Fluctuation in the price of base metals is a factor so much to be taken into consideration, that it is desirable in estimating mine values to reduce the working costs to a basis of a "per unit" of finished metal. This method has the great advantage of indicating so simply the involved risks of changing prices that whoso runs may read. Where one metal predominates over the other to such an extent as to form the "backbone" of the value of the mine, the value of the subsidiary metals is often deducted from the cost of the principal metal, in order to indicate more plainly the varying value of the mine with the fluctuating prices of the predominant metal. For example, it is usual to state that the cost of copper production from a given ore will be so many cents per pound, or so many pounds sterling per ton. Knowing the total metal extractable from the ore in sight, the profits at given prices of metal can be readily deduced. The point at which such calculation departs from the "per-ton-of-ore" unto the per-unit-cost-of-metal basis, usually lies at the point in ore dressing where it is ready for the smelter. To take a simple case of a lead ore averaging 20%: this is to be first concentrated and the lead reduced to a concentrate averaging 70% and showing a recovery of 75% of the total metal content. The cost per ton of development, mining, concentration, management, is to this point say $4 per ton of original crude ore. The smelter buys the concentrate for 95% of the value of the metal, less the smelting charge of $15 per ton, or there is a working cost of a similar sum by home equipment. In this case 4.66 tons of ore are required to produce one ton of concentrates, and therefore each ton of concentrates costs $18.64. This amount, added to the smelting charge, gives a total of $33.64 for the creation of 70% of one ton of finished lead, or equal to 2.40 cents per pound which can be compared with the market price less 5%. If the ore were to contain 20 ounces of silver per ton, of which 15 ounces were recovered into the leady concentrates, and the smelter price for the silver were 50 cents per ounce, then the $7.50 thus recovered would be subtracted from $33.64, making the apparent cost of the lead 1.86 cents per pound.



CHAPTER V.

Mine Valuation (Continued).

REDEMPTION OR AMORTIZATION OF CAPITAL AND INTEREST.

It is desirable to state in some detail the theory of amortization before consideration of its application in mine valuation.

As every mine has a limited life, the capital invested in it must be redeemed during the life of the mine. It is not sufficient that there be a bare profit over working costs. In this particular, mines differ wholly from many other types of investment, such as railways. In the latter, if proper appropriation is made for maintenance, the total income to the investor can be considered as interest or profit; but in mines, a portion of the annual income must be considered as a return of capital. Therefore, before the yield on a mine investment can be determined, a portion of the annual earnings must be set aside in such a manner that when the mine is exhausted the original investment will have been restored. If we consider the date due for the return of the capital as the time when the mine is exhausted, we may consider the annual instalments as payments before the due date, and they can be put out at compound interest until the time for restoration arrives. If they be invested in safe securities at the usual rate of about 4%, the addition of this amount of compound interest will assist in the repayment of the capital at the due date, so that the annual contributions to a sinking fund need not themselves aggregate the total capital to be restored, but may be smaller by the deficiency which will be made up by their interest earnings. Such a system of redemption of capital is called "Amortization."

Obviously it is not sufficient for the mine investor that his capital shall have been restored, but there is required an excess earning over and above the necessities of this annual funding of capital. What rate of excess return the mine must yield is a matter of the risks in the venture and the demands of the investor. Mining business is one where 7% above provision for capital return is an absolute minimum demanded by the risks inherent in mines, even where the profit in sight gives warranty to the return of capital. Where the profit in sight (which is the only real guarantee in mine investment) is below the price of the investment, the annual return should increase in proportion. There are thus two distinct directions in which interest must be computed,—first, the internal influence of interest in the amortization of the capital, and second, the percentage return upon the whole investment after providing for capital return.

There are many limitations to the introduction of such refinements as interest calculations in mine valuation. It is a subject not easy to discuss with finality, for not only is the term of years unknown, but, of more importance, there are many factors of a highly speculative order to be considered in valuing. It may be said that a certain life is known in any case from the profit in sight, and that in calculating this profit a deduction should be made from the gross profit for loss of interest on it pending recovery. This is true, but as mines are seldom dealt with on the basis of profit in sight alone, and as the purchase price includes usually some proportion for extension in depth, an unknown factor is introduced which outweighs the known quantities. Therefore the application of the culminative effect of interest accumulations is much dependent upon the sort of mine under consideration. In most cases of uncertain continuity in depth it introduces a mathematical refinement not warranted by the speculative elements. For instance, in a mine where the whole value is dependent upon extension of the deposit beyond openings, and where an expected return of at least 50% per annum is required to warrant the risk, such refinement would be absurd. On the other hand, in a Witwatersrand gold mine, in gold and tin gravels, or in massive copper mines such as Bingham and Lake Superior, where at least some sort of life can be approximated, it becomes a most vital element in valuation.

In general it may be said that the lower the total annual return expected upon the capital invested, the greater does the amount demanded for amortization become in proportion to this total income, and therefore the greater need of its introduction in calculations. Especially is this so where the cost of equipment is large proportionately to the annual return. Further, it may be said that such calculations are of decreasing use with increasing proportion of speculative elements in the price of the mine. The risk of extension in depth, of the price of metal, etc., may so outweigh the comparatively minor factors here introduced as to render them useless of attention.

In the practical conduct of mines or mining companies, sinking funds for amortization of capital are never established. In the vast majority of mines of the class under discussion, the ultimate duration of life is unknown, and therefore there is no basis upon which to formulate such a definite financial policy even were it desired. Were it possible to arrive at the annual sum to be set aside, the stockholders of the mining type would prefer to do their own reinvestment. The purpose of these calculations does not lie in the application of amortization to administrative finance. It is nevertheless one of the touchstones in the valuation of certain mines or mining investments. That is, by a sort of inversion such calculations can be made to serve as a means to expose the amount of risk,—to furnish a yardstick for measuring the amount of risk in the very speculations of extension in depth and price of metals which attach to a mine. Given the annual income being received, or expected, the problem can be formulated into the determination of how many years it must be continued in order to amortize the investment and pay a given rate of profit. A certain length of life is evident from the ore in sight, which may be called the life in sight. If the term of years required to redeem the capital and pay an interest upon it is greater than the life in sight, then this extended life must come from extension in depth, or ore from other direction, or increased price of metals. If we then take the volume and profit on the ore as disclosed we can calculate the number of feet the deposit must extend in depth, or additional tonnage that must be obtained of the same grade, or the different prices of metal that must be secured, in order to satisfy the demanded term of years. These demands in actual measure of ore or feet or higher price can then be weighed against the geological and industrial probabilities.

The following tables and examples may be of assistance in these calculations.

Table 1. To apply this table, the amount of annual income or dividend and the term of years it will last must be known or estimated factors. It is then possible to determine the present value of this annual income after providing for amortization and interest on the investment at various rates given, by multiplying the annual income by the factor set out.

A simple illustration would be that of a mine earning a profit of $200,000 annually, and having a total of 1,000,000 tons in sight, yielding a profit of $2 a ton, or a total profit in sight of $2,000,000, thus recoverable in ten years. On a basis of a 7% return on the investment and amortization of capital (Table I), the factor is 6.52 x $200,000 = $1,304,000 as the present value of the gross profits exposed. That is, this sum of $1,304,000, if paid for the mine, would be repaid out of the profit in sight, together with 7% interest if the annual payments into sinking fund earn 4%.

TABLE I.

Present Value of an Annual Dividend Over — Years at —% and Replacing Capital by Reinvestment of an Annual Sum at 4%.

======================================================= Years 5% 6% 7% 8% 9% 10% - - - - - - - 1 .95 .94 .93 .92 .92 .91 2 1.85 1.82 1.78 1.75 1.72 1.69 3 2.70 2.63 2.56 2.50 2.44 2.38 4 3.50 3.38 3.27 3.17 3.07 2.98 5 4.26 4.09 3.93 3.78 3.64 3.51 6 4.98 4.74 4.53 4.33 4.15 3.99 7 5.66 5.36 5.09 4.84 4.62 4.41 8 6.31 5.93 5.60 5.30 5.04 4.79 9 6.92 6.47 6.08 5.73 5.42 5.14 10 7.50 6.98 6.52 6.12 5.77 5.45 11 8.05 7.45 6.94 6.49 6.09 5.74 12 8.58 7.90 7.32 6.82 6.39 6.00 13 9.08 8.32 7.68 7.13 6.66 6.24 14 9.55 8.72 8.02 7.42 6.91 6.46 15 10.00 9.09 8.34 7.79 7.14 6.67 16 10.43 9.45 8.63 7.95 7.36 6.86 17 10.85 9.78 8.91 8.18 7.56 7.03 18 11.24 10.10 9.17 8.40 7.75 7.19 19 11.61 10.40 9.42 8.61 7.93 7.34 20 11.96 10.68 9.65 8.80 8.09 7.49 21 12.30 10.95 9.87 8.99 8.24 7.62 22 12.62 11.21 10.08 9.16 8.39 7.74 23 12.93 11.45 10.28 9.32 8.52 7.85 24 13.23 11.68 10.46 9.47 8.65 7.96 25 13.51 11.90 10.64 9.61 8.77 8.06 26 13.78 12.11 10.80 9.75 8.88 8.16 27 14.04 12.31 10.96 9.88 8.99 8.25 28 14.28 12.50 11.11 10.00 9.09 8.33 29 14.52 12.68 11.25 10.11 9.18 8.41 30 14.74 12.85 11.38 10.22 9.27 8.49 31 14.96 13.01 11.51 10.32 9.36 8.56 32 15.16 13.17 11.63 10.42 9.44 8.62 33 15.36 13.31 11.75 10.51 9.51 8.69 34 15.55 13.46 11.86 10.60 9.59 8.75 35 15.73 13.59 11.96 10.67 9.65 8.80 36 15.90 13.72 12.06 10.76 9.72 8.86 37 16.07 13.84 12.16 10.84 9.78 8.91 38 16.22 13.96 12.25 10.91 9.84 8.96 39 16.38 14.07 12.34 10.98 9.89 9.00 40 16.52 14.18 12.42 11.05 9.95 9.05 ======================================================= Condensed from Inwood's Tables.

Table II is practically a compound discount table. That is, by it can be determined the present value of a fixed sum payable at the end of a given term of years, interest being discounted at various given rates. Its use may be illustrated by continuing the example preceding.

TABLE II.

Present Value of $1, or L1, payable in — Years, Interest taken at —%.

=================================== Years 4% 5% 6% 7% - 1 .961 .952 .943 .934 2 .924 .907 .890 .873 3 .889 .864 .840 .816 4 .854 .823 .792 .763 5 .821 .783 .747 .713 6 .790 .746 .705 .666 7 .760 .711 .665 .623 8 .731 .677 .627 .582 9 .702 .645 .592 .544 10 .675 .614 .558 .508 11 .649 .585 .527 .475 12 .625 .557 .497 .444 13 .600 .530 .469 .415 14 .577 .505 .442 .388 15 .555 .481 .417 .362 16 .534 .458 .394 .339 17 .513 .436 .371 .316 18 .494 .415 .350 .296 19 .475 .396 .330 .276 20 .456 .377 .311 .258 21 .439 .359 .294 .241 22 .422 .342 .277 .266 23 .406 .325 .262 .211 24 .390 .310 .247 .197 25 .375 .295 .233 .184 26 .361 .281 .220 .172 27 .347 .268 .207 .161 28 .333 .255 .196 .150 29 .321 .243 .184 .140 30 .308 .231 .174 .131 31 .296 .220 .164 .123 32 .285 .210 .155 .115 33 .274 .200 .146 .107 34 .263 .190 .138 .100 35 .253 .181 .130 .094 36 .244 .172 .123 .087 37 .234 .164 .116 .082 38 .225 .156 .109 .076 39 .216 .149 .103 .071 40 .208 .142 .097 .067 =================================== Condensed from Inwood's Tables.

If such a mine is not equipped, and it is assumed that $200,000 are required to equip the mine, and that two years are required for this equipment, the value of the ore in sight is still less, because of the further loss of interest in delay and the cost of equipment. In this case the present value of $1,304,000 in two years, interest at 7%, the factor is .87 X 1,304,000 = $1,134,480. From this comes off the cost of equipment, or $200,000, leaving $934,480 as the present value of the profit in sight. A further refinement could be added by calculating the interest chargeable against the $200,000 equipment cost up to the time of production.

TABLE III. =========================================================================== Annual Number of years of life required to yield % interest, and in Rate of addition to furnish annual instalments which, if reinvested at Dividend. 4% will return the original investment at the end of the period. - - % 5% 6% 7% 8% 9% 10% 6 41.0 7 28.0 41.0 8 21.6 28.0 41.0 9 17.7 21.6 28.0 41.0 10 15.0 17.7 21.6 28.0 41.0 11 13.0 15.0 17.7 21.6 28.0 41.0 12 11.5 13.0 15.0 17.7 21.6 28.0 13 10.3 11.5 13.0 15.0 17.7 21.6 14 9.4 10.3 11.5 13.0 15.0 17.7 15 8.6 9.4 10.3 11.5 13.0 15.0 16 7.9 8.6 9.4 10.3 11.5 13.0 17 7.3 7.9 8.6 9.4 10.3 11.5 18 6.8 7.3 7.9 8.6 9.4 10.3 19 6.4 6.8 7.3 7.9 8.6 9.4 20 6.0 6.4 6.8 7.3 7.9 8.6 21 5.7 6.0 6.4 6.8 7.3 7.9 22 5.4 5.7 6.0 6.4 6.8 7.3 23 5.1 5.4 5.7 6.0 6.4 6.8 24 4.9 5.1 5.4 5.7 6.0 6.4 25 4.7 4.9 5.1 5.4 5.7 6.0 26 4.5 4.7 4.9 5.1 5.4 5.7 27 4.3 4.5 4.7 4.9 5.1 5.4 28 4.1 4.3 4.5 4.7 4.9 5.1 29 3.9 4.1 4.3 4.5 4.7 4.9 30 3.8 3.9 4.1 4.3 4.5 4.7 ===========================================================================

Table III. This table is calculated by inversion of the factors in Table I, and is the most useful of all such tables, as it is a direct calculation of the number of years that a given rate of income on the investment must continue in order to amortize the capital (the annual sinking fund being placed at compound interest at 4%) and to repay various rates of interest on the investment. The application of this method in testing the value of dividend-paying shares is very helpful, especially in weighing the risks involved in the portion of the purchase or investment unsecured by the profit in sight. Given the annual percentage income on the investment from the dividends of the mine (or on a non-producing mine assuming a given rate of production and profit from the factors exposed), by reference to the table the number of years can be seen in which this percentage must continue in order to amortize the investment and pay various rates of interest on it. As said before, the ore in sight at a given rate of exhaustion can be reduced to terms of life in sight. This certain period deducted from the total term of years required gives the life which must be provided by further discovery of ore, and this can be reduced to tons or feet of extension of given ore-bodies and a tangible position arrived at. The test can be applied in this manner to the various prices which must be realized from the base metal in sight to warrant the price.

Taking the last example and assuming that the mine is equipped, and that the price is $2,000,000, the yearly return on the price is 10%. If it is desired besides amortizing or redeeming the capital to secure a return of 7% on the investment, it will be seen by reference to the table that there will be required a life of 21.6 years. As the life visible in the ore in sight is ten years, then the extensions in depth must produce ore for 11.6 years longer—1,160,000 tons. If the ore-body is 1,000 feet long and 13 feet wide, it will furnish of gold ore 1,000 tons per foot of depth; hence the ore-body must extend 1,160 feet deeper to justify the price. Mines are seldom so simple a proposition as this example. There are usually probabilities of other ore; and in the case of base metal, then variability of price and other elements must be counted. However, once the extension in depth which is necessary is determined for various assumptions of metal value, there is something tangible to consider and to weigh with the five geological weights set out in Chapter III.

The example given can be expanded to indicate not only the importance of interest and redemption in the long extension in depth required, but a matter discussed from another point of view under "Ratio of Output." If the plant on this mine were doubled and the earnings increased to 20% ($400,000 per annum) (disregarding the reduction in working expenses that must follow expansion of equipment), it will be found that the life required to repay the purchase money,—$2,000,000,—and 7% interest upon it, is about 6.8 years.

As at this increased rate of production there is in the ore in sight a life of five years, the extension in depth must be depended upon for 1.8 years, or only 360,000 tons,—that is, 360 feet of extension. Similarly, the present value of the ore in sight is $268,000 greater if the mine be given double the equipment, for thus the idle money locked in the ore is brought into the interest market at an earlier date. Against this increased profit must be weighed the increased cost of equipment. The value of low grade mines, especially, is very much a factor of the volume of output contemplated.



CHAPTER VI.

Mine Valuation (Concluded).

VALUATION OF MINES WITH LITTLE OR NO ORE IN SIGHT; VALUATIONS ON SECOND-HAND DATA; GENERAL CONDUCT OF EXAMINATIONS; REPORTS.

A large number of examinations arise upon prospecting ventures or partially developed mines where the value is almost wholly prospective. The risks in such enterprises amount to the possible loss of the whole investment, and the possible returns must consequently be commensurate. Such business is therefore necessarily highly speculative, but not unjustifiable, as the whole history of the industry attests; but this makes the matter no easier for the mine valuer. Many devices of financial procedure assist in the limitation of the sum risked, and offer a middle course to the investor between purchase of a wholly prospective value and the loss of a possible opportunity to profit by it. The usual form is an option to buy the property after a period which permits a certain amount of development work by the purchaser before final decision as to purchase.

Aside from young mines such enterprises often arise from the possibility of lateral extension of the ore-deposit outside the boundaries of the property of original discovery (Fig. 3), in which cases there is often no visible ore within the property under consideration upon which to found opinion. In regions where vertical side lines obtain, there is always the possibility of a "deep level" in inclined deposits. Therefore the ground surrounding known deposits has a certain speculative value, upon which engineers are often called to pass judgment. Except in such unusual occurrences as South African bankets, or Lake Superior coppers, prospecting for deep level of extension is also a highly speculative phase of mining.

The whole basis of opinion in both classes of ventures must be the few geological weights,—the geology of the property and the district, the development of surrounding mines, etc. In any event, there is a very great percentage of risk, and the profit to be gained by success must be, proportionally to the expenditure involved, very large. It is no case for calculating amortization and other refinements. It is one where several hundreds or thousands of per cent hoped for on the investment is the only justification.

OPINIONS AND VALUATIONS UPON SECOND-HAND DATA.

Some one may come forward and deprecate the bare suggestion of an engineer's offering an opinion when he cannot have proper first-hand data. But in these days we have to deal with conditions as well as theories of professional ethics. The growing ownership of mines by companies, that is by corporations composed of many individuals, and with their stocks often dealt in on the public exchanges, has resulted in holders whose interest is not large enough to warrant their undertaking the cost of exhaustive examinations. The system has produced an increasing class of mining speculators and investors who are finding and supplying the enormous sums required to work our mines,—sums beyond the reach of the old-class single-handed mining men. Every year the mining investors of the new order are coming more and more to the engineer for advice, and they should be encouraged, because such counsel can be given within limits, and these limits tend to place the industry upon a sounder footing of ownership. As was said before, the lamb can be in a measure protected. The engineer's interest is to protect him, so that the industry which concerns his own life-work may be in honorable repute, and that capital may be readily forthcoming for its expansion. Moreover, by constant advice to the investor as to what constitutes a properly presented and managed project, the arrangement of such proper presentation and management will tend to become an a priori function of the promoter.

Sometimes the engineer can make a short visit to the mine for data purposes,—more often he cannot. In the former case, he can resolve for himself an approximation upon all the factors bearing on value, except the quality of the ore. For this, aside from inspection of the ore itself, a look at the plans is usually enlightening. A longitudinal section of the mine showing a continuous shortening of the stopes with each succeeding level carries its own interpretation. In the main, the current record of past production and estimates of the management as to ore-reserves, etc., can be accepted in ratio to the confidence that can be placed in the men who present them. It then becomes a case of judgment of men and things, and here no rule applies.

Advice must often be given upon data alone, without inspection of the mine. Most mining data present internal evidence as to credibility. The untrustworthy and inexperienced betray themselves in their every written production. Assuming the reliability of data, the methods already discussed for weighing the ultimate value of the property can be applied. It would be possible to cite hundreds of examples of valuation based upon second-hand data. Three will, however, sufficiently illustrate. First, the R mine at Johannesburg. With the regularity of this deposit, the development done, and a study of the workings on the neighboring mines and in deeper ground, it is a not unfair assumption that the reefs will maintain size and value throughout the area. The management is sound, and all the data are given in the best manner. The life of the mine is estimated at six years, with some probabilities of further ore from low-grade sections. The annual earnings available for dividends are at the rate of about L450,000 per annum. The capital is L440,000 in L1 shares. By reference to the table on page 46 it will be seen that the present value of L450,000 spread over six years to return capital at the end of that period, and give 7% dividends in the meantime, is 4.53 x L450,000 = L2,036,500 / 440,000 = L4 12s. 7d. per share. So that this mine, on the assumption of continuity of values, will pay about 7% and return the price. Seven per cent is, however, not deemed an adequate return for the risks of labor troubles, faults, dykes, or poor patches. On a 9% basis, the mine is worth about L4 4s. per share.

Second, the G mine in Nevada. It has a capital of $10,000,000 in $1 shares, standing in the market at 50 cents each. The reserves are 250,000 tons, yielding a profit for yearly division of $7 per ton. It has an annual capacity of about 100,000 tons, or $700,000 net profit, equal to 14% on the market value. In order to repay the capital value of $5,000,000 and 8% per annum, it will need a life of (Table III) 13 years, of which 2-1/2 are visible. The size of the ore-bodies indicates a yield of about 1,100 tons per foot of depth. At an exhaustion rate of 100,000 tons per annum, the mine would need to extend to a depth of over a thousand feet below the present bottom. There is always a possibility of finding parallel bodies or larger volumes in depth, but it would be a sanguine engineer indeed who would recommend the stock, even though it pays an apparent 14%.

Third, the B mine, with a capital of $10,000,000 in 2,000,000 shares of $5 each. The promoters state that the mine is in the slopes of the Andes in Peru; that there are 6,000,000 tons of "ore blocked out"; that two assays by the assayers of the Bank of England average 9% copper; that the copper can be produced at five cents per pound; that there is thus a profit of $10,000,000 in sight. The evidences are wholly incompetent. It is a gamble on statements of persons who have not the remotest idea of sound mining.

GENERAL CONDUCT OF EXAMINATION.

Complete and exhaustive examination, entailing extensive sampling, assaying, and metallurgical tests, is very expensive and requires time. An unfavorable report usually means to the employer absolute loss of the engineer's fee and expenses. It becomes then the initial duty of the latter to determine at once, by the general conditions surrounding the property, how far the expenditure for exhaustive examination is warranted. There is usually named a money valuation for the property, and thus a peg is afforded upon which to hang conclusions. Very often collateral factors with a preliminary sampling, or indeed no sampling at all, will determine the whole business. In fact, it is becoming very common to send younger engineers to report as to whether exhaustive examination by more expensive men is justified.

In the course of such preliminary inspection, the ore-bodies may prove to be too small to insure adequate yield on the price, even assuming continuity in depth and represented value. They may be so difficult to mine as to make costs prohibitive, or they may show strong signs of "petering out." The ore may present visible metallurgical difficulties which make it unprofitable in any event. A gold ore may contain copper or arsenic, so as to debar cyanidation, where this process is the only hope of sufficiently moderate costs. A lead ore may be an amorphous compound with zinc, and successful concentration or smelting without great penalties may be precluded. A copper ore may carry a great excess of silica and be at the same time unconcentratable, and there may be no base mineral supply available for smelting mixture. The mine may be so small or so isolated that the cost of equipment will never be justified. Some of these conditions may be determined as unsurmountable, assuming a given value for the ore, and may warrant the rejection of the mine at the price set.

It is a disagreeable thing to have a disappointed promoter heap vituperation on an engineer's head because he did not make an exhaustive examination. Although it is generally desirable to do some sampling to give assurance to both purchaser and vendor of conscientiousness, a little courage of conviction, when this is rightly and adequately grounded, usually brings its own reward.

Supposing, however, that conditions are right and that the mine is worth the price, subject to confirmation of values, the determination of these cannot be undertaken unless time and money are available for the work. As was said, a sampling campaign is expensive, and takes time, and no engineer has the moral right to undertake an examination unless both facilities are afforded. Curtailment is unjust, both to himself and to his employer.

How much time and outlay are required to properly sample a mine is obviously a question of its size, and the character of its ore. An engineer and one principal assistant can conduct two sampling parties. In hard rock it may be impossible to take more than five samples a day for each party. But, in average ore, ten samples for each is reasonable work. As the number of samples is dependent upon the footage of openings on the deposit, a rough approximation can be made in advance, and a general idea obtained as to the time required. This period must be insisted upon.

REPORTS.

Reports are to be read by the layman, and their first qualities should be simplicity of terms and definiteness of conclusions. Reports are usually too long, rather than too short. The essential facts governing the value of a mine can be expressed on one sheet of paper. It is always desirable, however, that the groundwork data and the manner of their determination should be set out with such detail that any other engineer could come to the same conclusion if he accepted the facts as accurately determined. In regard to the detailed form of reports, the writer's own preference is for a single page summarizing the main factors, and an assay plan, reduced to a longitudinal section where possible. Then there should be added, for purposes of record and for submission to other engineers, a set of appendices going into some details as to the history of the mine, its geology, development, equipment, metallurgy, and management. A list of samples should be given with their location, and the tonnages and values of each separate block. A presentation should be made of the probabilities of extension in depth, together with recommendations for working the mine.

GENERAL SUMMARY.

The bed-rock value which attaches to a mine is the profit to be won from proved ore and in which the price of metal is calculated at some figure between "basic" and "normal." This we may call the "A" value. Beyond this there is the speculative value of the mine. If the value of the "probable" ore be represented by X, the value of extension of the ore by Y, and a higher price for metal than the price above assumed represented by Z, then if the mine be efficiently managed the value of the mine is A + X + Y + Z. What actual amounts should be attached to X, Y, Z is a matter of judgment. There is no prescription for good judgment. Good judgment rests upon a proper balancing of evidence. The amount of risk in X, Y, Z is purely a question of how much these factors are required to represent in money,—in effect, how much more ore must be found, or how many feet the ore must extend in depth; or in convertible terms, what life in years the mine must have, or how high the price of metal must be. In forming an opinion whether these requirements will be realized, X, Y, Z must be balanced in a scale whose measuring standards are the five geological weights and the general industrial outlook. The wise engineer will put before his clients the scale, the weights, and the conclusion arrived at. The shrewd investor will require to know these of his adviser.



CHAPTER VII.

Development of Mines.

ENTRY TO THE MINE; TUNNELS; VERTICAL, INCLINED, AND COMBINED SHAFTS; LOCATION AND NUMBER OF SHAFTS.

Development is conducted for two purposes: first, to search for ore; and second, to open avenues for its extraction. Although both objects are always more or less in view, the first predominates in the early life of mines, the prospecting stage, and the second in its later life, the producing stage. It is proposed to discuss development designed to embrace extended production purposes first, because development during the prospecting stage is governed by the same principles, but is tempered by the greater degree of uncertainty as to the future of the mine, and is, therefore, of a more temporary character.

ENTRY TO THE MINE.

There are four methods of entry: by tunnel, vertical shaft, inclined shaft, or by a combination of the last two, that is, by a shaft initially vertical then turned to an incline. Combined shafts are largely a development of the past few years to meet "deep level" conditions, and have been rendered possible only by skip-winding. The angle in such shafts (Fig. 2) is now generally made on a parabolic curve, and the speed of winding is then less diminished by the bend.

The engineering problems which present themselves under "entry" may be divided into those of:—

1. Method. 2. Location. 3. Shape and size.

The resolution of these questions depends upon the:—

a. Degree of dip of the deposit. b. Output of ore to be provided for. c. Depth at which the deposit is to be attacked. d. Boundaries of the property. e. Surface topography. f. Cost. g. Operating efficiency. h. Prospects of the mine.



From the point of view of entrance, the cooeperation of a majority of these factors permits the division of mines into certain broad classes. The type of works demanded for moderate depths (say vertically 2,500 to 3,000 feet) is very different from that required for great depths. To reach great depths, the size of shafts must greatly expand, to provide for extended ventilation, pumping, and winding necessities. Moreover inclined shafts of a degree of flatness possible for moderate depths become too long to be used economically from the surface. The vast majority of metal-mining shafts fall into the first class, those of moderate depths. Yet, as time goes on and ore-deposits are exhausted to lower planes, problems of depth will become more common. One thing, however, cannot be too much emphasized, especially on mines to be worked from the outcrop, and that is, that no engineer is warranted, owing to the speculation incidental to extension in depth, in initiating early in the mine's career shafts of such size or equipment as would be available for great depths. Moreover, the proper location of a shaft so as to work economically extension of the ore-bodies is a matter of no certainty, and therefore shafts of speculative mines are tentative in any event.

Another line of division from an engineering view is brought about by a combination of three of the factors mentioned. This is the classification into "outcrop" and "deep-level" mines. The former are those founded upon ore-deposits to be worked from or close to the surface. The latter are mines based upon the extension in depth of ore-bodies from outcrop mines. Such projects are not so common in America, where the law in most districts gives the outcrop owner the right to follow ore beyond his side-lines, as in countries where the boundaries are vertical on all sides. They do, however, arise not alone in the few American sections where the side-lines are vertical boundaries, but in other parts owing to the pitch of ore-bodies through the end lines (Fig. 3). More especially do such problems arise in America in effect, where the ingress questions have to be revised for mines worked out in the upper levels (Fig. 7).



If from a standpoint of entrance questions, mines are first classified into those whose works are contemplated for moderate depths, and those in which work is contemplated for great depth, further clarity in discussion can be gained by subdivision into the possible cases arising out of the factors of location, dip, topography, and boundaries.

MINES OF MODERATE DEPTHS.

Case I. Deposits where topographic conditions permit the alternatives of shaft or tunnel. Case II. Vertical or horizontal deposits, the only practical means of attaining which is by a vertical shaft. Case III. Inclined deposits to be worked from near the surface. There are in such instances the alternatives of either a vertical or an inclined shaft. Case IV. Inclined deposits which must be attacked in depth, that is, deep-level projects. There are the alternatives of a compound shaft or of a vertical shaft, and in some cases of an incline from the surface.

MINES TO GREAT DEPTHS.

Case V. Vertical or horizontal deposits, the only way of reaching which is by a vertical shaft. Case VI. Inclined deposits. In such cases the alternatives are a vertical or a compound shaft.

CASE I.—Although for logical arrangement tunnel entry has been given first place, to save repetition it is proposed to consider it later. With few exceptions, tunnels are a temporary expedient in the mine, which must sooner or later be opened by a shaft.

CASE II. VERTICAL OR HORIZONTAL DEPOSITS.—These require no discussion as to manner of entry. There is no justifiable alternative to a vertical shaft (Fig. 4).



CASE III. INCLINED DEPOSITS WHICH ARE INTENDED TO BE WORKED FROM THE OUTCROP, OR FROM NEAR IT (Fig. 5).—The choice of inclined or vertical shaft is dependent upon relative cost of construction, subsequent operation, and the useful life of the shaft, and these matters are largely governed by the degree of dip. Assuming a shaft of the same size in either alternative, the comparative cost per foot of sinking is dependent largely on the breaking facilities of the rock under the different directions of attack. In this, the angles of the bedding or joint planes to the direction of the shaft outweigh other factors. The shaft which takes the greatest advantage of such lines of breaking weakness will be the cheapest per foot to sink. In South African experience, where inclined shafts are sunk parallel to the bedding planes of hard quartzites, the cost per foot appears to be in favor of the incline. On the other hand, sinking shafts across tight schists seems to be more advantageous than parallel to the bedding planes, and inclines following the dip cost more per foot than vertical shafts.

An inclined shaft requires more footage to reach a given point of depth, and therefore it would entail a greater total expense than a vertical shaft, assuming they cost the same per foot. The excess amount will be represented by the extra length, and this will depend upon the flatness of the dip. With vertical shafts, however, crosscuts to the deposit are necessary. In a comparative view, therefore, the cost of the crosscuts must be included with that of the vertical shaft, as they would be almost wholly saved in an incline following near the ore.

The factor of useful life for the shaft enters in deciding as to the advisability of vertical shafts on inclined deposits, from the fact that at some depth one of two alternatives has to be chosen. The vertical shaft, when it reaches a point below the deposit where the crosscuts are too long (C, Fig. 5), either becomes useless, or must be turned on an incline at the intersection with the ore (B). The first alternative means ultimately a complete loss of the shaft for working purposes. The latter has the disadvantage that the bend interferes slightly with haulage.

The following table will indicate an hypothetical extreme case,—not infrequently met. In it a vertical shaft 1,500 feet in depth is taken as cutting the deposit at the depth of 750 feet, the most favored position so far as aggregate length of crosscuts is concerned. The cost of crosscutting is taken at $20 per foot and that of sinking the vertical shaft at $75 per foot. The incline is assumed for two cases at $75 and $100 per foot respectively. The stoping height upon the ore between levels is counted at 125 feet.

Dip of Depth of Length of No. of Crosscuts Total Length Deposit from Vertical Incline Required from of Crosscuts, Horizontal Shaft Required V Shaft Feet - - - - 80 deg. 1,500 1,522 11 859 70 deg. 1,500 1,595 12 1,911 60 deg. 1,500 1,732 13 3,247 50 deg. 1,500 1,058 15 5,389 40 deg. 1,500 2,334 18 8,038 30 deg. 1,500 3,000 23 16,237 ========================================================================== Cost of Cost Vertical Total Cost Cost of Incline Cost of Incline Crosscuts $20 Shaft $75 of Vertical $75 per Foot $100 per Foot per Foot per Foot and Crosscuts - - - - $17,180 $112,500 $129,680 $114,150 $152,200 38,220 112,500 150,720 118,625 159,500 64,940 112,500 177,440 129,900 172,230 107,780 112,500 220,280 114,850 195,800 178,760 112,500 291,260 175,050 233,400 324,740 112,500 437,240 225,000 300,000

From the above examples it will be seen that the cost of crosscuts put at ordinary level intervals rapidly outruns the extra expense of increased length of inclines. If, however, the conditions are such that crosscuts from a vertical shaft are not necessary at so frequent intervals, then in proportion to the decrease the advantages sway to the vertical shaft. Most situations wherein the crosscuts can be avoided arise in mines worked out in the upper levels and fall under Case IV, that of deep-level projects.

There can be no doubt that vertical shafts are cheaper to operate than inclines: the length of haul from a given depth is less; much higher rope speed is possible, and thus the haulage hours are less for the same output; the wear and tear on ropes, tracks, or guides is not so great, and pumping is more economical where the Cornish order of pump is used. On the other hand, with a vertical shaft must be included the cost of operating crosscuts. On mines where the volume of ore does not warrant mechanical haulage, the cost of tramming through the extra distance involved is an expense which outweighs any extra operating outlay in the inclined shaft itself. Even with mechanical haulage in crosscuts, it is doubtful if there is anything in favor of the vertical shaft on this score.



In deposits of very flat dips, under 30 deg., the case arises where the length of incline is so great that the saving on haulage through direct lift warrants a vertical shaft as an auxiliary outlet in addition to the incline (Fig. 6). In such a combination the crosscut question is eliminated. The mine is worked above and below the intersection by incline, and the vertical shaft becomes simply a more economical exit and an alternative to secure increased output. The North Star mine at Grass Valley is an illustration in point. Such a positive instance borders again on Case IV, deep-level projects.

In conclusion, it is the writer's belief that where mines are to be worked from near the surface, coincidentally with sinking, and where, therefore, crosscuts from a vertical shaft would need to be installed frequently, inclines are warranted in all dips under 75 deg. and over 30 deg. Beyond 75 deg. the best alternative is often undeterminable. In the range under 30 deg. and over 15 deg., although inclines are primarily necessary for actual delivery of ore from levels, they can often be justifiably supplemented by a vertical shaft as a relief to a long haul. In dips of less than 15 deg., as in those over 75 deg., the advantages again trend strongly in favor of the vertical shaft. There arise, however, in mountainous countries, topographic conditions such as the dip of deposits into the mountain, which preclude any alternative on an incline at any angled dip.

CASE IV. INCLINED DEPOSITS WHICH MUST BE ATTACKED IN DEPTH (Fig. 7).—There are two principal conditions in which such properties exist: first, mines being operated, or having been previously worked, whose method of entry must be revised; second, those whose ore-bodies to be attacked do not outcrop within the property.

The first situation may occur in mines of inadequate shaft capacity or wrong location; in mines abandoned and resurrected; in mines where a vertical shaft has reached its limit of useful extensions, having passed the place of economical crosscutting; or in mines in flat deposits with inclines whose haul has become too long to be economical. Three alternatives present themselves in such cases: a new incline from the surface (A B F, Fig. 7), or a vertical shaft combined with incline extension (C D F), or a simple vertical shaft (H G). A comparison can be first made between the simple incline and the combined shaft. The construction of an incline from the surface to the ore-body will be more costly than a combined shaft, for until the horizon of the ore is reached (at D) no crosscuts are required in the vertical section, while the incline must be of greater length to reach the same horizon. The case arises, however, where inclines can be sunk through old stopes, and thus more cheaply constructed than vertical shafts through solid rock; and also the case of mountainous topographic conditions mentioned above.



From an operating point of view, the bend in combined shafts (at D) gives rise to a good deal of wear and tear on ropes and gear. The possible speed of winding through a combined shaft is, however, greater than a simple incline, for although haulage speed through the incline section (D F) and around the bend of the combined shaft is about the same as throughout a simple incline (A F), the speed can be accelerated in the vertical portion (D C) above that feasible did the incline extend to the surface. There is therefore an advantage in this regard in the combined shaft. The net advantages of the combined over the inclined shaft depend on the comparative length of the two alternative routes from the intersection (D) to the surface. Certainly it is not advisable to sink a combined shaft to cut a deposit at 300 feet in depth if a simple incline can be had to the surface. On the other hand, a combined shaft cutting the deposit at 1,000 feet will be more advisable than a simple incline 2,000 feet long to reach the same point. The matter is one for direct calculation in each special case. In general, there are few instances of really deep-level projects where a complete incline from the surface is warranted.

In most situations of this sort, and in all of the second type (where the outcrop is outside the property), actual choice usually lies between combined shafts (C D F) and entire vertical shafts (H G). The difference between a combined shaft and a direct vertical shaft can be reduced to a comparison of the combined shaft below the point of intersection (D) with that portion of a vertical shaft which would cover the same horizon. The question then becomes identical with that of inclined versus verticals, as stated in Case III, with the offsetting disadvantage of the bend in the combined shaft. If it is desired to reach production at the earliest date, the lower section of a simple vertical shaft must have crosscuts to reach the ore lying above the horizon of its intersection (E). If production does not press, the ore above the intersection (EB) can be worked by rises from the horizon of intersection (E). In the use of rises, however, there follow the difficulties of ventilation and lowering the ore down to the shaft, which brings expenses to much the same thing as operating through crosscuts.

The advantages of combined over simple vertical shafts are earlier production, saving of either rises or crosscuts, and the ultimate utility of the shaft to any depth. The disadvantages are the cost of the extra length of the inclined section, slower winding, and greater wear and tear within the inclined section and especially around the bend. All these factors are of variable import, depending upon the dip. On very steep dips,—over 70 deg.,—the net result is in favor of the simple vertical shaft. On other dips it is in favor of the combined shaft.

CASES V AND VI. MINES TO BE WORKED TO GREAT DEPTHS,—OVER 3,000 FEET.—In Case V, with vertical or horizontal deposits, there is obviously no desirable alternative to vertical shafts.

In Case VI, with inclined deposits, there are the alternatives of a combined or of a simple vertical shaft. A vertical shaft in locations (H, Fig. 7) such as would not necessitate extension in depth by an incline, would, as in Case IV, compel either crosscuts to the ore or inclines up from the horizon of intersection (E). Apart from delay in coming to production and the consequent loss of interest on capital, the ventilation problems with this arrangement would be appalling. Moreover, the combined shaft, entering the deposit near its shallowest point, offers the possibility of a separate haulage system on the inclined and on the vertical sections, and such separate haulage is usually advisable at great depths. In such instances, the output to be handled is large, for no mine of small output is likely to be contemplated at such depth. Several moderate-sized inclines from the horizon of intersection have been suggested (EF, DG, CH, Fig. 8) to feed a large primary shaft (AB), which thus becomes the trunk road. This program would cheapen lateral haulage underground, as mechanical traction can be used in the main level, (EC), and horizontal haulage costs can be reduced on the lower levels. Moreover, separate winding engines on the two sections increase the capacity, for the effect is that of two trains instead of one running on a single track.

SHAFT LOCATION.—Although the prime purpose in locating a shaft is obviously to gain access to the largest volume of ore within the shortest haulage distance, other conditions also enter, such as the character of the surface and the rock to be intersected, the time involved before reaching production, and capital cost. As shafts must bear two relations to a deposit,—one as to the dip and the other as to the strike,—they may be considered from these aspects. Vertical shafts must be on the hanging-wall side of the outcrop if the deposit dips at all. In any event, the shaft should be far enough away to be out of the reach of creeps. An inclined shaft may be sunk either on the vein, in which case a pillar of ore must be left to support the shaft; or, instead, it may be sunk a short distance in the footwall, and where necessary the excavation above can be supported by filling. Following the ore has the advantage of prospecting in sinking, and in many cases the softness of the ground in the region of the vein warrants this procedure. It has, however, the disadvantage that a pillar of ore is locked up until the shaft is ready for abandonment. Moreover, as veins or lodes are seldom of even dip, an inclined shaft, to have value as a prospecting opening, or to take advantage of breaking possibilities in the lode, will usually be crooked, and an incline irregular in detail adds greatly to the cost of winding and maintenance. These twin disadvantages usually warrant a straight incline in the footwall. Inclines are not necessarily of the same dip throughout, but for reasonably economical haulage change of angle must take place gradually.



In the case of deep-level projects on inclined deposits, demanding combined or vertical shafts, the first desideratum is to locate the vertical section as far from the outcrop as possible, and thus secure the most ore above the horizon of intersection. This, however, as stated before, would involve the cost of crosscuts or rises and would cause delay in production, together with the accumulation of capital charges. How important the increment of interest on capital may become during the period of opening the mine may be demonstrated by a concrete case. For instance, the capital of a company or the cost of the property is, say, $1,000,000, and where opening the mine for production requires four years, the aggregate sum of accumulated compound interest at 5% (and most operators want more from a mining investment) would be $216,000. Under such circumstances, if a year or two can be saved in getting to production by entering the property at a higher horizon, the difference in accumulated interest will more than repay the infinitesimal extra cost of winding through a combined shaft of somewhat increased length in the inclined section.

The unknown character of the ore in depth is always a sound reason for reaching it as quickly and as cheaply as possible. In result, such shafts are usually best located when the vertical section enters the upper portion of the deposit.

The objective in location with regard to the strike of the ore-bodies is obviously to have an equal length of lateral ore-haul in every direction from the shaft. It is easier to specify than to achieve this, for in all speculative deposits ore-shoots are found to pursue curious vagaries as they go down. Ore-bodies do not reoccur with the same locus as in the upper levels, and generally the chances to go wrong are more numerous than those to go right.

NUMBER OF SHAFTS.—The problem of whether the mine is to be opened by one or by two shafts of course influences location. In metal mines under Cases II and III (outcrop properties) the ore output requirements are seldom beyond the capacity of one shaft. Ventilation and escape-ways are usually easily managed through the old stopes. Under such circumstances, the conditions warranting a second shaft are the length of underground haul and isolation of ore-bodies or veins. Lateral haulage underground is necessarily disintegrated by the various levels, and usually has to be done by hand. By shortening this distance of tramming and by consolidation of the material from all levels at the surface, where mechanical haulage can be installed, a second shaft is often justified. There is therefore an economic limitation to the radius of a single shaft, regardless of the ability of the shaft to handle the total output.

Other questions also often arise which are of equal importance to haulage costs. Separate ore-shoots or ore-bodies or parallel deposits necessitate, if worked from one shaft, constant levels through unpayable ground and extra haul as well, or ore-bodies may dip away from the original shaft along the strike of the deposit and a long haulage through dead levels must follow. For instance, levels and crosscuts cost roughly one-quarter as much per foot as shafts. Therefore four levels in barren ground, to reach a parallel vein or isolated ore-body 1,000 feet away, would pay for a shaft 1,000 feet deep. At a depth of 1,000 feet, at least six levels might be necessary. The tramming of ore by hand through such a distance would cost about double the amount to hoist it through a shaft and transport it mechanically to the dressing plant at surface. The aggregate cost and operation of barren levels therefore soon pays for a second shaft. If two or more shafts are in question, they must obviously be set so as to best divide the work.

Under Cases IV, V, and VI,—that is, deep-level projects,—ventilation and escape become most important considerations. Even where the volume of ore is within the capacity of a single shaft, another usually becomes a necessity for these reasons. Their location is affected not only by the locus of the ore, but, as said, by the time required to reach it. Where two shafts are to be sunk to inclined deposits, it is usual to set one so as to intersect the deposit at a lower point than the other. Production can be started from the shallower, before the second is entirely ready. The ore above the horizon of intersection of the deeper shaft is thus accessible from the shallower shaft, and the difficulty of long rises or crosscuts from that deepest shaft does not arise.



CHAPTER VIII.

Development of Mines (Continued).

SHAPE AND SIZE OF SHAFTS; SPEED OF SINKING; TUNNELS.

SHAPE OF SHAFTS.—Shafts may be round or rectangular.[*] Round vertical shafts are largely applied to coal-mines, and some engineers have advocated their usefulness to the mining of the metals under discussion. Their great advantages lie in their structural strength, in the large amount of free space for ventilation, and in the fact that if walled with stone, brick, concrete, or steel, they can be made water-tight so as to prevent inflow from water-bearing strata, even when under great pressure. The round walled shafts have a longer life than timbered shafts. All these advantages pertain much more to mining coal or iron than metals, for unsound, wet ground is often the accompaniment of coal-measures, and seldom troubles metal-mines. Ventilation requirements are also much greater in coal-mines. From a metal-miner's standpoint, round shafts are comparatively much more expensive than the rectangular timbered type.[**] For a larger area must be excavated for the same useful space, and if support is needed, satisfactory walling, which of necessity must be brick, stone, concrete, or steel, cannot be cheaply accomplished under the conditions prevailing in most metal regions. Although such shafts would have a longer life, the duration of timbered shafts is sufficient for most metal mines. It follows that, as timber is the cheapest and all things considered the most advantageous means of shaft support for the comparatively temporary character of metal mines, to get the strains applied to the timbers in the best manner, and to use the minimum amount of it consistent with security, and to lose the least working space, the shaft must be constructed on rectangular lines.

[Footnote *: Octagonal shafts were sunk in Mexico in former times. At each face of the octagon was a whim run by mules, and hauling leather buckets.]

[Footnote **: The economic situation is rapidly arising in a number of localities that steel beams can be usefully used instead of timber. The same arguments apply to this type of support that apply to timber.]

The variations in timbered shaft design arise from the possible arrangement of compartments. Many combinations can be imagined, of which Figures 9, 10, 11, 12, 13, and 14 are examples.



The arrangement of compartments shown in Figures 9, 10, 11, and 13 gives the greatest strength. It permits timbering to the best advantage, and avoids the danger underground involved in crossing one compartment to reach another. It is therefore generally adopted. Any other arrangement would obviously be impossible in inclined or combined shafts.

SIZE OF SHAFTS.—In considering the size of shafts to be installed, many factors are involved. They are in the main:—

a. Amount of ore to be handled. b. Winding plant. c. Vehicle of transport. d. Depth. e. Number of men to be worked underground. f. Amount of water. g. Ventilation. h. Character of the ground. i. Capital outlay. j. Operating expense.

It is not to be assumed that these factors have been stated in the order of relative importance. More or less emphasis will be attached to particular factors by different engineers, and under different circumstances. It is not possible to suggest any arbitrary standard for calculating their relative weight, and they are so interdependent as to preclude separate discussion. The usual result is a compromise between the demands of all.

Certain factors, however, dictate a minimum position, which may be considered as a datum from which to start consideration.

First, a winding engine, in order to work with any economy, must be balanced, that is, a descending empty skip or cage must assist in pulling up a loaded one. Therefore, except in mines of very small output, at least two compartments must be made for hoisting purposes. Water has to be pumped from most mines, escape-ways are necessary, together with room for wires and air-pipes, so that at least one more compartment must be provided for these objects. We have thus three compartments as a sound minimum for any shaft where more than trivial output is required.

Second, there is a certain minimum size of shaft excavation below which there is very little economy in actual rock-breaking.[*] In too confined a space, holes cannot be placed to advantage for the blast, men cannot get round expeditiously, and spoil cannot be handled readily. The writer's own experience leads him to believe that, in so far as rock-breaking is concerned, to sink a shaft fourteen to sixteen feet long by six to seven feet wide outside the timbers, is as cheap as to drive any smaller size within the realm of consideration, and is more rapid. This size of excavation permits of three compartments, each about four to five feet inside the timbers.

[Footnote *: Notes on the cost of shafts in various regions which have been personally collected show a remarkable decrease in the cost per cubic foot of material excavated with increased size of shaft. Variations in skill, in economic conditions, and in method of accounting make data regarding different shafts of doubtful value, but the following are of interest:—

In Australia, eight shafts between 10 and 11 feet long by 4 to 5 feet wide cost an average of $1.20 per cubic foot of material excavated. Six shafts 13 to 14 feet long by 4 to 5 feet wide cost an average of $0.95 per cubic foot; seven shafts 14 to 16 feet long and 5 to 7 feet wide cost an average of $0.82 per cubic foot. In South Africa, eleven shafts 18 to 19 feet long by 7 to 8 feet wide cost an average of $0.82 per cubic foot; five shafts 21 to 25 feet long by 8 feet wide, cost $0.74; and seven shafts 28 feet by 8 feet cost $0.60 per cubic foot.]

The cost of timber, it is true, is a factor of the size of shaft, but the labor of timbering does not increase in the same ratio. In any event, the cost of timber is only about 15% of the actual shaft cost, even in localities of extremely high prices.

Third, three reasons are rapidly making the self-dumping skip the almost universal shaft-vehicle, instead of the old cage for cars. First, there is a great economy in labor for loading into and discharging from a shaft; second, there is more rapid despatch and discharge and therefore a larger number of possible trips; third, shaft-haulage is then independent of delays in arrival of cars at stations, while tramming can be done at any time and shaft-haulage can be concentrated into certain hours. Cages to carry mine cars and handle the same load as a skip must either be big enough to take two cars, which compels a much larger shaft than is necessary with skips, or they must be double-decked, which renders loading arrangements underground costly to install and expensive to work. For all these reasons, cages can be justified only on metal mines of such small tonnage that time is no consideration and where the saving of men is not to be effected. In compartments of the minimum size mentioned above (four to five feet either way) a skip with a capacity of from two to five tons can be installed, although from two to three tons is the present rule. Lighter loads than this involve more trips, and thus less hourly capacity, and, on the other hand, heavier loads require more costly engines. This matter is further discussed under "Haulage Appliances."

We have therefore as the economic minimum a shaft of three compartments (Fig. 9), each four to five feet square. When the maximum tonnage is wanted from such a shaft at the least operating cost, it should be equipped with loading bins and skips.

The output capacity of shafts of this size and equipment will depend in a major degree upon the engine employed, and in a less degree upon the hauling depth. The reason why depth is a subsidiary factor is that the rapidity with which a load can be drawn is not wholly a factor of depth. The time consumed in hoisting is partially expended in loading, in acceleration and retardation of the engine, and in discharge of the load. These factors are constant for any depth, and extra distance is therefore accomplished at full speed of the engine.

Vertical shafts will, other things being equal, have greater capacity than inclines, as winding will be much faster and length of haul less for same depth. Since engines have, however, a great tractive ability on inclines, by an increase in the size of skip it is usually possible partially to equalize matters. Therefore the size of inclines for the same output need not differ materially from vertical shafts.

The maximum capacity of a shaft whose equipment is of the character and size given above, will, as stated, decrease somewhat with extension in depth of the haulage horizon. At 500 feet, such a shaft if vertical could produce 70 to 80 tons per hour comfortably with an engine whose winding speed was 700 feet per minute. As men and material other than ore have to be handled in and out of the mine, and shaft-sinking has to be attended to, the winding engine cannot be employed all the time on ore. Twelve hours of actual daily ore-winding are all that can be expected without auxiliary help. This represents a capacity from such a depth of 800 to 1,000 tons per day. A similar shaft, under ordinary working conditions, with an engine speed of 2,000 feet per minute, should from, say, 3,000 feet have a capacity of about 400 to 600 tons daily.

It is desirable to inquire at what stages the size of shaft should logically be enlarged in order to attain greater capacity. A considerable measure of increase can be obtained by relieving the main hoisting engine of all or part of its collateral duties. Where the pumping machinery is not elaborate, it is often possible to get a small single winding compartment into the gangway without materially increasing the size of the shaft if the haulage compartments be made somewhat narrower (Fig. 10). Such a compartment would be operated by an auxiliary engine for sinking, handling tools and material, and assisting in handling men. If this arrangement can be effected, the productive time of the main engine can be expanded to about twenty hours with an addition of about two-thirds to the output.

Where the exigencies of pump and gangway require more than two and one-half feet of shaft length, the next stage of expansion becomes four full-sized compartments (Fig. 11). By thus enlarging the auxiliary winding space, some assistance may be given to ore-haulage in case of necessity. The mine whose output demands such haulage provisions can usually stand another foot of width to the shaft, so that the dimensions come to about 21 feet to 22 feet by 7 feet to 8 feet outside the timbers. Such a shaft, with three- to four-ton skips and an appropriate engine, will handle up to 250 tons per hour from a depth of 1,000 feet.

The next logical step in advance is the shaft of five compartments with four full-sized haulage ways (Fig. 13), each of greater size than in the above instance. In this case, the auxiliary engine becomes a balanced one, and can be employed part of the time upon ore-haulage. Such a shaft will be about 26 feet to 28 feet long by 8 feet wide outside the timbers, when provision is made for one gangway. The capacity of such shafts can be up to 4,000 tons a day, depending on the depth and engine. When very large quantities of water are to be dealt with and rod-driven pumps to be used, two pumping compartments are sometimes necessary, but other forms of pumps do not require more than one compartment,—an additional reason for their use.

For depths greater than 3,000 feet, other factors come into play. Ventilation questions become of more import. The mechanical problems on engines and ropes become involved, and their sum-effect is to demand much increased size and a greater number of compartments. The shafts at Johannesburg intended as outlets for workings 5,000 feet deep are as much as 46 feet by 9 feet outside timbers.

It is not purposed to go into details as to sinking methods or timbering. While important matters, they would unduly prolong this discussion. Besides, a multitude of treatises exist on these subjects and cover all the minutiae of such work.

SPEED OF SINKING.—Mines may be divided into two cases,—those being developed only, and those being operated as well as developed. In the former, the entrance into production is usually dependent upon the speed at which the shaft is sunk. Until the mine is earning profits, there is a loss of interest on the capital involved, which, in ninety-nine instances out of a hundred, warrants any reasonable extra expenditure to induce more rapid progress. In the case of mines in operation, the volume of ore available to treatment or valuation is generally dependent to a great degree upon the rapidity of the extension of workings in depth. It will be demonstrated later that, both from a financial and a technical standpoint, the maximum development is the right one and that unremitting extension in depth is not only justifiable but necessary.

Speed under special conditions or over short periods has a more romantic than practical interest, outside of its value as a stimulant to emulation. The thing that counts is the speed which can be maintained over the year. Rapidity of sinking depends mainly on:—

a. Whether the shaft is or is not in use for operating the mine. b. The breaking character of the rock. c. The amount of water.

The delays incident to general carrying of ore and men are such that the use of the main haulage engine for shaft-sinking is practically impossible, except on mines with small tonnage output. Even with a separate winch or auxiliary winding-engine, delays are unavoidable in a working shaft, especially as it usually has more water to contend with than one not in use for operating the mine. The writer's own impression is that an average of 40 feet per month is the maximum possibility for year in and out sinking under such conditions. In fact, few going mines manage more than 400 feet a year. In cases of clean shaft-sinking, where every energy is bent to speed, 150 feet per month have been averaged for many months. Special cases have occurred where as much as 213 feet have been achieved in a single month. With ordinary conditions, 1,200 feet in a year is very good work. Rock awkward to break, and water especially, lowers the rate of progress very materially. Further reference to speed will be found in the chapter on "Drilling Methods."

TUNNEL ENTRY.—The alternative of entry to a mine by tunnel is usually not a question of topography altogether, but, like everything else in mining science, has to be tempered to meet the capital available and the expenditure warranted by the value showing.

In the initial prospecting of a mine, tunnels are occasionally overdone by prospectors. Often more would be proved by a few inclines. As the pioneer has to rely upon his right arm for hoisting and drainage, the tunnel offers great temptations, even when it is long and gains but little depth. At a more advanced stage of development, the saving of capital outlay on hoisting and pumping equipment, at a time when capital is costly to secure, is often sufficient justification for a tunnel entry. But at the stage where the future working of ore below a tunnel-level must be contemplated, other factors enter. For ore below tunnel-level a shaft becomes necessary, and in cases where a tunnel enters a few hundred feet below the outcrop the shaft should very often extend to the surface, because internal shafts, winding from tunnel-level, require large excavations to make room for the transfer of ore and for winding gear. The latter must be operated by transmitted power, either that of steam, water, electricity, or air. Where power has to be generated on the mine, the saving by the use of direct steam, generated at the winding gear, is very considerable. Moreover, the cost of haulage through a shaft for the extra distance from tunnel-level to the surface is often less than the cost of transferring the ore and removing it through the tunnel. The load once on the winding-engine, the consumption of power is small for the extra distance, and the saving of labor is of consequence. On the other hand, where drainage problems arise, they usually outweigh all other considerations, for whatever the horizon entered by tunnel, the distance from that level to the surface means a saving of water-pumpage against so much head. The accumulation of such constant expense justifies a proportioned capital outlay. In other words, the saving of this extra pumping will annually redeem the cost of a certain amount of tunnel, even though it be used for drainage only.

In order to emphasize the rapidity with which such a saving of constant expense will justify capital outlay, one may tabulate the result of calculations showing the length of tunnel warranted with various hypothetical factors of quantity of water and height of lift eliminated from pumping. In these computations, power is taken at the low rate of $60 per horsepower-year, the cost of tunneling at an average figure of $20 per foot, and the time on the basis of a ten-year life for the mine.

Feet of Tunnel Paid for in 10 Years with Under-mentioned Conditions.

============================================================= Feet of 100,000 200,000 300,000 500,000 1,000,000 Water Lift Gallons Gallons Gallons Gallons Gallons Avoided per Diem per Diem per Diem per Diem per Diem - - - - - - 100 600 1,200 1,800 3,000 6,000 200 1,200 2,400 3,600 6,000 12,000 300 1,800 3,600 5,400 9,000 18,000 500 3,000 6,000 9,000 15,000 30,000 1,000 6,000 12,000 18,000 30,000 60,000 =============================================================

The size of tunnels where ore-extraction is involved depends upon the daily tonnage output required, and the length of haul. The smallest size that can be economically driven and managed is about 6-1/2 feet by 6 feet inside the timbers. Such a tunnel, with single track for a length of 1,000 feet, with one turn-out, permits handling up to 500 tons a day with men and animals. If the distance be longer or the tonnage greater, a double track is required, which necessitates a tunnel at least 8 feet wide by 6-1/2 feet to 7 feet high, inside the timbers.

There are tunnel projects of a much more impressive order than those designed to operate upper levels of mines; that is, long crosscut tunnels designed to drain and operate mines at very considerable depths, such as the Sutro tunnel at Virginia City. The advantage of these tunnels is very great, especially for drainage, and they must be constructed of large size and equipped with appliances for mechanical haulage.



CHAPTER IX.

Development of Mines (Concluded).

SUBSIDIARY DEVELOPMENT;—STATIONS; CROSSCUTS; LEVELS; INTERVAL BETWEEN LEVELS; PROTECTION OF LEVELS; WINZES AND RISES. DEVELOPMENT IN THE PROSPECTING STAGE; DRILLING.

SUBSIDIARY DEVELOPMENT.

Stations, crosscuts, levels, winzes, and rises follow after the initial entry. They are all expensive, and the least number that will answer is the main desideratum.

STATIONS.—As stations are the outlets of the levels to the shaft, their size and construction is a factor of the volume and character of the work at the levels which they are to serve. If no timber is to be handled, and little ore, and this on cages, the stations need be no larger than a good sized crosscut. Where timber is to be let down, they must be ten to fifteen feet higher than the floor of the crosscut. Where loading into skips is to be provided for, bins must be cut underneath and sufficient room be provided to shift the mine cars comfortably. Such bins are built of from 50 to 500 tons' capacity in order to contain some reserve for hoisting purposes, and in many cases separate bins must be provided on opposite sides of the shaft for ore and waste. It is a strong argument in favor of skips, that with this means of haulage storage capacity at the stations is possible, and the hoisting may then go on independently of trucking and, as said before, there are no idle men at the stations.



It is always desirable to concentrate the haulage to the least number of levels, for many reasons. Among them is that, where haulage is confined to few levels, storage-bins are not required at every station. Figures 15, 16, 17, and 18 illustrate various arrangements of loading bins.

CROSSCUTS.—Crosscuts are for two purposes, for roadway connection of levels to the shaft or to other levels, and for prospecting purposes. The number of crosscuts for roadways can sometimes be decreased by making the connections with the shaft at every second or even every third level, thus not only saving in the construction cost of crosscuts and stations, but also in the expenses of scattered tramming. The matter becomes especially worth considering where the quantity of ore that can thus be accumulated warrants mule or mechanical haulage. This subject will be referred to later on.



On the second purpose of crosscuts,—that of prospecting,—one observation merits emphasis. This is, that the tendency of ore-fissures to be formed in parallels warrants more systematic crosscutting into the country rock than is done in many mines.



LEVELS.

The word "level" is another example of miners' adaptations in nomenclature. Its use in the sense of tunnels driven in the direction of the strike of the deposit has better, but less used, synonyms in the words "drifts" or "drives." The term "level" is used by miners in two senses, in that it is sometimes applied to all openings on one horizon, crosscuts included. Levels are for three purposes,—for a stoping base; for prospecting the deposit; and for roadways. As a prospecting and a stoping base it is desirable that the level should be driven on the deposit; as a roadway, that it should constitute the shortest distance between two points and be in the soundest ground. On narrow, erratic deposits the levels usually must serve all three purposes at once; but in wider and more regular deposits levels are often driven separately for roadways from the level which forms the stoping base and prospecting datum.

There was a time when mines were worked by driving the level on ore and enlarging it top and bottom as far as the ground would stand, then driving the next level 15 to 20 feet below, and repeating the operation. This interval gradually expanded, but for some reason 100 feet was for years assumed to be the proper distance between levels. Scattered over every mining camp on earth are thousands of mines opened on this empirical figure, without consideration of the reasons for it or for any other distance.

The governing factors in determining the vertical interval between levels are the following:—

a. The regularity of the deposit. b. The effect of the method of excavation of winzes and rises. c. The dip and the method of stoping.

REGULARITY OF THE DEPOSIT.—From a prospecting point of view the more levels the better, and the interval therefore must be determined somewhat by the character of the deposit. In erratic deposits there is less risk of missing ore with frequent levels, but it does not follow that every level need be a through roadway to the shaft or even a stoping base. In such deposits, intermediate levels for prospecting alone are better than complete levels, each a roadway. Nor is it essential, even where frequent levels are required for a stoping base, that each should be a main haulage outlet to the shaft. In some mines every third level is used as a main roadway, the ore being poured from the intermediate ones down to the haulage line. Thus tramming and shaft work, as stated before, can be concentrated.

EFFECT OF METHOD OF EXCAVATING WINZES AND RISES.—With hand drilling and hoisting, winzes beyond a limited depth become very costly to pull spoil out of, and rises too high become difficult to ventilate, so that there is in such cases a limit to the interval desirable between levels, but these difficulties largely disappear where air-winches and air-drills are used.

THE DIP AND METHOD OF STOPING.—The method of stoping is largely dependent upon the dip, and indirectly thus affects level intervals. In dips under that at which material will "flow" in the stopes—about 45 deg. to 50 deg.—the interval is greatly dependent on the method of stope-transport. Where ore is to be shoveled from stopes to the roadway, the levels must be comparatively close together. Where deposits are very flat, under 20 deg., and walls fairly sound, it is often possible to use a sort of long wall system of stoping and to lay tracks in the stopes with self-acting inclines to the levels. In such instances, the interval can be expanded to 250 or even 400 feet. In dips between 20 deg. and 45 deg., tracks are not often possible, and either shoveling or "bumping troughs"[*] are the only help to transport. With shoveling, intervals of 100 feet[**] are most common, and with troughs the distance can be expanded up to 150 or 175 feet.

[Footnote *: Page 136.]

[Footnote **: Intervals given are measured on the dip.]

In dips of over 40 deg. to 50 deg., depending on the smoothness of the foot wall, the distance can again be increased, as stope-transport is greatly simplified, since the stope materials fall out by gravity. In timbered stopes, in dips over about 45 deg., intervals of 150 to 200 feet are possible. In filled stopes intervals of over 150 feet present difficulties in the maintenance of ore-passes, for the wear and tear of longer use often breaks the timbers. In shrinkage-stopes, where no passes are to be maintained and few winzes put through, the interval is sometimes raised to 250 feet. The subject is further discussed under "Stoping."

Another factor bearing on level intervals is the needed insurance of sufficient points of stoping attack to keep up a certain output. This must particularly influence the manager whose mine has but little ore in reserve.



PROTECTION OF LEVELS.—Until recent years, timbering and occasional walling was the only method for the support of the roof, and for forming a platform for a stoping base. Where the rock requires no support sublevels can be used as a stoping base, and timbering for such purpose avoided altogether (Figs. 38, 39, 42). In such cases the main roadway can then be driven on straight lines, either in the walls or in the ore, and used entirely for haulage. The subheading for a stoping base is driven far enough above or below the roadway (depending on whether overhand or underhand stoping is to be used) to leave a supporting pillar which is penetrated by short passes for ore. In overhand stopes, the ore is broken directly on the floor of an upper sublevel; and in underhand stopes, broken directly from the bottom of the sublevel. The method entails leaving a pillar of ore which can be recovered only with difficulty in mines where stope-support is necessary. The question of its adoption is then largely one of the comparative cost of timbering, the extra cost of the sublevel, and the net value of the ore left. In bad swelling veins, or badly crushing walls, where constant repair to timbers would be necessary, the use of a sublevel is a most useful alternative. It is especially useful with stopes to be left open or worked by shrinkage-stoping methods.

If the haulage level, however, is to be the stoping base, some protection to the roadway must be provided. There are three systems in use,—by wood stulls or sets (Figs. 19, 30, 43), by dry-walling with timber caps (Fig. 35), and in some localities by steel sets. Stulls are put up in various ways, and, as their use entails the least difficulty in taking the ore out from beneath the level, they are much favored, but are applicable only in comparatively narrow deposits.

WINZES AND RISES.

These two kinds of openings for connecting two horizons in a mine differ only in their manner of construction. A winze is sunk underhand, while a rise is put up overhand. When the connection between levels is completed, a miner standing at the bottom usually refers to the opening as a rise, and when he goes to the top he calls it a winze. This confusion in terms makes it advisable to refer to all such completed openings as winzes, regardless of how they are constructed.

In actual work, even disregarding water, it costs on the average about 30% less to raise than to sink such openings, for obviously the spoil runs out or is assisted by gravity in one case, and in the other has to be shoveled and hauled up. Moreover, it is easier to follow the ore in a rise than in a winze. It usually happens, however, that in order to gain time both things are done, and for prospecting purposes sinking is necessary.

The number of winzes required depends upon the method of stoping adopted, and is mentioned under "Stoping." After stoping, the number necessary to be maintained open depends upon the necessities of ventilation, of escape, and of passageways for material to be used below. Where stopes are to be filled with waste, more winzes must be kept open than when other methods are used, and these winzes must be in sufficient alignment to permit the continuous flow of material down past the various levels. In order that the winzes should deliver timber and filling to the most advantageous points, they should, in dipping ore-bodies, be as far as possible on the hanging wall side.

DEVELOPMENT IN THE EARLY PROSPECTING STAGE.

The prime objects in the prospecting stage are to expose the ore and to learn regarding the ore-bodies something of their size, their value, metallurgical character, location, dip, strike, etc.,—so much at least as may be necessary to determine the works most suitable for their extraction or values warranting purchase. In outcrop mines there is one rule, and that is "follow the ore." Small temporary inclines following the deposit, even though they are eventually useless; are nine times out of ten justified.

In prospecting deep-level projects, it is usually necessary to layout work which can be subsequently used in operating the mine, because the depth involves works of such considerable scale, even for prospecting, that the initial outlay does not warrant any anticipation of revision. Such works have to be located and designed after a study of the general geology as disclosed in adjoining mines. Practically the only method of supplementing such information is by the use of churn- and diamond-drills.

DRILLING.—Churn-drills are applicable only to comparatively shallow deposits of large volume. They have an advantage over the diamond drill in exposing a larger section and in their application to loose material; but inability to determine the exact horizon of the spoil does not lend them to narrow deposits, and in any event results are likely to be misleading from the finely ground state of the spoil. They are, however, of very great value for preliminary prospecting to shallow horizons.

Two facts in diamond-drilling have to be borne in mind: the indication of values is liable to be misleading, and the deflection of the drill is likely to carry it far away from its anticipated destination. A diamond-drill secures a small section which is sufficiently large to reveal the geology, but the values disclosed in metal mines must be accepted with reservations. The core amounts to but a little sample out of possibly large amounts of ore, which is always of variable character, and the core is most unlikely to represent the average of the deposit. Two diamond-drill holes on the Oroya Brownhill mine both passed through the ore-body. One apparently disclosed unpayable values, the other seemingly showed ore forty feet in width assaying $80 per ton. Neither was right. On the other hand, the predetermination of the location of the ore-body justified expenditure. A recent experiment at Johannesburg of placing a copper wedge in the hole at a point above the ore-body and deflecting the drill on reintroducing it, was successful in giving a second section of the ore at small expense.