The deflection of diamond-drill holes from the starting angle is almost universal. It often amounts to a considerable wandering from the intended course. The amount of such deflection varies with no seeming rule, but it is probable that it is especially affected by the angle at which stratification or lamination planes are inclined to the direction of the hole. A hole has been known to wander in a depth of 1,500 feet more than 500 feet from the point intended. Various instruments have been devised for surveying deep holes, and they should be brought into use before works are laid out on the basis of diamond-drill results, although none of the inventions are entirely satisfactory.



CHAPTER X.

Stoping.

METHODS OF ORE-BREAKING; UNDERHAND STOPES; OVERHAND STOPES; COMBINED STOPE. VALUING ORE IN COURSE OF BREAKING.

There is a great deal of confusion in the application of the word "stoping." It is used not only specifically to mean the actual ore-breaking, but also in a general sense to indicate all the operations of ore-breaking, support of excavations, and transportation between levels. It is used further as a noun to designate the hole left when the ore is taken out. Worse still, it is impossible to adhere to miners' terms without employing it in every sense, trusting to luck and the context to make the meaning clear.

The conditions which govern the method of stoping are in the main:—

a. The dip. b. The width of the deposit. c. The character of the walls. d. The cost of materials. e. The character of the ore.

Every mine, and sometimes every stope in a mine, is a problem special to itself. Any general consideration must therefore be simply an inquiry into the broad principles which govern the adaptability of special methods. A logical arrangement of discussion is difficult, if not wholly impossible, because the factors are partially interdependent and of varying importance.

For discussion the subject may be divided into:

1. Methods of ore-breaking. 2. Methods of supporting excavation. 3. Methods of transport in stopes.

METHODS OF ORE-BREAKING.

The manner of actual ore-breaking is to drill and blast off slices from the block of ground under attack. As rock obviously breaks easiest when two sides are free, that is, when corners can be broken off, the detail of management for blasts is therefore to set the holes so as to preserve a corner for the next cut; and as a consequence the face of the stope shapes into a series of benches (Fig. 22),—inverted benches in the case of overhand stopes (Figs. 20, 21). The size of these benches will in a large measure depend on the depth of the holes. In wide stopes with machine-drills they vary from 7 to 10 feet; in narrow stopes with hand-holes, from two to three feet.



The position of the men in relation to the working face gives rise to the usual primary classification of the methods of stoping. They are:—

1. Underhand stopes, 2. Overhand stopes, 3. Combined stopes.

These terms originated from the direction of the drill-holes, but this is no longer a logical basis of distinction, for underhand holes in overhand stopes,—as in rill-stoping,—are used entirely in some mines (Fig. 21).



UNDERHAND STOPES.—Underhand stopes are those in which the ore is broken downward from the levels. Inasmuch as this method has the advantage of allowing the miner to strike his blows downward and to stand upon the ore when at work, it was almost universal before the invention of powder; and was applied more generally before the invention of machine-drills than since. It is never rightly introduced unless the stope is worked back from winzes through which the ore broken can be let down to the level below, as shown in Figures 22 and 23.



This system can be advantageously applied only in the rare cases in which the walls require little or no support, and where very little or no waste requiring separation is broken with the ore in the stopes. To support the walls in bad ground in underhand stopes would be far more costly than with overhand stopes, for square-set timbering would be most difficult to introduce, and to support the walls with waste and stulls would be even more troublesome. Any waste broken must needs be thrown up to the level above or be stored upon specially built stages—again a costly proceeding.

A further drawback lies in the fact that the broken ore follows down the face of the stope, and must be shoveled off each bench. It thus all arrives at a single point,—the winze,—and must be drawn from a single ore-pass into the level. This usually results not only in more shoveling but in a congestion at the passes not present in overhand stoping, for with that method several chutes are available for discharging ore into the levels. Where the walls require no support and no selection is desired in the stopes, the advantage of the men standing on the solid ore to work, and of having all down holes and therefore drilled wet, gives this method a distinct place. In using this system, in order to protect the men, a pillar is often left under the level by driving a sublevel, the pillar being easily recoverable later. The method of sublevels is of advantage largely in avoiding the timbering of levels.



OVERHAND STOPES.—By far the greatest bulk of ore is broken overhand, that is broken upward from one level to the next above. There are two general forms which such stopes are given,—"horizontal" and "rill."



The horizontal "flat-back" or "long-wall" stope, as it is variously called, shown in Figure 24, is operated by breaking the ore in slices parallel with the levels. In rill-stoping the ore is cut back from the winzes in such a way that a pyramid-shaped room is created, with its apex in the winze and its base at the level (Figs. 25 and 26). Horizontal or flat-backed stopes can be applied to almost any dip, while "rill-stoping" finds its most advantageous application where the dip is such that the ore will "run," or where it can be made to "run" with a little help. The particular application of the two systems is dependent not only on the dip but on the method of supporting the excavation and the ore. With rill-stoping, it is possible to cut the breaking benches back horizontally from the winzes (Fig. 25), or to stagger the cuts in such a manner as to take the slices in a descending angle (Figs. 21 and 26).



In the "rill" method of incline cuts, all the drill-holes are "down" holes (Fig. 21), and can be drilled wet, while in horizontal cuts or flat-backed stopes, at least part of the holes must be "uppers" (Fig. 20). Aside from the easier and cheaper drilling and setting up of machines with this kind of "cut," there is no drill dust,—a great desideratum in these days of miners' phthisis. A further advantage in the "rill" cut arises in cases where horizontal jointing planes run through the ore of a sort from which unduly large masses break away in "flat-back" stopes. By the descending cut of the "rill" method these calamities can be in a measure avoided. In cases of dips over 40 deg. the greatest advantage in "rill" stoping arises from the possibility of pouring filling or timber into the stope from above with less handling, because the ore and material will run down the sides of the pyramid (Figs. 32 and 34). Thus not only is there less shoveling required, but fewer ore-passes and a less number of preliminary winzes are necessary, and a wider level interval is possible. This matter will be gone into more fully later.



COMBINED STOPES.—A combined stope is made by the coincident working of the underhand and "rill" method (Fig. 27). This order of stope has the same limitations in general as the underhand kind. For flat veins with strong walls, it has a great superiority in that the stope is carried back more or less parallel with the winzes, and thus broken ore after blasting lies in a line on the gradient of the stope. It is, therefore, conveniently placed for mechanical stope haulage. A further advantage is gained in that winzes may be placed long distances apart, and that men are not required, either when at work or passing to and from it, to be ever far from the face, and they are thus in the safest ground, so that timber and filling protection which may be otherwise necessary is not required. This method is largely used in South Africa.



MINIMUM WIDTH OF STOPES.—The minimum stoping width which can be consistently broken with hand-holes is about 30 inches, and this only where there is considerable dip to the ore. This space is so narrow that it is of doubtful advantage in any case, and 40 inches is more common in narrow mines, especially where worked with white men. Where machine-drills are used about 4 feet is the minimum width feasible.

RESUING.—In very narrow veins where a certain amount of wall-rock must be broken to give working space, it pays under some circumstances to advance the stope into the wall-rock ahead of the ore, thus stripping the ore and enabling it to be broken separately. This permits of cleaner selection of the ore; but it is a problem to be worked out in each case, as to whether rough sorting of some waste in the stopes, or further sorting at surface with inevitable treatment of some waste rock, is more economical than separate stoping cuts and inevitably wider stopes.

VALUING ORE IN COURSE OF BREAKING.—There are many ores whose payability can be determined by inspection, but there are many of which it cannot. Continuous assaying is in the latter cases absolutely necessary to avoid the treatment of valueless material. In such instances, sampling after each stoping-cut is essential, the unprofitable ore being broken down and used as waste. Where values fade into the walls, as in impregnation deposits, the width of stopes depends upon the limit of payability. In these cases, drill-holes are put into the walls and the drillings assayed. If the ore is found profitable, the holes are blasted out. The gauge of what is profitable in such situations is not dependent simply upon the average total working costs of the mine, for ore in that position can be said to cost nothing for development work and administration; moreover, it is usually more cheaply broken than the average breaking cost, men and machines being already on the spot.



CHAPTER XI.

Methods of Supporting Excavation.

TIMBERING; FILLING WITH WASTE; FILLING WITH BROKEN ORE; PILLARS OF ORE; ARTIFICIAL PILLARS; CAVING SYSTEM.

Most stopes require support to be given to the walls and often to the ore itself. Where they do require support there are five principal methods of accomplishing it. The application of any particular method depends upon the dip, width of ore-body, character of the ore and walls, and cost of materials. The various systems are by:—

1. Timbering. 2. Filling with waste. 3. Filling with broken ore subsequently withdrawn. 4. Pillars of ore. 5. Artificial pillars built of timbers and waste. 6. Caving.

TIMBERING.—At one time timbering was the almost universal means of support in such excavations, but gradually various methods for the economical application of waste and ore itself have come forward, until timbering is fast becoming a secondary device. Aside from economy in working without it, the dangers of creeps, or crushing, and of fires are sufficient incentives to do away with wood as far as possible.

There are three principal systems of timber support to excavations,—by stulls, square-sets, and cribs.

Stulls are serviceable only where the deposit is so narrow that the opening can be bridged by single timbers between wall and wall (Figs. 28 and 43). This system can be applied to any dip and is most useful in narrow deposits where the walls are not too heavy. Stulls in inclined deposits are usually set at a slightly higher angle than that perpendicular to the walls, in order that the vertical pressure of the hanging wall will serve to tighten them in position. The "stull" system can, in inclined deposits, be further strengthened by building waste pillars against them, in which case the arrangement merges into the system of artificial pillars.



Square-sets (Figs. 29 and 30), that is, trusses built in the opening as the ore is removed, are applicable to almost any dip or width of ore, but generally are applied only in deposits too wide, or to rock too heavy, for stulls. Such trusses are usually constructed on vertical and horizontal lines, and while during actual ore-breaking the strains are partially vertical, ultimately, however, when the weight of the walls begins to be felt, these strains, except in vertical deposits, come at an angle to lines of strength in the trusses, and therefore timber constructions of this type present little ultimate resistance (Fig. 30). Square-set timbers are sometimes set to present the maximum resistance to the direction of strain, but the difficulties of placing them in position and variations in the direction of strain on various parts of the stope do not often commend the method. As a general rule square-sets on horizontal lines answer well enough for the period of actual ore-breaking. The crushing or creeps is usually some time later; and if the crushing may damage the whole mine, their use is fraught with danger. Reenforcement by building in waste is often resorted to. When done fully, it is difficult to see the utility of the enclosed timber, for entire waste-filling would in most cases be cheaper and equally efficient.



There is always, with wood constructions, as said before, the very pertinent danger of subsequent crushing and of subsidence in after years, and the great risk of fires. Both these disasters have cost Comstock and Broken Hill mines, directly or indirectly, millions of dollars, and the outlay on timber and repairs one way or another would have paid for the filling system ten times over. There are cases where, by virtue of the cheapness of timber, "square-setting" is the most economical method. Again, there are instances where the ore lies in such a manner—particularly in limestone replacements—as to preclude other means of support. These cases are being yearly more and more evaded by the ingenuity of engineers in charge. The author believes it soon will be recognized that the situation is rare indeed where complete square-setting is necessarily without an economical alternative. An objection is sometimes raised to filling in favor of timber, in that if it become desirable to restope the walls for low-grade ore left behind, such stopes could only be entered by drawing the filling, with consequent danger of total collapse. Such a contingency can be provided for in large ore-bodies by installing an outer shell of sets of timber around the periphery of the stope and filling the inside with waste. If the crushing possibilities are too great for this method then, the subsequent recovery of ore is hopeless in any event. In narrow ore-bodies with crushing walls recovery of ore once left behind is not often possible.

The third sort of timber constructions are cribs, a "log-house" sort of structure usually filled with waste, and more fully discussed under artificial pillars (Fig. 31). The further comparative merits of timbering with other methods will be analyzed as the different systems are described.

FILLING WITH WASTE.—The system of filling stope-excavations completely with waste in alternating progress with ore-breaking is of wide and increasingly general application (Figs. 32, 33, 34, 35).

Although a certain amount of waste is ordinarily available in the stopes themselves, or from development work in the mine, such a supply must usually be supplemented from other directions. Treatment residues afford the easiest and cheapest handled material. Quarried rock ranks next, and in default of any other easy supply, materials from crosscuts driven into the stope-walls are sometimes resorted to.

In working the system to the best advantage, the winzes through the block of ore under attack are kept in alignment with similar openings above, in order that filling may be poured through the mine from the surface or any intermediate point. Winzes to be used for filling should be put on the hanging-wall side of the area to be filled, for the filling poured down will then reach the foot-wall side of the stopes with a minimum of handling. In some instances, one special winze is arranged for passing all filling from the surface to a level above the principal stoping operations; and it is then distributed along the levels into the winzes, and thus to the operating stopes, by belt-conveyors.



In this system of stope support the ore is broken at intervals alternating with filling. If there is danger of much loss from mixing broken ore and filling, "sollars" of boards or poles are laid on the waste. If the ore is very rich, old canvas or cowhides are sometimes put under the boards. Before the filling interval, the ore passes are built close to the face above previous filling and their tops covered temporarily to prevent their being filled with running waste. If the walls are bad, the filling is kept close to the face. If the unbroken ore requires support, short stulls set on the waste (as in Fig. 39) are usually sufficient until the next cut is taken off, when the timber can be recovered. If stulls are insufficient, cribs or bulkheads (Fig. 31) are also used and often buried in the filling.



Both flat-backed and rill-stope methods of breaking are employed in conjunction with filled stopes. The advantages of the rill-stopes are so patent as to make it difficult to understand why they are not universally adopted when the dip permits their use at all. In rill-stopes (Figs. 32 and 34) the waste flows to its destination with a minimum of handling. Winzes and ore-passes are not required with the same frequency as in horizontal breaking, and the broken ore always lies on the slope towards the passes and is therefore also easier to shovel. In flat-backed stopes (Fig. 33) winzes must be put in every 50 feet or so, while in rill-stopes they can be double this distance apart. The system is applicable by modification to almost any width of ore. It finds its most economical field where the dip of the stope floor is over 45 deg., when waste and ore, with the help of the "rill," will flow to their destination. For dips from under about 45 deg. to about 30 deg. or 35 deg., where the waste and ore will not "flow" easily, shoveling can be helped by the use of the "rill" system and often evaded altogether, if flow be assisted by a sheet-iron trough described in the discussion of stope transport. Further saving in shoveling can be gained in this method, by giving a steeper pitch to the filling winzes and to the ore-passes, by starting them from crosscuts in the wall, and by carrying them at greater angles than the pitch of the ore (Fig. 36). These artifices combined have worked out most economically on several mines within the writer's experience, with the dip as flat as 30 deg. For very flat dips, where filling is to be employed, rill-stoping has no advantage over flat-backed cuts, and in such cases it is often advisable to assist stope transport by temporary tracks and cars which obviously could not be worked on the tortuous contour of a rill-stope, so that for dips under 30 deg. advantage lies with "flat-backed" ore-breaking.



On very wide ore-bodies where the support of the standing ore itself becomes a great problem, the filling system can be applied by combining it with square-setting. In this case the stopes are carried in panels laid out transversally to the strike as wide as the standing strength of the ore permits. On both sides of each panel a fence of lagged square-sets is carried up and the area between is filled with waste. The panels are stoped out alternately. The application of this method at Broken Hill will be described later. (See pages 120 and Figs. 41 and 42.) The same type of wide ore-body can be managed also on the filling system by the use of frequent "bulkheads" to support the ore (Fig. 31).

Compared with timbering methods, filling has the great advantage of more effective support to the mine, less danger of creeps, and absolute freedom from the peril of fire. The relative expense of the two systems is determined by the cost of materials and labor. Two extreme cases illustrate the result of these economic factors with sufficient clearness. It is stated that the cost of timbering stopes on the Le Roi Mine by square-sets is about 21 cents per ton of ore excavated. In the Ivanhoe mine of West Australia the cost of filling stopes with tailings is about 22 cents per ton of ore excavated. At the former mine the average cost of timber is under $10 per M board-measure, while at the latter its price would be $50 per M board-measure; although labor is about of the same efficiency and wage, the cost in the Ivanhoe by square-setting would be about 65 cents per ton of ore broken. In the Le Roi, on the other hand, no residues are available for filling. To quarry rock or drive crosscuts into the walls might make this system cost 65 cents per ton of ore broken if applied to that mine. The comparative value of the filling method with other systems will be discussed later.

FILLING WITH BROKEN ORE SUBSEQUENTLY WITHDRAWN.—This order of support is called by various names, the favorite being "shrinkage-stoping." The method is to break the ore on to the roof of the level, and by thus filling the stope with broken ore, provide temporary support to the walls and furnish standing floor upon which to work in making the next cut (Figs. 37, 38, and 39.) As broken material occupies 30 to 40% more space than rock in situ, in order to provide working space at the face, the broken ore must be drawn from along the level after each cut. When the area attacked is completely broken through from level to level, the stope will be full of loose broken ore, which is then entirely drawn off.

A block to be attacked by this method requires preliminary winzes only at the extremities of the stope,—for entry and for ventilation. Where it is desired to maintain the winzes after stoping, they must either be strongly timbered and lagged on the stope side, be driven in the walls, or be protected by a pillar of ore (Fig. 37). The settling ore and the crushing after the stope is empty make it difficult to maintain timbered winzes.



Where it can be done without danger to the mine, the empty stopes are allowed to cave. If such crushing would be dangerous, either the walls must be held up by pillars of unbroken ore, as in the Alaska Treadwell, where large "rib" pillars are left, or the open spaces must be filled with waste. Filling the empty stope is usually done by opening frequent passes along the base of the filled stope above, and allowing the material of the upper stope to flood the lower one. This program continued upwards through the mine allows the whole filling of the mine to descend gradually and thus requires replenishment only into the top. The old stopes in the less critical and usually exhausted territory nearer the surface are sometimes left without replenishing their filling.

The weight of broken ore standing at such a high angle as to settle rapidly is very considerable upon the level; moreover, at the moment when the stope is entirely drawn off, the pressure of the walls as well is likely to be very great. The roadways in this system therefore require more than usual protection. Three methods are used: (a) timbering; (b) driving a sublevel in the ore above the main roadway as a stoping-base, thus leaving a pillar of ore over the roadway (Fig. 39); (c) by dry-walling the levels, as in the Baltic mine, Michigan (Figs. 34 and 35). By the use of sublevels the main roadways are sometimes driven in the walls (Fig. 38) and in many cases all timbering is saved. To recover pillars left below sublevels is a rather difficult task, especially if the old stope above is caved or filled. The use of pillars in substitution for timber, if the pillars are to be lost, is simply a matter of economics as to whether the lost ore would repay the cost of other devices.



Frequent ore-chutes through the level timbers, or from the sublevels, are necessary to prevent lodgment of broken ore between such passes, because it is usually too dangerous for men to enter the emptying stope to shovel out the lodged remnants. Where the ore-body is wide, and in order that there may be no lodgment of ore, the timbers over the level are set so as to form a trough along the level; or where pillars are left, they are made "A"-shaped between the chutes, as indicated in Figure 37.



The method of breaking the ore in conjunction with this means of support in comparatively narrow deposits can be on the rill, in order to have the advantage of down holes. Usually, however, flat-back or horizontal cuts are desirable, as in such an arrangement it is less troublesome to regulate the drawing of the ore so as to provide proper head room. Where stopes are wide, ore is sometimes cut arch-shaped from wall to wall to assure its standing. Where this method of support is not of avail, short, sharply tapering stulls are put in from the broken ore to the face (Fig. 39). When the cut above these stulls is taken out, they are pulled up and are used again.

This method of stoping is only applicable when:—

1. The deposit dips over 60 deg., and thus broken material will freely settle downward to be drawn off from the bottom.

2. The ore is consistently payable in character. No selection can be done in breaking, as all material broken must be drawn off together.

3. The hanging wall is strong, and will not crush or spall off waste into the ore.

4. The ore-body is regular in size, else loose ore will lodge on the foot wall. Stopes opened in this manner when partially empty are too dangerous for men to enter for shoveling out remnants.

The advantages of this system over others, where it is applicable, are:—

(a) A greater distance between levels can be operated and few winzes and rises are necessary, thus a great saving of development work can be effected. A stope 800 to 1000 feet long can be operated with a winze at either end and with levels 200 or 220 feet apart.

(b) There is no shoveling in the stopes at all.

(c) No timber is required. As compared with timbering by stulling, it will apply to stopes too wide and walls too heavy for this method. Moreover, little staging is required for working the face, since ore can be drawn from below in such a manner as to allow just the right head room.

(d) Compared to the system of filling with waste, coincidentally with breaking (second method), it saves altogether in some cases the cost of filling. In any event, it saves the cost of ore-passes, of shoveling into them, and of the detailed distribution of the filling.

Compared with other methods, the system has the following disadvantages, that:

A. The ore requires to be broken in the stopes to a degree of fineness which will prevent blocking of the chutes at the level. When pieces too large reach the chutes, nothing will open them but blasting,—to the damage of timbers and chutes. Some large rocks are always liable to be buried in the course of ore-breaking.

B. Practically no such perfection of walls exists, but some spalling of waste into the ore will take place. A crushing of the walls would soon mean the loss of large amounts of ore.

C. There is no possibility of regulating the mixture of grade of ore by varying the working points. It is months after the ore is broken before it can reach the levels.

D. The breaking of 60% more ore than immediate treatment demands results in the investment of a considerable sum of money. An equilibrium is ultimately established in a mine worked on this system when a certain number of stopes full of completely broken ore are available for entire withdrawal, and there is no further accumulation. But, in any event, a considerable amount of broken ore must be held in reserve. In one mine worked on this plan, with which the writer has had experience, the annual production is about 250,000 tons and the broken ore represents an investment which, at 5%, means an annual loss of interest amounting to 7 cents per ton of ore treated.

E. A mine once started on the system is most difficult to alter, owing to the lack of frequent winzes or passes. Especially is this so if the only alternative is filling, for an alteration to the system of filling coincident with breaking finds the mine short of filling winzes. As the conditions of walls and ore often alter with depth, change of system may be necessary and the situation may become very embarrassing.

F. The restoping of the walls for lower-grade ore at a later period is impossible, for the walls of the stope will be crushed, or, if filled with waste, will usually crush when it is drawn off to send to a lower stope.

The system has much to recommend it where conditions are favorable. Like all other alternative methods of mining, it requires the most careful study in the light of the special conditions involved. In many mines it can be used for some stopes where not adaptable generally. It often solves the problem of blind ore-bodies, for they can by this means be frequently worked with an opening underneath only. Thus the cost of driving a roadway overhead is avoided, which would be required if timber or coincident filling were the alternatives. In such cases ventilation can be managed without an opening above, by so directing the current of air that it will rise through a winze from the level below, flow along the stope and into the level again at the further end of the stope through another winze.



SUPPORT BY PILLARS OF ORE.—As a method of mining metals of the sort under discussion, the use of ore-pillars except in conjunction with some other means of support has no general application. To use them without assistance implies walls sufficiently strong to hold between pillars; to leave them permanently anywhere implies that the ore abandoned would not repay the labor and the material of a substitute. There are cases of large, very low-grade mines where to abandon one-half the ore as pillars is more profitable than total extraction, but the margin of payability in such ore must be very, very narrow. Unpayable spots are always left as pillars, for obvious reasons. Permanent ore-pillars as an adjunct to other methods of support are in use. Such are the rib-pillars in the Alaska Treadwell, the form of which is indicated by the upward extension of the pillars adjacent to the winzes, shown in Figure 37. Always a careful balance must be cast as to the value of the ore left, and as to the cost of a substitute, because every ore-pillar can be removed at some outlay. Temporary pillars are not unusual, particularly to protect roadways and shafts. They are, when left for these purposes, removed ultimately, usually by beginning at the farther end and working back to the final exit.



A form of temporary ore-pillars in very wide deposits is made use of in conjunction with both filling and timbering (Figs. 37, 39, 40). In the use of temporary pillars for ore-bodies 100 to 250 feet wide at Broken Hill, stopes are carried up at right angles to the strike, each fifty feet wide and clear across the ore-body (Figs. 41 and 42). A solid pillar of the same width is left in the first instance between adjacent stopes, and the initial series of stopes are walled with one square-set on the sides as the stope is broken upward. The room between these two lines of sets is filled with waste alternating with ore-breaking in the usual filling method. When the ore from the first group of alternate stopes (ABC, Fig. 42) is completely removed, the pillars are stoped out and replaced with waste. The square-sets of the first set of stopes thus become the boundaries of the second set. Entry and ventilation are obtained through these lines of square-sets, and the ore is passed out of the stopes through them.



ARTIFICIAL PILLARS.—This system also implies a roof so strong as not to demand continuous support. Artificial pillars are built in many different ways. The method most current in fairly narrow deposits is to reenforce stulls by packing waste above them (Figs. 43 and 44). Not only is it thus possible to economize in stulls by using the waste which accumulates underground, but the principle applies also to cases where the stulls alone are not sufficient support, and yet where complete filling or square-setting is unnecessary. When the conditions are propitious for this method, it has the comparative advantage over timber systems of saving timber, and over filling systems of saving imported filling. Moreover, these constructions being pillar-shaped (Fig. 44), the intervals between them provide outlets for broken ore, and specially built passes are unnecessary. The method has two disadvantages as against the square-set or filling process, in that more staging must be provided from which to work, and in stopes over six feet the erection of machine-drill columns is tedious and costly in time and wages.



In wide deposits of markedly flat, irregular ore-bodies, where a definite system is difficult and where timber is expensive, cribs of cord-wood or logs filled with waste after the order shown in Figure 31, often make fairly sound pillars. They will not last indefinitely and are best adapted to the temporary support of the ore-roof pending filling. The increased difficulty in setting up machine drills in such stopes adds to the breaking costs,—often enough to warrant another method of support.



CAVING SYSTEMS.—This method, with variations, has been applied to large iron deposits, to the Kimberley diamond mines, to some copper mines, but in general it has little application to the metal mines under consideration, as few ore-bodies are of sufficiently large horizontal area. The system is dependent upon a large area of loose or "heavy" ground pressing directly on the ore with weight, such that if the ore be cut into pillars, these will crush. The details of the system vary, but in general the modus operandi is to prepare roadways through the ore, and from the roadways to put rises, from which sublevels are driven close under the floating mass of waste and ore,—sometimes called the "matte" (Fig. 45). The pillars between these sublevels are then cut away until the weight above crushes them down. When all the crushed ore which can be safely reached is extracted, retreat is made and another series of subopenings is then driven close under the "matte." The pillar is reduced until it crushes and the operation is repeated. Eventually the bottom strata of the "matte" become largely ore, and a sort of equilibrium is reached when there is not much loss in this direction. "Top slicing" is a variation of the above method by carrying a horizontal stope from the rises immediately under the matte, supporting the floating material with timber. At Kimberley the system is varied in that galleries are run out to the edge of the diamond-iferous area and enlarged until the pillar between crushes.

In the caving methods, between 40 and 50% of the ore is removed by the preliminary openings, and as they are all headings of some sort, the average cost per ton of this particular ore is higher than by ordinary stoping methods. On the other hand, the remaining 50 to 60% of the ore costs nothing to break, and the average cost is often remarkably low. As said, the system implies bodies of large horizontal area. They must start near enough to the surface that the whole superincumbent mass may cave and give crushing weight, or the immediately overhanging roof must easily cave. All of these are conditions not often met with in mines of the character under review.



CHAPTER XII.

Mechanical Equipment.

CONDITIONS BEARING ON MINE EQUIPMENT; WINDING APPLIANCES; HAULAGE EQUIPMENT IN SHAFTS; LATERAL UNDERGROUND TRANSPORT; TRANSPORT IN STOPES.

There is no type of mechanical engineering which presents such complexities in determination of the best equipment as does that of mining. Not only does the economic side dominate over pure mechanics, but machines must be installed and operated under difficulties which arise from the most exceptional and conflicting conditions, none of which can be entirely satisfied. Compromise between capital outlay, operating efficiency, and conflicting demands is the key-note of the work.

These compromises are brought about by influences which lie outside the questions of mechanics of individual machines, and are mainly as follows:—

1. Continuous change in horizon of operations. 2. Uncertain life of the enterprise. 3. Care and preservation of human life. 4. Unequal adaptability of power transmission mediums. 5. Origin of power.

First.—The depth to be served and the volume of ore and water to be handled, are not only unknown at the initial equipment, but they are bound to change continuously in quantity, location, and horizon with the extension of the workings.

Second.—From the mine manager's point of view, which must embrace that of the mechanical engineer, further difficulty presents itself because the life of the enterprise is usually unknown, and therefore a manifest necessity arises for an economic balance of capital outlay and of operating efficiency commensurate with the prospects of the mine. Moreover, the initial capital is often limited, and makeshifts for this reason alone must be provided. In net result, no mineral deposit of speculative ultimate volume of ore warrants an initial equipment of the sort that will meet every eventuality, or of the kind that will give even the maximum efficiency which a free choice of mining machinery could obtain.

Third.—In the design and selection of mining machines, the safety of human life, the preservation of the health of workmen under conditions of limited space and ventilation, together with reliability and convenience in installing and working large mechanical tools, all dominate mechanical efficiency. For example, compressed-air transmission of power best meets the requirements of drilling, yet the mechanical losses in the generation, the transmission, and the application of compressed air probably total, from first to last, 70 to 85%.

Fourth.—All machines, except those for shaft haulage, must be operated by power transmitted from the surface, as obviously power generation underground is impossible. The conversion of power into a transmission medium and its transmission are, at the outset, bound to be the occasions of loss. Not only are the various forms of transmission by steam, electricity, compressed air, or rods, of different efficiency, but no one system lends itself to universal or economical application to all kinds of mining machines. Therefore it is not uncommon to find three or four different media of power transmission employed on the same mine. To illustrate: from the point of view of safety, reliability, control, and in most cases economy as well, we may say that direct steam is the best motive force for winding-engines; that for mechanical efficiency and reliability, rods constitute the best media of power transmission to pumps; that, considering ventilation and convenience, compressed air affords the best medium for drills. Yet there are other conditions as to character of the work, volume of water or ore, and the origin of power which must in special instances modify each and every one of these generalizations. For example, although pumping water with compressed air is mechanically the most inefficient of devices, it often becomes the most advantageous, because compressed air may be of necessity laid on for other purposes, and the extra power required to operate a small pump may be thus most cheaply provided.

Fifth.—Further limitations and modifications arise out of the origin of power, for the sources of power have an intimate bearing on the type of machine and media of transmission. This very circumstance often compels giving away efficiency and convenience in some machines to gain more in others. This is evident enough if the principal origins of power generation be examined. They are in the main as follows:—

a. Water-power available at the mine. b. Water-power available at a less distance than three or four miles. c. Water-power available some miles away, thus necessitating electrical transmission (or purchased electrical power). d. Steam-power to be generated at the mine. e. Gas-power to be generated at the mine.

a. With water-power at the mine, winding engines can be operated by direct hydraulic application with a gain in economy over direct steam, although with the sacrifice of control and reliability. Rods for pumps can be driven directly with water, but this superiority in working economy means, as discussed later, a loss of flexibility and increased total outlay over other forms of transmission to pumps. As compressed air must be transmitted for drills, the compressor would be operated direct from water-wheels, but with less control in regularity of pressure delivery.

b. With water-power a short distance from the mine, it would normally be transmitted either by compressed air or by electricity. Compressed-air transmission would better satisfy winding and drilling requirements, but would show a great comparative loss in efficiency over electricity when applied to pumping. Despite the latter drawback, air transmission is a method growing in favor, especially in view of the advance made in effecting compression by falling water.

c. In the situation of transmission too far for using compressed air, there is no alternative but electricity. In these cases, direct electric winding is done, but under such disadvantages that it requires a comparatively very cheap power to take precedence over a subsidiary steam plant for this purpose. Electric air-compressors work under the material disadvantage of constant speed on a variable load, but this installation is also a question of economics. The pumping service is well performed by direct electrical pumps.

d. In this instance, winding and air-compression are well accomplished by direct steam applications; but pumping is beset with wholly undesirable alternatives, among which it is difficult to choose.

e. With internal combustion engines, gasoline (petrol) motors have more of a position in experimental than in systematic mining, for their application to winding and pumping and drilling is fraught with many losses. The engine must be under constant motion, and that, too, with variable loads. Where power from producer gas is used, there is a greater possibility of installing large equipments, and it is generally applied to the winding and lesser units by conversion into compressed air or electricity as an intermediate stage.

One thing becomes certain from these examples cited, that the right installation for any particular portion of the mine's equipment cannot be determined without reference to all the others. The whole system of power generation for surface work, as well as the transmission underground, must be formulated with regard to furnishing the best total result from all the complicated primary and secondary motors, even at the sacrifice of some members.

Each mine is a unique problem, and while it would be easy to sketch an ideal plant, there is no mine within the writer's knowledge upon which the ideal would, under the many variable conditions, be the most economical of installation or the most efficient of operation. The dominant feature of the task is an endeavor to find a compromise between efficiency and capital outlay. The result is a series of choices between unsatisfying alternatives, a number of which are usually found to have been wrong upon further extension of the mine in depth.

In a general way, it may be stated that where power is generated on the mine, economy in labor of handling fuel, driving engines, generation and condensing steam where steam is used, demand a consolidated power plant for the whole mine equipment. The principal motors should be driven direct by steam or gas, with power distribution by electricity to all outlying surface motors and sometimes to underground motors, and also to some underground motors by compressed air.

Much progress has been made in the past few years in the perfection of larger mining tools. Inherently many of our devices are of a wasteful character, not only on account of the need of special forms of transmission, but because they are required to operate under greatly varying loads. As an outcome of transmission losses and of providing capacity to cope with heavy peak loads, their efficiency on the basis of actual foot-pounds of work accomplished is very low.

The adoption of electric transmission in mine work, while in certain phases beneficial, has not decreased the perplexity which arises from many added alternatives, none of which are as yet a complete or desirable answer to any mine problem. When a satisfactory electric drill is invented, and a method is evolved of applying electricity to winding-engines that will not involve such abnormal losses due to high peak load then we will have a solution to our most difficult mechanical problems, and electricity will deserve the universal blessing which it has received in other branches of mechanical engineering.

It is not intended to discuss mine equipment problems from the machinery standpoint,—there are thousands of different devices,—but from the point of view of the mine administrator who finds in the manufactory the various machines which are applicable, and whose work then becomes that of choosing, arranging, and operating these tools.

The principal mechanical questions of a mine may be examined under the following heads:—

1. Shaft haulage. 2. Lateral underground transport. 3. Drainage. 4. Rock drilling. 5. Workshops. 6. Improvements in equipment.

SHAFT HAULAGE.

WINDING APPLIANCES.—No device has yet been found to displace the single load pulled up the shaft by winding a rope on a drum. Of driving mechanisms for drum motors the alternatives are the steam-engine, the electrical motor, and infrequently water-power or gas engines.

All these have to cope with one condition which, on the basis of work accomplished, gives them a very low mechanical efficiency. This difficulty is that the load is intermittent, and it must be started and accelerated at the point of maximum weight, and from that moment the power required diminishes to less than nothing at the end of the haul. A large number of devices are in use to equalize partially the inequalities of the load at different stages of the lift. The main lines of progress in this direction have been:—

a. The handling of two cages or skips with one engine or motor, the descending skip partially balancing the ascending one. b. The use of tail-ropes or balance weights to compensate the increasing weight of the descending rope. c. The use of skips instead of cages, thus permitting of a greater percentage of paying load. d. The direct coupling of the motor to the drum shaft. e. The cone-shaped construction of drums,—this latter being now largely displaced by the use of the tail-rope.

The first and third of these are absolutely essential for anything like economy and speed; the others are refinements depending on the work to be accomplished and the capital available.

Steam winding-engines require large cylinders to start the load, but when once started the requisite power is much reduced and the load is too small for steam economy. The throttling of the engine for controlling speed and reversing the engine at periodic stoppages militates against the maximum expansion and condensation of the steam and further increases the steam consumption. In result, the best of direct compound condensing engines consume from 60 to 100 pounds of steam per horse-power hour, against a possible efficiency of such an engine working under constant load of less than 16 pounds of steam per horse-power hour.

It is only within very recent years that electrical motors have been applied to winding. Even yet, all things considered, this application is of doubtful value except in localities of extremely cheap electrical power. The constant speed of alternating current motors at once places them at a disadvantage for this work of high peak and intermittent loads. While continuous-current motors can be made to partially overcome this drawback, such a current, where power is purchased or transmitted a long distance, is available only by conversion, which further increases the losses. However, schemes of electrical winding are in course of development which bid fair, by a sort of storage of power in heavy fly-wheels or storage batteries after the peak load, to reduce the total power consumption; but the very high first cost so far prevents their very general adoption for metal mining.

Winding-engines driven by direct water- or gas-power are of too rare application to warrant much discussion. Gasoline driven hoists have a distinct place in prospecting and early-stage mining, especially in desert countries where transport and fuel conditions are onerous, for both the machines and their fuel are easy of transport. As direct gas-engines entail constant motion of the engine at the power demand of the peak load, they are hopeless in mechanical efficiency.

Like all other motors in mining, the size and arrangement of the motor and drum are dependent upon the duty which they will be called upon to perform. This is primarily dependent upon the depth to be hoisted from, the volume of the ore, and the size of the load. For shallow depths and tonnages up to, say, 200 tons daily, geared engines have a place on account of their low capital cost. Where great rope speed is not essential they are fully as economical as direct-coupled engines. With great depths and greater capacities, speed becomes a momentous factor, and direct-coupled engines are necessary. Where the depth exceeds 3,000 feet, another element enters which has given rise to much debate and experiment; that is, the great increase of starting load due to the increased length and size of ropes and the drum space required to hold it. So far the most advantageous device seems to be the Whiting hoist, a combination of double drums and tail rope.

On mines worked from near the surface, where depth is gained by the gradual exhaustion of the ore, the only prudent course is to put in a new hoist periodically, when the demand for increased winding speed and power warrants. The lack of economy in winding machines is greatly augmented if they are much over-sized for the duty. An engine installed to handle a given tonnage to a depth of 3,000 feet will have operated with more loss during the years the mine is progressing from the surface to that depth than several intermediate-sized engines would have cost. On most mines the uncertainty of extension in depth would hardly warrant such a preliminary equipment. More mines are equipped with over-sized than with under-sized engines. For shafts on going metal mines where the future is speculative, an engine will suffice whose size provides for an extension in depth of 1,000 feet beyond that reached at the time of its installation. The cost of the engine will depend more largely upon the winding speed desired than upon any other one factor. The proper speed to be arranged is obviously dependent upon the depth of the haulage, for it is useless to have an engine able to wind 3,000 feet a minute on a shaft 500 feet deep, since it could never even get under way; and besides, the relative operating loss, as said, would be enormous.

HAULAGE EQUIPMENT IN THE SHAFT.—Originally, material was hoisted through shafts in buckets. Then came the cage for transporting mine cars, and in more recent years the "skip" has been developed. The aggrandized bucket or "kibble" of the Cornishman has practically disappeared, but the cage still remains in many mines. The advantages of the skip over the cage are many. Some of them are:—

a. It permits 25 to 40% greater load of material in proportion to the dead weight of the vehicle. b. The load can be confined within a smaller horizontal space, thus the area of the shaft need not be so great for large tonnages. c. Loading and discharging are more rapid, and the latter is automatic, thus permitting more trips per hour and requiring less labor. d. Skips must be loaded from bins underground, and by providing in the bins storage capacity, shaft haulage is rendered independent of the lateral transport in the mine, and there are no delays to the engine awaiting loads. The result is that ore-winding can be concentrated into fewer hours, and indirect economies in labor and power are thus effected. e. Skips save the time of the men engaged in the lateral haulage, as they have no delay waiting for the winding engine.

Loads equivalent to those from skips are obtained in some mines by double-decked cages; but, aside from waste weight of the cage, this arrangement necessitates either stopping the engine to load the lower deck, or a double-deck loading station. Double-deck loading stations are as costly to install and more expensive to work than skip-loading station ore-bins. Cages are also constructed large enough to take as many as four trucks on one deck. This entails a shaft compartment double the size required for skips of the same capacity, and thus enormously increases shaft cost without gaining anything.

Altogether the advantages of the skip are so certain and so important that it is difficult to see the justification for the cage under but a few conditions. These conditions are those which surround mines of small output where rapidity of haulage is no object, where the cost of station-bins can thus be evaded, and the convenience of the cage for the men can still be preserved. The easy change of the skip to the cage for hauling men removes the last objection on larger mines. There occurs also the situation in which ore is broken under contract at so much per truck, and where it is desirable to inspect the contents of the truck when discharging it, but even this objection to the skip can be obviated by contracting on a cubic-foot basis.

Skips are constructed to carry loads of from two to seven tons, the general tendency being toward larger loads every year. One of the most feasible lines of improvement in winding is in the direction of larger loads and less speed, for in this way the sum total of dead weight of the vehicle and rope to the tonnage of ore hauled will be decreased, and the efficiency of the engine will be increased by a less high peak demand, because of this less proportion of dead weight and the less need of high acceleration.

LATERAL UNDERGROUND TRANSPORT.

Inasmuch as the majority of metal mines dip at considerable angles, the useful life of a roadway in a metal mine is very short because particular horizons of ore are soon exhausted. Therefore any method of transport has to be calculated upon a very quick redemption of the capital laid out. Furthermore, a roadway is limited in its daily traffic to the product of the stopes which it serves.

MEN AND ANIMALS.—Some means of transport must be provided, and the basic equipment is light tracks with push-cars, in capacity from half a ton to a ton. The latter load is, however, too heavy to be pushed by one man. As but one car can be pushed at a time, hand-trucking is both slow and expensive. At average American or Australian wages, the cost works out between 25 and 35 cents a ton per mile. An improvement of growing import where hand-trucking is necessary is the overhead mono-rail instead of the track.

If the supply to any particular roadway is such as to fully employ horses or mules, the number of cars per trip can be increased up to seven or eight. In this case the expense, including wages of the men and wear, tear, and care of mules, will work out roughly at from 7 to 10 cents per ton mile. Manifestly, if the ore-supply to a particular roadway is insufficient to keep a mule busy, the economy soon runs off.

MECHANICAL HAULAGE.—Mechanical haulage is seldom applicable to metal mines, for most metal deposits dip at considerable angles, and therefore, unlike most coal-mines, the horizon of haulage must frequently change, and there are no main arteries along which haulage continues through the life of the mine. Any mechanical system entails a good deal of expense for installation, and the useful life of any particular roadway, as above said, is very short. Moreover, the crooked roadways of most metal mines present difficulties of negotiation not to be overlooked. In order to use such systems it is necessary to condense the haulage to as few roadways as possible. Where the tonnage on one level is not sufficient to warrant other than men or animals, it sometimes pays (if the dip is steep enough) to dump everything through winzes from one to two levels to a main road below where mechanical equipment can be advantageously provided. The cost of shaft-winding the extra depth is inconsiderable compared to other factors, for the extra vertical distance of haulage can be done at a cost of one or two cents per ton mile. Moreover, from such an arrangement follows the concentration of shaft-bins, and of shaft labor, and winding is accomplished without so much shifting as to horizon, all of which economies equalize the extra distance of the lift.

There are three principal methods of mechanical transport in use:—

1. Cable-ways. 2. Compressed-air locomotives. 3. Electrical haulage.

Cable-ways or endless ropes are expensive to install, and to work to the best advantage require double tracks and fairly straight roads. While they are economical in operation and work with little danger to operatives, the limitations mentioned preclude them from adoption in metal mines, except in very special circumstances such as main crosscuts or adit tunnels, where the haulage is straight and concentrated from many sources of supply.

Compressed-air locomotives are somewhat heavy and cumbersome, and therefore require well-built tracks with heavy rails, but they have very great advantages for metal mine work. They need but a single track and are of low initial cost where compressed air is already a requirement of the mine. No subsidiary line equipment is needed, and thus they are free to traverse any road in the mine and can be readily shifted from one level to another. Their mechanical efficiency is not so low in the long run as might appear from the low efficiency of pneumatic machines generally, for by storage of compressed air at the charging station a more even rate of energy consumption is possible than in the constant cable and electrical power supply which must be equal to the maximum demand, while the air-plant consumes but the average demand.

Electrical haulage has the advantage of a much more compact locomotive and the drawback of more expensive track equipment, due to the necessity of transmission wire, etc. It has the further disadvantages of uselessness outside the equipped haulage way and of the dangers of the live wire in low and often wet tunnels.

In general, compressed-air locomotives possess many attractions for metal mine work, where air is in use in any event and where any mechanical system is at all justified. Any of the mechanical systems where tonnage is sufficient in quantity to justify their employment will handle material for from 1.5 to 4 cents per ton mile.

TRACKS.—Tracks for hand, mule, or rope haulage are usually built with from 12- to 16-pound rails, but when compressed-air or electrical locomotives are to be used, less than 24-pound rails are impossible. As to tracks in general, it may be said that careful laying out with even grades and gentle curves repays itself many times over in their subsequent operation. Further care in repair and lubrication of cars will often make a difference of 75% in the track resistance.

TRANSPORT IN STOPES.—Owing to the even shorter life of individual stopes than levels, the actual transport of ore or waste in them is often a function of the aboriginal shovel plus gravity. As shoveling is the most costly system of transport known, any means of stoping that decreases the need for it has merit. Shrinkage-stoping eliminates it altogether. In the other methods, gravity helps in proportion to the steepness of the dip. When the underlie becomes too flat for the ore to "run," transport can sometimes be helped by pitching the ore-passes at a steeper angle than the dip (Fig. 36). In some cases of flat deposits, crosscuts into the walls, or even levels under the ore-body, are justifiable. The more numerous the ore-passes, the less the lateral shoveling, but as passes cost money for construction and for repair, there is a nice economic balance in their frequency.

Mechanical haulage in stopes has been tried and finds a field under some conditions. In dips under 25 deg. and possessing fairly sound hanging-wall, where long-wall or flat-back cuts are employed, temporary tracks can often be laid in the stopes and the ore run in cars to the main passes. In such cases, the tracks are pushed up close to the face after each cut. Further self-acting inclines to lower cars to the levels can sometimes be installed to advantage. This arrangement also permits greater intervals between levels and less number of ore-passes. For dips between 25 deg. and 50 deg. where the mine is worked without stope support or with occasional pillars, a very useful contrivance is the sheet-iron trough—about eighteen inches wide and six inches deep—made in sections ten or twelve feet long and readily bolted together. In dips 35 deg. to 50 deg. this trough, laid on the foot-wall, gives a sufficiently smooth surface for the ore to run upon. When the dip is flat, the trough, if hung from plugs in the hanging-wall, may be swung backward and forward. The use of this "bumping-trough" saves much shoveling. For handling filling or ore in flat runs it deserves wider adoption. It is, of course, inapplicable in passes as a "bumping-trough," but can be fixed to give smooth surface. In flat mines it permits a wider interval between levels and therefore saves development work. The life of this contrivance is short when used in open stopes, owing to the dangers of bombardment from blasting.

In dips steeper than 50 deg. much of the shoveling into passes can be saved by rill-stoping, as described on page 100. Where flat-backed stopes are used in wide ore-bodies with filling, temporary tracks laid on the filling to the ore-passes are useful, for they permit wider intervals between passes.

In that underground engineer's paradise, the Witwatersrand, where the stopes require neither timber nor filling, the long, moderately pitched openings lend themselves particularly to the swinging iron troughs, and even endless wire ropes have been found advantageous in certain cases.

Where the roof is heavy and close support is required, and where the deposits are very irregular in shape and dip, there is little hope of mechanical assistance in stope transport.



CHAPTER XIII.

Mechanical Equipment. (Continued).

DRAINAGE: CONTROLLING FACTORS; VOLUME AND HEAD OF WATER; FLEXIBILITY; RELIABILITY; POWER CONDITIONS; MECHANICAL EFFICIENCY; CAPITAL OUTLAY. SYSTEMS OF DRAINAGE,—STEAM PUMPS, COMPRESSED-AIR PUMPS, ELECTRICAL PUMPS, ROD-DRIVEN PUMPS, BAILING; COMPARATIVE VALUE OF VARIOUS SYSTEMS.

With the exception of drainage tunnels—more fully described in Chapter VIII—all drainage must be mechanical. As the bulk of mine water usually lies near the surface, saving in pumping can sometimes be effected by leaving a complete pillar of ore under some of the upper levels. In many deposits, however, the ore has too many channels to render this of much avail.

There are six factors which enter into a determination of mechanical drainage systems for metal mines:—

1. Volume and head of water. 2. Flexibility to fluctuation in volume and head. 3. Reliability. 4. Capital cost. 5. The general power conditions. 6. Mechanical efficiency.

In the drainage appliances, more than in any other feature of the equipment, must mechanical efficiency be subordinated to the other issues.

FLEXIBILITY.—Flexibility in plant is necessary because volume and head of water are fluctuating factors. In wet regions the volume of water usually increases for a certain distance with the extension of openings in depth. In dry climates it generally decreases with the downward extension of the workings after a certain depth. Moreover, as depth progresses, the water follows the openings more or less and must be pumped against an ever greater head. In most cases the volume varies with the seasons. What increase will occur, from what horizon it must be lifted, and what the fluctuations in volume are likely to be, are all unknown at the time of installation. If a pumping system were to be laid out for a new mine, which would peradventure meet every possible contingency, the capital outlay would be enormous, and the operating efficiency would be very low during the long period in which it would be working below its capacity. The question of flexibility does not arise so prominently in coal-mines, for the more or less flat deposits give a fixed factor of depth. The flow is also more steady, and the volume can be in a measure approximated from general experience.

RELIABILITY.—The factor of reliability was at one time of more importance than in these days of high-class manufacture of many different pumping systems. Practically speaking, the only insurance from flooding in any event lies in the provision of a relief system of some sort,—duplicate pumps, or the simplest and most usual thing, bailing tanks. Only Cornish and compressed-air pumps will work with any security when drowned, and electrical pumps are easily ruined.

GENERAL POWER CONDITIONS.—The question of pumping installation is much dependent upon the power installation and other power requirements of the mine. For instance, where electrical power is purchased or generated by water-power, then electrical pumps have every advantage. Or where a large number of subsidiary motors can be economically driven from one central steam- or gas-driven electrical generation plant, they again have a strong call,—especially if the amount of water to be handled is moderate. Where the water is of limited volume and compressed-air plant a necessity for the mine, then air-driven pumps may be the most advantageous, etc.

MECHANICAL EFFICIENCY.—The mechanical efficiency of drainage machinery is very largely a question of method of power application. The actual pump can be built to almost the same efficiency for any power application, and with the exception of the limited field of bailing with tanks, mechanical drainage is a matter of pumps. All pumps must be set below their load, barring a few possible feet of suction lift, and they are therefore perforce underground, and in consequence all power must be transmitted from the surface. Transmission itself means loss of power varying from 10 to 60%, depending upon the medium used. It is therefore the choice of transmission medium that largely governs the mechanical efficiency.

SYSTEMS OF DRAINAGE.—The ideal pumping system for metal mines would be one which could be built in units and could be expanded or contracted unit by unit with the fluctuation in volume; which could also be easily moved to meet the differences of lifts; and in which each independent unit could be of the highest mechanical efficiency and would require but little space for erection. Such an ideal is unobtainable among any of the appliances with which the writer is familiar.

The wide variations in the origin of power, in the form of transmission, and in the method of final application, and the many combinations of these factors, meet the demands for flexibility, efficiency, capital cost, and reliability in various degrees depending upon the environment of the mine. Power nowadays is generated primarily with steam, water, and gas. These origins admit the transmission of power to the pumps by direct steam, compressed air, electricity, rods, or hydraulic columns.

DIRECT STEAM-PUMPS.—Direct steam has the disadvantage of radiated heat in the workings, of loss by the radiation, and, worse still, of the impracticability of placing and operating a highly efficient steam-engine underground. It is all but impossible to derive benefit from the vacuum, as any form of surface condenser here is impossible, and there can be no return of the hot soft water to the boilers.

Steam-pumps fall into two classes, rotary and direct-acting; the former have the great advantage of permitting the use of steam expansively and affording some field for effective use of condensation, but they are more costly, require much room, and are not fool-proof. The direct-acting pumps have all the advantage of compactness and the disadvantage of being the most inefficient of pumping machines used in mining. Taking the steam consumption of a good surface steam plant at 15 pounds per horse-power hour, the efficiency of rotary pumps with well-insulated pipes is probably not over 50%, and of direct-acting pumps from 40% down to 10%.

The advantage of all steam-pumps lies in the low capital outlay,—hence their convenient application to experimental mining and temporary pumping requirements. For final equipment they afford a great deal of flexibility, for if properly constructed they can be, with slight alteration, moved from one horizon to another without loss of relative efficiency. Thus the system can be rearranged for an increased volume of water, by decreasing the lift and increasing the number of pumps from different horizons.

COMPRESSED-AIR PUMPS.—Compressed-air transmission has an application similar to direct steam, but it is of still lower mechanical efficiency, because of the great loss in compression. It has the superiority of not heating the workings, and there is no difficulty as to the disposal of the exhaust, as with steam. Moreover, such pumps will work when drowned. Compressed air has a distinct place for minor pumping units, especially those removed from the shaft, for they can be run as an adjunct to the air-drill system of the mine, and by this arrangement much capital outlay may be saved. The cost of the extra power consumed by such an arrangement is less than the average cost of compressed-air power, because many of the compressor charges have to be paid anyway. When compressed air is water-generated, they have a field for permanent installations. The efficiency of even rotary air-driven pumps, based on power delivered into a good compressor, is probably not over 25%.

ELECTRICAL PUMPS.—Electrical pumps have somewhat less flexibility than steam- or air-driven apparatus, in that the speed of the pumps can be varied only within small limits. They have the same great advantage in the easy reorganization of the system to altered conditions of water-flow. Electricity, when steam-generated, has the handicap of the losses of two conversions, the actual pump efficiency being about 60% in well-constructed plants; the efficiency is therefore greater than direct steam or compressed air. Where the mine is operated with water-power, purchased electric current, or where there is an installation of electrical generating plant by steam or gas for other purposes, electrically driven pumps take precedence over all others on account of their combined moderate capital outlay, great flexibility, and reasonable efficiency.

In late years, direct-coupled, electric-driven centrifugal pumps have entered the mining field, but their efficiency, despite makers' claims, is low. While they show comparatively good results on low lifts the slip increases with the lift. In heads over 200 feet their efficiency is probably not 30% of the power delivered to the electrical generator. Their chief attractions are small capital cost and the compact size which admits of easy installation.

ROD-DRIVEN PUMPS.—Pumps of the Cornish type in vertical shafts, if operated to full load and if driven by modern engines, have an efficiency much higher than any other sort of installation, and records of 85 to 90% are not unusual. The highest efficiency in these pumps yet obtained has been by driving the pump with rope transmission from a high-speed triple expansion engine, and in this plant an actual consumption of only 17 pounds of steam per horse-power hour for actual water lifted has been accomplished.

To provide, however, for increase of flow and change of horizon, rod-driven pumps must be so overpowered at the earlier stage of the mine that they operate with great loss. Of all pumping systems they are the most expensive to provide. They have no place in crooked openings and only work in inclines with many disadvantages.

In general their lack of flexibility is fast putting them out of the metal miner's purview. Where the pumping depth and volume of water are approximately known, as is often the case in coal mines, this, the father of all pumps, still holds its own.

HYDRAULIC PUMPS.—Hydraulic pumps, in which a column of water is used as the transmission fluid from a surface pump to a corresponding pump underground has had some adoption in coal mines, but little in metal mines. They have a certain amount of flexibility but low efficiency, and are not likely to have much field against electrical pumps.

BAILING.—Bailing deserves to be mentioned among drainage methods, for under certain conditions it is a most useful system, and at all times a mine should be equipped with tanks against accident to the pumps. Where the amount of water is limited,—up to, say, 50,000 gallons daily,—and where the ore output of the mine permits the use of the winding-engine for part of the time on water haulage, there is in the method an almost total saving of capital outlay. Inasmuch as the winding-engine, even when the ore haulage is finished for the day, must be under steam for handling men in emergencies, and as the labor of stokers, engine-drivers, shaft-men, etc., is therefore necessary, the cost of power consumed by bailing is not great, despite the low efficiency of winding-engines.

COMPARISON OF VARIOUS SYSTEMS.—If it is assumed that flexibility, reliability, mechanical efficiency, and capital cost can each be divided into four figures of relative importance,—A, B, C, and D, with A representing the most desirable result,—it is possible to indicate roughly the comparative values of various pumping systems. It is not pretended that the four degrees are of equal import. In all cases the factor of general power conditions on the mine may alter the relative positions.

==================================================================== Direct Compressed Steam- Steam Air Electricity Driven Hydraulic Bailing Pumps Rods Columns Tanks - - - - Flexibility. A A B D B A Reliability. B B B A D A Mechanical Efficiency. C D B A C D Capital Cost A B B D D ====================================================================

As each mine has its special environment, it is impossible to formulate any final conclusion on a subject so involved. The attempt would lead to a discussion of a thousand supposititious cases and hypothetical remedies. Further, the description alone of pumping machines would fill volumes, and the subject will never be exhausted. The engineer confronted with pumping problems must marshal all the alternatives, count his money, and apply the tests of flexibility, reliability, efficiency, and cost, choose the system of least disadvantages, and finally deprecate the whole affair, for it is but a parasite growth on the mine.



CHAPTER XIV.

Mechanical Equipment (Concluded).

MACHINE DRILLING: POWER TRANSMISSION; COMPRESSED AIR VS. ELECTRICITY; AIR DRILLS; MACHINE VS. HAND DRILLING. WORK-SHOPS. IMPROVEMENT IN EQUIPMENT.

For over two hundred years from the introduction of drill-holes for blasting by Caspar Weindel in Hungary, to the invention of the first practicable steam percussion drill by J. J. Crouch of Philadelphia, in 1849, all drilling was done by hand. Since Crouch's time a host of mechanical drills to be actuated by all sorts of power have come forward, and even yet the machine-drill has not reached a stage of development where it can displace hand-work under all conditions. Steam-power was never adapted to underground work, and a serviceable drill for this purpose was not found until compressed air for transmission was demonstrated by Dommeiller on the Mt. Cenis tunnel in 1861.

The ideal requirements for a drill combine:—

a. Power transmission adapted to underground conditions. b. Lightness. c. Simplicity of construction. d. Strength. e. Rapidity and strength of blow. f. Ease of erection. g. Reliability. h. Mechanical efficiency. i. Low capital cost.

No drill invented yet fills all these requirements, and all are a compromise on some point.

POWER TRANSMISSION; COMPRESSED AIR vs. ELECTRICITY.—The only transmissions adapted to underground drill-work are compressed air and electricity, and as yet an electric-driven drill has not been produced which meets as many of the requirements of the metal miner as do compressed-air drills. The latter, up to date, have superiority in simplicity, lightness, ease of erection, reliability, and strength over electric machines. Air has another advantage in that it affords some assistance to ventilation, but it has the disadvantage of remarkably low mechanical efficiency. The actual work performed by the standard 3-3/4-inch air-drill probably does not amount to over two or three horse-power against from fifteen to eighteen horse-power delivered into the compressor, or mechanical efficiency of less than 25%. As electrical power can be delivered to the drill with much less loss than compressed air, the field for a more economical drill on this line is wide enough to create eventually the proper tool to apply it. The most satisfactory electric drill produced has been the Temple drill, which is really an air-drill driven by a small electrically-driven compressor placed near the drill itself. But even this has considerable deficiencies in mining work; the difficulties of setting up, especially for stoping work, and the more cumbersome apparatus to remove before blasting are serious drawbacks. It has deficiencies in reliability and greater complication of machinery than direct air.

AIR-COMPRESSION.—The method of air-compression so long accomplished only by power-driven pistons has now an alternative in some situations by the use of falling water. This latter system is a development of the last twelve years, and, due to the low initial outlay and extremely low operating costs, bids fair in those regions where water head is available not only to displace the machine compressor, but also to extend the application of compressed air to mine motors generally, and to stay in some environments the encroachment of electricity into the compressed-air field. Installations of this sort in the West Kootenay, B.C., and at the Victoria copper mine, Michigan, are giving results worthy of careful attention.

Mechanical air-compressors are steam-, water-, electrical-, and gas-driven, the alternative obviously depending on the source and cost of power. Electrical- and gas- and water-driven compressors work under the disadvantage of constant speed motors and respond little to the variation in load, a partial remedy for which lies in enlarged air-storage capacity. Inasmuch as compressed air, so far as our knowledge goes at present, must be provided for drills, it forms a convenient transmission of power to various motors underground, such as small pumps, winches, or locomotives. As stated in discussing those machines, it is not primarily a transmission of even moderate mechanical efficiency for such purposes; but as against the installation and operation of independent transmission, such as steam or electricity, the economic advantage often compensates the technical losses. Where such motors are fixed, as in pumps and winches, a considerable gain in efficiency can be obtained by reheating.

It is not proposed to enter a discussion of mechanical details of air-compression, more than to call attention to the most common delinquency in the installation of such plants. This deficiency lies in insufficient compression capacity for the needs of the mine and consequent effective operation of drills, for with under 75 pounds pressure the drills decrease remarkably in rapidity of stroke and force of the blow. The consequent decrease in actual accomplishment is far beyond the ratio that might be expected on the basis of mere difference of pressure. Another form of the same chronic ill lies in insufficient air-storage capacity to provide for maintenance of pressure against moments when all drills or motors in the mine synchronize in heavy demand for air, and thus lower the pressure at certain periods.

AIR-DRILLS.—Air-drills are from a mechanical point of view broadly of two types,—the first, in which the drill is the piston extension; and the second, a more recent development for mining work, in which the piston acts as a hammer striking the head of the drill. From an economic point of view drills may be divided into three classes. First, heavy drills, weighing from 150 to 400 pounds, which require two men for their operation; second, "baby" drills of the piston type, weighing from 110 to 150 pounds, requiring one man with occasional assistance in setting up; and third, very light drills almost wholly of the hammer type. This type is built in two forms: a heavier type for mounting on columns, weighing about 80 pounds; and a type after the order of the pneumatic riveter, weighing as low as 20 pounds and worked without mounting.

The weight and consequent mobility of a drill, aside from labor questions, have a marked effect on costs, for the lighter the drill the less difficulty and delay in erection, and consequent less loss of time and less tendency to drill holes from one radius, regardless of pointing to take best advantage of breaking planes. Moreover, smaller diameter and shorter holes consume less explosives per foot advanced or per ton broken. The best results in tonnage broken and explosive consumed, if measured by the foot of drill-hole necessary, can be accomplished from hand-drilling and the lighter the machine drill, assuming equal reliability, the nearer it approximates these advantages.

The blow, and therefore size and depth of hole and rapidity of drilling, are somewhat dependent upon the size of cylinders and length of stroke, and therefore the heavier types are better adapted to hard ground and to the deep holes of some development points. Their advantages over the other classes lie chiefly in this ability to bore exceedingly hard material and in the greater speed of advance possible in development work; but except for these two special purposes they are not as economical per foot advanced or per ton of ore broken as the lighter drills.

The second class, where men can be induced to work them one man per drill, saves in labor and gains in mobility. Many tests show great economy of the "baby" type of piston drills in average ground over the heavier machines for stoping and for most lateral development. All piston types are somewhat cumbersome and the heavier types require at least four feet of head room. The "baby" type can be operated in less space than this, but for narrow stopes they do not lend themselves with the same facility as the third class.

The third class of drills is still in process of development, but it bids fair to displace much of the occupation of the piston types of drill. Aside from being a one-man drill, by its mobility it will apparently largely reproduce the advantage of hand-drilling in ability to place short holes from the most advantageous angles and for use in narrow places. As compared with other drills it bids fair to require less time for setting up and removal and for change of bits; to destroy less steel by breakages; to dull the bits less rapidly per foot of hole; to be more economical of power; to require much less skill in operation, for judgment is less called upon in delivering speed; and to evade difficulties of fissured ground, etc. And finally the cost is only one-half, initially and for spares. Its disadvantage so far is a lack of reliability due to lightness of construction, but this is very rapidly being overcome. This type, however, is limited in depth of hole possible, for, from lack of positive reverse movement, there is a tendency for the spoil to pack around the bit, and as a result about four feet seems the limit.

The performance of a machine-drill under show conditions may be anything up to ten or twelve feet of hole per hour on rock such as compact granite; but in underground work a large proportion of the time is lost in picking down loose ore, setting up machines, removal for blasting, clearing away spoil, making adjustments, etc. The amount of lost time is often dependent upon the width of stope or shaft and the method of stoping. Situations which require long drill columns or special scaffolds greatly accentuate the loss of time. Further, the difficulties in setting up reflect indirectly on efficiency to a greater extent in that a larger proportion of holes are drilled from one radius and thus less adapted to the best breaking results than where the drill can easily be reset from various angles.

The usual duty of a heavy drill per eight-hour shift using two men is from 20 to 40 feet of hole, depending upon the rock, facilities for setting up, etc., etc.[*] The lighter drills have a less average duty, averaging from 15 to 25 feet per shift.

[Footnote *: Over the year 1907 in twenty-eight mines compiled from Alaska to Australia, an average of 23.5 feet was drilled per eight-hour shift by machines larger than three-inch cylinder.]

MACHINE vs. HAND-DRILLING.—The advantages of hand-drilling over machine-drilling lie, first, in the total saving of power, the absence of capital cost, repairs, depreciation, etc., on power, compresser and drill plant; second, the time required for setting up machine-drills does not warrant frequent blasts, so that a number of holes on one radius are a necessity, and therefore machine-holes generally cannot be pointed to such advantage as hand-holes. Hand-holes can be set to any angle, and by thus frequent blasting yield greater tonnage per foot of hole. Third, a large number of comparative statistics from American, South African, and Australian mines show a saving of about 25% in explosives for the same tonnage or foot of advance by hand-holes over medium and heavy drill-holes.

The duty of a skilled white man, single-handed, in rock such as is usually met below the zone of oxidation, is from 5 to 7 feet per shift, depending on the rock and the man. Two men hand-drilling will therefore do from 1/4 to 2/3 of the same footage of holes that can be done by two men with a heavy machine-drill, and two men hand-drilling will do from 1/5 to 1/2 the footage of two men with two light drills.

The saving in labor of from 75 to 33% by machine-drilling may or may not be made up by the other costs involved in machine-work. The comparative value of machine- and hand-drilling is not subject to sweeping generalization. A large amount of data from various parts of the world, with skilled white men, shows machine-work to cost from half as much per ton or foot advanced as hand-work to 25% more than handwork, depending on the situation, type of drill, etc. In a general way hand-work can more nearly compete with heavy machines than light ones. The situations where hand-work can compete with even light machines are in very narrow stopes where drills cannot be pointed to advantage, and where the increased working space necessary for machine drills results in breaking more waste. Further, hand-drilling can often compete with machine-work in wide stopes where long columns or platforms must be used and therefore there is much delay in taking down, reerection, etc.

Many other factors enter into a comparison, however, for machine-drilling produces a greater number of deeper holes and permits larger blasts and therefore more rapid progress. In driving levels under average conditions monthly footage is from two to three times as great with heavy machines as by hand-drilling, and by lighter machines a somewhat less proportion of greater speed. The greater speed obtained in development work, the greater tonnage obtained per man in stoping, with consequent reduction in the number of men employed, and in reduction of superintendence and general charges are indirect advantages for machine-drilling not to be overlooked.

The results obtained in South Africa by hand-drilling in shafts, and its very general adoption there, seem to indicate that better speed and more economical work can be obtained in that way in very large shafts than by machine-drilling. How far special reasons there apply to smaller shafts or labor conditions elsewhere have yet to be demonstrated. In large-dimension shafts demanding a large number of machines, the handling of long machine bars and machines generally results in a great loss of time. The large charges in deep holes break the walls very irregularly; misfires cause more delay; timbering is more difficult in the face of heavy blasting charges; and the larger amount of spoil broken at one time delays renewed drilling, and altogether the advantages seem to lie with hand-drilling in shafts of large horizontal section.

The rapid development of special drills for particular conditions has eliminated the advantage of hand-work in many situations during the past ten years, and the invention of the hammer type of drill bids fair to render hand-drilling a thing of the past. One generalization is possible, and that is, if drills are run on 40-50 pounds' pressure they are no economy over hand-drilling.

WORKSHOPS.

In addition to the ordinary blacksmithy, which is a necessity, the modern tendency has been to elaborate the shops on mines to cover machine-work, pattern-making and foundry-work, in order that delays may be minimized by quick repairs. To provide, however, for such contingencies a staff of men must be kept larger than the demand of average requirements. The result is an effort to provide jobs or to do work extravagantly or unnecessarily well. In general, it is an easy spot for fungi to start growing on the administration, and if custom repair shops are available at all, mine shops can be easily overdone.

A number of machines are now in use for sharpening drills. Machine-sharpening is much cheaper than hand-work, although the drills thus sharpened are rather less efficient owing to the difficulty of tempering them to the same nicety; however, the net results are in favor of the machines.

IMPROVEMENT IN EQUIPMENT.

Not only is every mine a progressive industry until the bottom gives out, but the technology of the industry is always progressing, so that the manager is almost daily confronted with improvements which could be made in his equipment that would result in decreasing expenses or increasing metal recovery. There is one test to the advisability of such alterations: How long will it take to recover the capital outlay from the savings effected? and over and above this recovery of capital there must be some very considerable gain. The life of mines is at least secured over the period exposed in the ore-reserves, and if the proposed alteration will show its recovery and profit in that period, then it is certainly justified. If it takes longer than this on the average speculative ore-deposit, it is a gamble on finding further ore. As a matter of practical policy it will be found that an improvement in equipment which requires more than three or four years to redeem itself out of saving, is usually a mechanical or metallurgical refinement the indulgence in which is very doubtful.



CHAPTER XV.

Ratio of Output to the Mine.

DETERMINATION OF THE POSSIBLE MAXIMUM; LIMITING FACTORS; COST OF EQUIPMENT; LIFE OF THE MINE; MECHANICAL INEFFICIENCY OF PATCHWORK PLANT; OVERPRODUCTION OF BASE METAL; SECURITY OF INVESTMENT.

The output obtainable from a given mine is obviously dependent not only on the size of the deposit, but also on the equipment provided,—in which equipment means the whole working appliances, surface and underground.

A rough and ready idea of output possibilities of inclined deposits can be secured by calculating the tonnage available per foot of depth from the horizontal cross-section of the ore-bodies exposed and assuming an annual depth of exhaustion, or in horizontal deposits from an assumption of a given area of exhaustion. Few mines, at the time of initial equipment, are developed to an extent from which their possibilities in production are evident, for wise finance usually leads to the erection of some equipment and production before development has been advanced to a point that warrants a large or final installation. Moreover, even were the full possibilities of the mine known, the limitations of finance usually necessitate a less plant to start with than is finally contemplated. Therefore output and equipment are usually growing possibilities during the early life of a mine.

There is no better instance in mine engineering where pure theory must give way to practical necessities of finance than in the determination of the size of equipment and therefore output. Moreover, where finance even is no obstruction, there are other limitations of a very practical order which must dominate the question of the size of plant giving the greatest technical economy. It is, however, useful to state the theoretical considerations in determining the ultimate volume of output and therefore the size of equipments, for the theory will serve to illuminate the practical limitations. The discussion will also again demonstrate that all engineering is a series of compromises with natural and economic forces.