Rock Hammer Drills

Rock Hammer Drills

The hammer drill rightly receives the credit for having made the one-man drill possible, and so many economies seem possible through the proper application of different types of hammer drills to various mining, quarrying, and excavating operations, that an indication of the economies effected by the New Jersey Zinc Co. at its Franklin mines may be of pertinent interest. When this company commenced its trials of hammer drills in 1907, these tools had not been developed to one-fourth the capacity and refinement which they have at present. At that time it was frequently stated that such a small tool, drilling holes of less diameter than the reciprocating rock drill, could not drill enough holes in a shift to permit the placing of sufficient explosive to break a tonnage of ore comparable with that produced by the ’“rock drills;” that the placing of small holes inclined upward, at angles steeper than 40° above the horizontal, could not be expected to produce results equal to the large flat, wet or dry, holes in the breasted back of an overhand stope, and would only shatter the ground so as to make the back unsafe. In spite of these adverse opinions, the hammer drills first showed their superiority over both heavy and light reciprocating drills in raising and in stoping, and then in drifting and quarry work. As a result, all of the reciprocating drills at the Franklin mines were scrapped three years ago, all of the mining work being accomplished with increased efficiency, as shown in detail in this article.

With the advent of the hammer drill in this property, it was considered advisable to make comparative tests of all the tools accessible, and it has since been the policy to investigate the merits of any advance of the drilling art in order to get the maximum amount of work from the tools. The necessity of devising some means of standardizing drill tests, and of measuring the consumption of compressed air as well as the drilling speed, was early realized.

The common test was to fill a measured air receiver with compressed air at a certain gauge pressure, run the drill until the pressure had dropped to too low a figure, then compute from the time, drop in pressure, and capacity of the receiver, the cubic feet of free air used. This was not considered a fair indication of the drilling capacity of a machine, since the performance of some drills did not vary directly with the absolute pressures of the compressed air.

It was, therefore, found expedient to build a water-displacement air meter with which the drill test could be carried on for any length of time without serious variations in the desired air pressure. This apparatus, as shown in Fig. 1, consists of two tanks, half-filled with water, and made, of 12-in. pipe with blank flanges, gauge glasses being mounted on one; a four-way cock connecting the compressed-air supply pipe with both tanks and the air line going to the drill. This device gives more accurate results than the common types of water or gas meters;and since any errors are due to the human element of reading the gauge glasses and reversing the four-way cock, they tend to be compensating throughout a number of tests.

The procedure is as follows: Air is drawn by the drill from the receiver C, which tends to trap any moisture carried over by the air from tanks A and B and assures a constant pressure while the four-way cock is reversed. In the arrangement shown in Fig. 1, the receiver draws its supply of air from the tank B and the water rises in this tank by virtue of the pressure of the air admitted from the air main, through the four-way cock to the top of tank A, where the water is being forced downward and through the 2-in. connecting pipe to tank B. When the water has risen to a certain point near the top of the gauge glass in tank B, the four-way cock is reversed and the inlet air is supplied to the top of tank B; the drilling air is then taken from the top of A, the reversal of the cock again being made when the water in tank B has fallen to a point near the.bottom of the gauge glass.

A pet cock is placed on the top of tank C so as to permit the bleeding of air to bring the water columns to the desired point for starting a run, and another pet cock is attached to the bottom of the same tank in order to permit the drainage of water. For convenience in measuring and computing, a run is made on the supply of compressed air indicated by a certain number of reciprocations of the water columns between fixed points on the gauge glasses; the pressures are measured on the air gauge, mounted on tank C, the length of time is taken by a stop watch, and the consumption of free air, per minute is computed. Unless a pressure regulator is installed between the four-way cock and the receiver C, or else a globe valve at this point is operated manually to throttle the air so as to maintain a constant pressure, it is evident that the air pressure at the drill will vary in accordance with the water column supported by the inlet air pressure, but since the gauge-glass marks are in this instance set 34¾ in. apart, the maximum variation in pressures is about 1¼ lb. per square inch; it is difficult to find a pressure regulator which will control a pressure of 90 or 100 lb. per square inch to a closer



degree of accuracy, and the sensitiveness of drills to air pressures and the accuracy of time and distance measurements rarely exceed this error.

It is obvious that in comparative drill tests the following factors must be considered: The nature of the rock drilled; the gauge of the drill bits and their form and condition; the maintenance of equal compressed-air pressures; similar inclination and approximate depths of drill holes; equal vigor in the rotation of hand-rotated tools; and proper fit of the drill shanks in the chuck bushings, as well as their construction so that the blows are delivered on a plane surface of proper size at right angles to the axis of the drill steel and at the center of its shank end. In the tests summarized in Table I, 1¾ in. has been taken as the standard diametral gauge of the bit, since it is a dimension which averages the gauges of drill steels used in reciprocating rock drills and is fair in determining the performances of such tools; it also represents almost the largest gauge necessary in hammer-drill stopers or block-holers, so that equal or even better performances may be expected from them as a hole is deepened with the smaller, gauges in a set of drill steel. At Franklin, the testing rock is a compact coarsely crystalline white limestone, which greatly resembles a marble, and this rock proves a fair average of the various qualities of ore met in the mining operations. Although it is not hard, for a well-tempered drill bit can drill 3 or 4 ft. of hole before its cutting edges are materially dulled, and although it seems to chip freely, yet it possesses a compactness and toughness which is likely to prove surprising to one who has not previously tested a drill in it. Tests with various machines in Franklin, and elsewhere, indicate that this white limestone does not cut quite so fast as a sharp drill can achieve in Cripple Creek granite; is about on a par with Barre granite; and cuts slightly faster than Quincy granite. The chief difference is that a good drill will cut this limestone as fast in the second or third minute’ of its run, while it would have been dulled by the first minute’s run in the granites, causing its cutting speed to fall off materially in the second and third minutes. Raised-center cross bits are the standard type used with solid steel in these test’s, and flat-faced six-point bits’ are generally used with the hollow steels, in general the cutting speed of these bits being about the same if the rotation of the drill steel is free. Fig. 2 shows the forms of the drill bits used at the Franklin mines.

The results in Table I are not complete to date, but indicate the improvements in hammer drills, from the operators standpoint of efficiency, during four years of advancement in the art; and it may be interesting to note that, so far as we know, no drills of the present day surpass in drilling speed and low air consumption the best drills listed in this table, although several makes of hammer drills are on a par with them. In order to avoid invidious comparisons between the different makes of drills, symbols have been used to designate each certain make and design of drill. The following abbreviations have been used in this table:


Auto aux V = automatic auxiliary valveless control of rotations.
Auto rifle = automatic rotation caused by a piston reciprocating as though it were controlled by a rifle bar.
Dir. air = direct-air feed, or one in which the feed cylinder is rigidly attached to the hammer cylinder and in which the feed piston or plunger extends from the rear end of the machine by virtue of the air pressure applied to it.
Rev. air = reversed-air feed, or one in which the feed piston is rigidly attached to the hammer cylinder and the feed cylinder is free, to extend backward, so readily adapts itself to the customary column mounting of stoping drills.

Some tests were included in this table for the consideration of points to be made later. Before studying the improvements in hammer-drill efficiency it seems wise to explain the reasons for offering the figures in the last column of figures as representing a factor of “drill desirability.”

In determining the relative merits of rock drills, whether of the reciprocating or hammer type, the logical basis is one of cost. Therefore, the drill which bores a foot of drill hole of standard cross-section at the
lowest cost rate- for drilling labor, power, and maintenance (including amortization), would have the highest “factor of desirability;” and a formula to express this may be developed as follows:


F be the “factor of desirability,”
D be the cost of drilling labor per foot of hole,
P be the cost of power per foot of hole,
M be the cost of maintenance per foot of hole.


F = 1/D + P + M


t = Period of time for drilling-speed test, in minutes.
d = Depth of hole drilled in time, t, in inches.
S = d/t = Drilling speed during actual running of machine, in inches per minute.
L = Hourly wage of drilling labor, in cents.
O = Percentage of time spent in drilling to total operating time including the changing of drill steels and shifting to new positions and starting of new holes.


D = L/60SO/12 = L/5dO/t = 0.2 tL/dO


p = Power cost to produce 100 cu. ft. of free air, compressed to standard drill-testing pressure, in cents.
v = Number of cubic feet of free air used in test by operating drill.
d = Depth of standard hole drilled, in. inches.


P = 12pv/100d = 0.12 pv/d

Substituting these values of D and P in the original equation we obtain

F = 1/0.20 tL/dO + 0.12 pv/d + M = d/0.2L/O t + 0.12 pv + dM

Since L is a constant for any particular mine, and 0 for a given number of steel changes, with any particular type of drill—such as a column-mounted reciprocating drill, a column-mounted hammer drill, an air-feed hammer drill, a block-holing drill, etc.—we may simplify the equation by substituting

k = coefficient of drilling = 0.20L/O

Also, since p is a constant for any particular mine we may further simplify by placing

k’ = coefficient of power = 0.12p

and we then have the general equation for any particular mining conditions and type of drill

F = d/kt + k’v + dM

However, the correct value for maintenance and amortization of any particular type and make of drill can be determined only after operations extending over months or years, so that this factor may well be left out of a formula which is to be used for classifying drills after drilling speed and power consumption tests, which may be completed in a short time. The consideration of the reduction of drilling speed and the increase of power consumption, which occur in a drill because of wear or any other normal results of service, may fairly be placed in the same class as maintenance. Judgment as to the materials, workmanship, and design of any drill, as well as reports of its satisfactory service elsewhere, will lead to a rough estimate of the final desirability of a drill if it has shown a high standard, on testing based on drilling speed and power consumption. The equation is thus simplified to the form

F = d/kt + k’v

But other highly important factors enter into the problem of selection of a drill, namely, the reduction in labor units, capital, and overhead , charges brought about by an increased drilling speed and increased tonnage per machine; the increased efficiency of supervision and work caused by the reduction and concentration of the number of working places; the possibility of producing a greater tonnage from any property with a limited number of working places; and the possibility of reducing the drilling equipment, with its attendant stock of spares, hoses, and connections, and extensive air mains, if a drill with a greater drilling speed may be employed. It therefore seems that the following formula is more indicative of the actual merits of drills, although theoretically it has no derivation, and must be considered empirical; it also possesses the virtue of reducing to a simple form. This formula for a “factor of desirability” has been used for the past six years at the Franklin Furnace mines of the New Jersey Zinc Co. All coefficients have been omitted since the following drill tests have all been under the same standard conditions.

F’ = 1/DPM

since M is treated separately, as has been previously suggested, the equation becomes

F’ = 1/DP

Now if the same values previously deduced for D and P are substituted,


where K is a new coefficient equal to the reciprocal of the product of k and k’.

Therefore, the “factor of desirability” equals the drilling speed, in inches per minute, divided by the power consumption, in cubic feet of free air, per inch drilled. It is quite evident that the factor gained from the quotient of inches drilled per minute divided by cubic feet of free air per minute (or the reciprocal of this quotient) gives merely the power consumption per inch of hole drilled and ignores the quantity of drilling which may be accomplished.

The application of both of these formulas for F and F’ to a hypothetical problem may be of interest to show the comparative; results within the limits of practice.

Let us assume that 30 h.p. is required to compress 100 cu. ft. of free air per minute to 100 lb. per square inch gauge pressure and deliver the same to a drill in the mine; that the power cost is 1c. per horsepower-hour; that a drill which shows a drilling speed of 10 in. per minute on test verages 20 ft. per hour under working conditions, and uses 60 cu. ft. of free air per minute on test; that another drill will show a drilling speed of 6 in. per minute on test with an air consumption of 36 cu. ft. per minute and will average 12 ft. per hour under working conditions, and that the wage scale for drill runners is 40c. per hour, then,


Thus the relative factors for the two drills by the first formula have a ratio of 0.424 to 0.271 or 1.56 to 1; and by the second formula (empirical) the ratio of factors is 1.389 to 0.833 or 1.67 to 1. In other words, by the empirical formula the fast drill is credited with about a 7 per cent, higher rating than by the theoretical formula, and this does not seem an undue allowance to cover the unestimated advantages previously enumerated.

Records made previous to July, 1909, have not been shown in Table I since much of the work done in 1907 and 1908 was distinctly experimental m determining the desirable cylinder diameters, lengths of strokes piston weights, valve weights, etc:, but such records show drilling speeds of about 2 to 3 in. per minute with air consumptions of from 40 to 70 cu. ft. of free air per minute at 90 lb. per square inch gauge pressure. The listed tests made during 1909 cover most of the well-known American makes of hammer drills at that time, and if one excepts the drills denoted by symbols G, G1, etc., since they were experimental tools, the design of which was developed by the New Jersey Zinc Co. at Franklin Furnace, N. J., it is noticeable that about 4½ and 5 in. were the highest drilling speeds obtainable at about 90 lb. pressure and with an air consumption of 60 to 90 cu. ft. of free air per minute; and for various drills the ‘‘factor” varied from 0.09 to 0.41. Those drills marked G, which were made exclusively for the New Jersey Zinc Co., increased the drilling speed about 40 per cent, above the best previous drill performances, and remained unequaled in drilling speed for a year and unsurpassed for about a year and a half. The fact that a number of these drills were included in the equipment at Franklin accounts for part of the increased stoping efficiency during the year 1911, as cited later. Although it was then the opinion of some unprejudiced persons, well versed in the drilling art, that such tools had reached their practical limit of drilling speed as well as the limit of strengths of materials, yet 18 months later a new type of drill was developed to achieve 20 per cent, more drilling with twice as good a factor, and a renewed equipment of these other drills again increased the mining efficiency. Again a period of 18 months sufficed for the production of a hammer drill which still further advanced the drilling speeds 20 per cent. and since the introduction of this drill we have been able to find several drills which surpassed it 10 to 20 per cent, in drilling speed.

In Table I some, seemingly freak runs are noticeable, which are included to call attention to the variability of results in presumably standard testing. For instance, under drills D it appears that a bit with two wings broken will drill faster and at a lower air consumption per minute than can be attained with a perfect bit; and again, with drill M1 a bit which has proved a little soft and battered drills one-fourth more per minute than bits in proper condition and with, the same air consumption. Furthermore, the tests of one person indicate that, when the size and form of the drill bits are the same, faster drilling can be done with short steels than long ones, while another investigator shows a greater drilling speed with long steel than with short. The use of tappets or anvil blocks between the shanks of drill steels and pistons is generally estimated as causing a reduction of 20 to 30 per cent, in the drilling ability, but some tests do not confirm this and show even an increased cutting speed with the use of anvil blocks in a machine otherwise the same. With some drills the use of water to clean the cuttings from the hole seems to cause a cutting speed below that obtainable through the use of compressed air for the same purpose, but in other instances the advantages are reversed. In short, there seem to be so many variables in the drilling problem as to warrant a 10 per cent, variation in the results of supposedly standard tests, and a number of runs should be made to gain a fair average; or strict judgment of machines should not be made within this limit.

Perhaps the consideration of the physical phenomena relating to the process of drilling may prove of interest and value. When rock is excavated by a drill bit three applications of forces seem to be involved- by abrasion, by crushing, and by severing or chipping. Although all of these must take place to a certain degree, the greatest amount of useful work is performed when the percentage of force applied to chip reaches a maximum. But in rock it appears that chips can be produced in radically different ways first, by the severing of molecules, and second by the reflex forces produced in an elastic medium. To illustrate this, consider the chipping of a comparatively in elastic substance such as lead. With a hammer and a chisel, whose axis is inclined considerably from the normal to the surface of a lead block, it is possible to sever the lead and roll up chips, but if the chisel is normal to the surface of a thick block only an indentation can be made and there probably will be a raised area about the indentation to accommodate a certain percentage of the displaced metal. On the other hand, with a highly elastic material, such as glass, the forces impressed by a normally positioned chisel will cause a compression of the molecules, whose elasticity will cause their expansion toward a free, unresisted surface. Since the greatest forces are developed at the surface, since the penetration of the chisel carries some forces to a depth below the surface; and since the chisel surface itself applies some forces at an angle to its axis and impedes the re-expansion of molecules to the space it occupies, therefore, the reflex forces produce more or less cone-shaped chips or flakes and leave a corresponding crater in the block of glass. Now, if the chisel is placed near the edge of a block of glass, the blow upon it will induce stresses to another free face and a correspondingly larger chip will be produced because of the tendency of the forces to seek relief in the shortest direction as well as because of the severing effect. The method of cutting of a drill bit is commonly shown as taking place in this last way with the progressive chipping of a series of benches or steps, but it is doubtful whether such a procedure exists, except in rare instances, for the speed and latitude of rotation between consecutive blows of the drill piston or hammer cannot be controlled with sufficient precision nor adjusted to the various rocks; and an inspection of the cuttings from a drill hole shows them to be flakes, or a crushed and abraded powder.

In the formation of these flaky chips there may be a limiting force of blow for each velocity of impression in order to gain the most useful work (i.e., in the production of flakes), for it appears that beyond certain limits the blows increase the percentage of crushed material and the drilling speed does not vary with the force applied, so that some heavy-hitting drills accomplish more in medium-soft ground when a portion of their blows are absorbed by a tappet at the shank end of the steel or by a cushion of water intervening between the bit and the rock. If the force of the blows was lessened by a reduction in air pressure the speed of the piston would be slowed up, and the drilling would suffer from the fewer number of blows per minute.

The transmission of the kinetic energy of the piston to the rock is also influenced by many factors. The blow may be delivered against the rock by the free drill steel which is driven forward through the intervening air or water by the impact of the piston and the velocity of the steel will depend upon the relative masses of the drill steel and piston, the velocity of the piston, and the coefficient of elasticity of the steel, in accordance with the well-known laws of mechanics which deal with elastic or partly elastic bodies and their impact. The drill steel in this way assumes the functions of a “jumper” drill which is driven against and rebounds from the rock at a high frequency, and its action is well seen in most all screw- feed hammer drills with the ringing or jingling of the steel in a drill hole. Another mode of force transmission is by compressional waves, traveling through the drill steel from the shank to the bit. This latter condition brings a cutting effect only when one end of the steel is tight against the rock, but then proves very efficient. Although the air-feed hammer drills usually chatter the steel against the rock, like a projectile shot from the chuck bushing by impacts of the piston, yet it seems possible to approximate the other working condition by designing the air feed so that the pressure is lowered as the piston is traveling on its back stroke (possibly by taking the supply air, for the back stroke, from the air feed) and so that the air-feed pressure builds up and forces the drill against the rock just before it is struck by the piston. The reversed-air feeds may sometimes approximate these conditions and then assist the machine to a higher drilling speed. If hammer drills were made so that the drill steels were always held firmly against the rock, when the piston strikes them, it seems unquestionable that the greatest efficiency of the blows of the piston would result, providing they were properly timed, for no energy would be lost by reason of the inertia of the drill steel, but only that due to heating, resulting from the imperfect elasticity of the metal. The question of the proper timing of the piston blows opens another phase of the matter, namely, the reaction of the rock upon the drill steel; and this effect is the more pronounced with harder rock. It tends to speed up the piston and is so noticeable in running a machine against a metal block as to invalidate, as too high, all air-consumption tests so conducted. The effect of these reactive vibrations upon the drill steel may prove very marked and serious. Where the reactive vibrations interfere with on coming compressional waves, considerable energy is dissipated, and at times one may be so fortunate as to detect points of increased temperature (probable nodes) upon a drill steel which is cutting ground; and it is no uncommon thing to see a drill steel, in service, break at two points (into three pieces) simultaneously, probably from fatigue because of these vibratory stresses. On the other hand, if these vibrations synchronize at the bit it is quite possible that the chipping forces are greatly augmented, and such an explanation may readily answer those puzzling drill tests in which a dull or broken bit exceeds a finely formed bit in drilling speed. For a long time at Franklin a tally was kept of the different individual drill steels which entered into the testing, with the hope of determining that some particular piece of steel produced the greatest cutting speed, but no conclusion could be drawn from the records, except that the changes in length due to resharpening probably masked any possibility of determining the suitable lengths for maximum efficiency. And it seems quite plausible that such a result should be expected if the possible wave lengths of the compressional vibrations in the drill steels are considered. Probably these reactive vibrations occur to a great extent, as well, in the process of drilling, where the steel dances in the chuck and against the rock, for steel breakage appears equally as high, if not higher, with such a type of machine as with the pneumatic feed, and tests comparing these two types for such effects might prove very interesting as well as instructive.

But still other factors influence the force delivered at the rock. If the anvil block or tappet is not in contact with the drill when the piston strikes, a considerable energy loss occurs through the transference of momentum to several pieces. If the steel is bound in the chuck bushing, a great amount of the energy is absorbed by the friction. If the steel is not straight, it loses energy because of the flexure. If the chuck is badly worn, the axis of the steel does not coincide with that of the drill and there is a loss due to the oblique, eccentric impact. If the steel is tight in the drill hole, or if the friction against the side of the hole is great because of its depth, the velocity of the steel, as a projected body, is lessened and the drilling speed is reduced.

The length of the drill steel is an item generally credited as an important influence, and common opinion supports the idea that the cutting speed falls as the length of the steel increases, although some people, on the contrary, feel sure that the long steels drill the fastest. The tests conducted at Franklin do not lend an unqualified support to either view, for the peculiarities of different types of machines play so important a part. For example, if the air feed is very strong in a stoping drill the additional counteracting weight of a long and heavy steel may so improve the working conditions as to indicate a superiority for the long steel, and if the air feed is weak the reverse may be true; if the drill steel cuts by virtue of a dancing or “jumper” action, the mass added with length may so reduce its velocity against the rock as to bring it below the amount required for efficient chipping; if the piston normally delivers too heavy a blow for the rock, the drilling speed may be improved by the added inertia of the long steel; and if the steel is always against the rock when a blow is delivered, it is doubtful whether the length of the steel plays an important part unless the permitted decrease in the gauge of the drill bits aids the cutting speeds. It is, of course, to be understood that the above considerations of drill-steel lengths refer to the performances with bit gauges of the same diameter.

The use of an anvil block is considered by some drill designers to necessitate a loss of from 20 to 30 per cent, of the power of a drill, but actual tests do not always indicate such a condition when the identical steel is tested in the same drill with and without a tappet. The results probably depend upon how frequently the tappet is struck when away from the shank of the steel, and also upon the suitability of the machine to the rock, for if its blows are too heavy the intervention of a loose tappet might reduce their force, with a benefit in drilling speed. The use of water at the bottom of the hole ordinarily consumes about 10 per cent, of the cutting speed if there is no tendency for the drill bits to lose their temper, and compressed air for cleaning the holes encourages a greater drilling speed, providing the cushion of water in the bottom of the hole does not have a benign influence in reducing too powerful a blow upon the rock.

The manner in which a drill is rotated has a bearing upon the amount of work accomplished, and with hand-rotated tools, a vigorous rotation with a rapid and wide arc of swing produces the best results; with power-rotated drills it is possible to reach such a speed as to abrade and dull the drill bits, and consequently lessen the drilling speed. It seemed that, with a positive and constant rotation, the axial planes of the cutting edges of the drill bits should be at the same angle with the cut surface as the resultant velocity vector, as estimated for the rotative and striking velocities; and such a bit was tried at Franklin without showing a change in cutting speed, probably because with either bit the chips came out in flakes, as previously described.

In view of the fact that the subject of hammer drills is more or less in its infancy and literature in regard to them is rather limited, it seems desirable to correct at the earliest opportunity any typographical or other errors which, if accepted without investigation, might work to the detriment of the art of drilling. In this connection it seems that some statements should be corrected in the 1910 edition of Eustace M. Weston’s book, Rock Drills, in the chapter Philosophy of Process of Drilling Rock, under the sub-heading of hammer drills. In considering the kinetic energy of a blow he states, on p. 139:

“In other words, to double the energy of a blow it would be necessary to double the mass, or weight, if the velocity is the same; but to double the energy, keeping the mass the same, the velocity must be increased four times. The weight of the piston hammer of the largest type of drill is 15 lb. The weight of piston, steel, etc., of a piston drill varies from 60 to 125 lb., so that a blow of equal force can be delivered by a hammer drill only by increasing the velocity of the hammer very greatly. This is acknowledged, for as one hammer-drill maker states, the weight of the piston is one-fourth that of a piston drill; but the velocity is four times as great. To give a blow equal in power it should be sixteen times as great.”

A mathematical error appears to have been made in the premises of Mr. Weston’s argument and his consequent deductions as to the practical impossibility of hammer drills being able to compete with piston drills are quite logical, but probably at fault.

If the kinetic energy of a body, such as a drill piston, is designated by K, its velocity by V, and its mass by M, then,

K = ½MV² and K1 = ½M1V1²

Now if M1 equals M and K1 is, say, twice the value K, then,

V1² = 2V² and V1 = V√2


V1 = 1.414V

So the velocity of the piston in a hammer drill need be only 1.4 times as great as that when the kinetic energy of the piston is cut in half.

Again, in the example comparing the piston weights of piston and hammer drills, Mr. Weston appears in error in stating that the velocity should be 16 times as great, for if the piston of a hammer drill is one- fourth the weight of that of the piston drill the velocity of the hammer-drill piston need be only twice as great as that of the piston drill in order to deliver blows of the same energy; and the hammer drill will also surpass the piston drill since it will strike twice as many of such blows per minute. The necessity of using high air pressures in hammer drills is only incident to the peculiarity of certain drill designs and is not dependent upon the divorcing of the piston from the steel. If we are to consider the shock upon the parts of two drills of equal capacity it is evident that with the shorter piston strokes in hammer drills, with the increased number of blows, whose final striking velocity is equal to that of a piston drill under comparison, the weight of the hammer-drill piston may be less and the energy in each individual blow may be less in order that the same amount of energy per minute be developed. Therefore, the shocks upon hammer-drill parts are more frequent but not as heavy as the shocks upon piston drills of equal capacity.

It is extremely difficult to get adequate figures as to the maintenance of drills unless some special forms are kept, which become to all intents a ledger account of each individual drill, for questions naturally arise as to the cost per foot of hole drilled, the length of time the machine has been in service and has been running, the number of holes drilled, the kind of rock encountered, and the supply of steel used, as well as the drill parts replaced. The New Jersey Zinc Co. uses a system of punched slips for shift bosses’ reports and “drill record” slips are a part of the scheme. These are punched in duplicate by the shift boss and filed at the mine office and main office, where the information is transferred to large sheets, of which each one accommodates the record of one machine for a month, and the footings are carried forward so as to indicate the total work accomplished and maintenance of any machine “to date.” The repair parts are designated as to whether they are new or old (second-hand) ones, and


original and subsequent drilling-test records are noted on the same summary sheet.

Figs. 3 and 4 show the record slip and sheet. The locations of drills are by stope-slice co-ordinates, the class of work (whether raising, drifting, stoping, block-holing, or drilling chutes) is indicated, the kind of rock (ore, limestone, gneiss, pegmatite, garnet, or feldspar) is punched, and if the machine is idle, broken, or being cleaned those points are recorded.

The compressed-air rock drill made revolutionary changes in mining methods and in the reduction of mining costs in units of labor per ton of ore, and at Franklin even more marked savings have been made through the development of hammer drills. There, in the days of hand drilling, a total of 8 ft. in three drill holes with varying diameters of 1¾ to 1¼ in. was considered a fair 10-hr. shift’s work, and possibly 8 tons of ore would be broken per drill-shift or 4 tons per man-shift. With 3-in. reciprocating rock drills from 20 to 40 ft. of drill holes, ranging in diameter from 2½ to 1½ in., would be the average work for a 10-hr. shift, although on rare occasions some men might drill as much as 80 or 90 ft. 0: holes in a shift, and possibly 20 tons of ore would be broken per shift, on 10 tons per man-shift, since two men were needed on a drill. It seems


that as a rule a greater tonnage per foot of hole was obtained with hand drilling because of the fact that, rather than dismount and reset heavy drill columns, machine men would tend to place as many holes as possible from one set-up, therefore many holes were placed disadvantageously for breaking efficiency. Another cause, which would contribute to the same results, would be the difficulty of starting holes with piston drills on uneven sloping faces, so that holes were frequently deflected from the direction in which they were supposed to be placed. These figures would lead to the rough estimate that 2½ times as much tonnage per drilling man-shift was accomplished by piston drills as by hand drilling.

With hammer drills 80 to 100 ft. of 1¾ to 1¼ in. drill holes are placed by one man in a 10-hr. shift, about 150 to 200 tons of ore will be broken per drill-shift and the same amount per drilling man-shift, or 15 to 20 times the amount broken per man with reciprocating drills. Of course the entire credit for such increase in tonnage cannot be given to the type of drill, for improved organization, system of working, and supervision have undoubtedly played an important part; but the greater mobility and flexibility of the light hammer drills have permitted and encouraged a more efficient placing of drill holes; have cut in half the labor necessary to run a drill; and permitted a more effective supervision and mining scheme. The actual tonnage broken per man working in a stope will not be so high, comparatively, since it has been found worth while to place additional men in stopes to sledge and block-hole large chunks of ore, which were formerly often allowed to become buried and so proved obstacles to high tramming efficiency by blocking chutes through which the shrinkage stopes were drawn into tram cars.

Table II shows the gains which have been made with the adoption of hammer drills by the New Jersey Zinc Co. It is to be regretted that no records of tonnage and labor were available for earlier years, so as to cover the average efficiencies before the advent of hammer drills and back to the days of hand drilling. The different divisions of mining work are classified in this table as drifting, raising, stoping, and open cut or quarry, and it may be interesting to summarize the important features, reducing the labor to an hourly basis, inasmuch as a change was made from a 10-hr. to an 8-hr. shift basis in July, 1913.


There have been no radical changes in the placing of the drill holes in drifts since the adoption of the air-feed hammer drill for this work, but one man with a single machine is now placed in a heading; he is instructed to “pull” a “round” each 8-hr. shift, stopping overtime if necessary, and to accomplish an advance of 3½ to 4 ft. per round. Two men operating a reciprocating rock drill formerly made an advance of a 5 to 6 ft. round in five 10-hr. shifts. So the drilling labor (runners and helpers) per foot of advance averages 18.4 hr. for the entire mine during the year 1910, when reciprocating rock drills were solely in use. As shown by the average for 1913, hammer drills have reduced this figure to 5.7 hr. per foot of advance, or about one-third the former labor of drilling and blasting. The explosive costs have also been reduced by the use of hammer drills from the figure of $1.84 per foot of drift during 1910 to $1.40 per foot in 1913, for two probable reasons.

First, hammer drills permit the placing of drill holes smaller in diameter than those bored by reciprocating drills, so that an unnecessary amount of explosive is not required merely to fill the holes sufficiently to distribute the force of the explosion.

Second, the flexibility and ease of rigging the light hammer drills permit and encourage a more efficient placing of drill holes. The almost exclusive use of 1 by 8 in. explosive cartridges now, as contrasted with the 1 1/8 by 8 in. cartridges formerly used, demonstrates the first contention, for in terms of 1-in. powder, the equivalent of 36.8 sticks per foot of drift was used in 1910, and 34.6 sticks per foot in 1913. The drill shifts per foot of advance have been lowered from 0.44 in 1910 to 0.20 in 1913, and the corresponding drill hours from 4.4. to 1.8.

The different drifts may vary in size from 6 by 7 ft. to 8 by 11 ft. in section, and perhaps 7 by 8 ft. is an average section. Because of the compact, tough nature of the ground, it requires from 20 to 30 drill holes in a round, and 24 would be a fair average, so the drilling operation is an important factor of the drifting costs. The following comparison of the average drifting costs for each year shows the saving which has been possible because of hammer drills; but only the cost of drilling labor and explosives is considered. The record drift in 1913 was driven for $2.06 per foot.



In 1908, using 2¼-in. piston reciprocating rock drills, 0.7 ft. of 6 by 6 ft. raise per 10-hr. drill shift was made with a labor expense of 28.5 man-hours per foot of raise. About 27 ft. of drill holes were placed per shift, 24 holes were placed in a round, and 16 lb. of explosives were used per foot of raise advance, at a cost of $2.70 per foot for supplies. Since labor was then paid $2 and $1.55 per 10-hr. shift, the total cost of raising was approximately $7.50 per foot of advance.

During the same year, 1908, hammer drills were introduced, and an advance of about 1.5 ft. per drill-shift was made with a labor expense of 13.3 man-hours per foot of advance. About 50 ft. of drill holes were placed per shift, 24 holes per 5-ft. round, and 10 lb. of explosives were used per foot of advance, at a cost of $1.75 per foot for supplies and a total cost of $4.10 per foot of raise, or only 55 per cent, of the cost with the reciprocating rock drills.

The development of hammer drills with increased drilling speed permitted the reduction of the drilling labor to 7.8 hr. per foot of raise advance, and the explosives cost to $1.54 per foot of raising done in the year 1910; and a further reduction to 4.8 hr. of drilling labor during 1912,





and an explosive cost of $0.93 per foot, although the wages were $2.20 and $1.70 per 10-hr. shift. These costs rose slightly in 1913, since wages rose to $2.25 and $1.85 for 10-hr. shifts, and in July of the same year the working hours were lessened from 10 to 8 and the hourly wage was increased to $0.281 and $0.231. However, the cost per foot was then only 5.2 hr. of drilling labor and $1.03 per foot for explosives. About 18 drill holes are now placed to pull a 5-ft. round and two men are expected to blast a round each 8-hr. shift and are each paid 11 hours’ time for performing the task.

The average raising costs for operating labor and explosives have been as follows:


The record short raise (of about 50 ft. in length) for 1913 had a cost of $1.65 per foot, and the record long raise (about 100 ft. long) had a cost of $2.09 per foot, with explosive costs, respectively, of $0.77 and $0.99 per foot of raise.


In 1909, when about 74 per cent, of those drills placing holes in the solid orebody were of the reciprocating type of 3-in. piston diameter, the ore production averaged about 20 net tons of ore broken from the solid per 10-hr. drill shift, with an equivalent of 1.1 sticks of 1 by 8 in. of 50 per cent, dynamite per ton of ore.

In 1910, when about 72 per cent, of the producing drills were air-feed stoping (hammer) drills, the tonnage per drill shift rose to 38 net tons with about the same amount of explosive (which cost $0.055 per net ton of ore broken), and 13.2 tons were broken per 10-hr. shift of men working in stopes, or 1.32 tons per hour. Although there is no record of the breaking labor prior to this year, the fact remains that in the actual running of the drills only one man was used with a hammer drill while two men were employed with each reciprocating drill.

In 1911, when the hammer drills were about 80 per cent, of the total, the stoping efficiency profited by the improvements in the drilling speed of the hammer drills, and 121 net tons were broken from the solid per drill-shift with about 0.8 stick of 50 per cent. 1 by 8 in. dynamite per ton (at an explosive cost of $0.041 per ton), and 2.03 net tons were broken per man-hour of men working in stopes.

In 1912, when about 98 per cent, of the stoping drills were hammer drills, 195 net tons were broken per drill-shift with about 0.87 stick of 50 per cent. 1 by 8 in. dynamite per ton (at an explosive cost of $0.045 per ton), and at the rate of 2.56 net tons per man-hour in the stopes.

In 1913, when all the stoping drills were hammer drills, the length of the working- shift was reduced from 10 to 8 hr. in the middle of the year and the tonnage broken per drill-shift fell proportionately to 170 net tons, but remained at approximately the same hourly rating as for the year 1912. However, the tonnage broken per man-shift in the stopes increased slightly to 26.7 net tons (at an explosive cost of $0.049 per ton), with the consumption of 0.89 stick of 50 per cent. 1 by 8 in. dynamite per ton. The tonnage broken per man-hour was 2.97, which showed a steady gain over previous years.

It should be noted that the explosives charged against stoping include those used by the trammers in blasting ore in the chutes, and thus represent all the dynamite necessary to reduce the ore to the proper size for being handled through chutes and in the mill.


In order to provide broken rock for filling material to fill empty stopes to support the remaining orebody, “mill-holes” are developed in limestone country rock at the surface. For some years it was the practice to use 30-ft. bench-holes in the opencut for quarrying the rock, both 3-in. and 3½-in. reciprocating rock drills being used. It took steady work for two men to sink one 30-ft. hole in a 10-hr. shift, and their work was hazardous because of the inconvenient localities where set-ups were made, and because of the clumsy weight of their machines and the long, heavy drill steels which were handled. After the success of hammer drills in the underground mining operations, they were tried in the opencut work in 1912. Small holes were drilled to an average depth of 16 ft., and were given lighter burdens than had previously been the practice, for the object was to distribute the dynamite more evenly in the rock, as contrasted to churn-drill or mammoth blasts. In the tough, crystalline Franklin limestone this application of hammer drills to quarrying has proved superior to the heavy or mammoth blasts, for the same tonnage can be produced from a bench with a saving of labor and powder, since a great amount of expensive block-holing is avoided. A machine will drill about 100 ft. of holes in a shift with a heavy hammer drill and two men can drill only 30 ft. with a rock drill.