 s* \ j g k > v- y fT *- '.fficienci/ Curves \ 40 % -f i S of- Ocrtherinq Motors V C f^ \ 1 i 1 2 2 4 5 6 7 8 9 JO 12 14 16 18 20 22 24 26 28 30 3 1 34 Months FIG. 30. Variation in the tonnage daily obtainable from the unit panel and the efficiency of gathering motors working in the panel when proceeding as outlined in Fig, 26. However, if one follows the history of the development of mining methods from the early-day single-entry system to the present-day panel system, it will be found that the square or nearly square panel meets sound mining practice more closely than any of the plans which have preceded it. Until methods of procedure are adopted which make the restrictions of the panel unnecessary, or until a plan of mining is devised with- out the objectional features of the panel, but retaining its may favorable features, the square panel will be accepted by many operators as the standard plan of mining. 66 COAL MINING COSTS For many years it has been the common belief that coal could be most economically cut and blasted by using a depth of cut equal to the height of seam. This erroneous idea fre- quently resulted in blasting down more coal than could be loaded in one day, and was the cause of allotting more than one room to a miner. That the height of seam does not bear any direct relation to economical cutting or blasting was demon- strated by the United States Coal & Coke Co. at Gary, W. Va., working with a Sullivan shortwall machine, and it has been found that miners are pleased to work two or more to a room, provided their earnings are as great as when they work in rooms by themselves. Probably the most marked results in devising more eco- nomical mining methods has been achieved by the officials of the above-mentioned company. They have realized the objec- tions to the mining methods outlined above and applied them- selves to working out a plan which would be simple, direct and efficient. They accepted it as axiomatic that any change in the prevailing plans of mining must be beneficial to the property owner, operator and miner alike, for any change that would benefit one or more of the interested parties at the expense of the others would not last. In this study difficulty was experienced because of the entire lack of systematized knowledge as to the proper relative rate of advance of room to retreat of pillar, the most economical width of room, and in fact what might be considered 100 per cent efficiency for any man, animal or machine about the mines. In order to determine these data, which were absolutely essential to an intelligent solution of the problem, a series of time studies was instituted and extended over a period of weeks, covering all the motions that make up certain underground operations that have to do with getting the coal from the working face to the railroad car. Thousands of observations were taken, properly checked, tabulated, collated, and used as a basis for a method of procedure, which has been put to the rigid test of practical use with remarkably good results. This method of procedure has for its object the maximum degree of safety, sanitation and opportunity to the miner and of security to the property owner, while at the same time offer- ing the greatest advantage to the operator. It combines a MINING COSTS 67 maximum degree of concentration with a minimum expenditure for labor, material and equipment, in such a manner that these quantities bear a constant relation to the output. Its use has resulted in a marked reduction in fatalities, increased earnings to the miners, decreased costs per ton for labor, material, equip- ment, and capital, and the recovery of practically all the coal in the seam. At Gary, W. Va., mules are used for gathering, and as a result of concentration their efficiency has, in some instances, been increased to over 200 per cent. At one of the mines, fewer day laborers are employed underground than are employed about the tipple. For the purpose of comparing the results obtained under this method with those from the several methods of procedure in the panel system, the method has been applied to the property and the problem under consideration. Fig. 31 shows the arrange- ment of the workings at the time the desired maximum out- put is reached. It also shows the details of the method of pro- cedure; the other data desired are given in Table II. Fig. 32 shows the tonnage curve and, for comparison, the total tonnage curve from the unit entry shows that the tonnage rises rapidly curve, from the unit entry shows that the tonnage rises rapidly until the maximum is reached and then continues indefinitely at that rate. By using available data, the proper length of room, angle of breakline and angle of advancing faces may be pre- determined, so that the total daily tonnage from the entry is the multiple of the tons that can be hauled by a mule or motor ; thus, the mules or motors are always working at the maximum efficiency. It is equally true that when the workings have advanced for a short distance, after reaching the maximum ton- nage from the entries, the estimated minimum number of day laborers required, may readily be confirmed, and once the entry reaches its maximum tonnage, and the quantities of labor, material and equipment have been accurately determined, these quantities remain constant throughout the entire extent of the entry, which may be as great as the property is long. The room space occupied per miner is less than in any of the other methods now in effect, which is an index of the rela- tive degree of safety a miner obtains for a given expenditure of time and energy. The excellent manner in which the rooms 68 COAL MINING COSTS 8 i S^^S* 5 C 6* w* * "" 2>V I II ! I II A, ^c^ \V'V^ -5-t, 5 ^ MINING COSTS 69 are timbered, shown in Fig. 31, is the minimum required ; where the mine foreman or miner has reason to believe that additional timber is required to make the place safe, the miner must place additional timber before doing anything else. As the entries advance, all rooms are driven and robbed immediately upon their completion, and rooms are opened up only fast enough to provide for the uninterrupted advance of 3000 10 15 20 25 30 35 "JO 45 50 55 60 65 Months at 25 Days per Month Flo. 32. Curve showing rate of development to the desired output, under the method of procedure, sketch F, Fig. 22, and the advancing plan of mining, Fig. 31. the robbing. Thus no barrier pillars are required, for the virgin coal protects the workings on three sides and the weight of the roof is resting on the bottom of the robbing. If a disturbed area of coal is encountered, or for some reason it is desired to discontinue the panel, a barrier pillar may be introduced at any time exactly where it is needed and the entries continued for the purpose of exploration. In order that the different methods of mining may be readily compared, Fig. 33, showing the relative amount of labor, 70 COAL MINING COSTS material, and equipment required to produce the tonnage desired from the property shown in Fig. 25 ; also the acreage of stand- ing pillars, the relative cost of production, and the estimated percentage of recovery. Any method of procedure that does not provide for the removal of pillars immediately on the completion of a room is fundamentally wrong, because it involves long-standing pillars open to the unfavorable influence of atmospheric agencies and other forces of nature ; the duplication of track work ; the clean- Fferiod Required Day Mining to Reach the Laborers Machines Dciired Output Required Required Mine Cars Required Mules Required I Motorii Mam EntmTrack Crois Mam Entry, Required and Trofley Trackand Trolley Wire Required W.re Required FIG. 33. Comparison of the amount of labor, material and property required when following the methods of procedure shown in Fig. 25 and the relative cost per ton, the recovery, and the period of time required. ing up of many slate falls that might otherwise have been avoided; and the scattering of workings, all of which increase the cost per ton for labor, material, and equipment, and cause the pillar coal to be badly disintegrated and low in domestic lump sizes. It sometimes happens in practice, however, that fundamentals must be sacrificed to adapt the method to peculiar conditions encountered, often resulting in lack of concentration and large areas of standing pillars. Where considerable tonnage is desired and a new property is being opened, skilled miners, experienced in robbing pillars, are hard to get and frequently the officials, MINING COSTS 71 mine foreman, and underbosses are not experienced. In order to keep up the tonnage under these conditions, . the workings must necessarily become distantly separated, because coal can only be obtained from room workings. It frequently happens also that the rates for mining pillar coal and room coal are not properly adjusted, so that the men can earn more in room work than in pillar work, naturally causing the pillars to lag behind, and requiring the introduction of barrier pillars to safeguard against squeezes; these barriers in turn cause a further separa- tion of the workings, and a decrease in the efficiency of labor, Room Entry and Roomsand Room Entry Room Track Track Required Recovery FIG. 34. Equipment required to produce an output of 2800 tons per day from the different plans of mining outlined. Also the acreage of standing pillars to reach the output. material, and equipment. The natural impulse of the mine fore- man, therefore, is to open up more rooms in advance of the robbing in order to increase the efficiency to something like a proper standard. For these reasons the territory for a given output during the development period should be as isolated as possible, and no greater in extent than is practicable. After the development period is passed and the organization perfected, there is no good reason why a mine operation should not be conducted with much the same regularity as a blast furnace or an industrial railroad. The fallacy that the average miner will load only so much 72 COAL MINING COSTS coal and no more has long since been exploded, and it is a matter of every-day observation that the miners are pleased to load coal if the mine cars are given to them with some degree of regularity and with some relation to the time required to load a car. When one considers that a coke loader, work- ing under the heat of the sun, and of the coke ovens, will FIG. 35. Standard plan of mine development adopted by the Pocahontas Coal & Coke Co. load from 35 to 40 tons as an ordinary day's work, there is no reason why a miner working under so much more favorable circumstances should not load at the same rate. In this con- nection the following observations that have to do with load- ing coal underground are interesting. These figures show that less than 47 per cent of the time spent underground was consumed in loading coal and over 12 per cent of the time was lost waiting for the empty mine MINING COSTS 73 cars. It may be stated further that these men were loading at the rate of 35 tons per day of 8 hr., and actually did load at the rate of 16 tons per man per day per year. Methods of working in the Pocahontas field. The entire Pocahontas field proper is practically all leased out on royalty FIG. 36. Double-entry system of mining used by the Upland Coal & Coke Co. by two large holding companies, the Pocahontas Coal & Coke Co. and the Crozer Land Association. Under the lease con- tracts, the holding companies have reserved the right to define the method of working, and the result has been satisfactory. A standard plan of mining, by Thomas H. Clagett, chief 74 COAL MINING COSTS engineer of the Pocahontas Coal & Coke Co., is shown in Fig. 35 ; this is largely followed by their lessees, although in instances materially modified, due to local conditions. This large hold- ing company owning or controlling some 275,000 acres of Pocahontas coal, has in active operation some 45 leases (in 1913), covering about 145,000 acres. One of the special advantages of the system of mining adopted by the Pocahontas Coal & Coke Co., is the relatively quick recovery of the pillars, and the panels are so driven that the rooms and all entries split the pitch; thus if the maximum pitch is 3 per cent, then the maximum for the work- ing will not exceed 2 per cent and may be even slightly less. Further particulars of this system appear on page 77. The method of working adopted by the Upland Coal Coke Co., on one of the Crozer leases, is shown in Fig. 36. Were the crop line shown on this plan it would be evident that the break line is carried in from the crop and does not involve, strictly speaking, breaks in the solid. There may be several " lifts" where the width of the lease is too great to admit of one lift only, as shown. This plan of mining was evolved from a number of years of revisions and has been found satisfactory under all conditions. The main entry is to be driven as near the line of strike as possible, in order that the reverse grade against the loads may be negligible. If the cross entries are turned off at more than 90 deg. from the main, and the rooms are less than 90 deg. of the cross entries, grades in favor of the loads may be obtained. The Pocahontas Consolidated Collieries Co.'s Angle col- liery, as of July, 1912, is shown in Fig. 37. Soon after taking up the work of the old Norfolk Coal & Coke Co. (which was essentially the nucleus of the Pocahontas Consolidated Col- lieries Co.) in 1904, the work of revising the systems of mining was taken up in detail. One of the first improvements adopted was the introduction of the multiple air-course system. Further interesting particulars regarding mining methods in this field were given in a paper presented before the West Virginia Coal Mining Institute in 1913, from which the follow- ing has been excerpted. It is bad practice to drive up a room and allow the pillars to stand in the expectation of drawing them later. A better MINING COSTS 75 76 COAL MINING COSTS method is to start to "stump off" the pillars as soon as the room is completed by cutting across the rib as at A, in Fig. 38. The track should be laid so that cars can be placed con- veniently for loading out both the coal in the cut and also that mined when the stump is drawn back. In this method, the miner is always protected by solid coal and the losses are reduced to a minimum. Room No. 6 shows the pillar and the heading stumps completely removed; room No. 5, a pocket just starting in ; room No. 4, a pocket finished and a stump partly drawn back. Room No. 3 shows the pocket finished and work just starting on the stump ; room No. 2 shows the pocket being driven, followed by a second pocket, which FIG. 38. Method of splitting the pillars used in the Pocahontas Field. is only extended as far as a man can conveniently load the coal without a track turn, in order to avoid the necessity for frogs and switch points. Room No. 1 shows the pocket just starting. The width of this pocket and the thickness of the stump depend largely on the nature of the roof and the mine equip- ment. With poor roof, which falls unexpectedly or within a few hours after the removal of the coal, the thickness of the stump should be such that a miner can reach all the coal safely and easily without venturing too far beyond the rib line of the pocket. If the roof is good and does not fall soon after the removal of the stump the thickness of the small pillar may be increased and the number of track turns required per pillar may be reduced. MINING COSTS 77 In the application of mining machines to the robbing of pillars, the distance between the centers of pockets should be such that the thickness of stump left will form, under bad roof, one machine cut, or under good roof, two cuts of the machine. The more common practice where the roof falls soon after the extraction of the stump is to leave a small shell of coal to protect the miner from the gob and also prevent him from loading fine slate into the car of coal. This results in a loss of coal that can be avoided at an expense for timber, which, under ordinary circumstances, is less than the value of the coal. A practice which has been advocated and proved success- ful, is to place a row of props on the lower rib of the pocket, before the removal of the pillar stump has begun. When the next pocket to the outby is driven, it will be found that prac- tically the entire stump may be loaded out without any admix- ture of gob and a greater percentage of lump coal will be obtained. This precautionary row of timbers is especially desirable where machines are used on the pillars. The roof over a robbing line exceeding 2400 ft. in length sometimes begins to sag in the middle and renders the removal of the pillars in that immediate section difficult. The breakline should be kept as uniform as possible at all times. A method in practical every-day use, which is to be recommended, is as shown in Fig. 39. The engineers, as they take their monthly measurements, mark the pocket centers on the robbing rib of the room, and the foremen are required to drive their crosscuts on the line of a pocket as at A. The object in keeping the breakline uniform is to insure against pillars extending back into the gob and acting as a fulcrum, or the knife-edge of a scale beam, upon which the roof teeters. This almost invariably causes additional timber expense and sometimes losses of coal, both of which could have been avoided had the break line been kept uniform. The essential features of the Pocahontas Coal & Coke Co.'s plan of mining, shown in Fig. 35 are : Provisions for tonnage during the development period; provision for meeting the market demand ; large barrier pillars, insuring against squeezes 78 COAL MINING COSTS and rendering impossible the destruction of coal over an extended area ; four-entry system for all extensive main entries, using two as intakes and two as returns with breakthroughs between only at points where the cross entries turn off, render- ing unnecessary the building of expensive masonry brattices every 80 ft. and insuring the maximum quantity of air for ventilation at a minimum cost for brattices and ventilating power; cross entries with narrow chain pillars, permitting the rapid advance of the entry. In general all robbing must be done retreating with rooms driven after the entry is nearing completion, insuring against slate falls and rendering possible the extraction of all of the FIG. 39. Locating crosscuts so as not to interfere with lifts from pillar. coal in one operation, thus combining first development and robbing. The depth and number of rooms on an entry vary greatly at different mines, depending on local conditions of the seam; whether the haulage is by mule or gathering motor, whether the undercutting is performed by pick or machine, and not infrequently on the personal equation of the mine executive, for sometimes the manager of a plant will contend that he obtains the best results when he drives 25 rooms, 500 ft, deep to the entry, and another manager working on an adjoining property under identically the same physical conditions and with the same type of equipment, not 1000 ft. away, will say that he gets the best results when his rooms do not exceed 300 ft. in depth and when there are only 15 rooms to the MINING COSTS 79 entry. The better policy is to encourage individual initiative and allow freely such modifications in any plan of mining as may be desired, provided that the modified plan embodies all the principles of modern methods and sound mining practice. In the successful operation of any mine some general scheme of mining must be agreed upon, subscribed to by all parties in any way concerned with the matter, including the land owner, if the property is a leased one. Then no omissions in, additions to, or deviations from that plan of mining should be permitted without the written consent of the lessee and lessor. RECOVERY OF COAL IN MINES OF POCAHONTAS COAL & COKE Co. Plant Thick- ness of Seam in Feet Acres of Entry- mined Acres of Rooms Mined Acres - of Pillars Mined Total Acres Mined Total Tonnage Mined Tons Mined per Acre Theo- retical Tons per Acre Per- centage of Re- covery Propor- tion of Seam Re- jected 1 6.15 3.06 4.57 11.03 18.66 165,254 8,856 9,922 89.3 0.24 2 5.65 4.40 4.80 14.80 24.00 188,391 8,185 9,115 89.79 0.22 3 5.16 2.68 6.52 15.80 25.00 180,386 7,215 8,325 86.6 0.22 4 4.42 5.88 8.65 13.09 27.62 192,437 6,960 7,131 97.6 0.23 5 5.94 7.00 10.09 19.20 36.29 334,005 9,203 9,582 96.0 0.22 6 4.32 2.11 3.64 9.20 15.04 94,427 6,278 6,969 90.0 0.31 7 5.34 3.31 6.34 0.00 9.65 83,000 8,601 8,614 99.8 0.20 8 5.42 3.72 6.06 9.72 19.50 144,769 8,181 8,777 93.2 0.20 9 4.65 8.10 16.80 2.34 27.24 201,044 7,380 7,534 98.0 0.18 10 8.03 5.20 8.47 10.09 23.76 262,975 11,068 12,923 85.6 0.23 After the general plan of mining has been decided and operations begun, its success or failure will depend largely upon the degree of watchfulness exercised. The mine should be accurately surveyed and mapped at least once every 90 days. Frequent inspections should be made of the mine, minute attention being given to the conditions of the working faces and the robbing line. At least once a year the theoretical yield of the property should be balanced against the actual tonnage delivered at the tipple. Accurate and complete records should be kept of the number of acres of entries, and rooms driven and pillars drawn each year and of both the percentage of recovery per acre and the state of exhaustion of the prop- erty. That the above method of mining will yield the maximum recovery is indicated in the accompanying table, the figures in 80 COAL MINING COSTS which are typical of the results obtained by the Pocahontas Coal & Coke Co., which are probably unexcelled anywhere. In this connection it should also be noted that the percentages of recovery are based on the total seam, including the portion rejected. The lower percentages of recovery in the table result from the fact that in some instances, pillars were being robbed that had been standing for many years. In the mines of the United States Coal & Coke Co., at and near Gary, W. Va., where all the work has been opened in recent years, the average percentage of recovery per acre since the beginning of opera- tions, has been better than 95 per cent, and of the area mined, over one-third has been final mining. The cost of production of room and entry coal by this method is the same as in other methods of mining, while pillar coal is produced at less expense than is incurred in other methods. Most operating companies have a statement showing the revenue derived from operations per ton of coal mined based on the net receipts from operations divided by the tonnage. By placing a value, which could be closely approximated, on the recoverable coal lost, and adding it to the net receipts a figure could be obtained showing what revenue would have been derived from the operations had the coal been mined without unnecessary waste. Dividing this by the tonnage a figure could be obtained for the statement which would show the profit that would have been derived per ton produced if the coal in the seam had been worked by the most conservative methods. Statements of the above nature have a further value from a financial point of view for if it can be shown to a bonding concern that a property contains, let us say $500,000 worth of coal in the ground, 90 per cent or more of which will be mined, it is certain that a greater asset value will be placed on the property than would be credited to it if the engineers of the bonding house report that under the methods of mining pur- sued, only 50 per cent of the coal in the ground will be mined and the rest irretrievably lost. Connellsville Systems. The system of mining practiced by the H. C. Frick Coke Co. in the Connellsville region as described in a paper presented before the Engineers Society of Western MINING COSTS 81 Pennsylvania in 1916, is the application of shortwall mining machines to the extraction of rib coal. The two salient factors effecting this result were, first, the effort to reduce accidents and second, the desire to obtain an increased output of coal per man per day. It has long been realized that the more intense the super- vision of working places and workmen the less liability there is to accident. In order to obtain the desired supervision without making the cost prohibitive, it was seen that the time spent by mine officials in traveling from working place to working place must be reduced to the minimum and the time actually spent in working places increased to the maximum. To obtain this result the working places were concentrated gradually, and it was soon found that, under the old method of pick mining, a limit was quickly reached, and it was realized that to obtain the desired intensive supervision it was necessary to decrease the number of working places and workmen. This could only be accomplished by an increased production from each miner and a consequent reduction in the number of work- ing places without affecting the total output of the mine. On account of the conditions in the Connellsville region, where it is necessary to drive narrow headings, narrow rooms and have large room centers, it was found that machines in the narrow work would not accomplish the result since the bulk of the coal comes from rib extraction. The use of electrically driven mining machines and the blasting of coal on the very rib line itself requires a system of ventilation that will insure gob gas being found only on the return. Such a system of ventilation necessitates ample, reliable fan equipment, airways of sufficient size and number, a generous provision of upcast openings, wise coursing of the air current and the existence of numerous bleeders from every gob into a return airway. It also demands the elimination of danger from dust by keeping it sprinkled and removing it before any dangerous accumulations are found. In addition it necessitates the use of permissible explosives and these only in the hands of selected competent shotfirers. It has been proved that working places cannot be concen- trated to as great an extent by any system yet tried in the Connellsville region as by the use of the H. C. Frick Coke Co. 's 82 COAL MINING COSTS system of machine mining in rib coal. On account of the intense concentration of working places and the output that is obtained it is necessary that the haulage arrangements and equipment be perfected beyond anything that had previously been necessary, and the transportation of coal cannot well be handled except by the use of electric gathering locomotives. The general plan by simple modifications can be made to suit all conditions; depth of cover; presence or absence of drawslate ; nature of coal, and the nature of bottom and roof. This. is divided into what we know as maximum, medium and minimum plans. The maximum plan is applicable where thickness of over- lying cover does not exceed 125 ft. and where the coal is hard and the general physical condition of roof and bottom is good. The medium plan is applied where the cover does not exceed 250 ft. with the same physical conditions of coal and bottom and roof as obtain under the maximum plan. The minimum plan may be applied to coal underlying any depth or thickness of cover, and whether or not the coal is hard or soft and the physical condition of roof and bottom good or bad, provided, of course, that mining machines in any form can be used. The H. C. Frick Coke Co. has always worked its mines according to a projection, carefully prepared, for the field of coal to be worked before actual excavations have been started. In this plan of concentrated mining it has been found of great advantage to supplement these general projections with a schedule, prepared on a scale of 20 ft. to the inch, showing in detail the daily operations. It should be understood that in the shortwall plan of min- ing the development is made on the face and the butt of the coal. After it has been determined as to what plan is to be followed for a given tonnage, the mining section is projected and developed and a fracture line established. Let us first consider the minimum plan of extraction. The main haulage headings are driven on the face as are also the return airways while the producing headings are on the butt. Off these producing headings main face rooms are turned, generally on 112-ft. centers. From these main face rooms, butt rooms are driven on 25-ft. centers. As the main face rooms MINING COSTS 83 advance the necks of the butt rooms to be driven are excavated to a depth of three machine cuts. After this main face room has been advanced 50 ft. there are available two places for the machine to cut that will yield 40 tons, and when it advances to a point where the first crosscut is turned off, there are three places to cut in each main face room, yielding 60 tons. This main room may continue to the end of the section or to the end of the coal field, turning butts or producing headings off at projected distances. The main face rooms being driven on 112-ft. centers and 12 ft. wide leave a pillar 100 ft. in thickness between the rooms. This pillar is considered ample to support any thickness of cover with a floor or bottom under the coal seam of any nature that may be found in the Connellsville region. On this minimum plan of extraction, where main rooms are advanced sufficiently far to begin the extraction of main face room pillars, the butt rooms are advanced in succession so that each room is 50 ft. behind the one next preceding. This plan provides for a tonnage output from three working places two butt rooms advancing furnish 40 tons and one butt rib retreating provides an additional 40 tons, or a total of 80 tons of coal while retreating; and the main face room advancing is yielding 60 tons, or a total of 140 tons of coal from one main face room properly prepared and developed on this minimum plan of production. A sketch of this method is shown in Fig. 40. Along the same lines the medium plan will not yield any greater tonnage from the advancing main rooms, but on the retreat the butt rooms are so driven as to maintain each face 30 ft. behind that of the preceding room. This allows three butt rooms to advance at a time, producing 60 tons, and neces- sitates two butt ribs retreating at the same time, giving a production of 80 tons, or a total from the butt rooms of 140 tons. This with the production of 60 tons from the advancing main room totals 200 tons for each main room. This method is shown in Fig. 41. In the maximum plan the main face rooms advancing pro- duce 60 tons while the butt rooms are so driven that the face of one is 15 ft. behind the face of the preceding room, thus necessitating four advancing butt rooms and the simultaneous 84 COAL MINING COSTS . ung MINING COSTS 85 withdrawal of four butt ribs. The four advancing butt rooms will produce 80 tons while the four retreating butt rooms will produce 160 tons. The sum of these, together with the 60 tons produced by the advancing main room, gives a total tonnage of 300 for each main room. The maximum plan is shown in Fig. 42. The work is thoroughly systematized and the routine can be described as follows: After the miner has cleaned up his place and the day's run is completed the machine crew enters and cuts the place to a depth approximately 7 ft. The timber- men follow the machine crew, resetting any posts that it has been necessary for the machine men to remove. They post up any crossbars that have been notched in the coal over the machine cut, and generally put the place in good condition, following out a prescribed system of timbering. The timber- men are followed by the driller, who bores the holes for blast- ing with an electrically operated power drill. The driller is followed by the shotfirer, who charges the hole, tamps it, and after his own personal examination of conditions, explodes the charge, blasting down the coal ready for loading. After the coal has been blasted empty cars are placed by the gathering locomotives preparatory for the next day 's work, so that when the loader arrives at his working place in the morning it is in a safe condition and every facility has been given him to load a maximum tonnage. Especial pains are taken through the day to see that wagons are changed as soon as loaded, thereby eliminating all unnecessary loss of time and allowing the men to load a maximum tonnage in the minimum time. Actual results obtained regularly with miners loading under these conditions are 18 to 20 tons per man per shift ; the average of all the loaders behind shortwall mining machines in all mines of the company for the month of August, 1916, was approximately 19 tons per shift. At mines where there is a full equipment of mining machines the proportion of machine-coal amounts to from 80 to 95 per cent of the total output. The concentration of work that has been obtained by this method resulted in a decrease in the cost of transportation, ventilation, track work and drainage because of the smaller area in active operation. There is also a considerable saving 86 COAL MINING COSTS in the amount money invested in track and materials generally. Some further interesting data on this system appeared in a paper presented before the Coal Mining Institute of America from which the following has been excerpted. Preparations for the adoption of this system are made well in advance by subdividing the panels into blocks 90 ft. square FIG. 43. Plan showing how the Connellsville system is used at the Con- tinental Mine No. 2. as shown in Fig. 44. Double butt entries on 50-ft. centers, 10 ft. in width and 1200 ft. long are driven in parallel across the panel with cutthrough connections every 100 ft. Other and similar butt entries are driven 350 ft. apart, dividing the panels into blocks 350 by 1200 ft. and the chain pillars between the double entries into blocks 40 by 90 ft. The 350 by 1200-ft. blocks are then subdivided into blocks 90 ft. square by driving rooms 10 ft. wide and 350 ft, long MINING COSTS 87 at right angles to the butt entries, the rooms being connected by cutthroughs at intervals of 100 ft. In this manner a whole panel can be developed and prepared to a reasonable distance in advance for the work of concentration in quickly withdraw- ing the pillar coal. Fig. 45 illustrates the concentration method, showing a part of a section when in full operation as worked in the mines of the lower Connellsville district. This shows the coal in 90-ft.- square blocks and oblong pillars 15 ft. by 90 ft., also the entries, Regulator..,^ FIG. 44. Method of laying out mine in 90-ft. blocks used in the Connellsville region. rooms, crosscuts and cutthroughs. The section shown cross- hatched represents that portion from which the coal has all been withdrawn and the overlying strata have subsided, or the "gob" section. This section of coal is developed by driving on the right hand side a pair of butt entries, 10 ft. in width, 500 ft. in length, on 50 ft. centers and connected by cutthroughs every 100 ft. Rooms 10 ft. wide and 350 ft. long, on 100 ft. centers, are driven at right angles to the butt entries, the rooms being connected by a straight line of cutthroughs at distances of 100 ft. apart for ventilation. Thus, the room pillars are divided into 3y 2 blocks 90 ft. square, as shown on the plan in prepara- 88 COAL MINING COSTS tion for the final operations of driving the crosscuts and the withdrawal of the pillar coal. The selection of the place to commence on the pillar work is important and must be determined by the persons directly interested largely from the local conditions such as the inclinations or pitch of the coal bed, convenience of transpor- tation or haulage, the size, area or extent of the panel or section available for operations or the required daily tonnage. Preparations are now completed for the essential part of the concentration work. It will be noticed on the map that the pillar work at the end of the last room, which was in the FIG. 45. The concentration method used in the Lower Connellsville region. upper left-hand corner of the plan, has been started. Each room having been divided into 3y 2 blocks, 90 ft. square, almost three and a half blocks from the last, or No. 5 room have been worked out; nearly one and a half blocks from No. 4 room; one fourth of a block from No. 3 room and crosscuts started in No. 2 room. This makes the angle of the gob line about 45 deg. with the butt entry. The plan shows that the pillar work of each room is 75 ft. in advance of the pillar work or gob fall of the room follow- ing. The room pillars are kept in this position for the purpose of breaking the roof falls in the advancing pillars and offering more resistance to thrust caused by the breaking of the strata ; also for the purpose of affording better protection against crushing in the pillars of the room following. MINING COSTS 89 The system so far as this pair of butt entries is concerned is now in full operation. The pillar work or gob line, however, may be, and often is, extended across several pairs of butt entries (see Fig. 45), leaving an offset of about 75 ft. as a breaker at each pair. In this manner the gob line may be extended clear across the panel at an angle of 45 deg. or even at 35 deg. The work on this plan was commenced at the top of the last room in the block, which was in the upper left-hand corner of the map. New crosscuts were started at intervals of two days, thus making each crosscut two days' work, or about 12 ft. in advance of the one following, until the crosscut first started has been driven through the 90-ft. block, for which 15 days were allowed at 6 ft. per day. The crosscuts being 10 ft. in width and turned off the rooms at distances of 25 ft. between centers makes the pillars, for the final operation, 15 ft. in width and 90 ft. in length. These pillars are divided and subdivided by lines drawn across and lengthwise of the pillar. The cross division lines divide the pillar into three sections of 15 X 30 ft. each the amount of coal to be taken out in one fall by three or four men working two days which makes three falls to each pillar in six days' work, as shown. Three of these pillars in each room are being worked at the same time and are started at intervals of two days, thus placing each pillar two days* work, or 30 ft. in advance of the one following. The proper time to start the first crosscut at the top of the next room in order that the 75-ft. offset may be maintained can be ascertained as follows : It takes 15 days to complete the first crosscut at the top of the last room, six days to withdraw the pillar, two days to finish the next pillar and two more days to finish the next one, making a total of 25 days to com- plete the withdrawal of the three pillars or cover a horizontal distance of 75 ft. Therefore, by starting the first crosscut in the next room 10 days later, the offset of 75 ft. will be main- tained as shown on the plan. By continuing the work on this schedule, leaving intervals of two days between the starting time of each crosscut and 10-day intervals between the starting time of the crosscut in 90 COAL MINING COSTS the next room, we shall have three room pillars of 90-ft. thick- ness retreating in a diagonal line on each pair of butt entries and three of the crosscut pillars in each room pillar or 9 working places when in full operation. In estimating the amount of daily output from this section, there are in operation 15 crosscuts and 9 pillars. Allowing one man in each crosscut and three men to each pillar makes a total of 42 loaders. The coal being undercut to a depth of 6 ft. by the mining machines, the tonnage of each crosscut 6 ft. undercut, 10 ft, in width and 7 ft. in height of seam, will be 6X10X7=420 cu. ft. Allowing 80 Ib. to each cubic foot and 2000 Ib. to the ton, we have, 420X80 ^000- = 16.8 tons, and 15 crosscuts equals 16.8X15 or 252 tons. Assuming a two- ton car, this gives us 252/2=126 loaded cars. From the 9 pillars having 6-ft. depth of undercut, 30 ft. in length of pillar and 7 ft. in thickness of seam, there will be 6X30X7X80 2000 = 50.4 tons, and for 9 pillars, 50.4X9=453.6 tons, or 226.8, 2-ton loaded cars, which makes the total number of tons from the whole section 252+453.6 = 705.6 tons, or 352.8 loaded cars from 42 loaders, being an average of 8.4 cars for each loader. This average may seem rather high until we take into consideration the facilities afforded by the concentration method and the preparations made to enable the miner or loader to attain a high efficiency. During the previous night all working places are under- cut to a depth of 6 ft. by mining machines. The shot holes are drilled with power drills by men employed especially for that purpose and are charged, tamped and fired by shotfirers using permissible explosives, clay for tamping and electric batteries for firing. By this system each miner, when he arrives at his work- ing place, has about eight or nine carloads of loose coal, which he can at once begin to load, provided no unusual difficulties MINING COSTS 91 arise to prevent him. He is kept constantly supplied with empty cars by the driver, who can also attain a high efficiency by reason of having the loaders within a comparatively small area and only a short distance from the side track. The tracks are kept in good condition and laid with steel rails, even in the miner's places. The trackmen, timbermen and laborers are also enabled to do more effective work, as there is no lost time in traveling long distances from place to place. For the same reason much better supervision can be given by mine foreman, assistant mine foreman and firebosses to the workmen and working places. They can make frequent visits and keep in close touch with the workmen and other matters influencing the success of the operations such as the machinery, transportation or haulage, ventilation, trackmen and laborers, timbering and timbermen, miners and drivers and see that defects interfer- ing with work or causing delays are remedied immediately. Comparative cost of mining different thickness of coal. An interesting study of the determination of the minimum thickness of anthracite coal that can be economically mined was presented in a paper read before the Engineering Society of Northeastern Pennsylvania in 1914. In deciding which beds are and which are not minable, we face, at once, the question of profitable operation, and it may be conceded that other things being equal, beds which are 6 ft. and over in thickness are more cheaply mined than those which are thinner. If we eliminate all variations other than cutting and loading in making our calculations, the relative cost of mining for varying thicknesses is a matter of simple calculation. Let a = allowance for rock in cents per inch per yard; h = normal required thickness, in inches, on which allowance is based; x = thickness of coal, in inches, as measured for allowance; x l =uei thickness of coal, in inches; /= capacity of mine car, cubic feet; iw= width of chamber in feet; s = thickness, in inches, which will give one mine car per yard of chamber. Assume loose coal occupies 1| times the volume of an equal weight of solid; c = cents per car allowance; m = mining price per car, no allowance. 92 COAL MINING COSTS Then (h x) a = allowance per yard of chamber; /. s = - = number cars per yard of chamber width; 3ws (hx)a hasasx From this, for any particular conditions, the cost for each thickness may be calculated, and a curve constructed show- ing the relation between cost and thickness, as shown in the diagram, Fig. 46. Thickness of Coal 2345 O 10 EO 30 40 50 60 70 80 90 100 110 120 130 140 Percent Increase m Output FIG. 46. Relation of thickness of seam and output to cost of coal. Unfortunately, all the costs which vary with the thickness of bed are not susceptible to calculation, but in general the inside costs increase rapidly with the mining of thinner coal, and the diagram showing the variation of cost with thickness is believed to represent fairly the average conditions in the anthracite fields. It was constructed by plotting a large num- ber of actual costs and then drawing the average curve. MINING COSTS 93 We find ourselves facing the dilemma whether we shall mine the coal which is profitable in itself or a mixture of profit- able and unprofitable coal which, through the preponderance of the latter, will result in a profitable output. Our first thought would naturally be that only the coal which is actually profitable should be mined, but when we make a more com- plete inquiry we find another condition in the relation of out- put to cost. As but approximately one-third of the inside cost is expended in actual cutting and loading, and as the greater part of the outside cost is but slightly dependent upon output, the cost per ton will be found to vary greatly with the coal production, even with a fixed unit cost for cutting and load- ing. How great this variation may be is indicated on the diagram, and it is apparent that a large output from beds which show a relatively high mining cost may be actually more profitable than a much smaller output exclusively from the larger and cheaper beds. Hence, it is apparent that, even from the standpoint of immediate profit, it may be advisable to mine the thinner beds with the thicker, and considering the ultimate yield of a prop- erty, there can be no question as to the advisability of such a course. The actual minimum minable thickness being dependent upon so many conditions is not susceptible to gen- eral determination and should be studied for each individual case. Conveyor system. The method of operation by conveyors herein described has been in use in a number of collieries working some comparatively thin measures in one of the coal fields in Scotland, and has proved its success and applicability through a period of at least 10 yrs. In some respects the method adopted was unusual in that while conveyors are in operation by themselves no coal-cutters are employed, under- cutting being done by hand. Present-day practice always con- siders conveyor work an adjunct to machine mining; but here is a case of conveyor practice by itself. Another distinctive feature is that the opening and development of the mine for additional faces, as well as running the usual longwall faces with the conveyor, are being done with a conveyor wall. The coal seam on the average is 3 ft. 9 in. thick, but owing to the presence of stone bands is rather broken up. This 94 COAL MINING COSTS means that after removing 31 in. of coal there remains about 14 in. of stone to be disposed of. The thick stone parting near the bottom of the coal is a yellow-white sandstone that breaks in flat squares, eminently suited for building the road pillars in longwall working. In working under the old system, "docks," or "deeps," were driven direct to the dip, the inclination being 8 deg. These "dooks" were driven in the solid coal, with pillars turned off every 150 ft. on the dip and 60 ft. on the level course. Every 300 ft. levels were broken off right and left, and a long- wall face commenced two pillar lengths from the center deep, in a fan-shaped fashion, which as it opened out gradually edged uphill until its upper corner worked along the waste of the level above, and the face stretched from one level to another. In driving the "deeps" three roads were allowed one for haulage and intake air current while the two on either side where needed for return air. In order to operate the long- wall faces at low cost, slants were driven uphill from the lowest level, and from the slant several parallels to the main bottom level turnecT into the coal face, these being a distance of 40 ft. apart. During the ordinary longwall methods of working the fol- lowing men were employed in the section: Miners, 14; brushers, 7; trackmen, 2; timbermen, 3. The output was 45 tons. The miners trammed their own coal to the main level. In the layout of the workings for the conveyor there was little actual difference in the direction of development. The faces still extended across the strike; but in place of the three parallel deeps a longwall face was laid out at an angle between the dip and strike, so that the left-hand end was the most advanced, which allowed the coal and water to gravitate to one point. The driving and formation of pillars were thus dispensed with, and the longwall system of extraction became a developing system as well. The conveyor in use is of the shaking type, and has been adapted from continental practice. The height from the floor to the edge of the pan is only 9 in. This means that the work- man is required to raise the coal only slightly over this height MINING COSTS 95 instead of a former 29 in., which accounts for more work with decreased effort. The width is only 18 in., and this allows of the distance from coal face to waste line of props being kept at a minimum. The principal dimensions of the driving engine are as fol- lows : Horsepower of engine, 12 ; air consumption at 60 strokes per minute and 60 Ib. pressure, 18 cu.ft. ; stroke of engine, 5 in. ; diameter of cylinder, 7 in. ; weight of engine, 572 Ib. The engine, driven by compressed air, is a simple, plain, air cylinder, with broad bed-plate, which may be bolted to planks, which are in turn wedged against the roof by timber. The total width is 18 in., and the length 24 in. Connection to the conveyer is through a lever action and rigid attach- ment, the cylinder being placed at right angles to the line of the conveyor. The pack walls on the top side of the driving road are uniformly built in line, 2 ft. or so back from the edge, this space being utilized for the engine. Air is furnished from a power-driven air compressor that stands in a small recess in the side of the road. The principal dimensions are as follows : Horsepower, 15 ; r.p.m., 960 ; amperes, 16.5 ; voltage, 500 ; cycles, 50; air pipe, l 1 /^ i n - ; cylinders, 4; strokes per minute, air cylinders, 480; pressure, 70 Ib. ; space occupied, 5 ft. 6 in. by 3 ft. 5 in. The air is passed through an 18-ft. hose to the air cylinder of the conveyor. The rate of advance changed from 160 ft. in six months under the old system to 270 ft. over the same period under the conveyor system. This is not remarkable compared to machine working advances, but it represents a considerably increased rate of extraction for handwork. The operation of each face in the colliery is regular and at the same rate, the only determining factor being the number of men employed. Shifting of the conveyor takes place every second night, so as to get under the fresh rock as soon as possible ; but this is governed by the rate of cutting and loading. The air engine is moved at the same time as the conveyor, but the compressor only when the length of hose is reached. The operation of shifting is accomplished by a night force of eight, who also shift the compressor when necessary and set all timber required. 96 COAL MINING COSTS A comparison of the costs of operation and performance accomplished by the old and the new systems has worked out much as follows: Hand Operation Conveyor Operation Tons per man 3 2 4 6 Length of face 300 ft 300 ft Tons per section 45 55 Length of face, per man Number of miners . . 43 ft. 14 40-50 ft. 12 Number of deadwork men 12 2 Number of conveyor men. . . 1 Shifting conveyor . . . (Average 4 Number of roads to maintain per night) 7 2 Tons per road Time stripping Distance bottom level in advance . . / 6 . 4 Main level I Bottom ' ' 9 hr. 49.00 6.00 9hr. 40ft. COSTS Cutting Shooting / 72 72 Loading j Brushing 41 (Done by squad Maintaining roads 08 shifting conveyor) 08 Tramming (Included with (Included for low- Shifting conveyor shooting and loading) er level in ton- nage rate) 15 Operating conveyor. 03 $1.21 $0.98 It will be seen that the saving in cost finally effected is entirely due to the elimination of roof troubles, which the con- veyor system made possible. There are now installed 10 con- veyor faces at this operation. Mining machinery. Man-power is about the most expensive energy purchasable. We pay a laborer, say $2 for 9 hr. work. This man is capable of exerting continuously about one-eighth MINING COSTS 97 of a horsepower. In other words, we have secured 1% hp.-hr. for $2, or we pay about $1.78 for man-power per horsepower- hour. In marked contrast to this high cost of energy is the cost of current delivered to the motor of a mining machine which should not exceed 2 to 2%c. per horsepower-hour. It is the realization of this discrepancy between the cost of power developed by man and that developed by a steam engine, for instance, that is driving the coal industry to employ machinery wherever such employment is possible. Furthermore, it is frequently the case as in undercutting, for example that the machine does the work better; that is, it cuts deeper and affords less resultant fine coal than when mining is done by hand. It is probable that most operations that may be performed by machinery require a greater expenditure of power than would the same operations performed by hand; nevertheless it has become almost axiomatic that it is economical to supplant manual power by machinery wherever possible. Consequently inventors are continually striving to perfect mining machines, and other power-driven devices that will do away with the employment of muscular energy. Cutting machines. Mining machines now produce about 65 per cent of the nation's coal output as compared with 35 per cent in 1907. The economies over hand mining may be summed up as follows: First, the actual cost of mining is lower, due to the fact that the greater cutting capacity of the machine makes possible a greater output with a given labor cost; second, the quality of the product is superior, due to the deeper and more uniform undercut of the machine, which increases the percentage of lump coal 10 to 30 per cent over hand mining methods; third, the mine may be more rapidly developed due to the much greater speed with which entries can be driven with machines insuring a rapid return on the capital invested; fourth, the ability to mine seams in which the height of the coal, or the character of the roof, has prevented mining by hand, on a commercial basis. During the period from 1891 to 1914, the average tons of coal mined per mining machine, per year, in the United States was about 13,700. In 1913 there were 2208 shortwall mining 98 COAL MINING COSTS machines in use and in 1914 there were 3024, an increase of 32 per cent in one year. In 1914 there were 6859 breast machines in use, which is an increase of 100 per cent over the year 1904. One of the most noticeable increases in coal production mined by machines has been in Kentucky, where, in 1912, 66.4 per cent of the coal was mined by machines, while in 1914 the production mined by this method was 77.2 per cent. In the accompanying table the unit of efficiency is given for the total production of bituminous coal in the United States for the years 1891 to 1915. There is also a column show- ing what percentage of the coal was machine-mined. By machine-mined coal is meant all coal won by the use of any of FIG. 47. Curves showing per capita production and per cent of machine-mined coal. the following types of machines: Punchers, radially mounted punchers and chain breast, shortwall and longwall cutters. The figures given in the table were taken from the reports of the Bureau of Mines on the yearly production of coal. From these statistics it is at once apparent that the increase in the number of tons mined per day per man has corresponded closely with the increase in the percentage of machine-mined coal. In order to show this more clearly the two curves shown in Fig. 47 have been drawn. The upper curve shows the tons per day per man and the lower the percentage of machine- mined coal. From these two curves, unless some radical changes are made, it can be estimated that about the year 1928 all coal will be machine-mined and that the efficiency of the miners will have increased to about 4.9 tons per man per day. MINING COSTS 99 PROPORTION OF MACHINE MINED COAL IN THE UNITED STATES AND PRO- DUCTION PER MAN Year Average No. of Men No. Days Worked Total Tonnage Tons per Man per Day Per Cent Machine- Mined 1891 205,803 223 117,901,238 2.57 5.26 1896 244,171 192 137,640,276 2.94 11.86 1897 247,817 196 147,617,519 3.04 15.35 1898 255,717 211 166,593,623 3.09 19.46 1899 271,027 234 193,323,187 3.05 22.74 1900 304,375 234 212,316,112 2.98 24.86 1901 340,235 225 225,829,149 2.94 25.61 1902 370,056 230 260,216,844 3.06 26.75 1903 415,777 225 282,749,348 3.02 27.58 1904 437,832 202 278,659,689 3.15 28.21 1905 460,629 211 315,062,785 3.24 32.82 1906 478,425 213 342,874,867 3.36 34.66 1907 513,258 234 394,759,112 3.29 35.11 1908 516,264 193 332,573,944 3.34 37.04 1909 379,744,257 37.52 1910 555,533 217 417,111,142 3.46 41.72 1911 549,779 211 405,907,059 3.50 43.89 1912 548,632 223 450,104,982 3.68 46.80 1913 571,882 232 478,435,297 3.61 50.07 1914 583,506 195 422,703,970 3.71 51.70 1915 557,456 203 442,624,426 3.91 55.00 Mining machines and the necessary equipment for success- fully operating them at the average colliery cost a large amount of money. If this is injudiciously spent in equipping a plant for cutting coal with machines, overhead charges for interest, maintenance, depreciation and taxes will be correspondingly heavy. Conditions may justify the expenditure in order to properly recover certain coals and at the same time safeguard life and property; but such moneys should be carefully and wisely expended and then only after exhaustive analysis of conditions surrounding the proposed operation. The fact that one's neighbor mines his coal with machines is not sufficient reason for one to so equip his own property. Usually each mine, and especially each coal, has its peculiarities that deserve careful consideration. Many pointers and suggestions that are worthy of serious deliberations may be had from the man at 100 COAL MINING COSTS the face. Such suggestions only cost when ignored or neg- lected. It is fair to assume that all up-to-date companies maintain accounting systems that are a criterion by which they may determine approximately the relative costs of pick-mined and machine-mined coal. However, there are many angles that afford viewpoints not generally considered in this connection. Interest on investment, maintenance, depreciation and taxes on all extra equipment over that necessary for the suc- cessful operation of the property as a pick mine should properly be charged to machine-mined coal. In this list should be included all extra housing, boilers, boiler settings, boiler fittings and accessories such as feed pumps, steam headers, etc., in addition to generating units, settings, switchboards and acces- sories, transmission lines and machines. Further, also, under the head of supplies should be included and charged to machine-mined coal all extra repairs, fuel, water, oil, tools and office supplies over and above that necessary for the suc- cessful operation of the property as a pick mine; and under the head of labor, should be included and charged to machine- mined coal all extra expenditures for electricians, wiremen, firemen, oilers, drivers, tracklayers, bit sharpeners or black- smiths and clerical force over and above that necessary for the successful operation of the property as a pick mine. Men are quite frequently required to timber after machines, clean slate and refuse at switches and turns on account of the additional space required for machines to turn; extra drivers are also frequently necessary Hue to the fact that they are required to get sharp bits to the machines and dull bits to the shop, must occasionally await the moving of a machine thereby losing time and in some mines they must drive further for their loads or past one extra place out of every three, due to the fact that three working places are allowed each two loaders. Extra track layers are frequently required for the same reason, viz., that they have more track to keep up and over a larger territory due to the fact that three places are allowed two loaders. Bit sharpener or extra blacksmith should properly be charged to machine-mined coal where the machine men are not charged for smithing. Purchasing agents and bookkeepers spend considerable time MINING COSTS in ordering machine supplies, checking freight bills and keep- ing track of supplies. Delays and shutdowns due to trouble with boilers or generating units should properly be charged to machine-mined coal where machines are responsible for such trouble. As an example, there are sometimes delays in both hoisting and haulage due to the generating plants being over- loaded by reason of having been required to furnish power to the machines. It is quite possible with the advice of the machine makers to buy equipment that will suit the underground conditions at the coal face, but the organizing of all the factors that make for successful operation of a mine to the new conditions created by the advent of the machine is a subject conveniently forgot- ten by the seller of the apparatus, and often not considered by the operator. To buy equipment without looking into this question is much like purchasing an automobile for which gasoline cannot be readily procured. Consequently, organiza- tion and reorganization of the mine are the most important factors in success. Consideration of the following table will serve to more clearly illustrate this point. CONTRAST OF HAND- AND MACHINE-MINING CONDITIONS Under Hand Conditions Under Machine Conditions Tons 100 360 Number of roads to be kept open 18 18 Tons per road 6 20 Men in section ... 40 59 Timber to handle, single pieces 3000 12 108 Cars of coal 100 360 Rate of advance, inches . . . 12 42 It is evident at once that the greatest feature is the traffic increase. Instead of 200 cars a day to deal with there are now 720, besides an additional number at night. Instead of 3000 pieces of timber to take in there are now 12,000 to supply. Other supplies have increased due to the use of the machine, and with the increase in traffic, ventilation of the mine work- ings has to be maintained at a higher state of efficiency. The finest machines may be worthless if the system of back- -? "\ - 1 V ? a *> o COAL MINING COSTS ing them up fails. Ordinarily in all coal sections the amount of work done is limited or controlled by one single factor. This may be the amount of cars provided, the size of the sec- tion or the capacity of the outside haul. In machine mining only one thing should control the section, and that is the ton- nage produced at the face each night. Everything should be subordinated and coordinated thereto. Before the men leave their places at the face, the coal should all be squared up properly so that the machine can get to work promptly. If there is any coal left behind not cleared up, broken down but not loaded, or ' ' noses " overhanging, a man should be sent round in advance of the machine to see that all these obstructions are cleared away. This extra cost is more than offset by the gain made in the time of the machine, as well as by the elimination of the risk of the possible loss next day. The tracks by which the machine travels from place to place should be so arranged that the time lost in travel is reduced to a minimum. In a thick coal, where the machine cuts a relatively high tonnage in each place, there is more coal to set against this waste but with thin coal this lost time runs up alarmingly, because the amount of coal in each place is small. It is obvious that all machines are doing their best work where the going is continuous. Idle time and idle men, or those employed on unproductive work, mean a loss. The traveling from place to place results in the loss of a certain amount of productive cutting time, and it should therefore be cut to the minimum. Tracks and curves should be easy and well placed, so as to facilitate traveling, and trouble in this connection should never occur twice running at the same spot. Nightmen should be at work on the bad piece of track that same evening. Machine supplies should be kept handy to the face. If the section is a long one, supplies should be kept at several points. Tool chests with proper keys should be provided, otherwise the oil is often found to have disappeared, together with necessary hammers, keys and similar tools. Machinemen should be capable of making reasonable repairs themselves, and the mate- rials for doing so should be kept on the ground. MINING COSTS 10$ There should be no hard and fast rule that only electricians are to repair machines, neither should every handy man be allowed to try his hand at machine troubles. It is a good thing to have spare machines. Each of the machines underground should be taken to the surface at regular intervals for over- hauling. Hardly any class of machinery, unless it is the pumps, receives worse usage than does the coal cutter, and in the dark and poor light underground defects may develop that will never be noticed until it is too late to make the proper repair. All machines should be operated steadily on the surface for a number of hours and thoroughly oiled and tuned up before being returned within the mine. Machinemen should be at their machines an hour or so before the time of starting work, in order that each cutter may be overhauled and oiled, the bits changed, and all such details given the proper attention. Spare bits should always be on hand and any that have been removed should be sharp- ened in the morning, to be sent down again at night. The proper place for all bits except those in stock in the warehouse is in the mine and not in the blacksmith shop. One smith should sharpen all these tools the first thing in the morning, regardless of any other work. It should never be necessary to stop the machine to hunt for cutters or telephone to the surface at two in the morning. When electric current is used, the ratio of power given out by the cutter motor to the power required to drive the gen- erator may be taken roughly as 70 per cent, while with com- pressed air the ratio of power given out by the machine to power required to drive the compressor may be taken as in the neighborhood of 35 per cent. In other words the steam consumption of a compressed-air plant for coal-cutting is about double that of an electric plant. The figure 70 per cent taken for the electric cutter will not differ much whether the installation is well or badly designed, but in a poorly planned and badly maintained compressed-air plant, a considerably lower efficiency will be obtained than the 35 per cent taken as representative of a moderately good instal- lation. Installation and operating costs. The cost of a 5-machine plant may be summarized as follows, figures as of 1905 : 204 COAL MINING COSTS Power plant, including boilers, air compressor, air receiver, feed pump and feed-water heater $ 5,181 .00 Mining machines, including all accessories, and freight 4,125.00 Installation of power plant, including freight, com- pressor foundation, boiler settings, wooden boiler and engine house, water tank, fittings, piping, labor, etc 2,260.00 Pipe lines above ground and in the mine, with all fittings, labor and freight 2,484 . 00 Total cost $14,050. 00 The expense of maintaining and operating this plant may be approximately estimated as shown below figures as of 1905 : Interest (6 per cent) on investment, depreciation (10 per cent), repairs and renewals on machines and power plant and extensions of pipe lines .... $ 2,250 . 00 Fuel, 6 tons per day of one 8-hr, shift, at 50c. per ton, and oil and waste, 50c. per day; per year of 200 working days 750.00 One engineer at $75 per month; one machine boss and pipeman at $75; one blacksmith to sharpen picks at $60 2,520.00 Total maintenance $ 5,520.00 The above figures are considered the maximum, so that in actual practice, the cost of maintenance of the plant will prob- ably be lower. If we assume the pick rate at this mine to be 60c. per ton, and the differential rate for the machine runner to be one- fifth of the hand rate, then the machine rate will be 12c., and the loading rate 30c. Adding 3c. for blasting makes 45c., leav- ing a margin of 15c. for profit and payment on the plant. With an output of 700 tons per day and considering 200 days in the year the annual gross saving will be $21,000. If we now deduct the total cost of plant, including maintenance, from this saving, we have, $21,000 minus $19,570 which leaves a net profit of $1430 at the end of the first year. Cost of machine mining. It is practically impossible to compare the actual cost of mining with the various types of cutting machines. The machine that would show a consider- MINING COSTS 105 able saving at one mine might prove inefficient at some opera- tion. However, when it comes to comparing machine mining with hand mining, there is no difficulty. The most important point to consider is the size of the differential favoring machine mining over hand mining. In many districts this differential, or margin, amounts to about 15c. (in 1910) and it is out of this differential that the operator makes his profit and pays for the plant. At a mine producing 1000 tons per day and having a 15c. margin in favor of machine mining, the gross saving would be $150 a day, or $30,000 per year of 200 days. In such a case, the company can maintain its output with 20 per cent fewer men than are required when hand mining is employed. Herewith is a statement showing the cost of machine min- ing with longwall machines at an operation where, owing to the tough and " woody" nature of the coal, which necessitated paying the miners excessive "allowances," the average cost of mining with picks was between 60c, and 63c. per gross ton, figures as of 1910: COAL CUT BY ELECTRICITY AND LOADED BY DAY LABORERS Hours Rate Amount Cost per Ton Cutting 530 531 255 260 222 2716 3067 270 30c. 20c. 25c. 20c. 25c. 20c. 20c. 30c. $ 159.00 106.20 63.75 52.00 55.50 543 . 20 613 . 40 81.00 $0.0393 0.0262 0.0158 0.0128 0.0137 0.1343 0.1516 0.0200 Scraping Trackmen Trackmen Shooting Slate Loading Foremen Total labor $1674.05 $136.01 36.00 17.00 5.00 24.04 $0.4137 $0.0336 0.0090 0.0042 0.0012 0.0060 SuDDlies Depreciation on machines Interest 6 per cent on $3400 Repairs and maintenance (est Power [mated) 4045 tons, 19 cwt., at total cost of $1892.10 $0.4677 106 COAL MINING COSTS The above cannot be taken as a typical or average state- ment, as the conditions at the operation referred to are much less favorable to pick mining than at most other operations in this field. The pick-mining rate along New River is 50c. per gross ton, and at most of the operations on Piney creek and Loup creek it is 40c. per gross ton, as compared with 69c. in central Pennsylvania, so that the economy by the use of machines is much less than would appear from the figures above quoted. The following are detailed figures of the mining cost at two operations, one where all of the coal is cut by electric chain undercutters, and the other where all of the coal is cut by air-driven punching machines. The pick-mining rate for the year when these statements were compiled was 62c. per gross ton. ELECTRIC BREAST MACHINES, OUTPUT FOR ONE YEAR 321,808 GROSS TONS Per Ton Labor ; $0.4134 Material 0.0259 Insurance and taxes . 0009 Depreciation 0.0091 Interest charges 0.0025 $0.4518 COMPRESSED-AIR PUNCHING MACHINES, OUTPUT FOR YEAR 99,207 GROSS TONS Per Ton Labor $0.4810 Material 0.0211 Insurance and taxes . 0095 Depreciation 0. 0132 Interest charges . 0036 $0.5284 These two examples cannot be taken as a general average, because in the instance quoted where chain machines are used, the mining conditions are rather exceptionally favorable. In the other case, where punching machines are used, the con- ditions are about the same as the general average in that field. The tonnage produced per machine of a given feed varies according to the thickness of the coal, its hardness, the width of the working places and the length of the transfers. MINING COSTS 107 The West Kentucky Coal Co. operating nine mines in the western part of that state with seams ranging from 4 ft. 7 in. to 8 ft. 6 in. thick gets about 150 tons per machine-shift. This is an average of nine mines and includes machines used on development work in areas practically worked out where the output of the machines is naturally limited. In one seam rang- ing from 5 to 8y 2 ft. thick the production is 200 tons per machine-shift. The record cut for all the western Kentucky field up to 1917 was 300 lineal feet of face in 9 hr. At the mines of the W. G. Duncan Coal Co. in this same field where five machines are in use the average production is nearly 250 tons per machine shift. At these mines the nature of the coal is such that they are able to use a 25-in. feed over a 61/^-ft. cutter bar or a 21-in. feed over a 7%-ft. bar; they are also getting a good tonnage per bit sharpened, all of which factors make a large production per machine possible. The maintenance cost is probably the most important item in connection with the mining machine. The successful and economical operation of a mining machine, like any other piece of machinery, depends largely on the human element. By using care in selecting hostlers who will later become run- ners, an efficient machine organization can be built up. There is an instance where one man operated a machine continuously for four years without calling on the machine boss except occasionally for some small repair part to replace one actually worn out. Unfortunately, men of this type are not numerous, and for the average runner some incentive is necessary to sufficiently interest him in getting the best results from his machine. At the mines of the West Kentucky Coal Co. a bonus system is in effect as follows : A general supply stock is kept at each division and all supplies as purchased are charged to this sup- ply account. One machine boss has charge of the machines and other electrical equipment in use at each division, with one or more helpers as conditions may require. All supplies are issued by either the machine boss or his helper and charged to the mining-machine account of the mine where used. The average maintenance cost for the year 1915 was taken as a standard. The maintenance cost per ton mined for the year, includ- 108 COAL MINING COSTS ing the month for which the bonus is to be figured, is sub- tracted from the standard described above. This difference, multiplied by the tonnage of the month, is divided equally between the company and the machine and repairmen. It can readily be seen that in this way the men can make a gradually increasing bonus, and this they have succeeded in doing. At the same time it is impossible by skillful manipu- lation of supplies on the part of machine and repairmen to show a large premium in one month followed by a high main- tenance cost and no premium the succeeding month, which could be arranged were each month figured separately. As an example, to show how this premium system works out: A mine having a machine maintenance-cost standard of 2 1 / 4c. per ton, determined as above, produces in a given month 20,000 tons. The cost per ton for the year, including the month in question, is 1^2^ making a net saving on the tonnage of the month of $150, which is divided as follows : $75 to the company and $75 among five machine runners, five hostlers and one repairman, in proportion to their earnings for that month. For the most part the men engaged in the operation of the machines have taken unusual interest in reducing maintenance costs, and the men at the different mines rival one another in the attempt to establish the best record. The cost of supplies for the year 1915 was 0.96c. per ton. This includes bits, bit boxes, cables and all supplies used in the operation of machines, but does not include depreciation. In this connection it is worth notice that the first machine purchased has been in continuous service for 11 yr. and is still in as good operative condition as one just out of the factory. From studying the results of several machine installations, particularly those of the United States Coal & Coke Co., and the Clinchfield Coal & Land Corporation, it was found that deep undercutting is a decided advantage, and where con- sistent with the mining conditions, nature of coal, etc., should be recommended since it reduces (1) first cost of machine installation, (2) powder consumption, (3) cutting cost per ton. Moreover, where a certain output is expected, deep undercut- ting should reduce the territory under development. This is an important feature when the maintenance of roads, ventila- tion, and supply of timber are taken into consideration. Fur- MINING COSTS 109 thermore, it should increase the percentage of lump and give the loader a more definite or dependable quantity of coal down, thus increasing his efficiency, and thereby raising that of the entire plant. Comparative cost of alternating and direct current for machines. In a mine where alternating-current machines are to be utilized, large tonnage should be developed in as small an area as possible in order to secure the most economical use of the machines. This is shown in the accompanying diagram of a typical room-and-pillar working, Fig. 48. High Tension Unes(Cable (3WRC.Wires) FIG. 48. Typical room-and-pillar development using alternating- current machines. In this mine the rooms available for the mining machines are possibly far in excess of what the mining machine is capable of cutting. In territory distributed as in this example one mining machine should cut at least 7 rooms in a working shift or 14 rooms on a double shift. This would give approximately 300 lin.ft. of coal per shift, depending, of course, upon cir- cumstances, the above example being under average working conditions. 110 COAL MINING COSTS The 'power cost of operating a single alternating-current mining machine is comparatively low, as with the arrange- ment laid out in the diagram one mining machine would not use over 2000 to 3000 kw.-hr. per month working single shift. This small consumption of power is principally due to the mining machine having sufficient voltage at all times, and, therefore, working at its highest efficiency. One mining machine in 4-ft. coal and with a 7%-ft. cutter- bar mining say, 300 tons to the shift, or 6000 tons a month, will have a power expenditure of less than y 2 kw.-hr. to the ton. This would be less than Ic. per ton on the average rate of central-station contracts. A comparison of maintenance between the alternating- and direct-current machines shows favorably for the former. With direct-current power in use, especially with a mine that is supplied from some isolated power house on the outside of the mine and a considerable distance from the workings, the voltage is usually low; consequently, much armature trouble is experienced in the motor on the cutting machine. With an alternating-current mining machine served from central-station power there is an assurance of good voltage, as at no place need the machine be at a greater distance than 1500 ft. from the distributing transformers. This distance in any mine will give the mining machine more territory than it is possible for it to work. Alternating-current power is probably more economical for general mining use, other than haulage. In fields where cen- tral-station power is available pumps, hoists, fans and tipple equipment are all run by alternating-current power and appear to give more satisfactory results than direct current. In mines where alternating current is available and electric haulage is required it would be an added expense to install wire for alternating-current cutting machines, as the con- ductors that supply power for haulage can be used for direct- current cutting machines. This accounts for the general use of direct-current cutting-machines in the large mines where electric haulage is employed. To the small operator with limited capital the alternating- current mining machine has shown the way to a greatly in- creased production without the large investment formerly MINING COSTS 111 required. It is possible to decrease mining costs and increase production at a comparatively small expenditure above the cost of the machine itself. Arc wall cutters. The Jeffrey-Drennen adjustable-turret coal cutter, was designed to make a cut at any elevation desired. This was installed at Jenkins, Ky., where the coal seam varies from 6 to 8 ft. in thickness. It is clear, bright and free from sulphur or other impurities with the exception of a band of shale located at a height of from 2 to 5 ft. from the bottom. This varies in thickness from nothing to 19 in. With the customary methods of undercutting it would be impossible when shooting to prevent this shale from becoming mixed with the coal, but by the use of a machine adapted to cutting out or removing this parting before the coal is shot down this difficulty is overcome. The machine is mounted on a turnable truck, which carries four heavy standards or uprights, on which the machine proper is raised or lowered or adjusted to the desired height at which to cut out the dirt seam. The cutter is designed for a minimum height of 2 ft. from the bottom, and can be adjusted to cut at any position between 2 and 5 ft. The raising or lowering of the machine is accomplished by power through a disk friction clutch, which enables the operator to absolutely control the elevation of the machine to a nicety, 3 ft. of a vertical move- ment being accomplished in about 25 seconds. The cutting is done in the shale at the bottom of the band with the lower nose of the bits cutting into the coal about !/4 in., which causes the shale to fall down in the kerf, after which it is cleaned out, loaded in cars and hauled out of the mine. This insures an absolutely clean product. A 15-ft. glace can be cut in 11 min. from the time the machine enters the room until it is ready to leave. Twenty- five rooms have been cut in a shift of 10 hr. The Utah Fuel Co. of Somerset, Colo., cut 258 lin. ft. of coal in 2 hr. 24 min. with this machine. The vein is 14 ft. thick. Post punchers. Compressed-air post mining machines of the radial type were installed at the Pacific Coast Coal Co.'s mines about 1909, and it was found that after a mining was put in with these machines the coal could be sent down the chutes 112 COAL MINING COSTS with only a little pick work, and without any powder, except an occasional light shot at a corner or to shoot out a " nigger head.'* In fact the powder consumption was reduced over 95 per cent. The lump coal was increased in this way from about 25 per cent to about 60 per cent ; and the practical elimination of powder made the mine much safer. The rooms are driven 45 to 50 ft. wide. The first cut is made from a post set 7 ft. from the left rib and about 18 in. from the face. A cut 8 ft. in depth is put in, using an exten- sion bar 80 in. long. The chuck enters the cut, which accounts for the mining being deeper than the length of the extension. After the 80-in. extension has been swung, a 100-in. bar is used to square up the cut. One man operates the machine, swinging it by means of the worm-crank with one hand, and feeding the cylinder forward two or three turns with the other at each end of the swing. The machine and the posts remain in the room at the face until the room is completed, for there is no shooting of the coal that can injure the machine nor any loading (as in a flat seam) that the machine would interfere with. Hence there is no waste of time due to moving, except from post to post, and little heavy lifting. Two men can set up the machine, with ease, as the heaviest parts (the machine and shell) weigh only 225 Ib. These machines will average about 300 sq. ft. in an 3-hr, shift, or from 45 to 55 tons per machine per shift, according to the height of the coal. While, as stated, the reason for the installation of these machines was solely to increase the pro- portion of lump coal, even if the cost of mining was increased, the results point strongly toward a material reduction in the cost of mining, after all interest, depreciation, power, pipe line, and maintenance charges have been made against the machines. In shooting off the solid, the former method, a yardage system of payment was used, the rate being $9.50 for a 50-ft. room. Three men worked together, furnishing their own pow- der. The reason for using a yardage and not a tonnage system was because of the impurities in the seams, and the pitch. Repair costs. It pays well to have the machinery in shape MINING COSTS 113 to run a full day when the mine is working. By this means whatever men are in the mine are given a chance to produce some coal, and the cost of production is considerably cheapened because repair costs are reduced to a minimum. This is important, for the cost for repairs is too high in almost every mine. Furthermore, by giving each machine the needed care there will be an increase in tonnage which will permit of a further saving in the cost of production. In order to lower the repair cost and increase the efficiency of the machinery, it is neces- sary to have a system simple but accurate, which will give an individual record of the performance and expense of each piece of machinery. There should be a book kept at each mine by the electrician for the purpose of recording the working hours of each machine, and in this the hours idle should be marked down. With a little attention it will be possible to get the average number of tons or cars each machine is capable of producing, and thus it will be possible to find at the end of each month the number of tons lost through the inefficiency of any machine. Every night the machineman should fill out a report show- ing the make and number of his machine and the hours it has been delayed, and he should state any defects he may have noticed in its operation. The electrician should then have these defects repaired and sign the report and forward it to the chief electrician. The mine electrician should also fill out a daily report on this order: Min e NAM: B OP COMPANY Dai Hours of Labor 3 2 Wage Rate 25 40 Make and Number of Machine or Locomotive SS No. 1620 Mining machine Repair Parts or Material Used 1 worm gear lkeyiXiX6in. Cost of Part $14.50 0.06 State if Broken or Worn Out and Cause Broken teeth worn too thin to stand strain REMARKS. Advise that this machine be changed to easier work as the section is full of rolls and the machine is too light for that work. S. G. MILLER. This report should be sent daily to the chief electrician, who would have each item charged against the machine on which it was used. Thus at the end of each month the exact cost of labor and material used on each machine could be easily found. 114 COAL MINING COSTS The wiremen and bondmen should report the section in which they have been working, the kind of work they did and the material used. This should be recorded, and it could then be seen at a glance how much copper or other material was used to work out a section. The mine electrician should at the end of each month send a report to the chief electrician, showing the horsepower, make and number of each motor, machine or pump at the mine, the number of hours in use, the amount of time out of com- mission for repairs and the amount of oil used on it. This, with the electrician's daily report, gives the chief electrician a chance to get down to facts. For example, if a certain mining machine gives much trouble, he can see at a glance just what part was at fault each time, and with these facts on hand he can proceed to investigate and remedy the trouble. Again, if a certain part wears out on each machine of a kind, he will know that it is a weak part in the construction of that machine, and he can take it up with the manufacturer who will probably be able to suggest some way to overcome it. The mine electrician should order his supplies once a month, and they should be charged against him and not against the repair cost until they are actually used and charged on his report. At the end of each month he should take an inventory of all the supplies he has at the mine and should be credited with them. For example, he has $500 worth of supplies on hand on the first of the month, and his requisition shows that he has received other supplies worth $500. At the end of the month, on taking his inventory he finds he has $400 worth left. This shows he has used $600 wortK of material. On checking up his daily reports they should balance with this figure to show that everything had been charged in its proper place. The mine foreman should mark the delays to the machinery on his report also, and this should check with the mine elec- trician's and efforts should be made to keep them correct. Each piece of machinery should be treated as a workman. The time in use, the amount lost, tons of coal handled or cut, cost of labor and supplies should all be completely recorded. This will show which is the most efficient machinery to buy MINING COSTS 115 for future use and which to discard, what parts wear longest and what parts are most liable to give out first. It also shows the amount of material, such as wire and hangers, bond, etc., in each section of the mine. In order to get results from this system it is necessary to have skilled and competent mechanics. One cannot reduce costs if the men fail to do their share. It is too frequently the case that companies using first-class machinery have in- experienced and underpaid mechanics tending their expensive equipment, while other companies having a poor grade of machinery have good mechanics. Many coal companies will buy an expensive piece of machin- ery and then let an inexperienced man experiment with it. The result is the repair cost is large for the first few years, though it ought to be practically negligible. There is nothing gained by letting someone experiment with machinery, and it is a costly practice. Through ignorance, most of such high repair costs are blamed on the equipment, whereas, if the machinery had been used in the right place and treated as it should have been, there would have been no trouble. The average cost of repairs per ton of coal mined is about lOc. while, with careful attention on the part of the chief electrician and the mine electricians, it would be an easy mat- ter to bring it down to 5c. per ton, and if all did their best, and the machines are in first-class condition it should be brought to less than 3c. with no interruptions to service during work- ing hours. Machine bits. The bit question is an important problem in connection with the operation of machines, for the bit is to the mining machine what the tooth is to the crosscut saw. When using chisel-point bits in connection with up and down pick points, the sharpening of these three classes of bits and their distribution to ftie different machines is quite difficult. The Sullivan Machinery Co. recognized this difficulty and fol- lowing the old 5-position chain it later introduced the 9 posi- tion superdreadnought. Straight pick-point bits in this chain give a good clean kerf, making comparatively coarse machine cuttings, while placing on the machine only a moderate load. Formerly all bits were sharpened by hand, but this is a slow and most expensive process. Small trip hammers will do 116 COAL MINING COSTS more than any other one thing to ease bit troubles. One man, with a boy to heat the bits, can sharpen 2500 to 3000 bits per day, making on an average a better bit than a man will make without the hammer, because a blacksmith sharpening by hand will endeavor to get the bits as hot as possible so as to save hammering. In doing this he not only makes a badly tem- pered bit but burns away valuable bit material. The die used in these hammers may be a home product that will probably be equal to anything on the market and that can be made for $20 per set. For tempering compounds a solution of cheap laundry soap and soft water gives good results at little expense. One cake of soap is used to 25 gal. of water, washing powder being added where the water is too hard to lather freely. At the beginning of the shift this mixture is heated to the boiling point and the sharpened bits while at a cherry-red heat are thrown into the tempering tub, the hot bits keeping the mixture boiling, insuring a slowly cooled and evenly tem- pered bit. The object is to furnish each machine with plenty of bits and thus encourage the runner to change them before they become dull enough to load the machine. It is cheaper to sharpen bits frequently than to run the risk of burning out armatures and wearing out cutter chains. In Nos. 9 and 11 seams the West Kentucky Coal Co. uses approximately 150 sharp bits to cut 100 tons of coal while in the No. 12 seam the tonnage per bit is more than double this amount. The number of bits required may seem high in this instance but it was frequently the case in these mines that 150 bits would be dulled in a single room where rock or heavy sulphur bands were encountered. In spite of these conditions the records over three years at these mines show an average production of 35,000 tons per 1000 bits purchased. Loading machines. Because of the decrease in the supply of labor throughout the country and the increasing wages more interest is being centered on ways and means for reduc- ing the amount of muscular energy required in the mining and loading of coal underground. Undercutting and loading are probably the two jobs least sought for in American coal mines, and it is increasingly difficult to procure men who will per- form this kind of work, MINING COSTS 117 Many machines have been designed for the loading of coal. Some of them both mine and load the material, while others merely load it. In general, no difficulty has been encountered in devising an apparatus to place the coal upon the cars, the main disadvantage of such devices being that it is impossible to feed cars to the machines with sufficient rapidity to make mechanical loading practicable. As a result such machines stand idle a goodly portion of the working day waiting for the loaded cars to be taken away and their places filled by empties. Furthermore, these machines are, as a rule, cumber- some and difficult to move because of their great weight. In hand shoveling the height through which the shovel must be raised determines the capacity of the loader. With the same expenditure of muscular energy to raise the coal, a man will load 29 tons into a wagon 52 in. above the rail, 44 tons into a wagon 32 in. above the rail, and 140 tons into a conveyor 8 in. above the floor, allowance being made for clear- ance in each case. Also it must be remembered that in raising the shovel, the foot pounds expended in raising the body is greater than the foot pounds exerted in lifting the coal. When loading into a conveyor in connection with a loading machine, the coal does not have to be lifted more than a few inches; some of it can be rolled or pushed on, and there is an important saving in muscular energy effected. The Jeffrey loader weighs one ton. It has the motor and machinery located in the center of the conveyor over the sup- porting rail, and therefore one man is able to bear down on the back end and slew the machine around at will, and one man can readily push the loader along the supporting rail to any desirable position. The supporting rail is ordinarily placed on two wooden horses. The loading machine is made strong enough to admit of shooting the coal down on top of the front end of the conveyor. It is provided with a self-propelling truck to move from one place to another. For best results with the loading machine the mine should be laid out systematically with just enough loaders in each section to keep one under cutting machine busy. For instance, in Fig. 49 is shown a system of ten rooms. In rooms 1 to 5 are loading machines; in rooms 6 to 10 a shortwall machine. 118 COAL MINING COSTS Each room is loaded out in half a shift, therefore all ten rooms are undercut and loaded out once a day. One gathering loco- motive will handle cars for the five loading machines, if fairly good size cars are used. The Jeffrey pit car loader is a simple conveyor, driven by an electric motor, so located as to add to the stability of the machine. Its novelty would appear to lie in its cost, which is about one-seventh that of a shoveling machine. It will doubtless appeal to those who contend that coal mining is attended with too many delays to tie up large amounts of capital in expensive machinery, and attract those who fear that the mine is no place for a machine designed to pick up the coal. FIG. 49. System of mining suggested for adoption with coal loading machines. The claim of its maker that the output from a room can be doubled by the use of this machine appears to be conserva- tive. The fact that it requires only two men to operate, whereas other machines require from four to eight, is vastly in its favor, for when the inevitable delays occur, it is not difficult for two men to find useful work in posting, extending track, etc. Quite a number of these machines are in operation at the mines of the Dominion Coal Co., Canada, some of them for a period of three years or more, and it has been proven that two men will average from 50 to 75 tons per shift. Three men using this machine can load out the coal from a 16-ft. entry, 6-ft. undercut, in one hour, where the height of coal is 42 in. with a 2 in. to 6 in. parting, the dirt from which has to be picked out. Three men have loaded 7 cars of slate in 35 min., the capacity of car being 3400 Ib. of coal. MINING COSTS 119 In 5 ft. coal with 4 in. binder, two rooms were drilled, shot and loaded out in 8 hr., making 39 cars of coal, each holding 3300 Ib. About 6 in. of draw slate and binder were cleaned out and gobbed. The cost of this machine in December, 1921 was $1500. The Westmoreland loader weighs 10 tons, is electrically operated and self-propelled. The truck is rigid, with 12-in. wheels, one front wheel loose axle, 4-ft. wheel base, and when in operation the truck is clamped rigidly to the rail. A 3-ft. geared turntable rests on the truck frame, and on this is mounted a solid steel case, 12 ft. long, 32 in. wide and I2y 2 in. high, which swings free above the wheels through an arc of 160 deg., 80 deg. right and 80 deg. left, the width of sweep being 16 ft. The steel case contains a ram in the form of a chute, 2 ft. wide, 6 in. deep and above the floor, and with an adjustable shovel end at the front. The ram-chute hangs on rollers and is moved forward and backward by a rack and pinion at the top. This represents the unique feature of the machine. The ready horizontal movement of the ram, together with its sweep of 8 ft. on either side of the machine, enables it to keep close to the loose coal throughout the full 16 ft., or slightly more, of width and thus facilitates the work of the laborer. The adjustable shovel end of the machine takes care of undulating bottom. It is rigid in operation for all sizes of coal, fine slack up to 200-lb. lumps being picked up at the rate of one ton per minute by the chain conveyor in the ram. The actual rated capacity of the loader under fair conditions is 20 tons per hour. Five men are required for its operation, including the driver. The machine will work in seams that are not less than 4% ft. thick, and will pass over curves of as low as 12-ft. radius. The total power requirement is 16 hp., consisting of three reversible motors. The Evans scraper loader is an adaptation of the old main- and-tail rope principle to the operation of a modified scraper as a means for conveying the coal from the face and out to the room neck, at which point it is dumped into the car. The apparatus consists of a double drum hoist, which may be driven by either air or electric power, two wire ropes of suitable lengths, and a V-shaped bottomless scoop, or drag. ID 120 COAL MINING COSTS addition to this major equipment, each room is provided with deflectors, loading pans and aprons. All coal is loaded on cars in the entry, the plan being to set a trip of empty cars above the room neck so that no time will be lost in spotting another car when one is loaded. It is figured that with rooms of 300 ft. length it is possible to load at least nine tons of coal per hour. The average haul in this case amounts to 150 ft., and the hoist being geared for a rope speed of 300 ft. per minute the traveling should be done in one minute actual running time. Assuming that a minute is lost at the face and another minute at the loading point, there would be 20 trips per hour, which on a basis of 900 Ib. per scoop amounts to nine tons. It requires five men to successfully operate the apparatus; one hoist man, one man on the entry to stop and trim the cars, two at the face and one timberman. The foremost saving is in the loading of cars. A full crew for the Myers- "Whaley machine consists of one runner and one car coupler, who load as much as 15 to 20 hand shovelers. With these machines the same working places can be loaded out twice a day, instead of once in two days, as is commonly the case. Hence a machine-equipped mine will produce a given tonnage with one fourth the development required where the loading is by hand. This concentration, with the consequent reduction of trackage, ventilation and mine car equipment, effects important economies. The machine is operated by one man, will work on elec- tricity or compressed air ; and is self propelling forward or backward. The machine runs on either. The machines operate at the rate of 13 to 18 strokes per minute, with a power consumption of .22 kw.-hr. per ton loaded. Three sizes are built for underground mining as follows : Height Length Width Net Weight Mine Height Required No. 2. 46 in. No. 3. 47 in. No. 4. 54 in. 19 to 21 ft. 20 to 22 ft. 22 to 26 ft. 4 ft. 9 in. 4 ft. 11 in. 5ft. 4Jin. 9,000 11,000 18,000 4| ft. 5ft. 6ft. MINING COSTS 121 The machines are all capable of loading at over 1 ton per minute. The smallest handle 30 to 45 tons per hour, and the larger sizes from 50 to 60 tons per hour, actual shoveling time. A test of one of the earlier makes of this machine was made at the mines of the United States Coal and Oil Co. at Holden, W. Va., about 1910, which is of interest. The coal at Holden LOADER'S DAILY REPOKT Sixth day of official test run. Holden, W. Va., Aug. 31, 1910 Time Started 7:30 a.m. Time Stopped 12:03 a.m. Total Time a.m. 4 hr. 33 min. Time Started 12:33 p.m. Time Stopped 5:32 p.m. Total Time p.m. 4 hr. 59 min. No. of Men Working: a.m. 4 p.m. 4 Working Place Time Loading and Shifting Cars Time Changing Machine Time Lost Total Time Consumed Cars Loaded No 6 room 1 h 3 m 1 h 1m 2 h 4m 24 tons No 7 room 2 h 29 m 25 m 23 m 3 h 17 m 54 tons No. 1 face room (neck) . 1 h. 45 m. 1 h 25 m 24 m. 11 m 26m. None 2 h. 35 m. 1 h 36 m 33 tons 39 tons Totals 6 h. 39 m. 1 h. 3m. 1 h. 50 m. 9 h. 32 m. 150 tons REMARKS. Time lost No. 6 room loose setscrew on conveyor sprocket tightened between 7:30 and 8:15 = 45 min. Off track 7 min.; pulling down coal 9 min. Total, 1 hr. 1 min. Time lost No. 7 room waiting on driver, 1 min.; off track, 5 min.; pulling down coal, 17 min. Total, 23 min. Time lost No. 1 face room neck off track, 9 min.; pulling down coal, 17 min. Total, 26 min. Time lost No. 1 room none J. B. HAILE, Machine Runner. Report by W. Whaley. is a peculiarly, hard, tough bituminous coal, known as No. 2 gaseous. Its coherent qualities make it somewhat difficult to shoot down and nearly as hard to shovel as large lumps of limestone. During six days of its operation a record was kept of all features of the run, cars loaded, time required to load and shift, time to change the machine and time lost for any cause. During the six days the machine loaded 768 tons of coal, the time of loading and shifting cars being 36 hr., 6 min. ; time changing the machine 4 hr. 21 min. ; time lost, 14 hr. 33 min. ; total time, 55 hr. The machine loaded out four rooms per day 122 COAL MINING COSTS and part of the lost time was due to delay in shot firing, in waiting on smoke to clear out of the room and other items not chargeable to the machine. A copy of daily report of the last day of this test run is shown herewith. In spite of the delays and lost time the machine averaged 128 tons per day. The lowest day's work (due to lack of coal) was 90 tons, and the highest 150 tons. The average time of loading and shifting a car was 8.4 min., or 21.3 tons per hour. The territory in which the machine worked consisted of seven rooms recently turned off the airway in No. 5 mine. Much of the work was done on curves. The rooms ranged in width from 21 ft. to 27 ft. and all but two had two tracks. The rooms were on the butt of the coal which greatly increased the difficulty of shooting. The company had no difficulty whatever in keeping the machine supplied with cars, which were hauled to and from the side track on the entry by a mule. The crew of the machine consisted of four men, as follows : One machine runner, one man in front, and two men to handle cars and pick slate. Mining and loading machines. The Ingersoll-Rand cutter and loader consists of a powerful air-driven puncher, mounted on a carriage over a conveyor. A lever is used to elevate or depress the puncher pick, a second lever moves it from rib to rib through the agency of a small air engine, while a third lever moves the whole puncher together with its truck bodily toward or away from the face, the conveyor remaining station- ary within certain limits, or accompanying the forward or backward motion of the puncher as desired. The puncher itself makes about 160 strokes per minute, and is controlled in the same manner as the ordinary hand- operated machine, going from rib to rib and making a cut extending the entire width of the entry. The conveyor is driven by a separate engine suitably controlled by a stop valve. The mine car is filled evenly by moving it about three times during the loading process. Making the undercut is the longest part of the operation, requiring from 25 to 40 min., according to the hardness of the coal. The knocking-down process ordinarily takes place faster than cars can be supplied. When it is completed the conveyor MINING COSTS 123 is pulled back about 7 ft., a section of track is laid and the conveyor again moved forward into position. Two set-ups during the shift constitute a 10-ft. advance of the entry. It is estimated that under suitable operating con- ditions, such as high coal of the quality found in the Pittsburgh seam, that the machine should average 250 ft. of advance per month of 25 days, working day shift only, the night shift being reserved for the laying of tracks, piping, etc. Where conditions have been suitable and the machine has had an adequate supply of cars, as much as 20 ft. of advance has been made in a single shift. The average advance in the entries of the Annabelle mine in West Virginia where several of these machines have been in operation since 1913, is somewhat over 10 ft. per shift. The cuts at this mine average 10y 2 to 11 ft. wide and 6y 2 to 7 ft. high. The Jeffrey cutting and loading machine undercuts, shears, breaks down, and loads the coal into the mine cars. No explosives are used. The machine stays in the entry until driven as far as desired. It is fed forward in a pan similar to a breast machine. The feed forward is 7 ft. Average depth of cut 6 ft. 6 in., width 5 ft. The time required to make the cut depends upon the con- ditions and nature of the coal. Ordinarily the sumping cut takes less than 30 min. ; the open cut less than 20 min. Mov- ing sideways to next cut, 3 to 4 min. The Jeffrey cutting and loading machine has two ver- tical shearing chains between which is mounted a frame carrying a number of heavy punching picks. This frame is readily raised and lowered by the operator, who can cause the picks to strike at any height he desires. The coal falls onto a conveyor which is made thin enough to go into the kerf cut by the undercutting chain. This conveyor carries the coal to the rear end of the machine and dumps it into a second con- veyor, which is mounted so that it can swing at any desired angle to the machine. After a cut is made, the machine is moved sideways by means of a rope hitched to a jack at the opposite rib and then another cut is made. When slate is to be piled up at the side of the room, the rear conveyor is swung sideways, the slate 124 COAL MINING COSTS rolled onto the front of the machine and gobbed by the two conveyors. In a test run the machine required an average of 13^2 min. to make a cut and about 3 min. to move the machine to the next cut, or about 16 or 17 min. time for each cut. The coal loaded averaged 21 tons per hour, although where roof conditions were good the machine had loaded 30 tons an hour in the same time. The height of the coal is 5 ft. 8 in. In another district an average of 60 tons per hour was made for the three months that the machine was in operation. In the driving of an 11-ft. entry it has averaged 20 ft. advance per shift of 8 hr. under rather unfavorable circumstances. The crew for each machine consists of a machine runner, a helper and a driver. If the slate is heavy it may require an extra man for this handling. At the Valier mine in Illinois these machines have been introduced to accomplish rapid development work, promote safety through elimination of explosives and prevent the shat- tering action of explosives on the ribs and roof of the entry. The cutting in this mine is unusually hard, but under ordinary conditions an advance of as much as 150 ft. per week was made with each machine, by working three shifts. On single shifts the machines make a general average of 300 ft. per month. With regard to the economy of the use of these machines, it is the opinion of Mr. Carl Scholz, general manager of the company, that the cost of installation and operation will be repaid many times by the saving in timbering because the coal is not affected by the use of explosives when these machines are used. The roof will stand much better than it does when so shattered and timbering will be unnecessary in most parts of the entries. These machines are being used on the main west entries where a possible distance of 3^ m i- can he driven. After an experience of about 1 yr. in entry driving in this mine, a difference between the standing qualities of these entries and those driven with explosives can be observed. The O'Toole machine undercuts, breaks down the coal and loads it. It breaks up the coal completely and is thus designed specially for use in mines where the production of lump coal is of no advantage as for instance where the mine is producing solely for coke making purposes. The machine consists of a MINING COSTS 125 ffi TjH O5 O (N CO CO to 2 2 3 3 8 g g S3 g 3 rH > ,0 ex I I a _o e 1 *o c / A! nr MINING COSTS 155 ing. But seldom or never he gets enough cars to earn any- where near that amount. It has been stated that the men in central Pennsylvania earned more money when pick coal was 35c. a ton than they do now. But the past and present conditions have not been taken into account. The blame for the loss in earning power is often angrily put on the union, whereas it ought to be partly borne by operators for not maintaining correct conditions men in proportion to the equipment and bosses in proportion to the mine. There is an inflexibility in the rates paid for work which makes an inequality in the earning power of the men. Rates are generally the same for all parts of a mine or for all parts of a vein. In one anthracite mine $1.58 is paid for a 2-ton car of coal whether the vein is 8 ft., and free blowing or 4 ft. and hard coal. In the bituminous regions the same price is paid for pick coal whether it is in rooms or in crushed pillars. Business has been business in establishing the rates, and in any argument between employer and employee over what the rate should be, the attempt on both sides has been to charge all the traffic would bear. There has been no measurement of work by which attempts could be made scientifically to get an equitable basis for payment. Measurement of work is just as practical in mining as in manufacturing, where it is now being largely done. Although the payment is made according to the ton or car loaded, responsibility for the amount of work done cannot be wholly cast aside by the management. Men are used in a mine entirely too freely and expected to be on their own resources. They cannot do it, for they do not work alone, but are depend- ent upon others for the opportunity to work. The drivers, being paid by the day, are not anxious to deliver more cars than are necessary to keep their jobs and so neglect the miner. In consequence the miner does .not earn as much as he might. One can hardly blame the driver so long as it is a fixed principle in life to get as much for as little as possible. Universally a mine is so undermanned by bosses that no work can be followed up, and the neglect of duty and failure to cooperate can become so common that the miners frequently suffer from needless delays. 156 COAL MINING COSTS It is hard to stir up trouble at a mine where the men are being properly treated and are earning a day's pay. It will be found that the strength of a union increases at a mine in proportion as there is more cause for discontent. But this is not to be regarded as an argument that the men should be paid whether they do any work or not. Increased earnings of the individual can be reflected favor- ably on the cost sheet. A man who is working a piece of coal which has been formerly neglected and is so situated that he is given only a car or two a day and whose earnings in con- sequence are low is likely to be a source of trouble. When it is necessary to keep a man and a mule for only a few cars a day, the transportation charge will be high. The cost of gathering was taken in one mine and found to run from 2y 2 c. to 25c. The total cost for haulage was high because on only a few roads were there enough men to give a driver a full day 's work. When the work was apportioned so that there were enough men on each road, the cost fell. It would be more advantageous to the workmen if the union leaders, instead of making demands for further increase in the rates of pay, asked for a better organization of the work so that the men were furnished an equal number of cars. Rates are high enough so that no man need overwork himself in earning good pay. In a boiler-room a fireman will shovel into the furnace from 18 to 20 tons a day, which is more than a miner will load into cars. One well-known union leader has been quoted as saying that 15 tons was not too much for a man loading after machines, and the union representatives agreed to a task of 12 tons at one mine. When one gets into the complex changing condi- tions of an anthracite mine it might seem a little doubtful whether this could be done, but again it is a case of the arrange- ment of work rather than a crying need of an increase in rates. Under the present general manner of payment the only cure is to promulgate the idea that each man should produce a large tonnage per day. An increase in their output will be of advantage to the miners, as they are paid by the ton. It will also be to the interest of the company, for an increased output per handling unit will result from the proper organiza- MINING COSTS 157 tion of work by which the larger product is secured. Though there is no one cure for all ills, the application of the principles of efficiency to mining will greatly help to alleviate the con- ditions which are now a frequent and natural source of much irritation. The accompanying table gives the wage scale in effect in the Hocking District from 1898 to 1921, covering all classes of labor in and about the mines. This table will be found of use in computing the comparative cost of doing the various kinds of work cited throughout this book where the figures are not up to date. The Hocking scale as it is generally known has been used for this purpose for the reason that it dates back the farthest of any of the important scales and also because it is used as a base for all wages in the Central competitive field and has influenced wage settlements in all union fields generally. The break in the table in 1914 was occasioned by the change from the screened coal basis of payment to the mine run. Losses from idle time. The yearly average of days of activity in the mines of the United States from 1908 to 1914 was 217. Thus the mining plants averaged only about 72 per cent of full time. In the year 1914 the number of working days, according to United States Geological Survey report, was only 207 for all coal mines throughout the country. Unfortunately even this time is not by any means evenly distributed. Some mines are idle a large percentage of the year, while others work with regularity. A few mines in agri- cultural states work mostly in the winter, and in the summer the men must find farming jobs to keep them busy, though it is doubtful whether many of them do. A miner is not always willing to work under a hot sun. Arkansas and Oklahoma in 1913, for instance, only worked 174 days, or barely 55.5 per cent of the possible working time. Perhaps under no circumstances will it be possible to keep coal mining at an even pace throughout the year, but it should be possible to do better than the 1914 season. Such depression as occurred in that year would be helped by a combination embracing all the mine operators, but the conditions of unem- ployment were nation-wide, and something broader than a com- bination of mine owners would be necessary as a stabilizer. 158 COAL MINING COSTS WAGE SCALES IN THE HOCKING DISTRICT, CENTRAL 1892 to 1894 Feb. to April, 1894 1894 to 1895 June to Oct., 1895 1895 to 1896 March to Oct., 1896 Pick Mining Screened lump, per ton Mine run, 5-7 lump price Entry, .5511, .5950 .6330 .6755 .03 .03 .03 .03 .03 .03 .03 .03 .03 .03 .014 014 .01} 02 .021 .024 021 02* .0263 .0276 4.20 4.50 . V/A'g v** j E. P. E. P. E. P. E. P. E. P. E. P. E. P. E. P. E. P. E. P. x Special prices according to nature of work. 160 COAL MINING COSTS WAGE SCALES IN THE HOCKING DISTRICT FOR THE YEARS 1892 TO 1921 Continued 1914 to 1916 1916 to 1918 April to Oct., 1917 1917 to 1920 1920 to 1921 Pick Mining Run of mine, per ton $0.676 2.4996 2.4996 1 . 7330 3.7884 2.84 2.62 1.32 1.50 2.84 2.84 2.84 2.84 2.78 2.84 2.84 $0.6764 2.6245 2.6245 1.8196 3.978 2.98 2.75 1.40 1.59 2.98 2.98 2.98 2.98 2.92 2.98 2.98 2.98 2.75 .074 . 10235 .1084 . 12016 .406 .426 .49945 .51945 .51945 .489 .02 .0195 3.27 2.95 2.75 2.95 2.75 2.28 1.64 $0.7764 2 . 6245 2.6245 1.8196 3.978 3.60 3.35 1.90 2.19 3.60 3.60 3.60 3.60 3.52 3.60 3.60 3.60 3.35 .0890 .1173 $0.8764 3.0181 3.0181 2.0925 4.5747 5.00 4.75 2.65 3.59 5.00 5.00 5.00 5.00 4.92 5.00 5.00 5.00 4.75 .1040 .1365 .1384 .1518 .5760 .5960 .6835 .7035 .7035 .6684 .02 .0195 5.27 4.95 4.75 4.95 4.75 3.24 1 3 R >> i # 3111* qiiii $1.1164 3.6217 3.6217 2.5110 5.4896 6.00 5.75 3.18 4.59 6.00 6.00 6.00 6.00 5.92 6.00 6.00 6.00 5.75 .14 .1790 .1744 .1905 .80 .9290 .9290 .9290 6.27 5.95 5.75 5.95 5.75 4.24 Entries, dry, per yard Inside Day Labor (Where old men are employed) Bottom cagers, drivers, trip riders, per day. . Snappers on gathering locomotives, per day. . Water haulers, machine haulers, per day. . . . Timber men, per day Pipe men, for compressed air plants, per day. All other inside day labor, per day Spike team drivers, 25 cents per day extra. . . Machine Cutting By Jeffrey style of machines, in room, per ton. By Jeffrey style of machines, in entry, per ton. By punching machines, in rooms, per ton By punching machines, in entries, per ton. . . Loading 2.62 .07 .0970 .1044 .1156 .38 .40 .4690 .4890 E. P. .46 .44 .02 .0186 E. P. 3.12 2.81 2.62 2.81 2.62 2.18 1.56 .4910 .5110 .5845 .6045 .6045 .5740 In rooms with hand drilling, per ton. In entry with hand drilling, per ton Breakthroughs in entry, per ton Breakthroughs in rooms Drilling By hand per ton By machine, per ton 3.87 3.55 3.55 3.55 3.35 2.88 2.24 ^IJ - ;N i^3 Room turning, cutter and loader Outside Day Wage Scale First blacksmith, per day Second blacksmith, per day Blacksmith helpers, per day Carpenters, per day Dumpers and trimmers, per day Slack haulers, per day Greasers and couplers, per day Where engineers and firemen are employed by the day, the minimum rate is $4.75 for 8 hrs. This does not apply to men employed at a monthly rate. This also applies to coal washers. MINING COSTS 161 The whole nation must get together to produce a stability in business which will make steady work in coal mines and in every other form of activity. The case of the miner against irregular operation has already been forcibly set before the public. What is not so generally realized is that the case of the operator is just as damaging to him. His capital is idle and his mine equipment, instead of benefiting by a rest, is rapidly depreciating. Al- 1.0 U 1.50 1.40 UO 1 on | / / / Percentage Increase in Cosf 3 S s! 2 S g 5 ; / / / / / / / / / > .40 .30 .W /I x X ^ ^ .10 \ ^ ' ^ ^-' l ^ H-O^ r**^ 95 90 85 80 15 TO 65 60 55 50 45 4 35 30 ZS Z IS 1C 5 Percen-fags Decre&ecil Car Supply FIG. 53. Percentage of increase cost due to irregular working time compiled from 830 observations in the New River field. though the mine shuts down, his fixed charges run on not only interest charges and salaries, but a host of maintenance charges as well. And in the end the coal consumer pays the bill for idleness of miner and mine. In this connection we may find instruction in an exceed- ingly valuable study made by Messrs. Garnsey, Allport, and Norris of the costs of production as effected by interruptions of working time. Fig. 53 is taken from their " Report of the Engineers' Committee of the United States Fuel Administra- 162 COAL MINING COSTS tion, 1918-19. ' ' It represents an analysis of the monthly records of seventy-three operators in the New River district of West Virginia. Each of these was carefully analyzed, and the per- centage increase of cost for each of the 830 observations thus obtained was plotted; weighed averages were then taken at each 2.5 per cent from 70 to 100 per cent working time, and for each 5 per cent below 80 per cent. The result of this study is shown in Fig. 53, which has been checked by numerous observations from practically every field and has been found, within reasonable limits, to be correct. This diagram can and has been used in reducing to normal cost the reported cost of collieries shut down during parts of months. The reason for the increased cost per unit of output is, that the smaller the number of tons produced the larger is the share of fixed over- head expenses that must be borne by each ton. F. S. Peabody testifying before the Frelinghuysen committee in 1919 stated that: "the earning of the laborer and the cost of coal depend entirely on continuous work. Our costs will vary from month to month, dependent on the running time of our mines. There will be a variation of between 50 and 60c. a ton from month to month, depending on the number of hours the mines are idle." The average number of productive days worked per annum in Illinois and Indiana is only about 175 out of a possible 300 or more. This idle time of the miners is not confined to one season or period during which they can find employment else- where. The men are always subject to call, for which reason they urge a greater daily wage so that that their annual in- come may be sufficient for their needs. This causes these operators to grant abnormal wage advances, which are directly reflected in coal cost. Many industrial plants which produce standard or basic commodities find it possible to operate 24 hr. per day by using different shifts of men. They work also for 310 or more days a year, or a total of 7440 hr. per annum. Still other industries, on two 8- or 10-hr, shifts per 24 hr. work 300 to 310 days per year, thus operating 5000 to 6000 hr. every 12 months. Even one 8-hr, shift in each 24-hr, period with 310 days per year gives 2480 working hours in every 12 months. Be- cause of the unrestricted competition the mine operators of MINING COSTS 163 Illinois and Indiana have built more plants than are needed and can only operate for 8-hr, out of every 24, and for 175 days per year, or 1400 hr. It will be seen, therefore, that as against 100-per cent plant utilization (24 hr. for 310 days or 7440 hr. per annum) pos- sible in some industries, and as against an average by all industries of 33 per cent to 45 per cent (one 8 or 10-hr, shift per 24-hr, period for 310 days), a coal plant is in actual pro- ductive use only about 18 per cent of the mine. This makes plant, interest and depreciation charges six times as heavy as for some other industries. The 97,000 miners of Illinois and Indiana who are prevented from working 125 days per year might at the 1915 wage have earned an additional $36,400,000, or $371 per man per year, had their employers been able to give them work or had their efforts been expended in other directions. Because the operators can give their miners work only a part of the time, these operators must pay higher daily wages than are warranted by the current selling prices. Their labor cost (in 1915) is 92.44c. per ton, whereas the selling price is but $1.14 and $1.11 respectively for the states of Illinois and Indiana. Statistics of the coal-mining industry for 1912, give a total production of bituminous coal of 450,000,000 tons, with a total production cost per ton as follows : Direct labor per ton $0 . 5425 Indirect labor (day work) 0. 2383 Salaries 0. 0575 Supplies 0. 1305 Royalties 0.0320 Miscellaneous expenses 0. 0530 Total per ton $1.0538 Assuming that as claimed by competent authorities, the output from the same workings might have been 600,000,000 tons, an increase of 33V 3 per cent, with little advance in the expenditures, let us see how this increase in output would affect the production cost per ton. Direct labor and royalties probably would remain about 164 COAL MINING COSTS as before, but the other expense items would show a reduction due to the increased output about as follows : Direct labor per ton $0 . 5425 Indirect labor (day work) 0. 1790 Salaries 0.0432 Supplies. 0. 0978 Royalties 0. 0320 Miscellaneous expenses . 0400 Total per ton $0.9345 This shows a decrease in production cost per ton of 11.93c., or 11.4 per cent, due simply to increase of tonnage produced. Economic aspects of conservation. It has been customary to refer to the unlimited coal resources of the United States, but while our country has been wonderfully blessed in this respect, the exhaustion of some of the choicest coal beds is already in sight, as for instance, the high-grade coking coals of the Connellsville region. Conservative estimators realize that the coal supply should now probably be measured by hundreds rather than by thousands of years, and that it be- hooves us to conserve our fuel resources. Large areas of high- grade fuel undoubtedly yet remain, and it is not too late to utilize these deposits much more fully than has been done with similar deposits in the past, both by more skillful mining and by a more thorough utilization of the coal after it is brought to the surface. The Anthracite Coal Waste Commission in 1893 estimated that probably not over 30 to 35 per cent of the coal originally contained in the areas mined over has been saved, a.nd that even by working over the old culm banks and reworking the area already mined, not over 10 per cent additional would be obtained, thus giving a loss of 50 to 60 per cent of the original coal. By means of the very close washing and sorting of the small sizes, and the better removal of the pillars through the crushing of culm and other methods, the amount of waste may be somewhat less than is estimated by the Coal Waste Com- mission, but still an enormous waste is going on, of a material which is not duplicated anywhere else in the country, and of which the supply is comparatively limited. Dr. I. C. White, has estimated about 1909, that in mining MINING COSTS 165 bituminous coal in the United States not over 50 per cent of the coal in the ground has been obtained. This figure is exces- sive for present-day practice, still the amount of waste is entirely too great and should be decreased. That great portions of our coal deposits have been skimmed over, leaving rich territory abandoned because of poor and inadequate methods, which aimed at the recovery of the easily accessible and abandoned the more difficult, is apparent to every observer. Much valuable coal property has thus been wasted or ruined. It is in a sense a moral obligation on the operator to recover the pillar and top coal that the loss to the country may be lessened, but where this involves an additional expense, it can- not be undertaken. Still, for every two acres of coal land which the operations exhaust they leave, one acre of coal unrecovered and unrecoverable in the ground. This means that in Illinois each year 12,000 acres of coal land are exhausted, whereas the exhaustion should be but 8000 acres. In Indiana the depletion is 3000 acres, whereas it should not be more than 2000. In the whole country 100,000 acres are exhausted, whereas not more than 65,000 or 70,000 acres should have been thus made of no mineral value. A summary of the recovery effected in the different mining fields will give a basis for fixing a standard recovery to work towards. Conditions prevailing in most of the important min- ing fields in 1914 were described in a paper presented before the West Virginia Mining Institute in that year. In order to get an idea of what is being accomplished in other fields, inquiries were sent out to different sections of the United States. The results are shown in a condensed form by the accompanying table. It will be noted on this table, the wide variation of per- centages given for different districts in various states. All are large producers of coal, with one or two exceptions, and employ what are presumed to be modern methods of mining. It is noticeable that the thin seams usually are overlaid with good roof and the percentage of recovery is high. Also, that but one operator expects the ultimate recovery to fall below his present percentage. In the southern Colorado field, where the roof and bottom 166 COAL MINING COSTS conditions are favorable for pillar drawing, no roof coal is left for protection and the recovery is given as 80 to 90 per cent, working on the room-and-pillar system. It is claimed that in the Canon City district, where the longwall system is used, that 100 per cent of the seam is recovered. This is in the thinner seam, which measures about 3 ft. Rooms in the southern Colorado district are driven 16 to 18 ft. wide on 45-ft. centers, while in the Walsenburg district, where the coal is harder and has less cover, rooms are driven 35 to 40 ft. wide, leaving the same thickness in pillars which are recovered by machine and pick work. Track is laid on each side of the room and frequently one or two cuts are taken off the side of the pillar with a machine before beginning pick work. The bottom of the Colorado seams is usually slate of a soft character, which heaves when weight is thrown onto the pillars, making it necessary frequently to drive a skip along the pillar in order to reach the back end before beginning to draw it. There are districts where both roof and bottom conditions are unfavorable and much difficulty is encountered in breaking the overlying strata. In these sections the recovery is esti- mated to be 60 to 65 per cent ; 15 or 20 per cent is lost in roof coal because the strata next overlying the coal cannot be propped. Another company, operating in practically the same field, states its recovery runs 75 to 80 per cent of the entire seam, while a higher ultimate recovery is expected. This firm is now driving room entries to the boundaries and the last rooms are worked first, thus making it possible to draw the pillars on the retreat. In the Michigan field, especially in the Saginaw district, the coal is in pockets rather than a continuous seam. The basin lies for the most part in a low, flat country, and shafts about 200 ft. deep are necessary to reach the coal. The bed averages about 3 ft., and is of poorer grade than the Ohio and Pennsylvania fuels, so that its market is somewhat limited. The top in these mines is usually black slate, while one mine has a fireclay roof, making it necessary to leave top coal. Yet rooms are driven 40 ft. wide with track along each rib. The length of the room is 150 ft., as the miner pushes his cars MINING COSTS 167 from the working place to the entry. With the conditions just given, the recovery claimed is between 80 and 90 per cent. The 65 per cent recovery given in the table (Item 4) repre- sents the result of leaving pillars for surface protection within the city limits. Going into central Illinois fields, where the No. 6 seam is operated extensively, there are adverse public feelings and unsettled industrial and labor conditions, which materially affect the percentage of recovery. Surface costing $100 to $250 per acre cost the operator two or three times these values in cases of subsidences, if the mining rights do not clearly cover the property. Besides these factors, the companies operating in the Glen Carbon, Mt. Olive and Divernon fields, state that owing to thick, soft clay under the coal or great overburden (300 ft to 400 ft.) that they do not recover more than 50 per cent. In the southern fields better results are claimed, since the cover is about 110 ft. thick and all soft. The slightly inferior seams of coal above or below the Nos. 5 and 6 seams are now receiving considerable thought as to future values, and for this reason they are trying to prevent roof movements by leaving sufficient pillars. In the Sherrard field of Illinois, the recovery is reported at 90 per cent. Here the top and bottom are good and conditions propitious for drawing pillars. The seam of coal is only 3 ft. 8 in. thick, with many clay veins running through it, which evidently must effect recovery to some extent. An inquiry sent into the southwest section of Pennsylvania shows a recovery of 72.5 per cent. Here 10 in. of roof coal is allowed to remain on account of drawslate and the operations for the past three years have been under the plant and town. The bottom in this mine is fireclay of rather soft character. The rooms are driven from both sides of entries on 60-ft. centers and widened to 21 ft., leaving 39-ft. pillars to be drawn by machine and pick work. By this method, the ultimate recov- ery is expected to show a material increase over that given. The company reporting from Westmoreland County, Pennsylvania, where conditions seem favorable, both in the steam- and gas-coal fields, shows a recovery of from 82 to 86 per cent of the entire seam ; and expects the ultimate recovery to fall below these figures. 168 COAL MINING COSTS In Somerset County, Pennsylvania, where coals of the Allegheny series are worked, the recovery is given as 94.75 per cent of the entire seam. Here, excellent roof and bottom conditions prevail, and most room headings are driven to the limit before any rooms are driven at all. Then, the rooms are started at the rear and pillars drawn as soon as these are finished. Practically the same conditions exist in the George's Creek field in the Sewickley seam, only the recovery is reported as 97 per cent. In this same field, the results obtained in the TABLE OF PRINCIPAL FACTORS GOVERNINO Per Cent Ultimate Period 5 Operating District and State of Recovery of Recovery Compared to of Opera- tion, Average Height of Seam Roof Coal Carried Entire Seam Present Years i Southern Colorado 80-90 Same 5 to 30 8' 6" to 24" 2 Colorado (other Districts) . . 60-65 Same 5 to 30 8' 6" 18" to 24" 3 Colorado (other Districts) . . 75-80 Better 10 to 35 3' to V Few places 4 Saginaw District, Michigan. 65, 80-90 Better 15 3' 1 of 10 5 Central Illinois 50 Same 20 8' None 6 Southern Illinois 65-70 Same 20 8' Yes 7 Springfield District, Illinois. 55-75 Increase 4 20 to 25 6' to 7' Some places 8 Franklin, WiUamson and Saline Counties, 111 55-75 Increase 4 18 5*' Some places .9 Sherrard Field, 111 90 Same 20 3' 8" None 10 Extreme Southwest Section, Pennsylvania 72* Better 3 T 6" 10" 11 Pennsylvania-Westmoreland Co 84 Below 25 to 35 6' 8" None 11 Pennsylvania -Somerset Field 95 Same 8 3' 11" None 13 Maryland-Georges Creek Field . ... 97 Same 12 3' 0" None 14 Maryland-Georges Creek Field 88 Same 94 9' 0" 18" 15 Ohio, Belmont County. . . . 60 Same 5' 6" None 16 Eastern Ohio, Harrison County 70-75 Same 3' 8" to 5' 0" None 17 West Virginia 90 Same 50 8' Some places 18 Alabama 19 Tennessee 20 Kentucky No reply 21 Kansas 22 Iowa 1 Length of room 150 feet. 2 Advocates retreat mining. 3 No. 6 seam. MINING COSTS 169 "Big Vein" or Pittsburgh seam, show 88 per cent. Consider- able propping is necessary, owing to the drawslate and the wild coal just above it. The systems of mining the coal in this field have changed from time to time until now headings are driven 9 ft. wide, rooms only 13 ft. wide and the distance between room centers maintained at 100 ft., thus providing against squeezes. Under this process, 90 per cent extraction is expected. In Ohio, as well as some parts of West Virginia, no attempt is made to draw pillars at all. Rooms are driven 25 ft. wide RECOVERY OF COAL IN DIFFERENT DISTRICTS Nature of Top Nature of Bottom System of Mining Are Pillars Drawn Clay Veins En- countered Slate Very soft Soft slate Soft slate Room and pillar Room and pillar Yes Yes Dikes None Sandstone, poor shale Same s s top Room and pillar Yes None Black slate . Fire clay Room and pillar Where allowed None i Slate, clod and limestone Fire clay Panel system None None 2 Sandy shale Fire clay Panel system To some extent None 3 Hard shale Fire clay Room and pillar Where allowed None Hard shale Fire clay Room and pillar Where allowed None Blue rock and cap rock Slate and sand rock Room and pillar Yes Yes 18" to 3' ' draw slate Soft fire clay Room and pillar Yes None Slate Fire clay Room and pillar Yes None Hard black slate Limestone Room and pillar Yes Very few Sand rock Sand rock Room and pillar Yes Yes {Gray shale, coal, } dark shale J Hard gray shale Room and pillar Yes None Slate and shale Fire clay Room and pillar None None 10" firm slate Fire clay Room and pillar None None Varies Fire clay Room and pillar Yes Yes 4 Adjustment of labor situation. Projected work adhered to. 1 Sewickley seam. 170 COAL MINING COSTS with 8 to 12-ft. pillars between. In one of the largest mines of Belmont County, rooms were driven from both sides of the headings and it was no infrequent occurrence to have a ter- ritory squeeze shut, leaving considerable blocks of coal between the ends of unfinished rooms. In this mine, 50 per cent would approximate the recovery. In Harrison County, Ohio, it is nearly as bad. The recov- ery is reported as 70 to 75 per cent, but the same conditions exist in this section as in Belmont County, excepting perhaps the driving of rooms both ways from the same entry. The Ohio Mining Commission found in its investigations that 30, 40 and as high as 50 per cent of coal is being left underground as pillars in that state. There are mines in West Virginia which show recovery from 85.6 per cent to 99.8 per cent, the highest percentage resulting from the fact that all the work was in the solid. The average result of the figures presented for 10 mines showed about 92.6 per cent. The foregoing figures reveal what is possible, at the same time showing what is actually, presumably, being accomplished. It would not be proper to accept an average of the per- centages here given as a fair maximum, nor even an average of the same field, as it is unfair to compare ultimate recovery of mines now drawing to a close with those at the best of their production. No doubt, the systems under which they were inaugurated were considered modern, but they would not be considered so now. From reports sent in, it is apparent that there are five factors limiting the possible recovery in these fields as follows : 1. Mining rights and public feeling. 2. Roof and bottom conditions. 3. Weight and character of overburden. 4. Labor conditions. 5. Market value of the coal. 1. Where the mining rights do not allow breaking of sur- face, the recovery naturally varies inversely in some ratio to the overlying weight. 2. Where roof and bottom conditions make it necessary to recover as quickly as possible, market conditions will affect recovery for pillar work of this kind will not wait. MINING COSTS 171 3. Weight and character of overburden require systematic mining and competent supervision. 4. It is a matter of what is next best when unions insist on conditions which increase both cost of operating and loss of coal. 5. The market value of the coal dictates how far it is pos- sible to go toward its recovery. These points are mentioned because there is a tendency to compare straight figures of recovery without taking into con- sideration the conditions under which they are derived. The Ohio Mining Commission, for instance, uses the mines and operations at Gary for an example of what Ohio should follow. Conditions, however, are so different in these two localities, that to secure the same results in recovery would require several radical changes. Generally the roof in Ohio is poor, union scales require rooms entirely too wide for economic pillar drawing, the general labor situation is always more or less unsettled and the selling qualities of the coal are inferior to those of the Pocahontas seam at Gary. Many of the difficulties incident to the adoption of adequate conservation measures were described in a paper presented before the West Virginia Mining Institute in 1908. There are four factors, any one of which is sufficient to cause a serious loss in the percentage of coal recovered and an increase in cost of production, reducing the ultimate earnings of the property : First Insufficient or incompetent engineering : Until very recently the engineer was regarded by many coal operators as a luxury and an unnecessary refinement. This unreasonable conservatism, or prejudice, still exists to some extent; but the rapid depletion of properties which have been regarded for many years as practically inexhaustible is finally bringing the operator to a realization of the necessity for the careful plan- ning and scientific projection of his mine by a competent and sufficient engineering force. Second Incompetent management: There are some mine managers or superintendents who produce better results from a poorly designed mine than others can obtain from a first-class plant. Many managers have been entirely satisfied with a low cost-sheet, neglecting the conditions both inside and outside the mine, which ultimately resulted in an abnormal increase in 172 COAL MINING COSTS cost of production or abandonment of valuable acreages of coal, in order to maintain lower cost. This method would be repeated until, finally, after millions of tons of coal had been ruthlessly buried, which could have been recovered by a prudent and careful manager without materially affecting his cost of production, he was brought to a sudden realization of an unnecessarily high cost and a diminished tonnage, result- ing in loss of prestige and position for him and thousands of dollars to the owners. The necessity for a mine manager to be familiar with not only the outside, but also the inside, con- dition of the property, either personally or through tried and competent assistants, cannot be underestimated. Unless he is thoroughly familiar with these conditions, how can he know if the individual mine superintendent or foreman is giving him the desired results of low cost with maximum recovery, and the best conditions for a continuation of that low cost and recovery? Third Unfavorable labor conditions : The employment of unskilled miners renders it impossible to obtain a good recov- ery. Strikes and shut-downs have often interfered with the application of economic methods of extracting coal. Labor unions sometimes insist on conditions which, while operating for the convenience of the miner, increase the expense or induce an unnecessary loss of coal to the operator. For instance, in some districts of the country the track must be laid in the center of the room, with the gob on either side. Few or none of the pillars are recovered, and many acres of coal have been lost through squeezes and creeps because of wide rooms and small pillars. Nor is interference with the methods of mining the only manner in which the unions sometimes operate against the maximum recovery of coal. There have been men dis- charged for incompetency who were reinstated by the manage- ment on demand of the union. It is thus that discipline, which is such an important factor in the economic administration of a mine with a view to the best ultimate results, is destroyed. Fourth Impatience of owners for quick returns on invest- ment: The deleterious effect on the economic development of a property by the demand of the investor for an immediate profit can hardly be overestimated. Many a rich and valuable property has been irreparably damaged by the insistence of MINING COSTS 173 owners for immediate and large profits before its proposed economic development has been fairly launched. This impa- tience and greed has at times resulted in the changing of slow, but good plans of development for bad ones, and the poor results thus obtained were further accentuated in later years by the demand for big tonnage, thereby causing the loss of millions of tons of coal and thousands of dollars to the investor. In- deed, this demand for large tonnage from a poorly developed mine has been the greatest factor in encouraging careless methods of recovering coal both from rooms and pillars, and it is no exaggeration to state, that the abandoning of many acres of good coal, which could have been recovered by more thorough and known methods, can be traced directly to this cause. Use of the longwall system to effect conservation. The adoption of the longwall system of mining, where possible, will be the ultimate solution to obtaining the maximum recovery of coal and the subject is, therefore, one for the serious considera- tion of the coal economist. To obtain a comparison of the results of the different sys- tems a 1000-acre tract of land with 6 ft. of coal lying at a depth of 400 ft. will be assumed. This depth is taken to make allowance for the possibilities of the longwall system on sur- face caving, one of its chief disadvantages. A conservative estimate of the recovery to be obtained from such a tract under present systems of mining would be about 60 per cent. The total tonnage in this acreage would be 9,680,000 short tons, 60 per cent of which would be 5,800,000 tons which means a loss of nearly 4,000,000 tons. At a value of $1 per ton this would mean a loss in the national wealth of $4,000,000. In removing the entire seam some damage has perhaps resulted to the farmer, but not much, certainly, at a depth of 400 ft., and nothing beyond easy repairs. Where a total extrac- tion has been effected on a 5-ft. seam having 100 ft. cover in certain of the Pennsylvania fields, the break has extended to the grass roots. Where this extraction has "been several acres in extent and the break has been general over the entire area at once, it cannot be said that any appreciable damage has resulted. In this instance there was no packing whatever, as 174 COAL MINING COSTS would ordinarily be the case in longwall workings. The con- ditions were simply 100 ft. of easily breaking roof with a clean fall of 5 ft. Taking our previous example again, of 6 ft. of coal at a depth of 400 ft. worked by the longwall system and packed carefully, the result would by no means be so serious. Further- more, it should not be forgotten that there are approximately 1000 sq. mi. of coal in which the seams are 600 ft. or more below the surface.* At this latter depth it is doubtful if the strata would break to the surface, for it has been shown in the British mines that at a depth of 700 or 800 ft. work can be carried on successfully under the sea. This proves that in average strata the highest fall reaches an apex well below that distance, and the miner who has had extensive experience in pillar drawing and is familiar with the quick oblique line that the ragged rock edges traverse toward a common juncture, will place the safety point much lower. Turning again to our example of a 1000-acre tract, and assuming this to have been worked by the longwall system which has resulted in certain damage to the surface, a com- parison of this surface damage, with the additional extraction obtained, may be made. Taking the mine on a royalty basis of 5c. a ton, the farmer has received $193,600 above what he would have received by the 60 per cent pillar-and-room method of working. This amounts to $193 per acre, or about twice as much as the average farm value of land. It should also be remembered that the expense of tipple erection, compressors and power plants and the general surface arrangement of a mine opened to develop a tract of 1000 acres, is the same for 60 per cent extraction as for 100. It is said that in longwall working continuous operation is necessary, in order to properly control the roof, but this applies nearly, if not quite as well, to room-and-pillar work, particu- larly when the mine contains water. The statement of a cer- tain large Western operator, a man who is both practical and theoretical, may be taken, apropos of this. He changed a num- ber of his mines to the longwall system, and during a strike *See Twenty-second Annual Report, U. S. Geological Survey, page 178. MINING COSTS 175 lasting over a period of about five months and a half, all of his mines were shut down. On the resumption of operations he had careful records kept of the relative cost of opening the longwall and the room-and-pillar mines. These records showed the cost of opening the room-and-pillar mines to be nine times that of the longwall. SECTION II SHAFT SINKING The investment involved in shaft sinking is heavier than that encountered with any other improvement. Mistakes can be neither rectified nor lived down. Time and first cost being the essence of the opening of a new property, important fea- tures are often sacrificed underground instead of on the sur- face, where future remodeling is practicable. It is, therefore, obvious that the preliminary engineering and estimating rela- tive to a shaft operation deserve serious study. In planning a shaft mine opening the following points come up for consideration: Avoidance of all unnecessary narrow- work at the bottom increasing the risk of loss from squeeze, compliance with all present and possible future legislation, minimum first and maintenance cost, the connecting up of the two shafts in the shortest time in order that a regular circula- tion of air may be obtained and the restrictions of the law limiting the number of men allowed in the mine before this is done complied with. The arrangement of the work should be such that during development cars may be placed with the utmost facility, that mining machines may be employed and steel timbering placed as the entries advance; that when operations begin, cars, supplies, waste, men, water, air, dam- aged equipment, etc., may be handled with safety, economy and speed; that protection is secured against coal-dust explo- sions, mine fires, flooding, freezing, etc. Operations. The universal method of shaft sinking in rock is to drill a number of holes in the bottom, charge them with dynamite and shoot them, and to load the broken rock by hand into buckets which are then hoisted out. When all the loose rock has been removed the process is repeated. Shafts are drilled on the " center-cut " principle. Eight or ten holes are drilled on a slant, separated at the top but con- verging, thus forming a wedge known as the "sump." "Re- 176 SHAFT SINKING 177 liever," or bench, holes are drilled back of the sump holes, each row being more nearly vertical; the end or outside holes point slightly away from the vertical and toward the wall line of the shaft. The sump is first shot and the broken rock removed or " mucked " out, forming a cavity into which the bench rounds can be successively shot. All muck should be removed before each succeeding round is shot. Two systems of drilling and mucking exist. In the first the holes for the entire cut sump and benches are drilled at one time, the sump is shot, and then the benches as required. In the second, the sump only is drilled and shot, and the benches are drilled while the sump is being mucked. The first plan is particularly applicable to small shafts and to circular shafts ; a rectangular or elliptical shape is needed to give room for simultaneous drilling and mucking. Fumeless, or gelatine, dynamite should in all cases be used for underground work. The fumes from ordinary glycerine dynamite make it impossible for the men to get back to work promptly after a shot. The strength of the dynamite used depends on the character of the rock, but 40-per-cent and 60- per cent gelatine are the most common strengths used. The number and depth of the holes and the quantities of powder loaded vary so greatly with the size of the shaft and the nature of the rock that no general rules can be stated. The systems actually used at several shafts were as follows: Shaft 13 X 26 ft., through Western Pennsylvania coal measures: Shale, slate, and limestone; horizontal stratifica- tion; 40-per-cent gelatine: Number Depth, Feet Inclina- tion with Vertical, Degrees Loaded with Pounds Sump . 8 10 45 4 Relievers Benches 8 8 8 8 30 3 2i End 8 8 10 back 2* Total charge 96 Average gain per cut, 6 ft. Average gain per week of 19 shifts, 24 ft. (no timber). Mucking and drilling simultaneous; 2 drills used on 1 bar, double. 178 COAL MINING COSTS Shaft 14 X 48 ft., through anthracite measures : Red sand- stone; stratification horizontal; 40-per-cent gelatine: Number Depth, Feet Inclina- tion, Degrees Loaded with, Pounds Sump 8 10 45 5 Relievers 8 8 30 4 Benches 24 8 10 3 End 8 8 10 back 3 Total charge per round 168 Average gain per cut, 6 ft. Average gain per week of 18 shifts, 16 ft. Mucking and drilling simultaneous; 2 drills used on 1 bar. Shaft 10 X 22 ft., through quartz conglomerate (Shawan- gunk grit) ; horizontal stratification, but very few bedding planes; 60-per-cent gelatine: Number Depth, Feet Inclina- tion, Degrees Loaded with, Pounds Sump . 8 10 45 31 Sump 4 8 o 31 Relievers 8 9 30 2 Benches 8 8 o 2 End 8 8 10 back 2 Total charge per round 94 Average gain per cut, 5 ft. Average gain per week of 20 shifts, 22 ft. Mucking and drilling simultaneous; 5 drills used on 2 bars. The four additional sump holes shown were used on account of extra hardness of the rock. Shaft elliptical, 19 ft. 4 in. X 33 ft., through West Virginia coal measures: Hard gray sandstones; 40-per-cent gelatine; horizontal stratification : SHAFT SINKING 179 Number Depth, Feet Inclina- tion, Degrees Loaded with Pounds Sump 10 12 45 5 Relievers 8 10 30 4 Benches 14 10 10 4 End 6 10 10 back 3 Total charge per round 156 Average gain per cut, 8 ft. Average gain per week of 20 shifts, 18 ft. Mucking and drilling simultaneous; 3 drills used on 1 long bar, 1 short bar. Shaft circular, 17 ft. diameter, through Hamilton and Mar- cellus shales: Rock distorted; stratification irregular; but about 45 deg. ; 60-per-cent gelatine : Number Depth, Feet Inclina- tion, Degrees Loaded with Pounds Sump 6 8 45 2i Relievers 8 6 20 14 Rib 16 6 10 back 1 Total charge per round 43 Average gain per cut, 5| ft. Average gain per week of 19 shifts, 33 ft. All drilling on one shift, mucking on two shifts; 5 drills used on 5 tripods. While the hand drilling has been displaced almost entirely by power driven drills, there are still occasions when a small job does not justify the installation of power equipment and it is more economical to resort to hand drilling. This is par- ticularly so in foreign countries where labor costs are fre- quently quite low. The customary procedure is the use of a 1-in. drill, turned by one man and struck by one or two others with 8-lb. hammers. Two strikers should always be used where practicable as they can obviously drill twice as fast as a single striker at three- fourths the cost, Three capable men can drill 1^-in. holes in hard sandstone at the rate of 2 ft, per hour. 180 COAL MINING COSTS In soft material, churn drills 6 to 12 ft. long with a bit at each end are sometimes used with satisfactory results. These drills are sometimes weighted to give additional striking power and they are usually operated by two or three men. Shaft sinking is usually carried on 24 hr. a day. The inside work is done by three shifts of men working 8 hr. each, the outside by three 8-hr, or two 12-hr shifts. The 12-hr, outside shift is customary in the coal fields; elsewhere, the 8-hr, shift for every one is prevalent. Shifts are usually changed at 7 a.m. and 3 and 11 p.m., sometimes an hour later. The men are given 20 min. for lunch in the middle of each shift. Wages vary with the locality, but in general men are paid better for drilling and mucking in a shaft than in any other kind of rock excavation. On account of the high wages paid in America machine drilling is universal, and the shifts are limited to the number of men that can be worked to the best advantage. Speed is not attempted at the expense of efficiency. In South Africa, on the other hand, Kaffir labor is cheap, hand drilling is usual, and as many men are worked as the shafts will hold. : The great clepth of the shafts on the Rand makes the high- est possible speed desirable, even at an increased cost. In both countries speed is increased without an increase of cost by the payment of a bonus to the sinkers as a reward for additional progress. The size of the shifts for any given shaft depends upon the number of drills required and upon the experience and ability of the sinkers obtainable. With first-class men, the men on each shift for a 13 X 26 ft. shaft would be as follows, wages as of 1909 : Inside men, 8 hr. : One shiftboss, at $3; two drillers, at $2.75; two helpers, at $2.50; six muckers, at $2.25. Outside men, 12 hr. : One engineer; one head tender; three car-men on dump ; one firemen ; one compressor man. General outside, 10 hr.: One foreman; one mechanic; two carpenters (on timber) ; one blacksmith and helper. A 17-ft. circular shaft would require: Drilling shift: One shiftboss; five drillers; five helpers; one extra man. Mucking shift; One shiftboss; nine muckers. SHAFT SINKING 181 Outside: Same as above. In South African shafts, which are usually 9 X 26 ft., drill- ing is always done by hand and each shift consists one white shiftboss and 35 Kaffir laborers who drill or muck as may be required. Thorough organization is essential to progress and economy. Each man must know his place and take it without losing time in getting started. Any system that prevents systematic work is fatal to economy. Circular or Rectangular. From a construction standpoint the circular or elliptical and rectangular types are equally feasible, and the choice depends upon the cost. In several cases a compromise has been effected by shaping the shaft as a quadrilateral with sides formed of circular arcs. For a single compartment air-shaft the circular shape is in every way the most desirable, not only because the circular shaft is cheaper to sink than a square shaft of equal area, but also because a circular ring of plain concrete is the strongest lining possible with a given amount of material. In the case of a shaft with two or more compartments, the selection of the most economical shape requires some calcula- tion. At first sight it would seem that a simple rectangular shaft surrounded by a concrete wall only thick enough to be as strong as the usual timber lining, would be a satisfactory, as well as a cheap, shape, but this is not the case. A concrete lining, even when provided with weep holes, must resist some hydrostatic pressure ; a timber lining has none to resist. Fur- thermore, permanent weep holes are most undesirable; the concrete should exclude the water entirely, and hence must be designed to bear very great pressure at considerable depth. Just what amount the theoretical pressure is reduced, by the adhesion of the concrete to the shaft walls and by the block- ing of the fissures with grout, cannot be calculated. In solid rock, where the water enters in well-defined springs, the proper grouting of the springs will relieve the lining of all pressure. In very seamy rock, on the other hand, the lining may have to bear practically the full hydrostatic pressure. In order to compare the costs of the different shapes, let us consider in detail three designs for a shaft with two 7 X 10 ft. hoistways and an airway with an area of 100 sq. ft. As the 182 COAL MINING COSTS QUANTITIES AND COSTS OF RECTANGULAR SHAFT Depth in feet 20 50 100 150 2ftft Total thickness of lining in inches . . . 14 21 28 34 MHI 39 Quantities per linear foot: Concrete to neat line in cubic yards. 3.90 5.70 7.60 9.30 10.70 Concrete actual in cubic yards 5.80 7.70 9.70 11.50 13.00 Excavation to neat line in cubic yards 12.80 14.60 16.50 18.20 19.70 Excavation actual in cubic yards . . 14.70 16.60 18.60 20.40 22.00 Weight reinforcing steel in pounds. 256 443 650 845 1030 Cost per linear foot: Forms $25.00 $25 . 00 $25 . 00 $25.00 eof) nn Concrete at $5 cubic yard 29.00 38.50 48.50 57.50 O~> . UU K Oft Excavation (see note *) 49.60 53.20 57.00 60.40 oo . uu AQ Aft Reinforcing steel at $0.02 pound. . 5.10 8.90 13.00 16.90 OO . ^U 20.60 Total $108.70 $125.60 $143.50 $159.80 $174 OO IP 1 1 1 . UU QUANTITIES AND COST OF ELLIPTICAL SHAFT Depth in feet, to 100 150 200 250 300 400 Thickness of lining in inches, ends .... 12 12 12 12 12 12 Thickness of lining in inches, sides .... 12 18 24 29 34 42 Quantities per linear foot: Concrete to neat line, cubic yards. . 2.60 3.40 4.30 5.00 5.70 6.80 Concrete actual in cubic yards 4.40 5.20 6.10 6.80 7.50 8.60 Excavation to neat line in cubic yards. $15.20 $16.00 $16.90 $17.60 $18.30 $19.40 Excavation actual in cubic yards. . . 17.00 17.80 18.70 19.40 20.10 21.20 Costs per linear foot: 15 00 15 00 ifi on 15 00 15 00 15 00 Concrete at $5 cubic yard 22.00 26.00 30.50 34 00 37 00 43 00 54 40 56 00 57 80 59 20 60 60 62 80 Total $91.40 $97.00 $103.30 $108.20 $113.10 $120.80 QUANTITIES AND COSTS OF QUADRILATERAL SHAFT Depth in feet to 100 12 2.70 4.50 14.90 16.70 $15.00 22.50 53.80 150 19 4.40 6.30 16.60 18.50 $15.00 31.50 57.20 200 26 6.20 8.20 18.40 20.40 $15.00 41.00 60.80 250 32 7.90 10.00 20.10 22.20 $15.00 50.00 64.20 300 39 9.90 12.10 22.10 24.30 $15.00 60.50 68.20 400 52 13.90 16.20 26.10 28.40 $15.00 81.00 76.20 Thickness of lining in inches Quantities per linear foot: Concrete to neat line in cubic yards . Concrete actual in cubic yards Excavation to neat line in cubic yards. Excavation actual in cubic yards . . . Costs per linear foot: Forms Concrete at $5 cubic yard Excavation (see note *) Total $91.30 $103.70 $116.80 $129.20 $143.70 $172.20 * Cost of excavation figured on basis of $4 per cubic yard for section containing 12 yards per linear foot; additional excavation at $2 per cubic yard. Thus cost of 16 cubic yard section = 12 X $4 +4 X $2 = $56. SHAFT SINKING 183 whole area of a hoist shaft is ordinarily used for the passage of air ; the size of the air compartment may be reduced if the rest of the shaft is enlarged; the airway must however be large enough to contain pipes and ladders and to provide in addition an ample passage for air if the hoistways are tem- porarily closed. Let us assume a minimum thickness of 12 in. of concrete for a water-tight lining ; also that in each case the lining carries the entire hydrostatic pressure ; then the specifications for the three forms of shafts will be as follows: Rectangular Shaft. Fig. 1. Two hoistways 7 X 10 ft., one airway 10 X 10 ft. Ten-inch concrete dividing walls in place of buntons. Extreme inside dimensions 10 X 25 ft. 8 in. Area .1 2 ! j'-~ . > .-/o'-0". . so * ^,'-0- 1 . FIG. 1. Rectangular concrete-lined shaft. airway 100 sq. ft., total clear area 240 sq. ft. Thickness of lining at any point made equal to depth of simple beam of 10 ft. span required to sustain hydrostatic pressure at that point. Resisting moment and weight of reinforcement calculated by Johnson's formula, factor of safety 3. (Ultimate tensile strength of steel 65,000 Ib. per square inch, compressive strength of concrete in beam 2500 per square inch.) Reinforcing steel set 3 in. inside of face of wall. Cost of forms, given in the accompanying table, includes cost of forms for dividing walls, and is therefore greater than the cost in the elliptical shafts. Excess of actual over theoretical quantity of excavation is estimated as 15 per cent for 28-ft. shaft. This excess increases with the length of the shaft only, as the ends are drilled to line. Elliptical Shaft. Fig. 2. Extreme inside dimensions 16 X 27 ft. Area of airway, 78 sq. ft. Total clear area, allowing for 10-in. buntons, 304 sq. ft. 184 COAL MINING COSTS Strength of lining calculated on the assumption that the stress in the elliptical cylinder at any point is equal to that caused in a circular cylinder with a radius equal to the radius of curvature of the ellipse at the given point, by the same hydrostatic pressure acting upon it. The lining is therefore made thicker at the sides than at the ends. To prove this proposition assume the lining to be constructed of a number of small portions, each the arc of a circle. The stress in each portion caused by the hydrostatic pressure of the film of water between it and the rock is directly proportional to the radius, and the thickness of each section should therefore be made proportional to the radius. Considering any portion, FIG. 2. Elliptical concrete-lined shaft. as o-&, Fig. 4, the skewback toward the side of the ellipse is formed entirely by the adjoining portion, while the skewback toward the end is formed partly by the adjoining portion and partly by the rock. If the number of circular portions is in- definitely increased, the unbalanced end thrust of each will be taken up by the irregularities of the rock. Ultimate compressive strength of concrete, 3000 Ib. per square inch ; factor of safety, 3. Excess of actual over theoretical excavation assumed as 12 per cent for smallest section. As the length of the shaft does not vary, this excess is constant. Quadrilaterial Shaft. Fig. 3. Inside dimensions, 16 X 24 ft. 8 in. Radius of ends and sides, 23 ft. Area of airway, 94 sq.ft. Total clear area, allowing for 10-in. buntons, 294 sq. ft. For calculating stresses, sides and ends are considered as SHAFT SINKING 185 portions of a 46-ft. circular cylinder. Ultimate compressive strength of concrete 3000 Ib. per square inch, factor of safety, 3. Excess of actual over theoretical quantity of excavation assumed to be 12 per cent for minimum length and to increase with the length. FIG. 3. Quadrilateral concrete-lined shaft. Germany is committed to the circular or elliptical form of shaft, the German engineers being of the opinion that the square or rectangular form is more expensive due to the extra work involved in excavating and keeping the corners squared up. FIG. 4. Design of the elliptical form of shaft. It is evident, that assuming the same hoisting capacity in either form of shaft, the excess area, which makes ventilation possible, should be the same in either a circular or a rectangular shaft. A circular shaft of 20 ft. net diameter would be roughly 186 COAL MINING COSTS equivalent to a rectangular shaft 12 X 20 ft. English mining engineers claim that the cost of lining is as 5 to 9 in favor of circular shafts, and it is generally conceded that where great pressure is encountered the circular form is the only safe one. A circular shaft, when once properly lined with iron or masonry, is a permanent affair, while timber lining under the best conditions cannot be expected to last more than 18 or 20 yr. and rarely more than 15 yr. It is also well known that for a given area, a circular shaft presents less rubbing sur- face, or resistance to the passage of the ventilating current, and the segments at the side of the cages furnish space for this air current without additional enlargement of the shaft. The principal arguments advanced for rectangular shafts are that less material needs to be removed for a given cage space, and that in sinking, the permanent lining is at once put in place, as the work progresses. It is probable that the matter of keeping a shaft in aline- ment during construction, by either method, is largely a matter of experience and character of labor employed. While engineers have claimed cheapness of construction as an argument for both forms of shaft, the data at hand would indicate that the circular shaft may be excavated fully as cheaply as the rectangular, but costs more to line; while on the other hand the upkeep and repairs on a circular shaft properly constructed, are very much less than on a rectangular shaft of same capacity, and the danger, especially in deep mines or in quicksand, is very materially reduced, and water much more easily kept out, effecting a saving in pumping. Equipment. A two-stage compressor is the best for shaft sinking. With a steam consumption of 45 Ib. per i.h.p. at half cut-off the simple compressor has for each indicated horse- power, a capacity of 5 cu. ft. of free air per minute compressed to 100 Ib., while the two-stage machine will deliver 15 per cent more air with the same steam consumption. For 500 cu. ft. of free air per minute, the saving of the two-stage, over the simple type, will amount to 15 per cent X i X 500 cu. ft. X 45 =675 Ib. steam or 150 Ib. of coal per hour. With the com- pressor operating to capacity 20 hr. a day, 6 days in the week, the saving in three months with coal at $4.50 per ton would thus be $525. SHAFT SINKING 187 The cost of a plant for a single shaft, assuming a depth of about 500 ft. and a moderate inflow of water, say 30 or 40 gal. a min, was estimated in 1909 as follows : Sinking engine $1,000 Two 80-hp. boilers and setting 1,800 Pipe and auxiliaries 500 150-hp. heater 300 14-in. compressor 1,750 Three drills and steel 1,000 Shaft bar and clamps '. 100 Derrick 400 Head-frame 500 Two buckets 150 Rope 150 Buildings 500 Dump cars and rail 300 Electric plant, 10 kilowatts 750 Two pumps 500 Small tools.. 500 Total $10,200 These figures are based on the cost of new machinery, and are large enough to include the necessary accessories. The cost of erecting and dismantling such a plant will be from $1000 to $2000, depending on location, labor conditions, etc. Some of the most serious pumping problems in shaft sink- ing in this country have been encountered in the Pocahontas field along the Norfolk & Western Ry. Except in two instances, sinking has been very expensive in this territory owing, it is thought to the fact that the water-bearing rock is a very hard sandstone carrying a considerable volume of water. When a shaft is started beyond the toe .of the mountain and in the valleys the sinking always encounters plenty of water, but in the two exceptions noted the location of the shaft in each case was back of the toe of the mountains and away from the valleys. The quantity of water in the wet workings, varies from 200 to 1100 gals, per min. The heavy charge of high explosive necessary to break the dense sandstone, coupled with the large volume of water, gives the worker constant pump trouble; moving the pumps during the blasting is out of the question, because the water would flow in so fast that the 188 COAL MINING COSTS pumps could not be replaced for service. Having this constant pump trouble makes progress very slow, reducing it to 10 ft. per month in several instances. The following is a list of some of the shafts and amount of water encountered: The Middlestate Coal Co. started an old shaft, abandoned 90 ft. down, which flowed 1100 gals, of water per minute, and they were about 18 months getting down the additional 90 ft. The Pocahontas Collieries' 16 X 32 ft. shaft, at Boissvain, was 11 months in sinking 200 ft., hindered as it was by 500 gals, of water per minute. The most expensive shaft proposition was the shaft for the Jed Coal & Coke Co. about 2 miles from Welch. This property's main shaft encountered 1100 gals, per min. while its air shaft met 500 gals. The coal expense for power has been as high as $1500 per month with coal at about $1.25 per ton. Sinking costs. The Nokomis Coal Co. sunk a shaft through the soft Illinois shales in 1913, the cost figures on which are of interest. The shaft was 631 ft. deep, timber lined and 17 ft. 5 in. X 11 ft. 5 in. inside the timbers, with an airshaft of the same dimensions 500 ft. distant. Eight hand-feed hammer drills, weighing 40 Ib. each were used on the work and com- pressed air was furnished by a 9 X 10 X 12-in. compressor supplying 174 cu. ft. of free air per minute at a terminal pres- sure of 100 Ib. per square inch. After passing through the upper capping of hard rock, shale of various degrees of hardness was encountered with an occasional layer of limestone 10 or 12 ft. thick. The rock at times consisted of slaty bands, sandy shale or soft gray material, more like indurated clay than shale. In sinking through limestone, from 28 to 32 holes constituted a round, the corner holes being bottomed 2 in. outside the line of curbing. As the shafts were timbered throughout, the break lines were 12 ft. 5 in. by 18 ft. 5 in. The holes were 4y 2 ft. deep and were connected and fired with an electric battery in the ordinary manner. Three 8-hr, shifts were worked, a round being drilled, blasted and mucked out on each shift. Four drillers, four muckers and a shift leader or boss con- stituted a sinking crew. It was found that in trying to bore holes in the soit shale with these hammer drills, the tool cut so rapidly as to choke SHAFT SINKING 189 the passage in the bit with muck, stopping the flow of exhaust air and preventing proper cleaning of the drill holes. Hand drilling was temporarily substituted, but on adopting certain recommendations of the manufacturers, the hammer drills worked satisfactorily, and their use was resumed with a marked increase in speed over the hand work. These changes gave the following results in soft shale, as compared with hand drilling. Hand drillers using 2 1 / 4-in. steel worked at the rate of 4y 2 ft. per man per hour, or four men put in a round of 54 ft. in 3 hr. With the air drill, four men drilled 18.9 ft. per man per hour, or a round of 54 ft. in 45 min. thus accomplishing a saving of 2 hr. and 15 min. per 54-ft round. Owing to the variation in the time required for the shoot- ing and mucking, this increase in drilling speed meant an increase in depth per day of 1 ft. or 4% ft. per 24 hr., with hand drills. Sinkers, including drillers and muckers, were paid $3.39 per 8-hr, shift, the shiftboss receiving $4, making a total labor cost per day of roughly $93, or $26.50 per foot by hand drilling. The saving by using the hammer drills therefore, amounted to one foot in each shaft or $53 per 24 hr. for both shafts. Interesting data on the comparative cost of sinking a ver- tical shaft and an inclined slope to accomplish the same pur- pose were disclosed in a request for tentative bids made to two leading contracting companies early in 1921. The conditions under which the work was to be done were as follows : The shaft to be in Ohio, 5 miles from present railroad. Fuel, supplies, and equipment to be hauled in by the com- pany. Company will furnish all timber. Strata shale and sand- stones. Plenty of water available. Company will provide housing accommodations. Contractors to furnish all equipment except, ties, rails, and such supplies as can be used in perma- nent mine equipment. Waste disposal within radius of 500 ft. Assume an average amount of water in sinking. First condition. Shaft three compartments 10 X 26 ft. in clear 170 ft. deep to coal. Timbered with 8 X 10 in. sets on 5 ft. centers with 2 in. lagging. Second condition. Single track slope on 30 deg. pitch 5 ft. clear above top of rail by 7 ft. wide, 340 ft. long, timbered 190 COAL MINING COSTS with 6 X 8 in. crossbars and posts with sets spaced 5 ft. centers and lagged with 2 in. planks for 170 ft. Airshaft within 500 ft. of above slope mouth 10 X 10 ft. and 170 ft. deep timbered and lagged for 100 ft. with 8 X 10 in. sets on 5 ft. centers. Third condition. Same as first paragraph of second con- dition except duplicate slopes with 35 ft. horizontal pillar be- tween. No air shaft. Fourth condition. Same as second condition except double track slope 12 ft. wide. Quote both with and without air- shaft. The first company submitted bids as follows: The price of the shaft in the first condition will be $200 per vertical foot timbered. Under the second condition, the price of the single track slope timbered will be in the neighborhood of $100 per lineal foot. An airshaft under this condition would cost in the neighborhood of $140 per vertical foot timbered and slopes under the third condition would be $100 each. Under the fourth condition, the double track slope would cost approximately $125 per lineal foot sunk at the same time as an airshaft is being sunk. If no other opening is sunk at the same time, that is, with the same main plant and the same overhead, the slope would cost $140 per lineal foot. The price for cement used as grout will run in the neighbor- hood of $12 per barrel. The bids of the second company, which were gross, were as follows: First condition. Three compartment shaft, 10 X 26 ft. in the clear, 170 ft. deep, to cost $45,500. Second Condition. Single track slope, 5 ft. clear above top rail, 7 ft. wide, 340 ft. long and airshaft 10 X 10 ft., to cost $42,500. Third condition. Same as second condition, except duplicate slopes, no airshaft, to cost $42,400. Fourth condition. Same as second condition, except double track slope, 12 ft. wide with airshaft, $47,000 ; without airshaft $33,250. There are certain items entering into the cost of such a short job as this which run the cost up considerably; for SHAFT SINKING 191 instance, the item of transportation in the first proposition, would make a cost of about $8 per foot of shaft. Sinking costs given below were fairly representative for the different methods of work in 1909 : Shaft excavated 14 X 20y 2 ft. through 6 ft. of soil and 14 ft. of quicksand, not very wet. Sides supported by 2-in. oak sheeting driven by mauls and braced by five sets of 10 X 12 in. timber : Per Foot Per Cubic Yard Labor $27 25 $2 57 Lumber, 6600 ft. B.M. at $30 9 90 93 Erection of derrick, etc 3 00 29 Superintendence 3 00 29 Sundry 2 00 18 Coal and pumping 5 00 47 Total $50 15 $4 73 Shaft excavated 12 X 20 ft. 3 in. through 45 ft. of clay and gravel. Sides supported by sets of 10 X 10 in. pine timber spaced 4% ft. centers and hung from top. IV^-ni- lagging: Per Foot Per Cubic Yard Labor $19 50 $2 17 Lumber, 240 ft. per foot at $25 Bolts, 15 Ib. per foot at $0.03 Erection of head-frame, etc 6.00 .45 2 00 .66 .05 22 Superintendence .... 2 00 22 Power . 1 50 17 Sundry 1 00 11 Total $32 45 $3 60 Shaft excavated 15 X 37 ft. through 21 ft. of dry sand. Sides supported by interlocking steel sheet piling driven with steam hammer and braced with sets of 8 X 10 in - timber : 192 COAL MINING COSTS Labor Costs Only Per Foot Per Cubic Yard Driving sheeting $6.55 $0 32 Removing sheeting Timbering 1.85 2 05 .09 10 Fixcavation 8.20 40 Total . . $18.65 $0.91 Caisson 26 ft. outside diameter, 21 ft. inside diameter, sunk through 56 ft. semiliquid mud and boulders: Per Foot Per Cubic Yard Excavation (Materials $27 . 00 $1.35 Labor 7 00 35 Forms and shoe . 23.00 1.15 Sinking caisson 38 00 1.90 Plant erection 3 00 15 Superintendence 5.00 .25 Sundry 5 00 25 Coal and power 6.00 .30 Total $114.00 $5.70 The H. C. Frick Coke Co. put down two shafts nearly 600 ft. deep near Brownsville, Pa., about 1909 that presented some interesting cost data. The main shaft is elliptical in shape and 13 X 28 ft. in the clear, the inside circumference being 69 ft. with a clear opening of 310 sq. ft. The airshaft is the same shape, 14 X 34 ft. on its main center lines, measures 81 ft. around the circumference and has a clear opening of 390 sq. ft. All concrete for lining the shafts is composed of one part Portland cement; two parts clean, sharp, river sand, and five parts of stone crushed to pass through a IVk-ni- ring. About 50 per cent of the stone used for concrete was obtained from the materials excavated from the shafts. About 30 per SHAFT SINKING 193 cent was shipped in, crushed ready for use, while about 20 per cent was obtained from a quarry on the grounds. The average amount of concrete to each batch was % cu. yd. Through solid strata the proportion of the mixture is one part cement, two parts sand, and five parts crushed stone, while through softer strata the amount of cement was increased 50 per cent, the other ingredients remaining the same. The mini- mum thickness of the concrete lining wall is 12 in. In soft strata the concrete is as much as 33 in. in thickness, as no voids were left between the rock and lining, all such being carefully filled with concrete. During the excavating of the shafts, blasting was done within 10 ft. of the concrete, but at no time did this blasting have any appreciable effect upon the concrete lining. Sixty- per-cent gelatine was used in the blasting, and the maximum charges were from 150 to 200 Ib. per shot. The holes in the "cut," or sump, were usually drilled to a depth of 12 ft., while those in the side rounds were 10 ft. The sinking and concreting were carried on separately. The work was pushed continuously from 12 o'clock Sunday nights until 12 o'clock the following Saturday nights. No Sunday work was done, except in rare cases, other than that necessary for pump operation. All men employed in the shafts worked 8-hr, shifts. No forms were removed under 72 hr., allowing ample time for the concrete to set. The average time for completing one 50-ft. section of the concrete lining wall occupied 6 days, during which operation the services of 9 men at the top and 4 men on the bottom platform were required. In sinking, 5 men were used on top and an average of 12 men on the bottom. An auxiliary hoisting engine was used in one compartment for the handling of the steel forms; the operation of placing and removing these forms required the services of 4 men at the top of the shaft and 6 men on the bottom. A special feature of the work was the construction of a 4-in curtain wall in the airshaft. The reinforcing in this wall consisted of %-in. round steel frames, 5 ft. high by 13 ft. 6 in. long, stiffened with two additional %-in. round steel placed equal distance from the ends. Around the bars forming the 194 COAL MINING COSTS frame, No. 10 gauge wire netting was laced, this netting having a 2-in. mesh. The concrete used in this curtain wall was made of one part of cement, and two parts of clean coarse river sand. In the excavating of the shafts a record was kept of the muck taken out and of the materials entering into the work. From the main hoisting shaft 21,168 buckets of muck were taken, while the ventilating shaft gave 28,442 buckets, approxi- mating 50,000 cu. yd. of loose excavation through earth, fire- clay, shale, slate, sandstone, limestone, and coal. The increase of the actual muck excavated, after the blasting, over the calculated yardage in the solid shows 135.2 per cent for all the materials passed through, or 35.2 per cent more than double the amount in the solid. There were used in the construction of the concrete lining in the shafts, 8217 barrels of cement, 4410 tons crushed stone and 2528 tons sand, made up into 12,951 batches of concrete, containing approximately 6500 cu. yd. Exclusive of the archways at the bottom landings, 539 ver- tical feet of concrete lining wall was placed in the hoisting shaft. The time to complete the same covered a period of 46 weeks from the start of the work of excavating, showing a progress of 48 ft. per month. Omitting archways at the bottom landing, 561 vertical feet of concrete lining was placed in the ventilating shaft. The time consumed in completing same covered a period of 62 weeks from the commencement of the work of excavating; a progress of 37.2 ft. per month, for the sinking, timbering, and placing of the concrete lining. Taking into consideration that this shaft has an area of 80 sq. ft. more than the hoisting shaft the progress of the work averaged about the same as at the other shaft. The first cost of concrete-lined shafts over that of timber lined is about one-third more, which amount would probably be spent upon the first renewal of timbers and from the aver- age run of timber now on the market this would likely be necessary in about 10 yr. In a concrete-lined shaft, the lining being indestructible, the only renewals required would be replacing of the buntons and guide rails from time to time. SHAFT SINKING 195 COMPARATIVE QUANTITIES IN THE CONCRETE LINING FOR THE VERTICAL FOOT OF SHAFT HOISTING SHAFT Calculated Yardage Actual Yardage of Material Placed in Work Increase of Actual Yardage over the Yardage Calculated Thickness, Inches Cubic Yards Cubic Yards Percentage 12 15 18 24 2.66 3.36 4.08 5.56 4.6 5.0 6.4 9.0 73 49 57 62 VENTILATING SHAFT 12 3.10 5.40 74 15 3.92 5.62 43 18 4.74 5.84 23 24 6.45 9.50 47 The following summary gives the principal figures in regard to the work on both shafts: HOIST SHAFT Total depth, feet 565 Total number of weeks worked 46 Average thickness of concrete lining, inches 16 Average number of batches per 5-ft. form 56 Average cubic yards concrete per 5-ft. form 28 Average depth of lining placed per week, feet 12 . 3 Average depth of sinking and lining placed per month, feet 48 VENTILATING SHAFT Total depth, feet 591 .2 Total number of weeks worked 62 Average depth sunk per week, feet 9.5 Average depth of lining placed per week, feet 9.1 Average depth of sinking and lining placed per month, feet 37 . 2 Size of shaft, inside concrete lining wall 14 ft. X 34 ft. Area, square feet . 390 Average thickness of concrete lining, inches 17 Average number of batches per 5-ft. panel 66 Average cubic yards of concrete per 5-ft. panel 33 196 COAL MINING COSTS Some interesting cost figures were obtained in the sinking of a 570-ft. inclined metal mining shaft in the Poverty Gulch region of Colorado about 1910. The shaft was 12 X 7.5 ft. in the clear and after it had been sunk a short distance two skips were added to the sinking equipment which weighed 550 Ib. each, held 10 cu. ft. and cost $70 a piece, including a water valve in the bottom which cost $10. The sinking equip- ment consisted of the following: Six Little Giant drills, 2|-in. diameter. $900.00 Six columns and arms up to 8 ft 240 . 00 Six sets drill steel, each 15.4 Ib., at $14.40 86 . 40 Buffalo exhaust fan, diameter outlet 24 in., price with bed and countershaft 700 . 00 460 ft. of 24-in. pipe, at $54 per 100 ft 248.40 1200 ft. of 12-in. pipe, at $27 per 100 ft 324 . 00 Ten Leyner No. 5 stoppers, $135 1350.00 Drill steel, 10 sets, at $15 150 . 00 Sixteen ore cars, at $45 720 . 00 Compressor, motor and pipe, freight 2689 . 14 Freight on 19,553 Ib., at $0.55 per 100 107.54 Total $7515.48 In the development work two drills were required. The drills are 2% i n - i n diameter and use air at 90 Ib. pressure per square inch at drill. According to the catalog specifications, each drill will need 67.2 cu. ft. of air per minute. The factor to determine a compressor capacity for four drills at 10,000 ft. altitude is 4.49; hence, 301.5 cu. ft. of air per minute will be required, but deducting 5 per cent for leakage and allowing the compressor a volumetric efficiency of 80 per cent, the total air required is nearly 400 cu. ft. per minute. E. A. Rix allows 20 hp. for every 100 cu. ft. of cylinder displacement, to compress air to 90 or 95 Ib. gauge pressure at sea level. Although 20 hp. is higher than the value given by Peele for the theoretical horsepower required, and figuring efficiency the figures would then be below 20; but, since com- pressors are usually purchased for excess power to supply possible additional uses, and the use of 20 hp. would only add a small percentage on the safe side, the power necessary for four drills is taken at 80 hp., and it was decided to purchase a two- SHAFT SINKING 197 stage air compressor 18 X 11 in. diameter, 12-in. stroke, 125 r.p.m., with a capacity of 440 cu. ft. at 10,000 ft. elevation. This sized compressor gives a reserve of 21 per cent, and costs with freight from Denver to Cripple Creek $1903.' The motor for the compressor is 80 hp. ; 900 r.p.m., 440 volts, and costs delivered $710.84. The following calculations give the power consumed during development by two piston drills: The catalog multiplier is 2.39 and as each drill will need 67.2 cu. ft. of air, 67.2 X 2.39 = 159. Allowing for air loss in pipe line and efficiency of compressor, 210 cu. ft. are necessary, or 42 hp. per minute. Each stope drill requires 25 cu. ft. of air per minute and the factor for seven drills is 7.55 ; therefore, 25 X 7 = 189 cu. ft., to which 63 cu. ft. is added to allow for loss and efficiency, and this is equivalent to 50.4 hp. The power for two stope drills is 25 X 2.5 (multiplier) = 62.5 cu. ft. of air, and if to this be added 20.7 cu. ft. for pipe loss and efficiency, the power required is 16.68 hp. The diameter of the pipe needed for carrying air 800 ft. is 3 in. and will cost at Cripple Creek $75.30. The total cost of com- pressor, motor, and pipe is $2,689.14. In sinking, 18, 4-ft. holes were used. Drilling was done at the rate of 39 ft. in 8 hr. per drill or 72 ft. in 7.4 hr. The rate of advance was 3 ft. per round which equals 3 X 12 X 7.5 = 270 cu. ft. solid material which divided by 12.4 gives 21.8 tons per round and multiplying this by 21.5 gives 469 cu. ft. or 17.4 cu. yd. of loose material. Estimating the cost of muck- ing on the basis of 1.2 cu. yd. per man per hour it will take two men 7% hr. to clean up the rock after each round, so that allowing % hr. for delays, changing buckets, etc. it will be seen that one round can be drilled, fired and mucked in 16 hr., which eliminating other delays would be equal to a progress of 90 ft. per month. For a 4-ft. hole it was found that six sticks of 40 per cent dynamite were required each weighing 0.6 Ib. or 3.6 Ib. per hole so that for the 18 holes 64.8 Ib. were required. The shaft was sunk 570 ft. The time required to sink the shaft was: 5-77 - -7 = 190 rounds, or days. 3 ft. per round 198 COAL MINING COSTS The detailed cost of sinking 570 ft. of a 90-sq. ft. inclined shaft is as follows: Two machine men, 190 shifts, at $4.50 $1,710.00 Two muckers (also top men) 190 shifts at $3 1,140 . 00 Two hoistmen, 190 shifts at $4.50 1,710.00 One blacksmith, 190 shifts at $4.50 855.00 One blacksmith helper, 190 shifts at $4 760 . 00 One foreman, 190 shifts at $4 . 50 855 . 00 One superintendent, at $175 per month, 6^ months 1,108. 35 One timberman, 190 shifts, at $3.50 665 .00 Powder, 12,312 lb., at $1 .27* 1,563 . 62 Fuse, 190 rounds, 7-ft. lengths, 23,940 ft., at $0 . 0035* 83 . 79 Caps, 3420, at $0.007* 23 .94 Depreciation on steel 14 . 40 Operation compressor plant (power), 336 hp. hours per day: Installation charge $20 . 50 40,000 kw. hr., at $0.013 520.00 7700 kw. hr., at $0.005 38.50 579.00 Timber, 95,440 ft., at $20 per thousand 1,908 . 80 Electric power, for hoist, average 6.23 hp. hr. per hour, 49.84 hp. hr. per day, 9470 kw. hr. (190 days) at $0 . 013 123 . 11 Coal for blacksmith, 28.75 tons, at $20.75 243 .20 Candles, 950, at $0. 0145 13 . 78 Rails (30-lb.), 570 ft., 5.08 tons, at $50 254.00 Cost for 570 ft $14,207.55 Or cost per ft $24.93 * These low costs of powder are those of the Portland Gold Mining Co. and include freight and unloading charges. The following costs of sinking a mine shaft through ande- site at the Esperanza Mine at El Oro, Mexico, are given by W. E. Hindry in the Mining and Scientific Press in 1910. The shaft was a three-compartment vertical shaft, having two 5X5 ft. hoisting compartments and a 5X7 ft. pump and ladderway. The timbering was 10 X 10 in. with 2-in. lagging ; sills 5 ft. center to center, and 6 posts per set. The total depth of the shaft was 679 ft., of which 101 ft. were sunk by wind- lass and hand work, and 578 ft. by steam hoist and machine drills. The work was done in 1899 and the prices of materials and wages were as follows : SHAFT SINKING 199 Materials Prices Timber per M ft. B.M $13.58 Wood per cord 3 . 15 Coal per ton 7.27 Powder, 60 per cent, per pound 0. 14f Fuse per foot 0.0055 Caps, each 0.0058 Candles, each 0.0194 Labor Superintendent, per 24 hr :'.'.. ' $4.850 Shaft men, foreign, per 8-hr, shift 3 . 220 Shaft men, native, per 8-hr, shift 0.528 Top men, per 8-hr, shift 0.422 Fireman, per 8-hr, shift 0.485 Hoistmen, per 8-hr, shift 0.970 Blacksmiths, per 8-hr, shift 1 .455 The cost of excavation was as follows : Labor Per Linear Foot Superintendence $2 . 529 Shaftmen, foreign 3 . 510 Shaftmen, native 7.043 Top men 0. 678 Blacksmiths 0.718 Firemen 0.317 Hoistmen 0. 894 Miscellaneous 0.936 Total $16.525 Materials Timber $3.961 Wood, fuel 3.781 Coal 0. 179 Powder 2.853 Fuse . 014 Caps 0.294 Candles 0.223 Oil, grease, etc 0.025 Miscellaneous . 051 Total $11.381 Grand total.. $27.906 200 COAL MINING COSTS The above costs are converted from Mexican money assum- ing the peso to have a value of 48^c. An exhaustive study of shaft sinking costs in the Michigan region was prepared about 1910. The figures and computa- tions were made upon the assumption that the shaft would be 6 X 16 ft. within timbers and reach a vertical depth of 1000 ft. For a vertical shaft, its total length would be 1000 ft. ; if in- clined at an angle of 45 deg. to follow the dip of the formation, its total length would be 1400 ft. to reach a total vertical depth of 1000 ft. A contract price of $40 per foot for sinking would apply only if the shaft was put down in the jasper formation. If diorite was encountered the contract price for sinking alone would be between $50 and $55 per foot, while all other items would remain the same. A slight difference in cost of timber would appear in the amount of timber used in a vertical or inclined shaft, as the latter requires 10 X 10 in. stringers, while the former would take 6X8 in. skip runners, but this has not been taken into account. The maximum flow of water to be handled (800 gal. per min.) is probably somewhat high. DETAILS OP SINKING CONTRACT, SHAFT, ETC. Maximum Minimum Contract price for sinking, $40 per foot, includes drilling, blasting, powder, caps, fuse, etc. For a vertical shaft, 1000 ft For an inclined shaft, 1400 ft Computing all work for a shaft 6X16 ft. within timbers, three compartments, shaft sets, 5-ft. centers; and assuming 40 ft. per month as average sinking. 1000 ft. would take 25 months, allow 26 months. 1400 ft. would take 35 months, allow 36 months. Allowing 25 working days per month, 300 per year 1000 ft. of sinking would take 650 days. 1400 ft. of sinking would take 900 days. Sinking contract would be worked on three 8-hr. shifts. All other labor, one or two 10-hr, shifts. Cutting pump station 10X15X20 ft. at a depth of 500ft A maximum flow of 800 gal. per min. when bottom of shaft is approached is used as the basis of pumping expenses. Total.. $56,000 $40,000 350 350 $56,350 $40,350 SHAFT SINKING DETAILS OF LABOR 201 Maximum Minimum Blacksmith, at $2.25 per day; helper at $1.65 per day; $3 . 90 per day for 900 days $3.90 per day for 650 days Two landers at $1 . 70 per day for 900 days Two landers at $1 . 70 per day for 650 days . Two timbermen at $1.75 per day. Assuming one set of timber can be cut and framed in 1 day by two men. 1400 ft. of shaft, 280 sets, 280 days 1000 ft. of shaft, 200 sets, 200 days Two brakemen at $2 . 20 per day, for 900 days Two brakemen at $2 . 20 per day, for 650 days Two firemen at $1 . 70 per day for 900 days Two firemen at $1 . 70 per day for 650 days Allow one-fourth of mining captain's time, $25 per month, for 36 months Allow one-fourth of mining captain's time, $25 per month, for 26 months Surveyor and helpers, allow $20 per month, for 36 months Surveyor and helpers, allow $20 per month, for 26 months . . $3,510 '3,060 980 3,960 3,060 900 720 $2,535 2,210 700 2,860 2,210 650 520 Total $16,190 $11,685 DETAILS OF TIMBER Board Measure, Feet Maximum Minimum 2 Plates, 12X12 in.XlS ft. contain 2 End pieces, 12X12 in. X6 ft. contain. . 4 Corner posts, 12X12 in.X4 ft. contain 2 Diyidings, 10X12 in.X6 ft. 4 in. con- tain 432 144 192 126f 4 Stringers, 10X10 in. X5 ft. contain. . . . 4 Center posts, 10X10 in.X4ft, contain. Sheathing, 3 in. X 5 ft, X 44 ft. contain. . . Boards, 1 in.X5 ft.X6 ft. 4 in. contain. . 166f 133| 660 31f Total amount of timber for 1 set 1 886 Total amount of timber in 280 sets Total amount of timber in 200 sets 535,173 377,267 At $14 per thousand for hemlock timber: for 280 sets $7 492 42 allow $7500 for 200 sets $5 181 7* allow $5 200 Ladders 17 ^c per ft 1400 ft 245 Ladders 17^ c per ft 1000 ft 175 Total $7745 $5 375 202 COAL MINING COSTS DETAILS OF RAILS, PIPE, TIE-RODS, ETC. Maximum Minimum .45-lb. rails for double skip road, allow 1500 ft., mak- ing 6000 linear ft. or 2000 yds., 45 tons at $25 $1 125 200 pairs fish-plates, at 21c 42 500 Ib. rail spikes, at 3c 15 1120 1| in.X6 ft. 6 in. round iron tie-rods, at 2c, $487 46 allow 500 2240 nuts and washers, $278 . 60, allow 300 All of the above computed for 1400 ft. of shaft. For a vertical shaft of 1000 ft. depth, rails, fish-plates and spikes would not be used. Tie-rods, nuts, and washers for 1000 ft. of shaft, allow . $550 1400 ft. of 10-in. water-column pipe 1 820 1000 ft. of 10-in. water-column pipe 1 300 Drill steel 40 30 2-in. steam pipe; allow 250 ft. in excess of length of shaft For 1650 ft 150 For 1250 ft 110 4-in. air pipe; allow 250 ft. in excess of length of shaft. For 1650 ft. 750 For 1250 ft 575 Allow a maximum of 5 tons of coal per day for 26 months 2280 For 36 months 3285 3000 Ib. 60d. spikes at lOc. per pound for 1400 ft ... 200 Ib. lOd. nails at lOc. per pound, for 1400 ft .... 2160 Ib. 60d. spikes at lOc. per pound, for 1000 ft ... 145 Ib. lOd. nails at lOc. per pound, for 1000 ft 300 20 216 15 Total $8 347 $5 081 DETAILS OF PLANT Maximum Minimum Shaft house and pockets Engine house Boiler house Powder house Coal trestle Boilers (4) Hoisting engine Compressor Skips (2) Two No. 3 Rand drills, complete Auxiliary pump at 500-foot depth Sinking pump Temporary equipment at start, small hoist, bucket, rope, tripod, etc., allow Hoisting cable, If -in. diameter Incidentals Teaming, pipe fittings, air hose, picks, shovels, hammers, wrenches, timber cutter's tools, axes, saws, oil, waste, candles, temporary bell signal system, etc Total . . $14,800 9,625 4,200 250 3,000 11,660 15,000 11,500 1,000 375 6,000 1,000 1,500 900 5,000 $85,810 $6,750 5,000 3,500 100 2,500 10,500 12,000 10,625 300 375 5,000 900 1,000 700 4,500 $63,750 SHAFT SINKING RECAPITULATION 203 Inclined Shaft, 1400 ft. Vertical Shaft, 1000 ft. Sinking contract .... $56 350 $40 350 Blacksmithing . . . 3 510 2 535 Landers 3 600 2210 Timber cutters 980 700 Brakemen 3 960 2 860 Firemen 3 060 2 210 Captain and surveyors 1 620 1 170 Timber and ladders 7745 5 375 Rails, fish-plates, spikes 1 182 Air and steam pipes and water column . 2720 1 985 Tie-rods, nuts, and washers 800 555 Nails and spikes 320 231 Coal 3285 2280 Drill steel 40 30 Total $88 632 $62 491 Total plant maximum $85,810 Total plant, minimum 63,750 Inclined shaft with minimum plant 152,382 Inclined shaft with maximum plant 174,442 Vertical shaft with minimum plant 126,241 Vertical shaft with maximum plant 148,300 In spite of very difficult sinking problems, as compared with conditions in this country, shaft-sinking costs in Europe have been substantially less than in this country. Shafts in the Taff and Rhonddha valleys in England, which are circular and from 17 to 21 ft. in diameter were, about 1910, sunk at a total cost of $30 to $50 per foot including the lining. In the north of England the shafts are somewhat larger, varying from 20 to 24 ft. in diameter and are usually lined with steel tubbing. Some excellent speed records have been made at these shafts, at the Sherwood colliery for instance a shaft was sunk 858 ft. in 21 weeks, an average of 40.8 ft. per week. In Belgium brick lining was used almost exclusively at one time, though the use of reinforced concrete is becoming more general. In 1910 sinking costs there were about $60 to $75 per meter and the lining $5 to $6 additional. It is frequently necessary to put down small prospect shafts for depths up to 100 ft. and the approximate cost of such equipment as of 1907 was as follows: 204 COAL MINING COSTS One 25-hp. vertical boiler $300 One 5 X5 in. Bacon type of hoist 350 One 2f-in. steam drill 180 One 7-ft. bucket 30 One 18-in. sheave and bearings 20 200 ft. of ^-in. wire rope 13 Lumber for head-frame, hauling, and labor 107 Total $1000 The items for blacksmith shop, bunk house, cook house, etc., must usually be added, but this amount will, of course, depend upon the size of the prospect and if they are needed. To provide a moderate equipment and to allow a certain amount of working capital another $1000 should probably be provided. Shaft lining's. An interesting example of the costs of a concrete lined metal mining shaft is that of the Brier Hill shaft at Vulcan, Mich., sunk about 1909. This shaft is cir- cular, 14 ft. in diameter and 850 ft. deep. Steel sets made up of 8-in. channels, 13% Ib. per foot, placed on edge and spaced 10 ft. 8 in. on centers were used. Between the sets there are studdles of steel channels to which the wooden runners for the cage are bolted. The ladders are built of steel and the ladderway and skip compartment are lined with galvanized corrugated sheet steel. The line of the shaft was through some old workings so that it was possible to make preliminary openings throughout the entire length from these different levels ; this opening was made 6 X 8 ft. and a 20 X 20-ft. square shaft sunk from the surface to connect with this. Concreting was started at a depth of 79 ft. from the surface. The concrete used was mixed in the proportion of one cement, three sand and six stone and the average thickness of the lining was 18 in. with a minimum of 6 in. Measure- ments of the actual excavation, taken every 3 ft., showed an average thickness of the lining of 19 in. which was equivalent to a little less than 3 cu. yd. of concrete to a vertical foot of shaft. Forms made of %-in. sheet steel were used, there being two sets of forms, each 5 ft. 4 in. high and each set consisting of four segments. SHAFT SINKING 205 After the excavation of the shaft to the proper size was carried downward for a distance sufficient to put in three or four sets of steel, a platform or "curb" is laid on the outside edge of the shaft all around near the bottom at the proper distance from the concrete above to allow for the number of sets proposed. One round of forms is then set upon this plat- form. Empty boxes are put in at the bottom to form "chute holes" for putting concrete into the form below at the proper time and a platform of 3-in. plank is laid over the top of the form. The concrete is then lowered in the kibble, dumped on the platform, slushed off and tamped in around the out- side of the forms completely filling the space between the forms and the rock. A set of steel is then lowered, laid in the hitches left in the concrete and cemented in. The round of forms is lowered upon this new set, expanded to the proper size and the second round of forms is set up and bolted to the first so that this time concrete is deposited for a height of 10 ft. 8 in. This can be done in an eight-hour shift, and 12 hr. are sufficient for the concrete to set so that it is possible to concrete 10 ft. 8 in. of shaft and put in the steel work in less than two days. As the space between the steel sets is 10 ft. 8 in. and in starting only half this amount of concrete is put in before a set is laid, the work is joined to the older concrete above by one round of forms of 5 ft. 4 in. This is filled through the spaces or "chute holes" left by the empty boxes previously mentioned. No attempt was made to make a record of speed. The best work was done in September and October, 1909, which resulted in the excavating (enlarging the original opening) concreting and putting in the steel of 138 2 / 3 ft. of shaft or an average of 69V 3 ft. per month. At that time the men were working two shifts a day or 11 shifts a week of 10 hr. each. Six men con- stituted the regular shift in the shaft either for excavating or constructing. In addition to the shaft work proper three sta- tions have been cut out and concreted. In keeping the record of the cost, the shaft has been divided into three parts: from surface to ledge, 62 ft.; from top of ledge to 7th level, 549.5 ft.; and from the 7th level down, ultimately, about 238.5 ft. 206 COAL MINING COSTS COST OF CONSTRUCTION Surface to Ledge, 62ft. Ledge to 7th Level, 549.5 ft Preliminary excavation i $13 07 $18 46 Final excavation 18 19 15 10 Steel shaft frames 7 90 7 90 Steel forms 83 83 Temporary surface structures and equipment . . . Construction 10.18 56 29 10.18 25 26 Estimated charge for compressed air 1 00 1 00 Total per foot $107 46 $78 73 Estimated salvage on shaft timbers Estimated salvage on temporary surface struc- tures and equipment 0.50 2.95 0.50 2 95 Net total per foot $104.01 $76.28 The item "construction" includes the handling of the steel forms, depositing concrete, and setting and erecting the steel frames. These figures include estimates of power and every charge except for general management and engineering. The cost of the first 85 ft. below the 7th level was $87.19 per foot. The cost of fhis circular concrete-lined shaft is about the same as a rectangular shaft of the same capacity with steel fram- ing. The small additional cost over a rectangular shaft with timber framing is abundantly justified by the increased safety and permanence. One of the earliest concrete shaft linings in this country was put in by the River Coal Co. near Bridgeport, Pa., about 1905. The shaft lining measures 23 ft. on its major axis parallel to the railroad tracks and 15 ft. on the minor axis, inside measurements, the thickness of its concrete walls varying with the depth below the surface. The following table gives approxi- mately the cost of the construction both per foot of depth and per cubic yard: SHAFT SINKING 207 Per Foot of Depth Per Cubic Yard Stone $5.90 $1.00 Sand 1 77 30 Cement 19 18 3 25 Labor: Mixing $3 83 $0 65 Placing 3 40 58 Firemen and pumpmen 2.19 0.37 940 1 (\c\ Forms : Lumber, $13 per thousand $1 83 $0.31 Making $21 per thousand. 2 95 50 Placing 4 82 81 9R1 1 AO Platform for starting upper section 92 16 Superintendence 3.03 0.51 Plant 28 0.05 Oil 21 04 Sundry 1.06 0.18 Tools 24 04 $51.62 $8.75 A novel shaft lining in the form of concrete blocks was used in a shaft in Belgium in 1912, the estimated cost of which was $8.65 per running foot of shaft. This lining was found to be equal in strength to a 32-in. masonry lining the cost of which would have been $13.50 per foot. The shaft was 133 ft. deep and 13 ft. 4 in. in diameter, the excavation being about 17 ft. in diameter. The concrete blocks were 30 in. high with a minimum thickness of 3.2 in. and 14 were required to make the circumference of the shaft. They were set with joints staggered and the successive rings of blocks were joined by 12 X 0.6-in. dowels there being two of these to each block. Concrete with additional reinforcing was filled in between the back of the blocks and the excavation. The increased cost of timber and inferior product being offered in recent years, has tended to cause the use of other materials for shaft linings, especially cement which has come into rapid favor because it has not increased in cost so much 208 COAL MINING COSTS as the timber and because it gives a more permanent job and is fireproof and more or less watertight. The average timber lining lasts from 12 to 15 yr. and in 6 to 8 yr. it becomes necessary to replace individual timbers and sections of lining, causing temporary shutdowns. In the life of a mine of any considerable size, say 30 yr., it will be necessary to re-timber the whole shaft at least once besides making many minor repairs. The cost of re-timbering a shaft (owing to the removal of the old lining, and the increasing price of timber and labor) will be much higher than the cost of the original lining; to this direct cost must be added the loss of income from the mine during the time of repairs. The chief advantages of timber lining are its lower first cost, greater speed in placing it and the fact that timber is better adapted to the rectangular or square form of shaft. Timber lining can be placed in about one-third the time that concrete can which at the average rate of sinking would amount to a saving in time of about 13 days for each 100 ft. of shaft. A comparison of a hoisting shaft and an air shaft of the usual design and with timber lining, with corresponding ellip- tical and circular concrete-lined shafts, will present an example which will closely approximate the conditions obtained in the mining region of western Pennsylvania; for other localities, the local conditions governing the cost can be substituted. The figures are as of 1905. Two shafts of the United States Coal & Coke Co. (a sub- sidiary company of the United States Steel Corporation) at Tug river, West Virginia, were sunk through one seam of coal at 100 ft. in depth, and continued through a second seam at 175 ft. depth. The air shaft was 14 ft. 2 in. on the short axis, by 20 ft. on the long axis. The shaft was lined for a depth of 45 ft. in order to shut off the surface water. The concrete was 12 in. thick. The main hoisting shaft was 17 ft. 4 in. on the short axis by 33 ft. on the long axis, the concrete being 12 in. thick at the sides and 18 in. thick at the ends. It was a four-compart- ment shaft, including a downcast airway, two hoisting-ways and a pipe-way. It was concreted throughout on account of the downcast air-way and the desire to shut off all the water SHAFT SINKING 209 COSTS OP TWO SHAFTS FOR THE U. S. C. & C. Co. AT TUG RIVER, 1905 Main Shaft Elliptical RECTANGULAR Timber-lined Concrete-lined Concrete, Excavation, Timber, per foot of depth, per foot of depth, per foot of depth. 4| cu. yd. 13.5cu. yd. 90 ft. B.M. 5.9 cu. yd. 15 cu. yd. 80 ft. B.M. 12 cu. yd. 500 ft. B.M. Cost per Foot of Depth Concrete, $9 00 per cu yd $40 50 $53 10 Excavation, 5 . 50 per cu. yd. Timber, 60. 00 per M 74.25 5.40 $66.00 30.00 82.50 4.80 Total cost of main shaft. . . $120.15 $96.00 $140.40 Air Shaft Elliptical RECTANGULAR Timber-lined Concrete-lined Concrete, Excavation, Timber, per foot of depth, per foot of depth, per foot of depth. 3 cu. yd. 8 cu. yd. 70 ft. B.M. 8 cu. yd. 450ft. B.M. 4.3cu. yd. 9.9cu.yd. 70 ft. B.M. Cost per Foot of Depth Concrete, $10.00 per cu. yd. $30 00 $43 00 Excavation, 6 .00 per cu . yd . Timber, 61 .00 per cu. yd. 48.00 4.20 $48.00 27.00 59.40 4.20 Total cost of air shaft $82.20 $75.00 $106.60 210 COAL MINING COSTS in the rocks; this was successfully done. The cross buntons were held by cast-iron boxes built into the concrete, but these boxes were probably unnecessary. In the hoisting shaft an average progress of 16 ft. per week was made, 20 ft. being the maximum. The total excavation in this case amounted to 21 cu. yd. per ft. of depth. A paddle concrete-mixer was placed at the head of the shaft, and the concrete was lowered directly from the dis- charge spout of the mixer without further handling. While one bucket was lowering, the other was filling; thus no time was lost in delivering concrete to the placing gang. All form work was done at night, and the concreting on the day shift. The forms were built in 5-ft. vertical sections and were used repeatedly. The cost of labor, mixing, placing forms, lumber, carpenters, hoisting engineers, oil, waste supplies, sundries and superin- tendence amounted to about $4 to $5 per yd. depending upon the size of the shaft and the thickness of the concrete. To this must be added the cost of materials. The table, given herewith, illustrates the relative cost of elliptical concrete-lined shafts, and also the usual rectangular shaft lined with concrete and with timber. It illustrates the great economy of the more permanent concrete shaft. An example of an air and main shaft, each 200 ft. in depth, would represent an outlay of $34,200 for a timber-lined rect- angular shaft; the equivalent elliptical concrete-lined shaft would cost $40,440, both figures including all materials. The difference due to the increased cost of $6240 is more than off- set by the fact that it would take over $15,000 to re-timber both shafts, not to mention the loss of time and repairs. Rates of progress. The best record in shaft sinking up to 1907 was made at the Dixon-Pocahontas Mine in West Virginia. This shaft at Olmstead, W. V., was completed in nine work- ing weeks, although not in nine weeks* continuous work, as an explosion in which four men were killed disorganized the forces and interrupted th work. This work was practically free from water except the last 50 ft. when a No. 7 Cameron sinking pump was required, but after the sump was com- pleted a No. 10 Cameron was installed for temporary use. The shaft is 14 X 22 ft. and 180 ft. deep to coal and was SHAFT SINKING 211 9 weeks in sinking. It is timbered with 8 X 10 in. wall plates and 6 X 10 in. buntons and lagged with 2-in. plank. The tim- bers have bearing sets about 30 ft. apart while the other sets are spaced on 5-ft. centers. The following is the equipment installed: Two 50-hp. Erie City Economy and one 40-hp. Atlas internally fired boilers, a six-drill compressor, a 10 X 12 in. Exeter hoist having a 4-ft. drum and %-i n - plow-steel cable, a 70-in. fan coupled to a 10-hp. engine, and a blacksmith shop outfit. The work was carried down with three 8-hr, shifts for shaft men and two 12-hr shifts for outside men. The number of men and their wages was as follows: 1 shift foreman, $3; 3 drill runners, each $2.50; 3 helpers, each, $2.25 ; 8 muckers, each, $2 ; 1 blacksmith, $2.75 ; 1 hoister, $2 ; 1 compressor man, $2; 1 carpenter, $2.50; 2 helpers, each $1.75. The coal for this 9 weeks* work at $1.25 per ton cost $460. The cost of installing the sinking plant was about $1000 while that of dynamite was $700. Hence, exclusive of machinery, the risk from accidents and the cost of moving the excavated material after dumping the buckets, we may get a close esti- mate of the cost of this work as follows : Installing plant $1000 Coal 460 Dynamite 700 Labor per day, $85 for 54 days 4590 Timber for shaft at $20 per M, 60 M 1200 Miscellaneous expenses 1000 Total $8950 This is practically an expense of $9000 for 180 ft. of 14 X 22 ft. shaft, a net cost of $50 per foot. The speed of sinking will be governed by the quality of the rock, size and shape of the shaft, amount of water present, class of labor available and the efficiency of the plant. The fastest sinking on record up to 1909 was on the Rand in Transvaal, South Africa. These shafts are sunk 9 X 26 ft. in the rock and the work is carried in three, 8-hr, shifts per day, seven days in the week. The labor is cheap and can be used under conditions that the white man will not work under so the shafts are filled up with every man that is possible to use. The records made at some of these shafts are given in 212 COAL MINING COSTS S o r>-io co 10 p p p p ^S9 si JSnisgpfc O5 'CO CO O5i>. s| ;ss ; i3 '!>- TjH '00 >O OOOO O (NtfSt^Ci l> OfNOCNo i i (N O (N O -M oooooo oo co ^H co S^ooocbcb o S 03 CO tt) J OH \-;;i Cal E. d , S. M ^ tfgp-g g, ^ OKj> oo o' ooooooooooo " co" oo" oT " oo" ci" o" o~ HAULAGE COSTS 227 ing to increase or decrease its velocity. If this grade is say 1 per cent then the fractional resistance of the car would be 1 per cent of 2000 or 20 Ib. per ton. When the frictional resistance of the cars is not given it should be assumed at 30 Ib. per ton unless they are newly equipped with roller bearings of an approved make. Roller bearings when installed and looked after properly, will no doubt give frictional resistance ranging from 15 to 20 Ib. per ton. However, a number of roller bearings which had been neglected and not properly lubricated were once tested and showed the average resistance of several cars was 36 Ibs. per ton. On account of the short wheel base of a mine car the fric- tion may be considerably higher when pushed than when pulled. When a string of cars are pushed they are liable to wobble more or less, and considerable binding of the flanges against the rails may take place. When the cars are pulled they are stretched out straight and but little wobbling will take place. The frictional resistance* of new cars will largely depend upon the type of bearing, while for the old cars it will be decidedly influenced by the manner in which the bear- ings are kept up. The locomotive resistances will range from 12 to 20 Ib. per ton. It is safe to take 15 Ib. as an average since the friction of the locomotive is such a small percentage of the total trac- tive effort, and a change of several pounds in either direction will not affect the weight of the locomotive appreciably, and the effect on the capacity of the motors will be negligible. The effect of the frictional resistance of the load will vary with the length and severity of the grades. If the track is practically level throughout, then a small change in the fric- tional resistance may have considerable effect on both weight and equipment. If, however, the grades are long and severe the effect will be small. When a locomotive is operating at a constant speed on a straight, level track, the drawbar pull available for hauling a trailing load (provided there is sufficient motive power) is limited only by the adhesion that can be obtained between the driving wheels and the rails. When starting, the drawbar pull available is reduced, depending upon the rate of accelera- 228 COAL MINING COSTS tion. As this rate is seldom more than 0.2 to 0.25 mi. per hour per second, the drawbar pull will be reduced from 19 to 24 Ib. for each ton weight of locomotive. If there are no grades the weight of the locomotive will be affected considerably by the rate of acceleration. With heavy grades, however, the acceleration will have little effect since the rate can be kept low if it becomes necessary to start on the heavy grade. Accordingly with the low rate of accelera- tion common to mine service this factor can be considered negligible as regards the weight of the locomotive, in view of the fact that a greater percentage of adhesion can be allowed for starting by the use of sand. It has been found in practice that with cast-iron wheels a running drawbar pull equivalent to an adhesion of 20 per cent of the weight on the drivers can be obtained with clean dry rails on level track, without the use of sand. A steel-tired or rolled-steel wheel seems to obtain a better grip on the rails, and a drawbar pull equivalent to an adhesion of 25 per cent can be obtained under the same conditions. When starting heavy trips and when on steep grades it is permissible to use sand, in which case a drawbar pull equivalent to 25 to 30 per cent for cast-iron wheels and 30 to 33V 3 per cent for steel wheels can be expected. Where grades are short the higher rates of adhesion may be used, but for long grades it is not the best practice. Dynamometer tests have given adhesion values as high as 40 to 45 per cent by the use of sand. The average of the tests was, however, much lower so that it is not good practice to count on such high values. These high percentages require the liberal use of sand on both rails, a practice which should not be encouraged as the sand increases the fractional resist- ance of the locomotive and cars, and may work into the bear^ ings and gears. Where no grades exist the weight of the locomotive should, therefore, be five times the drawbar pull for cast-iron wheels and four times for steel wheels, unless the rate of acceleration is such that additional weight is required. When, however, a locomotive with a trailing load is ascending a grade the draw- bar pull is necessarily greater than that required to overcome the friction of the trailing load as the weight of the load has HAULAGE COSTS 229 to be lifted up the grade. For every 1 per cent grade 20 Ib. per ton should be added to the drawbar pull required on straight level track since 20 is 1 per cent of 2000 Ib. The effect of grade on the locomotive as well as on the load must be considered. The heavier the grade the less will be the drawbar pull of the motor. This becomes evident when an abnormal grade is considered on which a motor will be barely able to .propel itself, and if any trailing load is added the wheels will slip. The greater tendency for the wheels to slip on a grade is due to the increased tractive effort necessary to propel the motor itself and the weight transfer due to grade. The weight transfer due to grade will depend on the wheel base and height of the center of gravity. With a short wheel base and a high center of gravity the weight transfer will be considerable. The modern mine motor, however, is constructed with a low center of gravity and a fairly long wheel base so that with the ordinary grades encountered the weight transfer is not serious. The weight transfer due to height of drawbar will also effect the drawbar pull if the wheel base is short and the drawbar high. In Fig. 2 a represents the height of drawbar and 6 the FIG. 2. Influence of drawbar height on weight transfer of car. wheel base. A horizontal force at the drawbar will act as a bell crank, one of whose arms is a and the other &. If the horizontal force represented by the drawbar pull is in the direc- tion of the arrow, an upward force will be exerted at x and a downward force at y. If the drawbar pull is represented by D then the moment about y will be Da. This moment divided by b will give the lifting force at x. The adhesion of the wheel x will of course be lessened by the lifting force. Assume a locomotive weighing 12 tons with a wheel base of 230 COAL MINING COSTS 5 ft. and the height of drawbar 10 in. At 25 per cent adhesion Da will be 6000 X if = 5000 ft. Ib. The upward pull at x will be 5000 ~ 5 = 1000 Ib. The normal weight at x is 12,000 Ib. therefore the weight when the drawbar pull is 6000 Ib. will be 11,000 Ib. so that the adhesion will be 3000 / 110 oo = 27.2 per cent. It is thus to be seen that with the ordinary height of drawbar and wheel base the drawbar pull is not seriously affected. In metal mining a much higher drawbar is sometimes required so that the wheel base must be lengthened to lessen the tilting effect. By tying the axles together by means of side rods, chains or gears the effect of weight transfer due to grade and height of drawbar can be eliminated, but these devices have not proven successful from an operating standpoint. Ordinary mine cars will require from 30 to 40 Ib. pull per ton to move them on a level track. The horizontal tractive resistance of modern cars in good condition may drop down as low as 20 Ib. per ton, but it is not safe to figure much less than 30 to 35 Ib. per ton on the level. For each 1 per cent grade against the load, 20 Ib. per ton must be added. For example, if a car will move on 30 Ib. per ton pull on a level, it will require 30 plus 20 or 50 Ib. pull against a 1 per cent grade, or 30 plus 20 plus 20 or 70 Ib. pull against a 2 per cent grade. A. M. Wellington found that it required 5 Ib. per ton to pull a loaded freight car and 7 Ib. per ton to pull an empty freight car over a level railroad track at 10 mi. per hr. R. Van A. Norris made 980 tests some years ago, which gave a resistance of 26 Ib. per ton for 20 loaded mine cars and 42 Ib. per ton for 20 empty mine cars of the same size over the same track, traveling 4% mi. per hr. One reason for the wide difference between the two sets of experiments men- tioned is found in the size of the car wheels. The freight cars had 32-in. diameter wheels, while the mine cars had 16-in. diameter wheels, thus making it possible for the former to ride over the inequalities in the roadbed with more ease than the latter. Mr. Norris 's experiments brought out another important matter not usually mentioned in haulage articles ; namely, that it requires more power per ton to haul short trains of mine cars than longer ones. This is of added importance to those HAULAGE COSTS 231 mines which have adopted the two-unit locomotive in order to pull heavier loads. One factor that probably reduces the ton resistance in longer trips is that the forward cars clean the track; and another is that the greater momentum of longer trips aids in keeping the cars moving. Professor Baker made some experiments on clean and dirty tracks, the results of which briefly were as follows: On a perfectly clean track the resistance was 19 Ib. per ton; on the same track with % in. of fine dust the resistance was 28 Ib. per ton; while with % in. of powdered stone the resistance was 40 Ib. per ton. This should comfort superintendents in the soft-coal mines who begrudge spending money on road cleaning, because when they clean the haulage entries, they not only lessen the dangers from explosions, but decrease the cost of haulage. It is also a lesson for some anthracite superintendents, for fine anthracite is nearly equal to sand in offering resistance to traction effort. A gasoline motor will exert a drawbar pull equal to one- fifth its weight in pounds when working on a level and on dry rail of proper weight for the motor, but from this drawbar pull exerted on a level we deduct 1 per cent of the weight of the motor in pounds for each 1 per cent grade. For example, a 5-ton motor will exert a tractive effort, or drawbar pull, on the level of one-fifth of 10,000 Ib., or 2000 Ib. Against a 1 per cent grade we would have 2000 Ib. less 1 per cent of 10,000 Ib., or 2000 less 100 Ib., or 1900 Ib., net; or 1800 Ib., net against a 2 per cent grade. It is not only advisable for the operator to prefer the locomotive with the higher tractive effort and low speed to the one with a lower tractive effort and higher speed having the same horsepower rating, but he should even be cautious in buying a locomotive with a high horsepower rating if such a rating is obtained on account of high speed. The reason for this is that a high horsepower rating means increased power consumption without increasing the amount of work done by the locomotive, if the high rating is obtained through high speed. This may be made clearer by stating, for instance, that a 15-hp. locomotive with a rated speed of, 232 COAL MINING COSTS say 9 miles per hour, will take 50 per cent more current for starting a certain train than a locomotive having a 10-hp. rating with a rated speed of 6 miles per hour. It is evident from this how misleading a mere consideration of the horsepower rating of the locomotive may be. In the particular instance given above, the purchaser might think that he is getting a locomotive which will do 50 per cent more work when he buys a 15-hp. locomotive instead of a 10 hp., while actually he would not be able to pull any more with such a machine and would at the same time pay for his mis- take in higher current consumption. As a matter of course, 10 15 40 20 25 30 35 Radius of Curve, Feet FIG. 3. Curves showing the relation between wheel base and radius of curve. there are certain limitations in speed, below which it would not be advisable to go, because if this is chosen below certain limits, the work done by the locomotive would be reduced. In view of the numerous possibilities for being misled, it seems advisable for the buyer of a mining locomotive to request from the manufacturer the following information regarding the rating: 1. Weight of locomotive. 2. Will the motor be able to slip the wheels of the locomo- tive at all conditions of rail? (These two questions will definitely determine the maximum tractive effort which the locomotive is able to exert.) HAULAGE COSTS 233 3. What is the one-hour rating of the locomotive according to the standardization rules of the A.I.E.E. bases on a tem- perature rise of 75 deg. C. on stand test of the motors? 4. What are the tractive effort and speed of the locomotive at the one-hour rating? 5. What is the continuous ampere rating of the motors on stand test at one-half and three-fourths of the rated voltage, and what tractive efforts correspond to these ratings; all bases on 75 deg. C. temperature rise on the stand according to the standardization rules of the A.I.E.E. ? By securing the above information and comparing same for the various locomotives in the market, the purchaser will be in a position to know what he is actually buying. Without it, he is liable to purchase almost anything without knowing just exactly what he is getting. It may be advisable to cite here one of the many cases where purchasers have been misled by not taking a little time in getting the above information. At a certain mine, tests were made on three locomotives of different manufacture. One was a 15%-ton machine with motors rated at 185 hp. total, one was a 17-ton locomotive with motors rated at 210 hp. total. After an all day run, in which the number of car-miles hauled were practically the same, the temperature rises on the motors were as follows : 15!/^-ton locomotive, rated 185 hp., field 52 deg. C., armature 63 deg. C. 17-ton locomotive, rated 200 hp., field 56 deg. C., armature 65 deg. C. 16-ton locomotive, rated 210 hp., field 86 deg. C., armature 85 deg. C. In other words, the locomotive with the smallest horse- power rating in this case proved to be the best of the three in actual service, while the machine with the highest horsepower rating not only showed up to be the poorest of the three, but even had temperature rises exceeding safe limits. This would mean a short life for the motor insulation and windings. Number and size motors required. In order to determine the proper equipment for a mine locomotive it is necessary to have the following information : 234 COAL MINING COSTS Plan and profile of the road. Number of cars to be handled per trip. Number of cars to be handled per hour. X Weight of empty cars. Length of cars. Weight of load. Frictional resistance of cars. Time of layover, including switching and making up trip. Voltage of circuit. Gage of track. Weight of rail. Eadius and length of minimum curve. Spread of track on minimum curve. Limiting dimensions which locomotive can have. Position and range of trolley wire. It is seldom that all of the above information can be obtained, and in many cases it is necessary to make certain assumptions to supply the missing data. This can only be done by one having considerable experience in working out mining problems. Motors for mine locomotives are rated on the one-hour basis with a 75-deg. C. rise in temperature. This rating does not indicate the capacity of the motor for all-day service, and is not used in determining its ability to meet with a certain set of conditions. The capacity of a motor for all-day service depends upon the temperature which the windings will attain. This in turn depends upon the average heating value of the current. Since the heat generated by an electric current is proportional to the square of the current value, the average heating for all- day service must depend upon the square root of the mean square of the current. Two motors may have the same one-hour rating, but one may have a much larger continuous capacity than the other, due to better design and the proper distribution of the losses. A poorly ventilated motor will in some cases have hot spots, which will lower the capacity of the machine. This is due to the fact that, in order to keep these spots within a safe tem- perature rise, the average temperature of the windings must be kept much lower than would be necessary if such spots were eliminated by proper design. That the real capacity of a motor is its continuous capacity HAULAGE COSTS 235 for all-day service and not the rating for one hour is apparently not generally appreciated among mine operators. The one- hour rating depends largely upon the thermal capacity of the motor, while the continuous rating depends on the ventilation, distribution of the losses and the capacity of the machine to radiate heat. The one-hour capacity is not a fair rating of a motor for the foregoing reason and also because the speed of the motor- is not taken into account. A fairer way would be to rate the machine on the pounds tractive effort at the wheels, irrespective of the speed, provided it is not considered essential for com- mercial reasons to capitalize the increased horsepower ratings due to increase in speed. If the length of haul, the grade, curve, running time and time of layover are known, the current for each part of the run can be computed. In most main-haulage cases the locomo- tive will have a definite cycle to go through, this cycle being repeated throughout the working day. If the square root of the mean square current for one cycle can be found, this will, of course, determine the suitability of the motor selected for the all-day service as regards heating capacity. To illustrate the working out of the above principles, the following conditions may be assumed to exist at a mine which desires to install electric haulage: Locomotive required 1 Profile as follows: 1300 ft. 2 per cent grade against load. 1400 ft. 1 per cent grade against load. 2200 ft. level. Number of cars to be handled per trip 20 Number of cars to be handled per hour 50 Weight of empty car 2000 Ib. Weight of load 4500 Ib. Total weight of loaded car 6500 Ib. Frictional resistance of cars 30 Ib. per ton Time of layover, including switching and making up trip. 5 min. each end Voltage of circuit 250 Gage of track 36 in. Weight of rail 30 Ib. per yd. Radius of minimum curve 25 ft. Length of minimum curve 20 ft. Spread of gage on curve \ in. Limiting dimensions of locomotive, 5 ft. wide, 4 ft. high. Trolley wire 6 in. outside of rail; height above rail, 4 ft. 6 in. to 6 ft. 236 COAL MINING COSTS The total weight of the trip will be 65 tons. The limiting condition in regard to weight is, of course, on the 2 per cent grade. The weight of the locomotive will be found as follows: 30 X 65 + 20 X 2 X 65 + 20 X 2 X W = 400 W W = 12.6 tons if cast-iron wheels are used ; W = 9.88 tons if steel wheels are used. It would, therefore, be necessary to use a 13-ton locomotive with cast-iron wheels or a 10-ton locomotive with steel wheels. Steel wheels should be used unless the customer specifies cast- iron. A locomotive to negotiate a 25-ft. curve should have a wheel base not more than 55 in. with 33-in. wheels or 65 in. with 30-in. wheels. With motors tandem hung no trouble will be experienced in keeping below 55 in. or 65 in. for a 10-ton locomotive. See Fig. 3. 50 Speed M.P.H. 01 O W 40- Hf. MINE MO TOK Continuous Capacity, 64 Amperes c t+ISt ' ZOC IVo/t S '/* 4000 "I 8000 1 1000 \ (& / \ J r \ \ at 's ^ ^/ ( 2& / f~ ^J/ Lt i. / / 00 250 100 150 Amperes FIG. 4. Characteristic curves for a 40 hp. mine locomotive motor. The number of cars to be handled per hour being 50, the trips per hour will be 50 -j- 20 = 2y 2 . The total time per trip, including layover at each end, will therefore be 60 -7- 2y 2 24 min. Allowing 5 min. layover at each end will make the actual running time 14 min. For a locomotive of a given weight there are, as a rule, two or more motors to choose from. For a 10-ton machine these motors range from 40 to 50 hp. in capacity, although larger ones are sometimes required for special cases. A 40-hp. motor HAULAGE COSTS 237 is selected for the first trial. The locomotive will have 30-in. wheels with a gear ratio of 4.78 to 1. The highest gear reduc- tion is always selected unless a greater speed is required and can be obtained without overloading the motors. The charac- teristics of this motor are shown by the curves in Fig. 4. This curve is made from an actual test, and the tractive effort given includes gear losses, so that to obtain the drawbar pull only the locomotive friction should be deducted. A table of data covering the case, such as shown in Table I or II, should be prepared, the values inserted being calculated from the motor curves and weights to be handled. Since the curves give values for one motor, the locomotive and trailing weight should be divided by two to give the weight each motor will be required to handle. TABLE I Speed, Miles per Amp. Total Tr. Eff. Loco. Res. Train Res. Grade Res. Dis- tance, Feet Time, Sec. Amp. 2 Amp.2x Time Hr. - 8.5 87 1050 75 975 2200 177 7,569 1,340,000 6.5 165 2550 75 975 1500 1300 136 27,225 3,700,000 7.2 125 1800 75 975 750 1400 133 15,625 2,080,000 Returning with Empty Trip 10 25 75 75 300 -300 1400 96 625 60,000 10 -225 75 300 -600 1300 89 10 45 375 75 300 2200 150 2,025 303,000 Amp. 2 X time = 7,483,000 Plus 10 per cent for accelerating, switching, etc .... 748,300 Total amp 2 X time 8,231,300 Total running time, sec 781 Total time at both ends, sec 659 Total time including lavover. sec . . 1.440 8,231,300-^- 1440 = 5700 = mean squared current. The square root of 5700 = 75 . 5 = square root of mean square current. Capacity of 40-hp. motor is 60 amp. For the above project the weight of locomotive is five tons per motor; the loaded trailing weight, 32% tons per motor, and the light trailing weight, 10 tons per motor. Assume that the locomotive starts with a load on a level track and runs 2200 ft., when it encounters a 2 per cent grade. After ascend- 238 COAL MINING COSTS ing this grade for 1300 ft. the grade changes to 1 per cent for 1400 ft. The return trip will be with empty cars. Starting with the loaded trip on the level, the locomotive resistance per motor will be 5 X 15 = 75 Ib. This value is placed under "locomotive resistance" in the table. The train resistance will be 32.5 X 30 975 Ib. The grade resistance will be zero. The total tractive effort will be 1050 Ib. Consulting the motor curves of Fig. 4 the current for a tractive effort of 1050 Ib. is 87 amp. and the speed 8.5 miles per hour. The time to cover 2200 ft. at 8.5 miles per hour will be 177 sec. The amperes squared will be 7569 and the amperes squared multiplied by time will be 1,340,000. These values should be recorded in their proper place in the table. When the 2 per cent grade is reached, the train and loco- motive resistance will remain the same, while the grade resist- ance will be 40 X (32.5 + 5) =1500 Ib. The total tractive effort will be 2550 Ib., which corresponds to a motor current of 165 amp. and a speed of 6.5 miles per hour. At this speed it will require 143 sec. to travel 1300 ft. By the same process the values for the 1 per cent grade are calculated and filled in the table. On the return trip with the empty cars the locomotive resistance will be the same, the train resistance 300 Ib. for 10 tons, and the down-grade resistance 300 Ib. for the 1 per cent and 600 Ib. for the 2 per cent grade. It will be noted that, running down the 2 per cent grade, the net tractive effort is 225 Ib., which means that the brakes must be applied in descending this grade. On the 1 per cent grade and on the level the speeds shown on the curve of Fig. 4 are too high for most mines unless the track is in good shape. It is likely that the operator will not care to run faster than 10 miles per hour, which he can do by operating the motors in series on low notches, or by cutting off the power and coasting before the speed becomes too high. The total running time is 781 sec., or 13 min. 1 sec. The actual running time will be close to 14 min., due to time taken to start and stop the trip and for possible slowing down at cross-overs. As the total time for a round trip is 24 min., a layover of five minutes is obtained at each end. The product of the total current squared by the time is HAULAGE COSTS 239 7,483,000. To this, 10 per cent should be added to allow for acceleration and switching when making up trips, making a total of 8,231,300. The total time for making a round trip, including layover, is 1440 sec. Dividing 8,231,300 by 1440 = 5700 as the mean square of the current. The square root of 5700 is 75.5 amp., which is the square root of the mean square current for one trip or cycle. The continuous capacity of the motor is 68 amp. at 150 volts and 64 amp. at 200 volts. The class of service is such that the average voltage applied to the motor will be near 200, so that the rating of the motor is about 65 amp., which shows that it is not of sufficient capacity for the service. TABLE II Speed, Miles per Hr. Amp. Total Tr. Eff. Loco. Res. Train Res. Grade Res. Dis- tance, Feet Time, Sec. Amp. 2 Amp.2x Time 10.1 7.6 8.5 100 187 146 1050 2550 1800 75 75 75 975 975 975 1500 750 2200 1300 1400 149 117 113 10,000 34,970 21,316 1,490,000 4,090,000 2,410,000 Returning with Empty Trip 75 75 300 25 75 300 75 75 300 accelerating, switching -300 -600 , etc. ... 1400 1300 2200 96 89 150 1,225 3,136 117,500 470,400 8,577,900 857,790 9,435,690 714 726 layover, sec . . 1.440 Amp. 2 X time: Total amp. 2 X time Total running time, sec . . Total time at both ends, sec 9,435,690 -5- 1440 = 6550 = mean squared current. The square root of 6550 = 81 = square root of mean square current. Capacity of 50 hp. motor is 80 amp. at 150 v. Care should be taken that a motor is not selected in which the commutating limit is exceeded when the wheels are slipped while using sand. The motor curves are generally stopped at the commutating limit, although with the modern com- mutating-pole motor it is rather difficult to find the commutat- ing limit. 240 COAL MINING COSTS A larger motor should be selected, and Table II shows the results of the calculation using two 50-hp. motors. The curves of Fig. 5 show the characteristics of this motor. The total calculated running time is 714 sec., or 11 min. 54 sec. The square root of the mean square current is found to be SO amp. The capacity of the motor at 200 volts is 78 amp. and 82 amp. at 150 volts, so that this motor will be of just about the proper capacity to meet the conditions. The actual running time will be about 12 to 14 min., allowing 5 to 6 min. at each end for 16 50-HP MINE MOTOR Continuous CapacHy, 82 Amperes erf- ISO Vo r y 78 r 00 . 5000 4000 3000 2000 1000 40 60 120 160 200 Z40 280 320 360 Amperes FIG. 5. Characteristic curves for a 50-hp. locomotive motor. switching and making-up trips. The higher root mean square current obtained for the 50-hp. motor is due to the fact that it is a higher-speed machine than the 40-hp. This shows the importance of having a low-speed motor. It is not safe to figure on a short layover, as in many cases the average line voltage is much less than 500 or 250 volts, which means that the speed will be less than is figured on. A low line voltage signifies that a given current will be re- quired for a much longer time than with the normal voltage, which in turn means additional heating. Where the voltage is likely to be poor, a margin should be allowed in the motor HAULAGE COSTS 241 capacity, since the value of the square root of mean square current will be greater than that calculated. The conditions outlined above in regard to profile are typical of what may be expected in mines. Sometimes it is possible to lay out a mine so that all of the main haulage will be with the grades ; other mines will have mixed grades that is, some inclinations in favor of the load and some against it ; while too often a mine will be found with little or no level track and all grades against the load. If the profile instead of that given above were 2 per cent against load 300 ft., level track 2300 ft., 1 per cent in favor of load 2300 ft., the same weight of locomotive would be required, but the heating would be much less. Table III shows that the root mean square current is only 51.7 amp., which gives a large margin of safety when using the 40-hp. motor. TABLE III Speed, Miles per Hr. Amp. Total Tr. Eff. Loco. Res. Train Res. Grade Res. Dis- tance, Ft. Time, Sec. Amp. 2 Amp.2x Time 6.5 10 8.5 165 40 87 2550 300 1050 75 75 75 975 975 975 1500 -750 300 2300 2300 31.5 157 184.5 27,225 1,600 7,569 870,000 251,200 1,395,000 10 10 10 Returning with Empty Trip 75 75 75 75 25 75 iccelerating, g 300 2300 157 300 300 2300 157 300 -600 300 20 witching, etc 2,025 4,225 318,000 664,000 3,498,200 349 820 3 848 020 ec 707 ids, sec 733 layover, sec. 1.440 Amp. 2 X time Total amp. 2 X time Total running time, sec Total time at both ends, sec 3,848,020 -=- 1440 = 2670 = mean squared current. The square root of 2670 = 51.7 amp. = square root of mean square current. On the other hand, the grade may be 2 per cent for the entire distance against the load. In this case the root mean square current would be about 105 amp., corresponding to a motor having an hour rating of 75 to 80 hp. As it would not be practical to put two such large motors on a well designed 242 COAL MINING COSTS 10-ton locomotive, it would be necessary to go to a 12- or 13-ton machine. In another computation carried out along somewhat dif- ferent lines, it will be assumed that it is desired to haul coal from a siding at the bottom of the slope, in trips of 30 cars. The empty cars each weigh 1 ton and have a capacity of 1.5 tons. This will make the total weight of the loaded trip 30 (1 -|- 1.5) =75 tons. The accompanying profile of the road, Fig. 6, extending from the siding at the slope bottom in the mine to the tipple where the coal is loaded into the railroad cars, shows the length and grade of each portion of the track and the weight of rail in use. It is desired to know the size of electric motor that will be required for this haul; also the weight of rail and style of rail bond that should be used, the size of trolley wire required for the transmission of the power from the generator to the mine, and the required horsepower of the generator and boiler. The first step is to estimate the weight of locomotive re- quired to haul a loaded trip of 75 tons up a 3-per cent grade, under ordinary mining conditions. It is not safe to estimate on the track resistance as being less than 50 Ib. per ton. To this must be added 20 Ib. per ton for each per cent of grade or, in this case, 3 X 20 = 60 Ib. grade resistance, which makes the total resistance 50 -f- 60 = 110 Ib. per ton of total moving load including the locomotive. The coefficient of adhesion of the wheels to the rails will be estimated at 0.16, which makes the tractive effort that the locomotive can exert to move itself and the loaded trip 0.16 X 2000 = 320 Ib. per ton. Then, since this tractive effort of the locomotive must be equal to the total resistance of the entire moving load including the locomotive, and ^ 110TF, 110X75 UQ = ~2I6~ Ar . =Sa 4 This shows conclusively that, under the adverse conditions common in coal mining, a 40-ton locomotive would be required to haul a loaded trip of 75 tons up a 3-per cent slope. This of course provides for the worst conditions that are liable to HAULAGE COSTS 243 8. 8 &' 244 COAL MINING COSTS exist, in respect to both the track and rolling stock in the mine. Before going further, it will be well to estimate the weight of locomotive that will be required to haul the same loaded trip on the 1%-per cent grades outside of the mine. In this case, the track resistance, as before, is 50 Ib. per ton, but the grade resistance is 1.5 X 20 = 30 Ib. per ton, which makes the total resistance 50 -f- 30 = 80 Ib. per ton. This gives for the required weight of the locomotive : 80X75 320-80 TTr O\J /\ I fj c\f , W m = ^^ = 25 tons. The same weight locomotive will haul practically two-thirds of the number of cars up the slope that it can handle on the outside grades, as shown by transposing the formula pre- viously given and finding the load that a locomotive can be expected to haul regularly up grades of 1% and 3 per cent, respectively, under the conditions named, thus, On a 3-per cent grade: tractive effort of locomotive, 320 Ib. per ton; track and grade resistance, 80 Ib. per ton. qon _ 1 in Loaded trip, W t = W m = 1.9TF W . On a IJ-per cent grade: tractive effort of locomotive, 320 Ib. per ton; track and grade resistance, 80 Ib. per ton. Loaded trip, W t = 32 8 ~ 8 W m = 3W m . For this reason, it might be well to provide a siding at the top of the slope that will hold 20 or 40 cars as desired, and make three trips on the slope for every two trips to the tipple, using a 25-ton locomotive for the entire work. So far, we have only considered relative weights of the locomotive and the load it can be expected to handle regularly and satisfactorily on different grades, under the worst con- HAULAGE COSTS 245 ditions that are liable to exist or arise in mining practice. It is important to remember that having assumed a coefficient expressing the adhesion of the wheels to the rails ft is the weight of the locomotive that determines the tractive effort the machine can exert, regardless of the power of the motors or engines with which it is equipped. It is a fatal mistake to equip a locomotive with more power than its weight will permit it to utilize. Having decided on the weight of the locomotive necessary to haul the desired number of loaded cars over the given grades, the next step is to determine the power of the motors that will be required to produce a given speed of haul, say from 6 to 8 mi. per hr. As a basis of this calculation, it is important to observe that the effective power, in foot-pounds per minute, is equal to the tractive effort in pounds, multiplied by the speed in feet per minute, as expressed by the formula. Power = tractive effort X speed. (ft.-lb.p.m.) (lb.) (ft.p.m.) Now, estimating the horsepower required per ton on drivers, per mile-hour, assuming a tractive coefficient c = 0.16, we have: Tractive effort (per ton) =0.16X2000 = 320 lb. Speed (per mi.-hr.) = 5280 ^ 60 = 88 ft.p.m. QOQ \/ CO Horse-power (per ton-mi. -hr.)-. =0.85 h.p. ooUUU This is the effective horsepower per ton mile-hour, or the power that must be available to produce a speed of 1 mi. per hr., for each ton of weight resting on the drivers. Assuming an efficiency of 85 per cent, the input to the motors should be 1 hp. per ton-mi.-hr. Using a 25-ton locomotive for this haulage and assuming that the entire weight of the locomotive is on the drivers, the actual horsepower of the motors required for a speed of 6 mi. per hr. will be 25 X 6 = 150 hp., which corresponds to a wat- tage of 150 + 746 = 111,900 watts delivered to the motors. It is interesting to note here that the effective wattage per mile-hour, in electric mine haulage, is practically twice the tractive effort of the locomotive, in pounds. This is shown by the following simple calculation, since a speed of 1 mi. per 246 COAL MINING COSTS hr. is equal to 88 ft. per min. and 1 hp. is equivalent to 746 watts : /OO y rp p \ Effective watts (per mi.-hr.)74Q= ( 33QQQ j =sa y 2 T - E - Under the assumed conditions, it was previously estimated that the total tractive effort when hauling up a 3-per cent grade is 110 Ib. per ton of moving load, which makes the effective wattage required in that case 2 X HO 220 watts per ton-mi.-hr. In like manner, it was estimated that the total tractive effort when hauling up a 1%-per cent grade was 80 Ib. per ton, which makes the effective wattage required to haul up that grade 2 X 80 = 160 watts per ton-mi.-hr. In electric mine haulage, it is customary to estimate on hauling at a speed of, say from 6 to 8 mi. per hr. Taking the actual power of the motors at 1 hp. per ton-mi.-hr., to give an input of 111,900 watts on a 500-volt circuit will require a current of 111,900 -r- 500 = say 224 amp. The weight of rail in pounds per yard, in locomotive mine haulage, may be taken as twide the weight of the locomotive in tons, or, in this case, 2 X 25 = 50 Ib. per yd. In estimating the size of wire required to transmit this current from the generator to the siding at the foot of the slope in the mine, we will assume a rail-return, using 50-lb. rails for the track, bonded with compressed-terminal bonds, 24 in. long. The length of track from the end of the siding at the foot of the slope to the tipple is about 5400 ft. The length of trolley wire extending to the power plant will be assumed as 5600 ft. Referring now to the diagram shown in Fig. 7, first find the resistance of 5200 ft. of the two 50-lb. rails in the track- return, including the resistance of the bonds, assuming a rail to copper ratio of 1 : 11 and rails 30 ft. long, making 5400 -4- 30 = say 180 bonds in a single length of rail. Following the horizontal line marked 50 on the left of the diagram, to its intersection with the curved line indicating a rail-to-copper ratio of 1 : 11 ; and, from that point, following the vertical line to the scale at the bottom of the diagram, we find a resistance of 18 microhms per foot of rail or 30 X 18 = 540 microhms per rail. In like manner, for a 24-in., 4-0 bond, follow the vertical line marked 24 on the top scale down to its intersection with HAULAGE COSTS 247 the diagonal 4-0; and, from this point, follow the horizontal line to the right-hand scale, which shows a bond-resistance of 100 microhms. The total resistance for 180 bonded rails is, therefore, 180 (540 + 100) =115,200 microhms, or 0.1152 ohm. The resist- ance for a two-rail return is one-half this amount or 0.0576 ohm. The voltage absorbed in the rail-return is E = CR = 224 X 0.0576 = 12.9 volts. The line drop for 5600 ft. of T. B., 4-0, copper wire, carrying a current of 224 amp., as taken. - from wire tables, is 5.6 (224 X 0.0489) = 61.3 volts. The dif-K Length of Bond, Inches 12 16 20 24 26 32 36 approximately 150 per cent of that of 3600 Ib., with a current, consumption of 500 v. of 115 amp. The ragged current and volt- age curves are noted to occur during the whole of the round-trip. 262 COAL MINING COSTS The current serves also as a direct measure of the torque, which can be computed from the actual characteristic curves. As the entry in this mine often runs on the strike of the coal, one side of the track is ballasted and the other rests on bed rock. This means a poor, uneven track, sometimes dirty from loose coal falling off cars or from rock falls. Then, too, wet spots and small increases of grades are often met with, requiring fre- quent sanding. Although the couplings are long, the grade I of 6 TOWtf'JVe LOCOHOT/VC.. ?- SOO V: /V- Motors. 4O-t(^W \ *5 C. ftfse - 43/tfnp. for. / Hour. \&f?-6M.PH. Whee/ O/0/n. 28 /nchts. Gear ffaf/'o /S/o 94. L/'fte /fes. 2 Oftma. 60 .70 SO 9O JOO A/np. O 10 ZO 30 FIG. 13. Typical round-trip performance of a mine locomotive working under actual mine conditions. keeps them tight and requires a greater accelerative torque than when running on the level. As this entry is one of four in a mine about 2 miles from the power station, and each of them has a locomotive, besides using four electric booster fans and an electric hoist, no further ex- planation is necessary of the apparently inconsistent variation of voltage. It is typical of mine service. The generators at the power house are not overcompounded, and there is, therefore, no compensation for line loss. This results in this particular case in quite a drop of pressure when the electric locomotive is pulling HAULAGE COSTS 263 heavily ; consequently, the characteristics obtained at 500 v. can not be very well used for correct calculation at 100 amp. con- sumption. The changes in such characteristics due to two ohms drop in the feeder line and return are sketched in roughly. The decrease in electrical and brake horsepower is very apparent, resulting in great loss of speed at heavy loads, but increasing the tractive effort at this lower speed. It is interesting to note the enormous difference of perform- ance between an ordinary day trip such as is shown here, and one with very cold weather outside. The empties coming from the outside run with great difficulty because of frozen boxings, and often require more than a 50-per cent increase in torque; by the time they are loaded, however, the mine temperature has thawed them sufficiently to make the outgoing trip a fairly normal one. Where the inertia of the trip of cars at starting, or increased resistance due to a dirty track, causes the locomotive to pull heavily and run slowly, the effect of a long feed line causing an appreciable voltage drop at the locomotive is of interest. The drop, of course, is proportional to the current consumption. The motor torque is proportional to the current only, and is inde- pendent of the impressed electromotive force. With a heavy pull, the motor 's speed need not be so great; as an increase in current flow means less speed for the necessary current electro- motive force. As the applied electromotive force decreases (due to line loss) and the current electromotive force increases with the greater current flow, the torque, and therefore the current, does not increase as in the case of constant applied voltage. There is, therefore, less danger of overloading the motors. As the tractive effort also is higher at lower speeds, such line resist- ance allows of larger trips to be started by a certain weight loco- motive without an excessive output at the power house. As the trip resistance is decreased after starting, the speed rises with the decreased current and increased applied electromotive force making up for time lost on the heavy pull when starting. Another advantage of a long-feed line is decreased harmful- ness of short-circuits. A short-circuit on this 2-mile line will blow the circuit breaker without causing a violent flashing of the gen- erator brushes, or excessive arching at the contacts of the circuit breaker, since the current cannot exceed 300 amp. 264 COAL MINING COSTS Line costs and losses. Line losses can best be shown by a practical example. For purposes of demonstration it will be assumed that the haulage road at a certain mine is 6000 ft. long and laid with 30-lb. rail, which are bonded with 00 wire, and bonding caps with cross bonds every 30 ft. There is a sub- station at the inside end, 250 volt current is used, the trolley wire is 0000 and a 10-ton motor using current at 150 amp., is used. Let it be desired to find: What will be the drop in volts at 2000 ft., 4000 ft., and 6000 ft. from the substation. What is the percentage of loss in each case. What would be the saving, if any, in installing 0000 feeder wire, allowing say 6 per cent interest on the investment and com- puting the cost of power at 2c. per kw.-hr. In estimating the drop at distances of 2000, 4000 and 6000 ft. from the substation, it is necessary to calculate the trolley- drop and rail-drop separately and add the results. Then, in order to estimate the possible advantage of installing a feeder-line of the same size as the trolley-wire (0000), it is necessary to cal- culate further the combined trolley and feeder-drop for the same distances. The first step in the calculation is to take from an electric- wire table the circular mils of a 0000 copper wire, which is 211,600 circ.mils. The combined trolley and feeder wires, the circular mils is double this amount, or 423,200 circ.mils. Now, to find the copper equivalent of two 30-lb. rails, properly bonded and cross-bonded, assume a rail-to-copper ratio of 11 :1 ; that is to say, take the rail resistance here as 11 times that of copper, for the same sectional area. Then find the combined sectional area of two 30-lb. rails as follows: Since the weight of wrought iron is 480 Ib. per cu.ft, and two 30-lb. rails weigh 60 Ib. per lineal yard (36 in.), the cubic contents of the two rails is 60/480X1728=216 cu.in. per yd. The corresponding sectional area for the two rails is therefore 216-^-36=6 sq.in. Again, since 1 mil is 1/1000 in., and a circ. mil is the area of a circle whose diameter is 1 mil, 1 circ.mil = 0.7854 (1/1000) 2 = 0.0000007854 sq.in. and 1 sq.in. = 1/0.0000007854 = 1,273,200 circ. mils. The sectional area of the two 30-lb. rails in circular mils is therefore 6X1,273,200 = 7,639,200 circ.mils. For a rail-to- HAULAGE COSTS 265 copper ratio of 11 : 1, the copper equivalent is 7,639,200 -^ 11 = say, 694,500 circ.mils of copper. It is now possible to calculate the drop of potential for the trolley line, feeder and rail return for each required distance by using the general formula: ~ . ,. , 10.8 (distanceX current) Drop of potential = -- - - : -- = - -. ctrc.mils The drop of potential is expressed in volts and the current in amperes, while the distance is given in feet. Applying this formula to the present case gives for the drop at a distance of 2000 ft. from the substation the following: 10.8(2000X150) Trolley-drop = -- -- = 15 - 3 voUs - It is evident that the combined trolley- and feeder-drop would be one-half of this amount, or 7.65 volts, since there are two wires of equal size for the transmission of the current to the locomotive. The rail-drop is: D , 10.8(2000X150) Rail-drop = - ^=4.6 vote. The above results give the drop of potential for a distance of 2000 ft. For distances of 4000 and 6000 ft. the drop will be two and three times the above amount respectively. A comparison of these results gives the following : Total drop without feeder, 4.6+15.3 = 19.9 volts Total drop with feeder, 4.6+ 7.65= 12.25 volts Voltage saved by feeder, 19.9 - 12.25 = 7.65 volts Wattage saved by feeder, 150X7.65 = 1147.5 watts The percentage of drop in each case above is: Without feeder (2000 ft.), 19.9X100-^250 = 8.0% With feeder (2000 ft,), 12.25X100^-250=4.9% Assuming 250 working-days of 8 hr. each (2000 working hours) in a year, the saving in that time at a cost for electricity of 2c. per kw.-hr. would be 2000X1. 1475X0.02 = $45.90. Against this saving must be reckoned the interest or fixed charges on the investment for the erection of 2000 ft. of feeder wire. The weight of 0000 copper wire is practically 640 Ib. per 1000 ft. making the costs for 2000 ft. of this wire, at 20c. 266 COAL MINING COSTS per lb., 2 X 640 X 0.20 = $2.56. The cost of erection, including poles, will be about $4 per 100 ft., or $80 for a distance of 2000 ft., which makes the total cost of this length of feeder wire erected, $336. Estimating the fixed charges on this investment as, interest, 6 per cent; depreciation, 4 per cent, and taxes and insurance, 2 per cent, making a total of 12 per cent, gives 0.12 X 336 = $40.32. This estimate makes the net saving per year, $45.90 40.32=$5.58. Since the drop in potential the saving in kilowatt-hours and the fixed charges are all proportional to the distance, the net saving per year will also be proportional to the distance, or $11.16 for a distance of 4000 ft., and $16.74 for a distance of 6000 ft. In actual practice, however, the locomotive .will not be run- ning more than from one-half to three-quarters of the actual time, depending on the length of the haul and other conditions, which will reduce the saving in kilowatt-hours at the same rate, while the fixed charges must remain constant. Although the figures would then show an excess of fixed charges over the saving in the cost of electricity, it would still be good practice to add the feeder wire. The gain in the operation of the loco- motive will exceed what the actual saving in power would indicate. Grounding losses can sometimes be discovered on the ammeter when the mine is shut down or at any other time when all the power is supposed to be off. In one instance an investigation of a 75 amp. reading of the ammeter when the mine was supposed to be closed for the day disclosed the fact that 25 trolley wire hangers were grounded. When the trouble was remedied the ammeter read zero. A simple calculation will show the loss from this leakage of 75 amp. under a pressure of 275 volts, being .continuous for the time the switch on the supply current is open, which is 16 hr. a day. This loss in six months of 125 working-days of 16 hr. each, or say 2000 hr., would amount to (75 X 275 X 2000) -4- 1000 = 41,250 kw.-hr. At a cost of 2c. per kw.-hr., the loss would be 41,250X0.02=$825 in six months. In the present case, the loss being due to the leakage of current by the grounding of 25 trolley- wire hangers, the loss is 825 -f- 25 = $33 per hanger HAULAGE COSTS 267 in six months, or $5.50 per hanger per month an amount that is almost inconceivable. However, it shows the importance of carefully testing an electric circuit for grounds and locating and stopping the leak. Besides this direct loss of current, there is the danger from fire where the hangers are in the coal roof or insulator pins are used instead in the entry ribs. The following is an approximate estimate of the cost of 1000 ft. of trolley construction in coal mines, including rail bonds as estimated in 1908: 36 single-bolt roof suspensions $15 30 straight ears for grooved wire . 6 6 curved ears for grooved wire 6 1000 ft. of 0000 grooved trolley wire (641 Ib.) 94 210 ft. of 00 bond wire (85 Ib.) 12 175 ft. channel pins, GE Catalog, No. 17,315 5 Labor and other material 50 Additional labor, if necessary, to drill rails for bonds with a hand drill. . 15 Total cost per 1000 ft $197 Other sizes of trolley and bond wires may be readily sub- stituted in the above estimate. Bonding. While a few operators are using modern methods of bonding, many are depending on channel pins and wire to carry the return circuit. The channel pin when first installed is more or less efficient, but as three-fourths of its contact is between steel pin and steel rail, it is impossible to obtain a union that will exclude* air and moisture. Therefore it is only a short time until corrosion has started and a high resistance is intro- duced at the points .of contact. The method of testing prevalent at most mines consists of examining the bond to see that the wire and pins are intact. The return circuit of a mine that was bonded partly with channel pins and partly with compressed-terminal flexible-cable bonds was tested with a direct-reading bond tester which showed the resistance of each joint as equal to the resistance of a certain length of solid rail. Thirty-one per cent of the channel-pin bonds showed a resistance equal to, or greater, than that of a 30 ft. rail, or practically an open joint; the average resistance of the balance was equal to that of 13 ft. of rail. These channel- 268 COAL MINING COSTS pins had been installed about 2% yr. and on an exceptionally good roadbed. The compressed terminal bonds had been installed 4 yr. on a roadbed that was in bad condition. The drainage was bad and the soft roadbed permitted a considerable rising and sinking of each joint when a car passed over, thus imposing a severe strain on the bond terminals; 16 per cent, of these bonds were found defective, and the balance showed an average resistance of 6.6 ft. Had these bonds been installed in track similar to that in which the channel-pins were used there is no doubt but that the depre- ciation would have been cut in half. A point of special interest in this test was the fact that the majority of the compressed-terminal bonds were in good condi- tion after 4 yr. of service and under rather unusually un- favorable conditions. Had this company tested these bonds at certain intervals, and replaced defective ones as found, they would have had, at a small expense for labor and material, a highly efficient return circuit at all times. This operation had been suffering from an excessive drop in voltage, and the results of this test proved conclusively that it was caused by a defective transmission of the return current. The relative resistances of 21 joints selected at random along the haulage system of this mine, bonded first with channel- pin bonds then with compressed-terminal bonds, are clearly shown in the accompanying curve, Fig. 14. One mine manager reported that, with sufficient copper over- head, he found a drop of 100 volts, at less than 3000 ft. from the generator on 250-volt direct-current circuit. The channel- pin bonds in this mine were tested, and 90 per cent of them found to have a resistance greater than that of 30 ft. of rail. The channel-pins were replaced by compressed-terminal bonds, the line voltage went up to normal and the efficiency of the locomotive increased to a marked extent. Computing losses in bonds. It is essential, both in making tests of bond installations, and in estimating new work that the resistance of a bonded joint be known. The following tables and formula show just what resistance a well bonded joint will have. To find the resistance of a rail joint bonded with compressed terminal bonds, the following formula is used: LXR+2XCR ~RF~ =JR HAULAGE COSTS 269 in which L = length of bond in inches; R = resistance of one inch of cable or strands compos- ing the bond; CR = Contact resistance of bond terminals; RF = Resistance of one foot of rail; JR = resistance of bonded joint expressed in feet of rail. As most bond-testing instruments show the resistance of the bonded joint, as compared with the equal resistance of a certain FIG. 14. Results of 21 tests of channel-pin and compressed-terminal bonds. number of feet of unbroken rail, it is more convenient to express this value in such terms than in ohms. To illustrate the use of the above formula and tables, assume that a 40-lb. T-rail is to be bonded with compressed terminal flexible cable bonds, 4/0 capacity, %-in. terminals, 26 in. in length. The resistance of each joint when the bond has been installed is desired. By referring to the tables, the resistance of one inch of 4/0 cable is found to be 0.00000414 ohm, and the resistance of a 26-in. bond is 26 times 0.00000414, or 0.00010764. The resistance of a %-in. terminal is 0.00000053 ohm, and of the two terminals is 0.00000106 ohm ; adding the cable resistance and the contact resistance the total ohmic resistance of the in- stalled bond is 0.0001087 ohm. To express this in terms of 270 COAL MINING COSTS RESISTANCE AND CARRYING CAPACITY OF RAILS Figures Based on Rails Having a Ratio of 12 to 1, as Compared with Copper, and at 70 F. Weight, Pound per Yard Resistance, Ohms per Foot Carrying Capacity in C.M. 16 0.0000622 169,764 20 0.00004923 212,206 25 0.00003935 265,257 30 0.00003321 318,309 35 0.00002844 371,360 40 0.00002489 424,412 45 0.00002212 477,463 50 0.000019355 530,515 60 0.0000166 636,618 RESISTANCE OF SOLID TERMINALS Figures Based on a Pressure of 15 Tons per Square Inch of Contact Surface Diameters Resistance, Ohms 1" 0.0000008 f" 0.00000064 *" 0.00000053 f" 0.00000045 1" 0.0000004 RESISTANCE OF BOND CABLES PER INCH OF CONDUCTOR AT 75 F. Size Resistance, Ohms per Inch Capacity in Amperes 1/0 0.00000829 210 2/0 0.00000657 265 3/0 0.00000521 335 4/0 0.00000414 425 250,000 C.M. 0.0000035 500 300,000 C.M. 0.00000275 600 350,000 C.M. 0.0000025 700 400,000 C.M. 0.00000219 800 HAULAGE COSTS 271 equivalent rail length, divide by the resistance of 1 foot of 40-lb. rail, and the resistance of the bonded joint will be found to equal that of 4.2 ft. of unbroken 40-lb. rail. Trials made with standard bond-testing instruments show that when bonds are installed with reasonable care, the resistance of the joint will very closely approximate this calculated re- sistance. Abutting rail ends and clean tight splice bars may slightly lower the resistance of the joint, but this is quite neg- ligible and should be disregarded. Numerous tests made of com- plete haulage roads, installed under usual mining conditions, show that an average will vary but a fraction of a foot from the calculated resistance. The use of the tables will not only be found of benefit in learn- ing standards to test individual joints for their efficiency, but they can be used to great advantage in figuring voltage drop on proposed work. In calculating voltage drop on a circuit com- posed of a trolley and a rail return, a formula is generally used which is correct for the copper loss, but does not take into consideration the weight of the rail and the size and length of the bonds. By calculating the voltage drop on the trolley and feeders only, and then on the rail, when properly bonded, the exact drop for a given load is secured. This method has been found particularly advantageous where a potential of 250 volts is used and the current transmitted over long dis- tances, as is common with bituminous mines. For instance, taking the example mentioned above, assume that a road 3000 ft. long is to be bonded, and the actual drop on the return side of the circuit is desired. With 30-ft. rail lengths, there will be 100 joints on one rail, having a total resistance of 420 ft. About 10 per cent should be added to the joint resistance to take care of short rail lengths and bond- ing at switches. Then the resistance of one rail will be equal to that of 3460 ft. of unbroken rail, or 0.00836394 ohm, or for the two rails in parallel 0.00418197 ohm. The voltage loss on the return side is the load times the resistance, then by calculating the drop on the trolley side, the exact drop on the circuit is easily ascertained. In the same manner, these tables can be used to test the efficiency of the bonding of an entire haulage road on a certain 272 COAL MINING COSTS section. With voltmeters at both ends of a section and an ammeter on a locomotive, the voltage drop at a certain load can be obtained. It is a simple matter to calculate the drop on the positive side of the circuit and on the return side, assum- ing the bonds are in good condition. Any difference found will be an increased joint resistance, and the individual bonds should be tested with a bond tester and the defective ones replaced. A number of companies have adopted this method, as an entire mine can be gone over on an idle day or night, and the individual joints need only to be tested in those sections which show defective bonding. Mine operators who are desirous of obtaining efficient and economical results from electrical min- ing equipment will find it well worth their while to make such tests following them up with any necessary repairs. One company recently tested in this manner a newly bonded road. The test showed that the joint resistance was 69 per cent of the total circuit resistance, where it should have been only 8 per cent. Upon making an examination of the road, it was found that the track men had neglected to install bonds at two of the switches. As their load was heavy, this unnecessary resistance would have cost a considerable sum in a month's time for power, which in this instance was purchased. The following is another example of computing bonding losses it being assumed that it is desired to know what is the difference in resistance between a 4-0 round copper wire 3000 ft. long and the two 25-lb. steel rails, in a track, 2000 ft. long, followed by two 30-lb. steel rails, in 1000 ft. of track. The rails are bonded with pressed-terminal all-wire 2-0 bonds. . Also determine how many kilowatt-hours will be available at the end of a 250-volt line, where a 30-hp. motor, 3000 ft. from the generator, is taking 100 amperes. The resistance of 1000 ft. of a 4-0 copper wire at, say 68 deg. F. (20 deg. C.), as taken from a table giving the resist- ances of copper wires for different gages and temperatures, in ohms per thousand feet, is 0.04893 ohm. The resistance for such a conductor 3000 ft. long is then, 3 X 0.04893 = 0.14679 ohm. An approximate rule for calculating the resistance of copper wire per thousand feet, in ohms, is to divide 10,000 by the HAULAGE COSTS 273 size of the wire in circular mils. Thus, for a 4-0 wire (211,600 circ.mils), the resistance is, approximately, #=^^=0.04726 ohm per 1000 ft. This rule should only be used in rough calculations. When accuracy is desired, the resistance for the wire should be taken from electrical tables, as above stated. The resistance of steel rails, for the same cross-section and length, varies with the composition of the steel. The presence of sulphur and manganese, particularly the latter, greatly modifies the resistance of the steel. It has been found that this resistance will vary from about eight to thirteen times that of copper of the same sectional area and at the same temperature. This has given rise to what is termed the "ratio of rail to copper" or the " rail-to-copper ratio. " The results of numerous experiments have made it possible to calculate the equivalent circular mils of copper corresponding to any given weight of rail in pounds per yard, for any rail-to-copper ratio. To do this, the weight of rail, in pounds per yard, is multiplied by the constant corresponding to this ratio, as determined by the composition of the steel. The value of this constant, for the several ratios, is as follows: Rail-to-Copper Ratio Constant Rail-to-Copper Ratio Constant 8 15,550 11 11,360 9 13,820 12 10,360 10 12,500 13 9,590 Applying this method and assuming a rail-to-copper ratio of 10, the constant for this ratio, as taken from the above table, is 12,500. Then, for a 25-lb. rail, the equivalent circular mils of copper is 25 X 12,500 = 312,500. Electrical tables are not generally extended to include as large a wire as this area indicates. The resistance in ohms per thousand feet, however, can be calculated, approximately, by the rule previously given. Thus, 10,000 -j- 312,500 = 0.032 ohm. The resistance of the two 274 COAL MINING COSTS rails, in the first 2000 ft. of this track, is the same as the resist- ance of a single 25-lb. rail 1000 ft. long, or 0.032 ohm. Again, for a 30-lb. rail of the same composition, the equiv- alent circular mils of copper is 30 X 12,500 = 375,000. The corresponding resistance is, therefore, approximately, 10,000 -f- 375,000 = 0.0267 ohm. The resistance of the two rails, for 1000 ft. of track, is one-half of this amount, or 0.01335 ohm. The total rail resistance in this track is, therefore, 0.032 -\- 0.01335 = 0.04535 ohm ; and the difference, in favor of the iron rails, is 0.14679 0.04535 = 0.10144 ohm. The diagram, Fig. 7, taken from the Ohio Brass Co.'s catalog, shows graphically the circular mils of copper, of equal electrical resistance to steel rails of different weights and " rail- to-copper ratios. " The curved lines show the resistance, in microhms, of steel rails of different weights and ratios. The diameter, in inches, of a copper wire that is the elec- trical equivalent of a steel rail of a given weight (Ib. per yd.) may be calculated by multiplying the respective constant taken from the above table, by the weight of the rail, extracting the square root of the product and dividing that result by 1000 Thus, for a rail-to-copper ratio of 10, the constant is 12,500. Then, the diameter of the copper-wire equivalent is: V 25X12,500 As regards the second problem, a 30-hp. motor consumes 30X746 = 22,380 watts or 22.38 kw. If this motor is taking, as stated, 100 amp., the voltage, at the full capacity of the motor, is 22,380-^100 = 223.8 volts. The drop in voltage for this line is, therefore, 250-223.8 = 26.2 volts. The work performed by this motor, in each hour, when working at its rated capacity, is 22.38 kw.-hr. Tests made at a colliery in Pennsylvania showed that the bonding was so bad that the return current was leaving the rails and finding its way to the generating plant by way of the ditches and water pipes, which, of course, had a high resist- ance. The locomotives with a rated speed of 6.3 miles per hour were found to be traveling at about 2.5 miles and making an average of 12 trips per day. The actual load on the generator HAULAGE COSTS 275 was 30 per cent over the rated load on the line. The main haulage roads were bonded with compressed terminal bonds of sufficient capacity to equal the size of trolley and feeders. The first month following this installation ike production was the largest in the history of the mine. The locomotives instead of making 12 trips per day were averaging 18 trips. Later another locomotive was added and the whole load on the generator was less than that carried before the bonding was changed. At another colliery the bonding resistance of the main haulage road was reduced 80 per cent by the use of compressed terminal bonds. During the last six months of the channel- pin installation the track bonder had averaged three days per week on this road replacing broken and defective bonds. Dur- ing the first seven months of the new installation the bonder spent two hours on the road replacing some bonds broken by a wrecked trip. Investigation of bad haulage conditions at another mine disclosed a reading on the voltmeter of 180 volts on a 250-volt circuit at a point 6000 ft. from the mine mouth. A 10-ton motor was used on the main haulage in to this point, beyond which there were three motors engaged on secondary haulage. When the main haulage motor started out with a trip the current would drop to 80 volts and remain at this for 8 to 10 min. during which time the gathering motors could not move. It was estimated that the haulage charges at this mine were increased about lOc. per car or 4c. per ton from these causes which on the basis of a working schedule of 20 days per month would entail a loss of $800 per month. TABLE SHOWING COST OF ENERGY-LOSS IN RETURN CIRCUIT PER 100 AMPERES LOAD, PER YEAR OF 240 9-HouR DAYS, POWER COSTING ONE CENT PER KILOWATT-HOUR (1913) One Mile. Two 42-lb. rails in parallel, assuming continuous joints. . $15.78 One Mile. Two 42-lb. rails in parallel, bonded with channel-pin bonds, resistance found by actual test 37 . 80 One Mile. Two 42-lb. rails in parallel, bonded with compressed termi- nal bonds, resistance found by actual test 17 . 99 Fixed cost of rail resistance 15 . 78 Increased cost with compressed bonds 2 21 Increased cost with channel-pin bonds 22 . 02 Saving per year with compressed bonds for power alone per 100 amperes 19.81 276 COAL MINING COSTS A method to overcome bonding troubles and to reduce the bad effects to a minimum is to put in ground wires along the track and to tap these, at suitable distances, to the rails as shown in Fig. 15; this insures the bonding and provides an uninterrupted metallic circuit. The high price of copper prac- tically prohibits its use for this purpose. Wornout wire hoisting ropes, in lengths of 500 ft. and over, which make few joints necessary, having a scrap value of approximately $8 per ton (as of 1912) and a specific resistance of 8 to 1 as compared with copper, may be utilized for this purpose. It is a better conductor than rails and a rope of, =: r: EEI FIG. 15. Method of improving the bonding' by the use of old wire cable. say, l l /2 in. diameter has about the same current-carrying capacity as a No. 000 B. & S. gage, or about a %-in. copper wire. Smaller ropes may also be used, the size required depending upon the conditions. The wire rope may also be used where ground wires are required, as between substations and the mines. For this work, copper cables 1 in. in diameter are quite frequently used. An installation requiring 1000 ft. of such copper cable was esti- mated at approximately $550 in 1912, and the same result would be obtained with wire rope at a cost approximating $100. For such outside work 1%-in. and 2-in. ropes, which cannot be 'handled very well inside, may be utilized. The installation should be made in the following manner: From the substation to the entrance of the mine, where no rails are used, 1%-in. or 2-in. rope, either single or in multiple, HAULAGE COSTS 277 according to the load on the circuit and the distance, should be used. Along the track in the main gangway 1%-in. rope, and for extensions 1^4 and 1-in. single ropes are suitable. The rope is stretched out by attaching it to a mine motor or locomotive and in this manner several thousand feet may be laid in a few hours. As the ropes are seldom shorter than 500 ft., only a small number of joints, which should be made with sub- stantial 10-in. clamps, are necessary. Sheet-lead lining should be used between the clamps and rope, thus providing a reliable contact. One of the bolts used with the clamps serves as a bond terminal. Where the rail taps make connections with the ropes, smaller clamps are used at distances of about 200 to 300 ft. It is advisable, in order to increase the durability, to put the rope on the high side of the track to keep it out of the water as much as possible. The efficiency of the return track, either single or double ^Ki* bonded, and provided with such a ground wire, will be very little affected by a few defective bonds, and the practical results will be far superior than with the regular bonding only. On account of the stable conditions obtained. with the ground wire, in many cases single bonding with some cross taps between the rails will be sufficient; where distances are short and the load light, satisfactory results will be attained by the ground tapped to the rails at suitable distances with- out the regular bonding. The inspections of the track in regard to bonding will also be reduced to a minimum. At one operation where three SV^-ton electric mine loco- motives were working there was considerable complaint due to the inefficiency of these motors and in fact a request was made for an additional motor. An examination of the place showed that there was lost time for power, a large amount of sand was being used on the track and trolley wheels did not last longer than a day or two. The track at this operation was laid with 40-lb. rail at the bottom of the shaft and 30-lb. rail in the gangways. The bonding was about equal to the average con- dition of bonding in mines. As promptly as possible, 7500 ft. of wire rope was installed, 2000 of which was iy 2 -in., 4000 114-in. and the balance 1 in. in diameter. The results were entirely satisfactory. The motors could then do good work, the amount of sand used on the 278 COAL MINING COSTS track has been reduced 65 per cent, which also reduces to a minimum the wear on the motor wheels. There is no further complaint regarding power or poor trolley wheels. Track costs. The determining factors in fixing grades in the mines are drainage, haulage and sometimes loading. The grade sought after in the mines of this country is 4 to 6 in. per 100 ft. There is a theoretic grade at which the inclination of the haulageway will compensate the added resistance due to loading the cars and thus equalize the load on the motor for both empty and loaded trips. When this condition has been realized the motor is able to handle the same number of loads and empties going in their respective directions. The greater the difference in weight between the loaded and empty car, the steeper should be the grade in favor of the load, to overcome the extra resistance. Consequently when- ever a larger capacity mine car is adopted, a corresponding change in the standard grade should be made for all haulage- ways not influenced by other considerations. This, however, is seldom done, the original grade being maintained. On haul- ageways where the 0.35 per cent grade is used with cars of 3 tons capacity, the number of empties it is possible for the motor to haul invariably exceeds the number of loads. An all-steel car (115 cu. ft. capacity at water level full) empty weighs 5080 Ib. and loaded with coal weighs 12,230 lb., while loaded with rock it would weigh 17,000 lb. Assuming the sum of the resistance due to friction and track to be 30 lb. per ton for an empty car and 25 lb. per ton for a loaded car, the coefficient of friction per short ton for loaded and empty cars will be 0.015 and 0.0125 respectively. With an 8-ton motor having a rated drawbar pull of 3000 lb. on the level, letting N = the number of cars and g = the theoretic grade for equalizing the drawbar pull, then when the motor is handling coal-loaded and empty cars: 3,000+16,000g 3,000-16,000^ (12,230 X 0.0125) - 12,2300 ~~ (5080 X 0.015) + 50800 = 5 in. per 100ft. With the same motor handling empty and loaded rock cars: 3000+16,0000 3000=16,0000 17,000X0.0125-17,0000 = 5080X0.015+50800 = 7 in. (almost) per 100 ft. HAULAGE COSTS 279 The foregoing is a somewhat involved quadratic equation not always easy of solution. A method which is perhaps simpler is as follows: Let 6 = desired angle of grade to give equal drawbar pull in both directions, then the percentage of grade, or the rise in unit length of track, is practically equal to sin 6. If W weight of empty car; W' = weight of loaded car; C = coefficient of friction for empty car; C' ~ coefficient of friction for loaded car; then WC+ W sin 6 = W'C' - W sin 8. Substituting values in the above case (5080X0.015) +5080 sin =(12,230X0.0125) -12,230 sin 76.2+5,080 sin 0=152.875-12,230 sin 17,310 sin = 76.675 0.00443 Multiplying by 100 to secure the rise in 100 ft., we have 0.443 ft., or about 51/4 in. With straight track, then, the 8-ton motor on a grade of 7 in. per 100 ft. could pull in 28 empty cars and return with the same number loaded with rock; but as the resistance of the curves, of which a major portion of every gangway con- sists, is a corollary of the weight, and determined grade should be augmented somewhat in favor of the loaded cars to partly compensate for curve resistance. The value of the track resistance per ton on straight track will have to be determined experimentally for any type of car. Equipping the cars with patent bearings will have salutary effects, should this resistance prove too high. On a recent test of nine cars, three each of the three various types ordinarily employed, on fairly clean straight track, on the surface, with cars weighing 12,230 Ib. each; loaded with coal, capacity 115 cu. ft. loaded water level; wheel base, 3 ft. 6 in. ; gage of track 3 ft. 6 in. ; 40-lb. rail ; both wheels tight on the axle ; center of drawbar pull to top of rail, 18% in., the tractive effort required was as follows: 21.2 Ib. per ton for cars with 160 deg. iron case, babbitt COAL MINING COSTS lined; 10.5 Ib. per ton for cars with 160 deg. brass lining; 6.5 Ib. per ton for cars with Hyatt bearings. The cars were of all-steel construction identical in every respect, except the bearings. The brass-bearing cars were new and in first-class condition, while the Hyatt and babbitt-bear- ing cars had been in use almost 2 yr. The results, if secured on tracks underground, in their usual dirty condition and with inferior ballast, would no doubt have been higher. A contingency not always considered in planning a haulage is that, since the abandonment of mule haulage, the loaders are compelled to move the cars at the loading chutes by their own exertions. Four cars per trip are not infrequently loaded from one chute. On the prevalent 0.35 per cent grade to move the cars the distance necessary to accomplish this would require an additional man. To dispense with this, the foreman resorts to increasing the grade immediately beneath the chutes with an equalizing diminution between them. This expedient per- mits loading the cars, but makes the roadbed a series of undula- tions with pools of water in the "dead spots." It raises the haulage cost, destroys the rolling stock and sets the cars bumping and jerking over tlie entire working section. It is needless to mention that a heavier grade would abate, if not wholly remove, this condition. To be in congruity with the preceding, a grade of 7 to 8 in. per 100 ft. for the haulage and up to 8^ in. for the primary gangways would seem to have more to recommend it than the lesser grade. To be sure, the first installation ,of the heavier grade would reduce the available lift as the gangway advanced, but this objection would remedy itself in all sub- sequent levels. Again, a ditch averaging 18 in. wide edge to edge and 6 in. deep in the center, on a 0.35 per cent grade, would give 38.4 ft. velocity and almost 108 gal. capacity per minute, against 54.4 ft. and 153 gal. for an 8%-in. grade. If an analysis of any haulage problem indicates that a saving is possible by cutting down grades, it becomes at once necessary to determine the approximate amount that can be profitably spent on the proposed improvement. The advantages accruing from the improvement will be: (1) Reduction in general labor costs; (2) reduction of power HAULAGE COSTS 281 cost per ton hauled; (3) reduction in general expense, includ- ing the saving made possible by the postponement, either tem- porarily or permanently, of expenditures for additional equip- ment such as motors, generating units, engines and boilers. If a haulage motor pulls a certain number of extra cars per trip as a result of a grade reduction, the total number of tons produced at the mine will be increased without a pro- portionate increase in the expense. This saving is expressed by a formula in which let : t = tons produced per shift before increasing the tonnage; jP=tons produced per shift after increasing the tonnage by pulling extra cars per trip; c labor cost to produce t tons, in dollars; C = labor cost to produce T tons, in dollars; S = saving per ton in dollars resulting from the increased tonnage. We then have the formula: c C s= t~r For example, a mine with one main-line haulage motor making 16 trips in 8 hr. from two partings produces 950 tons per day at a cost of $145.38 for inside labor and $52.69 for outside, making a total of $198.07. Investigation of the haulage profile shows that by reducing the grade on one of the runs the motor will be able to handle 1000 tons per day. To handle this increased tonnage an extra driver inside will be required and a track layer, one hour per shift, bringing the total labor cost including full allowance for feed, car and depreciation of the mule up to $202.45. Sub- tituting these values in the above formula, we have : 198.07 202.45 "950" "TOOO" "When the summit in a grade is lowered, the power required to overcome grade resistance is less and by increasing the number of cars per trip less power is required per ton of coal hauled since the proportionate ton-mileage of the motor itself is reduced. The weight of the motor inbound may be about 282 COAL MINING COSTS one-third of the total weight of the trip, and outbound about one-fifth, so that it is evident that considerable power is con- sumed in moving it alone. A heavy pull exerted on a stiff grade will tend to increase maintenance and repair charges for both rolling stock and track. On the other hand a reduction in the grade may result in an increase in the car repair bill, due to the greater num- ber of cars handled, but the charges per car-mile and the maintenance per ton hauled will remain the same and will not effect the unit cost per ton of coal produced. The elimination of bad grades also reduces the hazard to operations and though the danger to life is not directly cal- culable, it is of vital importance and must not be overlooked in considering the possibility of any proposed improvement; in other words it is simply applied safety and could properly be included under the charge for insurance. If the advantages resulting from a certain grade elimination can be reduced to cents per ton handled over the section of track improved and this is multiplied by the number of tons so handled, the result will be the amount that may be expended on the proposed work. Or expressed in a formula, let: 2/=the estimated number of tons available; C = cost of grade reduction, including interest on money invested; S = summation of all savings, in cents per ton; V = value of safe operation. Then (yS + V) should be equal to, or greater than, the value of C. If the value of yS (total saving in dollars) is less than the cost of the improvement, and the value of safe opera- tion does not, in the opinion of the management overcome the difference, then the project should be abandoned. As a matter of fact the judgment of the financier must be relied upon throughout the whole study of the question. False assumptions may be made in some cases, leading to erroneous conclusions, but by following carefully the steps indicated, it is possible to mafce a fairly accurate estimate of the results to be obtained in any contemplated work of this description. Haulage grading estimates. When all the data for a pros- pective change of grades has been assembled, the estimated HAULAGE COSTS 283 cost of the various plans, routings and schemes should be made for purposes of comparison. An intelligent estimate of cost must consider detail and be based on accurate knowledge of the proposed requirements, together with the application of the unit costs of similar work formerly completed. A con- FORM No.l ABC COAL COMPAls General Estimate Propose IY d Extension of Motor Road Pn,H~a Ft. anil Construction pf ft. MINE N<- ITEM Tool Rail* Laid Taken up 35*; ii*"i** "wliki** prti *cS" air * Tou Rail. UN Taken op Kegs Small Spikes Site Kegs Motor Spikes Site Site Sm.ll Tlea Pn FUB Plate* No. Rail* No. Rails Ken Track Bolu No. Rail* No. Rail* Bono* . Bonding Cap* Bonding Sleeves Trot* anil Switege* No. Rail* No. Rails 2|o Trolley Wlrs 4|0, Trolley Wire Hancor* and Clamps Wire Splicing, Sleeve* Trolley Fret* Automatic Cut-out Swltcasr Trolley wire Guard* Intulated Telephone Wire Porcelain Insulator* and Pitt Material to os DeliTsrsd Cleaning Oob. Etc. _ Griding.. Yd*. Top Griding. . .'.Yd*. Bottom. Setting Props Setunc Timber Sets and Cross Ban Mlae. Eixnie Total Eitlmited COM Credits of Material not applied 19 estimate. Small Tit, Total Value Credits to be Deducted Grand Total FstlmAtiJ Cost The extension of this Motor Road will AftMftsffe JffMMrffe FIG. 16. Form for assembling estimate of cost for grade revision in a motor road. venient form for assembling an estimate of cost of a motor road extension or revision is shown in the form Fig. 16. The detail required is not exacting, yet a fairly complete record is indi- cated. When it is finally decided to carry through the work a final estimate is made and a copy bearing the official signature 284 COAL MINING COSTS of approval sent forward to all departments concerned. The record is thus made complete. Economy in the execution of the work will depend on the evolution of a systematized method and a strict adherence to two fundamental principles of good management : First the FORM No. 2 DAILY TIME AND PROGRESS CHART MOTOR ROAD CONSTRUCTION Name Cefvn Amoll E2 Ma nws Hgyr MATERIAL RECEIVED PROGRESS CHART Note: Place a cross n -the squareslo indicate the portion completed between each station Cleanin Takin u Track Drillin and Shoot n Loadinqjfock Drivers or Motorman Unloadinqjtock 3onding &, Hanging Wir Timberin and Misc. XXX XIX Correct ^n tAmitfi, Foreman- Checked Wi vam> Supt. FIG. 17. Progress chart for use in analyzing grade revision costs. number of men employed in any one section in any period shall be adjusted to the amount and classification of the work to be done and, secondly, only experienced men should be employed. A careful study of the progress made from day to day will plainly show any weak points in the organization and will often indicate a probable remedy. Progress charts, coupled HAULAGE COSTS 285 with the report of the daily time, are useful in judging the comparative efficiency of the organization as a whole or in part and also gives a record of the unit costs for making future estimates and comparisons. The accompanying form, Fig. 17, shows a good method of accomplishing this. The progress for any particular section can be indicated in the daily report by filling in the squares of the progress report at the bottom of the form with crosses to express the estimated amount of work completed in that section. Thus a daily chart can be sent to the administrative department as a record of both the work done and the efficiency of the organization. Haulage track curves. Viewed solely from the haulage standpoint, the determining factors of the curve radius can be covered by two heads: 1. The cost of resistance due to curvature on the total estimated number of cars that can be hauled. 2. The probable number of cars to be hauled in each trip and the speed of haulage. The amount of resistance due to curvature varies with each type of car, and to a lesser degree with each car of a certain type. This resistance expressed in terms of grade, with curves of from 30 to 100 ft. radius, will run 0.015 ft. to 0.025 ft. per 100 ft. of track for each degree of curvature. That is, with a 50-ft. radius, or 115-deg. curve, moderately clean track, fair running cars with both wheels keyed on the axle, approximately a 1.8 per cent down grade would be necessary to secure the same drawbar pull as on a tangent. The value of a curve expressed in degrees can be obtained by dividing 5730 by the radius in feet. This formula will have to be employed especially in small radius curves; the actual arc is used to find the degree, rather than the central angle subtending the 100-ft. cord, the practice on standard-gage roads. By using the actual arc, a 50-ft. radius = _ =115 deg. curve; ou 50 by using a 100-ft. cord, a 50-ft. radius = . 1 , = 180 deg. curve, sin -^oL showing a disparity of 65 deg. Assuming a curve with a central angle of 90 deg. a ruling grade of 0.5 per cent and allowing the same rate of resistance per degree on a 25-ft. and 50-ft. radius curve, the motor in traveling over the two would have to work equivalent to 286 COAL MINING COSTS mounting a 4.5 per cent grade for 39 ft. and a 2.5 per cent grade for 78 ft. respectively. From the beginning of the 50-ft. radius to the point of tangent there would be a total of 1.96 ft. ^\ vertical, while to travel between the same points by way of the 25-ft. radius curve, including the 25 ft. of tangent on each end of the curve, there would be a total of 2.02 ft. vertical, or essentially the same vertical rise in either case. While actually with the smaller radius curve there would be a lower rate of resistance per degree, this would be more than balanced by the increased resistance due to the slower speed compelled by the sharper curve. If the resistance due to grade and curvature between the similarly located points is accepted as equal, then there remain in favor of the 50-ft. radius the greater speed at which the trip can travel, the reduced danger from cars jumping the track, the better adher- ence of trolley to wire, and 11 ft. shorter haul, and under -% A some conditions 11 ft. less of track. With a gangway produc- ing six trips per day of twelve 5-ton cars each, this 11 ft. twice per trip would consume enough power to draw 1 ton 7920 ft. each day, or 375 mi. per year. A self-recording dynamometer will reveal the frictional resistance of any type of car, and the monetary value per unit of haulage can be readily ascertained. In estimating the number of cars per trip, the future out- put, as well as the length of haul, must be considered. A haulage over which 100 cars travel per day may be increased threefold by a tunnel to the veins. This will mean the instal- lation of a larger motor if the haul is long, or possibly the use of two motors. The maximum speed of haulage under- ground being fixed by law (6 mi. per hour), nothing can be expected from faster transportation. A 15-ton locomotive will not traverse curves possible to an 8-ton machine, and this fact will demand an extra motor, with its attendant expense. For obvious reasons no compensation is allowed for curva- ture underground; and if a motor is required to work at its capacity, the additional resistance to be overcome, because of curvature, will be the factor limiting the length of the trip. With the large curve a locomotive may pull through on its potential velocity, but on a curve of 25 or 30 ft. radius, the velocity will have to be reduced before reaching the curve. HAULAGE COSTS 287 Rails. Much extra expense can be incurred by not having the weight of rail used in the track properly proportioned to the weight of the motor operating on it. Where the rail is too heavy, there is an unnecessary expenditure in first cost and where too light, as is more frequently the case, costs of track maintenance, together with extra wear and tear on motors, due to poor track conditions, will mount up rapidly, though perhaps not be so evident. Working under average conditions, the Baldwin-Westinghouse Co. recommend the fol- lowing minimum weight rails for general mine service with motor haulage : Weight of Motor in Tons Weight of Rail in Pounds per Yard 4 to 6 16 6 to 8 20 8 to 10 25 10 to 13 30 13 to 15 40 15 to 20 50 Mines having average size mine cars customarily use 40 to 60 Ib. rail on the main haulage, 20 to 40 Ib. on secondary haulage and 16 to 25 Ib. in the rooms. An approximate rule sometimes used for this purpose is to have a rail that will weigh at least 4 Ib. per yard for each ton of weight in the locomotive. For example, a 4-ton motor should run on a 16-lb. rail; a 5-ton on a 20-lb. rail, etc. This rule gives somewhat excessive rail weights when applied to the heavier types of motors and the above table is preferable if available. In purchasing rail for mine use the buyer will require: (1) Stiffness, (2) strength and (3) durability rather than tons of steel. If the strength of various sections is compared, it will be found that these requisites can be purchased at a lower unit rate in the larger sections. In " stiffness" we have that property which allows the rail to span the ties and support the load without bending, affording thereby a smooth running surface for the cars; in "strength" we have that quality which 288 COAL MINING COSTS bears the load without yielding or breaking, while in "dura- bility ' ' we have the ability to resist wear over extended periods of time. The stiffness varies as the square of the weight, and the strength as the 3/2 power, while the price per ton is nearly constant. If the unit weight is assumed as being 30 Ib. per yd., then the stiffness will increase as follows: THIRTY POUNDS PER YARD STIFFNESS = 1 16f per cent increase in weight 35 Ib. per yard stiffness = 1 . 36 or a 36 per cent increase. 33 1 per cent increase in weight 40 Ib. per yard stiffness = 1 . 78 or a 78 per cent increase. 50 per cent increase in weight 45 Ib. per yard stiffness = 2 . 25 or a 125 per cent increase. 66f per cent increase in weight 50 Ib. per yard stiffness = 2 . 79 or a 179 per cent increase. 100 per cent increase in weight 60 Ib. per yard stiffness = 4 . 00 or a 300 per cent increase. The ultimate strength will increase as follows: THIRTY POUNDS PER YARD ULTIMATE STRENGTH = 1 16f per cent increase in weight 35 Ib. per yard ultimate strength = 1 . 26 or a 26 per cent increase. 33 per cent increase in weight 40 Ib. per yard ultimate strength = 1 . 54 or a 54 per cent increase. 50 per cent increase in weight 45 Ib. per yard ultimate strength = 1 . 84 or a 84 per cent increase. 66f per cent increase in weight 50 Ib. per yard ultimate strength = 2. 15 or a 115 per cent increase. 100 per cent increase in weight 60 Ib. per yard ultimate strength = 2 . 83 or a 183 per cent increase. The advantages of the heavy section over the light, as regards stiffness and strength, would show a higher comparison as the rail wears or wastes away from any cause whatsoever. In determining the durability of rail, it is obvious that a great amount of wear cannot be expected if the weight selected conforms closely to the immediate duty it has to withstand. We can assume for practical purposes that half the total weight is in the head, and that about half of this weight, or one-quarter the weight of the rail, can be worn away before the rail is discarded, if a sufficient margin of metal has been HAULAGE COSTS 289 allowed; otherwise, the rail will fail before it has attained much more than a high polish. In mining work, particularly underground, with the track- men in absolute charge, trackwork, derailments, rail breakage, etc., are taken as part of the day's routine and pass unnoticed, except that part which appears indirectly in the high main- tenance charges. If we assume that a wear of 1 / 5 the weight of the head was allowed as a safety factor in the lighter rail, then the durability of light and heavy sections will compare as follows : AVAILABLE FOR WEAB Spare Times Increase Metal in Increase in Left in Head Next of Wear Weight Weight in Pounds per Yard Weight in Head Only Maximum Half Head Minimum One-fifth Head After Minimum Wear Heaviest Rail before Head Becomes as Light by Adding 5Lbs. to Section by Adding 5Lbs. to Section 30 15.0 7.5 3.0 12 5.5 1.830 1/6 35 17.5 8.75 3.5 14 6.0 1.710 1/7 40 20.0 10.00 4.0 16 6.5 .625 1/8 45 22.5 11.25 4.5 18 7.0 .550 1/9 50 25.0 12.50 5.0 20 7.5 .500 1/10 55 27.5 13.25 5.5 22 8.0 .454 1/11 60 30.0 15.00 6.0 24 8.5 .420 1/12 Or, using 30-lb. rail as a unit, the metal available for wear would compare as follows: AVAILABLE FOR WEAR BEFORE Weight in Pound Weight in Head HEAD WOULD BECOME AS LIGHT Increase in Weight per Yard Only Maximum Minimum per Yard Per Cent Per Cent Per Cent 30 15 7. 5 or 100 3.0 or 100 35 17| 10.0 or 133| 5. 5 or 183! 16f 40 20 12. 5 or 166| 8.0 or 266| 33| 45 22! 15.0 or 200 10. 5 or 350 50 50 25 17. 5 or 233| 13.0 or 433! 66f 55 27* 20.0 or 266| 15.5or516f 83! 60 30 22. 5 or 300 18.0 or 600 100 290 COAL MINING COSTS Briefly, if we were about to build a permanent (so-called) narrow-gage road for mine traffic, for which 30-lb. steel would ordinarily be used, we would gain, by using a 60-lb. section, the economy in maintenance, a more easily operated road with its attendant benefits, fewer ties, fewer derailments and a larger scrap value when the rail was reclaimed. Furthermore, we would have a stiffness four times, an ultimate strength 2.83 times and a durability three to six times as great, for a rail expenditure but double that for 30-lb. steel. Some concerns, by purchasing "second" rail from the rail- road companies, obtain the heavier rail for the same price per lineal foot as for new sections one-half to two-thirds their weight. This quality of rail for most mining purposes will serve as well as new sections. In localities where acid water abounds the corroding of the steel is frequently the limiting factor in the life of the rail. It would be futile to lay heavy section rail in locations where the water would soon destroy it. As the web and edges of the flange are the portions destroyed first, an inspection of the standard dimensions will evidence that by increasing the weight we do not secure a proportionate increase in the acid- resisting properties of the rail. Rail weighing 25 Ib. per yd. has been taken as the basis of unity. Weight of Rail Increase in Weight, Per Cent Thickness of Web Increase in Thickness, Per Cent Thickness Ends of Flange Increase in Thickness, Per Cent 25 it ... H 30 20 ft 11 ft 35 40 H 21 9 40 60 it 32 H 27 45 80 H 42 H 35 50 100 If 47 H 36 60 140 ft 63 H 64 In the standard tee rail, adopted by the American Society of Civil Engineers, 42 per cent of the metal is in the head, 21 per cent in the web and 37 per cent in the flange. The top corners are curved to a 5 /i G -i n - radius, and the car wheels are designed to give on this as little friction as possible; as the HAULAGE COSTS 291 rail more nearly wears to the shape of the flange the friction is augmented. The height of the rail is identical with the width of the flange, so if this dimension is measured the weight can be determined. The table shows the weight of rail per yard corresponding to the height of flange width. Track frogs. Standardization of switches and frogs at mines to a limited number of sizes to meet requirements will substantially lower the cost of making these. The accompany- ing illustration, Fig. 18, shows a standard frog and switch used by the O'Gara Coal Co. in 1916. iof7Je. OB End Elevation Side Elevation Switch FIG. 18. Standard frog and switch used by the O'Gara Coal Co. The designs were made after considering both simplicity and economy of construction. Any ordinary blacksmith or ironworker will make these parts without difficulty. The cost of making a No. 5 frog and two 6-ft. switch points at a well- equipped mine shop was about 1916: Material, $6.37; labor of blacksmith and machinists, $6.58 ; total, $12.95. A cast-steel frog supplied by manufacturers at a cost of about $6.50 is inherently more rigid than the riveted frog, but it is difficult to fasten it securely to the ties. In order to stiffen the riveted structure, cast-iron fillers may be added which also support the flange of the wheels in passing over the throat of the frog, thus relieving the jar to the rolling stock. 292 COAL MINING COSTS DIMENSIONS OF STANDARD FROGS AND SWITCHES FOR NARROW-GAGE INDUSTRIAL AND MINE TRACKS Standard Frog for Motor Turnout (Right or Left) Frog No. Frog Angle, X Rail, per Yard Length of Frog, A Wing Rail, B Heel Distance, C Length of Throat, D Straight Rail, E Deg. Min. Pound Ft. In. In. Ft. In. In. Ft. In. 3 18 55 30 4 16 2 8 3 i| 2 3 4 14 15 30 4 8 20 3 5| 2 9 4 14 15 40 4 8 20 3 6| 3 5 11 25 30 4 10 20 3 2 7& 3 5 11 25 40 5 20 3 4 81 3 Standard Switch for Motor Turnout (Right or Left) Weight Length Distance Length Length of of between of Rail Gage of Rod Rail, Point, Bridle Rail Punching, of Bridle Punch- Pound F Rods Planed, Track, Rod, ing, per Ft. In. L H J K G M N Yard Ft. In. Ft. In. Ft. In. In. Ft. In. In. 20 4 (1 rod) 1 4 4 2 36 5 SI 25 25 4 (1 rod) 1 6 4 2 40 6 Oi 29 30 4 (1 rod) 1 9 4 2 42 6 21 31 30 6 2 3 2 6 4 2 44 6 4| 33 30 7 6 3 3 4 2 48 6 8| 37 40 6 2 3 2 9 5 2i 40 7 6 3 3 6 5 2| The throw of switch point is 3^ in. all for cases. In designing various parts of the turnouts it was kept in mind that all such turnouts may be of only temporary useful- ness in one particular location and that the constituent parts may be used many times before being cast aside as useless. The standard frog is somewhat shorter than one designed to the specifications ' of the American Railway Engineering As- sociation, but the saving in weight and bulk, with the conse- quent saving in making the several installations, will more than offset any loss due to instability. HAULAGE COSTS 293 The cost of laying and ballasting a No. 5 turnout complete, as shown in Fig. 19, was about 1916 as follows : One 30-lb. No. 5 frog and two 6-ft. points $12.95 40 ties, 5X6 in., at 20 cts 8.00 Spikes, bolts, tie-plates, etc .75 1 low switch stand and rods 2 . 25 2 headblocks, 5 X6 in., 8 ft. long 1 .00 Total material $24.95 Laying, 16 hr. at 35? cts 5.68 Ballasting and surfacing, 8 hr. at 35| cts 2 . 84 Total labor $8.52 Total cost of material and labor $33 . 47 K- FIG. 19. Standard turnout used by the O'Gara Coal Co. The dimensions of the standard frogs, switches and turn- outs are given in the accompanying tables. The formulas used for the turnouts are as follows : G-B sin X-F sin Y Chord length U-- Radius R- sin% (X+Y) G-B sin X-F sin Y -K7; cos Y cos X Lead S=(R+%G)(sin X-sin Y) +BcosX+F+0 ; in which X=frog angle; F = angle of point rail; B= length of wing rail; F = length of switch rail; = distance from actual to theoretical frog point: G = gage of track; R = radius of turnout. 294 COAL MINING COSTS The dimension of was taken as 2 in. and the heel distance of switch points as 4^4 in. The spacing of the ties depends on the size of the tie and the style of the turnouts. If the regular set of switch ties is used as in standard-gage trackwork, 5 X 6-in. ties spaced 18 in. center to center will give good results for track laid with rails of up to 40 Ib. in weight. If the turnout is laid with ties of even length staggered in as shown in Fig. 19, a spacing of 16 to 18 in. for each branch has proved satisfactory. This style of construction is specially well adapted to underground turnouts, where headroom is limited and flat ties 3 X 5 in. or 3 X 6 in. are used. All switch ties should be of hardwood and treated if possible, as decay will set in before mechanical wear destroys their usefulness. DIMENSIONS FOR TURNOUTS IN NARROW-GAGE TRACKS Gage of Track, 36 In. Frog No. Length of Switch Points, Radius of Turnout, Length of Lead, Length of Straight Rail, Chord Curved Rail, Mid- ordinate Curved Rail, F R S T U V Ft. Ft. In. Ft. In. Ft. In. Ft. In. In. 3 4 42 7& 16 3H 10 9H 10 10& 4& 4 4 81 1| 19 3 13 5 13 8 3| 4 6 75 8 22 6 14 8 14 lOf 4& 5 4 141 71 22 2f 16 41 16 7A 2| 5 6 126 7 26 0| 18 2| 18 41 m 5 71 122 91 28 4 19 19 2| 4& 6 n 184 9f 30 9f 21 5| 22 9 4A Gage of Track, 42 In. 3 4 52 Iff 18 91 12 11| 13 31 4H 4 4 99 2| 22 3 16 5 16 8| 4| 4 6 92 6| 25 9 17 11 18 2| 6J 5 4 172 0| 25 9| 19 lit 20 2 31 5 6 153 8f 29 11| 22 \\ 22 31 4H 5 71 149 If 32 5| 23 \\ 23 4 5^ 6 7 223 51 36 1\ 27 3| 27 6 5^ Track. The following is an interesting example of estimat- ing the cost of laying 5000 ft. of track where the grade is 1 per HAULAGE COSTS 295 cent in favor of the loads ; a 12-ton motor is used and 3*/2 ton cars assuming that the rails cost $26 per ton, ties lOc. each, spikes $3.75 per keg of 200 lb., labor for trackmen $2.50 per day, and helpers $1.75 per day, these figures being as of 1911. The rails for a 12-ton motor haulage should not be lighter than 40 lb. per yard, which would require (2 X 5000 X 40) -f- (3 X 2240) = 59.5 tons at a cost of 26 X 59.5 = $1547. For 40 lb. rails, use 3V 2 X Vie in. spikes, 12 kegs, at $3.75 per keg = $45 ; and 4 X 6 in. cross-ties, spaced 2 ft. center to center, 2500 at lOc. each = $250. There will be required also, using 24-ft. rails, 832 angle or fish-plates, 6240 lb., at iy 2 c. per pound $93.60, and 8 kegs bolts, nuts, and washers at $5 per keg = $40; making the total track material $1975.60. The laying and surfacing of 5000 ft. of track in mine entries, under ordi- nary conditions, including the handling of the material in the shaft and its distribution in the entry will require, approxi- mately 120 days' labor for helpers at $1.75 per day = $210; 50 days, trackmen, at $2.50 per day = $125 ; and 6 days, drivers, at $2 per day = $12 ; total for labor $347. The total cost of the track laid is therefore $2322.60, making no allow- ance for special grading which might be required at some points in the entry. A number of interesting figures on the comparative cost of track laid with steel and wood mine ties under varying con- ditions were given in a paper presented before the West Virginia Mining Institute, in 1913 from which the following have been excerpted: The Peyton Block Coal Co. used four steel mine ties per rail length which at a cost of 32c. each made a total of $1.28 per each pair of rails. Under the same conditions, 11 wood ties would be required, which at a cost of 5c. and allowing lOc. for spikes brought the total cost per pair of rails to 65c. Off- setting this difference, however, it was found that the miners would lay track with steel ties in their working places them- selves while they always insisted on the regular mine track layer performing the work when wood ties were used because of the special tools and labor required. It was believed that this saving in the time of the track layer compensated for the difference in the cost of the material involved, so that the extra life of the steel tie could be regarded as clear profit. The Allegheny River Mining Co. advances the opinion that 296 COAL MINING COSTS the life of the steel mine tie is about six times that of the wood tie and since only one-half as many are required per foot of track, the ratio of comparative utility was 1 to 12. On the basis of 32c. for the steel ties and 8c. for wood, which was the delivered cost to this company, and disregarding cost of the spikes required for wooden ties, the cost ratio is 4 to 1. Track costs at the Dartmore Mine of the Davis Coal & Coke Co., when laid with steel ties were found to be $1.28 per 30-ft. length for material. When using wooden ties, 10 of these were required per 30-ft. length of track which, at a cost of lOc. each and allowing 12c. for spikes, brought the cost of material for the track with wooden ties up to $1.28 per 30 ft. The difference in the cost of material for wooden and steel ties at this mine amounted to %c. per lineal foot of track, disregard- ing economies in laying, salvage, etc. as indicated above. At mines Nos. 14 and 20 of the same company, it was found that the steel tie saved sufficient head room to eliminate the necessity of brushing the top and effecting a computed saving of 63c. per yard of track. The Hutchinson Coal Co. compiled an interesting study of the comparative cost of steel and wooden ties at its Kirkwood Mine, near Bridgeport, Ohio, during the years 1907 to 1909 inclusive when it was changing from the wood to the steel ties. In the 1907 period all wood ties were being used; in the 1908 period 75 per cent steel ties were in use and in 1909 all steel ties were in use. The comparative figures are as follows : 19 07 19 08 19 09 Output Cost in Cents per Ton Output Cost in Cents per Ton Output Cost in Cents per Ton September. . 16,083 4 80 12,620 3 98 21,556 2.84 October 26,216 4.02 12,620 3.36 22,540 2.81 November 22,617 3 70 16,281 3.21 25,191 2.82 Average cost per ton 4 10 3.48 2.82 HAULAGE COSTS 297 The saving per ton with all-steel track thus appears to be 1.28c., which in the three months period of 1909 amounted to $886.85. This same company also compiled an interesting compara- tive estimate of the cost of track in a room, computed on the basis of a width of 24 ft., length of 200 ft. and a thickness of coal of 5 ft. 4 in. which worked out as follows : WITH WOOD TIES Ties, 1\ ft. apart, 80 ties at 12 cts $9 . 60 320 spikes equals 40 Ib 90 Laying and removing track, labor 7 . 50 Depreciation of ties and spikes 3 . 50 Total $21.50 Salvage 7.00 Net cost.. . $14.50 WITH STEEL TIES 35 ties at 33 cts $11 . 55 Labor, removing (track laid by the miners) 1 . 50 Depreciation 1 . 00 Total $14.05 Salvage 10. 55 Net cost.. $3.50 Estimating the output of coal from this room at 1000 tons on which a saving of $11 will be effected, this amounts to l.lc. per ton. After an exhaustive series of tests the Carnegie Steel Co. prepared the accompanying comparative cost of track laid in rooms with steel and wooden ties. The table is based on rooms 280 ft. long with steel ties spaced at 4 ft. center to center and wood ties 2 ft. center to center. This estimate contemplates using each wood tie in two consecutive rooms and after the second year renewing annually 15 per cent of steel ties, or 10 new ties per year, per room. 298 COAL MINING COSTS Number of Ties in One Room Steel, 70 Wood, 140 Cost of ties f.o.b. mine in carload lots 560 spikes @ 0.29 $20.30 @ 0.06 $8.40 @ 00| 2 80 @ 02 $1 40 @ 04 5 60 @ 01 70 @ 02 2 80 Maintenance cost for first room steel ties $2 10 2 . 10 Total cost at end of life of first room $22 . 40 $19 60 @ 02 $1 40 @ 04 5 60 560 spikes @ 0.00$ 2 80 Taking up when room worked out Maintenance cost for second room steel ties @ 0.01 .70 $2 . 10 2 . 10 Total cost at end of life of second room $24 . 50 $28 00 Saving per room in favor of steel ties $3 . 50 @ 0.29 $2.90 @ 05 50 $2.40 @ 0.06 $8 40 560 spikes @ 0.00^ 2 80 @ 02 $1 40 @ 04 5 60 @ 01 70 @ 02 2 80 $4 50 $4 50 Total cost at end of life of third room $29 . 00 $47 . 60 Saving per room in favor of steel ties $18 60 $2 40 560 spikes @ 0.00 2.80 @ 02 $1 40 @ 0.04 5.60 Taking up when room worked out @ 0.01 .70 Maintenance cost for fourth room steel ties . $4 . 50 $4 . 50 Total cost at end of life of fourth room . . $33 . 50 $56.00 Saving per room in favor of steel ties $22 . 50 The matter of track friction is important, and most mining men realize that there are material advantages in a good track. Few, however, really comprehend the reduction in power requirements that can be effected on much-used roads by mak- ing a strictly firstclass track in every respect. When we con- sider that it is possible to have a track friction as low as 12 Ib. of drawbar pull per ton, while as many roads show as much as 40 Ib. per ton, it is apparent that this is an extremely HAULAGE COSTS 299 important item. The use of heavy rails is an essential feature, but it is even more necessary that the track be kept clean. Mine cars. In 1911 a wooden car of 49 cu. ft. capacity, without brakes, cost at a certain mine in the neighborhood of $45 each. The same company was offered steel cars of the same outside dimensions, but having greater capacity, as fol- lows: Capacity Cost, Delivered Increase, Per Cent 7 cu. ft. $65.00 45 1 cu. ft. 62.50 38 1 cu. ft. 75.00 66 5 cu. ft. 68.50 52 The cost of an all-steel car thus ran from 38 to 66 per cent more than a wooden one, in some cases more, possibly, espe- cially in case of the addition of improvements in the way of draft gear, running gear or wheels. The manufacturers them- selves state the increase of cost to be from 50 to 100 per cent. Steel cars have a somewhat greater capacity than the wooden car, varying somewhat with the design of the car, but ranging generally between 10 and 20 per cent. The accom- panying table gives the comparative capacities and weights of some steel and wooden cars of the same dimensions, taken from actual practice: CAPACITY, CUBIC FEET WEIGHT IN POUNDS Wood Steel Increase, Per Cent Wood Steel Per Cent Gain in Capacity Saving in Weight, Per Cent 49 58 18 2010 1800 18.0 11.5 94 105 111 53 60 13 1920 1985 13.2 3.0* 14 19 35 1085 855 35.0 26.0 * Increase 300 COAL MINING COSTS Weights of both wooden and steel cars vary widely and it is difficult to make any accurate comparison. Steel cars are sometimes heavier than the wooden ones of the same capacity, but generally they appear to run from 10 to 20 per cent less in weight for the same capacity. The chief advantages of the steel car are : That although the first cost of the steel car is greater, the increased life and decreased cost of maintenance, together with increased capacity, thereby necessitating a fewer number of cars to handle a given output, more than make up for the difference. That the advantage of the steel car having a greater capacity with the same outside dimensions, or the same capacity with smaller dimensions, is of great value, especially in low veins. The increased capacity of the steel car should materially reduce the cost of haulage, and incidentally tend to increase the output of the miner. The saving it is possible to effect in the tare weight of the car itself would also be a factor in the reduction of costs by reducing the proportion of dead to live load. The steel cars will not warp, shrink or split, which are advantages that are apparent to all, besides preventing the leaking of dust coal on haulways not only a nuisance, but a constant source of danger and expense. On a special type of steel car, repair charges have been estimated at Ic. per 50 ton-miles. The total ton-miles on which this was computed amounted to 120,000 on which the repair charges were $22.50. The manufacturers claim that the cost of maintenance of steel cars is only about 20 to 25 per cent of that for wooden cars. From the experience of the railroad companies with their steel-car equipment, it would seem this cannot be regarded as an underestimation. Four different manufacturers estimate the life of the steel car to be two to four times the life of the wooden cars. The relation of the tare, or weight of the car, to its capacity is an important consideration in the economics of haulage, the tare of course representing dead load haulage. The accom- panying table, compiled from bulletins of the Illinois Coal HAULAGE COSTS 301 Mining Institute, and data received from car manufacturers give average ratio of capacity to tare for the normal mine car. LIGHT CARS MEDIUM CARS Tare Capacity of Cars in Pounds Ratio of Capacity to Tare Tare Capacity of Cars in Pounds Ratio of Capacity to Tare 800 900 1000 1100 1200 1300 1400 1500 1600 1700 1750 1900 2000 Averag 2000 3000 4200 2600 2600 4200 3000 3500 3000 5000 5600 4000 5000 e ratio 72-28 77-23 80-20 70-30 68-32 76-24 68-32 70-30 65-35 75-25 76-24 68-32 72-28 2200 2400 2525 2665 2850 Averag 6000 6000 4000 4800 4800 e ratio.. 73-27 71-29 61-39 64-36 63-37 66-34 HEAVY CARS 3240 3330 3500 3700 3780 Averag 6500 6000 6500 8000 6750 e ratio 67-33 65-35 65-35 68-32 64-36 72-28 66-34 At mines having cars already equipped with plain bear- ings which it is desired to change to roller bearings the cost under average conditions will be about $50 per car, this figure being as of 1917. Computing on the basis of 500 cars this would involve an expenditure of $25,000 less the salvage value of the old axles, journal boxes, and wheels which would be about as follows (figures as of 1917) : 1000 axles, 100 Ib. each, 2000 boxes, 25 Ib. each, 2000 wheels, 135 Ib. each, $20 per ton $1000 $15 per ton 375 $15 per ton 2025 $3400 The actual first cost of the roller-bearing installation would then be $25,000 $3400 = $21,600. The interest on this invest- ment is $21,600 X 6 per cent = $1296 per year. 302 COAL MINING COSTS Let it now be assumed that we can wipe out the initial invest- ment by creating a sinking fund. To provide a sinking fund for $1000, for example, over a period of 10 years, which is the con- servative life of the bearings if properly used, we must lay aside $75.87 each year. This is based on an interest rate of 6 per cent. For $21,600, we must lay aside $1638.79 each year. The interest on $21,600 and the sinking fund amounts yearly to $1638.79 + $1296 = $2934.79. Let us calculate what this saving amounts to in the case of 500 cars. Power costs, say Ic. per ton of coal hauled. Drawbar pull per ton (plain bearings) equals 32 Ib. -|- 20 Ib. for every 1 per cent of grade. Drawbar pull per ton with roller bearings equals 13 Ib. + 20 Ib. for every 1 per cent of grade. These figures are from dynamometer tests made by P. B. Lieber- mann, chief engineer of the Hyatt Roller Bearing Co. The saving in drawbar pull equals 52 Ib. 33 Ib. = 19 Ib. per ton hauled up a 1 per cent grade. The saving in power = if X Ic. = 0.4 mill per ton hauled. Each car, let us assume, hauls 6 tons per day. The saving in power per year with 500 roller bearing cars = 0.4 mill X 6 tons X 300 days X 500 cars = $3600. Plain-bearing cars require oil once a day. It takes two men at least to attend to the oiling. The cost of oil and waste for 500 plain-bearing cars per year equals $525. Cost of two men per year equals $1200. Total, $1725. The cost of oil and labor for 500 roller-bearing cars is $290. The saving is thus $1435. These figures are the results of carefully made tests. With plain-bearing cars it is necessary to use two men at the tipple, in order to push the cars. Since roller-bearing cars push with one-half the effort required for plain-bearing cars, it takes but one man to handle the roller-bearing cars at the tipple. For the same reason, the services of one eager can be dispensed with at the foot of the shaft. The saving on these two men equals $1875 per year, figur- ing that the man at the bottom costs $2.75 per day and the one at the top costs $3.50 per day. The total yearly saving on 500 roller-bearing cars is thus HAULAGE COSTS 303 estimated at: Power, $3600; lubrication, $1435; labor, $1875. Less a yearly cost of $2944.79, or $3965.21. The Hyatt Roller Bearing Co. conducted experiments with a dynamometer car to determine accurately the actual train resistance of cars equipped with their type of roller bearing and those equipped with plain bearings. One of these tests conducted at Greensburg, Pa., on a track with an average grade showed an average drawbar-pull of 12.8 Ib. at a speed of 5.98 miles per hour for the roller bearing and 24.3 Ib. at 5.59 miles per hour for the plain bearing. The advantage in favor of the roller bearings works out at 47.25 per cent. A second test at Carbondale, Pa., on an average grade of 0.45 per cent showed an average drawbar-pull of 13 Ib. at 7.8 miles per hour for the roller bearing as compared with 32 Ib. at 8 miles per hour for the plain bearing. The saving in drawbar-pull in this case amounts to 59.3 per cent. The diameter of the car wheel has an important influence on the power required to move the car. The smaller the wheel the more difficult it is to move. Cars move with less power on the narrower track gages as well, other conditions being equal. Wheel bases on mine cars rarely exceed 42 in. and the shorter this is the sharper the curve the car will negotiate. On a good clean, level track a man can push a car thai weighs iy 2 t ns an( i carrying a load of 2y 2 tons, making a total load of 4 tons, though the car will be difficult to start and , stop. It is probably inadvisable to have cars that have to be handled by men weigh when loaded over 3 tons and then the track should be in good condition. Many progressive operators are now providing a stretcher car for emergency purposes, a move that will be commended by all who have ever had occasion to assist in bringing a badly injured man from the mine. Fig. 20, shows an excellent type of car for this purpose, the cost of which was $80, including material and labor, in 1916. The design is quite simple, being an ordinary mine truck with the stretcher box supported on carriage springs. A box is slung from the truck portion of the car. In this will be kept bandages, salves and stimulants and a lungmotor. The car will be kept in a dry place specially 304 COAL MINING COSTS oo O oo o H 00 ? | ? 5 % O ^ <0 *x fcZ > h coo o o o * 00 oo *KO ( UI HAULAGE COSTS 305 provided for the purpose. This place will be at some point near the face of the workings. The length over all is 8 ft., the width 2 ft. 8 in., and the height of the box 12 in. The bill of material is as follows : Lumber : 2 pieces 4 X 6 in. by 8 ft. long, yellow pine. 2 pieces 2 X 12 in. by 7 ft. 8 in. long, yellow pine. 1 piece 2 X 10 in. by 7 ft. 8 in. long, yellow pine. 4 pieces 2 X 6 in. by 2 ft. 8 in. long, yellow pine. 4 pieces 2 X 8 in. by 2 ft. 8 in. long, oak. 2 pieces 2 X 4 in. by 8 ft. long, yellow pine. 4 pieces 1 X 12 in. by 7 ft. 8 in. long, yellow pine. 2 pieces 1 X 12 in. by 2 ft. 8 in. long, yellow pine. 1 piece 1 X 10 in. by 7 ft. 8 in. long, yellow pine. Iron and bolts: 60 % X 1%-in. carriage bolts. 40 l /2 X 4%-in. carriage bolts. 35 1/2 X 9-in. carriage bolts. 20 y 2 X 3-in. carriage bolts. 10 % X 3-in. carriage bolts. 1 iron % X 4 in- X 8 ft. 2 irons % X 4 in. X 20 ft. 10 irons % X 1% in. X 14 ft. 1 %-in. chain, 21 in. long. 2 irons i/ 2 X 2-in. X 30 ft. 10 irons % X 1%-in. X 14 ft. 2 irons % X 2-in. X 30 ft. 5 %-in. cut washers. 10 %-in. cut washers. 1 1-in. pin., 8 in. long, with 12 in. of %-in. chain attached, to couple to 4 common two-leaf buggy springs made of % X 1%-in. spring steel; length over all 3 ft. with 8 in. between springs. 1 set 14-in. car wheels. 2 2-in. standard axles, 3 ft. 1 in. gage. 306 COAL MINING COSTS Rope haulage. The Chicosa Fuel Co. in Colorado installed a rope haulage system about 1910, operated by a number of small electric hoists. Single-drum hoists are used on the cross- entries and double-drum tail-rope engines on the levels. The pitch on the cross-entries averages 8 1 / 3 per cent, the highest is 13% per cent. No trouble has been experienced in hauling 6 cars per trip, and at times 8 cars per trip. The car and coal combined weigh about 4900 Ib. The speed of the single-drum hoists working on the cross-entries is 300 ft. per min. on empty drum. This is what the manufacturer calls the starting speed. The 300 ft. per min. was considered the best speed for switching purposes, after several speeds had been tried. On the double-drum hoists, from 12 to 24 cars are hauled with the rope speed of 400 ft. per min. One engineer, one rope rider and one hoist can move as much coal as 8 or 10 drivers and mules, besides eliminating risks of killing both mules and drivers. It is a well-known fact that the depreciation on machinery is much less than that on mules. What it takes to feed three mules will pay one engineer, and what it costs to keep up two sets of mine harness will keep in repair an electric hoist. There are at present five 50-hp., single-drum, electric hoists and two 25-hp., double drum, tail-rope hoists in operation. The five single-drum hoists work on cross-entries, and the two double-drum hoists haul the coal from the crosses to the main- slope partings. The entire output of the mines which amounts to 1100 tons of coal per 10 hr. is moved with this machinery. None of the hoists are handling at present more than 50 per cent of their capacity, as the equipment inside the mine could easily handle 2200 tons of coal per 10 hr. With the present output the cost of hauling coal inside the mine is 35 per cent cheaper than if hauled by mules, and 15 per cent cheaper than if hauled by mine locomotives. On the present basis of the output the cost per ton inside the mine for hauling from the rooms to the main slope partings is 3c. per ton; by increasing the output to 2200 tons the cost would l%c. per ton. For each 1000 ft. the level entries are driven in, the cost will only increase 3 per cent. These figures are based on actual tests which are made daily, and all expenses, labor, oil, power, interest on investment, depreciation, maintenance HAULAGE COSTS 307 of hoists, power lines, ropes, and bell wire system, are taken into account. The electric-haulage system which is now being used is far ahead of the electric locomotive. Even where a room-gather- ing locomotive is used, the expense of keeping up trolley wires for running locomotives is 70 per cent higher than keeping up power wires in back entries. The danger to both men and mine from trolley wires is 100 per cent greater than from power wires for electric hoists. The troubles, expenses, and power losses in rail bonding for electric locomotives are entirely done away with in the electric hoist, and the cost of keeping up the electric locomotive will be more than keeping up the electric hoist and rope. With electric locomotives, heavy steel rails are necessary in the tracks, while ordinary 20-lb. steel rails answer the purpose where' the electric hoist is used. This item makes a difference of 50 per cent in track cost in favor of the electric hoist. The electric locomotives cannot be used in gassy mines, while the electric hoist can be used with perfect safety. Electric locomotives cannot be used to an advantage on grades exceeding 5 per cent, while with the hoist the grade makes no difference. Gravity planes. Some interesting cost figures on gravity plane haulage were contained in a paper presented before the "West Virginia Mining Institute in 1914. The figures applied to an installation at the mines of the LaFollette Coal, Iron & Ry. Co. in Tennessee. The plane operates through a vertical head of 613 ft. and a horizontal distance of 3640 ft. and the average inclination is 16.8 per cent. It is operated by seven men, two men and the drumman at the head house to attach the trips and two at the tipple to attach the empties with two others on tho plane to test the grips, oil rollers, etc. The rope for the plane lasted two years and cost $2000. In 1913 the plane handled 140,497 tons in 268 days of 9 hr. or an average of 524 tons per shift. The maximum tonnage handled in one day was 885 tons in &y 2 hr. which is at the rate of 104 tons per hour. It will be seen from the above that the depreciation on the 308 COAL MINING COSTS rope amounts to 0.7c. per ton and the labor of the seven men, which includes oiling the cars, amounts to $11.83 per shift or 2i/4c. per ton. Maintenance costs for new ties, sheaves, grips, etc., was found to average $25 per month or 0.2c per ton of coal handled; this includes cleaning up wrecks, replacing derailed cars, etc., there being an average of one wreck per month which took about two hours to clean up. The total cost of operating the plane in cents per ton, is: Depreciation on rope, O.TOc. ; labor, 2.25c. ; maintenance, 0.20c. ; total, 3.15c. Endless rope haulage. The expression face haulage, indi- cates a means of disposing of the output from the working face by a haulage system sufficiently flexible to be capable of rapid extensions. A system conforming to these requirements, upon which some valuable cost data are available, and which has been in use for a number of years, consists briefly of the following : The method is applicable to either room-and-pillar or long- wall mining. At the special mine under consideration, the output was 600 tons per day and endless rope haulage was used the empty cars entering the section at one end and the loads passing out at the other, the empty cars being taken off along the rope and the loaded ones attached. The cars can pass around curves with a relatively small radius and they are automatically detached at any point that may be desired. The haulage is a side rope system, the rope being on the side farthest away from the working places. When first started the miners were required to push their cars to the rope and attach them, but it was found that this led to an unequal distribution of the cars, the men at the beginning of the section taking the most of the cars and finish- ing their day's work first. To obviate this six boys were put on pushing the cars whose duty it was to see that the cars were equally distributed. The following are the particulars of a typical installation : Number of miners in section 98 Number of tons per man 6 Number of men per place 2 Cars per day 860 Tons per day 600 Weight per car 14 to 13 cwt. Tare of car 4| cwt. Size of car 4X2X3 ft. Rails handled in 12-ft. lengths weight 24 Ib. per yd. HAULAGE COSTS 309 Gauge of road 24 in. Rope, plow steel.. . . weight f Ib. per ft. and 2 in. in circumference Height of coal 72 in. Nature of roof Good, sandstone Nature of section, flat, little water, good roof and floor, coal easily mined, roof weight helping considerably. Grade, practically horizontal, about one-half of 1 per cent. Labor in Section per Day Deputy, at $3, for one-third time $1 . 00 Haulage man, at $2.50 2.50 Six drawers, at $1 . 75 10 . 50 Two roadmen, at $2, for one-half time 2 . 00 One boy, at $1 .50 1 . 50 One night roadman, at $2 2 . 00 Shifting and haulage cost, about $210 a month, per day. ... 3 . 50 Total $23.00 Of these only part time of the roadmen and very little (about one-third) of the deputy's time are charged against the haulage, which gives a cost on the tonnage named of 3.83c. per ton. The distance traveled is 2.08 miles, which works out at a rate of 1.84c. per ton-mile, which for a face haulage com- pares favorably with the larger and more permanent of end- less-rope haulages. Cost of wire rope. List prices of crucible cast steel rope of either standard or lang lay were quoted in 1920 as follows : Price per Foot Diameter in Inches Approximate Weight per Foot Approximate Strength in Tons Working Load in Tons Diameter of Drum in Feet $0.60 li 3 55 63 12.6 11 0.51 H 3 53 10.6 10 0.43 U 2.45 46 9.2 9 0.36 li 2 37 7.4 8 0.29 1 1.58 31 6.2 7 .22^ i 1.20 24 4.8 6 .17 4~ .89 18.6 3.7 5 .141 H .75 15.4 3.1 4f .12 1 .62 13 2.6 41 .10 T6 .50 10 2 4 .08 i .39 7.7 1.5 31 .06^ A .30 5.5 1.1 3 .05^ ! .22 4.6 .92 2f .04? T6 .15 3.5 .70 2* .04 A .125 2.5 .50 If 310 COAL MINING COSTS List prices of plow steel scale lay rope were quoted as follows : Price per Foot Diameter in Inches Approximate Weight per Foot Approximate Strength in Tons Working Load in Tons Diameter of Drum in Feet $1.30 If 4.85 112 22 7 1.08 If 4.15 94 19 6.5 .93 i| 3.55 82 16 6 .79 H 3 72 14 5.5 .65 ij 2.45 58 12 5 .54 if 2 47 9.4 4.5 .43 l 1.58 38 7.6 4 .34 1 1.20 29 5.8 3.5 .26 I .89 23 4.6 3 .19 I .62 15.5 3.1 2.5 .16 A .50 12.3 2.4 2.25 .14 1 .39 10 2 2 .13 A .30 8 1.6 1.75 .121 1 .22 5.75 1.15 1.50 12| A .15 3.8 .76 1.25 .12 1 .10 2.65 .53 1 Wire rope lubrication. Wire rope deteriorates with use, but not with, age when properly cared for; but the rate of deterioration depends in large measure upon the character of the metal used, the construction of the rope, the diameter of the drums, sheaves and pulleys over which it operates, and to a still greater degree upon how it is lubricated. A rope may be made with great accuracy and meet every specification that human ingenuity can devise, but if not properly protected from the elements which may attack its constituent parts, value and the desired economy cannot be secured. Consequently, the protection and lubrication of wire ropes are of much impor- tance. This applies to cables lying idle as well as to those in service, since rust destroys as effectively as hard work. The question of efficient lubrication has recently assumed a position of considerable importance. The manufacturer may use the best technical knowledge at his disposal and be most careful in selecting the material which goes into his product, but he cannot be expected to estimate beforehand the rapid HAULAGE COSTS 311 and varying degrees of deterioration of the rope when it is in the hands of the user, who it often happens does not fully appreciate that conservation of the life of the cable is entirely under his direction. Even the most perfect rope can be used under such severe conditions and with such lack of attention that it will have a short life, while one of much inferior quality, used under the same conditions of service but carefully taken care of in the way of lubrication, will outlive the higher grade rope. A careful investigation of all steel cables used ir mine service has shown that no two manufacturers are using the same grade of lubricant, and an analysis of the materials most commonly employed and recommended for this class of work shows that tar, graphite and other fillers are used in large proportions, this no doubt for the purpose of developing heavy, adhesive mixtures, which are commonly sold under the names of ''rope shield," "rope dressing" and "protectors." Such materials have the effect of only partially protecting the external parts of the rope, and at low temperatures will crack and peel. This may be noticeable only in spots, but it has the effect of permitting the deteriorating elements to attack the internal portions of the cable, and instances are not infre- quent of ropes suddenly breaking while the visible wires show no signs of deterioration. It is invariably found upon examina- tion in such cases that the internal wires have perished by corrosion. In some instances ordinary black oil, or what is commonly known as "waste oil," is used. Both of these materials are worthless as a rope lubricant, as neither will cling to the outer surfaces, penetrate to the core, nor resist the effects of moisture or other damaging elements. Where such materials are used, frequent applications are necessary ; and owing to their charac- teristics, they afford no lubrication to sheaves, drums and pulleys, and are either thrown off or evaporated within a few hours after being applied, thus leaving the rope at the mercy of the elements and of frictional wear. Many high-grade and expensive greases have been used which perhaps by laboratory analysis are shown to possess the qualifications necessary for resisting the effects of moisture and chemicals, but which on account of the high speed at which 312 COAL MINING COSTS ropes are often run and the consequent stress and vibration to which they are subjected are readily thrown off and at no period after application afford more than a temporary pro- tection to the external surfaces of the rope. It rarely occurs nowadays that the full efficiency of a rope is developed, owing to service conditions requiring it to come into contact with water containing deteriorating elements which have the effect of quickly producing corrosion and con- sequent brittleness of the wires. Furthermore, the higher the carbon content of the metal the more susceptible are the wires to this corrosion. A series of practical tests conducted by a number of high- grade technical men has demonstrated that the use of a pooi or unsuitable lubricant will often do more to lessen the dura- bility of a rope than using no lubricant at all and that the cost, of a lubricant that will meet all operating conditions is trifling as compared with the saving which its use makes possible. An efficient wire rope lubricant must be free from any material that will attack the constituent parts of the rope, must remain soft and pliable under all atmospheric conditions and must not be subject to evaporation. It must be insoluble in water, so as to prevent water from coming into direct con- tact with the surfaces of the rope, and must be unaffected by water heavily charged with the chemicals that are encountered in mine operation. It must be of such a nature that it will penetrate between the wires in the strands and between the strands to the core of the rope preserving the latter as well as the metal which surrounds it. It must not be subject to decomposition under the severest conditions of wear and must possess great adhesiveness so that it cannot be thrown off by any force. It must be a material that will not harden or peeJ either through too frequent application as is often the case or under low temperatures, and must be of such a nature as to permit of it being liquefied to a consistency to permit of appli- cation being made while quite thin. In addition to meeting these requirements, the lubricant used for wire ropes should possess characteristics which permit of its showing equal efficiency as a lubricant for sheave wheels, drums, pulleys or other machine elements over which ropes are liable to pass. And to secure the greatest efficiency this same material should be used to lubricate each wire in the strand HAULAGE COSTS 313 and to saturate the core during the process of manufacture, as it often occurs that with the use of one material in the manufacture and another after the rope has been put in service, the two being of decidedly different nature, no lubrication can be effected on account of one material preventing the other from adhering or penetrating to the interior of the rope. Too little attention has been given to the initial lubrication of wire rope, particularly as regards proper saturation of the hemp core. It has been the general practice to pass the hemp center through the lubricant used at the time the rope is laid up. When applied in this manner, the greater part of the lubricant is forced out between the strands and, in the case of some grades of lubricant, drips entirely away from the external surfaces. The object of inserting the hemp center is to increase the flexibility of the rope, and the deterioration of this element has the same effect as the deterioration of the material surrounding it. If the hemp center is thoroughly lubricated before being inserted into the rope, it will act as a container of the lubricant and will assist in distributing it to all parts of the cable ; also if the proper lubricant is period- ically applied it will penetrate and maintain the hemp center as a continuous lubricator. Another important feature in connection with wire rope lubrication which does not generally receive the proper atten- tion is the method of application. In some instances the lubricant is poured onto the rope either at the sheave wheel or at the drum; in other instances it is applied with a brush. Both of these methods are crude and wasteful. The proper way to apply the lubricant is to use a split box large enough to hold about 25 Ib. of the lubricant to be used. This should be constructed with a hole in the center large enough for the passage of the largest rope in the mine, and when coating smaller cables an old rubber pump valve or a piece of ordinary burlap wrapped around the rope may be employed to act as a wiper and regulate the thickness of the application. These boxes can be constructed so as to be used on either horizontal or vertical ropes. The lubricant should be thoroughly liquefied in a metal container, and after it is poured into the split box the rope should be permitted to pass slowly through it. In this manner a uniform and economical application can always be made. 314 COAL MINING COSTS When possible the external surface of the rope should be dry when the application is made. Comparative costs of different systems of haulage. Many of the opinions expressed in regard to relative economy of the different systems of haulage are founded largely on prejudice, with little or no basis for accurate comparison. The relative costs of haulage by all systems depend upon their intelligent installation and handling, and with equal energy and experi- ence the difference in the cost of haulage by any system will often be only a fraction of a cent. If the animal haul can be kept short and the grades not too steep, the mule or horse in gathering service is a close competitor with the locomotive except where the three- or four-ton mine car enables the loco- motive to get more coal each time it makes a trip to the room. An important factor in securing maximum economy in haulage lies in correctly delimiting the line of secondary haul- age, or as it is more commonly known, the gathering. This will vary somewhat according to local conditions at the dif- ferent mines but a good maximum to set between the working face and the main haulage switch is 1000 ft. Where long distances have to be covered on secondary haulage motors are substantially the most economical. One instance is on record of a 6-ton motor on secondary haulage work handling 60 cars per shift over grades ranging from 5 to 17 per cent and distributing to 14 different entries, none oi which were in, over 500 ft. Where approximate values are required for general pre- liminary estimates, the following figures were those commonly used by the German engineers about 1912 : Working costs for continuous current locomotives with overhead contact lines, 0.9 to 1.2c. per ton-mile; for single-phase locomotives with overhead contact line, 0.9 to 1.2c. per ton-mile ; for accumulator locomotives, 1.8 to 2.1c. per ton-mile. The working costs are here based upon the output in ton- miles; i.e., the costs stated are those incurred in hauling one ton of material over a track one mile in length. The above figures, which can be reduced under favorable conditions, enable any expert acquainted with his own working costs to ascertain by comparison whether he would be able to effect economy in his own installation by the introduction of electric haulage. HAULAGE COSTS 315 Comparison of all systems. One of the most exhaustive comparative cost studies of all types of haulage (except animal) that has come to the attention of the author was that appearing in a German technical journal of a number of years ago, the results of which are given herewith together with a chart, Fig. 21, showing graphically the percentage of power, wages, interest and depreciation, repairs and supplies for each of the different systems. 1 1. Interest & Depreciation f^wer GRAPHIC COSf CHART FIG. 21. Distribution of costs for all kinds of haulage except animal. ELECTRIC LOCOMOTIVE ConditionsTonnage, 2530 per day of 16 hr.; distance 3f mi.; speed, 7| mi. per hr. Material said to be ore. Cost of Plant Four locomotives, including transformer apparatus a.nd trolley wire, $54,748. Working Expenses $25,911, distributed as follows: Approximate Per Cent of Total Interest and depreciation, 10 per cent $5474 21 Upkeep of locomotives 2623 10 Upkeep of trolley wires, etc 1810 7 Wages 5975 23 Power 8399 33 Oil and Waste 1630 6 The operation cost worked out at %c. per ton-mile. 316 COAL MINING COSTS STORAGE BATTERY Conditions Tonnage, 2400 per day of 16 hr.; distance, 1200 yd.; speed, 6| mi. per hr. Cost of Plant Cost of five locomotives (20-hp.), weighing 6| tons, also of transformers, switchboard and reserve $14,600 Accumulators 10,219 Cable, rooms for transformers and charging 7,542 Total $32,361 Working Expenses Annual amount, $14,133. Approximate Per Cent of Total Depreciation. $1460 10 Parts for batteries 1469 lOf Interest on total first cost 1095 8 Wages of drivers 3581 25| Brakemen's wages 992 7 Attendance on transformers and batteries.. . 2671 19 Upkeep and cleaning 569 4 Oil and waste 141 1 Power, 186,349 kw.-hr 1790 12| Acid, distilled water, etc 355 2| The operation costs were 3c. per ton-mile, but it is con- sidered that the conditions were not favorable. BENZOL (GASOLINE) LOCOMOTIVE PLANT Conditions Tonnage, 1250 per day of 16 hr.; distance, 1000 yd., sloping toward shaft at 4 per cent grade; speed of 3| to 5| mi. per hr. Cost of plant $8930 Four locomotives, 8-hp $6813 Filling apparatus 73 Engine shed for six machines 2044 Working expenses for six months, $587. Approximate Per Cent of Total Interest and depreciation $87 15 Upkeep of locomotives 180 31 Wages of engine drivers .- 136 23 Wages of brakemen, etc 68 11 Benzol 97 17 Oil and waste 19 3 This operation works out at 3c. per ton-mile. HAULAGE COSTS 317 COMPRESSED-AlR HAULAGE Conditions Tonnage, 1242 tons in 16 hr.; distance 1| nil. Cost of Plant Four 12-hp. locomotives weighing 5| tons, hauling 40 to 50 cars of 17 cwt. gross, compressor, plant and accessories, $14,500. Working Expenses Annual amount, $9391. Approximate Per Cent of Total Interest and depreciation at 10 per cent $1460 16 Repairs, upkeep, oil and waste 608 6 Wages, brakemen and switchmen 851 9 Wages of drivers (5) 1971 21 Cost of attendance on plant 365 4 Consumption of power (170 hp., 16 hr. per day, 300 days) 4136 44 This works out at 2^c. per ton-mile. ROPE HAULAGE Conditions Tonnage, 1690 in 18 hr.; distance, 1400 yd., main road served by a number of branches. Gradient of 5 per cent for 700 yd., including right-angled turn. Cost of Plant $5092. Total cost of installing $2929 Rope 973 Share of outlay on haulage engine and buildings 2190 Working Expenses Annual amount, $7526. Approximate Per Cent of Total Depreciation and interest $496 6 Wear and tear on rope 486 6| Wages of hookers-on 3504 46 Two engine drivers 734 10 Roadman 365 5 Upkeep, repairs, oil, cleaning, waste 116 2 Cost of power 1825 24 This cost works out at l%c. per ton-mile. 318 COAL MINING COSTS SECOND ROPE HAULAGE Conditions Tonnage, 2100 in 10 hr.; distance, 3 mi. on a level track. Cost of Plant Including engine, 250-hp., and accessories, $48,666. Working Expenses Annual amount, $15,069. Approximate Per Cent of Total Interest and depreciation at 10 per cent. . . . $4866 32 Wear and tear of rope 851 5 Upkeep, repairs, cleaning, oil waste 618 4 New parts for rope haulage and rope clips .. 837 5^ Wages of hookers-on 4088 27 Engine driver 296 2 Cost of power 3513 23 This works out to %c. per ton-mile. OVERHEAD CHAIN Conditions Tonnage, 1750 per day of 16 hr.; level track 1 mi. long. Cost of Plant Machinery, including 65-hp. engine, $6082; chain, $2929; total, $9011. Working Expenses Annual amount, $13,828. Approximate Per Cent of Total Interest and depreciation $876 6 Interest and depreciation of chain 525 4 Wages of hookers-on 7878 57 Repairs and upkeep of plant Ill 1 Oil and waste 141 1 Cost of power 1786 13 Upkeep of haulage road 2501 18 This is almost exactly 2c. per ton-mile. Animal, compressed-air and electric haulage costs. A most interesting and valuable comparison of the cost of animal, com- pressed air and electric haulage at a mine in Western Pennsyl- vania, producing 35,000 tons a month was worked up in 1912. The grades in this mine are variable, some of them in favor of, and some of them against the loads, and are in places as steep as 6 per cent. The capacity of the mine cars is two tons, and their empty weight 2700 Ib. In the case of the elec- trically operated mine, the car capacity was not definitely stated, but presumably it was about 3000 Ib., and the weight of the car itself about 1300 Ib. A comparison of figures for the two cases is given in Tables I and II. HAULAGE COSTS 319 w ! 4 I 8 if ,s o 5 11 1 a ji ss? 1 1 O IO CO OO l> co o I-H o o o dodo P 8 E 6 ^w *C 3 l> (N O O O* IO iO O CO -00 o dodo d TH d 02 (H 3] CO CO cTcT CO CO CO . CO . CO o" o" o" o" o" CO 1 cT r < rH 1 OJ O5 o^ o^ o^ o^ o^ c^ OS 1 w o ^o I Illii! i i K- a d jj feJ2 c^S i-H i 1 S 8 g :S :S 00 00 I s * * O^ * O^ (N rH (M Tj< i < Q _ 1 ^^ 5! 8 i-l rH CO C od d d 10 CO i-( rH C<1 O -CO 6 foo|as (H .fa H -? C^ o3 g o> .2 Fl J3 c^ S w -^ N CO CO 52 <^ IO JO i i ll CO j^ n3 ?^x J^x O O f* p o3 f^ -^ * . H . -H ' g 02 " v i ~ ^ o3 o3 8 "g H ^ < |-2-^o < & HAULAGE COSTS 323. TABLE V COSTS OF MULE HAULAGE Average distance of mule haul 1200 ft. Drivers, 250 hr. at 30c $75 . 00 Drivers, 398 hr. at 18c 71 . 64 Drivers, 294 hr. at 19c 55.86 Cost per Drivers, 310 hr. at 20c 62 . 00 Ton $264.50 $0.0646 Stable boss 35.00 0.0086 Blacksmith, shoeing 30 . 00 . 0074 Feed 79.20 0.0193 Miscellaneous supplies 18 . 52 . 0043 Depreciat on, 5 years 25 . 00 . 0062 Interest, 6% on $1500 7. 50 0. 0018 Total $459.72 $0. 1122 Total cost per ton mile 49 . 4c. The coal gathered by the mules in this mine was hauled to the pit mouth, an average distance of 2500 ft., by an electric locomotive. The comparative figures for main haulage in the two mines under consideration, one using electricity and the other com- pressed air are given in Table VI. The electric locomotive moved the coal an average distance of 2500 ft. and the compressed-air locomotive hauled it for au average distance of 3400 ft., so the costs per ton-mile, are as follows : For electricity: = 1938 ton-miles, 1939 =10.83c. per ton mile. 3400X19,688 For air : = 12,677 ton-miles, 5280 386 31 12677 = 3.05c. per ton mile. In explanation of the correction of the figure of $20 per month for power as given in Table I and increased to $50 in Table IV, it should be noted that the costs of power per ton-mile 324 COAL MINING COSTS 1 o O P 2 > i .fa 00 rH II d d p o d d oT oT o o o 8 d d O5 ' O5 ' O5 o o o o o CO TH (N I> S S388 * ^ * 8 ^ t^ CO t^ Tj< (M CO S-l ' S S ^ ^ iiiil o i! s j n - ,9,33 HAULAGE COSTS 325 for gathering and main haulage as previously considered were as follows : Cost for power in gathering $20.00 Ton mileage as calculated 2758 $20 - 00 = $0 . 00726 per ton-mile. 2758 And for main haulage: con 00 ^P = $0 . 01032 per ton-mile. 1939 It will be seen from these figures that the cost per ton-mile is 42 per cent greater for the main haulage than for the gather- ing service. The costs of power per ton-mile for the same two classes of service at the compressed-air operated mine were : For gathering : ' = $0 . 0284 per ton-mile. 41oo $80 18 For main haulage : : = $0 . 0063 per ton-mile. 12677 These figures show the cost for gathering with compressed- air locomotives to be 4.48 times as great as for main haulage. There are many reasons why the cost for power per ton-mile for gathering should be much more than for main haulage, but none why it should be less, unless under such a condition as where all the main haulage grades are against the loads and all the gathering is down hill. In all the above figures it is the ton of coal moved one mile in the desired direction that is considered. In actual service, the cars and the locomotive must be moved as well as the coal, and because the car and the locomotive must move approximately twice as far as the coal in order to get the empty cars back to the loading place, the gross ton-mileage is nearly twice the net ton mileage for main haulage. For gath- ering work the gross ton-mileage will be from four to six times the net ton-mileage because in gathering cars from the rooms the locomotive must make many movements with only one car or without any cars. While doing this work the per- centage of net to gross tonnage is exceedingly low because the 326 COAL MINING COSTS locomotive itself forms a large part of the total weight of the moving train. Any remaining inconsistency in the relative costs of main haulage and gathering service with the compressed-air loco- motives is readily accounted for by considering the following adverse conditions of gathering as compared with main haul- age: Ordinarily the gathering locomotives has poorer track than the main-haulage locomotive, a greater percentage of curves, more frequent stops and starts, and a somewhat lower efficiency, because of handling only one car at a time instead of being loaded approximately up to its capacity. From the foregoing considerations it seems certain that the figure of $20 as the cost of power for the electric gathering locomotive needs correction; more especially as this locomotive operated two shifts per day, while the main-haulage locomotive operated only one shift. The conclusion that, as it cost only lOc. per ton to gather and haul 4045 tons of coal per month with one locomotive working two shifts, and as it cost 16.34c. per ton to gather 4095 tons of coal with mules and haul it 2500 ft. to the pit mouth, there is, therefore, an advantage of not less than 4c. per ton in gathering by electric locomotives, does not seem to have sufficient support to make it generally true. In the mine using compressed-air haulage, coal was gathered and hauled by mules an average distance of 900 ft. at a cost of 5.2c. per ton and was afterward hauled an average distance of 3400 ft. by main-haulage compressed-air locomotives at a cost of 1.97c. per ton, giving a total cost for gathering and hauling of 7.17c. per ton, a lower figure than that in the electric haulage mine. This figure is also slightly lower than the results achieved with compressed-air gathering and main-haulage loco- motives in sections of the same mine where the gathering locomotives worked but one shift, as is shown by the follow- ing costs: Gathering by compressed-air locomotives 6.72c. a ton Main haulage 1 . 97c. a ton Total cost 8 . 69c. a ton Thus it is seen that the coal was delivered to the shaft bottom more cheaply by mules and main-haulage locomotives HAULAGE COSTS 327 than by gathering locomotives and main-haulage locomotives; but in order to achieve these results with the mules it was necessary to keep the mule haul down to 900 ft. or less in order to enable the animals to gather, as they did, 41.6 two-ton cars per mule. If the mules had to haul the coal an average dis- tance of 1200 ft., as did the gathering locomotives, and had been forced to encounter the adverse grades that the locomo- tives did, the cost for mule haulage would have been 50 per cent greater, throwing the balance again in favor of the gath- ering locomotive. In other words, it was possible to obtain the good results that were obtained with mules in this mine only by working them in selected places under the most favor- able conditions, all of which goes to show that it is extremely dangerous to draw general conclusions in regard to the relative economy of the various types of haulage unless the conditions are strictly comparable, or accurate corrections are made to compensate for differing conditions. Single- and two-stage air motors. The two-stage com- pressed-air motor will do from 40 to 60 per cent more work with the same amount of air than the single-stage machine. The cost of power for this type motor will thus be about 30 per cent less than with the single-stage, the compressor and boiler capacity will be reduced by about the same amount and the first cost of the installation will be about 15 per cent less, while the motors will travel substantially further on one charge. A practical test of the difference between single-expansion and two-stage compressed-air locomotives was once made as follows: The same train was hauled by the single-expansion and two-stage locomotives the same distance over the same piece of track, with the same operator, giving the locomotives as nearly as possible the same work to do under the same con- ditions. In all cases the trains were started from a given point and were allowed to come to rest as near as possible to an- other given point without the use of brakes. Air consumed was determined by the difference in pressure recorded at the beginning and end of the trip by the gauge on the main reservoir. Running conditions were the same as would obtain in the regular hauling of coal in the mine. Trial 1 was conducted at the Susquehanna Coal Co.'s No. 328 COAL MINING COSTS 1 colliery, Nanticoke, Pa., in the presence of Mr. McMahon, Chief Engineer, Susquehanna Coal Co., and Mr. C. B. Hodges, of the H. K. Porter Co. TRIAL 1 LOCOMOTIVE DATA Single- Expansion Two-Stage Weight 10,000 Ib 10,600 Ib Cylinders high-pressure, diameter 6 in 5 in Cylinders, low-pressure, diameter 10 in Cylinders, stroke 10 in 10 in Driving wheels, number and diameter .... Working pressure, high-pressure cylinder . Maximum charging pressure 4-23 in. 150 Ib. 900 Ib. 4r-23 in. 250 Ib. 900 !b. Capacity of main reservoir 41 cu. ft. 41 cu ft Age of locomotive 8 months 1 month TRAIN AND ROAD DATA Length of trial run, feet 2200 Average grade, per cent . 96 Maximum grade, per cent 2 . 00 Track gage, inches 42 Train consisted of loaded coal cars weighing about 9500 Ib. each. The excessively high efficiency indicated on lines Nos. 4 and 6 of the table of log of runs, hauling four and five cars down grade, may be due to the fact that considerably more air is wasted in the single-expansion locomotive when giving the train just a little assistance when the grade is nearly steep enough to cause it to run down by itself. Lines Nos. 7 and 8 show additional trips up grade with the two-stage with five-car trips. The trip with the two-stage (line 5) was made with the reverse lever "in the corner, " using the air full stroke all the way. The superior efficiency achieved on trips shown on lines Nos. 7 and 8 was the result of using the air as expansively as possible. Taking the total pressure reduction of the three round trips with the single-expansion locomotive and two-stage locomotive (lines 1, 2, 3, 4, 5, and 6), we find that the two-stage locomotive HAULAGE COSTS 329 used 1245 2225 = 56 per cent of the air used by the single-expansion engine under exactly the same conditions of service, this average per cent for the entire test showing a saving of 44 per cent of the air used by the single-expansion machine. In making the test every care was used to obtain reliable results. The pressure gauge on the two-stage locomotive was removed and placed on the single-expansion locomotive during the test of this locomotive, and then shifted back to the two- stage for testing it. This gauge was a comparatively new one, and presumably correct, and if any error did exist, it would have been the same for both locomotives. The tanks were exactly of the same capacity. The same engineer operated both locomotives alternately during the trials. Trial 2 was conducted at Orient Mine of the Orient Coke Co., Orient, Fayette County, Pennsylvania, in the presence of Mr. Chas. Opperman, of the Orient Coke Co.; Mr. G. E. Hut- telmaier, of the H. C. Frick Coke Co. ; Mr. C. B. Hodges, of the H. K. Porter Co. TRIAL 2 LOCOMOTIVE DATA Single- Expansion Two-Stage Weight . 9600 Ib 10 500 Ib Cylinders, high-pressure, diameter 6 in 51 in Cylinders, low-pressure, diameter.. 11 in Cylinders, stroke 10 in 10 in Driving wheels, number and diameter Working pressure, high-pressure cylinder. Maximum charging pressure 4r-23in. 150 Ib. 800 Ib 4r-23in. 250 Ib. 800 Ib Storage tanks 1 1 Capacity of main reservoir 40 26 cu ft 40 26 cu ft Age of locomotive 6 months 2 months TRAIN AND ROAD DATA Length of trial run, feet 2500 Average grade, per cent 52 Track gage, inches 44 In this run there was a reverse curve in a chute leading from one heading to a parallel heading. The train consisted of four loaded wagons, each about 7000 Ib.; and six empty wagons, each about 2200 Ib. 330 COAL MINING COSTS LOG OP TRIAL RUNS TANK PRESSURES TIME (P.M.) No. of Run Type of Locomotive At start At finish Amount of Start Finish Elapsed Drop 1 Single-expansion . 705 265 440 7:23 7:27^ .04 ?, Two-stage 740 420 320 8:06* 8'12 05 i 3 Two-stage 685 385 300 8:48 8:52f .04| No. 1 run: Very satisfactory. No. 2 run: Very irregular; operator not so familiar with two-stage machine, hence decided to rerun. No, 3 run; Much better and smoother than second. DEDUCTIONS Calculated by Mr. G. E. Huttelmaier, H. C. Frick Coke Co. Trial No. 1 Trial No. 2 Free air consumed, cubic feet 1,206.92 930.16 2,325,400.00 581,350.00 625.00 7.10 17.61 1,926.00 301 . 70 17.13 100 per cent 100 per cent 877.76 :} 2,348,350.00 426,973.00 454.50 5.17 12.94 2,676 . 00 159.59 12.33 72 per cent 1.38 per cent 28 per cent Drawbar effort \ . \ . . I Trip 832. 24 / Total work performed in foot-pounds Foot-pounds performed per minute . . . . , / Feet per minute "1 Average speed \ , ,., > > . . 1 Miles per hour / Average horsepower developed . ... Foot-pounds work per cubic foot free ah* Free air consumed per minute Air per minute per horsepower Percentage of air consumed as compared with Trial No 1 Amount of work per unit of air compared to Trial No 1 Saving of air effected over Trial No. 1 .... HAULAGE COSTS 331 LOG OP RUNS WITH PRESSURE REDUCTIONS AND PERCENTAGE USED BY TWO-STAGE LOCOMOTIVE Line 1 2 3 4 5 6 7 8 SlNGLE-E X PANSION TWO-STAGE Percentage Used by Two-Stage No. Cars Gage Pressure Pres- sure Re- duction Time No. Cars Gage ? Pressure Pres- sure Re- duction Time i Start Stop Start Stop 3 3 4 4 5 5 1 i 690 280 770 760 805 730 \>tal p 2225 It . 280 140 180 540 200 470 ressvm >. 3 r 410 140 590 220 605 260 3:15 3:30 2:55 an 1 3 3 4 4 5 5 1 5 5 680 400 570 250 710 290 ^otal p 1245 It 610 735 400 320 250 175 290 220 ressurt . 3 r< 220 360 280 80 320 75 420 70 68.3 57.1 54.2 34.1 69.5 26.9 Up grade Down grade Up grade Down grade Up grade Down grade Up grade Up grade 2:40 2:50 n 1 2252 ; reducti sund tri] 1245 reductic >und trip 390 375 3:15 Compressed-air and animal haulage costs. The Consolida- tion Coal Co., in 1902, conducted an investigation into the relative cost of mule and compressed-air haulage, the results of which were described in the transactions of the A.I.M.M.E., Vol. 34, p. 144. Five of these machines were placed in the company's Ocean No. 3 mine (Hoffman), displacing a number of mules, but leav- ing 19 still working. This opportunity was embraced to make a close comparison between the two methods of gathering. The mules working in the North Heading and the South Heading deliver their cars directly to the rope on the slope The other mule routes deliver to the heavy motors, as do all the motor routes. The mules used weigh from 1200 to 1400 lb., and pull an average of 2.4 long tons. The following table shows the work performed by the mules during a period of 18% working days in the month of Decem- ber, 1902: 332 COAL MINING COSTS Route Cars Moved Average Haul, Feet Constant Tons Moved 1000 Feet South heading 1119 2900 2.4 7788 24 North heading 268 1300 1000 do 836 16 First cross . . 1629 2100 do 8210 16 Second cross 3042 1100 do 8030 88 Third cross 747 400 717 12 This represents a total of 339 days' work for one mule. The company's accounts show a cost of $1.15 per day for each day worked by a mule, including expense of replacing worn- out animals. Drivers are paid $1.98, and there is one with each mule. This makes a cost of $3.13 per day for each day worked by a mule. The cost per ton hauled 1000 ft. would . therefore, be: 339XS3.13 =4.15c. 25,582.56 For the work of the motors during the same time we have . Route Cars Moved Average Haul, Feet Constant Tons Moved 1000 Feet. Tippens 1122 2300 2.4 6 193 44 Scobies 1073 2050 1000 do 5 279 16 First Klondyke 1147 1835 do 5 046 80 Second Klondyke. 1032 1800 do 4458 24 Third Klondyke 1147 1865 do 5 138 56 Fourth Klondyke 114 1992 do 544 92 Total 26 661 12 This work was done by the five small motors operated by compressed air, working a total of 94 days. This plant is supplied with steam by a battery of boilers, which also supplies steam to the large pumps. The plant consists of the following items, with their approximate first cost: HAULAGE COSTS 333 One straight-line Norwalk air-compressor, 18 and 28 compound steam, 18, 13J, and 6 three- stage air, 30-in. stroke $5,300 5600 ft. of 5-in. pipe 5,600 3100 ft. of 2Hn. pipe 1,700 1000 ft. of U-in. pipe 300 2 motors, 30,000 Ib. each 6,000 5 motors, 8,000 Ib. each 10,000 Estimated proportion of boilers 1,000 Installation 4,000 $33,900 Allowing $3000 per year for interest and depreciation, to be earned in 300 working-days, would justify a charge of $10 per day from this source against the entire plant. This same compressor also drives the large motors men- tioned above, which weigh 30,000 Ib. each (60,000 Ib. for the two) ; the five small machines weigh 8000 Ib. each (40,000 Ib. for the five). Dividing the general expenses according to the weight would result in four-tenths being charged against the small motors. These general expenses may be summed up as follows per day: Coal, 4 tons at $1 $4.00 Fireman 2 . 00 Mechanic in charge of compressor 2 . 50 Interest and depreciation 10 . 00 $18.50 The cost of operation of the five small motors would then be: 5 motormen @ $2 . 67 $13 . 35 5 brakemen @ $2 . 03 10. 15 General expenses 0.4X$18.50 7.40 Repairs and oil 3 . 00 $33.90 Dividing this among the five machines would give $6.78 per day for each machine and the cost per ton moved 1000 ft. would , 6.78X94 334 COAL MINING COSTS In the matter of community of service the motors show to great advantage. A broken down motor can usually be re- paired over night, while an injured mule can only be replaced by a new one that must usually be broken in and inured to the work before he is thoroughly efficient, entailing loss of time and output in each case. Electric motor and animal haulage costs. An interesting and valuable comparison of the cost of electric and mule haul- age at one of the mines of the Peabody Coal Co. was worked up in 1907. The introduction of the electric haulage not only resulted in reducing the cost of production, but also made prac- ticable the development of more extended operations and increased the output from 1400 tons to a daily average of 2000 tons. The motors have pulled as much as 2570 tons in 8 hr. Prior to installing electric haulage there were 16 gathering mules and 17 mules working in spike teams, pulling to the bottom, producing 1400 tons of coal per day. Owing to the size of the cars, grade and average haul of 1800 ft. to the bot- tom of the shaft, the output had reached its limit with mule haulage, and it was decided to install electrical haulage. Two 15-ton traction locomotives with double-end control were in- stalled. The locomotives have pulled 17 loaded cars up a 2 l / 2 per cent grade 1200 ft. long. These cars weigh when empty 1950 Ib. and hold on an average 6600 Ib. coal, so the weight of the loaded trip would be over 72 tons. The power for operating the motors in the mine is supplied by a 175-kw. generator belted to a 200-hp. high-speed engine, 18 X 18 in. The generator also furnishes light for the under- ground haulage-ways. Steam to run the electric plant is furnished by a battery of four 150-hp. tubular boilers, which also furnishes steam for the large hoisting engines; but in order to make the proper comparisons between mule and electric haulage, the cost of two complete power units has been added to the electrical equipment. The machinery making up the electrical installa- tion is as follows: HAULAGE COSTS 335 COST OF ELECTRIC INSTALLATION Two 15-ton locomotives @ $2300 $4,600 . 00 One 175-kw. generator and switch-board 2,400.00 One McEwen engine, 18 X 18 in., 200 hp 2,000 . 00 Foundations and placing engine and generator .... 300 . 00 Two 72 in. X 18 ft. tub. boilers, 150 hp ., complete . . 2,800 . 00 9000 ft. trolley wire 1,019 .90 200 ft. 400,000-cm. lead cable @ 55c 110.00 665 trolley hangers @ 65c 432.25 768 bonds @ 35c 268.80 75 crossbonds @ 35c 26 . 25 18 interchangeable trolley frogs @ $2 . 75 49 . 50 1 extra 250-volt armature 375 . 00 2 motor jacks @ $12.80 25.60 Extra fittings for motors 86 . 24 116 tons 40-lb. rail @ $28 . 25, $3,291 . 13. Cr. for 25-lb. rails, $2,056.75 1,234.38 6055 white-oak ties @ lOc 605.50 65 kegs 4XHn. spikes @ $3.75 244.50 22 split switches, material and labor @ $17.00. . 374.00 Fish plates and bolts 280.00 Lumber for trolley supports 76 . 11 Sundries 3,810.21 Entire labor cost 3,810.21 Total of complete installation $21,172.79 COST OF MULE HAULAGE Mules, average cost $225 . 00 Mules, depreciation 20 per cent Mules, interest. . < 6 per cent COST PER MULE, DIVIDED FOR 275 WORK DAYS, PER WORK DAY Depreciation $0. 163 Interest . 049 Feed 0.20 Shoeing and stableman .'. . 158 Total per day, 275-day basis $0 . 57 336 COAL MINING COSTS TEAM HAULAGE, No. 3 MINE DAILY AVERAGE TONNAGE, 1400 TONS 17 mules $0.57 : $9.69 9 drivers 2.56 24.24 Total $33.93 Team drivers, 15c. extra, or $1.20 extra for 8. Cost per ton outside mule haulage, 2.4c. COST OF OPERATING ELECTRICALLY, 275 WORK DAYS 2 locomotive runners @ $3 . 20 $6 . 40 per day 2 trip riders @ $2.56 5. 12 per day electrician @ $75 per month 1 . 08 per day fireman @ $2 . 02 0. 67 per day Fuel, 5 tons @ 75c 3 . 75 per day Total fixed labor, etc $17 . 02 per day Interest on investment, $21,172 . 79 @ 6 per cent $4 . 62 Depreciation and repairs @ 8 per cent 6.16 Oil and waste : . 30 Taxes.. 0.50 Total others $11 . 58 Total daily operating cost electrically $28 . 60 Cost per ton based on 2000 tons 1.4c It will be seen from the foregoing that the plant, besides increasing the output and saving Ic. per ton, which means $20 per day, practically pays for itself in four years. The following figures show a comparison of cost previous to 1910, between mules and an electric locomotive at a mine where 14 mules were replaced by one locomotive. The output of the mine averages 1500 tons per day for 245 working days per year. The cars weigh 2400 Ib. empty and hold 3600 lb., making a total weight of 6000 lb. MULE HAULAGE 14 mules $180 each $2,520.00 14 sets harness @ $25 each 350.00 $2,870.00 HAULAGE COSTS 337 INTEREST AND DEPRECIATION 20 per cent depreciation on $2870. 6 per cent interest on $2870 $574.00 172.20 $746.20 WORKING EXPENSES FOR 245 DAYS 14 mules feeding, shoeing, repairing harness and care @ 50c. per mule, per day $1,715.00 6 drivers @ $2.80 per day 4,116.00 $5,831.00 $5,831.00 746.20 $6,577.20 Fifteen hundred tons per day for 245 days equals 367,500 tons per year at a haulage cost of 1.8c. per ton. ELECTRIC HAULAGE Engine, locomotive, boiler and generator $9,000 . 00 Switches, insulators and wire 1,200.00 Cost of erecting, etc 1,000 . 00 $11,200.00 INTEREST, DEPRECIATION, ETC. Interest at 6 per cent on $11,200 $672.00 Depreciation on boiler, engine, etc. @ 9 per cent. . 810.00 Repairs on boiler, engine, etc., @ 9 per cent 810 . 00 Depreciation on switches, wire, etc., @ 5 per cent. 110.00 Repairs on switches, wire, etc., @ 5 per cent 110. 00 $2,512.00 Engineer power house @ $75 per month $900 . 00 Motorman @ $2.80 per day 686.00 Oil and waste 100 . 00 Nipper on motor @ $1 . 50 per day 367 . 50 Sand 50.00 $2,103.50 From above $2,512 . 00 Total $4,615.50 338 COAL MINING COSTS Fifteen hundred tons per day for 245 days equals 365,700 tons per year. This makes the haulage on each ton of coal, where electric locomotives are used, cost 1.27c. per ton. These estimates, taken from an actual case, show a considerable dif- ference in favor of electric haulage. The cost of installing mechanical haulage is greater than when a mine is supplied with mules; however, when we consider the cost of erecting a stable and the great loss due to mules killed in accidents, the initial expenditure is not so favorable to the use of mules. Cost and care of mules. Frank Amos, of the Fairmont Coal Co., in 1911, made the statement "that the average life of a mine mule was 3y 2 yr., and unless conditions were changed to prolong life, the use at the present cost of the animal was unprofitable. ' ' When an animal of this kind lives indefinitely on the farm it seems incredible that his life should be shortened to 3y 2 yr. in the mine. There are mules to-day in anthracite mines that have been working 20 yr., yet it is probable that the aver- age life of such animals in all anthracite mines does not exceed 5 yr. In purchasing stock for underground haulage, activity, eye- sight, feet, temperament, strength, and wind are considera- tions, but if the animal lacks intelligence he has no place in the mine. In most instances it is better to deal with those who make a business of furnishing mules to mining companies, and if it is possible, to go to the stock yards and pick out the animals rather than trust to the dealer's judgment. This sug- gestion is made because the dealer scarcely knows more about the animals than the purchaser, and does not know the con- ditions under which the animals must work. Many animals which act and look right on the surface are most unsatisfactory underground, for which reason the sug- gestion is made that after the animals are picked out an agree- ment should be made with the dealer that in case any of them do not act rightly in the mine they may be exchanged. While a mule's heels are not to be trusted, nevertheless the humane driver and his mule become good chums. If, after a 3-day training period the mule does not appear active on his feet and to use judgment, he should be taken from the HAULAGE COSTS 339 mine, as he is unsuited to the work and cannot be depended upon to look out for himself when occasion demands. According to two authorities the animals should be fed as follows: First, hay, next water, and then grain. Animals should eat hay at least half an hour before being given grain. If the water is given last it washes the food into the intestines before it is acted upon by the gastric juices. If the hay is given after the grain it carries the grain with it, for the hay is prin- cipally digested in the intestines, while the grain is acted upon by the stomach for the most part. Corn is richer in fat than oats ; therefore, for strength, feed corn, and for speed, feed oats. For an illustration, race horses are fed oats, and the experienced teamster will favor the feed- ing of corn. Dr. I. C. Newhard, Chief Veterinarian of the Philadelphia & Beading Coal and Iron Co., in 1911, experimented with various feeds and found that two-thirds crushed oats and one- third cracked corn the most reliable. "A handful of coarse ground pure salt should be fed to each mule twice a week.' : Dr. Frank Amos, who is in West Virginia, suggests a coarse- crushed feed, about two-thirds corn and one-third oats. Mine stock will consume about 12 Ib. per head per day of this feed and about 15 Ib. of hay. If a horse or a mule has not cleaned up its former feed the troughs should be cleaned and less put in the next time, until it is ascertained just how much it takes to keep them. The animal should have about all it will eat, but it is better to give not quite enough than too much. Too much grain will cause acute indigestion, paralyze the walls of the stomach, and usually results in death. A stable boss will make a great mistake by feeding too much and allowing food to stand before the animals all the time. While this method will increase flesh for a short period the animals eventually break down through their digestive organs being destroyed. Grain should not be placed in the animals* troughs ready for them when they come in from work, and it is better for them to be without grain at noontime than to be without water, but by all means give them three feeds a day. Plenty of water will keep the digestive organs in good condition, while large quantities of grain and no water will destroy them. 340 COAL MINING COSTS To give a feed of good bran once a week will aid the con- ditioning of stock, keep the bowels open, and reduce fever, which is caused by strong grain. The Fairmont Coal Co. 's records for 1905 show that 26 per cent of their stock either died, was killed, or had to be dis- posed of at practically nothing, on account of being crippled and worn out. There is probably no part of the company's business in which the loss is so great, and one-half of this is brought about by carelessness and neglect. There is probably no other business that requires the use of stock in which the loss is so great, and this in face of the fact that the facilities furnished, with the exception possibly of good roads inside the mines, are the best that can be had. The maintaining of live stock is no little item, and in cases there is an average of 5 per cent of the total number of animals standing in the barns all the time unfit for service on account of having been crippled. The feed for this stock, besides other expenses and the loss of their work, cost one company in 1911, over $6000 per year. Where mine stock is given good attention, the upkeep is reduced to a minimum, more work is obtained, and the animals are more valuable. Professor Ihlseng estimates the capacity for work of the mine mule on a level track as between 30 and 80 gross ton- miles in 8 hr. and the work of the average mule for the same time as between 40 and 50 gross ton-miles (5 to 6 ton-miles per hour) with a limiting grade of 3 per cent. Hughes gives examples varying all the way from 3 to 16 ton-miles per hour on level track. Other conditions being equal the condition that really determines the work that a mine mule can do is the length of the haul, since a large part of the time is con- sumed in changing and waiting for trips. The average of a number of actual working conditions shows that 4 to 6 ton- miles per hour is a good average for the work of an ordinary mine mule, in a flat seam, under usual mine conditions, though this will increase slightly with the length of haul. With a normal load the limiting grade under which a mule can bo worked to an advantage is about 3 or 4 per cent. As loaded mine cars will slide on iron rails, with 4 sprags, on grades HAULAGE COSTS 341 varying from 6 to 8 per cent the safe down grade for a mule to work on is limited to this. Figures gathered from a number of years experience in a level seam 5 l / 2 ft. thick show that a good 2-mule team, with an efficient driver and helper, will handle from 50 to 60 cars of 2y 2 tons capacity in a 10-hr, shift when the haul does not exceed 2000 ft. Allowing two hours lost time and assuming the weight of the empty car to be 1^ tons each mule has accomplished 7 ton-miles of work per hour. An additional expense that must be allowed for in animal haulage in low seams is the cost of brushing to obtain suffi- cient headroom. Haulage expenses have been increased as much as 3c. per ton (about 1912) from this cause and the decreased efficiency of the animals. I I I I I I I I I I I I I -A LOCOMOTIVE,! EN6INEER.2SRAKEMEN $2442 PER DAY - ///, " ! 7 " 1BPAKEMAN 20.07 \-\-JDRIVER, 2 RUNNERS, 3 MULES 17.60 \\1 I 1 RUNNER, 3 13.24 ' v * ! / Z 11.74 _ \..-l DRIVER, S 7.37 n JMULE 5.67 . Q 0.10 0.10 0.30 0.40 050 060 0.70 0.60 090 1.00 1.10 1ZO 1.30 1.40 150 1.60 1.70 IflO Cost of Car (^c.v C6Jr). FIG. 22. Chart showing relative economy of mule and motor haulage. Additional ventilation must also be provided where animal haulage is used, most state mining laws providing that 500 cu. ft. per min. be allowed for each animal in the mine. In a mine using 50 mules this would mean that an additional 25,000 cu. ft. of air per minute be provided. The working life of the average mule is seven to ten years. Because of the hard working conditions and generally fre- quent accidents it is necessary to keep a good reserve supply of animals available which increases the investment, as well as costs of maintenance. The chart shown in Fig. 22, will give a quick approximate solution of the relative economy of mule and motor haulage where the conditions are known. Other factors, such as availa- bility of power and the possibility of securing locomotives and 342 COAL MINING COSTS equipment are matters for consideration after or before the relative costs of the different organizations are determined. The curves are made up from figures similar to those appear- ing in Table I: TABLE I COST PER CAR FOR ONE DRIVER AND ONE MULE Cars Cost Cars Cost Cars Cost Cars Cost 4 $1.47 8 $0.73 16 $0.37 30 $0.195 5 1.17 12 .49 18 .33 40 .147 6 .98 14 .42 21 .28 60 .098 These when plotted give the basis for the regular curve for an organization or transportation unit of one driver and one mule. The other curves shown are plotted in a like manner. In the case of locomotive haulage, the power, installation and repair costs given or estimated for any single locality will not be cor- rect for every installation. This is one weak point in the tabu- lation. However, because unforeseen conditions usually develop in mine electric installations, it is best to be on the safe side and figure locomotive maintenance high. A close inspection of these curves will also show that tho handling of four or five more cars makes much more difference in the cost per car than does a few cents in the figures repre- senting power, repairs, etc. For instance, in the case of a crab locomotive with one engineer and two brakemen working in pitching rooms, let us estimate that this machine will handle 40 cars per day. This, from Fig. 22, would cost 60c. per car. If another place can be added so that the machine will handle 44 cars, it will do so at a cost of 56c. per car a saving of four cents per car, or $1.76 per day. We cannot always estimate mine haulage possibilities closer than four cars, but our cost figures certainly will not be in error to the amount of $1.76. Two representative problems are given below to illustrate the use of the curves : Problem 1 Assumption by colliery superintendent that his transportation gathering cost shall not be over 60c. per car. HAULAGE COSTS 343 How many cars must each unit handle in order to give him this cost? Follow the dash line on Fig. 22: It crosses one driver and one mule at 10 cars; one driver and two mules at 12 cars; one driver, one runner and two mules at 19 cars; one driver, one runner and three mules at 22 cars ; one driver, two runners and three mules at 29 cars; one locomotive, one engineer and one brakeman at 34 cars ; one locomotive, one engineer and two brakemen at 40 cars. Locomotive Data: Per Day Power, $90 per month $3.60 Installation, $2000, 5 per cent of $2000 1 $2200 Depreciation Locomotive, 15% of $8000 Repairs Locomotive lines, etc., $900 per year J 7.33 Labor Engineer at $0.28, plus $0.25 allowance for 9 hrs. 4 . 77 Brakeman at $0.2351, plus $0.25 allowance for 9 hours 4.36 Mule Data: Mule value, $250; 20% depreciation; 5% interest 0.23 Feed and care 1 .25 Labor Driver at $0.2351, plus $0.25 allowance. Runner at $0.2351, plus $0.25 allowance. One driver, two runners and three mules in a certain sec- tion where there are pitching rooms will handle 25 cars per day. This section due to its development can produce 50 cars per day with two-mule organizations. Which relatively is the cheaper plan, to use two-mule organizations or one crab loco- motive to handle 50 cars? By following the two dotted lines in Fig. 22 it will be seen that one driver, two runners and three mules will handle 25 cars at a cost of 70c. per car ; one locomotive, one engineer and two brakemen will handle 50 cars at a cost of 48c. per car. In this case, of course, the locomotive is the better installation. Close inspection of the curves in Fig. 22 will also show that there is a certain fairly definite limit to the number of cars handled, below which the cost per car increases rapidly. This limit can be taken from the curves by noting the point at which the curve from right to left deviates from practically a straight line. Thus, for instance, in the case of the one driver and one mule curve, the line is practically straight from the right hand 344 COAL MINING COSTS side of the figure to the 10 or 12 car point, and from here turns rapidly into a curve. This possibly is better illustrated by Table II. Taking the limit for the differences or decrease in cost per car at 4c., we see that 12 cars must be handled by one driver and one mule, 18 cars by one driver, one runner and three mules, etc., before this limit is reached. Taking 2c. as the limit 16 cars must be handled by one driver and one mule, and a corresponding number by other units before reaching this limiting figure. TABLE II COST PER CAR FOR DIFFERENT UNITS FOR INCREASING NUMBER OF CARS . 1 . .2 fc I *1 c 2 1 'II I g! 1 Ill V QtfS 0> Ill HI I III | Q 1 2 3 4 6 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 5.87 2.93 1.96 1.47 1.17 .98 .84 .74 .65 .59 .53 .49 .45 .42 .39 .37 .35 .33 .31 .29 .28 .27 .26 2.94 .97 .49 .30 .19 .14 .10 .09 .06 .06 .04 .04 .03 .03 .02 .02 .02 .02 .02 .01 .01 .01 7.37 3.68 2.46 1.84 1.47 1.22 1.05 .92 .82 .74 .67 .61 .57 .53 .49 .46 .43 .41 .39 .37 .35 .33 .32 3.69 1.22 .62 .37 .25 .17 .13 .10 .08 .07 .06 .04 .04 .04 .03 .03 .02 .02 .02 .02 .02 .01 11.74 5.87 3.91 2.94 2.35 1.96 1.68 1.47 1.31 1.17 1.07 .98 .91 .84 .78 .73 .69 .65 .62 .59 .56 .53 .51 .49 .47 .45 .43 .42 5.87 1.96 .97 .59 .39 .28 .21 .16 .14 .10 .09 .07 .07 .06 .05 .04 .04 .03 .03 .03 .03 .02 .02 .02 .02 .02 .01 13.24 6.62 4.41 3.32 2.65 2.21 1.89 1.65 1.47 1.32 1.20 1.10 1.02 .95 .88 .83 .78 .74 .70 .66 .63 .60 .57 .55 .53 .51 .49 .47 6.62 2.21 1.09 .67 .44 .32 .29 .18 .15 .12 .10 .08 .07 .07 .05 .05 .04 .04 .04 .03 .03 .03 .02 .02 .02 .02 02 17.60 8.80 5.87 4.40 3.52 2.94 2.52 2.20 1.96 1.76 .60 .46 .35 .26 .17 .10 .03 .98 .93 .88 .84 .80 .76 .73 .70 .67 .65 .63 8.80 2.93 1.47 .88 .58 .42 .32 .24 .20 .16 .14 .11 .09 .09 .07 .07 .05 .05 .05 .04 .04 .04 .03 .03 .03 .02 .02 20.07 10.03 6.69 5.02 4.01 3.35 2.87 2.51 2.23 2.01 1.83 1.67 1.54 1.44 1.34 1.25 1.18 1.11 1.05 1.00 .96 .92 .88 .84 .80 .77 .74 .72 10.04 4.34 1.67 1.01 .66 .48 .36 .28 .22 .18 .16 .13 .10 .10 .09 .07 .07 .06 .05 .04 .04 .04 .04 .04 .03 .03 .03 24.42 12.21 8.14 6.11 4.89 4.07 3.49 3.06 2.72 2.44 2.22 1.04 1.88 1.74 1.63 1.52 1.44 1.36 1.28 1.22 1.16 1.11 1.06 1.02 .98 .94 .90 .87 12.21 4.07 2.03 1.22 .82 .58 .43 .34 .28 .22 .18 .16 .14 .11 .11 .08 .08 .08 .06 .06 .05 .05 .04 .04 .04 .04 .03 HAULAGE COSTS 345 The following are the results obtained through the instal- lation of five storage-battery locomotives in the Red Ash vein at Exeter Colliery of the Lehigh Valley Coal Co.* In the fifth vein, one locomotive, operating from inside slope to chambers, gangways and airways on roads Nos. 1002 and 1006, with chambers pitching both ways and grades in some places as high as 9 per cent, handles fifty cars with a maxi- mum run of about 1800 ft. When replacing the empty cars this locomotive is assisted by a second locomotive. The latter also covers chambers, airways and gangways on roads Nos. 1001 and 1006 and handles sixteen cars with a run of 1300 ft. in each gangway with grades ranging as high as 7 per cent. To replace these two motors with mule power would take fifteen mules, five drivers and five runners. In the Babylon vein, one motor hauls from the various working faces on roads Nos. 144 and 147 and handles twenty- five cars daily, having a maximum run of 2100 ft., and delivers coal to the head of No. 9 plane. It would take nine mules, three drivers and three runners to replace this motor. The fourth locomotive collects thirty-two cars from roads Nos. 39, 46 and 50 and delivers coal to a big turnout, having a maximum run in each road of 3500 ft., 2700 ft., and 3000 ft. respectively. Nine mules, three drivers and three runners would be required to replace this motor. The fifth locomotive, operating as a collecting locomotive on road No. 5 and in working faces between chambers Nos. 33 and 47, and assisting in concentrating coal from roads Nos. 65 and 4, handles thirty cars daily over a run of 1000 ft. having grades up to 5 per cent against loaded trips. It would require six mules, two drivers and two runners to replace this motor. As compared with mule haulage there is ease on record in the Ohio fields where one storage battery 6-ton motor re- placed 12 mules and four drivers, handling 85 to 100 cars on grades up to as high as 17 per cent it being necessary to sand the track on the steeper grades of course. The economy effected in this case, after generous allowances for deprecia- tion, is obvious. * Extract from "The Storage Battery Locomotive for Gathering Pur- poses" in Employees Magazine of the Lehigh Valley Coal Co. 346 COAL MINING COSTS Gasoline motor vs. animal haulage. The Long Branch Coal Co. installed a gasoline-motor haulage system at its mine in West Virginia, that materially reduced haulage costs and at the same time increased the output. The supplanting of the old system of mule haulage was done in 1913 with the idea of cutting down operating costs. The company purchased a 7-ton gasoline locomotive which was put into service in the summer of 1913. At the end of the month of August, a com- parison was made with the month of February, the most pro- ductive month during the regime of mule haulage. A summary of this comparison is given below : COMPARATIVE COSTS OF THE Two SYSTEMS OF HAULAGE Total cost of mule haulage, per month $810 . 00 Total cost of gasoline haulage, per month . . . ." 529 . 63 Decrease in haulage cost, per month $280 . 37 Total coal tonnage by gasoline locomotive, tons 11,601 Total coal tonnage by mules, tons 7,848 Increase in coal output, tons 3,753 The analysis of the above summary is shown by the follow- ing detail comparison between the cost of haulage by mules and by a combination of gasoline locomotive and mules for gathering. Mule Haulage Month of February, 1913 Length of haul one way, feet 2,000 Maximum grade against loads, per cent 5 . 625 Average grade against loads, per cent 3.0 Total tonnage per month of 24 days 7848 15 mules feed and upkeep per day @ 60c $ 9 . 00 11 drivers wages per day @ $2 . 25 24 . 75 $33.75 24 working days @ $33 . 75 cost per month $810 . 00 Total haulage cost per ton of coal . 103 Total haulage cost per ton-mile . 272 Gasoline Locomotive Haulage in Connection with Gathering by Mules Month of August, 1913 Length of haul one way, feet 3000 Maximum grade against loads, per cent 5 . 625 Average grade against loads, per cent 3.0 Total tonnage per month of 25 days 11,601 Expense of mules and drivers for gathering, 25 working days $313.20 HAULAGE COSTS 347 COST OF OPERATING LOCOMOTIVE, 25 DAYS 1 motorman @ $3 per day $75 . 00 1 trip rider @ $2 per day 50 . 00 300 gal. gasoline @ 20c. per gal 60 . 00 38 gal. engine oil @ 40c. per gal 15 .20 3 gal. black oil @ 16c. per gal 0.48 Cup grease . 50 Waste 0.25 Repairs on motor 15 . 00 Total operating expense of locomotive $216 . 43 Total haulage cost per month $529 . 63 Total haulage cost per ton of coal . 0456 Total haulage cost per ton-mile . 08.03 Cost per ton of coal mule haulage . 103 Cost per ton of coal locomotive haulage 0.0456 Saving per ton of coal $0 . 0574 Cost per ton-mile mule haulage . 272 Cost per ton-mile locomotive haulage . 0803 Saving per ton mile $0 . 1917 On a basis of 12 months the cost by mule haulage for one year ($810X12) $9720.00 By locomotive for one year ($529 X 12) 6348 . 00 Yearly saving $3372.00 With the gasoline haulage system the tonnage of the mine has been increased on an average of 25 per cent per month, and the company has dispensed with 6 double teams or 12 mules and 6 drivers, a total monthly saving in expense of $496.80. The management of the Midvalley Coal Co. in Pennsyl- vania, in 1911, effected a saving of 32.2 per cent on coal hauled by substituting the gasoline locomotive for mules. This comparison seems too conservative, because the loco- motive was in a position where its full capacity could not be demonstrated, in fact was idle a large part of the time, mak- ing only 24 miles per day when it is capable of doing more than twice as much. The management estimates that a second locomotive that was to be installed would save practically 50 per cent over' the present system of haulage as it was to be 348 COAL MINING COSTS placed on a level and have sufficient work to keep it moving, displacing 15 mules. By comparing the cost of haulage with the gasoline loco- motive at Midvalley and the average cost of electric locomotive haulage as furnished in the Coal and Metal Miners' Pocket- book, it was found that there was a lessened cost of 27.9 per cent in favor of gasoline. The locomotive uses naphtha for fuel, it being less dangerous and better to handle than gasoline. The consumption of naphtha is about 15 gal. per day ($1.50 per day), where if gasoline were used the consumption would be about 12 gal. for the same work while the cost would be 30c. more. The Midvalley locomotive, rated as a 9-ton locomotive, has .about the following dimensions : Length, 150 in. ; width, 59 in. ; height, 60 in. ; wheel base, 48 in. ; and diameter of driving- wheels, 24 in. The following table gives the detail of the average work performed daily by the first locomotive in 6 months, during which period approximately 2 hr. out of a 9-hr, day were devoted to switching, a feature which fails to show on the cost sheet: Average tonnage of loaded cars per day 550 tons Average tonnage of empty cars per day 250 tons Average mileage of loaded cars per day 12 miles Average mileage of empty cars per day 12 miles Weight of one loaded car 5 tons Weight of one empty car 1\ tons Average number of cars per train 8 cars COST OF OPERATION Wages of operator and helper per day $3 . 35 Cost of fuel per day 1 . 50 Cost of lubricating oils per day .12 Consumption of fuel in gallons daily 15 MAINTENANCE Cost of maintenance of locomotive for six months, including repairs and labor $65 . 14 The substitution of gasoline-motor haulage for mules at some mines near Rockwood, Tenn., in 1911, presented some interesting figures on the comparative costs of these two HAULAGE COSTS 349 methods of haulage. The coal in this mine is collected on side- tracks on the main entries by mules or by rope from cross- entries, the nearest parting being iy 2 miles from the slope. Mule haulage had been used on this long haul and this had been found so expensive that it was decided to install three gasoline motors. The total output of the mine was between 600 and 700 tons a day all of which passed over this long main- haul distances varying from 1% to 2 miles. All the extra work in the mines, necessary for the instal- lation of these motors, was some slight trimming of the rib and top in places, so as to give ample clearance for the motors and going over the track to replace with 20-lb rail, the places on the entry where a lighter rail had heretofore been used. There was no difficulty found by reason of the many curves, as the motors have a 4-ft. wheel base, and can take a curve of 25-ft. radius. The locomotives are 6 tons each, and were built for the mine gage of 33 in. They are designed with 4-cylinder engines, of ample power to slip the wheels, and all parts are well protected, as is necessary for mine use. The mine cars used are about 1400 Ib. in weight, and carry ! 1 / 5 tons of coal. As the grade is in favor of the loads, the empty cars up the entry make the load for the motor. The regular 20-car trips are handled without difficulty, and on trial trips 40 cars have been taken up the entry. These three motors replaced 23 mules. The comparative estimate of mule and motor haulage on one entry was as follows : 10 twenty car trips equals 224 tons By mules: 4 drivers, at $1.65 $6.60 9 mules, at 50c 4.50 $11.10 By motor: 1 motorman, per day $2 . 05 1 coupler, per day 1 . 65 13 gal. gasoline, at ll^c 1 . 50 2 Ib. carbide, at 4c 08 \ gal. gasoline engine oil, at 23c 12 1 gal. transmission case oil 24 $ 5.64 Saving by motor $ 5 . 46 Or. 49 . 1 per cent 350 COAL MINING COSTS These motors use 12 to 13 gal. of gasoline each, per shift. The Connellsville Central Coke Co. converted from horse to gasoline motor haulage in 1915 and comparison of the results obtained are of value. Twenty-nine horses had formerly hauled the output of 700 cars, to the haulage rope or to the shaft bottom as the case might be. The coal from some of the flats was handled independently of the haulage rope. This repre- sented a net tonnage of only 41.5 tons per horse, the tonnage per unit being low, not only because of the excessive length of haul, but because of the heavy grades, which averaged 6.5 per cent in the butt headings. The situation evidently needed corrective treatment. Four hundred mine cars are used in hauling the coal. These are 44-in. track gage and have a capacity of approximately 4000 Ib. per car, which weigh when empty about 2000 Ib. As the daily output is 1700 tons, it is about twice the capacity of the mine cars. On all main haulageways 40-lb. rail is used, and on the flats and subsidiary butts the rails weigh 25 Ib. per yard. The joints are all fishplated, and the track is well ballasted and carefully graded. A 5-ton gasoline locomotive was selected and put in opera- tion in September, 1914. It pulls 15-car trips in No. 5 flat right and 20-car trips in F flat on the left, the latter being a one- way haulage of approximately 2000 ft. and the other being roughly half that length. The cars are delivered to the rope at the main butt entry. The grades on both headings are partly in favor and partly against the load. Thus in the No. 5 flat there is a grade which averages 1% per cent against the load extending for the whole distance between two of the butt entries, and in F flat there is a grade averaging about 1.2 per cent against the load and nearly 1200 ft. long. It is easy to see that conditions more favorable might have been chosen. The operating expenses are as follows: Locomotive runner per 9-hour day $2 . 75 Trip rider per 9-hour day 2.60 11 gal. of gasoline at 12c. per gal 1 . 32 J gal. lubricating oil at 22c. per gal 06 Cup grease, waste, oil, etc .05 Total. . $6.78 HAULAGE COSTS 351 The results obtained from this locomotive were so satisfac- tory that a duplicate was purchased and shipped to the mine in January, 1915. This locomotive is working in C flat on a haul which measures 2700 ft. one way and in D flat where the haul is 3000 ft. one way. Both flats have grades favorable to the load. However, in D flat there is a grade about 300 ft. long which runs about 1.1 per cent against the load. At present each of these locomotives is hauling about 250 cars per day, and handling about 500 tons. It is expected that when certain grades are made more even and when delays are eliminated 300 cars loaded with 600 tons will be handled by each unit. For the months of April and May, 1915, the locomotive in F flat and No. 5 right averaged 260 cars daily; on this basis the comparative haulage costs per day are as follows : Horse haulage: 11 drivers at $2.60 $28.60 Feeding 11 horses at 50c. each 5 . 50 $34.10 Motor haulage: 6 drivers at $2.60 $15.60 Feeding 6 horses at 50c. each 3 . 00 1 motorman 2 . 75 1 snapper 2 . 60 Gasoline, oil, grease, etc 1 . 51 25.46 Savings per day accomplished by use of gasoline motor $ 8.64 The locomotives when fully equipped weigh 10,500 Ib. Over all they are 144 in. long, 55 in. wide and 46 in. high and their wheels are 18 in. in diameter. They are each equipped with 5 X 6-in. four-cylinder four-cycle engines of the vertical type, specially designed for mine-locomotive service and cap- able of delivering 25 hp. to the wheels at 800 r.p.m. The volume of tonnage is the principal determining factor in deciding when to substitute motor for animal haulage. Care must be exercised not to make a change before the motor be- comes the most economical method. It is not possible to prescribe exact limitations for animal haulage but as a general rule where the haul exceeds one-half mile or where the cost of hauling 300 352 COAL MINING COSTS tons on the main haulage way amounts to $1800 a year (these figures as of 1910) it is usually economy to install some form of motor haulage. A comparison between gasoline motor and mule haulage was compiled at the mines of the Shade Coal Mining Co. in Pennsylvania, the results of which are given herewith. The coal is handled in 185 mine cars of 2500 Ib. capacity and weigh- ing 1000 Ib. each. The outside and main entry haulage track is laid with 30 Ib. rail to the first gathering point in the mine, Total lengfh of hauJ in No.!, heading -4,400rt. Average grade in favor of load 2' Average grade, again empties approx. 2.2% l"< CO. HOUSES Kails Main haulage SO Ib. per yard Room and entries 20 Ib. per yard =Side Tracks, Main Haulage FIG. 23. Haulage layout at Shade Coal Co.'s Mine in Pennsylvania. beyond which 20 Ib. rail are used. A map of the haulage arrangement is shown in the accompanying drawing, Fig. 23. The first motor was installed in October, 1911, and was a 7-ton machine, with a 4-cylinder, 4-cycle engine, having a 6 in. bore and stroke and developing 35-hp. at 800 r.p.m. The second motor was a duplicate of the first and was put in service in March, 1913. The coal at this mine is hauled in one hundred and eighty- five 1000-lb. mine cars, having a capacity of 2500 Ib., all of these cars being equipped with plain bearings. The outside and main-entry track consists of 30-lb. rail, and back of the first gathering point in the mine, the rail is 20 Ib. The mini- mum radius of the curves is approximately 30 ft. HAULAGE COSTS 353 The first gasoline machine was put in operation, October, 1911, and was 7-ton, 36-in. gage locomotive. It has a vertical, 4-cylinder, 4-cycle engine, 6-in. bore and 6-in. stroke, which develops 35 hp. at 800 r.p.m. Mule haulage was used before the gasoline locomotives were introduced and at that time the following conditions pre- vailed : OPERATING CONDITIONS Length of haul one way, 2640 ft. Maximum grade in favor of loads, 2| per cent Maximum grade against empties, 5 per cent Tonnage per month of 24 working days, 9600 tons COST OF MULE HAULAGE 10 mules to handle tonnage @ 60c. per day $ 6.00 Above day rate includes feed, harness and shoeing expenses Eight drivers @ $2 . 25 per day 18 . 00 Investment in 10 mules @ $200 is $2000. Assuming the average life of a mule is 5 years, this gives a 20 per cent depreciation per year, 24 working days per month, the cost for depreciation per day will be .... 1 . 38 Interest on investment at 6 per cent per annum per day . 41 Mule haulage cost per day $25 . 79 Total mule haulage cost per month of 24 working days 618 . 96 Mule haulage cost, per ton . 064 Mule haulage cost per ton-mile traveled by loads .... . 128 GATHERING COST USING MULES AT MINE No. 1 Tonnage per month of 24 working days, 9000 tons 3 mules for gathering @ 60c. per day $1 . 80 Three drivers @ $2.50 per day 7. 50 Mule depreciation . 417 Interest on investment at 6 per cent per annum, per day 0.125 Total gathering expense per day $9 . 842 Total gathering expense per month $236 . 208 Gathering expense per ton . 0262 354 COAL MINING COSTS MAIN ENTRY GASOLINE HAULAGE AT MINE No. 1 Length of haul one way, 4400 ft. Maximum grade in favor of loads, 6 per cent Average grade in favor of loads, 2 per cent Maximum grade against empties, 5 per cent. Average grade against empties, 2.2 per cent Tonnage per month of 24 working days, 9000 tons Motorman for 24 days @ $2 . 75 $66 . 00 Trip rider for 24 days @ $2.75 66.00 Total labor $132.00 384 gal. of gasoline per month @ 15c . . . $59 . 52 24 gal. engine oil per month @ 30c 7 . 20 6 gal. black oil per month @ 15c . 90 6 Ib. cup grease per month @ 15c . 90 Waste per month 1 . 00 Total supplies 69.52 Repairs: Material $28.95 Labor.. 6.86 Total repairs 35 . 81 Depreciation on locomotive at 10 per cent per annum, per month 29 . 16 Interest on locomotive investment at 6 per cent per annum, per month 17 . 50 Total operating cost per month $283 . 99 Operating cost per day $11 . 83 Operating cost per ton 0.0315 Operating cost per ton-mile traveled by loads.. 0.0379 GATHERING COST USING MULES AT MINE No. 3 Tonnage per month of 24 working days, 3,500 tons 1 mule required for gathering at 60c. per day $0.60 1 driver required per day 2 . 40 Mule depreciation per day . 138 Interest on investment at 6 per cent per annum, per day 0.0416 Total gathering expense per day $3 . 1796 Gathering expense per month $76 . 308 Gathering expense per ton 0.0218 HAULAGE COSTS 355 MAIN ENTRY GASOLINE HAULAGE AT MINE No. 3 Length of haul one way, 3100 ft. Motorman for 24 days at $3 per day $72 . 00 Trip rider for 24 days at $2 . 75 per day 66 . 00 Labor $138.00 120 gal. gasoline per month @ 15c $18.60 4 gal. engine oil per month @ 30c 1 .20 2 gal. black oil per month @ 15c . 30 2 Ib. cup grease per month @ 15c 0.30 2 Ib. waste per month @ 15c 0.30 Total supplies 20.70 Repairs : Material $28.95 Labor.. 6.86 Total repairs 35 . 81 Depreciation on locomotive @ 10 per cent per annum, per month 29 . 16 Interest on locomotive investment @ 6 per cent per annum, per month , ? 17 . 50 Total operating cost per month $241 . 17 Operating cost per day $10.04 Operating cost per ton 0.0691 Operating cost per ton-mile 0. 1173 Mine No. 3 is under development, and the motor, as well as the gathering mule and driver, are not working to exceed 40 per cent of the time. The gasoline consumption, and the amount of sand used, varies with the weather. The gasoline consumption runs from 4 to 7 gal. per day, bad weather caus- ing more wheel slippage and a higher engine speed, and, there- by, increasing gasoline fuel consumption. SUMMARY BASED ON EXPERIENCE IN MINE No. 1 Mule haulage cost per ton $0 . 064 Gasoline haulage per ton . 0315 Saving per ton 0. 0325 Mule haulage cost per ton-mile . 128 Gasoline haulage per ton-mile . 0379 Saving per ton-mile . 0901 The following is a typical example of gasoline consumption 011 a 7-ton motor, developing 50 hp. at a speed of 500 r.p.m. On a break test this motor would consume about a pint of 356 COAL MINING COSTS gasoline per horsepower-hour, equal to 50 pints or 6*4 gal. per hour. Theoretically, therefore, this motor would consume 50 gal. of gasoline per 8-hr, shift, while in actual practice the consumption varies from 15 to 18 gal. indicating that the aver- age horsepower developed varies between the same figures. The following are some of the results obtained with an Otto internal combustion mine locomotive, working at the Bar- ton mines in Nottingham, England, about 1910. The length of haul inside the mine is about 700 yd. and on the surface about l 1 /^ miles. With the previous horse trac- tion one round trip on the surface line occupied over two hours, including shunting at the tipple, hauling a train of 10 loaded wagons of about 20 tons total gross load. The locomotive makes one trip with the same number of wagons in three- fourths of an hour regularly, and in some cases in 35 min. The line is for the greater part level but partly in favor of the loads. The heaviest grades are 0.77 per cent against loads and 4 per cent against empties. The gasoline consumption during 27 working days was 38 gal. which equals 1.4 gal. per day from 7 a.m., to 5 p.m. Dur- ing this time 1274 net tons of stone were hauled. This shows that with one gallon of gasoline, 33% net tons were covered; at a tare of 10 cwt. per wagon this represents a total gross load of 47.6 tons, this being over a line of 1% miles, so that 71.4 ton-miles were covered with one gallon of gasoline. The locomotive is fitted with two speed-gears in either direction, i.e., for 3% and 7% miles per hour. This locomotive has replaced six horses which cost $14.40 per week in fodder alone, while also requiring four boys. The locomotive only requires one driver and one boy for shutting the gates, when crossing the roads. The gasoline consumption per week of about Sy 2 gal. at 16c. exclusive of rebate, amounts to about $1.34 per week. . The Germans and Austrians were the pioneers in the use of gasoline motors for mine use. In 1910 it was estimated that there were about 300 of these in use in various parts of the world. Gasoline motor haulage costs were estimated in 1910 on underground haulage work where the tracks were inferior and curves sharp at 2.4 to 2.6c. per ton-mile. Actual working costs HAULAGE COSTS 357 taken over a sufficiently long time to give reliable results were found on the surface track to be as low as 1.2c. per ton-mile after deducting 20 per cent for amortization. Fuel consumption on one type of gasoline motor was found to be slightly less than 0.1 gal. per hp.-hr. when working under full load. As, however, the motor is never run at full load except when starting and running up grade, it is found that 0.05 gal. per hp.-hr. is the normal consumption. Storage battery and trolley motor haulage costs. Some excellent figures on repairs and maintenance costs of storage battery motors on metal mine work at the Bunker Hill and Sullivan Mine are given on p. 229, Vol. 51 A.I.M.M.E. It was found there that low voltage, as compared to the 500-volt d.c. used on the trolley-type locomotive, practically eliminates brush and commutator troubles, which always have been a source of heavy expense. About the only charge against the batteries is the time of one man for a few minutes each morn- ing, giving them the daily inspection and refilling the cells with distilled water to replace that evaporated during the previous day the amount of distilled water required for three batteries being about 20 gal. per week. In addition, there is a monthly charge, not exceeding $10 per battery, for cleaning and overhauling. The principal source of repair expense on the locomotives was for new wheels. The figures given below are the total average monthly cost of repair and upkeep from the date of installation to Novem- ber 1, 1914, and also include the cost of installation, which was quite large, because the Battery boxes had to be altered and partly rebuilt, to adapt them to the company's charging system, and to protect them from the water issuing from the chutes under which they pass. Average Average Repair Monthly Monthly Cost Repair Tonnage in Cost. Hauled. Cents. 2|-ton Jeffrey, No. 11 level $48 513 13,501 0.359 4-ton Westinghouse, No. 12 level 55 456 12,755 0.434 4-ton Gen Electric No 13 level 28 612 2,406 1 189 358 COAL MINING COSTS The last figure is high because the costs are figured only on the tonnage of ore hauled, and most of the material handled by this motor was waste. A comparison of these costs with the trolley motors which they replaced is interesting. They cover a period of two months in the first case and four months in the second, so that the figures must not be taken to represent an average cost over a long period. Separate repair costs were kept for all motors until January, 1913, which accounts for the short period taken for the above motors, when it is remembered that they were replaced by storage-battery motors in March and June respec- tively, in the same year. The 2%-ton Jeffrey trolley motor on the No. 11 level had an average monthly repair cost of $39.88 as compared with $93.912 for the 4i/2-ton General Electric working on the No. 12 level ; the average monthly tonnage handled by the two machines was 9154 for the 2y 2 -ton and 14,645 for the 41^-ton motors; and the repair costs per ton were 4.35c. and 6.41c. respectively. These figures do not include the initial cost of upkeep and the trolley wires and track bonding, which kept two men busy practically all the time, and which was consequently a heavy expense. No separate costs were kept for the two levels, how- ever, so they may be omitted in this connection. It has been estimated by the company's electrical engineers that, with a few minor improvements in the charging system, and a "better understanding of, and more careful attention given to the operation of these motors by the motormen, the cost of repairs and operation would be 75 per cent less than the trolley motors doing the same work. Repair records on a 4-ton Westinghouse storage-battery motor at the Big Five Tunnel in Colorado, covering a 6-months period, showed a cost of $12.60 per month, or 0.8c. per ton- mile. The motor was doing very light work during three months of this time so that the repair costs may be relatively high ; for two months during which time it was working under a more nearly full load, the repair cost per ton-mile was 0.65c. One of the chief objections advanced to the storage battery motor is the cost of the batteries and their comparatively short life which ranges from two to four years. It is doubtful, how- ever, if this extra charge against the motor will exceed the HAULAGE COSTS 359 cost of copper wire, bonding and twin cables and cost of upkeep for the trolley type motor. It must also be remem- bered that the storage battery motor uses less than half the power required to operate the cable and reel motor which it usually replaces on secondary haulage. Constant fluctuations in output in different sections of the mine makes it difficult to operate trips on any regular time- table as in the case of railroads. This is due largely to the fact that there are so many elements in the movement of the cars, a delay or accident in any one of which would derange the whole schedule. The system here described was in use in a mine of the Sterling Coal Co. in Ohio. The coal in this mine is handled over one main entry off which there are 11 butt entries having from 12 to 15 working rooms each. The motive power consists of one, 8-ton trolley motor on the main haul, two S^-, one 4-, and one 4%-ton motors for the butt entry hauls and eleven, 2!/^-ton storage battery motors for gathering. The 8-ton motor on the main haul handles 36-car trips taking all the loads from, and placing all the empties at a sidetrack along the main haul. The butt-entry locomotives take the empties from the side- track in trains of 12 cars and deliver them to one of the 2y 2 - ton storage-battery locomotives. Each butt-entry locomotive takes care of three butt entries. When the butt-entry locomotive brings in a trip of 12 empties, this trip together witK the 21^-ton storage-battery locomotive is pushed into a room. In the meantime the butt- entry locomotive makes into a train the 12 loaded cars which have been placed on the entry by the storage-battery locomo- tive. After the loaded trip is coupled up it is pulled down to the room where the empty trip and storage-battery locomotive are waiting and coupled onto them. The empty trip is then pulled with its battery locomotive out on the entry, after which the empty trip is uncoupled, the loaded trip taken to the part- ing by the butt-entry locomotive, and the storage-battery loco- motive then proceeds to distribute the empty cars and push them to the face of the rooms. The reason for pushing the empty trip into a room and pulling it out again with the trolley locomotive is to save the storage battery from handling the trip of empties while the 360 COAL MINING COSTS trolley locomotive is on the entry. The twelve loads together with two other similar trips taken from two other butt entries are taken to the siding to make up a trip of 36 loads for the main-haulage locomotive. With this system of haulage, a schedule is maintained approximately as follows: One entry produces 12 cars per .hour. This means that the 21^-ton storage battery locomotive in each hour places 12 empties at the face and takes 12 loads away to the entry leaving them just outside the room neck. When these loads have been placed on the entry, the butt-entry trolley locomotive comes along, leaves 12 empties and picks up the trip of 12 loads, taking it to the sidetrack. This trolley locomotive comes into each butt entry once every hour, and inasmuch as it has three entries to take care of, there are 20 min. available for taking care of each entry. The main-haulage locomotive, handling 36 cars per trip, must make four round trips per hour, or a round trip every 15 min. The average weight of coal loaded into each car is 2700 Ib. and at a rate of production of 12 cars per hour per butt entry for 11 butt entries, the capacity of this haulage system is approximately 178 tons per hour. On this basis, the haulage capacity per day of 8 hr., is 1424 tons. Prior to the installation of these locomotives the cars were handled in rooms by pushers, three being required on each butt entry. Now the motorman on the locomotive is the only man required to handle these cars. It is estimated that the saving thus effected will amount to approximately $15,000 per year. A comparison between this three-element and a two-element haulage system which might be used in its place, may be of interest. The capacity of each storage-battery locomotive in an 8-hr, day under the present system is 96 cars. If these loco- motives were eliminated and the butt-entry locomotives pro- vided with cable reels, to enable them to go up into the rooms to place the empties and pull the loads, a cable-reel locomotive would be required on each entry. The cost of a storage-bat- tery locomotive, such as is used here, and that of a cable-reel locomotive, such as would be required, are practically the same. With the two-element system, then, only 11 locomotives HAULAGE COSTS 361 would be required as compared with 15 under the present system, saving the installation of four locomotives representing an investment of approximately $7200. Assuming depreciation and interest at 25 per cent, this investment costs approximately $1800 per year. On the other hand, with cable-reel locomotives, two men would be required on each locomotive, or a total of 22 men for 11 locomotives. With the present system, only one man is required on each storage-battery locomotive and two men on each butt-entry locomotive, making a total of 19 men. At $60 per month per man, the present system thus effects a saving of $2160 per year, which more than offsets the interest and depreciation on the added investment. To this should be added the freedom from cable trouble and expense, freedom from the expense of more carefully laid and bonded track in rooms and the possi- bility of operating on a smaller generator equipment. Thus the economic advantages of this three-element system become evi- dent. No doubt there are many operations where this plan, which has proven its advantages and economy at this mine, can be applied with equal success. While the storage-battery locomotive is of course not abso- lutely safe in the presence of gas, the sparks it makes are not near the roof, so the danger is lessened, for the gas must be in large quantity if it is to settle so low as to be ignited. Another increase of safety with the battery locomotive results from its shortened electric circuit. The current does not pass from the power house to the motor and back through the rails to the generator, but the circuit is contained within the locomotive not even the wheels are included in its range. Thus if it is true that there is a risk of stray currents pre- maturely igniting shots, the current of the storage-battery loco- motive can meet the accusation with a perfect alibi. And, of course, as there is no conducting wire, there can be no short circuits to ignite gas, coal dust or wooden structures. The travel of the electric current along the drawbars and couplings of a train of mine cars has always been an objection to the trolley locomotive, as it has occasionally shocked men and ignited powder. For this reason operators have sought to improve the grounding of the traction rigging and to make 362 COAL MINING COSTS powder cars relatively nonconducting throughout. In other cases the powder car has been hauled around by a mule. The battery locomotive, however, having its current self- contained, does not offer any such risk. It seems that it would serve admirably for hauling men and transporting material, explosive and otherwise, into and out of the mines. By switch- ing off the electric current on the main haulage road, the load- ing of men on the man trip and their passage along the road at the beginning and end of the day would be robbed of its dangers, some of which, though unnecessary, are unavoidable so long as careless and ignorant men are employed. The use of a section insulator at the point of embarkation or disembarking has been occasionally adopted and has prob- ably saved many lives. In mines where all or part of the night load is quite light, it is customary to delay pumping till the mines are closed down. If undercutting is done at night and the coal is loaded and hauled out by day, there is a third shift during which the storage-battery locomotives can be charged. This tends to keep the load curve even and to save in expense. Additional boosting of the locomotive batteries can be performed during the lunch hours and when shifts are being changed, should the batteries need it and should the men walk to their work. It has been generally thought that the storage-battery repairs would be excessive, but offsetting this there are no trolley harps, wheels and supports to be maintained in con- dition. The chance of the armature bearings and poles heat- ing or rubbing is about the same in both battery and trolley locomotives. The battery renewals might, however, cost so much that the cost of harps, wheels and supports would appear only a trifling drawback to electric locomotives. Compressed-air and electric haulage costs. The general superiority of electricity for underground haulage has been too well established by its widespread popularity during the last two decades to admit of any serious controversy as to the relative economy of it and the compressed-air haulage methods. Certainly in any event, the economic possibilities of the air-motor are limited to specific and unusual conditions such in gaseous mines where there would be danger in the use of the electric motor and even here the motor would find its HAULAGE COSTS 363 greatest application for secondary haulage or gathering pur- poses. So long as it is still used, however, a few examples of its costs of installation and operation will be of value. The accompanying table excerpted from Vol. 34, A.I.M.M.E., gives the comparative cost of electric and compressed-air haul- age as worked up by the H. K. Porter Co., covering results with the fourth and fifth motors built by that concern. They represent the results of operations at the Glen Lyon plant in 1898. The table gives the cost per ton in preference to the ton-mile basis, because the delays at terminals forms so large a part of the time lost that it remains a fixed quantity, regard- less of the length of haul so that the best showing on a ton- mile basis is only obtained on long hauls. Com- pressed Air ELECTRICITY Estimated Actual Number of working days during year 160 *2362 $1.16 4.20 3.20 200 989 $1.20 4.23 3.20 1.67 5.95 14H 989 $2.84 9.31 3.61 3.68 8.42 0.46 0.61 2.50 J8.17 4.41 0.35 0.74 Output per day in tons Cost per day: Engineer, powerhouse Motorman Helpers (brakeman) Electrician Repairs to motors 74 Repairs to line Repairs to generator 0.57 Fireman Depreciation 14.74 4.73 1.63 0.25 0.47 2.32 J5.20 Interest ... . Interest, repairs and depreciation, 174 hp. boiler 0.22 Oil and waste for motor Oil and waste for generator Steam (fuel and firing) Totals $24.01 0.01015 $21.67 0.02192 $45 . 10 0.0456 Cost per ton * Tons of coal hauled, t At 5 per cent, t At 3 per cent. 364 COAL MINING COSTS In this table interest on the compressed-air motors is cal- culated at 5 per cent and at 3 per cent on the electric motors. The column headed "Actual" gives the results accomplished at the electric-haulage plant of the No. 2 Mine of the Hillside Coal & Iron Co.; under the column " Estimated" those results are given as re-calculated on the assumption of 200 working days in the year instead of 141.25. About 1900 one of the larger bituminous coal companies installed a compressed-air haulage system consisting of: 1 compound-steam three-stage air compressor, 800 Ib. pres- sure. 5300 ft. 5-in., triple strength, pipe-line. 3600 ft. 2 in., triple strength, pipe-line. Three 14-ton motors. The cost of this installation, exclusive of boiler plant was approximately $37,000. The time consumed for each trip was 45 min. There were two charging stations and the compres- sor was intended for the operation of all three motors but under actual operating conditions there was very little margin when one motor was in operation. The cost of maintenance of the high-pressure plant and the motors was $8 and $7.50, respectively, per day. An electric-haulage system was later installed at this mine, consisting of two, 150-kw. generators, directly connected to two 22 X 20 in. simple engines and six, 13-ton electric mine motors, feeder and trolley lines, etc., the entire cost of which was about $42,000, exclusive of the boiler plant. One of these motors performs all the work of the above described com- pressed-air plant, making the round trip in 30 min. and in addi- tion is used on other entries for two or three hours per day. That part of the cost of the electric plant properly chargeable, for the purpose of comparison, against the old compressed-air haulage plant would be about $7000. Also a few months before the compressed-air plant was abandoned, one of the motors was overhauled at an expense of $2000. SECTION IV TIMBERING COSTS The enormous quantity of timber, poles, lagging, mine ties, plank and lumber in general used in mine operations, makes this subject of great importance, and upon the intelligent handling of this material in the future will depend in a great- measure the cost per ton output of coal. Although wood has been in universal use since creation, there is a remarkable lack of knowledge as to its structure and behavior in its various uses, by those who might be ex- pected to know its properties, thereby using species totally unfit for certain purposes, and consequently expensive to the company. In the past, when timber was plentiful and cheap, it mattered little in many cases how long it lasted, as the ser- vice it gave was ample return on the cost; smaller quantities being consumed then, few thought it necessary to study the structure and behavior of wood in order to lengthen its service. How to buy timber. The cost of wood has increased enormously in the past decade, while the quality is steadily decreasing. The greater care in its selection and use is there- fore self-evident, in order to lengthen its life by elimination of infected, or poor quality. The same care should be used in the selection of the proper species for the various uses. It is necessary, therefore, to use improved methods in select- ing prop timber, mine plank and various other wood, first, by determining the species meeting the most important require- ments, or several qualities in combination as shown by actual experience, and tests. Second, the methods of procuring the supply. Third, the handling and preparation, and finally the placing in the mines. Conceding that the present prop material in healthy con- dition (such species as Southern short leaf, loblolly pine and spruce pine) is probably as good as can be secured at a reason- 365 366 COAL MINING COSTS able first cost, the next question is, when, how and where to purchase same. Timber cut between September and March is preferable to timber cut during the spring and summer months. Timber that is cut during hot weather is subjected to attack through a period when all insect life, such as worms, wood borers, beetles and other low plant life (fungus parasites) are most active in their work of wood destruction, and the timber during this period is in its most favorable condition, with fresh green sap, inviting attack of fungi spores and borers of all kinds, thereby causing the first signs of decay. Experience has shown that summer-cut timber does not give the satisfaction, nor will it give as long service as winter-cut stock. Black oak and red oak are approximately of equal value with short-leaf, loblolly and spruce pine, all of these being an easy prey of fungi, when in contact with soil, while white oak is of greater strength and durability. A white-oak plank, one inch full thickness, is equal to an inch and a half pine plank of equal width at an approximate decrease of one-third in cost per surface foot. Because of the use of heavy electric motors instead of mule haulage in mines, it is essential that improved provision be made to take care of mine tracks, by the purchase of mine ties, manufactured to a rigid specification, from live winter- cut white oak, chestnut oak and young chestnut properly seasoned before using, thereby reducing purchases and track repairs. Sound pine trees cut down in the winter season and cut into log lengths, stripped of their bark and piled in layers with sticks between each layer so that a free circulation of air can pass through the pile, will harden the exterior juices. This will form a coating which will, to a great extent, furnish protection from exterior checking and materially resist the attack of the fungus, spores and wood-destroying insects, which protection it does not have if cut during the summer season. Green and unpeeled pine timber placed in mines for gang- way use is sure to give short service and minimum strength. Consequently such timber is the most expensive for the service it renders. In such condition crushing is most liable and decay sets in quickly. Tests should be made to determine the most efficient species TIMBERING COSTS 367 for each particular use, bearing in mind cost and service and specifications covering such needs should be prepared and purchases and inspection made accordingly. Timber used and costs for United States. The following statistics on the timber used in the mines of the United States in 1905 are based upon data gathered by the Forest Service in cooperation with the United States Geological Survey. Nearly 14,000 mines were selected in which the use of timber seemed certain or possible, and from more than 5000 of these, reports were received showing that timber had been used. It will be noted that 2940 bituminous coal mines used nearly $6,400,000 worth of timber, while 216 anthracite minetf used over $4,400,000 worth, the average cost of timber per mine being $2170 for the bituminous and $20,524 for the anthracite mines. It will also be noted that the timber used in the 216 anthra- cite mines was of slightly greater value than that used in 1718 mines for precious metals. The much higher cost of the tim- bering required for the anthracite mines is due to several causes. In the first place, many of the anthracite workings lie at great depths, and some of the larger properties have many miles of gangways which have to be carefully main- tained. They are below water level, and, as a result of the combined action of air and mine water, the timbers decay rapidly. Some of the beds are of enormous thickness, and require vast quantities of timber in the construction of * l square sets" to support the roof and preserve the workings in over- lying coal. Moreover, since the hills in the immediate vicinity of the anthracite mines have been largely denuded of timber suitable for mine supports, operators are obliged to obtain their supplies from considerable distances. Pennsylvania, with 524 mines, used 37,826,000 cu. ft. of round timber and 55,716,000 board feet of sawed timber, cost- ing altogether $2,290,053. The average cost of the round tim- ber was 3.5c. per cubic foot and that of the sawed timber $17.39 per thousand board feet. In Illinois 400 mines used 10,342,300 cubic feet of round timber, costing 6c. per cubic foot, and 7,025,000 board feet of sawed timber, costing $22.04 per thousand board feet, the total cost being $778,186. The 325 mines in West Virginia used 6,716,000 cu. ft. of round 368 COAL MINING COSTS timber and 19,645,000 board feet of sawed timber. The total cost was $561,061; that of the round timber being 4.6c. per cubic foot, and that of the sawed timber $12.76 per thousand board feet. Next in order of total outlay for timber is Ohio, with $471,730 ; Iowa with $232,148 ; Indiana with $220,209 ; and Alabama with $216,221. None of the other states used over $200,000 worth of timber. Timber is more expensive in Colorado than in any other state. The average cost of the round timber in that state was 11. 6c. per cubic foot and that of the sawed timber $33.76 per thousand board feet. The large amount of mining has made a heavy demand for timber, and, although most of the round timber is obtained locally, much of the sawed timber must be shipped in from considerable distances at high freight rates. Round timber in Wyoming and New Mexico cost 10.4 and 10.5c. per cubic foot, respectively, or nearly as much as in Colorado ; but the fact that sawed timber was obtained locally kept its price down to $16.93 per thousand in Wyoming and $12.18 in New Mexico. The lowest average price reported for round timber was 3.3c. per cubic foot in Indiana, and for sawed timber $5.58 per thousand in Washington. It must be borne in mind, however, that in many cases the timber used was cut from land belonging to the mine operators, and the cost in- cludes only cutting and hauling. Reports were received from 216 collieries in the anthracite regions producing approximately 83 per cent of the total anthracite tonnage of the United States. Figures for the remaining 17 per cent were computed, using as a basis the reports actually received, assuming that conditions and require- ments were uniform throughout the state. The results of the tabulation show that 121,565,000 ft. board measure of sawed timber (equivalent to 10,130,000 cu. ft.) and 52,440,000 cu. ft. of round timber were used during 1905. The total value of the sawed timber was $1,842,000, or $15 per thousand feet board measure. The total value of the round timber was nearly double that of the sawed timber, being $3,468,000, or $6.60 per 100 solid cubic feet the approximate equivalent of the average . standard cord of 128 cu. ft. The total value of the round and sawed timber combined was $5,310,000, or about Sy 2 c. per long ton of coal, using as a basis TIMBERING COSTS 369 for the calculation the production in 1905 in round numbers 61,000,000 long tons. So far as reported, the kinds of wood have been tabulated separately, but in many cases the operators were unable to furnish information in regard to the quantity of each species used, and it has therefore been necessary to classify a large amount as " mixed" or * 'miscellaneous." ROUND TIMBER SAWED TIMBER Kind Cubic Feet Kind Board Feet ' Yellow pine 9,250,000 6,220,000 1,180,000 590,000 444,000 236,000 165,000 115,000 10,263,000 477,000 23,500,000 Hemlock 63,600,000 14,200,000 2,860,000 1,740,000 371,000 328,000 84,000 28,642,000 1,370,000 8,370,000 Oak Yellow pine Hemlock Oak Pitch pine Maple Chestnut Spruce Beech White pine Jack pine Pitch pine Spruce . . Mixed hardwoods .... Mixed softwoods Miscellaneous Mixed hardwoods Mixed softwoods IVliscellaneous Total Total 52,440,000 .121,565,000 Of the species used for round timber, yellow pine, of which a large amount is loblolly pine from the South, furnishes one- half. Oak ranks next, but furnishes a much smaller propor- tion, according to the reports. The proportion of oak would unquestionably be increased if the large items reported as 11 mixed hardwoods" and "miscellaneous" could be separated into species, and it is not improbable that oak would then dis- place yellow pine in rank. For sawed timber hemlock holds first place in quantity, while yellow pine ranks next. The amount of oak reported is doubtless too small, but an explanation is found in the classifi- cation for " mixed hardwoods" and ''miscellaneous," which contains over 37,000,000 ft. board measure, of which probably a large amount is oak. 370 COAL MINING COSTS Computing size of timber. There seems to be no definite rules by which the size of mine-gangway timbers may be cal- culated. Experience largely governs their use and, as both experience and judgment differ widely, in many cases the timbers are either too small or too large for the work required of them. The compressive stress of timbers wood or steel is often neglected, resulting in economic waste. Owing to the adjustment of stresses underground, it is impossible to calculate the loading of the timbers with mathe- matical accuracy; however, rules may be formulated that will give approximate results. The experienced miner knows the size of collar best adapted for certain conditions in the gang- way, so that by taking an average of the strength of collars used in several different gangways we arrive at a value that may be used in calculating other gangway timbers. Where the strata are horizontal or the dip is light the loads are applied normal to the collar, and produce therein bending stresses, while compressive stresses result in the legs, each leg taking one-half of the total load on the collar. Where the dip of the strata is great, however, legs and collars may be subject to both bending and compressive stresses at the same time. The compressive stresses in the collars may safely be ignored, as they are taken care of in the average bending load that a collar will support, but the legs must be calculated for both bending and compression. This bending load per foot of length will be assumed as being equal to that applied to the collar. When the diameter and length of an existing collar are known the safe load per foot of length that it will support may be calcu- lated by the formula: (1) in which w = safe load per foot of length of span (lb.); d = least diameter of collar (in.); 1 = length of clear span (ft.); /=safe unit fiber stress (1200 lb. per sq. in. of section, for long-leaf yellow pine and white oak, and 900 lb. per sq. in., for short-leaf yellow pine). TIMBERING COSTS 371 In calculating legs for compressive stresses only, the formula LJ700+15C+C2) Jc ~ ' a(700+15c) ' should be used, in which L = total load on leg (Ib.) ; a = area of cross-section of leg (sq. in.) ; c = length of leg in inches, divided by its least diameter in inches. The diameter of the leg must be assumed for trial, the most economical section being that in which the safe unit fiber stress is not exceeded. The following practical examples will show the correct method of using these formulas: Example 1. What size legs would be required for a long-leaf yellow-pine timber set in a gangway 12 ft. wide, with clear head- room of 8 ft.; the strata being horizontal? Assume 2000 Ib. per lineal foot for load on collar. Solution 2000X12 = 24,000 Ib., total load on collar, and 24,000-^2 = 12,000 Ib., total load on each leg. Assuming the least diameter of the leg to be 4f in., the area of section is 0.7854 (4f) 2 = 15.03 sq. in.; and c = 8X!2^4f = 21.94, which substituted in Formula 2 gives for the compressive stress in the leg, 12,000 (700+15X21.94+21.94 2 ) fc= 15.03(700+15X21.94) l172 *' per sq ' m ' Since 1172 Ib. is less than the unit stress for long-leaf yellow pine (1200 Ib.), a 4f-in. stick will support the load; however, in this case, it would be advisable to use a larger stick, say 8-in. diameter, to allow for defects in the wood and unforeseen bending stresses. Example 2. Design a long-leaf yellow-pine timber set for a gangway 12 ft. wide, with a clear headroom of 8 ft., the strata having a dip of 45 deg. It has been found that 2000 Ib. per lineal foot of collar is the average load in districts where the dip of the strata is very heavy; so we will assume this value in solving the problem. Solution. Reversing Formula 1, so as to give the value of d 372 COAL MINING COSTS and substituting the values given for the load per lineal foot of collar, unit fiber stress and length of span, we have wl 2 3 I 2000X12 2 Say The load producing compression in each leg is 24,000-^2 = 12,000 Ib. Assuming as before the same bending load for the leg, as for the collar 2000X8 = 16,000 Ib. bending load on each leg. Then, assuming 12^ in. as the least diameter of leg and solving for the allowable unit stress due to compression, we have by For- mula 2, since c = 8X!2-^12| = 7.68. L_f (700+ 15c) _ 1200 (700+15X7.68) _ a~700+15c+c 2 700+15X7.68+7.68 2 ~ MH'Jju The unit stress due to bending, for a total load of 16,000 Ib., as found above, is calculated by solving Formula 1 for /; thus, wl 2 2000 X8 2 m ' The unit stress due to compression is L 12,000 Jc 0.7854 d 2 0.7854 X12.5 2 The total stress in the leg due to bending and compression is therefore 1008.2+97.8= 1106 Ib. per sq. in. As the unit stress produced by the loads is thus shown to be nearly equal to but less than the allowable unit stress, a 12|-in. stick is the most economical section to use. When calculating beams of special cross-section, as steel I-beams, it is customary to employ an expression called the " section modulus " of the beam. The section modulus (S) of a beam is a value that multiplied by the unit fiber stress (/) of the material gives the bending moment (M) of which such beam is capable of supporting, as expressed by the formula: M=fS ...... .: .. . (3) But, for a beam uniformly loaded and supported at each end, the bending moment (M), in inch-pounds, is ,, I2wl 2 1 c 72 ,. N M= g =1.5wl 2 ..... . . (4) TIMBERING COSTS 373 Combining equations 3 and 4 gives for the section modulus 5=1.5^ ........ (5) Example 3. Design a steel timber set for a gangway 12 ft. wide, with a clear headroom of 8 ft., the strata having a heavy dip, assuming a fiber stress for steel, /= 16,000 Ib. per sq. in. Solution. The average load per lineal foot on the collar being assumed, as before, w = 2000 Ib. and the span being 1=12 ft., the required section modulus, in this case, is K wl 2 1.5X2000X12 2 :L5 T = ~Woo~ Any one of the following sections, taken from steel manufac- turers' handbooks, will be satisfactory: 8-in.-32.5-lb. Bethlehem girder beam, section modulus, 28.6; 8-in.-32-lb. Bethlehem H-col- umn, section modulus 26.9; 10-in.-30-lb. standard I-beam, section modulus, 26.8. A 10-in.-30-lb. I-beam is the most economical section, but an H-beam, or girder beam, is to be preferred as it has a broader bearing surface. Assuming the leg to be a 10-in.-25-lb. I-beam, which has an area a = 7.37 sq. in.; radius of gyration, r = 0.97; and section modulus, $ = 24.4; using Gordon's formula for medium steel, fixed- end column and solving for ultimate strength (P), we have 50,000 50,000 n + 36,000 r> 36,OOOX0.97 2 39 308 With a safety factor of 4, the safe unit stress is ' = 9827 Ib. Again, assuming the same unit bending load for the leg as for the collar, w = 2000 Ib. per lineal foot, and the length of the leg being I = 8 f t. the bending moment, applying Formula 4, is M= 1.5 wl 2 = 1.5X2000X8 2 = 192,000 in.-lb. This makes the unit fiber stress due to bending , M 192,000 7 _. 07 , j = - = n.. = 7869 Ib. per sq. in. The unit stress due to compression, for a 12-ft. span, the sec- tional area of the leg being a = 7.37 sq. in. is L 2000X12 2X7.37 374 COAL MINING COSTS which makes the total unit stress due to bending and compression 7869+1628 = 9497 Ib. per sq. in. This actual stress in the leg is less than the safe stress, there- fore a 10-in.-25-lb. I-beam will satisfy the conditions assumed in this case. The accompanying table sums up the results of a series of tests made by the U. S. Forest Service, to determine the effect of knots of different classifications on the crushing strength of certain varieties of timber. It will be noticed that in some cases the presence of knots seems actually to increase the strength. RATIO OF RESULTS OF STRENGTH TESTS ON KNOTTY TIMBER TO RESULTS ON CLEAR TIMBER, STRENGTH OF CLEAR TIMBER TAKEN AS UNITY Compressive Strength at Elastic Limit per Square Inch Crushing Strength at Maximum Load per Square Inch Modulus of Elasticity per Square Inch Douglas fir: Pin knots . 95 94 1 06 Standard knots Large knots 0.87 78 0.86 78 0.90 71 Western larch: Pin knots Standard knots . . . 1.12 98 1.04 89 1.19 1 00 Large knots 0.98 0.85 Western hemlock: Pin knots 96 97 1.00 Standard knots 94 91 97 Large knots 0.86 0.83 0.81 Pin knots are denned as sound knots % in. or less in diameter. Standard knots are defined as sound knots ranging from 1/2 to 11/2 in. in diameter. Large knots are also sound knots from l 1 /^ in. in diameter, up. The accompanying table shows what different sizes of steel rails will support when uniformly loaded and from these loads, the sizes of equivalent standard I-beams and the different TIMBERING COSTS 375 TABLE COMPARING STRENGTH OF STEEL AND WOOD FOR SUPPORTING MINE ROOFS TEN-FOOT SPAN Uni- form Load inLb. Re- quired Std. T-rail Standard I-beam In. Lb. White Oak Chestnut White Pine Sawed Round Sawed Round Sawed Round 1,015 16 3 16.5 5X3 5 6X4 5 6X4 5 1,385 20 3 16.5 5X4 5 6X5 6| 6X5 ftj 1,920 25 3 19.5 6X4 6| 7X5 71 7X5 71 2,450 30 4 22.5 6X5 7* 7X7 8 7X7 8 3,090 35 4 22.5 7X5 71 8X7 8| 8X7 8| 3,840 40 5 29.3 7X6 8 8X8 9 8X8 9 4,475 45 5 29.3 7X7 8| 10X6 9| 10X6 $i 5,225 50 5 36.8 8X6 9 10X7 10 10X7 10 6,290 55 6 36.8 9X6 9| 10X8 IQi 10X8 101 7,140 60 6 36.8 9X7 10 10X10 11 10X10 11 7,890 65 6 44.5 10X6 10 12X7 ill 12X7 iij 8,955 70 7 45.0 10X7 Hi 12X8 12 12X8 12 9,700 75 7 45.0 9X9 11 12X9 12i 12X9 m 10,765 80 7 45.0 10X8 11 12X10 13 12X10 13 11,940 85 7 52.5 10X9 ni 12X11 13 12X11 13 13,110 90 8 54.0 10X10 12 12X12 131 12X12 13| 14,180 95 8 54.0 12X8 12 12X13 14 12X13 14 15,670 100 8 60.8 12X9 12| 12X14 14| 12X14 141 TWELVE-FOOT SPAN 845 16 3 16.5 5X3 5 6X4 6 6X4 6 1,155 20 3 16.5 5X5 N 6X5 6| 6X5 6| 1,600 25 4 22.5 6X4 7X5 7| 7X5 7* 2,045 30 4 22.5 6X5 7 7X7 8 7X7 8 2,580 35 5 29.3 7X5 7| 8X6 8| 8X6 81 3,200 40 5 29.3 7X6 8 8X8 9 8X8 9 3,735 45 6 36.8 7X7 81 2 10X6 9| 10X6 9| 4,355 50 6 36.8 8X6 9 10X7 10 10X7 10 5,245 55 6 36.8 8X8 9| 10X9 11 10X9 11 5,955 60 7 45.0 9X7 10 11X8 11 11X8 11 6,580 65 7 45.0 10X6 10 10X10 111 10X10 111 7,470 70 7 45.0 10X7 10| 11X10 12 11X10 12 8,090 75 7 52.5 10X8 11 12X9 12| 12X9 12i 8,980 80 8 54.0 11X7 11 12X10 13 12X10 13 9,955 85 8 54.0 10X9 111 12X11 131 12X11 13| 10,935 90 8 60.8 10X10 12 12X12 14 12X12 14 11,825 95 9 63.0 12X8 12* 13X11 14 13X11 14 13,070 100 9 63.0 11X10 12| 13X12 14| 13X12 14| NOTE. Loads given in table are the safe uniform loads that T-rails will carry: Other members show sizes necessary for these loads. Timber presumed as seasoned. For green timber use loads. Factor of safety 6 (about). Fiber stress white oak, 1200 lb.; white pine and chestnut, 800 lb. Timber to be placed narrow side against roof. 376 COAL MINING COSTS wood beams have been calculated. Thus, taking a 20-lb. mine rail it is seen that the safe uniform load it will carry on a 10-ft. span is 1385 lb., then following this line to the right, under the same conditions it is seen that the 3-in. 16.5-lb. I-beam will do the same and save 3.5 lb. per yard, or about 14 lb. to a beam of 12-ft. length. Under white oak is found a 5 X 4-in. sawed or S^-in. round timber for this load, while chestnut and white pine require a 6 X 5 in. sawed or 6^-in. round timber. Quite frequently a requisition will call for a certain sized timber, say, an 8-in. round timber, to be 14 ft. long, to be used for a 12-ft. span. It is seen by the table that a 4-in. 7.5-lb. I-beam will carry the same load. The importance of the place and the length of time it is to be maintained should govern which should be used. Timber in a mine, if it carries approxi- mately the safe load, will seldom last much longer than 18 months, and the replacement plus the first installation will be more than the cost of placing the steel. This table is presented with the idea that it will be used by mine officials in the proper selection of material when re- placing or retimbering, and to give them some idea what their selection of sizes will carry in the weight of roof supported. The Scranton Mine Cave Commission conducted a series of interesting experiments to determine the comparative strength of various materials used for supporting mine roofs, a summary of the results of which are given in the accompanying table. Timber framing equipment. Where a large amount oi timber framing is necessary, say sufficient to require the ser- vices of four to five framers continuously, the introduction of machinery to replace these men should be considered. This will be something to eliminate the use of the two-man cross- cut saw, hewing, gaging and framing the ends as required. A regular timber framing machine for this work will cost from $2500 to $3000 (in 1914) and weigh about 9000 lb. It will require about a 50-hp. engine and boiler to drive it and will then only cut the framing gages at the ends. The total cost in place will be about $7000 allowing for a suitable build- ing to house it. A good and economical substitute for this is a slabber and swinging cut-off saw which can be purchased complete for TIMBERING COSTS 377 $420 (figures as of 1914). This equipment can be erected as shown in the accompanying illustration, Fig. 1, in which it will be noted that the cut-off saw is set horizontal instead of vertical and another saw introduced, though these do not both operate at the same time. This particular outfit is driven by an old 7 X 10-in. engine, connected up as shown. Description of Test MAXIMUM LOAD MAXIMUM SETTLEMENT Total Pounds Pounds per Square Foot Inches Per Cent Rectangular pillar of mine rock. . . . Timber crib filled with mine rock . . . Circular pillar of mine rock Pile of broken sandstone, small sizes . Pile of broken sandstone, large and small sizes 489,150 900,000 361,600 581,000 417,000 600,000 800,000 200,000 300,000 300,000 300,000 300,000 300,000 600,000 42,000 63,300 85,000 209,000* 150,000* 216,000* 228,000* 886,000 1,330,000 1,330,000 1,330,000 1,330,000 1,330,000 * 5.26 7.08 4.51 4.36 4.61 5.00 4.69 2.73 3.66 2.42 5.33 3.00 3.35 2.43 29 30 31 46 41 63 45 30 35 23 51 33 32 27 Pile of river sand Pile of small broken sandstone and sand Wet culm in cylinder Broken sandstone in cylinder Broken sandstone and sand in cylin- der Cinders in cylinder Wet culm in cylinder River sand in cylinder Pillar of mine gob * Pressure under 20X20 in. bearing plate. The posts are first sawed on the slabber, and then squared on the ends by running through the machine, say one hundred of them, and the saw is then raised up so as to cut just 2 in. deep. The horizontal saw remains stationary, the stick is shoved through and the cut made on top ; the carriage is then shoved on the horizontal saw and a slice taken out of the bot- tom; it is then pulled back and rolled over one quarter and the operation repeated; it requires four cuts to finish the end. 378 COAL MINING COSTS TIMBERING COSTS 379 The complete operation averages 3y 2 min. to frame both ends of an 8 X 8-in. post. The slabbing saw takes logs up to 8 ft. in length and the company made a set of dogs and some per- forated plates to hold pins, and bought an inserted tooth saw to replace the thin saw which came with it. The machine can also be used for sawing straight lumber of any dimensions up to 8 ft. in length, this particular instal- lation reducing lumber costs at the mine about one-half. A small wedge saw was added for cutting scraps into wedges which were made at a cost of %c. each as compared with a cost of 6c. when made by hand. Some of the other advantages of the machine are that the slabs made can be generally used for lagging in the less critical places and the machine fram- ing has been found more accurate and giving better fits in the mine. Timber preservatives. In 1906 the United States Forest Service, in cooperation with the Philadelphia & Reading Coal & Iron Co. carried on a series of experiments to determine the best method of prolonging the life of mine timber. It was found as a result of these studies that 45 per cent of the mine timber is destroyed by decay, while breakage, wear and insects accounted for the balance. Germs or spores which produce decay may gain access to the timber at any time before or after it is cut, though for the most part the disease is contracted in the mines from decaying timber near by. In untreated timber, rough surfaces of bark and wood furnish a foothold for the spores, which subsequently germinate and attack the wood tissues. Spores may also enter timber only superficially treated through checks, cracks, or nail wounds. For a fungus to exist it must have a definite amount of air and water, food, and heat. If mining conditions were such that the timber would be kept always wet or always dry, it would never decay. It is the alternating wet and dry condi- tions or continuous dampness which produce rot. Tentilation is a very large factor in the life of mine timber. Poorly ven- tilated gangways and air passages, with a fair degree of mois- ture and a fairly high temperature, are favorable to fungous growth, and hence to rapid decay. It is probably impossible to exterminate disease, sufficiently to wholly prevent decay in mine timber. Right preservative treatment, together with care- 380 COAL MINING COSTS ful handling of the timber will, however, reduce both to a minimum. The important part which insects play in the destruction of mine timber is seldom realized. They are for the most part brought into the mines with the timber. Regular and thorough inspection and the rigid condemnation of insect-infested timber would therefore greatly reduce the loss from this source. Insects bore into the sound wood and greatly weaken it and, moreover, leave holes which encourage the entrance of wood-destroying fungi. A good preservative treatment will protect the timber from insect attack, as well as prevent decay. If the bark is removed from the timber soon after it is cut, it will not be attacked by wood-destroying insects until the wood becomes old arid dry, after which it may be attacked by "powder post" and other borers. Sets of round gangway timber averaging 13 in. in diameter were chosen as the basis for the experimental treating work. These sets in the anthracite regions consist of two legs, usually 9 to 10 ft. long and a collar 6 to 7 ft. long, and they are usually placed on 5-ft. centers. The sets contain 26 cu. ft. of timber and the average life in the anthracite mines is two years. Experiments have shown that peeled timber is superior in durability to unpeeled timber. The space between the bark and the wood especially favors the development of wood- destroying fungi, and is a breeding place for many forms of insect life. When, after placement in the mines, the bark begins to flake off, the timber has already begun to decay. The cost of peeling timber before it goes into the mine ranges from 20c. to 50c. per ton of wood (figures as of 1905), accord- ing to local conditions and the kind of timber. Seasoning or drying gives mining timber greater strength and durability. A stick of wet timber has only one-half the strength it has when thoroughly dry. Though it is not prac- ticable for mining companies to hold their timber until it is absolutely air dry, peeled timber will dry out sufficiently in a few months to gain in both strength and durability. From two to four months is necessary for proper seasoning. To determine the possible loss in weight in round timber, due to peeling and seasoning, a test was conducted at one of the collieries of the company. Representative sticks of South- TIMBERING COSTS 381 ern loblolly pine, averaging 11 to 13 in. in diameter and from 9 to 10 ft. in length were chosen. This timber was weighed immediately before and after peeling, to determine the weight of the bark. It was then weighed every two weeks until, seasoned, to learn the weight of the water evaporated. The time of the year greatly favored rapidly seasoning. The short lengths into which the timber was sawed gave a large drying surface in proportion to volume and longer sticks would season more slowly. The accompanying diagram, Fig. 2, shows the I i, I* Z& 42 56 70 84- 98 Number of Days Seasoned FIG. 2. Percentage of loss of green weight in seasoning. average percentage of loss from the green weight in seasoning, a synopsis of the results of the tests being as follows : PEELING AND SEASONING TEST, SOUTHERN LOBLOLLY PINE ROUND TIMBER? APRIL 17 TO JULY 24, 1906 Per Cent Total loss of green weight by peeling 8.1 Total loss of green weight by seasoning 35 . 1 Peeling and seasoning 43.2 RATE OF SEASONING Number of Days Percentage of Green Number of Days Percentage of Green Seasoned Weight Lost Seasoned Weight Lost 14 16.2 70 31.4 28 20.5 84 33.7 42 26.2 98 35.1 56 30.3 If a mining company handles its own timber from the woods to the mines, the saving in freight made possible by peeling 382 COAL MINING COSTS and seasoning can readily be estimated. Labor is the principal factor in the cost of peeling, while the cost of seasoning must be represented by the loss of interest on the capital invested in the timber during the seasoning period. However, these additional items of expense are more than offset by a maxi- mum reduction in freight of from 30 to 40 per cent and by the far better condition of the timber with regard to both its life at the mines and the readiness with which it will take preservative treatment. The peeling of timber at the mines has been unsatisfactory and expensive, because of the limited amount of yard room and the accumulation of bark. The following considerations favor peeling in the woods: (1) The saving in the cost of freight due to peeling and seasoning; (2) the saving of yard room at the mines; and (3) the prevention of fungous disease and insect attack by early peeling. Peeling and seasoning mine timber unquestionably increase its durability. However, in order to prolong its life to the fullest extent, a preservative treatment is necessary. The in- creased life necessary to justify the cost of applying a preserva- tive by the several methods in vogue is indicated by the accompanying diagram, Fig. 3. Creosote ^j fl BRUSH if Soft Sot uf ion 7 TANK a 41 ?7 CYLINDER Creosote ; SJ FIG. 3. Increased life necessary to pay cost of preservation treatment. Impregnated wood resists decay because the preservative is antiseptic and excludes the moisture necessary for fungous growth. Timber used in mines was treated with a variety of preservatives under several methods of application. Both green and seasoned timbers were treated to determine both the rela- tive value of the treatments and the best method of handling preparatory to treatment. If treated at all the timber musl be peeled. The accompanying table shows the method of treat- TIMBERING COSTS 383 merit, the preservative applied, the cost of same and the cost of the treatment for an average gangway set and per cubic foot: COST OF TREATMENT Absorp- Cost of tion Method of Preservative Pre- Per Set per Treatment Applied servative of Ppr Cubic Gangway Timber JT ci Cubic Foot, Foot (25.8 Per gal. Cu. Ft.) Pounds f Creosote (dead oil of Brusn < coal tar) $0.09 $0.40 $0 . 015 1 Carbolineum .70 1.15 .045 fSalt solution, magne- sium, chloride 15 Open tank with- per cent .01 .50 .020 out pressure . Zinc chloride solution, 6 per cent .04 .90 .035 1 Creosote .09 2.85 .110 10 Cylinder with pressure (Zinc chloride solution, 6 per cent Creosote .04 .09 1.90 3.85 .075 .15 10 Brush treatments with both creosote and carbolineum were applied in two coats to the Pennsylvania and Southern pines. A large flat brush and kettle of the hot preservative are all that is required for this treatment. A very small amount of the preserving fluid suffices, but the cost of application in pro- portion to the results obtained is considerable. For small individual operators who cannot afford the cost of a large plant, brush treatments are feasible and economical. The disadvantages of brush treatments are : (1) The difficulty of completely covering the timber and filling all checks and cracks. (2) The very slight penetration secured. The subsequent checking or opening of the timber may often allow disease to pass through the shallow exterior band into the untreated interior wood. 384 COAL MINING COSTS Pitch pine and loblolly pine have been most successfully treated with both creosote (dead oil of coal tar) and a 6-per cent solution of zinc chloride by the open-tank process. The open-tank treatment as given in this experiment was briefly as follows: Green, partially seasoned, and thoroughly seasoned timber was lowered into the tank and immersed in creosote, or in a zinc chloride or salt solution, at a temperature of from 90 deg. to 120 deg. F. The temperature of the creosote was raised by the coils to from 212 deg. to 220 deg. F., and that of the zinc chloride or the salt solution to about 212 deg. F. In no case, however, was the temperature allowed to go above 240 deg. F. for fear of injuring the fiber of the timber and so decreasing its strength. When this hot bath was over the steam was turned off, and the timber was allowed to stand until the liquid cooled to a temperature of from 170 deg. to ^100 deg. F. The periods of heat and of cooling were varied for each kind of timber and for each stage of its seasoning. The time required for the cooling operation, which depended largely upon the temperature of the atmosphere, was usually from 3 to 12 hr. For the whole treatment the time varied from 6 to 20 hr. Loblolly and pitch pine can be successfully and economically treated by simple immersion in successive hot and cold baths in an open tank, at a cost of about lie. per cubic foot. Green timber is treated with far more difficulty than seasoned timber. AVERAGE AND REPRESENTATIVE TREATMENTS OF LOBLOLLY AND PITCH PINE BY THE OPEN-TANK PROCESS CREOSOTE AVERAGE ABSORPTION AND PENETRATION, LOBLOLLY PINE (PINUS T/FDAI) Absorption Depth of Condition of Timber per Cubic Foot, Penetration, Pounds Inches Green 5- 7 T-li Seasoned (1 to 2 months) . ... 12-15 2 -4 Seasoned (3 to 4 months) 20-25 5-complete TIMBERING COSTS 385 REPRESENTATIVE INDIVIDUAL RUNS, SEASONED LOBLOLLY PINE (NEARLY ALL SAP WOOD) Time Seasoned, Total Length of Treat- ment, Duration of Hot Bath, TEMPERATURE Absorption per Cubic Foot, Penetra- tion, Average, Maxi- Months Hours Hours F. F. Pounds Inches 3 24 7 230 240 22.0 4-5 3 24 4| 225 235 21.5 4-5 3 6* if 178 220 10.7 2-3 3 6 2 173 210 10.7 2-3 3 H 11 174 198 10.2 2-3 REPRESENTATIVE INDIVIDUAL RUNS, SEASONED PITCH PINE (HEART WOOD AND SAP WOOD) Time Seasoned, Total Length of Treat- ment, Duration of Hot Bath, TEMPERATURE . Absorp- tion, Pounds per Cubic Pene- tration, Width of Sap Wood, Aver- age, o jp Maxi- mum, O "IT* " Font Months Hours Hours -1? UU L Inches Inches 4 22 71 215 240 6| 3 4 1 4 22 7 218 240 12 If 11 4 22 7| 209 232 2U 3 2| i i SOLUTION OF ZINC CHLORIDE (6-8 PER CENT) THOROUGHLY SEASONED LOBLOLLY PINE (NEARL\ ALL SAP WOOD) Total Length of TEMPERATURE Length of Treatment, Period in Hot Solution, Absorption, Pene- tration, Average, Maximum, Pounds per Hours Hours F. F. Cubic Foot Inches 20 6 200 210 20 3-5 20 6 200 210 35 4-6 386 COAL MINING COSTS The difference in weight of green timber before and after treatment is by no means indicative of the amount of the preservative absorbed. The simple application of the hot liquid to green timber slightly reduces its weight and yields no penetration. The same application to seasoned timber slightly increases its weight and gives a slight penetration. Green timber after treatment may show a penetration of 1 inch without an increase in weight. Heart wood of both loblolly pine and pitch pipe is pene- trated with far more difficulty than is the sap wood of the same species. This is especially the case with pitch pine which clearly shows after treatment a distinct division between the treated sap wood and the untreated heart wood. Experiments indicate that for pine timbers of the same degree of dryness, or containing equal proportions of heart wood and sap wood, impregnation can be regulated by increas- ing or decreasing the duration of the cooling bath. During the year 1906-7 gangway timber of various species, peeled and unpeeled green and seasoned, and treated and untreated was placed in gangways in the collieries of the com- pany. Each and every kind and condition of gangway timber has been compared with the timber in most general use in the southern anthracite region, namely, green, unpeeled loblolly, and pitch pines. The object of this comparison is to prove exactly to what extent the experimental timber is superior to that at present used. In the course of the experimental work the following comparisons have been made: Species Compared Treatments Compared Loblolly pine (Pinus Tceda) . \ Creosote Brush Pitch pine (Pinus rigida) Longleaf pine (Pinus palustris) Chestnut (Castanea dentata) Red oak (Quercus rubra) Carbolineum {Creosote Solution of zinc chloride Solution of sodium chloride and magnesium chloride -, ,. , Creosote Cylinder S i r i i i Solution of zinc chloride The history of each set of gangway timber and each part of each set has been recorded in writing and on maps. These records include: (1) The date of setting; (2) the colliery; TIMBERING COSTS SUMMARY OF EXPERIMENTAL SETS OF TIMBERS . 387 COLLIERY Silver Creek Eagle Hill Wadesville Total Untreated : 11 loblolly 26 loblolly 37 44 loblolly 16 loblolly 98 Green unpeeled ^ 16 longleaf ' 112 loblolly 31 pitch pine 8 longleaf )36 loblolly f 26 pitch pine 221 8 black oak 7 pitch pino Total untreated 356 Brush treatment: Green 14 loblolly 9 loblolly 27 f 18 loblolly > 9 loblolly 38 Seasoned Carbolineum \ 5 chestnut 7 loblolly 9 loblolly 9 loblolly 28 loblolly 6 pitch pine 5 loblolly 22 42 Total brush treatment. . 129 Tank treatment: Green creosote (104 loblolly 7 chestnut I 6 loblolly 124 Seasoned 5 pitch pine 2 black oak f 20 loblolly 31 Salt Zinc chloride \ 11 pitch pine f 6 loblolly f 17 pitch pine \ 14 loblolly }" 11 \ 5 longleaf Total tank treatment. . . 97 Cylinder treatment, seasoned: Creosote 23 loblolly 23 Zinc chloride 50 loblolly 50 Total cylinder treatment. . 73 Grand total . 755 (3) the gangway; (4) the position in the gangway relative to the nearest chute and adjacent set of timber. The accompanying table gives a summary of the experi- 388 COAL MINING COSTS mental sets of timber placed in the mines of the Philadelphia & Beading Coal and Iron Co. in 1906. PRESERVATIVE TREATMENT APPLIED Method of Application Preservative Used Approximate Cost of Preservative Approximate Cost per Set of Gangway Timber of 26 Cu. Ft. Cost per Cubic Foot The preservative heated to 180 F. and applied in two coats with a brush Creosote (dead oil of coal tar) Avernarius carbo- lineum $0.09 per gal. 0.70 per gal. $0.40 1.15 $0.01* 0.04* Immersion in an open tank without pressure succes- sive baths of hot and cold fluid. Plant of simple con- struction Solution of common salt (15 per cent) Solution of zinc chloride (6 per $0.009 per Ib. 0.04} perlb. . 09 per gal. $0.50 0.90 2.85 $0.02 0.03* 0.11 Creosote (dead oil of coal tar) In a closed cylinder under vac- uum and pressure. Plant of complex construction Solution of zinc chloride (6 per cent) Creosote (dead oil of coal tar) $0.04} perlb. 0.09 per gal. $1,90 3.85 $0.07 0.15 Steel timbering. The following are the comparative costs of steel and frame timbering as found under actual working- conditions : In 1908 at their Maxwell colliery the Lehigh & Wilkes- Barre Coal Co. timbered a double-track gangway with 20 in. 65-lb. I-beam collars 17 ft. long between supports, and 8 in. H-beam legs 10 ft. 6 in. high in the clear, weighing, with base plates, 1720 Ib. per set. These took the place of wooden sets made of 24 in. round yellow pine timbers, the cost of which erected was $15 per set, weight 5040 Ib. and the life of which was two and one-half years. In view of their probable dura- bility, the steel sets were erected on concrete bases and this added to the cost of installation, which reached a total of $40 per set. Capitalized at 6 per cent interest, the value of the steel sets at the end of 15 yr. will be $95.86 each, while the capital- ized value of the six wooden sets needed in that time will be TIMBERING COSTS 389 $153.56. At the end of the 15 yr., the steel will have a scrap value per set of $12.03, while the wood will be worth nothing, a saving by the use of steel of $69.73 per set or $4.65 per year. The pump house at the Dodson colliery, of the Plymouth Coal Co., is 100 ft. long, 8 ft, high in the clear and 18 to 22 ft. wide. The roof is exceedingly tender and has caused all kinds of trouble in the pump house, especially in connec- tion with the pipes. Before retimbering with steel, 18 to 22 in. round timbers were used, on 2-ft. centers, practically skin to skin. It is estimated that the pump room was retimbered once a year. Beginning with April, 1910, the 70 sets of wood tim- bers were replaced by 48 sets of steel. The last steel set was placed December 15, 1910. According to figures furnished by John C. Haddock, president of the company, the relative costs were: 1. Wood 70 sets; weight per set 4150 Ib. ; cost per set f.o.b. cars at mine, $12; cost, erected in place, $34.50; total cost, $2415 ; life, one year. 2. Steel 48 sets ; weight per set 1483 Ib. ; cost per set f.o.b. cars at mine, $31.47; cost erected in place, including concrete footings, $61.47 per set; total cost $2889.09, or not quite 20 per cent more than wood. Based on its life, the cost per month of a wood set without interest was $201.25. The cost of the steel sets at the end of 16 months without interest was $180.57 per month. At the end of that time they had shown no signs of failure. At the No. 8 mine of the West Kentucky Coal Co., steel timbers are used in a slope, both for the main entry and air course. The sets are composed of a 10 in. 25-lb. I-beam collar and 4 in. H-beam legs. They are spaced 3 ft. centers and lagged with oak plank 3 in. thick on top, and 2 in. thick on the sides. Between the sets, concrete is placed up to a height of 4 ft. This makes a solid reinforced concrete slope from the entrance to the point where the ribs are hard and top good. According to figures furnished by W. H. Cunningham, general manager of the company, the comparative costs of wood and steel for his mine were (about 1912) : Wood Yellow pine creosoted ; size 12 X 12 in., 264 ft. b.m. ; cost at Sturgis, $10.56 per set; cost in place, $15.70; weight 1575 Ib. 390 COAL MINING COSTS Wood Native white oak; size 12X12 in., 264 ft. b.m.; cost at Sturgis, $7.92; cost in place, $13.06 per set; weight 1340 Ib. Steel Cost of steel at Sturgis, $9.75 per set; cost of plac- ing $1; cost of concrete per panel $5.16; total cost in place per set, steel alone $10.75, steel concreted $15.91 ; weight of steel sets 425 Ib. Saving in the use of steel without concrete, over native white oak, $2.31 per set, over yellow pine $4.95. Excess cost of steel with concrete, over white oak $2.85 per set, over yellow pine 21c. This favorable comparison is due to the high unit cost of the wood and to the elimination of waste. The safe uniformly distributed load on the wood collar is 1200 Ib., on the steel collar 26,000 Ib. The safe compressive strength of the steel leg is 43,200 Ib., while that of the wooden leg is 105,100 Ib. ; in the one case more than ample, in the other case out of all proportion. In some cases it has even been found, where transportation costs are not excessive, that the first cost of the steel timber- ing erected in place is equal to or but slightly more than the cost of a wooden installation of similar strength. It has also been found that it is possible to fit the steel exactly to the structural necessities, whereas it is impracticable in all cases to do so with wood. This circumstance eliminates an economic waste due to the use, for practical considerations, of larger sized sticks of wood than are really necessary. The accompanying table was prepared by a mining company in western Illinois and compares the relative cost of steel beam collars, square-sawed white oak beams and square-sawed yel- low-pine beams delivered underground. It also estimates the comparative maintenance costs for a period of 20 yr., based on the most favorable and the least favorable probable con- ditions. The safe working loads are based on the 1915 values for the materials under bending stress. The smaller table gives the first cost of such beams and their cost at the end of the 20-yr. period under the conditions assumed. It, therefore, represents what might be called a reasonable expectation of relative service in a particular dis- trict where the conditions as to transportation, cost of steel and the relative availability of wood are normal. TIMBERING COSTS 391 BJ ^ >" " 8 I o o w fe M W J g ** y y M ll-oi- Hfe M* ^ PQ 1-3 4 CO i-< O IN 00 t^ * Tji ooooooo GO * 00 CO t^ t^ lO O CO rH (N CO cococofo r-( rH T-KN O * Tj< CO CD rf< -* O IN (N O CO l> O IO O IN 00 CO rH O5 Tj< (N I-H CO CO IN I-H OOINOOOOCOCO T}< * CO (N 00 -^ IN * IN a . 0) O ^PQ, *" S ci^l* ^a J ( 43 ^ '43 ^ '43 '43 ^ '43 ^ '43 ^ '43 ^ CO CO CO CO CO > CD O fl> ^ O ^O ^_ 8 S 8 00 Oi (N O O ^ "^ ^^iiiisiiiiiii (N C (N 10 CO C5 co"co" ^HCO O <*< "5 COIN 1C * (N 00 t^ -< 1C TJ< CO TfiN O (N (N CO 00 XX X X X X X i-*OJ CO O5 C35 00 00 COIN 00 392 COAL MINING COSTS S - ^i S.8I a^ "53 o co oo CO O "5 CD 10 CO CO Ci t~ 10 CO CO 3 CO fci *2 * '+* * '^ *' '+* * CO CO CD ** CO N 00 (N GO 00 CO Oi CC W S a Unfavo Condit CM O GO W 00 T- OS 00 iO CO CO O * CO t a 1 i W H 3 o el 00 rj< CO CO * 00 T3 * H a T-I O CM I-H N O 3 11 i> * Tt< co * 06 r^ CO CM fH CN CS 1-1 i-i 3 Hi 8-g 3o'8o' : 8 i S3 OS 10 (N 10 lO CO CO Is ll COCOKOOCDO a Si >73 o3 C 3 coc5c5coc5SN s 3_o T* ^H CM O O 00 00 OQ O'S * O5 CO e eS o feO 2 tf t-H CXI CN I>(N 00 COCOt^- OOCOO i ^0 O5 10 (N 00 1^ -^ ^ C 1. CO 03 00 * ^-3^ Lining Slab FIG. 9. Reinforced-concrete sets or timbers for shaft lining. in the mix was such that when the batch was piled, it settled rapidly without agitation. A drier mix was attempted by way of experiment, but owing to the amount of reinforcement employed, it was found impossible to ram this concrete into place. The labor force required was six men, as follows: Two carpenters, setting up forms and keeping them in repair; one man wheeling forms onto skidways ready for filling, return- ing used forms to shop and cleaning the forms ; one man feed- TIMBERING COSTS 401 ing the mixer from stick piles of rock, sand and cement; one man delivering mix to forms and shoveling material into place ; and one mason ramming charge into final position. With this combination of men as many as four complete sets, consisting of 64 separate pieces, have been molded in one day of nine hours. The shafts lined in this way are of the three-compartment type (with two skipways and one manway), dipping at an angle of 80 deg. The compartments are 7 ft. 6 in. high inside, with a width of 6 ft. 10 in. for the skipways and 3 ft. for the manways, with the end plates and dividings making the great- est span 7 ft. 6 in. The weights of the different pieces comprising the set are as follows: Lb. Long section of wall plate 1035 Short section of wall plate 700 End plate 600 Dividers 645 39-in. studdles 268 Complete set, 16 pieces 8104 Taking the weight of No. 1 Western fir, which has been exposed to the weather in stock piles, as 33 Ib. per cu.ft., the above concrete set weighs almost three times that of a 12 X 12-in. timber set which the concrete set is intended to replace. Because of this additional weight of the concrete set, it was t'ound necessary to increase the erecting gang from the usual 5 or 6 men on the timber sets to 7 men for the concrete sets. In a vertical shaft to which the concrete sets are especially adapted, the number of men per gang might be reduced. The comparative cost of the concrete set and timber set, delivered, at the shaft collar is striking. The concrete set was delivered for $22.50, the timber set for $37.50. These figures are based on the following prices : Western fir $28.00 per M., f.o.b. car Crushed rock 35c per yd., f.o.b. shaft Conglomerate sand 60c per yd., f.o.b. shaft Portland cement $1.15 per bbl., f.o.b. works Reinforcement $12.00 per set, f.o.b. factory 402 COAL MINING COSTS Cement gun. The cement gun has come into favor for cer- tain uses in connection with mine timbering and the following are some representative costs of this work as of 1920: In the examples given, details of cost such as are available have been presented, and are incomplete, in that power and replacements are usually omitted. In the cement gun itself the gaskets for the cone valves require replacement from time to time, as does the outlet-valve body liner. Liners are used for the nozzles. Compressed air and water hose are subject to the wear which comes from frequent handling, and would require replacement no more frequently than drill hose. The material hose is subjected to considerable wear. Estimates of its life range from four to six months with continuous use. Nozzle liners last eight days and cost 80c. for renewal. The upkeep cost on the work in the Anaconda properties at Butte amount to 50c. per eight-hour shift per machine operated. The elements of cost of a single job in summary are : 1. Assembly of machine, compressed air and water piping, materials, mixing apparatus. 2. Preparatory cleaning of surfaces, placing wire reinforc- ing, staging. 3. "Guniting": Labor, power, water, cement, sand, lubri- cants. 4. Wear and replacement of liners, gaskets, hose, gun parts. 5. Tearing down, cleaning up and removal of apparatus. H. V. Croll has given the cost data shown in the table. In this example "gunite" was used to prevent the walls of a mine " tunnel" from slaking. M. S. Sloman described the coating of a coal mine slope with gunite in 1918. The surface was first cleaned and scaled. No reinforcement was used and the coating averaged y 2 in. in thickness, 1 :3 mixture. Timbers were covered with ^-in. wire mesh. The slope was 12 X 12 X 625 ft. The total cost was $7488.58, or 30c. per sq. yd. (3.3c. per sq. ft.). A 900-ft. section of this slope required 13!/2 8-hr, shifts for a working crew of eight men ; 2376 sq. ft., or approximately 100 cu.ft., was averaged per shift. Materials required were 540 sacks cement at 60c. and 1620 sacks sand at 12c. per sack. The working crew comprised one mechanic, one engineer on hoist, one operator on cement gun, two mixers, one nozzle man, one TIMBERING COSTS 403 man drying sand, and the part time of one man hauling sand. No mention is made of power cost. COST DATA OF PLACING "GUNITE" IN TUNNEL AT UNITED VERDE EXTEN- SION MINING Co., JEROME, ARIZ. 1 Gun man, also motorman $ 7 . 00 1 Nozzleman 5 . 60 1 Man holding lights 5 . 60 1 Man loading gun 5 . 60 1 Man cleaning roof 5 . 60 3 Men mixing at $5 . 60 16 . 80 Total labor .'.$ 46.20 Cement, 77 bags at $1 77 . 00 Sand 9 cu. yd. at $1 9 . 00 Air and supplies 10 . 00 Superintendence 5 . 00 Total $147.20 3750 sq. ft. at $147.20 = 4c. per sq. ft. Above crew placed 125 running feet of tunnel in one 8-hour shift, equiv- alent to 3750 sq. ft.; average thickness | in. Stephen Royce described the use of gunite at the Gary "A" shaft at Hurley, Wis. This is a steel, five-compartment shaft with the steel sets blocked in place with wooden block- ing and lagging of 3-in. tamarack planks wedged into the flanges of the I-beams. "Gunite" was applied to fireproof the lathing and wooden blocking and to protect the lagging from decay by keeping it from contact with air; also to prevent corrosion of shaft sets and water-proof the shaft. The surface to be covered was cleaned thoroughly. This was done partly with water under heavy compressed air pressure, partly by sand blasting, and partly by chipping the rust and accumulated coating from the steel. Reinforcement consisted of No. 7, A. S. & W. Co.'s triangular-mesh reinforcing wire for the side walls. This was stapled directly to the I-beams and to the lagging at intervals. To keep the wire mesh about one-eighth of an inch away from the surface to be covered was ac- complished by stapling the reinforcing wire on with nails underneath it. The I-beams, before the cement was applied, were covered with 1%-in. mesh chicken wire, clamped on with wire clamps. The total area of wall surface was 14,260.9 404 COAL MINING COSTS sq.ft.; of steel covered 3749.96 sq.ft., or a combined total of 18,010 sq.ft. The materials required were sand, 102.5 cu.yd. ; cement, 173 bbl. ; reinforcing, 14,260.9 sq.ft., and chicken wire, 3750 sq.ft. Linear feet of shaft was 263.13. The work required one foreman and six men for thirty-two working days. The cost was given as $9.30 per linear foot of shaft. As the area per linear foot equaled 68.4 sq.ft., the per-square-foot cost is found by calculation to be 13. 6c. The thickness of the coat- ing was given as 1% in. The "gunite" was applied in from one to three coats. Reclaiming timbers. Although mine props and sets of tim- ber are often broken a short time after being set, the broken ends are valuable as they can still be utilized for the purpose of cap-pieces, wedges, track ties, or for the building of "cogs" or "chocks." Also, post timber broken in a thick seam can often be used again, in a thinner seam at the same colliery. In some mines there is a considerable loss of timber, through the carelessness of miners who will let them lie in the waste where they are finally buried. By keeping a careful watch in their daily rounds through the mine, the mine officials can do much toward reducing this loss or waste of timber. It is well to emphasize the importance of drawing all kinds of timber, as the work proceeds, using, if necessary, some suitable appliance for this purpose. Timbers left standing in the waste often cause a loss greater than their own value, by preventing the roof from caving and frequently making it necessary to build extra packwalls or timber cogs to keep the roads open. The material for these packwalls often has to be transported a considerable distance; whereas, if the timber was drawn and the roof allowed to fall, there would be plenty of material for the building of all necessary packwalls in most cases. Again, under many conditions, when the roof does not fall but a large standing area is kept open a great weight is thrown on the timbers standing next to the face of the coal, with the result that these timbers are broken more quickly, or they kick out and the roof is ruptured at the face. When this occurs, the condition is bad, as the influence of the roof in breaking the coal after the latter is undermined, is destroyed. When the roof is of such a nature that it breaks readily, it TIMBERING COSTS 405 is a very good policy to set a line of large breaking posts, with good sized cap-pieces, on one side of the track, which should be carried along the straight rib of the room. As the face of the room advances, up to the last crosscut, there is not only a saving of timber, but the caving of the roof prevents the crushing of the pillar coal when the roof "weights" and cannot fall. In heavy pitching seams, the recovery of timbers is much more difficult and dangerous than in flat seams ; because the worked-out portion, from which the timbers are drawn, is located up the pitch, and any loose pieces of rock that fall when the post is drawn are liable to roll or slide down upon the men engaged in drawing the timber who are unprotected. The danger may be avoided, in part or wholly, by using a long i^-in. steel cable or chain that will reach from the tim- bers to the first crosscut, in which the drawing machine should be placed. This will not only afford the necessary protection for the men, but will enable them to recover a larger percentage of timbers. The cost of timber in pitching seams is much greater than in flat seams, owing to the labor required in handling the timber on steep pitches. In a seam pitching 35 deg. the cost of timber frequently amounts to an average of about S^c. per ton of coal mined. This was in a mine where the roof conditions were fairly good. V MISCELLANEOUS INSIDE COSTS TUNNELING COSTS American and foreign tunneling records compared. The accompany table gives some of the most creditable American tunnel records made up to 1909. The ranking order of some of the tunnels is probably open to discussion. That given, based on the 31-day record, is by no means necessarily the order of merit. There is a tendency among mining men to mention the highest one month's record, however, rather than to recall record figures extending over longer periods. A comparison of foreign records with our own is shown in another table. On the face of this comparison, American records appear to rank ahead of the Continental records. Thus the best month's record on the Gunnison tunnel actually exceeds the best Simplon record by 87 ft. While we have no explicit data concerning the kind of rock in which the record progress for the Simplon was made, we know that the Gun- nison record was made using air-driven augers in the soft shale of the east heading, in which 7500 ft. of progress was made for the first year, making the high average of 625 ft. per month. In the granite of the west heading, the best Gunnison record 449 ft. per month so that this is the figure that should in all fairness be compared with the Simplon record of 755 ft. While no long-time average figure is available for the Gunnison progress in granite, we can hardly anticipate a serious rival to that remarkable record in the Simplon of 426 ft. per month for 76 months. The top notch American record in granite, however, was that made in October, 1908, by the Elizabeth Lake tunnel at Los Angeles, Calif., where a progress of 466 ft.. was made in the 31 days ending Oct. 31. 406 MISCELLANEOUS INSIDE COSTS 407 SSJ | ||| 1 1 a n CG W 1 tf 1 Sip *o n a> +a . s * 1 ssi s? K^8 M S ~3'3 ' CS I o GO S " w 00 ill per month months in Driven ft || p || || ss S(S s| S &i &l i ftoi Mft^ 10 a a fl .IN gj 3 .1-1 c .(N CO ., oT oj +j oo oo S ^ ^'i^ " " 1 s^ I g 43 aj-C^ 2 2 O O s Si ** *'3 c8 I 8,1-5 a s 15-8 ^ -s < > OQ OQ 410 COAL MINING COSTS In America, the heavy Sullivan drills made the first Ophelia tunnel record at Cripple Creek, and then beat it by 54 ft. on the west heading of the Gunnison. The Leyner No. 9 hammer drill beat the Ophelia record at the Roosevelt tunnel work, and finally captured the American record previous to 1910 by its run of 466 ft. at the Elizabeth Lake tunnel, Los Angeles, Calif., October, 1908. In Europe the Ferroux percussion drill and the Brandt rotary core drill, both of foreign make, have established a long series of records that threw the performance of American drills far in the rear until the appearance of the Ingersoll- Rand drill at the Loetschberg tunnel in 1906. Since then, this American drill has not only eclipsed all American records by its September, 1907, record progress of 574 ft., but has exceeded all the previously established foreign records except three, namely, the Simplon (755 ft.), the Arlberg (641 ft.), and the Albula (607 ft.). But, if any one individual machine is to have the credit for the world record performance of tunnel driving up to 1909, that credit belongs to the Jeffrey A-2, air-driven auger with which 842 ft. of soft shale were removed in one month's time from the east heading of the Gunnison tunnel. When the Brandt drill accomplished the wonderful records of the Simplon, some engineers went so far as to say that its success spelled the finish of air-driven percussion drills. Others, however, pointed to the fact that in driving the Arlberg tunnel with Ferroux percussion drills in one heading and with the Brandt rotary drills in the other, the best month's record of the two differed by less than 1 per cent. As a matter of fact, an average pf the four best months * records of the Ferroux drills was 613 ft. as against 576 ft. of the Brandt drills. The top-notch record of each was 637 ft. for the Ferroux and 641 ft. for the Brandt. The secret of the great speed made in the Alpine tunnels appears to have been in the very careful study made of the various causes of delay in the successive operations of drilling, blasting, mucking out, and setting up the drills again. As a result, a radically different method of mounting the drills in the heading has been employed from that practiced in MISCELLANEOUS INSIDE COSTS 411 America. In the mounting of the drills, the return to the old carriage drill is seen, but there is no such cumbersome affair as the drill carriage used in the early tunneling operations in this country. The drill carriage at the Loetschberg consists simply of a small truck whose wheel base is about 4 ft., running on the regular heading track. Mounted on this truck in the longi- tudinal axis of the tunnel, and hinged to swing vertically, is an I-beam set with its web vertical and reinforced to give it lateral stiffness. On the forward end of the I-beam there is mounted what would, in America, be called a shaft bar, set transversely while the beam is pivoted so as to swing the bar horizontally. On the opposite end of the I-beam is mounted a counterweight. Four drills are mounted on the shaft bar, the compressed-air connections from these drills running back to one hose connection in the rear of the truck. When not in use, the shaft bar is swung so that it lies directly over the I-beam and the carriage can then be run anywhere over the heading tracks occupying no more room in the heading than two muck cars. This carriage is practically the same as that used by Brandt on the Simplon tunnel, the main difference between the two systems of work being found in the drills. Before blasting, a %-inch plate of steel about 6% X 3% ft. is laid down just ahead of the track end. After blasting, the gases are sucked out of the heading through a pipe of 24 in. diameter, the fan being capable of running either as an ex- hauster or blower. A cut is then mucked through the center of the muck pile down to the steel plate sufficiently wide to allow the arm of the drill carriage to introduce the bar carry- ing the drills into the top 01 the heading, the drill carriage, of course, running on the steel plate laid down ahead of the track. The shaft bar is then jacked firmly against the sides of the heading and drilling immediately begins on the top holes. Mucking out then continues while drilling is in progress. On account of the drills remaining on the carriage all the time with the air connections at the drills undisturbed, there is little chance for grit to get into the working parts and so impair their efficiency. The following table gives the approximate time required for the various operations in the heading. 412 COAL MINING COSTS Minutes Setting up drill carriage in the heading 20 Drilling 12-14 holes, 4 ft. deep, 2 in. diameter, at bottom . 60 Removing drill carriage from heading 20 Loading and firing holes 30 Clearing out smoke 20 Cutting center of muck pile to admit carriage 90 240 = 4 hours As before mentioned, work is carried on in three 8-hr, shifts, each shift being expected to drill two rounds and shoot twice making 7 ft. per shift advance or 21 ft. per day. Here we have an American drill adapted to a European system of work, beating all American records and threatening to rival those of the Simplon before construction is completed. It is apparent that system rather than the various European makes of drills is to be credited with the Continental records. It is interesting to compare these time items with those of the Simplon tunnel which were described in a paper on tunneling, read before the Eoyal Institute of Great Britain, May 5, 1900: Bringing up and adjusting the dri Drilling Minutes 11.... 20 105 Minutes 25 150 Charging and firing 15 15 Cleaning UD rock debris . . 120 120 Total, 260 305 = 4 hours, =5 hours 20 minutes 5 minutes The average advance made by this method is given as 3 ft. 9 in., or about 7% ft, per 8-10-hr shift. Perhaps the first American engineer to successfully apply a bonus system of payment for tunnel driving in addition to the usual wages was D. W. Brunton, of Denver. The follow- ing were the rates of payment used by him in driving the Cowenhoven tunnel at Aspen, Colo., in 1889, when the progress made per month exceeded 150 ft. MISCELLANEOUS INSIDE COSTS 413 Bonus for Progress, Feet every Foot 150-200 $1.00 200-250 1.50 250-300 2.00 300-350 2.50 Over 350 3.00 With these payments drillmen often made $120; helpers, $112 ; and laborers, $95 per month in addition to their regular wages. Wherever tried in America, the bonus system has generally proved a success. This is for the simple reason that an unruly giant will not especially exert himself to earn a $3 or $4 per shift. But make him a bonus-system proposition whereby exertion of wits as well as muscles may bring him from $6 to $8 per shift, and we have a transformation from halfway in- difference to keen interest in his work. Stimulating the spirit of competition between the various crews has in some cases given good results in increased work but in no case has the glory of beating the other fellow proved so satisfying as that extra $2 or $3 per shift. Cost of rock tunnel at a coal operation. As a general rule the only literature on driving rock tunnels is that descriptive of work in metal mines, or for irrigation, water supply and railroads. Such tunnels are usually driven by organizations specializing in this work and much effort is expended toward making records. In coal-mining operations and the opening of new coal fields it is often necessary to drive rock tunnels but only small men- tioin is made of them and the special methods used and equip- ment available. The tunnels here described were used to open a new develop- ment adjacent to the Utah Fuel Co.'s Clear Creek mine No. 1, about 1914, and were driven as haulage and air-course entries. In order to systematize and expedite the work as much as possible a separate organization was created made up of expert hard-rock men. Although kept distinct from the regu- lar operating organization it was necessary to coordinate their work with that of the mine in order not to interfere with the production of coal. 414 COAL MINING COSTS The main haulage rock tunnel was driven with a rectangu- lar cross-section having a minimum height of 8 ft. and a minimum width of 10 ft., while the air course, of a similar cross-section, had minimum dimensions of 8 ft. in height and 12 ft. in width. Over breakage and trimming have increased these dimensions on an average of one foot each way. Two tunnels were driven in order to provide for adequate ventila- tion of the new mine as it was impracticable to sink a shaft or drive an adit from the outside for air. The rock strata penetrated were sandstones and shales of medium hardness. The shales were the hardest to drill as the gumming of the cuttings made it difficult to keep the holes clean. The first halves of the tunnels, driven on a y 2 per cent grade, were parallel with the bedding planes of the strata and thus much harder to break. The last halves, driven on a 7^/2 per cent grade, took the tunnels off the bedding planes some- what but increased difficulties were encountered in the numer- ous small faults which were cut through as a main faulting plane was approached. On the average the ground was fair drilling but in some places it was necessary to use water under pressure to clean out holes so as to render it possible to drill them at all. In such cases the water, at 100 Ib. pressure, was forced into the holes through %-in. pipe 16 in. long used as a nozzle. The time necessary to drill the rounds varied from one hour and 45 min. to the full 8-hr, shift. The outside back holes were drilled first, then the outside breast holes, then the inside breast holes, and so on as shown in the accompanying sketches, Fig. 1, the numbers designating the order of drilling. The source of power was the power plant of the mine, con- sisting of six 125-hp. 72-in. by 16-ft. horizontal return-tubular boilers hand fired with slack or the poorest grade of run-of- mine coal coming from the mine, and two Thompson-Ryan generators of 175-kw. capacity, 320 amp. 500 volts, driven by 19 X 18-in. McEwan engines of 260 hp. each at 200 r.p.m. Power was charged against the work at the rate of 1.24c. per kilowatt-hour. Current from the generators was conducted about 200 ft. to a 24 X 64 ft. combination compressor and blacksmith shop, 40 longitudinal feet of which was used for the compressors. MISCELLANEOUS INSIDE COSTS 415 Two Ingersoll-Rand compressors of the Imperial type No. 10, were used. These compressors had a rating of 427 cu.ft. of free air per minute at 175 r.p.m. at sea level. Each machine was driven through a belt connection from a 500-volt 116-amp. 75-hp. General Electric continuous-current motor operating at 850 r.p.m. B-^ec oncf position 4? bar; double shi ft. SIDE ELEVATION Holes large so charge could be placed welf toward back of holes [>ouble primers used in Lifter and Cut holes, also m wethotes\ 'Fuse cut 2'difference for 'different kinds of holes, ^"difference tor Cut holes. HOLE DATA 16 HOLE ROUND KIND OF HOLE LENGTH INSIDE RD. LENGTH OUTSIDE RD. STICKS OF 60% POWDER 8xlV Back 6 ( -4 v _5-6" 8 to 10 Breast 6-8" 5-10' 9 toll Cut 7- Z" 6V 4* 12 to 15 Lifters 6-4" 5'- 6" IZtolB r STEEL DATA KIND MATERIAL LENGTH sire: GAGE Starters Cruciform Steel Z'-6" Z' 3" Seconds CruafbrmStieel 4-6" iV 2&* Thirds Cruciform Steel 6-6" I YZ' 2!^" Fourths Cruciform Steel ! L &"-&'-6 [ 2i4 r FIG. 1. Method of drilling rock tunnel at the Sunnyside Mine. Air was conducted to two receivers in the compressor house, one being 3 ft. in diameter by 8 ft. 3 in. long and the other 3 ft. 6 in. by 8 ft., then into the mine through a 4-in. pipe 3725 ft. long to a 3-ft. 6-in. by 8-ft. receiver, thence 350 ft. through a 3-in. pipe line to the starting points of the tunnels and through 3-in. branch lines to within 50 ft. of the faces, 1-in. air hose being used from thence to the drills. Drill steel was sharpened with a Numa rock-drill sharpener. For heavy blacksmith work the smith set up an old Sullivan 416 COAL MINING COSTS coal puncher for a hammer. This required only the fitting on of a hammer-block 4 X 6 X 6 in. in place of the bit of the puncher, setting the machine on a suitable frame and making a foot control. The expense of this makeshift was slight but the saving effected in the cost and in the grade of work done was quite noticeable. For forge fires a combination of coke breeze and slack coal from the Sunnyside mines of the com- pany gave good results. Chicago giant rock drills, size, 3% in. were used in driving, while jackhamers and Sullivan stoping drills were used for trimming and widening. Two drills were used at each face and these were mounted on 7-ft. double screw columns. On double shift work, the muck was always in the way during the first drilling so it was necessary to mount the drills on 10%-ft. single screw bars, having a single screw brace to the face to take up the vibration in the bar. In this way the upper rows of holes were drilled first with the bar set near the roof and by the time these were completed the muck was cleared away from the face and the bar could be set up in a lower position and the remainder of the holes drilled. A 3-in. water main having %-in. hose connections was used along the back entry to within 50 ft. of the face. From these connections to the faces, %--in. pipe was used with % X 16-in. special reducing nozzles. The explosive used was mainly 60 per cent dynamite with sticks 8 in. long and l 1 /^ in. in diameter. On single shift the machine men worked during the day, that is from 7 :30 a.m. to 4 p.m., and the muckers and drivers during the night shift beginning at 8 p.m. and ending when the muck was all cleaned up. The shift for the machine men, muckers and drivers, while nominally 8 hr., was considered finished whenever their work was completed. A bonus was given of 10 per cent of the day's wage for each foot over 3 ft. of tunnel driven that they averaged per shift. The bonus was calculated at the end of the month instead of each day, except in the case of men quitting before the end of the month. The average wages earned by this system by the machine men was $4.17 per shift. These men were required to do the drilling, extend both the air and water lines, make up primers, load and shoot holes. MISCELLANEOUS INSIDE COSTS 417 Timbermen were paid the same wages as machine men in- cluding the bonus. There was comparatively little timbering to be done so that machine men were used for this work and it was thought fair to allow them the same wages as they would have earned if running machines. There were four muckers to each heading who were paid $3.25 per 8-hr, shift and a bonus of 10 per cent of the day's wage for each foot over three feet of tunnel driven per shift, just as paid the machine men. By this system the average wage earned by muckers was $3.84 per shift. The muckers laid all track, not including switches, put down the mucking sheets, loaded and unloaded drill steel and kept track clean for 100 ft. from the faces. Drivers were paid $3.15 per 8-hr, shift without any bonus. They were required to take muck away from the faces to the parting, bring in empties and haul steel, powder, caps, fuse, rails, ties, etc., from parting to faces. The dumpers outside were paid $3 and $3.15 per 8-hr, shift without any bonus. They were required to handle the rock but the extending of the tipples was done by the regular mine carpenters paid $3.40 and $3.15 per 8-hr, shift. The blacksmith was paid $4 per 8-hr, shift and straight time for overtime. His duties were the sharpening of drill steel, general repairs and other blacksmith work. The com- pressor attendants were paid $3.50 per 8-hr, shift and were required to run the compressors and assist the blacksmith. The head foreman was paid $175 per month and his assistant $4.50 per 8-hr, shift. The work cf widening the main haulage tunnel for the pass- by and parting was done after the tunnels were driven to the coal. For this work Ingersoll-Rand Jackhamer drills with %-m- hollow steel were used. The men were paid for straight time at the same rate as for tunnel driving, no bonus being given. The shots were detonated by electricity, using 6X detonators. A total lineal distance of 700 ft. 8 ft. high was widened from 3 to 5 ft., the unit cost for the work being given on p. 418. If the cost of the widening operations is added to that of driving the tunnels the total expense per foot would be about $17, but if salvage is allowed on pipe and material left over 418 COAL MINING COSTS UNIT COSTS OF STRAIGHT TUNNEL Engineering $0. 0503 Superintendence . 9885 .,,. f Machine men lmg \ Machine men bonus Muckers Muckers' bonus Dumpers Extending dump Drivers Motor men Shafts Stable expense Horse killed Ditch and track Tracks and switches Blacksmithing 0.3844 P.H. charge Running compressor Handling rock Hauling . 3 . 8221 4.0331 1.3193 Power. . 1.8604 Repairs, compressor Pipe lines Repairs, drills Timbering 0. 4389 Blasting material 2. 7707 Housing Setting Machinery... Depreciation Miscellaneous Total cost per foot of tunnels . . . 0.6655 $16.3332 COST PER LINEAL FOOT OF WIDENING OPERATIONS Superintendence $0. 4675 Drilling 0. 3656 Muckers Handling rock \ Dumpers J. 1 . 3243 Extend dump Drivers Hauling \ Stable expense } 0. 3171 Motor men Blacksmithing 0. 1712 ( Compressor men \ ., Power (Pipelines } ' 1681 Blasting material 0. 5292 Miscellaneous. . . 0486 Total cost per linear foot . $3.3916 MISCELLANEOUS INSIDE COSTS 419 and subsequently used for coal-mining operations the cost per foot is brought down to $15, which would be the net cost. The average progress per shift for both single- and double- shift work during the whole time of driving the tunnels was 5.11 ft. The highest average per shift per month was 5.56 ft. The greatest distance driven in a tunnel in a single month was 323 ft. in 60 shifts. Single-shift work was found to be less efficient than double- shift work, although to a certain extent this was due to the forming of a working organization and to the necessary experi- menting with the drilling and shooting of the ground during the earlier stages of the undertaking. Some examples of tunnel costs. A tunnel located at Idaho. Springs, 37 miles west of Denver, in the lower Clear Creek mining district of Colorado started in 1893 and extended at intervals over a 10-yr. period, presents some interesting cost data. The original purpose of this tunnel was to cut the well- known veins on the line of the tunnel so as to receive royalties from the property owners for drainage and for transportation of ore. The yearly progress made is as follows : 1893, 80 ft. ; 1894, 1405 ft. ; 1895, 1992 ft. ; 1896, 2061 ft. ; 1897, 1080.5 ft. ; 1898, 99 ft; 1899, 455 ft.; 1900, 2285.3 ft.; 1901, 2923.5 ft.; 1902, 761.1 ft. ; 1904, 1389.3 ft. ; 1905, 672.5 ft. A summing up of the above footages gives a total length of 15,154.7 ft. from the original starting point. The first 80 ft. of the tunnel were driven by hand, but as there were no available data concerning this hand work, very little can be said about it. This method necessarily made the progress of the work very slow, and it was deemed advisable to install a compressed-air plant to supply power drills. The equipment was in duplicate, consisting of two Norwalk com- pound 14 X 16 in. high-altitude compressors and two 80-hp. boilers. These compressors supplied air to the drills, which were of the 3-in. Leyner percussion type. From 100 Ib. to 175 Ib. of dynamite were used to the round, and the greatest footage made was 160 ft. per month by a working force of 26 men. In October, 1899, there was a change in the management, and the work was done in a more systematic manner. The 420 COAL MINING COSTS equipment of the power plant was increased by the addition of another 80-hp. boiler and a 22 X 24 in. compound Norwalk compressor. The holes were placed in the breast according to the Ameri- can center-cut system with an extra plunger hole at the upper center of the cut. The side and cut holes were drilled by two 3-in. Leyner sluggers mounted directly on separate columns placed on each side of the tunnel, while one model 5 Water Leyner, mounted on an arm, drilled the back and plunger holes. One Slugger was started at the bottom and worked up, while the Water Leyner put in the back holes on that -side. The other Sluggers started at the top and worked down so that it was out of the way when the Water Leyner machine was ready to shift to the other column. This system, together with the use of high pressure air, 160 lb., resulted in deeper holes in less time, and therefore greater progress. In blasting, the cut holes were electrically fired first, then others until the whole cut was taken out clean, using about 100 lb. of 60-per-cent gelatin powder. The side and back holes were then fired, using from 50 to 70 lb. of 40-per cent gelatine powder. In no case were the holes tamped. The greatest footage made in any one month under this system and management was 267.6 ft., and a total of 2760 was made from September, 1899, to August, 1900, at an average cost of $28 per foot, or a total of $79,470. The premium system of wages for employes was used. This consisted of paying $6 for every foot driven over 160 ft. per month. This $6 was divided proportionately between the drill gang, powder gang, and muckers, according to their wages. The success of the rapid progress of the tunnel under this management can be attributed to the following points: 1. The use of high-pressure air, 160 lb., being the minimum pressure kept at the breast. 2. The use of the Water Leyner machine for the back holes ; for by putting in these holes to a greater depth, it was possible to put in a much deeper round throughout. 3. The premium system of paying, which offered an incen- tive to employes to do their best work. MISCELLANEOUS INSIDE COSTS 421 Below is given a typical monthly footage expense during the year above mentioned taken from the 1900 annual report : Drill-crew men and foreman $3.11 Trammers and drillers : . . . . 3 . 88 Blacksmith shop 1 . 13 Engineers 1 . 08 $9.20 Ammunition 4 . 09 Oil and waste .18 Coal for power . 3 . 78 Feed and shoes for mules .23 Drill repairs .81 Premiums on footage .61 Track equipment and repairs .78 Mine timber .31 Labor 1.92 Material 1 . 10 Labor, repairs along tunnel 1 . 29 Surveying .21 Legal expense '. .19 Insurance .07 Salary and office expense 2 . 39 Minor expenses .58 Total $27.74 Some interesting cost data were obtained in driving a rock tunnel at the property of the Iron Mountain Tunnel Co. at Superior, Mont., in 1906 and 1907. The new tunnel is 7 X 6 ft., on a grade of 4 in. per 100 ft., and has a drainage flume 12 X 12 in. laid in the flooring. The work of driving the tunnel which is to be 5602 ft. long was begun February 10, 1906. In October, 1906, it had been driven 1500 ft., and 4345 ft. had been excavated by September, 1907, leaving uncompleted a distance of 1255 ft. Excepting four stretches of ground, aggregating in all about 800 ft., the rock encountered was very hard, requiring a heavy expenditure of explosives. In the distance the tunnel has thus far been driven, it has been necessary to timber only 535 ft., the formation requiring support being encountered in four different places, one of them 325 ft. long. The accompanying table gives the monthly costs of the work up to September, 1907 : 422 COAL MINING COSTS Time of Driving Feet COST PER FOOT FEET DRIVEN Gross Driving and Equipping Driving Side- track Cross- cuts Prior to June, 1906 559 231 215 229 227 285 264 260 288 267 244 251 254 251 206 228 170 $20.11 14.81 14.96 13.29 12.95 12.66 15.77 12.08 12.09 13.81 13.03 14.35 13.37 14.65 18.11 15.95 $18.70 11.81 14.85 12.31 12.59 12.30 15.36 11.79 11.76 13.44 12.68 13.99 13.00 14.22 17.67 15.55 $15.58 11.01 14.15 11.59 11.80 11.00 13.39 10.64 10.69 12.75 11.53 13.03 11.88 13.02 16.27 14.45 40 80 85 100 60 50 70 43 50 64 100 10 9 4 1906 June . . . July August September October November December 1907 January February March April May June July August September. . Totals Averages 4429 233.1 $231.99 14.50 $222.02 13.88 $202.78 12.67 742 46 23 H NOTE. No report was made of costs per foot of the first three months' operation. The city of Los Angeles, in California, started constructing an aqueduct about 217 miles long in 1909, including 105 tunnels whose aggregate length is 28 miles. The Elizabeth tunnel, which is 26,860 ft. long, was driven from both ends, termed the north and south portals. This tunnel has a cross-section 12 ft. 4 in. X 12 ft. 9 in., a grade of 1 ft. in 1000 ft., and a water capacity of 1000 sec.-ft The general equipment at this tunnel consists of the follow- ing: Two compressors, 520 ft. capacity, belt driven from elec- MISCELLANEOUS INSIDE COSTS 423 trie motors; two motor-generator sets, 150 hp. ; one 50-hp. electric locomotive; one 30-hp. electric locomotive; nine water Leyner drills ; 38 rocker dump cars, 32 cu.f t. capacity ; one drill sharpener. In addition, the machine-shop equipment included a lathe, drill press, saws, blowers, motors, and the necessary tools for such work. Most of the destructible equipment was supplied in duplicate, and extra machine drills were furnished. Each shift was supplied with a tool box and all the tools necessary for its members' work. These tool boxes were locked and one man on each shift was held responsible. A station was cut in the tunnel where all repairs to machines, hose, tools, etc., were made. Wherever possible each individual was held responsible for the tools he used. Each shift was required to drill and blast, the length of round being regulated by the nature of the ground encoun- tered in the first hole drilled. Discipline and strict attention to business while on shift were required of everyone, while at the same time a spirit of friendliness was fostered and every man made to feel that in a large measure the success of the project was due to his own efforts and the interest he took in the work. A bonus of 40c. per foot per man for every foot driven over 2 2 / 3 ft. per shift was paid, bringing the wages up to a good figure above the scales usually paid in other mining camps and consequently bringing a better grade .of men than could have otherwise been obtained. The rock encountered was mainly granite, which in places shades into both gneiss and schist. The granite is composed mainly of feldspars and biotite with only a small amount of quartz usually present. Its texture varies at different points along the tunnel, the finely crystalline rock usually being with- out joints or seams and the coarsely crystalline rock usually being full of seams. The finer crystalline rock, therefore, allows much the slower progress in tunnel driving, while the blocky ground gives about as near ideal conditions for record- breaking drives as one is apt to find. The rock in the south heading is hard, unaltered, and in general blocky enough to afford fine breaking qualities. It requires no timbering except one or two sets at long intervals 424 COAL MINING COSTS J Timber- ing N O >O 00 H M O5 00 O5 00 QW 0000^00 O O5-* O'* O>O O O5O OO5 IO U5 O ^- CO O CD t~ t~- t- O3 O5 -fOCC l> t^ (N O . CO r-t rj< ^H ^H (O (O lO IO ^ 10 >c CO C Tf >O t> 00 CO O t-- * O5 Tj* (O (N 00 T}< COOOiO 1-1 OS CO 00 00 O 00 r}< CO T}H 1C O3 Tf< CD CD d M -COCOOO Oi i-l O3 CO -^H> CO 00 Oi 00 O5 O5 O O -i 1 c-' PORTAL Timber- ing (NlMOO M OOCOWO T-I i-l ^-1 1^ rt< Oi OliCO^OOIN t>Ol^t^ CO -H O5 (N >O CO 00 O >O OOCO t^ 09 3 Timber- ing ooooo looiooosoo oco or^rjHco 01 oocoeT}HO lO -^ Tf lO CD *O i i e i SOOO (N iH t^ CO O5 CO * X t~ t-i CO r-i Oi CO --H Tf I-H CO CO Tj< O CO Tf< O COCDOCD O5 CN b- CO (N CD O OO ^ t>- t^ 00 O IN CO O O O >O CO 03t> -^ t^ O O5 0 rHCOOOOirHCO c C: 34 H cc 13 Ij lO (N IO IN lO CN1 O O OOCO CO 00 CO CO , CO 1C -H CO ^H OS >O TJH (N O s o EH 11 d os oo d d * O^TtOrM CD CO lococo^oJ^oo^fo^^S M ""^ ^ N 5 2 2 2 co "* N * ^ '"'Sos lu SCO^tf OS CO T}< COi-H rH CO Tj( 8 ^H(N O 1 * O OSTt< r|< M CO iOOO>OOOOOSOO CO^H 5^3 !r s ^ ^^^QOOOO^S 00 ^^CO ^HCOOS i-H> NCOOO 1* * e* us i-H -* 10 1C i-H COO-* OS O TJ* CO CO t^ O>O(N S5 1 ?2 a OOO O IO 00 N I ^ i OS OS H O IN Tj< CO O Tf^OOOO iO CO I-H (M O 00 O CO C. 1 * 10" t^ EH IM f* co" &f o O co'cfco" i-Tco" co'oT C<) ?H CO CO 3 $ O a _j s r O < tn S 1 ^ 1 03 coco c^ o 00 H oot>o o o o iO OSi-H CO 00 CO rj< T ) l>. OS CO T}* 00 CO CO 00 t~ 03 ^ ^ ^ s% t>lN" T}< C^l CO co _^ > w O OO^OOg 00 CO OOOi-KNt^COiOiMi-HOOcO s 1 iO(N t^ IOIOOS CD s CO g I co"'*" c OQ > 00 ^ ^ : .a "a ' ' ' a *J 00 o 03 fl C3 M M ^2 *.S bi^ai.2^ '"55 1 JS CO g ?-S c3^ * fe 03 1 IHlIiflliil 1 1 H W ^ P t-" h3 H lHSSa>3onilH S -^cqoQfeJi*,Ci!&3 Qs~Bq 426 COAL MINING COSTS where the roof is heavy, and there is comparatively little water encountered. In the north heading, the rock has been shattered by scores of fissures, and in general is so kaolinized by percolating waters that it is for the most part soft, friable and frequently pasty. Considerable water enters the tunnel, and the ground is so heavy it requires timbering, except at one or two places in the tunnel of small extent where hard unaltered rock was penetrated. Long extents of heavy, treacherous ground are continually encountered, so that the timbering must be kept well up to the face. As a result of these widely varying conditions neither the costs, nor the record performances of work at one heading may be fairly taken as a criterion of the work at the other heading of this same tunnel. In spite of this diversity of work, however, the progress made at each heading was surprisingly uniform as shown in the accompanying table, until at the end of April, 1910, 30 months after starting work, the north heading had penetrated 9754 ft. and the south heading 9738 ft., totaling 19,492 ft., or nearly 72 per cent of the entire length. To those interested in the rivalry between the crews of the north and south Elizabeth tunnel headings, the accompanying tables will be replete with interest as well as with valuable data. These give the official costs for both headings during the month of April, 1910, when the south portal advanced the American record to 604 ft., while the north portal progressed 561 ft. in spite of exceptionally heavy ground. It will be noted that, as might be expected, the cost of explosives was the heavier for the south heading, while the cost of timbering- is mainly responsible for increasing the cost per foot of the north heading to $4.36 more than the $25.25 of the south head- ing. It is also significant that the cost of mucking in the two headings is very nearly the same. The appended study covers unit costs of driving a timbered tunnel (North Portal, Tunnel No. 7) in the little Lake sub- division of the Grapevine division of the Los Angeles aque- duct. Ninety feet were driven in 15 eight-hour shifts, the period covered by detailed cost keeping. UNIT COSTS 8js MISCELL t" CO CO I"- Tt< OS O rH 00 O O CO ;ANEC WS &> INSI 1 i o o i fi a i < 1 O J 1 DE ( ?OSTS O O5 ^ I s " rH 1C OS r-i 1C rH 1C O O * s Powder 19.15 Ib. per ft. Feet to be driven, 13,430. Estimated cost, $750,570. Feet so far driven, 11,693. Actual cost so far, $343,114. fe W. C. Aston, Superintendent. 1C (N CO Tf< ^ ?3 t> O rH O O CO CN 8 ss="" CO i 08 .S |s 5 W S "*. I CO (N CO OS 1C CN -COcN 8 CO 1 COCOOSOOCOPOOOS ' 00 *' t>i CO r-I C P CO ^ rH O5 00 O5 (N $232,921.96 Classification 1 Totals ^ '.'.'.:'. : : ::::::: S :::::: fl (fiT) Back trimming Timbering g :::::: | M i I ::::; 1: : : : :| 428 COAL MINING COSTS SOUTH PORTAL ELIZABETH TUNNEL. DETAILS OP ANNUAL SUMMARY OF TUNNEL REPORTS FOR 1909 Totals Units Total Cost Unit Cost Footage to date 7,585 ft. Footage during 1909 Daily footage (365 days) . 4,476 ft. 12 26 Footage of untimbered section Actual cost of untimbered section . . . 3,274 ft. 121 787 45 37 198 Footage of timbered section Actual cost of timbered section 1,202 ft. 51 941 11 43 20 Average cost of timbering per foot progress 1 16 Total bonus footage 1,563 ft. Total bonus pay roll . . 23440 23 Average cost of bonus per foot progress Actual expenditure 173 728 56 .523 Number of shifts worked Number of shifts lost Number of men days. . . 1,055 40 21,452 Men per foot progress Average progress per shift 4.79 4 24 Number of holes drilled 21 066 Total feet drilled . . 135,896 Feet drilled per foot progress 30 36 Total time drilling (in hours) 3,656 25 Average time drilling per foot progress (in minutes) 49 01 Average time drilling one hole 1 foot deep 1 61 Pounds of powder used (including trimming) 143,659 Average pounds per foot. . . 32 09 Number of cars mucked (32 cubic feet). Cars mucked per foot 28,561 6 40 Leyner No. 9, drill repairs (3 machines per shift) 4 116 99 Average cost of repairs per machine per foot . 30 Leyner No. 2 drill sharpener repairs (for 3300 feet) 387 13 117 Energy used (kilowatt hours) 608,762 135.00 MISCELLANEOUS INSIDE COSTS 429 SOUTH PORTAL ELIZABETH TUNNEL. SUMMARY OF TOTAL EXPENDITURE DURING 1909 Totals Unit Cost A. Engineer and superintendent B. Excavation C ]M uc king $3,851 . 24 79,607.15 55,556 . 33 $0.86 17.78 12.41 E Drainage 421.01 0.09 F Ventilation 2,313.44 0.52 G Light and power 24,750.71 5.53 Total cost untimbered $166,499 . 88 $37.19 D Timbering 7,228 68 6.01 Total cost of timbered tunnel $173,728.56 $43.20 SOUTH PORTAL ELIZABETH TUNNEL. DETAILS OF TUNNEL REPORTS, APRIL, 1910 Totals Units Total Cost Unit Cost Total footage to date Footage during April, 1910 9,738 ft. 604ft. Daily footage (30 days) 20 13 604 ft. Actual cost of untimbered section. . . 15,257.61 25.25 Footage of timbered section 000 ft. Bonus footage 364 ft. Bonus pay roll . ... 2,377.48 3.93 78,379 129.76 Cost of energy, at 1.85 c. per kilo- watt hour . . 1,450.02 2.40 *Estimated expenditure 24,238.52 Actual expenditure . 15,257.61 *Amount saved over estimated ex- penditure 8,980.91 Number of underground men days: Foreman and heading crew Timber, pipe, track, car, and machine repair men 1,860 458 3.079 .758 Mechanics, electrician and helpers. 154 .255 430 COAL MINING COSTS SOUTH PORTAL ELIZABETH TUNNEL. DETAILS OF TUNNEL REPORTS, APRIL, 1910 Continued Totals Units Total Cost Unit Cost Number of mule davs 90 81.00 0.134 Number of shifts worked 90 Average progress per shift Number of holes drilled 1,924 6.711 Total feet drilled 16,079 Feet drilled foot progress 26.62 Total time drilling heading (hours). . Average time drilling heading per foot (minutes) ... 324 32.18 Average time drilling one hole one foot (minutes) 1.209 Pounds of powder used (including trimming) 16,100 26.65 Number of cars mucked (32 cu. ft.) heading and ditch 3,216 5.324 Drill repairs (three No 9 Leyners) 248 . 77 Average cost of repairs per foot (three machines) 0.411 Average cost of repairs per foot (one machine) 0.137 Drill sharpener repairs (Leyner No. 2) 61.70 0.102 Drill steel broken 380 .629 Drill steel sharpened Drill steel welded 6,865 445 11.36 .736 Cars repaired 38 .062 Cars repairs 48.52 0.8 Car equipment (changing from 12-inch to 14-inch wheels) 260 68 431 NOTE. New American hard rock tunnel record established April, 1910, 604 ft. MISCELLANEOUS INSIDE COSTS SUMMARY OF GENERAL EXPENDITURES DURING 1909 431 Total Cost Miscellaneous structures $165 79 Tunel construction 15 348 27 Miscellaneous construction equipment 5,168.85 Miscellaneous camp eouipment 19 22 M. & O. live stock 183 75 M. & O. local telephone lines 31 73 M. & O. water supply. 2 34 Division administration ... 740 86 M. & O. roads and trails. 116 05 Total expenditure $21,776 86 Pay roll 9 419 76 Bonus roll 2 377 48 Material issues 6 915 83 Material receipts. . 2.841.54 SUMMARY OF TUNNEL EXPENDITURE DURING APRIL, 1910 Totals Units E. W. O. 33 A. Engineering and superintendent. . . . B. Excavation. $162.89 7 163 30 .27 11 86 C. Mucking E. Drainage. . . . 5,275.78 44 49 8.73 07 F. Ventilation 85 10 14 G. Light and power 2,109 55 3 49 D. Timbering. . . 14,841.11 90 66 24.56 K. Back trimming. 416 50 69 Total expenditure $15348 27 25 25 The tunnel is approximately 10 X 10 ft. in section ; cu.yd. in place per linear foot to pay line ; overbreakage about 17 per cent, making a total of 61/2 cu.yd. of broken material per foot of tunnel. The heading was in 800 ft., lighted by electricity at 110 volts, ventilated by a No. 3 Champion blower through 12-in. pipe, the heading being cleared in 15 min. after shooting. 432 COAL MINING COSTS Drilling is done by one No. 7 Leyner drill, water being forced through hollow steel; drill uses approximately 66 cu.ft. free air per minute at 83 Ib. pressure per square inch, drilling holes to 10 ft. in depth. Mucking is accomplished by use of steel sheets laid down before shooting; No. 3 D-handle, square-point, shovels, and 32-cu.ft. rocker dump cars pulled by a 3y 2 -ton locomotive, running on a 24-in. gauge single track laid with 25-lb. steel. The rock is a close-grained, hard, gray granite with numer- ous seams, causing the drill to run from alignment, but breaks well. The seams and water combined make it necessary to timber all this ground. The ground carries enough water to make disagreeable mucking, and has to be pumped out. Timbers are of 6 X 8 in. Oregon pine spaced 5 to 8 ft. apart, as ground permits, lagged with 2 X 6 in. plank. Sets of timbers consist of two vertical posts and a four-segment arch. The crew consisted of 1 shift boss at $3.50 per day; 4 miners at $3.50 ; 5 muckers at $2.50, and 1 trammer at $2.50. The blacksmith doing repair work was paid $4 per day. The four miners worked on day shift drilling the ground, timbering and shooting, the muckers following on night shift, resulting in a clean heading for the drill crews, and nothing interfering with the mucking crew. In February, 1908, the Chamber of Mines and the Trans- vaal Government jointly offered a prize amounting to 7500 for the best drilling machine that could be produced for mine work on the Witwatersrand. There were 23 entries for this prize, 19 of which started in the competition. In the elimination trials, which were carried out on the sur- face at the Transvaal University College and underground at the Ferreira Deep, Ltd., nine drills were eliminated and 10 entered for the competition of 300 shifts. These were the Hoi- man 2% in., Holman 2% in., Siskol, Climax Imperial, New Cen- tury 00, Konomax, Chersen, Waugh, Murphy, Westfalia. In the surface elimination test the most rapid rate of drill- ing was accomplished by the Westfalia machines; namely, 4.996 in. per minute of actual drilling time. The largest air consumption recorded was that of one of the Konomax ma- chines; namely, 125.9 cu.ft. of free air per foot drilled. MISCELLANEOUS INSIDE COSTS LABOR COSTS 433 Class of Work Total Hours Labor Total Labor Costs Cost per Foot of Tunnel Inside Labor: Squaring heading. Setting up and tearing down machine . Drilling, one No. 7 Leyner; including shift boss' time 23.50 36.00 55.33 $ 9.03 16.59 43.21 $0.100 0.184 0.480 Number of holes drilled 150 Total footage of holes 1,202.30 Feet drilled per hour, includ- ing lost time 15 . 84 Feet drilled per hour, actual drilling time 21 . 74 Average depth of holes, feet . . 8 . 00 Cost per foot of hole, cents ... 3 . 60 Fastest hole 9 ft. 6 in. in 10 min Slowest hole 8 ft. 6 in. in 1 hour 18 min. Average hole 8 ft in 22 min Blowing out holes 5.75 4.15 0.046 Loading and shooting . . . 56.25 22.31 0.248 Mucking 412 cars . . 835 . 00 268.44 2.980 Trimming, stulling, caves, etc Timbering (cost per M. ft. =$11.32).. . Lost time 102.00 107.25 40.75 39.24 40.82 15.66 0.436 0.453 0.174 Bonus, 30 c. per man per foot in excess of 2 3 ft per shift 112.08 1.240 Repairs, to trolley, pump, etc 3.25 1.20 0.013 Totals inside labor 1,265 08 572 73 6.354 Outside Labor: Sharpening steel, with Leyner No. 2 machine 44 00 17 83 0.198 Repairing drill 7 50 2 88 0.032 Framing timbers, at shop, per M ft =2 42 8 75 0.097 Light and power 90.00 33.75 0.375 Totals outside labor 141 50 63 21 702 Auxiliary Labor: Laying track 90 ft 31 50 12 75 141 Drainage line 90 ft 43 50 14 33 160 Ventilation line 72 ft 11 00 3 44 0.048 Trolley line 95 ft 18 00 6 35 0.067 Air line 80 ft ... 2 50 0.80 0.010 Water line 80 ft .... 2 50 79 0.010 Lights line, 90ft 8.00 2.86 0.032 Totals auxiliary labor 117 00 41 32 0.468 Local Administration and Engineering: Proportion of division engineer and assistant's time 50.40 0.560 Total labor costs 1,523 58 $727 66 $8 . 094 434 COAL MINING COSTS COST OF MATERIALS AND SUPPLIES Total Material Costs Construction Materials and Supplies: Drill repairs, 2 side rods, $1.40; 1 chuck, $15; 2 rings, $2.02; 1 oil can, .15; 1 belt, .16; 20 per cent, freight, $3.74 $ 22.47 Cost per foot of hole = .018 Drill supplies, machine oil, .58; drill steel, 45 ins., $4.12; 412 Ib. blacksmith coal, $3.14; 20 per cent freight, $1.57 9.41 Cost per foot of hole = .008. Mucking supplies, car oil, .20; pick handle, .26; hammer handle, .15; 20 per cent freight, .12. ... .73 Power, machine, 2052 kw. h.; blower and lights, 355 kw. h.; locomotive, 1800 kw. h.=4207 kw. hours, at 1.7 c 71 . 52 Explosives, tamping stick, .40; 2700 ft. fuse, $11.54; 306 15 gr. Lion caps, $2.08; 650 Ib. 1| inch, 40-per-cent gelatin powder, $69.88; 250 Ib. 1 in. 40-per-cent gelatin powder, $26.88; 150 Ib. 1 in., 60-per-cent gelatin powder, $20.63; 20 per cent freight ,$26.28 157.69 1050 Ib. powder = 11. 66 Ib. per foot of tunnel. 1050 Ib. powder = 3. 3 Ib. per cubic yard in place. Explosive cost = $.050 per cubic yard in place. Explosive cost = $.27 per cubic yard broken. Timbers, 3597 feet B. M. lumber, $59.24; freight on same $55.01; 775 wedges, $5.43; 50 dowel pins, .42; nails, .73; freight, $1.33 122. 16 Timber per foot of tunnel, B. M.=40. Lighting, candles, $7.15; 14, 16, and 32 candle- power globes, $2.58; 20 per cent freight, $1.94. . 11.67 Totals for construction materials 395 . 65 Auxiliary Material: Trackage, 180 feet, 25 Ib. rail, $15; splices and bolts, .83; spikes, .18; ties, $1.42; 20 per cent freight, $3.68 21 . 11 Drainage 90 ft., 250 ft., wire, $2.13; knobs, .07; 2-inch pipe, $4.36; 20 per cent freight, $1.31 7.87 Ventilation 72 ft., 12-inch pipe, $24.34; 20 per cent freight, $4.87 29.21 Trolley 95 ft., wire, $6.32; lumber, .37; fittings, $1.42; 20 per cent freight, $1.62 9 . 73 Air line 80 ft., pipe, $5.28; fittings, .76; 20 per cent freight, $1.21 7.25 Water line, 80 ft., pipe, $5.28; fittings, .76; 20 per cent freight, $1.21 7 . 25 Light line, 90 ft., wire, $1.48; fittings, .37; 20 per cent freight, 37 _ Total auxiliary materials 84 . 64 Total material costs 480 . 29 Live Stock: Mule 15 days at 90 c $13.50 Total direct and auxiliary field charges $1,523.58 $727.61 $480.29 MISCELLANEOUS INSIDE COSTS RECAPITULATION. COSTS PER FOOT OF TUNNEL 435 Labor, direct charge $ 7 . 056 Material and supplies, direct charge 4 . 399 Local administration and engineering . 560 Stock service 0. 150 12.165 Labor on tracks, etc 468 Material for tracks, etc 1 . 034 1.502 As this work will salvage at about 66 per cent, we deduct . . . 690 Net charge for auxiliary work . 812 Estimated proportion of charge for roads and trails on division . . 1 . 500 Estimated proportion of charge for buildings on division . 200 Estimated proportion of charge for water supply on division .... . 220 Estimated proportion of charge for machinery and tools 1 . 060 Total field charges 15.957 Add 3 per cent to cost for executive office administration . 475 Total cost of tunnel ready for lining $16 . 432 In the underground elimination air trials the quickest drill- ing speeds were attained by the Chersen and Holman 2%-in. machines. The former drilled 1.81 in. per minute over drilling and changing time and 1.56 in. per minute over total time, which consisted of three 8-hr, shifts. The figures recorded for the Holman 2%-in. machine on this trial were 1.94 and 1.47 in. per minute, respectively. As the competition proceeded the following machines with- drew : Climax Imperial, Konomax, Murphy, Waugh, and West- falia. It was the intention to run 300 shifts, but owing to lack of air pressure, 215 shifts were run. The stopes drilled in were 24 to 45 in. wide. The number of feet drilled by each pair of the four leading competing machines were as follows: Holman 2y 8 -in., 12,779 ft. ; Siskol, 14,083 ft. ; Holman 2%-in., 11,744 ft. ; Chersen, 11,781 ft. This was drilled in 215 shifts of 8 hr. each. The drilling speeds attained over total times by the four machines are as follows : 436 COAL MINING COSTS Holman, 2 inches 742 in. per min. Siskol 818 in. per min. Holman, 2f inches 682 in. per min. Chersen 684 in. per min. At the surface elimination trials the average rate of drilling of each pair of machines was as follows : Machine Total Time Inches Per Minute Actual Drilling Time Inches Per Minute Holman 2| in Siskol 1.566 2 058 2.393 4.337 Holman 2f in Chersen 1.988 2.515 3.110 4.110 The cost of drilling per foot with these four leading ma- chines amounted to: Holman 2%-in., 9.77d. ; Siskol, 9.90d. ; Holman 2%-m., 10.91d.; Chersen, 11.94d. These figures com- pare favorably with the cost of hand drilling on the Rand. In the competition, two sets of machines, the Holman 2%-in. and the Siskol have cost approximately 8.3d. per foot drilled plus the cost of steel and drill sharpening, which would come to, at the most, 1.5d. per foot, making 9.8d. per foot. But it has taken 1.282 mine shifts to make one 8-hr, shift, therefore the wage cost must be increased in that ratio. Fur- ther, instead of assuming 10s. per shift for two machines as the white wage, 25s. per shift were taken for four machines. This alteration means an increase of y 2 d. per foot drilled, mak- ing the total cost per foot of these two sets of machines prac- tically 11. 8d. per foot. This is as near as one can get to the actual cost per foot for 28,528 ft. drilled by these machines. This cost of 11. 8d. per foot compares favorably with the cost per foot by native labor, for which Is. Id. per foot would be an exceedingly low figure. The tonnage broken would be as 6 to 5 in favor of the machines. The cost of explosives would be as 6 to 5 in favor of native labor. The tendency is for native wages and cost to increase and this is the heavy item, lOd. per foot, in hammer work, but a much smaller item, only 3d. per foot in machinery, With a proper size of bit a hole 4 ft. 8 in. MISCELLANEOUS INSIDE COSTS 437 or 5 ft. deep is much better as regards tonnage than the ordi- nary hand-drill hole, for it will carry more explosives and so a greater burden. Inches Drilled Per Pound. Weight of Machine Over Drilling Depth of Holes Drilled Percent- Period of 4 Hours age In- Weight crease in Name of Drill of Drill in Pounds Depth Using 60 Pounds Air Pres- sure 50 Air Pres- sure 60 Air Pres- sure 50 Air Pres- sure 60 Pounds Pounds Pounds Pounds Air Per Per Per Per Pressure Square Square Square Square Inch Inch Inch Inch Ft. In. Ft. In. Flottman 52.250 4.33 5.83 18 lOf 25 4f 34.4 Gordon 72 . 625 4.69 6.08 25 41 36 9 29.5 Holrnan 97.500 1.77 2.09 ***** ^*-$ 14 4J 17 Oi 18.6 Kimber 100.000 1.55 2.03 *-4: 12 llf W 4 16 11 30.6 Little Kid 102.500 1.85 2.63 15 9i 22 6} 42.6 Chersen 113.625 2.79 3.50 26 5f 33 2 26.0 Little Wonder.... 118.500 1.68 2.02 ArfVF V 4 16 71 19 llf 19.7 Baby Ingersoll. . . 129.500 2.34 2.74 25 3 29 6} 17.0 Drill steel. There is perhaps no single item in rock drill- ing that will affect the cost of the work so quickly as the effi- ciency of the drill steels. The essential qualities of a drill steel as laid down in a paper presented at the February, 1921, meeting of the Am. Inst. of Min. and Met. Engr's., are: First, it must be easily forged; second, the forged bit end must be such that it can be easily heat treated to obtain hardness to resist chipping; third, the bar or body must be stiff to resist bending or twisting and yet tough to resist shock and vibra- tion, with resulting breakage, and fourth, the forged shank end must be such that it can be easily heat treated to obtain some hardness with great toughness. Drill steel must be properly forged either by hand or machine, and this operation requires pyrometric control. Two 438 COAL MINING COSTS causes of trouble can be eliminated at this point, as there is no doubt "that drill steels as a rule are forged at too high tem- peratures, and the forging operation is continued after the temperature has dropped below the critical. Furthermore, little if any annealing is done on drill steel after forging; hence the forging operation must be conducted with great care. Assuming the forging temperature is correct, the other minor requirements are that the bit and the shank be in align- ment with the body ; that the shank shall be of the proper shape and length and the shank collar or lugs be of the proper diameter and length ; that the hole, if any, be free from obstruc- tion; that the striking end of the shank be flat and square; that the bit be of the proper shape, with the cutting and ream- ing edges formed full and to the required size ; that the ream- ing edges are concentric with the axis of the steel, and that there are no sharp corners at the shoulder where the bit blends into the body. It is obvious that the above requirements can best be obtained day in and day out by means of a drill steel sharpen- ing machine. The efficiency of such a machine will accordingly depend, first, on the initial forging temperature required, for the lower the initial forging temperature the better the steel structure ; second, on the accuracy of the forged bit and shank ; third, on the speed of operation ; fourth, on restriction of the hole in the shank and bit when using hollow drill steel ; fifth, on the number of heats or times required to heat the steel be- fore securing the finished bit or shank, and sixth, on the air consumption of power required. The means of heating for forging will be considered later. The drill steel must have an ideal bit and shank. The essential qualities of such a bit, regardless of the conditions under which it is operated, are that its shape be such that maximum cutting speed can be maintained for as great a dis- tance as possible before wear of the gauge and cutting edge reduces the speed of penetration to a point where a change of steel is made necessary, and that the size or diameter of the drill hole corresponding to the gauge of the bit can be main- tained with the least possible reduction as the depth of the MISCELLANEOUS INSIDE COSTS 439 hole increases, and also that the shape of the bit is such that it can be correctly and readily formed and heat treated. The following features of bit design require attention: Shape, total length, and angle of cutting edge; length and area of the reaming edges or surfaces ; size and shape of clear- ance grooves for ejection of cutting, and lengths and angle of the wings and the manner in which they are blended into the body of the steel. The combined length of the cutting edge and the manner in which it is applied is a big factor in the drilling operation regarding the speed and the life of the drill steel and the drill. The longer the cutting edge the greater the amount of rock cuttings per blow, assuming that the cutting edge is and remains sharp or sharp enough for the conditions. Further- more, the drilled or blunt edge, besides decreasing the penetra- tion per blow, lessens the cushioning effect, and this causes the drill steel to rebound from the rock, which may cause break- age of the drill steel or parts of the drill. In a radial cutting- edge bit the work done is greatest at the extreme cutting and reaming edges, which accounts for the unequal wear along the cutting edge. Therefore it is apparent that the only way to improve this condition is to so shape the cutting edge that the work is evenly distributed throughout its length, and so that the extremities of the cutting edge have suitable reaming surfaces properly tapered back to improve the wearing quali- ties of the gauge. A good deal has been written about the angle and shape of the cutting edge, and many kinds of bits have been brought out, such as the bull bit, rose bit, double cross bit, chisel and double chisel bit, and other types ; but the bit that approaches the ideal design is the double-arc, double-taper bit. The accompanying illustration, Fig. 2 shows the characteristic dimensions of this bit. As to the ideal shank, there is no doubt that the so-called shankless steel approaches .the ideal, and it is regrettable that this type is not suitable for all con- ditions. Its advantages and disadvantages are evident. The third and last requirement of the drill steel which has ideal qualities is that it must be properly heat treated. This without doubt is the most essential operation in securing best results, and yet it is safe to say it receives the least attention. 440 COAL MINING COSTS The good results which are expected from all previous opera- tions can be obtained only through proper heat treatment. Theoretically, this operation of heat treating should be simple; practically it is not, for all too much depends on the equip- ment. It is interesting to note that answers to a recent questionnaire brought out the following facts and are representative of all fields and conditions of mining. To the question, what type of bit was giving the best service, returns showed that the double-arc, double-taper bit and cross bit were far in the lead. This end rrtusfbe ' FIG. 2. Standard design for double-arc, double-taper bit. To the question as to what necessitates the resharpening of the drill steels most frequently, wear of gage was unanimously first choice ; chipping of bit was unanimously, with one excep- tion, second choice; breakage of shank was unanimously, with one exception, third choice; upsetting of bit, upsetting of shank and breakage of body were about on a par for fourth choice. To the question as to what was the direct cause of the necessity of resharpening, poor heat treatment was unani- mously first choice, and faulty steel and severe rock conditions were about on a par for second choice. Explosives. The selection of explosives for tunnel blasting, probably requires a more careful study of conditions than for MISCELLANEOUS INSIDE COSTS 441 any other kind of excavating. Maximum speed and economy in driving cannot be attained unless the explosive best adapted to the work is used. When starting a tunnel or drift, it is a good plan to thoroughly try out several explosives, which are distinctly different in action, before finally adopting any one of them. The results, however, from this preliminary trial will be of little or no value, unless each different explosive is used under exactly the same conditions. Care must be taken to see that no change occurs in the character of the rock, number and direction of the bore holes, strength of the de- tonator, kind and quantity of tamping, amount of water encountered, method of connecting up the bore holes for firing, and that the explosive is always thoroughly thawed. If a material change in any of these conditions occurs as the work progresses, further tests should be made to determine whether a quicker or slower, a stronger or weaker, explosive might not break the ground, or bottom the bore holes better, or make it possible to bring out the cut with fewer holes or deeper ones. The speed at which rock can be drilled does not indicate how it will break, and not infrequently that which can be easily drilled is very difficult to blast. High explosives suitable for tunnel blasting should not give off objectional fumes on detonation, and accordingly gelatin dynamite, blasting gelatin, or ammonia dynamite should always be selected. Gelatin dynamite is made in various grades of strength, from 25 to 80 per cent, inclusive. It is comparatively slow in action, the higher grades being little, if any, quicker than the lower ones. Blasting gelatin is manufactured in only one strength, which for comparative purposes may be said to be 100 per cent. It is more powerful and quicker acting than any other blasting explosive. It should be used sparingly, therefore, until the maximum safe charge has been learned from experience. Good results will often be had in hard ground, if a few cartridges of blasting gelatin are used in the point of the boro hole, with gelatin dynamite on top. When this is done, it is best to put detonator in one of the cartridges of blasting gelatin. Ammonia dynamite is made from 25 per cent to 75 per cent 442 COAL MINING COSTS strength. All grades are quicker than gelatin dynamites, and generally speaking the quickness increases with the strength. That is, the stronger grades are quicker, and the lower grades slower, in action. The various grades of these three high explosives, offer a wide range in strength and quickness to select from, and it is always possible after a few trials to find an explosive exactly suited to the conditions. Railroad tunnels, mine tunnels and drifts, highway tunnels, and irrigation tunnels, are being driven daily through various kinds of "ground." Often it is a matter of first importance to finish them quickly, and consequently details in regard to methods and equipment are matters of general interest. In Engineering Contracting of October 20, 1909, Mr. J. B. Lippincott, assistant chief engineer of the Los Angeles aqueduct, gave an interesting account of the driving of the Red Rock tunnel of the Los Angeles aqueduct system. In August, 1909, this tunnel, which is 9-ft. 10 in. X 10 ft- Sy 2 in. in section, was advanced 1061.6 ft. Mr. Lippincott states that the explosives used were Du Pont 40-per-cent ammonia dynamite and blasting powder. In the Engineering News of November 18, 1909, the Red Rock tunnel is again referred to, and details are also given by Mr. C. H. Richards, division engineer, in regard to a tunnel on the Little Lake Division of the Los Angeles acqueduct. The explosives used in this tunnel were Hercules 40-per-cent and 60- per-cent gelatin dynamite, the average weight of explosives per cubic yard of rock, place measurement, having been only 3.3 lb., or about 35 lb. per linear yard of tunnel almost 10 ft. X 10 ft- in section. A short time before, accounts were given in several engineer- ing magazines, of a record driving speed made in the Roosevelt drainage tunnel, at Cripple Creek, Col. The explosives used in this tunnel were 40-, 50-, and 60-per-cent Repauno gelatin dyna- mite and Du Pont blasting gelatin. A very interesting description of the Rondout pressure tunnel of the Catskill aqueduct, written by John P. Hogan, assistant engineer of the New York City Board of Water Supply, was published in the January 1, 1910, number of the Engineering Record. Very rapid progress was made in this tunnel, and also MISCELLANEOUS INSIDE COSTS 443 in the Moodna pressure tunnel of the same system, described in an article in the Engineering Record of June 4, 1910. The explosive which gave best results, and which was used exclusively in both of these tunnels was 60-per-cent For cite a gelatin dyna- mite. Reference to a paper by B. H. M. Hewett and "W. L. Brown, on the land section of the Pennsylvania Railroad North River tunnels, published in Vol. XXXVI of the Proceedings of the American Society of Civil Engineers, and reprinted in part in Engineering Contracting of May 11, 1910, shows that 40-per- cent. Forcite was used in blasting on the Manhattan section, and 60-per-cent Forcite on the Weehawken section. The records of many other tunnels recently constructed, further illustrate how many kinds and strengths of explosives are used for blasting under the different conditions encoun- tered in one class of work. The specific cases referred to above, were all connected with large and important contracts, where equipment and methods were at the best, and several of these tunnels were driven at record speed. The fact that so many different explosives were used in the several tunnels, goes to show that care was taken to use the explosive which was best adapted to the conditions, and it is not unlikely that the speed of driv- ing these tunnels, was largely due to the attention given to the selection of the explosives. VENTILATING COSTS There are three forms of efficiency that must be considered in all fans, and the value of the fan depends largely upon its ability to give that efficiency, which is most valuable for the particular mine on which the fan is to operate. The mechanical efficiency is to be considered first. Mechan- ical efficiency is the relation the useful work performed by the fan, bears to the power required to propel the fan. The manometric efficiency is of second importance. Manometric efficiency is the relation which the depression caused by the fan, is to the theoretic depression which the fan would make if it were a perfect machine, and working against a closed air- way. The volumetric efficiency is the third to be considered. 444 COAL MINING COSTS This is the relation which the volume of air discharged by the fan, bears to the volume or cubic contents of the fan, taken the number of times of rotation during the given interval of measurement. A fan may be high in one of these efficiencies and low in the others. For instance, a fan may produce a large volume of air at a very low pressure, and be excellent in volumetric efficiency, and yet be low in manometric efficiency and mechani- cal efficiency. In fact, fans high in mechanical efficiency are usually high in manometric efficiency, and almost all fans high in mechanical efficiency, when working under favorable con- ditions, are low in volumetric efficiency. A fan may be high in volumetric efficiency and low in mechanical efficiency, and also low in manometric efficiency; these conditions all depend on the construction of the fan, and the conditions under which it works. The mechanical efficiency of the fan may be determined if the volume of air passing through the mine is known, and the pressure (water-gage) reading is taken at the same time the air reading is taken, together with the indicated power of the motor (either steam engine or electric motor) when resolved into foot-pounds. The manometric efficiency may be determined if the tangen- tial speed of the fan is known, and the actual pressure as read by the water-gage. The volumetric efficiency may be determined if the rotations of the fan are known, and the vol- ume of air discharged is known for the same interval. A fan may have high manometric efficiency, and not be useful for mine ventilation; for instance, a cupola fan can produce the required pressure for a mine and not give any volume worth consideration. A fan may have a large volumetric efficiency under favorable conditions, and yet under unfavorable con- ditions be unable to produce pressure, and hence be a poor fan for mine ventilation. Testing a fan is a proceeding requiring a considerable degree of skill and care. It is customary in making an ane- mometer test to organize the force so that each main airway shall have two men allotted, one to take the anemometer read- ings, and one to hold the watch and light and call the time. The time of readings should be one minute, two minutes, three minutes, or any number of minutes that may be agreed upon. MISCELLANEOUS INSIDE COSTS 445 It is difficult, however, to hold the anemometer in a swift air current longer than three minutes, and for this reason this is usually the time limit. The readings should be taken about 100 ft. away from the fan in a drift, or about 100 ft. from the bottom of the shaft in a shaft mine, and should be taken at a place where the section of the airway is nearly uniform and as smooth as possible, and not close to any turns, where the air may be deflected into eddies. It is well to adopt a schedule of readings at intervals of fifteen or twenty minutes so that the velocity of air, the water- gage, the speed of motor, and temperature and barometer may be taken at exactly the same time as the motive power applied is indicated. With large fans it is customary to use steam engines, and the indicator is used to ascertain the motive power applied. In the usual tests at coal mines, the barometer and temperature are neglected, as not having sufficient bear- ing to warrant the calculations necessary to apply them to the test. It is essential that a certain time be set for each and every reading, that they may be all taken simultaneously, and at least three readings should be taken at varying speeds of the fan. It is preferable to test under such speeds as may be employed under the conditions which the mine may require. To get proper readings of the water-gage, it should be so placed that it will record the pressure of the air immediately in front of the fan, and in the full current where there is no possibility of an eddy in the current. To obtain the proper position it is necessary to pipe from the water-gage in the engine room to a place in front of the fan where the air current is in one body. The end of this pipe should be from 10 to 25 ft. from the periphery of the fan and should be bent so the air current will flow directly into it. If it is hung down a shaft, the pipe should be curved returning upward, and have a small hole in the lower side of the curve to drain the moisture that would collect and form a water seal, if there were no drainage. This water seal will destroy the true reading if the pipe is not drained. The indicator should be operated by a competent engineer who understands the theory of steam consumption of a steam engine as well as the mere knowledge of the operating of a steam-engine indicator, as there may be defects in the valve 446 COAL MINING COSTS gear that the indicator will show to the practiced eye, which might be overlooked by the novice. It should be borne in mind that the engine or motor is having its efficiency tested along with the fan, and any loss of efficiency in the engine or motor will reflect on the fan. In taking the readings it is well to begin with the even hour, and allow five minutes to each party detailed to take the various readings to measure the air velocity, the water-gage and to take the indicator dia- grams. The fan should be maintained at a constant speed during this interval. At the end of the first five minutes, the fan should be accelerated to a desired speed during the succeed- ing ten minutes, and beginning with the even quarter hour, be maintained at a constant speed during the succeeding five minutes, so the various readings may be taken at the desired speed of the fan. At the end of this five minutes, the fan may be accelerated again, and other readings made at intervals of one quarter hour, and as long as may be desired. The following table gives a test taken by the West- moreland Coal Company, of Irwin, Penn., which is a fair sample of the manner of taking the various readings on a 20-ft. cen- trifugal fan driven by a steam engine: FAN TEST AT WESTMORELAND SHAFT, IRWIN, PA. SPLIT No. 1 SPLIT No. 2 SPLIT No. 3 Time. P.M. per Volume W. G. H.P. in Air H.P. Ind. Mech. Eff. Area Vel. Area Vel. Area Vel. 4 : 00 100 50 1428 58 1593 44 500 195,134 2.40 73.8 114 65 4 : 15 132 50 2245 58 2200 44 740 286,791 4.20 190 265 71.7 4 : 30 150 50 2853 58 2383 44 822 335,914 5.10 272 344 79.0 4 : 45 150 50 2800 58 2463 44 833 337,171 5.20 276 360 77.7 The table shows the general method of a test for practical purposes, where boiler test is omitted, and also the variation of the barometer. For all practical purposes, this test is suffi- cient, and shows the mine engineer where he can practice economy on his ventilation, through the mechanical efficiency of his fan. Power required. The power required to drive air through a given mine increases as the cube of the volume, provided no MISCELLANEOUS INSIDE COSTS 447 change is made in the air courses: Let V represent volume of air, in cubic feet of air per minute ; let P represent pressure of air as represented by inches in water-gage; Let W represent pressure in foot-pound performed on the air ; then W = V X I" X 5.2, as 5.21 Ib. is the pressure of air per sq.ft. as represented by 1 in. of water-gage. When reduced to horsepower, the formula then becomes FXPX5.2 h.p. 33,000 To illustrate this formula, suppose a mine passing 100,000 cu.ft. of air per min. against a 2.3-in. water-gage. The work performed in passing the quantity of air at the given pressure, will be in terms of 100,000X2.3X5.2 hp.= 33,000 which is equal to 36.24 hp. Now the pressure of air increases as the square of the volume, so that if we desire to increase the volume of air from 100,000 cu.ft. of air per min. in the above instance to 200,000 cu.ft. of air per min., we will re- quire eight times as much power as to produce the 100,000 cu.ft. of air per min., as in the first case, for having doubled the velocity, the water-gage will have increased from 2.3 in. to four times 2.3, or 9.2 in. Now substituting in our second formula, we have , _ 200,000X9.2X5.2 P ' 33,000 which equals 289.92 hp., which is eight times 36.24 hp., the amount of power required in the first example. It will be seen that to get the increased 100,000 cu.ft. of air, it will require 252.68 hp., whereas in the first example only 36.24 hp. was required, so that the second 100,000 cu.ft. of air required seven times as much power to pass it through the mine as the first 100,000 cu.ft. When we consider that the installation of a ventilator, capable of producing 289 hp. in the air, would cost eight times as much as a ventilator producing 36 hp. in the air, we can understand why the management hesitates to purchase such 448 COAL MINING COSTS expensive apparatus. Furthermore, if a horse-power in the most favored coal regions at the mine costs about $50 per year, then to increase the ventilation as above cited from 100,- 000 cu.ft. of air per min. to 200,000 cu.ft. of air per min., would cost for power alone, about $12,634 per year. This annual cost at many mines would be a serious consideration. What, then, would be required to get the necessary amount of air without making such an outlay of money in running expense? In most mines the quantity of air could be doubled without in- creasing the power eight times, by cleaning up, enlarging and increasing the number of air courses, thereby reducing the velocity of air and consequently the pressure. The pres- sure is the great absorber of power, and it is well to bear in mind that for a given amount of air, a certain pressure will be necessary to propel the air through the mine, and regardless of what form of ventilator is used, this pressure will be the same. It is a common but mistaken belief among mining men, that one favored form of ventilator will propel the current through a mine at a less pressure than another ventilator. This fallacy should not be entertained by any mine manager. Specification of fans. The engineer is sometimes called upon to prepare specifications for a fan to be used at a par- ticular mine and to obtain bids for the erection of same. Cer- tain requirements will be laid down,, say that the fan will furnish 250,000 cu.ft. per minute against a 6-in. water gage. A manufacturer in presenting his bid will sometimes state that he can do better than what the specifications ask, often claim- ing that the dimensions given are too great and the fan un- necessarily large. Where the manufacturer is left to make his own estimate of the size of fan required, he will often make no inquiries in reference to the size of the mine airways or the present circulation in the mine ; but will take chances in refer- ence to these important data. In nine cases out of ten the new fan is installed at a new mine, where, owing to the short airways of ample size, it continues to do good work for a few years. In the later development of the mine, however, there is experienced a scarcity of air, and, referring to the guaranteed capacity of the fan, the superintendent orders the mine fore- man to speed it up, claiming that it is not run at its full MISCELLANEOUS INSIDE COSTS 449 capacity. The results are still far from satisfactory, but the fault is not with the fan, which is now forced to operate under conditions for which it was not designed. A short familiar calculation of the pressure required to pass a given quantity of air through an airway of given dimensions, using the Fairley coefficient (k = 0.00000001) , shows that a water gage of 1.9 in. is required to pass 100,000 cu.ft. of air per minute through an airway 7 X 10 ft. in section, for each 1000 ft. in length. Suppose the main airway in the mine is 3000 ft. long to the first point of split. The distance the main air cur- rent must traverse, including the return, is then 6000 ft., and the water gage required, for the passage of the main airways only, is 6 X 1-9 H-4 in. It is clear at once to any mining engineer or foreman that it would be impracticable to attempt to pass 100,000 cu.ft. of air through a single airway of this size for a distance of 6000 ft., including the return. By pro- viding two air-ways of this size for the intake and two air- ways for the return, both the water gage and the power on the air would be reduced to one-fourth of the previous amount. By this means, a main air current of 100,000 cu.ft. per min. can be conducted a distance of 3000 ft. to the first point of split and returned to the upcast, under a water gage of 2.85 in., requiring 45 hp. This is the horsepower on the air con- sumed in the main airway. These are the more important calculations in estimating the requirements of the mine in respect to ventilation. To these amounts, however, must be added the water gage and power consumed in the splits. Finally, to determine the power of the engine driving the ventilating fan, it is customary to assume an efficiency of 60 per cent. For example, if the entire horsepower of the air, for the main airways and splits, is, say 60 hp., the required horsepower of the engine would be 60 -r- 0.60 = 100 hp. A similar calculation shows that a water gage of 5.08 in. and 160 hp. on the air are required to pass 200,000 cu.ft. per min. in three airways, a distance of 6000 ft. ; but, using four air- ways, the same air volume, circulated the same distance, will only require a water gage of 2.85 in. and 90 hp. on the air. Assuming a consumption of 5 Ib. of coal per hp.-hr., the saving of 160 90 = 70 hp., by employing four instead of three main airways, would correspond to a saving in fuel of 450 COAL MINING COSTS 5 X 70 = 350 Ib. of coal per hour, or 1% tons in 10 hr., whicl) figures are a fair approximation to fact, in actual practice. Assuming, then a saving of 3 tons of coal per day of 24 hr. (less fuel being consumed during the night, when the mine is not working), at a cost of $1 per ton for coal at the mines, the total saving per month would be practically 3 X 30 = 90 tons or $90 per month, at the least computation. Experience suggests that economical ventilation requires that the water gage be kept down to at least 1 in. per 100,000 cu.ft. of air in circulation. As has often been suggested before, this can be done by properly splitting the air current in the mines. Experience indicates further, that for the circulation of from 80,000 to 100,000 cu.ft. of air per minute, two main intakes and two return airways should be provided; or three main airways for a circulation up to 150,000 cu.ft. per min. ; and four main airways for a circulation up to 200,000 cu.ft. per min., the sectional area being 70 sq.ft. If this plan is followed, a great saving in coal will be found to result. It is important, in overcasting the air, to see that all air bridges on main airways are substantially built so as to prevent the leakage of air. Such overcasts and all stoppings on main air- ways should be built of concrete, for the same reason. It has been the policy of the Consolidation Coal Co., when opening a new mine, to make the projections on the maps of that mine, with enough intakes, returns and overcasts, to ventilate the mine under a low water gage. Where old mines were contending with a high water gage, air shafts were sunk, in order to shorten the airways and lower the water gage. At Mine No. 26, an air shaft was sunk 11,700 ft. from the mouth of the mine, and a new 6 X 20-ft. fan and a boiler planl installed. By installing the fan at this last location, it shortened the airway one-half, besides permitting the two return airways to be used as outlets or intakes, as required, thus providing four inlets in place of two, as before. Previous to the change, this mine had but two inlets and two returns. By moving the fan, the water gage was reduced from 3 in. for 100,000 cu.ft.. to 1.36 in. for the same air volume, a splendid demonstration of the practical value of shortening airways. It will be noticed that 1.36 in. water gage is too high for a circulation of 100,000 cu.ft. of air, as, in case 200,000 cu.ft. were required the neces- MISCELLANEOUS INSIDE COSTS 451 sary gage would be 5.44 in., which is far too high for every- day practice. It may be possible that this is the best that can be done with an old mine sometimes. In estimating the saving of coal effected by moving this fan, the calculation is based on the difference in actual total horsepower required of each engine, which is 123 hp. for 150,- 000 cu.ft. of air. Five pounds of coal per hp. per hr. then means 615 Ib. of coal per hr., or 14,760 lb., or, say 7.4 net tons per day of 24 hr. Coal at $1 per ton then means $7.40 per day. or $222 per month and $2664 per year. The water gage on the new fan will be 3.06 in. for 150,000 cu.ft. of air ; while on the old fan, it would require a 6.75 in. water gage for the same air volume. The Consolidation Coal Co. has adopted a different method for the ventilation of its new coal field in the Elkhorn Division, Kentucky. The mines are all projected on the map of the coal field, showing the mines about as they will be, and giving the amount of coal they expect each mine to produce and the amount of air they expect to put into each mine. Then the proper number of airways are projected on the map, together with all overcasts, so as to handle this amount of air with a given water gage. The large mines from which it is intended to produce not less than 2000 tons of coal per day, are expected to require about 200,000 cu.ft. of air per minute when the mine becomes fully developed. In these large mines, there will be four return airways and four intakes, each of which will average 6^2 ft. high by 10 ft. wide, making an area of 65 sq.ft. each. From the fan to the first split is generally about 1200 ft., and here two double- face headings are turned off two to the right at about 1200 ft. and two to the left at about 1320 ft. so the switches will not interfere, as the face headings are turned with a 112-ft. radius curve. Each of these two face headings will have about four room headings, making eight room headings for each 25,000 cu.ft. of air. The four room headings will be ventilated with one split of air, say about 12,500 cu.ft. to each four room headings, which will be about 1000 ft. long. Face headings will be turned every 2100 ft., center to center, and room head ings are then to be turned both ways from the face heading. The water gage that will be required in passing 200,OOG 452 COAL MINING COSTS GU.it. of air to the first, second, third and fourth splits, there being four return airways and four intake airways, is esti- mated as follows: The distance to the first split of air is, say 1260 ft., an average between the two face headings where 50,000 cu.ft. of air goes to the four face headings. Using Fairley's coefficient of friction (k = 0.00000001), the water gage for 200,000 cu.ft. of air up to this first split will be 0.00000001 X 1260X2(6.5+10) X200,000 2 n _ . W ' g= 5.2(6.5X10)3 =(X72m - To the second split is a distance of 2100 ft., the air volume being 150,000 cu.ft., the water gage for this section is found in the same manner, and is 0.68 in. Likewise, to pass 100,000 cu. ft. of air to the third split, a distance of 2100 ft. again, will require a water gage of 0.30 in., and finally, to pass the remaining 50,000 cu. ft. the same distance to the last face headings will require a gage of 0.075 in. The sum of these four gages must then be doubled to provide for the return, in each case, which gives for the total water gage absorbed in the main headings 1.775 X 2 = 3.55 in. It is well to estimate on a loss of at least 10 per cent of the air, which means a loss in water gage of 19 per cent, which taken from 3.55 in. leaves only 2.87 in. water gage for 200,000 cu.ft. of air on the whole mine, to this point. The splits in the face headings show a water gage of only 0.075 in. for 2100 ft. with 12,500 cu.ft. of air. All regulators on face headings should be placed near the main airway and must not be set beyond the first room head- ing, under any consideration. In reference to the percentage of water gage spoken of as being lost (19 per cent), it has been found in practice this loss reaches as high as 25 per cent, so that it is safe to call it 19 per cent, as 10 per cent loss in air means 19 per cent loss in gage. This is not all due to loss of air by leakage, but from air passing crosscuts and other wide places, which reduce the velocity. Most of the old style paddle-wheel and screw-propeller fans are working with less than 20 per cent mechanical efficiency, while the latest high-grade speed fans are working above 60 per cent mechanical efficiency, and in the great majority of cases, are working with a mechanical efficiency of between 70 MISCELLANEOUS INSIDE COSTS 453 and 80 per cent. Assuming that the old-style fan is working at 20 per cent mechanical efficiency, and the new high-speed fan at 60 per cent efficiency, also that 100,000 cu.ft. of air per min. at a 2-in. water-gage is required, the horse-power in air 100,000X2X5.2 33,000 = 31 - 51 ^' Now 31.51 hp. is 20 per cent of 157 hp. ; 31.51 hp. is 60 per cent of 52.51 hp. It will be seen that 157.55 hp. will be required to drive the old-style fan to produce the required ventilation, while only 52.51 hp. will be required to drive the new high- speed fan. Thus the new high-speed fan will save the dif- ference between 157.55 hp. and 52.51 hp., or 105.04 hp. Now, as one horse-power is produced at a cost of $50 per year, it follows that the saving effected with the new high-speed fan is 105.04 X 50 = $5252 per year. This amount of money will install a fan to produce the required 100,000 cu.ft. of air per min. at a 2-in. water-gage, under the worst conditions likely to be found, and would therefore return to the owner the entire value of the fan each year. Change in volume of air required. Ventilating conditions at all mines change from week to week. As the workings extend underground, the amount of air required is increased, the resistance runs up, the number of stoppings multiply, the current of air may be lengthened or shortened. All of these things take place, resulting in a change in the amount of work required of the fan. If, by making repairs, extensions and changes, costing, for example, $100, the fan is relieved of 10 per cent of its work, we have something to show on both sides of the account, and can decide whether or not the money has been well spent, if we know the amount of power saved and its cost per unit. There are few cases where the cost of power per horse- power-hour is known in installations where a steam engine is used for driving the fan. The practice usually is to operate at a certain number of revolutions per minute, this speed being maintained by opening and closing the throttle valve. This critical speed is governed by the needs of the mine. If there is "bad air" on "Third Right," the foreman may " speed 'er up" a few revolutions. Or, again, the drivers are "not 454 COAL MINING COSTS able to keep a light," in another part of the mine. In this case the fan may be ' ' slowed down a little. ' ' Under such conditions, and they exist more generally than one would imagine, the cost of power used cannot always be figured accurately. In many instances it is not known at all. The mine, however, may be amply ventilated at some one of the different speeds at which the fan is run. This particular speed, whatever it is, is the proper one at which the fan should be driven during the time when the underground operations are in full activity. Where a mine is equipped with a fan of ample capacity, and has airways as well as underground structures in fair con- dition, a certain number of revolutions of the fan will furnish a sufficient amount of air to comply with the mining law and to fully ventilate the mine when it is producing its maximum tonnage. Such ventilation will require the maximum horse- power applied to the shaft of the fan. During the hours when the mine is not producing coal, say at night, on Sundays and on holidays, only a fraction of this ventilation is needed. In mines where no inflammable gas is found, tests have shown that at night and on idle days, about half the full ventilating current is sufficient for all needs and all the demands for safety. Under conditions such as we have outlined, a steam-driven fan can be slowed down to, say, half the speed used in tKe daytime, and the volume of air will be reduced in almost the same proportion. The steam, however, will vary in pressiire, and this, too, will further vary the speed of the fan as well as the volume of air passing. In nongaseous mines and others where only half the normal ventilation is required when the mine is idle, it is an inexcus- able waste to use more. The responsible officials at the mine should decide what amount of air is necessary when the mine is not working, making due allowance for any men who may be required in the mine at such times and this should be rigidly adhered to. If a speed-regulating device were supplied as in the case of a steam engine, for example, the control of the power con- sumption, and therefore to a certain degree the cost of ventila- tion, would be in the hands of a comparatively irresponsible MISCELLANEOUS INSIDE COSTS 455 attendant who cannot be depended on to do one thing the same way every time. A motor capable of running at full speed, whatever that may be, and at half speed fills the require- ments of fan drive fully. Once it is installed it assumes all responsibility as to the speed of the ventilating apparatus. It operates at either one or the other of its two possible speeds, and the fan, as a result, is either handling its full prescribed capacity or approximately half of it. Occasions arise which demand that the maximum rotation of the fan be changed, either increased or decreased. Such changes can be well and cheaply made by substituting a dif- ferent diameter of pulley on the motor shaft, when a belted installation is being considered. The power saved by making the high-speed rotation exactly what is needed to fully ventilate the mine, and no more, when underground conditions are reasonably good, soon amounts to enough to pay for an extra or different driving pulley, or even several of them. Money can frequently be saved and better mine ventilation assured by the use of canvas piping. Wherever blind entries, such as water courses, air courses, tunnels and shafts are contemplated or wherever gases will not clear, this system can be used to an advantage. Every mine should carry at least one of these outfits on hand all the time. Briefly, the system consists of a flexible, pliable, treated canvas tubing and a fan directly mounted on the armature shaft of a motor. Protection for the tubing is afforded by suspending it from a wire attached to pegs driven in the roof at 15-ft. intervals, or to a wire fastened to upright posts. Addi- tional sections are coupled within 15 seconds with a special coupling furnished as an integral part of each section. The outfit provides a compact, portable, practical, "fool- proof" blowing unit, to be used as an auxiliary to the large mine fan. In nearly all mines, places occur where the air has been short-circuited to such an extent it does not provide suffi- cient ventilation to carry off gases and smoke in blind head- ings. By installing a booster outfit at this point, the air can be carried for distances up to 1000 ft. in blind entries in suffi- cient volume to work up to eight men. This added circulation 456 COAL MINING COSTS of air strengthens the main air system to permit work of any class without the slightest handicap or loss of time. It is often possible to drive headings 25 per cent faster by eliminating cross-cuts, and save $50 to $100 through the elimi- nation of each stopping. These units are capable of giving volumes of air varying from 1900 to 1200 cu.ft. per minute for distances from 100 to 1000 ft. and are especially designed for driving single entries in coal mines. Spiral riveted galvanized pipe is also often used in ventilat- ing tunnels and gangways in which there are no return air- ways. A system in general use is to place an electric booster fan to drive air through an 18-in. pipe to the face of the tunnel or gangway, and allow the air to return out the tunnel or gangway, until a hole is driven to the surface or the upper level, when the fan is moved nearer the face and pipe used again to advance the tunnel or gangway. Gangways and tunnels four miles long have been driven by this method ; they are usually ventilated a distance of 2000 ft., when fan must be advanced. Mine lighting. To measure accurately the illuminating power of a lamp, we must consider not only the intensity of light (or candlepower), but also the solid angle over which the intensity is maintained. A lamp which gives an intensity of light of one candlepower all around gives twice as much light as one which gives a light of equal intensity half way round it. The term "flux" is used by illuminating engineers to designate the product of intensity and the angle over which it is exhibited, since this product most represents the light which flows from the lamp. The unit of flux is called a lumen and is about 8 / 100 of the total flux of light produced by a source of one spherical candlepower. The term candlepower used without qualification is not only confusing but really meaningless. If all sources of light distributed light equally in all directions, then a single measure- ment of their candlepower would suffice to compare them. Practically, however, sources of light differ a great deal in the way they distribute light, and this is especially true if re- flectors are used. Therefore, if a lamp is stated to give two candlepower, the statement should also explain whether " head-on " candlepower MISCELLANEOUS INSIDE COSTS 457 is meant, or average candlepower over the stream of light, or average candlepower in a given plane such as, for instance, the horizontal. A lamp that uses a reflector may have a * * head- on" candlepower 3 to 10 times the average candlepower over its entire stream of light. Generally it is best to state the average candlepower of a lamp instead of the candlepower at a single point or group of points. A statement of the candlepower of a lamp does not suffi- ciently define its light-giving capacity. A 100-cp. lamp is seemingly 33 times as desirable as a 3-cp. lamp and yet a 100-cp. lamp shining through a hole % in. in diameter gives less actual light and much less useful light than a 3-cp. lamp shining through a hole 3 in. in diameter. Therefore, in order to define properly the light-giving capacity of a lamp, a state- ment must be made regarding both the candlepower and the total flux of light (or lumens) produced by the lamp. The selection of proper lower limits for intensity of light and its flux is, aside from safety, the most important con- sideration in selecting portable electric lamps. Without these standards of reference accurate and intelligent comparison of lamps is not possible. In an attempt to establish such lower limits the Bureau of Mines searched for some time for stand- ards which should be fair, not too low in value, not arbitrarily selected, and which should bear an easily recognized relation to something already in use. It was finally decided to prepare a standard Wolf safety lamp to give its best performance and, after adjusting the flame height to 1 in., measure the average intensity of the stream of light and also the total flux of light in the stream. This was accordingly done at two different times, using dif- ferent lamps, prepared by different men, and tested with dif- ferent instruments of different types. The first measurements were made by Dr. L. 0. Grondahl, of the Carnegie Institute of Technology, and the second measurements at the Bureau of Mines. The results of the two tests checked within a very few per cent. The lamp used was a Wolf miner's safety lamp, 1907 model, round burner, burning 70-72 deg. naphtha, and prepared and trimmed in accordance with the standard practice of the Bureau of Mines. The average intensity of light stream, as determined 458 COAL MINING COSTS by these tests, was a trifle under 0.4 cp. and the total flux of light was found to be not quite 3 lumens. The bureau therefore concluded that a satisfactory lower limit of flux of light for hand lamps would be 3.0 lumens and a satisfactory lower limit of average intensity would be 0.4 candlepower. The bureau suggests that lamps designed to be worn upon the cap should give the same intensity of light as that required for hand lamps, but that the minimum flux of light required from cap lamps should be not more than half the minimum demanded from hand lamps, because when a lamp is worn upon the head any light that is thrown to the rear is wasted. If the equivalent of a safety lamp were mounted upon a man's head, one-half of its light would fall behind the man and thus could not be used. Therefore the bureau concluded that 1.5 lumens would be a satisfactory lower limit for the flux of light produced by a cap lamp. Twelve hours was selected by the bureau as a reasonable time of burning. This length of time was chosen after con- sultation with several people outside of the bureau, who were competent to express an opinion in regard to the subject. The folowing table prepared by E. M. Chance, gives com- parative candlepower of various types of lamps. These data were accumulated during about eight years and are general averages. The photometric determinations were made upon a United Gas Improvement Co. 60-in. bar photometer. The photometric standards used were 10-volt tungsten lamps, pre- CANDLEPOWER OF VARIOUS TYPES OF PORTABLE MINERS' LAMPS Candle- Cost per power Shift, Cents Miners' open oil cap lamp 1 . 50 2.4 Miners' open acetylene cap lamp 5.00 1.5 Electric cap lamp 1 . 10 Davy safety lamp . 12 Clanny safety lamp . 35 Wolf -type safety lamp . 65 Akroyd and Best safety lamp 1 . 10 T. M. Chance acetylene safety lamp 3.80 NOTE. The above candlepowers are in no sense maximum, but are the average values over the field illuminated by the lamp in question and have been obtained from many determinations. These are the value that may be expected to be realized in practice under working conditiona MISCELLANEOUS INSIDE COSTS 459 pared and calibrated by the National Lamp Works, and stand- ard sperm candles. No estimate of the cost per day of electric cap or flame safety lamps is given in the table. The labor charge on both the electric cap and flame safety lamps is so large and varies so much with the size of the installation that such figures as could be given would have but little meaning. Portable electric lamps. The qualifications of portable electric lamps can be grouped under three main heads as fol- lows : Weight, cost and capacity. The weight of a lamp can be easily ascertained and each prospective user of a lamp must decide for himself whether or not its weight is excessive. Under the head of cost would be included the first cost of the equipment, as well as all proper charges for operating and maintaining the lamp. Some of these charges will vary with each installation and whether or not the cost is excessive will depend somewhat upon the conditions which surround each case. The capacity of the lamp is taken to mean its ability to produce a certain amount of light for a definite number of hours per day, every day in the year if need be. A lamp that can do this with the fewest interruptions has the greatest capacity for performing the duty for which it is intended. The capacity of a lamp as thus defined takes into consideration not only the ampere-hour capacity of the battery and the efficiency of the lamp bulb, but also the life of battery plates, the mechanical strength of parts and the resistance to wear and tear. We need to define (1) what is the proper amount of light for a lamp to give; (2) the proper time it should burn each day, and (3) what are reasonable interruptions of service and how often they may occur. Proper care of the lamps has considerable effect on the reliability of the service. One of the large German mines, having several thousand electric lamps in daily use, reports that at first about 5 per cent of all lamps taken into the mine at the beginning of the shift were returned at the end of the same shift, either burning poorly or not at all. By a careful study of all details in the lamp house and by putting a skilled man in charge of the lamp-house work, this percentage was 460 COAL MINING COSTS reduced to less than 1.5, with the expectation that it would eventually drop below 1 per cent. While the first cost of electric lamps is undoubtedly higher than that of those burning benzine, the cost of operation, in- cluding maintenance, is claimed to average from 10 to 15 per cent less. The cost of the electrical energy is small and the cost of maintenance consists about one-third of labor and two- thirds of renewal of parts, and depreciation. Of especial importance is the cost of renewing the electrodes of the storage batteries, replacement of complete lamps, which are broken on account of rough handling and accidents, and renewing the incandescent bulbs. The life of the electrodes for lead cells ranges from about 100 to 400 shifts, depending entirely upon the treatment which they receive. The replacement of complete lamps which are broken on account of rough handling and accidents undoubtedly varies more or less in accordance with the character of the work per- formed in the mine. European practice shows that about 0.1 per cent of all lamps per shift are lost in this manner. The incandescent-lamp renewal already has been expressed in figures, in connection with the reliability of service. Excel- lent results have been obtained, the average life of the lamps being approximately 1000 hours. The Manlite lamp complete, including the battery, cable and head piece, weigh 3y 2 Ik- ; f this the head piece weighs 4 oz. The light of this lamp is brilliant, but soft, and at the same time absolutely steady and unflickering for at least 12 hr. per single charge of battery. The curves of discharge of the latter shown in Fig. 3, show how it improves in service. On the first discharge the voltage falls to 1.8 in a little more than 8 hr. The second discharge reaches the same point in a little over half as long again. The progress is thereafter not so rapid, but by the twelfth discharge the voltage is main- tained at over 1.8 for more than 153/2 hr. By the thirty-seventh discharge that period is extended over the sixteenth hour. But these curves are based on a discharge of 1 amp., whereas the lamps only use 0.78 to 0.83 amp., and consequently the voltage is maintained for much longer periods than those men- tioned. The curve of charge with 2 amp. shows a slow increase MISCELLANEOUS INSIDE COSTS 461 in voltage till the fifth or sixth hour, when the rise becomes quite rapid till the seventh hour. The batteries will supply light after 10 charges and dis- charges for 20 hr. at a stretch. This has been shown by tests duly authenticated by mining companies which have used the outfits. The bulb employed, and approved by the Bureau of Mines, has a guaranteed life averaging 600 burning hours. It is held in the burning position by a perfectly ground crystal lens carried securely in place by the lens holder, the whole being sealed so that the miners cannot tamper with it. Accurate figures of the actual cost of upkeep of the lamp in continuous practical service for periods ranging from six months to over a year at various large mines have been col- 18 I I 3 4 5 6 Time in Hours FIG. 3. Curves showing how battery maintains its voltage for a lengthened period after repeated use. lected which show that the cost of material per lamp-shift does not exceed li/2 c - In a lamphouse operating 320 of these lamps during five months no battery repairs or renewals were required aside from a small quantity of electrolyte. The cost of other mechanical repair parts for the 320 lamps amounted to only $15.25. The equipment is sold at a reasonable figure, and the cost of renewal and repair parts is equally low in price; for in- stance, the cost of a positive plate approximates 50c., while a set of negative plates costs $1. At the Merchants Coal Co., Orenda No. 2 mine, Boswell, Penn., 250 Edison lamps were installed on February 15, 1916, at a total cost of approximately $3200. These lamps were in continuous service for about 5 months during which time they were used the equivalent of 30,450 lamp-shifts. During this time the total cost for bulbs, cords, lenses and all other repair parts was $126.49, or a trifle less than 15c. per lamp 462 COAL MINING COSTS per month. On the basis of 30,450 lamp-shifts, the cost per lamp-shift was 0.4c. Add to this the cost of lamp tender, current, approximate depreciation and interest on original investment, and the total is 2c. per lamp per shift. This installation shows the approximate range of cost, fixed charges, depreciation, interest on original investment, service and current. The figures given herewith were taken from the invoices covering material shipped to the mine during the period men- tioned, and no account has been taken of the material on hand on June 1, 1916, in either case; so that there is a satisfactory margin, and the costs shown are slightly in excess of the actual replacements during the period outlined. No expense has been incurred for repairs or replacements of the batteries them- selves. A 12-months test of electric lamps was made at the Vulcan mine, Newcastle, Colo., of the Rocky Mountain Fuel Co., about 1915. It is important to note that the installation was not large, and so the figures well represent what might be readily duplicated in a similar small station. Moreover it may be noted that the lamps were tended by the regular lampman of the mining company and not by an experienced and specially trained man. The Vulcan mine dips at an angle of over 45 deg., and the lamps are operated under the most unfavorable conditions, being subjected to the roughest usage. On December 23, 1914, 27 lamps were put into service, and monthly statements were made out, showing not only the number of delivered lamp shifts and burning hours of the different lamps, but every repair necessary and every spare part consumed. The total number of delivered lamp-shifts during the 12 months was 6350, making a total of burning hours of 61,700. The number of lamp-months was 324, and the cost of upkeep excluding labor was, as shown, $65.52 net or 20.22c. per lamp per month. This figure included all material needed to keep the lamps in good working condition. Even lye solution, vaseline and acid received due consideration. The average number of lamp-shifts was 19.6 per month. Thus the upkeep per shift per lamp was only 1.03c. Including the 27 bulbs furnished with the lamps, 113 bulbs were rendered MISCELLANEOUS INSIDE COSTS 463 valueless during the year. Of these 34 were destroyed by the miners. These may be figured as being half-consumed before they were destroyed, and the same assumption is made about the 27 bulbs in use at the end of the period. COST OF MATERIAL AT SELLING PRICES USED AT VULCAN MINE DURING 12 MONTHS ABOUT 1915 Material Consumed and Destroyed by Carelessness : 86 bulbs, 45c. each $38! 70 49 cable lengths, 4 ft. each, 20c. each 9 . 80 20 spring terminal sockets, 12c. each 2.40 8 connection pieces, 20c. each 1 . 60 2 lens holders, 20c. each .40 100 lead seals for lamps, 60c. per 100 60 15 rubber corks, 5c. each .75 10 lead check nuts with washer, 13c. each .... 1 . 30 17 lamp holders, 15c. each 2.55 2 lenses, 20c. each 60 3 lamp bodies, 70c. each 2 . 10 8 celluloid battery casings, $1.40 each 11.20 43 celluloid battery tops, 30c. each 12.90 9 oz. celluloid strips, 25c. per oz 2 . 25 5i pt. celluloid paste, $1.75 per pt 9 . 62 Gross total of material $96 . 77 Less 20 per cent trade discount 19 . 35 Net total of material $77 . 42 Material for Care of Batteries: 10 cans of lye solution, lOc. each $1 .00 8 Ib. of vaseline, 8c. Ib ... .64 19 gal. of acid, 18c. gal 3.42 5.06 Net total $82.48 Material Destroyed by Carelessness of Miners: 34 bulbs, 45c. each $15.75 17 lamp holders, 15c. each 2 . 55 3 lenses, 20c. each 60 3 lamp bodies, 70c. each 2 . 10 1 lens holder, 20c. each 20 Gross total of material $21 . 20 Less 20 per cent trade discount 4 . 24 16.96 Material consumed by use during 12 months $65.52 464 COAL MINING COSTS Thus 61 bulbs may be figured as half-consumed, which is equivalent to about 31 bulbs wholly burned out. Thus it is fair to assume that the requirements of the year's running were 82 bulbs. These delivered, as stated, 61,700 burning hours, or 753 hr. per bulb, which is a good result. It is interesting to note that no battery plates had to be renewed during the whole year, and the mine reports that all the electrodes were still in first-class condition. It will be noted that 82 bulbs were sufficient for 27 lamps, so that less than three bulbs served for each lamp for one whole year. Forty-nine cable lengths had to be renewed during the trial year. Let it be assumed that the 27 cables were half worn out when the year was concluded. This will make the equiva- lent number of cables a trifle under 63. In other words, as the total number of delivered lamp shifts was 6350, a cable lasted for over 100 shifts, or as 19.6 shifts were delivered per month, it lasted for over five months, a long life for a part under such stress. This estimate does not cover the maintenance charges though such a small installation does not give representative results in this respect. On the other hand a new fastening has been adopted since this lamp was put in service which will greatly lengthen the life of these. It has been found that 250 lamps can easily be tended by one lampman at $75 per month and one assistant at $45 per month or a total of $120 per month. Assuming that each lamp is operated for 19.6 shifts, the 250 lamps deliver 4900 shifts for a labor cost of $120 or 2.45c. per lamp-shift. Adding the cost of upkeep to this the total cost including all charges will be 3.48c. or roughly 31/2^. At the Keystone Coal and Coke Co.'s Salem mine, New Alexandria, Penn., 200 Edison electric safety mine lamps are in service. One hundred of these were installed September 17, 1915, 50 were added November 18 of the same year and an- other 50 on January 6, 1916. Including the erection of charging racks, switchboards, etc., the cost of installing these 200 lamps was approximately $2000. From September 17, 1915, when the first lot was installed, until June 1, 1916, the only expense for maintenance and MISCELLANEOUS INSIDE COSTS 465 upkeep, all necessary supplies and repair parts, such as cords, bulbs and lenses, was $383.87, which is only a trifle over 24c. per lamp-month. During the period from September 17, 1915, to June 1, 1916, there was a total of 34,250 lamp-shifts, so that the actual cost per lamp-shift for maintenance and upkeep during that time was l%c. The total cost for service, including salary of lamp tender, current, interest on investment and deprecia- tion, was 2%c. per lamp-shift. As these lamps were in continuous service for a period of 6 mo., and some for more than 8 mo., these figures are interest- ing, in view of the fact that they represent the maximum cost at any time during the operation of this type of lamp, since from its very construction, there can be no further deprecia- tion or necessary repairs except those included in the amount given. This same company also had 250 lamps in operation at its Crow's Nest mine in 1915, the operating costs of which are of interest. In a six-months' period the lamps were burned for 34,419 lamp-shifts. The list cost of the spare parts totals $521.83. A discount of 20 per cent may be figured on these prices, leav- ing the net cost of parts and acid consumed $417.45. But of this some material was broken while in the care of the miners: 6 lamp bodies, at 70c $ 4 . 20 20 lenses, at 20c 4.00 218 bulbs, at 45c 98. 10 3 safety devices, at 25c .75 2 lamp holders, at 15c .30 $107.35 Less 20 per cent discount $21 . 47 Net cost of material $85 . 88 This leaves $331.57 of the cost chargeable against the lamp- station expense, and dividing the sum thus obtained by the number of lamp shifts 34,419, the cost of material obtained is 0.963c., or less than Ic. per lamp-shift. It will be noted that the figures given include not merely what are known as spare parts and repairs, but sulphuric acid 466 COAL MINING COSTS and celluloid paste also. Thus all possible material charges are included, though it will be observed that the cost of current, amortization, interest and labor of lampman are not figured. It is noticeable that the cost of electric bulbs and cable are the two principal items in the list. Schedule 6A of the United States Bureau of Mines requires that the average life of lamp bulbs shall be not less than 300 hr. for primary and acid storage batteries and not less than 200 hr. for storage batteries using an alkaline solution. Not more than 5 per cent of the bulbs examined shall give less than 250 hr. life with acid batteries, nor less than 170 hr. life with batteries having an alkaline electrolyte. If the number of working days in the month is taken at 25 and the number of shift-hours at 12, there would be 300 hr. during which the lamps would be in use in any month. Under such circumstances a bulb of 300-hr, capacity would have to be renewed every montfr, or 12 times in a year. If the price of each bulb is 36c., the cost per annum would be $4.32 per lamp per year. This maintenance would be far too heavy, as it covers the cost of bulbs only. At Crow's Nest the total number of approved bulbs re- ceived up to September 18, 1915, was 995. Some of these of course were mounted in new lamps and others were placed in stock. On September 18 there were 157 bulbs in reserve. Thus there were 838 bulbs broken, consumed or in use. The miners had broken 218 bulbs in service. It seems conservative to rate half of these as burned-out bulbs. The bulbs were destroyed by the mishandling of the miners, but not being new they would not have burned as long as bulbs from stock, even if they had not been injured. The loss from negligence can therefore be calculated as being equivalent to 109 bulbs. This can be deducted from the loss as previously obtained, leaving 729 bulbs destroyed by burning out. The same assump- tion, that the bulbs when unbroken have still half their life unexpired, may be applied to the bulbs in the lamps. There are 250 of these. It will be assumed that their life is equiva lent to that of 125 bulbs fresh from stock. These bulbs can be deducted from the number already obtained and the bulbs actually consumed will be 604. The number of lamp-shifts of these 604 bulbs is 34,419. MISCELLANEOUS INSIDE COSTS 467 Each shift is taken at 11 hr., giving 378,609 lamp hours, or 627 burning hours per bulb, or 57 shifts of 11 hr., instead of the minimum average as required by the bureau for an acid battery, namely, 300 hr. Assuming the number of shifts in a year to be 300, then 5.26 bulbs will be needed per year, which at 36c. would entail an expenditure of $1.89. Where the Mannesmann Light Co. maintains its lamps at so much per shift it stipulates that the operating corporation shall charge its miners with the net amount of all material destroyed through the carelessness of these employees. It is fully justified in assuming therefore that the estimates of lamps broken by miners is not in any way excessive. The cable used is marked as 5 ft. long. As a matter of fact it need only be 4% ft. The additional 6 in. is added for conservative estimation. The cable is the one part of the lamp still furnishing a small problem to the manufacturers, and efforts are being made to solve it in a more satisfactory degree. MATERIAL DESTROYED BY MINERS AT CROW'S NEST MINE FROM MARCH 22 TO SEPTEMBER 18, 1915 Period Lamp Bodies Lenses Bulbs Safety Devices Lamp Holders March 22 to April 30 During May 1 3 5 4 50 52 1 1 1 During June 3 45 During July During August 1 1 6 34 23 1 September 1 to September 18 1 1 14 1 6 20 218 3 2 According to observations made, a lamp cable is bent on the miner 's back about 7000 times during the length of a single shift. It is easy to understand, therefore, to what a severe service it is subjected. It is not so much the kind and quality of the cable which is in question. The important matter is the manner in which the cable is attached to the lamp and battery casing. The tests mentioned are being made from this point of view. 468 COAL MINING COSTS 2 OO O 1C ^ * iOO ^ O 'i-I M -O -IN r-t&t I-H IM O IM ,, g ' -iC O -O O O "C O .(N Tt< -1C O (Nt^iN >C 1 -(N ' '^ '^ iC n rH 1C i-H "5 l-(INrH 1C -O . O -- 00 O5 O i-< *><- --*-- 1 Ui ^=m. ---* 3_"[)rainPij>e_ I -I Plan FIG. 5. Plan of two stalls of another type of underground stable. to contain one or any number of mules. Ample room is pro- vided in front and behind the stalls. It will be noticed that a good concrete floor is first laid, in which are imbedded the tracks for the feed cars. Facing and also at the back of the stalls are concrete gutters for carrying out the water daily used to cleanse the stable. Each stall is 8 ft. 4 in. long and 5 ft. wide, and so is commodious enough for the largest of mine mules. The con- crete piers for the center posts are clearly shown in the drawing. Fig. 6 is a cross-section of the stable. The thickness of the concrete floor, position of drain pipe, slight pitch of the 480 COAL MINING COSTS plank floor, car tracks, passage ways, gutters, manger, feed box, center posts, end arches, and the 2-in. gas pipe which alone forms the partition between the stalls, are all clearly outlined in the drawing. The single gas pipe is an improvement over the old style high board partition, as it offers no resistance to the free circulation of air, and the building is always free from obnoxious odors. When it is necessary to stable a fractious mule, he is usually placed in one of the end stalls, and the Cross Section on Center Line of Stall FIG. 6. Cross section of stable shown in Fig. 5 showing manger and drainage system. gas-pipe partition is reinforced by a strap-iron lattice-work, which effectually prevents him from annoying his neighbors. Fig. 7 is a detail drawing showing side view and plan of the center post. This exemplifies a method of supporting a stable roof which is both effective and inexpensive, and in the eight stables where it has been tried, no failures have been recorded. It may here be noted that the end arches extend throughout the entire length of the stable, serving the double purpose of sealing any coal measures and sup- porting the mine roof. Fig. 8 gives the construction details in a longitudinal sec- tion through the center posts. The method of supporting the MISCELLANEOUS INSIDE COSTS 481 T-rail laggings, and the position of the mangers are clearly shown. Fig. 9 is a plan of the arches with the T-rails in position. C-I4- *l 4 A " r ! f. j^- i* 5- r^r* IL 5? vs> | ^ oil rr *i> J^ ^= * I ^3 ,- HL V rrr |i w r r AJ3 71? FIG. 7. Cast-iron center post used in stables shown in Figs. 5 and 6. FIG. 8. End elevation of arches for stable shown in Figs. 5 and 6. Fig. 10 supplies all needed details of construction of the manger with its feed box, which have been designed with a 482 COAL MINING COSTS view to sanitation as a leading desideratum. They are usually constructed in the shops and sent to the mines with all parts FIG. 9. Plan of arches in Fig. 8 showing T-rail reinforcing. r* - 5/ - 1 K 4'- ol I 1 t ' \OOOOOOOCOOOOOOOOCOOOOOO booooooooooooooooocoooo \ooooooiooooooooooOooooo pOOOOO'OOOOOOOOOOOOOOOO poococj'ooooooooooqooooo 00 00 OOO CO 00 00000000000 /OOOCOQ'OOOOOOOOOOOOOOOO /6 O O O O O Q'O OOOOOOOOOQOOOOO A ! i ; i ' 1 3t j A fw v V" - p^r--- -j - - - jooooooojoooooooooo^ooooo "T""^~" 71 ill i S! s ^\ / ir 'j| : & jl*j y Plan 1 o o ; !; A o < ' 1 1 u || I" o 1 1 II QQ IP l 1 t? ,! \ \ Part Fronl Elevation End Elevation FIG. 10. Detail of manger and feed box for stable shown in Figs. 5 and 6. marked to facilitate in erecting. The work of installing them is thus simplified, and can be done by the average mine MISCELLANEOUS INSIDE COSTS 483 timberman. Particular attention is called to the hole provided in the bottom, affording a convenience for cleaning same. When construction of the building is completed, the sheet iron which forms the temporary support for the concrete arches is not withdrawn, but is left in position and when white- washed serves as an. efficient reflector for the electric lights which are provided for each stall. A 32-cp. lamp with a heavy guard is provided with a No. 14 (BX) extension cord 10 ft. long. With this the stableman can make a careful inspection of the mule as it returns from its day's toil. The following is the bill of material for one stall only, and an equal increase should be made for each additional mule for which accommodation is required. BILL OF MATERIAL FOR ONE STALL 40 bags of cement. 5 tons of sand. 14 T-rail laggings 4 ft. 6 in. long (40-lb. rail) 70 ft. odd lengths old T-rail (for reinforcing). 1 10 ft. length, 8 in. dia., cast-iron column pipe 13 pieces of sheet iron 4 ft. 6 in. long XI ft. 6 in. wide (for arches) 2 pieces of No. 8 sheet iron 5 ft. long X 3 ft. wide. 1 piece of No. 10 sheet iron 2 ft.X2 ft. for feed box. 5 floor planks 8 ft. 4 in.Xl ft.X4 in. thick. 1 bolt 1 ft. 83^ in. longXf in. diam. 1 bolt 1 ft. 5^ in.Xf in. diam. 1 bolt 1 ft. 2^ in. longXf in. diam. 3 bolts 1 ft. 3f in. longXf in. diam. 3 bolts 4 in. longXf in. diam. 2 bolts 3 in. longXf in. diam. 28 rivets 1 in.X^ in. diam. 1 piece 2 in. diam. gas pipe, 7 ft. 8 in. long (for partition between stalls). 1 piece 3 in. diam. gas pipe, 5 ft. 9| in. long, (for drain). 2 iron straps, 2 ft. 10 in. long X 2 in. wide X \ in. thick. 2 iron straps, 2 ft. 9 \ in. long X 2 in. wideX^ in. thick. 10 lin. ft. of strap iron 2 in. wideXs in. thick. 1 piece 1-in. mesh segment for tray 4 ft. longXl ft. 2 in. wide. 2 brackets. Overcasts. The Lehigh Valley Coal Co. in 1907, started sub- stituting for the wood overcast one of concrete and steel. These overcasts will be permanent and substantial, their destruction only being accomplished by a squeeze, in which event all other construction would be destroyed as well. The construction of the overcasts is shown in Fig. 11. Being located in an old portion of the mine, much gob and other 484 COAL MINING COSTS g i _ _L ^- rt -* FIG. 11. Plan View. MISCELLANEOUS INSIDE COSTS 485 H S 486 COAL MINING COSTS refuse is found 011 both sides of the road and must be cleared away at the immediate location of the bridges. After this trenches for the masonry walls to carry the bridge are dug and carried to sufficient depth below the bottom slate of the vein to obtain a solid foundation. The walls are then built 4 ft. thick, consisting of rock and bone of sufficient size selected by the mason from the gobbed refuse made and packed in the chambers during mining. Lime mortar is used and the faces of the wall are almost entirely surfaced off with a coat of mortar after completion. Next in order, timbers for carrying the concrete are placed by the timberman, and consist of second-hand 3 X 5-in. wooden rails taken from old chambers and of second-hand mine ties and props for uprights, on which are laid 2-in. planks double thick, which in turn carry 2 in. planks on edge. To these planks are nailed the 2-in. planks which carry the concrete. The time consumed is somewhat greater than one might con- sider and an explanation is necessary. The clearing away of the refuse accumulated is very laborious and progress is slow, and in excavating for the trenches considerable heavy rock and bone must be removed. The work being done in the return current makes it warm for men to work, and less is accom- plished than if the same were located in a fresh-air current. Again, the rock and bone constituting the material for the masonry wall are all collected in the old chambers in the vicinity and taken to the location for the wall. The broken stone used in the concreting is obtained from a chamber some distance away, the stone having been dumped there during the driving of the rock slope. The mason selects what material is fit for use. At the same time sand suitable for concreting is gathered from the same place, having been made during the blasting in the slope and loaded out with the rock. As will be seen by the end view, Fig. 11, two concrete walls form the sides of the overcast and become necessary where the entire vein has been mined out. The Baltimore vein, which in this particular location is one seam, more fre- quently splits into two distinct seams, the top split known as the Cooper or Upper Baltimore, and the bottom split known as the Bennett or Lower Baltimore vein. At the location in question the dividing slate is only about 1 ft. thick and in MISCELLANEOUS INSIDE COSTS 487 most instances both splits have been removed. At the location of No. 4 overcast only the bottom split was mined and it was necessary to take down the top vein to get sufficient height for the roof of the overcast. In taking down this top coal, or split, the ribs were neatly dressed and the concrete floor of the overcast was carried high enough to dispense with the walls. At the location for Nos. 1, 2 and 3 overcasts the total vein has been mined and concrete side walls will become neces- sary. The two masonry walls are built 4 ft. thick, the walls on the up-pitch side being on an average 7 ft. high, the one on the down-pitch side averaging 10 ft. high. The concrete floor is 12 in. in thickness and 22 ft. long b> 20 ft. wide, consisting of a 1-2-3 mixture. The breadth and length will of course be greater or less depending on the dimensions of the openings through the coal at the different locations. The T-iron rails are spaced 12 in. center to center, the bot- tom or base of the rail being embedded 2 in. in the concrete. During concreting the rails are supported by blocking under the head. For the side walls a 1-2-5 mixture will be used, the walls being 12 in. thick and cement worked well into the crevices of the coal and top rock. The details of cost of the No. 4 overcast are given here- with, the figures being as of 1907. The day's work consists of nine hours and the hourly rate includes the strike percentages and sliding scale. The proportions for the concrete mixture were 1-2-3. Sand and stone for the concrete and walls do not appear in cost for material, the only cost on these items being included in the labor. The total cost per cubic yard for the masonry wall is there- fore $1.63. The cost for labor and material is about equal for the overcast or $5.25 per cu.yd. for each, a total of $10.50 per cu.yd. The two walls needed for Nos. 1, 2 and 3 overcasts make an additional 11.5 cu.yd. at a somewhat smaller rate per cu.yd., say $8, so that the entire cost of the overcast, including two supporting masonry walls and two concrete side walls, in a vein of this thickness will be about $365, and about $270 without the side walls. 488 COAL MINING COSTS MASONRY WALLS, 60.5 Cu. YD. LABOR Clearing away refuse, digging trenches, getting stone, mixing mortar and building two walls, 2 masons, 20 days each, or 360 hr. at 25.4c $91.44 MATERIAL 25 bu. of lime at 30c . . 7 . 50 $98.94 CONCRETE OVERCAST, 16.5 CU.YD. LABOR Getting lumber for supporting concrete work and placing same, 3 men at 2 days each, or 54 hr. at 25.2c $13.61 Placing 2-in. plank for concreting, 1 man 2 days or 18 hr. at 25.2c 4.54 Getting broken stone, mixing and placing concrete, 2 masons at 14 days each or 252 hr. at 25 .2c 63.50 Bending and transporting rails approximate 5 . 00 $86.65 MATERIAL 120 bags portland cement at 45c $54 . 00 1000 ft. 2-in. hemlock plank at 22c 22 . 00 20 Ib. 20d. nails at 2c 50 1850 Ib. ( T %- ton) second-hand 25-lb. T-iron rails at $12.50 a ton . . . 10 . 00 $86.50 Tile stoppings and overcasts. Terra-eotta blocks have proven their usefulness in a mine, as well as for the purposes to which they are put outside. Not only have they been used economically for stoppings along the main air courses, but they are now being adopted successfully for the walls of overcasts. It appears also that scrap iron is to be displaced by a rein- forcement which gives a concrete structure not only more artistic but more economical and more quickly erected. The Consolidation Coal Co., as a preventative of fire at the same time as a safety-first measure, decided about 1914 to eliminate wood stoppings and wood overcasts where practi- cable and to allow no more new overcasts anywhere to be con- structed of combustible material, the same to apply to perma- nent stoppings along main entries. MISCELLANEOUS INSIDE COSTS 489 The most natural step was to use the cinders from the boiler house at the mine and to build these structures of con- crete. However, investigation proved that too much money was being spent for these improvements. The stoppings were reduced to a thickness of 4 in., but the cost still appeared too large. The terra-cotta blocks were first used for stoppings and their success was readily apparent. As the 4-in. cinder-concrete stopping had proven stable, a hollow block of 4-in. width was at first advocated ; but the 5 X 8 X 12 - in - block was finally adopted. On main entries where the stoppings are intended to remain for a considerable length of time they are bonded with a mortar composed of one part cement to two parts sand. On room headings where these stoppings have been built, lime was used in place of the cement, this making it possible to tear down the stopping when desired, without destroying the tile. Sufficient cost data have been obtained on the construction of this type of stopping at the mines using this material to show a saving of from 40 to 50 per cent in every instance as against the use of concrete. The cost compared to wood stop- pings for room headings shows about the same difference in favor of the wood. It was thought that by the recovery of the blocks and their reuse this difference could be more than off- set; but accurate costs kept on this work shows that it would be necessary to recover these. blocks several times over to do this. Therefore, it is not advisable to use the blocks for tem- porary structures unless for other reasons than economy. The crosscuts in the mines of the Fairmont region run about 10 to 12 ft. wide and 7 to 8 ft. high, and average about 80 sq.ft. in area. It has been found that two men will construct three or four stoppings per day, providing the material is placed ready for their use. It might be possible to reduce the cost of stoppings further by the use of ' ' Self-sentering, " "Hy-Rib," or "Rib-Lath," plastering these with a cement mortar an inch or two thick. A stopping made with any of these materials properly placed, should carry as much as, or more pressure than, the ordinary wood stopping. The plan of erection would be simple, con- sisting of trenching the ribs, roof and bottom an inch or two, 490 COAL MINING COSTS slipping the adopted metal in place and plastering with cement or lime mortar, the material depending upon the desired life of the stopping. Price and the strength which could be secured should govern the selection of the reinforcement. A heaving of the bottom would in all probability be cut by the stopping, since heaving is caused by the swelling of the fireclay, this material at the same time softening markedly. If the pressure comes from the top or roof there might be much deflection in the stopping without doing it any material injury, besides it could be as easily repaired as the wood stopping in case of sufficient pressure to cause cracks in the plaster. All well ventilated mines must contain overcasts, and the costs of these structures soon run into considerable amounts, if they are not carefully planned and constructed. There are few mine foremen or even superintendents who take the trouble to ascertain accurately just what it costs to erect an overcast so that he can tell you how much was spent per cubic yard for concrete or what the walls and roof cost separately. Fig. 12 shows the type of overcast now being constructed in the mines of The Consolidation Coal Co. in the Fairmont Region. Blank spaces are left in the blueprints, and dimensions to suit the location are supplied. Since the sheets of "Self- sentering" are sold in even foot lengths, the distances between walls are made to correspond. Thus for an 11-ft. sheet 2 in. are deducted for the 12-in. rise and two or more inches on each side for bearing spaces, leaving the distance between walls 10 ft. 6 in. This style of overcast makes use of the terra-cotta blocks, the 5 X 8 X 12-in. block being laid to give 8-in. walls. On top of each wall is placed a layer of concrete on which is set a small steel rail 2 in. from the edge of the wall. Then the reinforcing sheets of ' ' Self-sentering, " which come already shaped, are placed abutting on the flanges of the rails and the center line or middle of the reinforcement supported by a 3 X 5-in. wood stringer ; then the concrete is spread on the top as required on the plan. After the concrete sets, the wood stringer is knocked down and a coat of cement mortar plas- tered underneath. Not a single wood form is required on the whole structure, except around the iron rails on top of the w^alls, and even this can be eliminated by standing a row of MISCELLANEOUS INSIDE COSTS 491 the terra-cotta blocks on the 5-in. side just outside of the rails, thus making them serve as forms. f Cinder concrete. /?/ cinders to be thoroughly netted, concrete mined fo the consistency of paste and placed on "&elf-sente ring "remforcement &X..5" posts nitna3xS"cap fo be placed on center line to support arch while concrete roof is being placed \. [Self-sentering* reinforcemenr Ho. 24 gage ( 3 Concrete on top of rt in forcf- < ment, I coat-cement mortar (. underneath FIG. 12. Reinforced concrete overcast supported on terra-cotta blocks used by the Consolidation Coal Co. There are times when it is necessary to deviate from the arched-roof type on account of the skew of the air bridge, in which case it is more economical to use the flat roof. Fig. 13 x5 Stringer entire length ofovercasr I" Cement footing each pipe SECTION A.-A FIG. 13. The Consolidation Coal Co.'s reinforced concrete, hollow tile overcast with flat roof. presents a design of this kind where the roof is supported by three or four old boiler tubes, the reinforcement being sus- pended from these tubes or pipes by means of wire of sufficient 492 COAL MINING COSTS strength to carry the weight. This type of overcast also needs only temporary supports along the middle while the concrete is being placed, thus saving considerable labor and material in the construction and emplacement of forms. The Utah Fuel Co. constructed six concrete stoppings and one overcast in some work they did about 1914. On account of the height of the stoppings they were built of reinforced concrete with reinforcing wings as shown in the accompanying drawings, Figs. 14 and 15, old rails and wire rope being used for the reinforcing material. The reinforcement of the over- cast was made with the same class of material. The cost of these is given in the figures. Comparison of doors and overcasts. A mine door is a costly as well as dangerous item of equipment, yet this seldom receives the thought that it deserves. When it is necessary to install a permanent door the conditions should be carefully considered to see if it will not be cheaper and safer to put in an overcast. A permanent door involves much expenditure in addition to that for lumber. The wages of two men must be paid for its construction and erection. A trapper will have to be con- stantly employed to open and close it. A shelter hole must be constructed at the door and another hole also must be provided for a barrel of water, for all permanent doors should be protected against a possible fire. The continual breaking and smashing of the door by trips involves a further expendi- ture for lumber and for the workmen who replace it. A light should be provided at all doors. This involves the wiring of the light, the replacing of globes, and the employ- ment of wire men who must spend a day, more or less, in installing it. Trolley wires must be guarded on both sides of the door, which means the use of more lumber and the employ- ment of more labor. These boards often are knocked down by trolley poles and again the service of a high-class or highly- paid man is needed. Where permanent doors are used an extra door should be placed that will have the same effect on the ventilation. This door, of course, should be a full trip length away. Heavy blasting or heavy caving often damages doors. The weighting MISCELLANEOUS INSIDE COSTS 493 of the roof, the movement of the sides or the heaving of the bottom will affect them also. Looking at the matter from a safety viewpoint we often find the door frame cuts the clearance, frequently to as little 30-lb. T Rail Reinforcing, ' J5 C+oC. ELEVATION FIG. 14. Plan and elevation of the Utah Fuel Co.'s concrete overcast. .--.26-0*- - >i 2 ft- |< M-O- OOUBLE BRACED STOPPING DETAIL OF COST PER CU. YARD Stoppings and I overcast Total 119.0 Cubic Yards d work Labor $3.15 per 8Hrs. Carpenters$3.40 perSHr* ITEM LABOR 1ATERIAU TOTAL Reinforcinq Material (Scrap Iron) i> 0.278 $0.278 Qatherina & Distributing Gravel in Mine 2.683 Cement 0.076 $2.659 2 735 Forms 1.936 0.554- 2490 Redistribute Gravel.Mix & place Concrete 3.835 * 3.835 Totals. & e.&oa &3.ZI3 fT2.02T Gravel hauled in winter. Difficult to redistribute gravel and place concrete on account of old workings * INCLUDES COST OF MAKING OLD WORKINGS SAFE FIG. 15. Plan and cost of reinforced-concrete stoppings at the Utah Fuel Co.'s mines. as 12 in. A brakeman or driver, knowing he has 30 in. clear- ance, forgets the door frame is in his way and accordingly at this point a man may be squeezed to death. The required shelter hole is not always provided, and an instance is recorded 494 COAL MINING COSTS where a trapper stood back of his door and the heavy iron guard receiving the weight of the cars crushed him so severely that he died. Should the boy neglect his work the motorman is in great risk of losing his life by running into the closed door. The cross bar over the door often is far below the uniform height of the roof and may easily dash a man's brains out. Doors have been known to take fire and cause serious damage. One at Delagua, Col., is thought to have been set on fire by the trapper and been the cause of a destructive fire in which many lives were lost. Trappers, drivers and workmen leave doors fastened or thrown back and as a result that portion of the mine where the air is short-circuited may be endangered by an accumula- tion of gas, with a result that is not to be reckoned in dollars and cents but in humanity. It is not advisable to place doors at the foot of steep grades. Even automatic doors are danger- ous where a man cannot control his trip, as these doors are not always positive in action. The penalty under the compensation schedule for neglect to dispense with the door certainly makes it a costly and dubious economy. Doors must be so hung that they will close themselves, must be strongly built, tightly sealed into the roof, and a second door that has the same effect on the ventilation is required. For neglecting to comply with these specifications a charge of Ic. per $100 of payroll is the penalty provided. Under the rules of the compensation schedule shelter holes must be main- tained at all permanent doors. A penalty of 6c. can be charged the operator not complying with this requirement. Doors must be whitened or enclosed lights maintained. A 6c. penalty is provided for non-compliance with the requirement that the trolley wire at doors be guarded on both sides. A clearance of 30 in. must be provided between the widest part of the motor or cars and the frame of the door or a 12c charge can be made. When these charges are totaled we find that one door can create a maximum of 31c. per $100 payroll. When the total payroll of the mine is computed and this pro- portion deducted it will give some idea why a permanent door should not be placed, but even when conditions seem to warrant MISCELLANEOUS INSIDE COSTS 495 paying the penalty to save the first cost of an overcast it is generally more advisable to erect it. The initial cost is greater, of course, but in the end the overcast will be found to have paid for itself in saving the expenditures enumerated. With the overcast the clearance can be made ample. Cost of stone and wood brattices. An ordinary wooden brattice, single thickness, may be figured as follows, figures as of 1907 : 1-in. plank, -m. strips and waste, 100ft. B.M. @ $20 per thousand $2 . 00 Labor $1.75, 3 props and caps, 5c 1 .90 Daubing 35c., nails, etc., 25c . 60 $4.50 The life of this brattice will ordinarily be from 4 to 5 yr., so that $1 per year per brattice may be taken as the cost of maintenance. Where top is shot, or where there is much draw slate, it will usually be found, particularly in low coal, that there is considerable cleaning out to do to get at the old brat- tices; which work may readily cost as much as the brattice itself thus making the cost of maintenance $2 per year, instead of $1. Doors cost in 1907: Lumber, say 120 ft. B.M. @ $20, $2.40, nails $0.10 $2.50 Labor $3.75, hinges $0.75 4.50 $7.00 There are in most mines, at least one or two places where a door boy at 75c. per day, or say $200 per year, could be replaced by an overcast at $50 first cost, good for 5 yr., and thus costing only $10 per year, making a net saving of $190 per year for each door boy replaced. Where the old brattices have to be dug out before they can be replaced, this work may readily amount to $1 per year for each old brattice, or $90 additional. Where the mine work- ings are so arranged that the brattices on the cross entries will not be in service longer than the life of the first brattices put in, the $160 for maintenance may be dropped. It often occurs that shorter new lines are driven, but this is objectionable, as 496 COAL MINING COSTS it leaves no suitable air current* along the haul ways to carry off the dust from haulage and this is the class of dust most to be dreaded. Where the brattices have to be dug out the total cost may be $2250, or 1.25c. per ton, for a year 's output. Stone brattices are preferably from 1^ to 2y 2 ft. thick, depending upon thickness of coal, top, etc., and should be well mined or cut into the rib, particularly in the case of soft coals that spall off readily. Where the coal is low and hard, with good roof and bottom, a 1-ft. wall would seem ample. The follow- ing table shows the contents in cubic yards of different sizes of brattices: FOUR-FOOT COAL Width Crosscuts or Breakthroughs in 1 Foot Thick, 1.5 Foot Thick, 2 Feet Thick, Feet Cubic Yards Cubic Yards Cubic Yards 10.0 1.48 2.22 2.96 12.0 1.78 2.67 3.56 14.0 2.08 3.11 4.15 16.0 2.37 3.56 4.74 18.0 2.67 4.00 5.33 20.0 2.96 4.44 5.92 22.0 4.89 6.52 8.15 24.0 3.56 5.33 7.11 SIX-FOOT COAL Width Crosscuts in 1 Foot Thick, 1.5 Feet Thick, 2 Feet Thick, Feet Cubic Yards Cubic Yards Cubic Yards 10.0 2.22 3.33 4.44 12.0 2.67 4.00 5.33 14.0 3.11 4.67 6.22 16.0 3.56 5.33 7.11 18.0 4.00 6.00 8.00 20.0 4.44 6.67 8.89 22.0 4.89 7.33 9.78 24.0 5.33 8.00 10.67 MISCELLANEOUS INSIDE COSTS 497 Where the draw slate forms a durable building stone, double-stone brattices, filled with muck, may be built for $10 to $15 (figures as of 1907), part of the actual cost being prop- erly chargeable to " slate," or entry cleaning. Where stone must be quarried and brought in from the outside, a good brattice, single thickness, may cost anywhere from $15 to $35. Where coke cinder is available, cinder concrete may be used for brattices and overcasts. Disregarding the cross entries, which may be assumed to be the same in either case, and taking $25 as the cost of a stone brattice, which, particularly for low coal, may be con- sidered a liberal estimate, the costs of wooden and stone brat- tices may be figured as follows: Stone brattices: Main, 10 brattices @ $25.00 per year $250.00 Wood brattices: Main, 90 old brattices, maintenance @ $1.00 $90.00 Main, 10 new brattices @ $4.50 45 . 00 135.00 $115.00 The balance of $115 in favor of wood brattices is equivalent to 0.064c. per ton, for 180,000 tons. Figuring the cost of maintenance of wooden brattices at $2, instead of $1, the balance in favor of wood brattices is only $25, or equivalent to 0.014c. per ton for 180,000 tons. Figuring stone brattices at $25, and the maintenance of wooden at $1, a stone brattice will need to be in service 25 yr. to be as cheap as a wood brattice ; but figuring the maintenance at $2, will only need to be in service 12^ yr. Double wood brattices filled with muck may be figured at $10 to $11 each, with say $2 to $2.25 per year, for maintenance. Compared with this a stone brattice will only need to be in service 12% yr., or allowing $1 per year for digging, say 8 yr.. to be as cheap as a wooden brattice. Stone brattices cost somewhat more than the double air-course arrangement, where the latter does not involve shooting considerable top, or equiva- lent yardage for narrow work, but cost less where yardage must be paid. 498 COAL MINING COSTS Refuge Chambers. Subsequent to the Cherry (Illinois) mine fire there was a general feeling among the engineers of the country that underground refuge chambers should be established at all mines to prevent a repetition of this insofar as was humanely possible. A paper was presented before the West Virginia Mining Institute in 1910, dealing with this ques- tion some excerpts from which are given herewith. The maximum size of a district to be supplied by a refuge chamber depends somewhat on the geological and other phys- ical conditions presented by the seam and the system of work- ing same. It would seem desirable to have it bear some rela- tion to the maximum number of men employed in a district ventilated by a separate split of air. We will assume that the maximum number of men is one hundred, a not uncommon maximum allowed for a single split of air. As there will be new districts or panels forming while others are being worked out, the average number of men we will figure at 50. A medium-sized mine has about 200 men employed on the day shift, and a large mine about 500. Accepting the average of 50 men in a district, there would be from 4 to 10 live districts in a medium- to large-sized mine, and as many refuge chambers under the system proposed. COST OF REFUGE CHAMBER 500 ft. 2-in. common pipe casing, in place, say $50 50 ft. of excess room neck yardage and special entrance, say 50 5 room crosscuts, say, 100 ft. of yardage 50 5 masonry stoppings, at $10 50 6 masonry door frames, at $5 30 6 doors and frames, at $6 36 Sanitary closet and fixtures 15 Wall cases with glass fronts 20 Casks, pails and miscellaneous fittings 10 Food in tins and cans, say 25 6 dry cell electric lights, say $5 each 30 2 safety lamps, at $5 10 1 oxygen resuscitating box, with two cylinders 45 First aid box, medicines and disinfectants 25 Miscellaneous, say 54 Total $500 To establish these refuge chambers may appear to be a serious task, but if they are planned for in laying out the mine, MISCELLANEOUS INSIDE COSTS 499 the cost per ton would be insignificant. Nearly all modern coal developments, as a matter of good engineering, are, or should be preceded by thorough prospecting, both to knoiv the continuity of the seams and to properly plan the mine. In the following estimate, the room is not considered an added expense, except for the extra length of room neck. The cost of drilling the hole is considered part of the cost of pros- pecting; the cost of its casing for an assumed depth of 500 ft. is alone considered. The telephone is not regarded as an extra cost. The foregoing provides for a good equipment ; other appara- tus mentioned previously should be considered as part of the mine equipment. If a mine had 6 such stations, the cost underground would be $3000. On the surface the special equipment would vary widely with the physical conditions and regular equipment. If a mine used compressed air, the only additional cost for the stations would be the outside pipe lines. These pipe lines need not be large, as economy of operation would not enter into the calculations. It is probable that all such lines to drill holes of six refuge chambers could be supplied at from $1500 to $2000, under ordinary conditions. When the mine has an electric plant but not a compressor plant, the additional surface equipment would be the cost of the power lines to the various drill holes and the cost of the small motor-driven fans or compressors. Each drill hole sur- face instalment could probably be put in at a cost not exceeding $500. When a mine had neither compressed-air nor electric plant, the cost of instalment would, of course, be much greater, as it would involve a small central plant. However, it may be pointed out that such a plant would be extremely useful, and no doubt pay for instalment on other grounds. Let us assume that the average total cost of instalment of district refuge chambers figures as much as $10,000, or let us say 5 per cent of the total cost of the mine investment, the possibility of saving a considerable number of lives, if disaster comes, makes it seem a good investment. Mine sprinkling costs. At the Sunnyside, No. 2 Mine in Carbon County, Utah, quite an elaborate sprinkling system 500 COAL MINING COSTS was installed about 1908. In this system the smallest pipe used on the haulageways was 1% in. and in the rooms, % or 1 in. pipe is used equipped with a brass hose bibb and kept within 200 ft. of the working face. In operation, two men are continually employed and they are required to attend to all extensions and repairs, as well as keep the mine sprinkled. In an ordinary mine the necessary work on pipe lines will not occupy more than an average of 2 or 3 hr. of their time per day, the work including extensions in rooms and entries where work is advancing, taking up pipe where work is retreating and roof expected to cave, repair- ing broken pipe, packing leaky valves, etc. Each water man will carry with him 150 to 200 ft. of %-in. rubber hose, attaching one end to the hose bibbs which are opened by the key or lever, which he carries. He will thor- oughly wet down the roof, floor, and sides, or ribs, of all open- ings accessible, paying especial attention to the vicinity of working faces, brattice and timbers and from time to time wetting down abandoned rooms, etc., which are still open. The water men are instructed to keep all parts of the mine sufficiently damp so that upon taking a handful of debris from the floor, and subjecting it to pressure of the hand, it will cake and retain its shape after removal of pressure. All brat- tice must also be kept damp, this requirement alone demand- ing the presence of water men at least every other day. If fine dust is dampened and then allowed to dry, it is very diffi- cult to penetrate it with water, due to the tendency of this fine dust to form an almost impervious film of dust around globules of water, while if this dust is kept damp, there is no dry fine dust present to imprison and waste the water. Pour- ing water in quantity, as from pail or barrel, on dry fine dust, is wasteful of water and absolutely ineffective, due to the tendency of the fine dust to form the film above mentioned, while the use at frequent stated intervals of a stream with good pressure finely divided by hose or nozzle moves the dust and permits the water to penetrate the dust particles, not superficially, but through to the floor. At Sunny side No. 2 Mine, Carbon County, Utah, the coal vein is 6 to 10 ft. thick, and pitches about 10 per cent. Here the coal is absolutely dry and dust very inflammable, making MISCELLANEOUS INSIDE COSTS 501 the wetting of dust absolutely necessary. The above-described sprinkling system is here an unqualified success. The cost of the system was about as follows in 1907 : Labor Materials Total 50,000 gal. redwood tank in place Pressure pump in powerhouse $600 175 $ 900 825 $1500 1000 3000 ft. 3-in. pipe 300 750 1050 1000 ft. 2-in. pipe . ... 30 120 150 20,000 ft. IHn. pipe 300 1,600 1900 17,500 ft. 1-in. and f-in. pipe 130 770 900 Hose, hose bibbs, valves, elbows, etc 500 Total $7000 The cost of this plant is somewhat higher than would be necessary elsewhere, due to the fact that very seldom would both tank and pump be required. The above figures, more- over, are for a well-developed mine with distances somewhat well advanced from the surface. Here also, the following of a systematic plan from the start for laying pipe, instead of pro- ceeding hit or miss, would have diminished the amount of pipe required considerably. Even installed at the above cost and on a tonnage of about 300,000 tons per year, the entire plant could be paid for in one year at cost of 2 a / 3 c. per ton and if the cost were distributed over a period of five years, less than y 2 c. per ton would suffice, which is certainly not prohibitive. The operating cost in 1908 was about as follows : One pipe man, 275 days at $2.75 (he also sprinkles). . $ 756.25 One water man, 275 days at $2.75 756 . 25 20,000 gal. of water per day for 275 days at 12c. per 1000 gal 660.00 Powerhouse expense including labor running pump, coal for steam, etc 275 . 00 6 per cent interest on investment 420 . 00 Depreciation 10 per cent 700 . 00 Extensions, repairs, etc. (material) 500 . 00 Total $4,067.50 The sum of $4067.50 per year on 300,000 tons amounts to lVs c - P er ton > a mer e trifle compared to benefits derived as will 502 COAL MINING COSTS be quickly admitted by any company which installs tne system. In the above operating cost the 12c. per 1000 gal. is variable even for this mine, and covers the cost of bringing the water to the storage tank. This cost has at times amounted to $1.50 per 1000 gal., when water was hauled 40 miles in railroad cars and even at this cost sprinkling was kept up, though its extent was somewhat restricted. Depreciation at 10 per cent is liberal as there is little wear on tank or pressure pump, and the pipe will, under ordinary treatment, last 6 or 8 yr., even where laid on coal debris and subject to the corrosive action of acids pro- duced by leaching of coal. Material for extensions, repairs, etc., will amount to more than $500 per year when a mine is new and ground being opened up fast, but during the period of retreating there will not only be no new material needed, but pipe fittings, etc., will accumulate, hence $500 is placed as an average. Power-house expense is the cost of running the pump at the power house to pump water from the tank to the mine, and includes attendance of engineer, who also attends air compressor, dynamos, etc., as well as cost of coal, water, labor, etc., used in generating steam to run pump. INDEX Acceleration in haulage, 227 Adhesion of steel and cast-iron wheels to the track, 228 Adjustable-turret arc wall cutting machines, 111 Africa. See "South Africa" Air, compressed. See " Compressed Air" Alternating current: Layout for mine using alternating current cutting machines, 109 Anemometer tests, 444 Animal haulage costs. See "Haul- age" Anthracite, average and bulk line costs of, 17 Anthracite : Increase in costs after second War Bonus, 18 Percentage of different sizes of, 15 Prices fixed by the President, Aug. 23, 1917, 14, 19 Prices received for White Ash, average, 17 Prices fixed, Dec. 31, 1918, 20 Royalties on, 1, 16 U. S. Census Report on, for 1909, 1 Anthracite Coal Waste Commission estimates of percentage of recov- ery, 164 Apportionment of costs to cover future operation, 31 Arc wall cutting machines, 111 B Back haul, 222 Bearings, roller, for mine cars, 227. 301 Bits for mining machines, 102, 115 See also "Mining machines" Number required, 116 Sharpening, 115 Sullivan Machinery Co., "Dread- naught," 115 Tempering, 116 Blasting, 125 Dynamite, 128, 177 Guncotton, 130 Missed shots, 138 Nitroglycerin, 129 Shaft sinking, 177 Bonding rails, 267 Compressed terminal bonds, 267 Copper wire resistance, 272 Efficiency of, 271 Losses in bonding, 268, 272, 275 Rail to copper ratio, 273 Solid terminal bonds, 270 Wire rope for, 276 Brattices, 488, 495 Breakage of coal. See "Screenings" Buying mine locomotives, 232 Candlepower, 456 Portable lamps, 458 Capacity : Mining and loading machines. See under that title Mining machines. See under that title Capital invested: Anthracite mines, 2 Bituminous mines, 3 Illinois and Indiana mines, 5 Westphalia (Germany) mines, 33 Cars, mine, 299 503 504 INDEX Cars, Capacity of, 299 Frictional resistance of, 224, 227, 230 Oiling, 302 Repair costs, 300 Roller bearings for, 301, 303 Steel cars, 299, 300 Stretcher cars, 303 Tare and weight, 300 Wheels for, 303 Cartridges, hydraulic, 131, 134 Advantages of, 135 Capacity, 135 Commercial value of, 133 Percentage of large coal obtained with, 132 Test of, at the Hulton Colliery (England), 132 Cement gun for timbering. See "Timbering" Census reports, U. S. for 1909, pro- duction, costs, salaries, etc., 1 Charges. See subject as "Deple- tion," "Depreciation," "In- terest," etc. Coke, maximum capacity of drawing ovens at the operations of the U. S. Coal & Coke Co., 147 Colorado : Systems of mining and percentages of recovery in, 166 Detailed cost of sinking 570-ft. shaft, 198 Comparative cost of different systems of haulage. See "Haulage" Compressed air: Locomotive haulage costs. See "Haulage" Mining machines. See under that title Underground compressor for driv- ing conveyor, 95 Compressed terminal bonds. See "Bonding" Compressors (air) : for shaft sinking, 186 Single- and two-stage compared, 186 Concentration method: Connellsville district, 81 Gary, W. Va., 48 Concrete : Shaft linings. See under that title Timbers. See under that title Connellsville district : Frick Co. system of mining, 82 Method of laying out in 90-ft. blocks, 87 Method of working at the Conti- nental mine No. 2, 86 Systematizing work in rooms, 85 Systems of working, 80 Conservation : Anthracite Coal Waste Commission on, 164 Complete extraction required at German mines, 33 Economic aspects of, 164 Factors governing percentages of recovery in various fields, 168, 170 Percentages of recovery at the mines of the Pocahontas C. & C. Co., 79 Summary of results in different fields, 165 Use of longwall in, 173 Continuous panel system of working at Gary, W. Va., 52 Contract form for shaft sinking, 217 Conveyor system of mining, 93, 95 Copper wire, resistance of, in bond- ing, 272 Costs: See under various heads Apportionment of, to cover future operations, 31 Central Pennsylvania, 26 Illinois No. 6 District, 24 Indiana, 24 Increases in, from 1916 to 1920, 27 Increase in, due to intermittent work, 161 Influence of thickness of seam on, 22 INDEX 505 Costs: Labor. See under that title Middlewestern and German mines compared, 31 Ohio No. 8 district, 25 Pocahontas field, 25 Southwestern Pennsylvania, 23 Cottonwood Coal Co., shaft sinking report, form of, 215 Crozer Land Association system of mining, 72 Curtain wall for shafts. See "Shaft linings" Curves, track. See "Track" D Daymen, 139 Average number required at 454 mines, 140 Effects of the thickness of coal on number required, 142 Efficiency of, 60 Production per dayman, 141, 142, 145 Variations in, with capacity of mine, 142 Depletion charges at anthracite mines, 20 Depreciation : Charges at anthracite mines, 20 Depth of undercutting, 108 Developed coal property compared with an undeveloped, 28 Development work: Rate of, 53, 95 Shaft sinking. See under that title Direct current compared with alter- nating for operating machines, 109 Dominion Coal Co. loading machines, 118 Doors, underground, 492 Drainage. See "Pumping" Drawbar, determining height of, 229 Drawbar-pull of locomotives, 222, 224, 231 Drawing pillars. See "Pillars" Drills, 196 Churn, 180 for soft material, 180, 188 Tests of, on the Transvaal, 433 Steel for, 437 Drilling : Churn, 180 for shaft sinking, 176, 177 hand, 177, 189, 215 in soft material, 180, 188 records made in the Transvaal, 433 Dynamite : See also "Blasting," "Explosives" Fumeless, 177 Gelatine, 177 E Efficiency: See also "Loading machines" Determining maximum, 66 Hand and machine loading com- pared, 117, 120 Mining machines, 102 of daymen, materials and equip- ment, 60 of various companies in Central Pennsylvania, 145 Outside handling systems, 150 Per capita production and percent- age of machine mined coal, 98 Production per employee, 98, 148, 152, 154 Shoveling capacity of miners, 156 Electric lamps. See "Lighting" Electrical shotfiring. See "Shot- firing" Elizabeth tunnel on the Los Angeles Aqueduct, 422 Employees. See "Work," "Load- ing," "Efficiency," etc. Endless rope haulage, 307 Engineering, effects of in obtaining a maximum recovery, 171 Engines for driving underground conveyors, 95 English shaft sinking methods, 203 Equipment for shaft sinking. See "Shaft sinking equipment" 506 INDEX European shaft sinking methods : Belgium, 203 English, 203 Evans scraper loading apparatus, 119 Excavating for shafts. See ''Shaft sinking" Exhaustion of coal reserves. See " Conservation " Exploders used by the Utah Fuel Co., 137 Explosives: See also "Blasting" Exploders "Reliable," 137 for tunneling, 440 Giant powder used by the Utah Fuel Co., 137 Permissible, 130 Production and amount used in the U. S., 130 Fans, ventilating, 443 Anemometer tests of, 444 Efficiency of, 443 Manometric efficiency of, 444 Mechanical efficiency of, 444, 452 Power required for, 446 Specifications for, 448 Volumetric efficiency of, 444 Water gage requirements, 450 Fluctuations in tonnage as effecting costs, 22, 32, 34, 39, 161, 162 Forces in blasting, 127 Frick Co.: Shafts at Brownsville, Pa., 192 Systems of mining, 82 Frictional resistance. See "Resist- ance" Frogs and switches, 291 Cost of laying, 293 Spacing of ties, 294 Fuel Administration: Costs on anthracite, December, 1917, to May, 1918, 20 Costs, prices fixed and tonnages, August 12, 1918, 5, 8 Costs, prices fixed and tonnages, full year, 1918, 5 Fuel Administration: Costs, prices fixed and tonnages, previous to November, 1917, 7 Prices fixed on anthracite, Decem- ber 31, 1918, 20 Reported costs, prices fixed and tonnages for the full year, 1918, 13 Fuel and power costs at mines com- puted by the U. S. Census Bureau for 1909, 1 Future worth of coal properties, 29, 51 Gary, W. Va., percentages of re- covery at, 80 Gathering locomotives, 251 See also "Locomotives" Georges Creek field: Systems of mining, 39 Percentages of recovery, 168 Germany : Costs of mining and distribution of revenue, 31 Production per employee, 152 Shapes of shafts, 185 Grade, track: Effect of, on locomotives, 229 Weight transfer due to, 229 See also "Track" Gravity planes, 307 Great Britain: See also "England" Production per employee compared with the U. S., 154 Shaft sinking methods, 203 Grounding losses. See "Line costs and losses" Guncotton, 130 Hand drilling. See "Drilling" Hand shoveling compared with ma- chine loading, 117, 120 Haul, increasing entry driven to ob- tain shortest, 220 Back haul, 222 INDEX 507 Haulage, 219 Acceleration, 227 Animal, compressed air and electric haulage costs compared, 318 Animal haulage costs, 318, 319, 331, 334, 346 Capacity of locomotives, 224 Comparative cost of different sys- tems, 314 Compressed air locomotive costs, 317, 319, 331, 362 Compressed air single- and two- stage locomotives, 327 Compressed air and animal haulage costs compared, 331 Compressed air and electric haul- age costs compared, 362 Electric locomotive costs, 315, 319, 334, 357, 362 Electric locomotiv-e and animal haulage costs compared, 334 Gasoline locomotive costs, 316, 346 Gasoline locomotive and animal haulage costs compared, 346 Gathering methods, 58, 156 Gravity planes, 307 Number of cars per trip, 286 Preliminary considerations, 219 Rope, 306 Rope haulage costs, 317 Storage battery locomotive costs, 316, 345; 357 Storage battery and trolley loco- motive costs, 357 Horse haulage costs . See ' ' Haulage ' ' Horsepower of mine locomotives, 224 Horsepower rating, 231 Hydraulic cartridges. See also "Car- tridges," 131 Illinois: Costs in the No. 6 District, 1916 to 1920, 24 Efficiency of outside plants, 150 Increase in costs, 1916 to 1920, 27 Systems of mining and percentages of recovery, 166 Illinois: U. S. Census report of wages, sala- ries, royalties, etc., for 1909, 5 Inclined shafts and slopes. See " Shafts" and " Shaft sinking" Increases in costs, 1916 to 1920, 27 Indiana : Costs for years 1916 to 1920, 24 Increase in costs 1916 to 1920, 27 U. S. Census report of wages, sala- ries, royalties, etc., in 1909, 5 Ingersoll-Rand Mining and loading machine, 122 Inside men. See "Daymen," "Labor costs," "Work" Interest charges, increase of, due to intermittent work, 163 Intermittent work: at mines compared with other in- dustries, 162 Comparison of shifts worked at Middlewestern and Westphalia mines, 34 Increase in costs due to, 161 Influence of on costs, 22, 39 Losses to miners due to, 32, 157 Peabody, F. S., on, 162 Jeffrey-Drennan adjustable-turret cutting machine, 111 K Kentucky, machine mined coal in, 98 Labor, effects of union rules in obtain- ing maximum recovery, 172 Labor costs: See also "Daymen" Average proportion of cost, 139 Central Pennsylvania, 1916 to 1920, 26 Illinois No. 6 District, 1916 to 1920, 24 Indiana, 1916 to 1920, 24 508 INDEX Labor costs : Labor costs at 454 mines, 140 Maximum labor effort as developed in tests by the U. S. Coal & Coke Co., 145 Ohio No. 8 District, 1916 to 1920, 25 on anthracite, December, 1917, to May, 1918, 20 on anthracite, 1913 to 1916, 23 Pocahontas field, 1916 to 1920, 23 Production per dayman, 141, 142, 145 Production per employee, 148 Ratio to total, 21 Southwestern Pennsylvania, 1916 to 1920, 23 Variations in, with capacity of mine, 142 Lamps, miners: See also "Lighting" Acetylene and oil, 470 Cables for, 467 Charges to miners for, 469 Electric bulbs for, 466 Maintenance costs, 461, 470 Manlite, 460 Oil lamps, 471 Plant cost for installing electric lamps, 471 Lighting, mine, 456 Acetylene and oil lamps, 470 Bureau of Mines standards, 458 Charges to miners for, 469 Lumens, 456 Measurement of, 456 Oil lamps, 471 Portable lamps, 459 Line costs and losses, 246, 264 Cost of feed wire, 265 Cost of line construction, 267 Drop in voltage, 264 Grounding losses, 266 Linings, shaft. See " Shaft linings " Loading : See also "Loading machines" Hand and machine loading com- pared, 117 Loading: Maximum effort in hand loading developed at tests by the U. S. Coal & Coke Co., 145 Number of tons loaded per shift by miners in the Connellsville Dis- trict, 85 Shoveling capacity, 156 Loading and mining machines, 122 See also "Mining and loading ma- chines" Loading machines, 116 Capacity of, 118 Evans scraper, 119 Hand shoveling compared with, 117, 120 Jeffrey loader, 117, 118 Layout of mine for machine load- ing, 117 Meyers- Whaley loader, 120 Power required for, 120 Westmoreland loader, 120 Locomotives: Adhesion of cast-iron and steel wheels on, 228 Ampere rating of, 233 Buying, 232 Checking the work of, 253 Commutating and non-commutat- ing pole motors, 256 Comparison of one hour and all day rating, 234 Computing size of, 223, 236 Costs of, 252 Curves for a 40-hp. locomotive, 235 Curves for a 50-hp. locomotive, 240 Effect of long feed lines on, 263 Effective wattage, 224 Haulage capacity of, 224 Haulage costs with. See "Haul- age" Horsepower of, 224, 233, 245 Height limits of, 248 Large size, 250 Life of, 250 Motor losses on, 253 INDEX 509 Locomotives : Number and size required, 233, 249 One hour rating of, 233 Power costs for, 256, 260 Resistance, 227 Roundtrip performance of, 261 Speeds, 224 Tractive effort drawbar-pull and rating of, 222, 224, 235, 245 Tractive effort per inch of height, 248 Longwall, use of, in conservation, 173 Loss: Conditions where operations may be conducted at, and apparent, 33 Mining unprofitable seams, 93 Lump coal, increase of, with mining machines, 112 M Machines : Loading. See " Loading ma- chines" Mining. See "Mining machines" Mining and loading. See " Mining and loading machines" Management, effect of, in obtaining maximum recovery, 171 Mexico, details of sinking a 679-ft. shaft in, 198 Michigan, systems of mining and per- centages of recovery in, 166 Middlewestern mines: Comparison of costs and distribu- tion of revenue compared with German mines, 31 Comparison of prices realized with those of Westphalia, Ger- many, 34 Progress in shaft sinking, 213 Miners : See also "Loading machines" and " Daymen" Earnings of. See "Wages" Hand and machine loading com- pared, 117, 120 Miners: Maximum loadings of, in tests by the U. S. Coal & Coke Co., 145 Number of tons loaded per shift in the Connellsville region, 85 Shoveling capacity of, 156 Tons mined per employee, 98, 148, 152, 154 Mining and loading machines, 122 Capacity of, 123 Ingersoll-Rand type, 122 Jeffrey type, 123 O'Toole type, 124 Saving of timbering with the use of machines, 124 Mining machines: Alternating current for, 109 Arc wall cutters, 111 Bits for. See also "Bits," 102, 115 Bonus system in operating, 107 Buying machines, 101 Capacity of, 107, 111, 112 Care of, 102, 115 Charges against, 100 Considerations affecting their ad- aptability, 99 Cost of, 99 Cutting machines, 97 Economies of, 97 Hand and machine operating condi- tions compared, 101, 104, 105 Installation costs, 103 Jeffrey-Drennan adjustable-turret machines, 111 Labor and mechanical power com- pared, 96 Loading machines. See "Loading machines" Lump coal proportion with ma- chines, 112 Maintenance costs, 107 Maintenance costs for alternating and direct current installations, 111 Obtaining efficiency from, 102 Operating costs, 103 Post punching machines, 111 Power for, 97, 103, 109 r 510 INDEX Mining machines: Proportion of coal cut by, 97, 99 Repair costs, 112, 115 Repair reports, 113 Shortwall used in the Connellsville district, 81 Small capacity mines, high costs at, 142 Sullivan Machinery Co. "Dread- naught" chain, 115 Supplies for, accounting of, 114 Undercutting, depth of, 66, 108 Mining, systems, 39 Concentration method used at Gary, W. Va., 48 Connellsville district, 80 Continental mine No. 2 (Connells- ville district), 86 Continuous panel, 52 Conveyor system, 93 Crozer Land Association, 72 Driving rooms . See ' ' Rooms ' ' Frick Co. methods, 82 Georges Creek field, 39 in various fields with special refer- ence to percentages of recovery, 165 Laying out in 90-ft. square blocks, 87 Layout for machine loading, 117 Layout for alternating current mine, 109 Pocahontas field, 72 Square or rectangular panel, 52 Upland Coal & Coke Co., 72 Working thick soft seams, 45 Missed shots. See "Blasting" Mule haulage costs. See "Haulage" Mules: Cost and care of, 338 Depreciation of, 340 Work of, 340 Meyers- Whaley loading machine, 120 N New York Aqueduct, progress in shaft sinking at, 213 Nitroglycerin, 129 Ohio: Costs in the No. 8 District, 1916 to 1920, 25 Increase in costs, 1916 to 1920, 27 Percentage of recovery in, 169 Oil: Cost of for mine cars, 302 Illuminating, 471 Outside men. See "Daymen," "Labor costs," "Work" Outside plants, the most efficient, 150 Outputs : Comparison of shifts worked at Middlewestern and Westphalia (Germany) mines, 34 Estimating, from a mine section, 90 Influence of fluctuations in, on costs, 22, 32, 39, 161, 164 Overcasts, 488, 492 Pacific Coast Coal Co., post punchers at, 111 Panel system of mining: Continuous, 52 Square or rectangular, 52 Pennsylvania, U. S. Census report on anthracite, 1 Pennsylvania bituminous: Costs in the Central District, 1916 to 1920, 26 Costs in Southwestern District, 1916 to 1920, 23 Increase in Central District costs, 1916 to 1920, 27 Increase in Southwestern District costs, 1916 to 1920, 27 Systems of mining and percentages of recovery, 167 Percentages of recovery. See "Con- servation" and "Drawing pillars" Permissible explosives. See also "Explosives" Pillars: Barrier, 56, 60, 69 Size of. See various systems of mining INDEX 511 Pillars, drawing: See also various systems of mining Concentration method used at Gary, W. Va., 49 Georges Creek field, 47 Maintenance of breakline, 77 Methods in the Connellsville re- gion, 88 Percentage of recovery at the mines of the Pocahontas Coal & Coke Co., 79 Percentages of recovery at Gary, W. Va., 80 Pocahontas field, 74, 78 Use of mining machines in, 77 Pocahontas Coal & Coke Co., sys- tem of mining, 72 Pocahontas field: Methods of drawing pillars, 74 Methods of working in, 72 Shaft water in, 187 Portable lamps . See ' ' Lighting ' ' Post punching machines, 111 Power : Comparison of alternating and direct current for operating ma- chines, 109 Costs of for operating locomotives. See "Locomotives" Plant costs by various units, 261 Power and fuel, costs of reported by the U.S. Census Bureau of 1909, 1 Present and future worth of coal properties, 29 Prices: Anthracite, fixed by the president in 1917, 14, 19 Comparison of, at Middlewestern and Westphalia mines, 34 Fixed by the Fuel Administration. See under that head Method of fixing used by the West- phalian Coal Syndicate, 32 Production: Comparison of, at Westphalia (Germany) and Middlewestern mines, 33 Influence of, on costs, 22, 39 Profits: as affecting conservation, 172 Increase of, with increased output, 164 U. S. Census report on, for 1909, 4 Profit and loss, conditions where operations may be conducted at an apparent loss, 33 Pumping. See " Shaft sinking, pump- ing" Punching machines : See also "Mining machines" Cutting and punching machines compared, 106 Post punching machines, 111 Pulling pillars. See "Pillars" R Rails, track, 287 Durability, 289 Electrical resistance of. See "Lo- comotives" Frogs. See "Frogs" Minimum weight, 287 Purchasing, 287 Seconds, 290 Stiffness, 288 Strength, 288 Recovery of coal : See also "Conservation" Anthracite coal waste, Commission on, 164 Factors governing recovery in, various fields, 168, 170 Percentage of, in various fields, 165 Percentage of, at mines of the Pocahontas Coal & Coke Co., 79 Records of, 79 Use of longwall to effect maximum, 173 Refuge chambers, 498 Report form for shaft sinking, 215 Reports for repair costs to mining machines, 113 Resistance, track and grade, 223, 231 of cars, 224, 227, 230 Locomotives, 227 512 INDEX Resultant forces in blasting, 127 Revenue distribution at German and Westphalia mines, 31 Robbing pillars. See " Pillars" Rooms : See also various systems of mining Depth and number of, 78 Methods of driving commonly used, 49 Number of men in, 50 Placing track in, 49 Room space per miner, 67 Safety of, 51 Systematizing wor.i in, 85 Roller bearings . See ' ' Bearings ' ' Rope haulage, 306 Rope, wire, 309 Lubrication, 310 Royalties : on anthracite, 16, 20 U. S. Census report on, for anthra- cite mines in 1909, 1 U. S. Census report on, for bitu- minous mines in 1909, 3 U. S. Census report on, for Indiana and Illinois mines in 1909, 5 S Salaries: U. S. Census report on, for anthra- cite mines in 1909, 1 U. S. Census report on, for bitu- minous mines in 1909, 3 U. S. Census report on, for Indiana and Illinois mines in 1909, 5 Charges for, at anthracite mines, December, 1917, to May, 1918, 20 Scales, wage. See "Wage scale" Scraper loading machine, Evans, 119 Screenings : Breakage with hydraulic cartridge, 132 Proportion of, with machine min- ing, 112 Seams, costs of mining different thick- nesses of, 22, 91 Shafts: Circular, 179, 181, 185, 208 Elliptical, 178, 182, 183, 185, 192 Hydrostatic pressure in, 181 Inclined, comparison of cost with vertical, 189 Preliminary considerations for, 176 Prospect, 203 Quadrilateral, 182, 185 Rectangular, 177, 208 Shapes, 177, 208 Weep holes in, 181 Shaft linings, 204 Brick, 203 Comparison of different shapes, 182 Concrete, 181, 192, 195, 203, 208 Concrete blocks, 207 Concrete shaft for the River Coal Co., 206 Curtain wall, 193 Grouting, 181 Hydrostatic pressure against, 181 Timber, 201, 208 Weep holes in, 181 Shaft sinking, 176 Blasting, 177, 193 Caisson, 192 Circular, 179, 208 Contract form for, 217 Drainage. See "Pumping" Drilling, 176, 188 Elliptical, 178 Equipment for. See "Shaft sink- ing equipment" Excavation, 182, 184, 191, 194, 199, 205 Grouting, 181 Hand drilling, 177, 189, 215 Hydrostatic pressure in, 181 in quicksand, 191 in soft material, 180, 188 Inclined, comparison of cost with vertical, 189 Operation, 176, 193, 204 Report form for, 215 South Africa, 180 through surface material, 191 Wages in, 180 INDEX 513 Shaft sinking costs, 188, 191, 213 Clay and gravel, 191 Comparison of circular and rectan- gular, 208 Costs in Michigan, 200 in Great Britain, 203 in quicksand, 191 in soft material, 188 Inclined and vertical compared, 189, 200 Nokomis Coal Co. (Illinois), 188 per cubic foot and cubic yard, 191 Power costs, 197 Prospect, 203 Sinking 570-ft. shaft in Colorado, 198 Sinking 679-ft. shaft in Mexico, 198 Shaft sinking equipment, 186, 196, 202, 211 Boilers, 187 Compressed air plant for, 186, 196 Dump cars, 196 for sinking a shaft in Michigan, 200 for sinking a prospect shaft, 203 Headframe, 187 Hoist, 187 Plant for sinking a single 500-ft. shaft, 187 Pumps, 187 Single- and two-stage compressors compared, 186 Shaft sinking, progress, 194, 197, 205, 213, 209 Computed progress in Michigan, 200 Effects of large volumes of water on, 187 in Great Britain, 203 Middlewestern, 213 New York Aqueduct, 213 Report forms for, 215 Various American and South Afri- can shafts, 212 Shaft sinking, pumping, 187 Shifts worked: See also "Fluctuations in tonnage," "Intermittent work," "Out- puts" Shifts worked: at mines, compared with other in- dustries, 162 Comparison of, at Middlewestern and Westphalia mines, 34 in shaft sinking, 180 Shooting coal. See "Blasting" Shot firing, electrical: See also "Blasting" Cost of installing, 138 Power required for, 137 Safety of, 139 Wiring for, 136 Shoveling capacity. See "Effici- ency," " Loading, " " Miners, ' ' "Work" Sinking shafts. See "Shaft sinking" Sizes of anthracite, percentage of, 15 South Africa: Shaft sinking practice, 211 Wltwatersrand, 213 Speed of development work, 53, 95 Speed of shaft sinking. See "Shaft sinking, progress" Square panel used at Gary, W. Va., 52 Stables, underground, 475 Steel mine cars. See "Cars, mine" Steel for timbering. See "Timber- ing" Stoppings, 488 Storage battery locomotives. See "Haulage," "Locomotives" Sprinkling costs, 499 Stretcher car, 303 Switches, track. See "Frogs and switches" Systems of mining: Concentration method used at Gary, W. Va., 48 Connellsville district, 80 Continental Mine No. 2 (Connells- ville district), 86 Continuous panel, 52, 60 Crozer Land Association, 72 Driving rooms. See "Rooms" for working thick seams, 45 Frick Co., 82 Georges Creek field, 39 514 INDEX Systems of mining: Laying out in 90-ft. square blocks, 87 Layout for an alternating current mine, 109 Layout for machine loading, 117 Longwall used to affect conserva- tion, 173 Pocahontas field, 72 Square or rectangular panel, 52, 60 Upland Coal & Coke Co., 72 Taxes, computing returns for, 28 Thickness of seam, influence on costs, 22, 91 Ties, track: Comparison of steel and wood, 295 Life of steel ties, 296 Spacing of, for switches, 294 Timber: Computing size of, 370 Concrete for timbering, 393 Costs and amounts used in the U. S., 367 Framing equipment and costs, 376 How to buy, 365 Kinds used, 369 Knots, effects of, on strength, 374 Preservatives, 379, 382 Reclaiming, 404 Round and sawed, 369 Seasoning, 380 Timbering : Cement gun used for, 402 Computing sizes for, 370 Costs, 365 Linings for shafts. See " Shaft linings" Maintenance cost of wood and steel, 390 Rails for, 374 Savings in with mining and loading machines, 124 Time required to reach a certain out- put, 53, 95 Tonnage, influence of, on costs, 22 Track: Costs, 278, 294 Cost of grade revisions, 283 Curvature of, in relation to wheel base, 232 Curves. 285 Frogs. See under that title Grade revision, 280 Grades of, 278, 280 Methods of placing, in rooms, 49 Rails. See under that title Resistance, 223, 279, 298 Ties. See under that title Tractive effort of mine locomotives, 222, 231 Tunneling costs, 406, 419 American and foreign records com- pared, 406 Bonus system applied to, 412 Elizabeth tunnel on the Los An- geles aqueduct, 422 Explosives for, 440 Gunnison and Simplon tunneling records, 406 Utah Fuel Co. methods, 413 Turnouts. See " Frogs and switches" U Undercutting, depth of, 108 See also "Machine mining" U. S. Census report on production, costs, wages, salaries, 1 Upland Coal & Coke Co., system of mining, 72 Value: Developed and undeveloped prop- erties, 28 Present and future worth, 29, 51 Ventilation : See also various systems of mining Air required, 453 Canvas tubing for, 455 Connellsville district, 81 Consolidation Coal Co. 's methods, 450 Costs, 443 INDEX 515 Ventilation: Fans. See under that title Water gage readings, 450 W Wage, ratio to value of product, 5, 21, 139 Wage scales: Hocking district, 1898 to 1921, 158 How fixed, 22 Wages: Causes for inadequacy, 155 Inequalities of, 155 Mine car supply, effects of, on wages, 156 Miners, 153 on shaft sinking, 180 U. S. Census report on, for anthra- cite mines in 1909, 1 U. S. Census report on, for bitu- minous mines in 1909, 3 U. S. Census report on, for Indiana and Illinois mines in 1909, 5 Westmoreland coal loading machine, 119 Westphalian Coal Syndicate (Ger- many) : Costs and revenue compared with Middlewestern mines, 31 Increase in production, 1850 to 1907, 33 Westphalian Coal Syndicate (Ger- many) : Investment per ton of production, 33 Price fixing at, 32 Prices compared with Middle- western mines, 34 W T heels, mine car, 303 Wheels, adhesive characteristics of cast iron and steel for mine loco- motives, 228 Wheel base and radius of curve, 232 Wire rope, 309 Lubrication of, 310 Work: See also "Loading machines," " Day men" Hand and machine loading com- pared, 117, 120 Intermittent work and its effect on costs, 161 Maximum labor effort developed in tests by the U. S. Coal and Coke Co., 145 Number of tons loaded per shift in the Connellsville district, 85 Shoveling capacity, 156 Tons mined per employee, 98, 148, 153 World, production per employee in various fields of, 152 UNIVERSITY OF CALIFORNIA LIBRARY