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By O. B. Bennett

WHEN diamond drilling was introduced in the Rhokana mines in 1939 it was used principally for pillar removal and for completion of the upper portions of shrinkage stopes which were being affected by increasing pressure. This method of drilling long blastholes proved so successful that it was extended gradually to cover stoping, pillar recovery, and hanging cave work. BY 1949 virtually all the ~roduction of Mindola and Nkana was being obtained by this method. At the present time 87,500 ft are drilled each month by the 80 diamond drills in daily operation. Responsibility for control and issue of diamond drilling equipment and crowns, as well as tabulation of all performance figures, was taken over by a sPecially formed Roto drill department, which also investigated the problems encountered with this new method. To assist this department a fully equipped test chamber, Fig. 1, was established underground where performances of various types of machines and equipment could be studied under conditions as nearly uniform as possible. Since the establishment of this department, which was eventually taken over and incorporated into the study department, considerable experimental work has been done on every aspect of the subject. The problems can be classified broadly under four headings: improvement of drilling equipment, crown design, machines, and stoping layouts. One of the major problems with drilling equipment has been to eliminate vibration. Owing to flexing of rods in the hole, severe friction is set up on the back end of the 'Ore barrel and On any high spots in the rods, inducing harmonic vibration in the string of rods and causing the crown to chatter against the face. This not only causes premature crown failure but also reduces penetration speeds and increases wear on the machines and rods used. In the early days, when only holes of EX size were drilled, vibration was largely overcome by periodic greasing of rods and core barrel during each run, but with the change-over to the larger BX hole it became obvious that application of grease by hand was inefficient and time-consuming, and attempts were made to perfect a self-lubricating core barrel. A series of these core barrels was made up and tested and a number of the latest type were used under normal operating conditions, but although footages up to 120 ft were drilled without refilling the overall performance was inconsistent, and the idea was shelved in view of the success of the stabilizer rods referred to later in this paper. At the same time tests were made with barrels 5 ft and later 6 ft long instead of the normal 2 ft. Although a slight improvement was noticed, greasing was still necessary. It was found that rod vibration increased as the core barrel became worn, and in an early test chamber experiment crowns drilled with a worn core barrel averaged 95 ft with a diamond loss of 4.76 carats, whereas the same type of crowns with a new barrel averaged 228 ft with a diamond loss of 3.13 carats. until then all BX drilling had been done with B-sized rods, but during a test on a string of BX-sized rods it was noticed that vibration was negligible. Because of the larger surface area of metal bearing on the sides of the hole, however, the friction and resistance of rods of this size rendered them impracticable on any but the most powerful of the machines, The use of stabilizers spaced evenly along the rods was the next logical step, and for this B couplings, see Fig. 2, were set with three tungsten carbide inserts 1 in. long placed around the periphery equidistantly and at an angle of 45" with a right hand lead. These were placed immediately behind the core barrel and then at 12-ft intervals, as it was found that vibration still occurred when the stabilizers were more than 15 ft apart. The effect of these stabilizers was immediately noticeable; holes were drilled with a minimum of vibration, penetration speeds were improved, and as it was no longer necessary to grease the rods there was a marked decrease in the overall drilling time for each hole. While tests were being made with the stabilizer comeb periodic were taking place with a set of tapered threaded rods, and because there was marked improvement in efficiency it was decided to incorporate the stabilizers and tapered threading in all new rods ordered for Rhokana. The feature of these rods is that only four full turns are required to tighten the coupling as against nine for the present type of B rods. Also, as they are self-centering it is virtually impossible to crossthread them. Each rod has a male 5" tapered Acme thread, Fig, 3, on one end and a female at the other, so that separate couplings are unnecessary, and every fifth rod has an

Jan 1, 1955

By Ingvar Janelid

DURING the last ten years, in the effort to save manpower and costs, methods of drilling and blasting in Sweden have changed and developed in a revolutionary manner. These developments have been accompanied by extensive alterations in methods of tunneling and in mining technique. Since the end of the 1940's there has been a 100 pct change-over in Sweden to the use of tungsten carbide tipped drill steel. Both investigations and practical results have shown clearly that a tungsten carbide chisel bit fastened directly on the drill steel is the most economical for Swedish conditions. This applies to drill steel up to the longest size it is practical to handle. In terms of drilling footage, durability of the drill steel and the tungsten carbide bit is generally the same. In this respect, therefore, a detachable bit offers no advantage over integral steel. Transportation of integral steel to the grinding plant, instead of detachable bits only, involves heavier work, but this is counteracted by placing grinding plants closer together or by grinding at the working place itself. The drilling rate with integral steel is higher, the trouble of changing bits is avoided, and the difficulties which always arise with joints are eliminated. At present steel dimensions are usually 7/8 in. with a shank 41/4 in. long, whereas the international standard shank was formerly 31/4 in. The longer shank renders possible a better guide control of the steel and consequently longer life for the rotation chuck and the water-flushing pipe. After extensive experiments and investigations, agreement has been reached in Sweden on certain series of standards, shown in Table I. Other steel dimensions are, of course, available for special purposes. The standard drill-steel set begins with a 34-mm diam and has a length interval of 2 2/3 ft. On the other hand, for ordinary drifting it has been proved more economical to use greater length intervals on the steel, so only one steel change is required for a 61/2-ft round. Moreover, owing to the relatively close hole spacing necessary in a drift round, it has been possible to reduce the diameters of the tungsten carbide bits. This results in faster and cheaper drilling. As shown by the table, the lengths of drill steels for normal drifting are 33/4 ft and 71/2 ft, with 29 and 28-mm diam bits. During the last few years operators have begun to use 3/4-in. drill steel in place of the 7/8-in. drill steel chiefly employed earlier. This is a valuable development in drilling technique, particularly in view of

Jan 1, 1955

By L. Adler

Some estimate of the geologic pressures existine at the site of a proposed underground opening is essential for its economic design and construction. Regardless of the size of such an opening, drill hole, or mine stope, these primitive stresses need evaluation; at present a number of techniques are available for this purpose. However, a new method of analysind stresses is proposed in this paper. Some estimate of the geologic pressures existing at the site of a proposed underground opening is essential for its economic design and construction. Regardless of the size of such an opening, drill hole or mine stope, these primitive stresses need evaluation. At present a number of techniques are available for this purpose, and another one is to be proposed. These may be grouped into three general methods: 1) precedent, 2) instrumentation, and 3) theoretical analysis. The first two are covered in the literature where their advantages, as well as limitations, are discussed.'9' Generally a theoretical approach combines structural geology with engineering mechanics. The former deals with the geodynamic causes of residual stresses (stresses not related to overburden weight) and presents interesting, but controversial, views.'" A general and useful conclusion, however, can be drawn - vertical residual stresses are much less prevalent than horizontal. However, if for some reason, the values and effects of horizontal pressures were better known to engineers the following analysis

Jan 1, 1962

By J. Moore, D. Gault, R. V. Lugn

Impact of small projectiles with velocities between 0.9 and 7.3 km per sec on basalt produce craters chiefly by the ejection of fragments. Weight-size distributions of the ejecta are linear for part of the size range of the fragments. Cumulative weight-size distributions of the ejecta from different craters vary because of erratic formation of large fragments and a large percentage of ejecta below 0.1 mm. In the range of velocities studied, the crater depth is approximately proportional to the 0.75 power of velocity of the projectile. This relationship may represent a transition regime between the fluid impact regime and the undeformed projectile impact regime. The volumes per unit of projectile energy of the craters produced are comparable to those reported for lower speed projectile impact and drill bits. An experimental investigation of projectile impact on rock is being conducted jointly by the Ames Research Center of the National Aeronautics and Space Administration and the U.S. Geological Survey. Projectiles are launched with light-gas guns at the Ballistics Range of the Ames Research Center. The craters and ejecta produced in these experiments are being studied by the U.S. Geological Survey. Steel, aluminum and polyethylene projectiles weighing from 0.25 to 0.015 g were accelerated to velocities of 0.88 km per sec (2,900 fps) to 7.3 km per sec (24,000 fps) with a light-gas gun, using hydrogen as the propellant medium. Projectiles impacted the targets normal to plane rock surfaces in an atmosphere of air with a nominal pressure of 25 to 50 mm of mercury. The metal projectiles were mounted in supporting four-piece nylon sabots, which guided the projectiles down the launching tube. After launch, aerodynamic drag acted to separate the sabots from the projectiles. Impact velocities were determined with time measurements and spark photographs of the projectiles in flight. Ejecta were collected in lucite and aluminum containers placed around the target block. A small hole in the containers permitted projectile entry. Impact of projectiles of the size employed produced craters in the rock targets chiefly by ejection of fragments. The craters are small, roughly conical depressions with crushed and intensely fractured rock at the bottoms of the depressions. The ejecta are composed of pieces of the projectile and rock fragments ranging in size from a few tenths of a micron or less to several centimeters in maximum dimension. In this paper, the structural features of the craters, the weight-size distributions of the recovered ejecta, and crater depth and crater volume relationships are described. DESCRIPTION OF CRATERS Typical craters (Fig. 1) produced by experimental projectile impact in basalt are shallow, wide base

Jan 1, 1963

By R. A. L. Black, J. E. Jennings

The problems of slope stability in open cast mines are examined. A criterion, the instantaneous stripping ratio, is suggested for use in the design of pit slopes and as an index of control at all stages of the exploitation of the orebody. This criterion may also be related to safety and to the maximizing of profits. Two approaches to safe design of slopes, one from the point of view of soil mechanics and the other from the point of view of rock mechanics, are described and critically compared. The effect on slope stability of the regional tectonic stress field is discussed. Explanation is given of the application of the Terzaghi effective stress concept to problems in both soil and rock environments, including possible effects when borehole deformation techniques are used to measure stresses in rock in situ. A company embarking on open cast mining is faced with two basic questions: 1) will the venture be profitable and 2) can it be carried out with safety? The two questions are obviously interrelated because both depend upon the slope angles adopted for the pit. The cost of open cast mining depends largely upon the stripping ratio, defined as the volume of overburden that must be removed to the volume of ore recovered. The steeper the slope angles, the more favorable the stripping ratio, but the stability of the slopes, and therefore the safety of men and materials in the pit, require that the angles be made relatively flat. Consequently, the design depends upon a compromise: the slopes must be steep enough to result in economical mining and flat enough to ensure safety. Today there is a world trend towards deeper mining by open cast methods. This is mainly due to the higher cost of underground mining and the lower cost of open excavation resulting from the introduction of large modern earth and rock-moving machinery. Deeper open cast mining has led to a very great increase in the volumes of wasted overburden, and much thought is being given to the considerable economic benefits which would result if pit slopes could be made steeper by even one or two degrees. The result has been an increasing interest in research into slope stability problems in the expectation that the application of scientific principles will provide solutions leading both to economies and to more predictable operation of the mines. In at least four places, the U.S., Australia, Western Germany and Southern Africa, comprehensive investigations are at present in progress, and it is the object of this paper to review some of the more important aspects of these programs. RELATION BETWEEN ECONOMY OF OPERATION AND SLOPE ANGLES The main object of any industrial enterprise in a free economy is to make a profit. In mining, where profits must be made from a wasting capital asset, a further definition of the profit motive is imposed, namely that the economic aim should be to maximize the total possible profit. But since profits deferred for a number of years are not as valuable as profits that are immediately available, it is usual to discount deferred profits at some rate of interest which reflects their current or present value in the hands of the investor. Thus, in mining the logical aim can be defined more closely still: to maximize the present value of the total future profits. In commercial terms the present value represents the market value of the operation as a going concern. Many factors enter into the forecast of current and future profits. They include the volume of production, the price of the salable product and the cost of production. Current profits can be estimated with reasonable accuracy by the operating engineers; but it is the task of economists and management to assess future trends in the mineral markets and in the economy as a whole, and together with the engineers, they can then estimate the future cost and profitability of the mining operations. In open cast mining, the rate of ore production once decided on remains a virtually constant factor unless a major change in the capital structure of the mine is made. But one factor over which the operator has a considerable measure of control is the ratio of the volume of ore mined to the volume of waste over-

Jan 1, 1963

By L. Adler

A relatively simple and effective technique is proposed to deal with the failure of geologic masses containing weakness planes. This technique, termed irextended" rupture envelope, provides a means to clearly distinguish rock and soil from fractured rock and, on a theoretical level, considers a matrix-inclusion system. It may also constitute the link between Mohr's and Griffith's failure theories. Included in the article are five sample problems concerning site location, structural design, strengthening the plane of weakness, changing the magnitudes of major and minor principal stresses and changing the orientations of major and minor principal stresses. Engineers trying to solve design problems in rock and soil mechanics have often assumed an iso-tropic and homogeneous medium. For naturally occurring geologic masses this assumption has always been viewed with regret. The following analysis which deals with the effects of plane anisotropy on failure presents a solution to this. The planes of weakness are joints, faults, bedding planes, fracture and cleavage (natural and induced), foliation and schistosity. The physical properties of these weakness planes include an undulating surface, frictional resistance, discontinuous extent, filling material, tightness and presence of water to cause a hydrostatic pore pressure. All of these properties relating to failure can be summarized in Mohr's rupture envelope for the plane of weakness. This, along with the geometrical orientation, strike and dip, of the plane of weakness produces what is termed an irextended" rupture envelope. MOHR'S STRESS CIRCLE AND RUPTURE ENVELOPE The significant features of these concepts are listed below.' 1) A point on the u - r coordinate plane represents the normal and tangential stress components on a reference area in the material. This area can have any orientation. (Fig. la)

Jan 1, 1963

By J. B. Fletcher

The Miami mine first started operations in 1910. For convenience, the history of the orebody can be divided into the following categories (Fig. 1): 1) 1910 to 1925: 24.4 million tons of high grade ore, mined by top slicing, sublevel caving, etc. 2) 1926 to 1954: 102 million tons of low grade ore, mined by block caving.

Jan 1, 1961

By F. H. Buchella, J. F. Buchanan

The San Manuel copper deposit is located about 45 miles northeast of Tucson. The concentrator, smelter, administration building, and other plant facilities are located about seven miles southeast of the ,nine area at the new town of Sari Manuel as shown on Figure 1 (Sketch).

Jan 1, 1961

By L. Adler

Two basic cases involved in the design of an opening in bedded rock are: 1) where the beds deflect from each other so as to be separated; and 2) where the beds deflect onto their lower neighbor, loading the latter while reducing their own load. Analysis of case 1 is relatively simple, while the interference of the original deflection curves in case 2 complicates analysis of the latter. Figures and formulae are presented for determination of interference loading in this situation. The design of an opening in bedded rock involves two basic alternatives.1'2 In one case the beds deflect away from each other, so as to be separated from their neighbors; in the other case the beds deflect onto their lower neighbor, loading it, while reducing their own load. This is shown in Fig. 1. The first case lends itself to a simple analysis, since the upper and lower boundaries of each beam remain a free surface. In the latter case, however, due to the interference of original deflection curves, the analysis is more complex. Earlier investigators of this interference problem solved it by making the tacit assumption that this type of loading was uniformly distributed along the beam's length. While there is no intuitive justification for this assumption, its use does lead to an approximate solution. Still, a more exact solution does not require this apparently unrealistic restriction. To achieve initial simplicity, the analysis will at first be limited to a two-bedded sequence, as in Fig. 2. Since both beams have identical end conditions and spans, the original deflections of their center lines are members of the same family of curves, as shown in Fig. 3. Therefore, if interference occurs at one point along their center lines, it will occur at every point. For convenience, the midspan point is that usually examined. Deflection at midspan for a uniformly loaded beam regardless of end conditions is: where c is a constant depending on end conditions; w is the body load and equals yh, lb per ft; y is the specific weight of rock, lb per ft'; L is the span, ft; E is Young's Modulus, lb per ft2; I is moment of inertia (for a rectangular section), I = bh3/12 = h3/12, (ft4); b is the width of beam and equals one; and h is the thickness of beam, ft; or 3U > yB an Substituting the appropriate values into Eq. lb, interference between the two beds can be predicted. Should it occur, the additional loads along the lower boundary of the upper bed and the upper boundary of the lower bed must be determined. From the principle of action and reaction, these added loads are known to be equal and opposite, as indicated in Fig. 4. Since the final deflection curves for both beds are identical (Fig. 3), their successive derivatives of y with respect tox must also be identical. That is:

Jan 1, 1961

By C. D. Michaelson

Bridging the gap between have and have-not nations is one of the necessities of the present era. The responsibility for accomplishing this must be assumed by the affluent industrial societies of the world, both through their governments and through private enterprise. Despite a discouraging number of instances where companies established in underdeveloped countries have been expropriated, U.S. industry must not abandon such ventures. Instead, it should tailor them to the needs and aspirations of the host nations. One way in which industry can minimize the danger of expropnation is to enter into partnership with the government of the host nation. Kennecott Copper has recently done this in Chile, becoming 49% owner of a company formed to expand and operate the El Teniente mine. So far, the partnership has produced marked benefits not only for Chile, but also for Kennecott. The memory of D.C. Jackling serves chiefly to remind us how much we owe to the pioneers of our industry. Jackling was one of the giants. Orphaned at the age of two, he made his way through sheer persistence and plain hard work. His brilliance gave rise to a vision, and his tenacity made his vision a reality. Foreseeing that economies of scale would make the handling of low-grade ore profitable, he introduced the revolutionary concept of open-pit copper mining. During World War I, the defense effort of the United States was irnrneasurably strengthened by the tremendous supplies of copper from the western mines that Jackling had made possible. Again in World War 11, the demands for copper were supplied, and the nation once more had profound reason to be grateful to this rugged, stubborn man who had long ago ventured into the mountains and unlocked their wealth. We are now embarked upon another great era of change, and we must devise new concepts to meet it. The task is not only to create new techniques of mining, but to forge better ties between men and institutions which will let us work in harmony with the techniques we have. We might recall that while Jackling was a man of towering pride and formidable will, he was also adept at getting others to share his vision and join his purpose. Today our reach for cooperative endeavor must go far beyond our own shores as we seek to develop the buried riches of distant lands. And we would do well to remember that Jackling's remarkable insight would have come to nothing had not British investors been willing to provide much of the initial capital. As was true of so many early enterprises in America, European capital was absolutely essential. CORPORATIONS ABROAD - AN EXPANDING ROLE Today the situation is quite different. The United States is now a capital-surpIus nation and looks abroad for opportunities to employ its capital profitably. On the other hand, we have become dependent on foreign sources for supplies of vital raw materials, while we seek

Jan 1, 1970

By L. Adler

The proper transfer of roof loads from props and bolts to ribs and pillars can result in appreciable savings. The author shows how to plan such load reduction in underground mines. For openings in bedded rocks, analyzed by simple beam theory, it has been shown that roof loads can be shifted from one support to another.&apos; This transfer is effected by controlling the relative deflection of adjacent supports. If this action is used to reduce loads in the props or bolts and increase them in the ribs or pillars, it could be productive of economies averaging 30 pct. Such savings can be realized by: 1) using smaller artifical supports such as props and bolts, and 2) increasing loads in the rib to induce eventual failure for longwall caving. Since the rib is relatively load-insensitive, it provides a natural load sink. However, caution must be used in applying such permissive deflections, for while the supports are being handled properly, the stresses in the roof itself are very sensitive to such manipulations.1 An underground opening overlain by a relatively thin immediate roof, and a thick main roof, is usually laid out as shown in Fig. 1. The total open span (T) between pillars or ribs is calculated on the basis of the main roof, and the secondary spans (L) on the basis of the immediate roof, so that a complete accounting of the existing loads is made. For this purpose the immediate roof is treated as a uniformly loaded continous beam, which is an indeterminate structure as shown in Fig. 2(A). The term systematic supports places the following two restrictions on the continuous beam: 1) Spans between all supports are equal: uniformity of spans. 2) Loads on all supports (excepting ribs) are equal: uniformity of supports. The usual mode of analysis of such a structure by beam theory is to divide it up into a series of simply supported beams with end moments as shown in Fig. 2(B)2 The problem now is to transfer a maximum roof load to the rib from the span immediately adjacent to it by permitting its support to deflect, and yet not fail the roof. Once this has been done, the load can be equalized on all other supports by permitting further degrees of relative deflection. It is necessary, therefore, to determine an expression for the constant load on each support, and the required deflections, in terms of properties of the roof. From Fig. 1: T/L0 = No and T/L = N [1] Where: T = allowable span of main roof, ft. (Calculated from S = Mc/l) S is allowable modulus of rupture, psi. M is maximum moment, lb-ft (in this case wT2/12). c is distance from neutral axis to extreme fiber, ft. I is moment of inertia with respect to the neutral axis, ft4. L,, = allowable span of immediate roof, ft (calculated as was T). L = actual span of immediate roof, ft. N0 and N are pure numbers where No < N and N is the integer just higher than Nu. For example: If T =100 ft and L0 - 15 ft 100 N0 =100/15= 6.667 . . .N= 7 L = T/N = 100/7 = 14.28 ft Consequently N = No + n where n is a positive fraction i.e.: 0< n < 1 Since from Eq. 1, T = NL = N0L0 Lo /L = N/No = N/N-n

Jan 1, 1961

By William T. Folwell

The Lucky Friday mine east of Mullan, Idaho, is an outstanding example of a property in the Coeur dlAlene district where a small and insignificant-appearing silver-lead-zinc vein at the surface has changed at depth into a large vein of great importance. The Lucky Friday vein has little if any surface expression and above the 1200 level the ore shoots are small and discontinuous. Between the 1200 level and 2450 level, the lowest developed level, the main ore shoot has shown remarkable improvement on each succeeding lower level, and today the mine is one of the major lead-silver producers in the Coeur d&apos;Alene district. History: The Lucky Friday property is on the north side of the South Fork of the Coeur d&apos;Alene River in sections 25, 26 and 35, T. 48 N., R. 5 E., Hunter mining district, Shoshone County, Idaho. This is about one mile east of the town of Mullan, which serves the eastern portion of the Coeur d&apos;Alene district. The southern part of the property is crossed by a branch line of the Northern Pacific Ry. and by U. S. Highway 10. This highway is the principal road crossing the panhandle of Idaho and connects the district with Spokane, Wash., on the west and Missoula, Mont., on the east. The main portal and surface plant of the Lucky Friday mine is at an elevation of 3365 ft, only a short distance above the valley floor and a few hundred feet from U. S. Highway 10. so the mine is readily accessible for year-round operation. The property is comprised of six claims, known as the Lucky Friday group, owned outright by the Lucky Friday Silver-Lead Mines Co. There are four patented claims, Good Friday, Lucky Friday, Northern Light, and Lucky Friday Fraction No. 2 (Mineral Survey No. 3028), and two unpatented claims, Hunter and Creek. In addition, Lucky Friday owns an undivided one-half interest in the Hunter Creek property. which adjoins the Lucky Friday group on the north: a 90 pct interest in the mineral rights in the Jutila Ranch (160A), which adjoins the Lucky Friday group on the east; and a 60 pct interest in the Lucky Friday Extension claim group, which adjoins the Lucky Friday group on the west. The company also has a long-term mining lease on the Hunter Ranch, which adjoins the Lucky Friday group on the west. The claims of the Lucky Friday group were located between 1899 and 1906. The Lucky Friday Mines Co. was organized in 1906 and did considerable exploration work by surface trenching and shallow underground workings, only to see the property sold by the Shoshone county sheriff to satisfy labor claims totaling $2000 in 1912. Another firm, Lucky Friday Mining Co., bought the claims in 1914 and spent 12 years driving what is now known as the tunnel level crosscut. This tunnel intersected a vein previously exposed in a higher tunnel, but it was only a few inches wide. The vein was followed a short distance westerly but was so unpromising that the work was discontinued in favor of extending the main crosscut tunnel several hundred feet north. No ore was found and all work was discontinued. The property was held in such low esteem by the firm that it let taxes amounting to less than $15 a year go delinquent for nine years. Then the property lay idle for two more years until in 1938 John Sekulic, a Mullan service station operator, took a lease, with a $15,000 purchase option, on the advice of an old miner who had worked in the mine. Sekulic re-opened the tunnel level crosscut and explored the vein with an easterly drift for about 200 ft. The vein was too narrow to be of commercial value but was believed interesting enough to warrant further exploration at depth. Lacking funds to explore the vein at depth, Sekulic tried to get the district&apos;s larger operating companies to take over his lease and option. They were not interested because of the lean tunnel level showing and the fact that the property lay between the White Ledge fault on the north and Osburn fault on the south, an area which geologists always considered unworthy of exploration. Sekulic then organized the present company and assigned his lease and option to it for stock. This was in 1939. Enough stock was sold locally to finance sinking of a shaft 100 ft from the tunnel level east drift. The vein at this additional depth still was not commercial but showed some improvement. Treasury stock was offered at 5 to 10C a share

Jan 1, 1959

By T. A. O’Hara

SINCE May 1948, when tungsten carbide bits were introduced at the Flin Flon mine, they have been popular with the miners because of their fast drilling speed and low gage loss. The high cost of commercial carbide bits and tipped drill steel, however, prevented their use except for the hardest rock. In an effort to extend the use of tungsten carbide on a basis economically competitive with detachable steel bits, experimental work was begun in 1950 to test the feasibility of making tungsten carbide tipped drill steel in the mine drill steel shop. This work showed that tipped drill steel could be made locally at less than half the cost of the commercial product. The performance of the local tipped drill steel was comparable to that obtained with commercial carbide bits and tipped drill steel and the cost per foot drilled was much lower. Local tipped drill steel was adopted for all mine drilling in November 1951. Since then drilling costs per foot have been sharply reduced and footage drilled per manshift has increased markedly. Experience at Flin Flon has shown that production of satisfactory carbide tipped drill steel is not difficult and that highly skilled labor and costly equipment are not required. As long as wise selection of brazing materials is made and certain simple precautions are rigidly maintained, there is no reason why small mines with relatively unskilled labor cannot produce a satisfactory product. The following description outlines the technique used at Flin Flon for making carbide tipped drill steel and discusses characteristics of the brazing process that make special precautions necessary. Drill steel is forged to four-wing shape in a conventional steel sharpening forge. Standard steel dies are modified to minimize forging cracks around the central waterhole and to forge a blunt bithead on the steel. The steel is preheated to 1500°F and held at this temperature for at least 2 min. When the temperature has equalized throughout the steel section, the drill steel is transferred to the forging furnace and heated rapidly with a reducing flame up to 2000°F. This two-stage method of heating minimizes the grain growth and decarburization of the steel while ensuring that the steel temperature does not vary greatly throughout the forging zone. After forging the steel is allowed to cool in air to about 1600°F before being annealed in a bath of vermiculite. Despite the high hardenability of the 3 pct Ni-Cr-Mo drill steel used, this simple treatment anneals the drill steel sufficiently for milling. The forged and annealed drill steel is slotted on a plain horizontal milling machine that is equipped with a quick opening chuck and a slot depth stop. The full depth of the slot is milled in a single pass of the 3-in. milling cutter which is fed at 33/4 in. per min across the crown of each bit wing. The slots are cut to a width of 0.342 to 0.344 in. Maintenance of this slot width is necessary to ensure that the optimum brazing clearance of 0.002 in. will result after assembling of shims and carbide in the slot. Prior to March 1953, when the milling machine was installed, drill steel was slotted on a small manually fed ¾ hp milling attachment mounted on the bed of a lathe. Over 16,000 drill steels were slotted on this unit, and in view of its small size and low cost it gave excellent service. Brazing of Tipped Steel Drill steel that has been milled and cleaned in carbon tetrachloride is mounted in a rotating cradle holding six drill steels, the length of which may be from 2 to 12 ft. The slots in the drill steel, the shims, and the tungsten carbide inserts are thoroughly fluxed with a fluoride flux and assembled as shown in Fig. 1. Fig. 2 shows the brazing equipment in use. As the ring burner is lowered over the bithead a spring valve opens the gas lines, and the gas mixture, preset to give a slightly reducing flame, is fed to the ring burner where it is lit from a pilot flame. The ring burner heats the drill steel over a zone about 1 to 2 in. below the bithead, which becomes heated by conduction through the steel. By this means the bithead is heated rapidly and evenly, and contamination of the brazing joint with soot from the flame is avoided. The bithead is heated to the melting temperature of the brazing alloy within 1 min. This rapid heating minimizes the disadvantage of a non-eutectic brazing alloy. The brazing alloy, a nickel-bearing quaternary alloy, is placed at the bottom of the slot below the carbide insert, as shown in Fig. 1. As the brazing alloy melts it is drawn by displacement by the carbide and by capillary action into all parts of the joint to displace liquid flux from metal surfaces. As soon as the brazing alloy melts, each insert in turn is wiped by being moved back and forth along the slot. This action assists wetting of the carbide by the brazing alloy and assists in displacing molten flux from the joint. After continuous heating for about 75 sec, when the bithead has reached a temperature of about 1500°F, the ring burner is raised and the gas supply is shut off automatically by the spring valve. As soon as heating is stopped a hand press is placed on the bithead and the inserts are squeezed down firmly. This action minimizes the clearance between the bottom of the insert and the slot. Correctly brazed steel should maintain a clearance at the bottom of the slot of 0.001 to 0.002 in. After six steels have been brazed they are removed from the cradle and allowed to cool in air. As soon as each drill steel is cool it is dressed on a grinding wheel to remove excess flux and braze and is ground to the gage appropriate to the length of the drill steel.

Jan 1, 1955

By Harry C. Swanson

TRANSPORTING concrete from mixer to forms has always been a problem. Twenty-five years ago this task was generally accomplished by means of wheelbarrow or concrete buggy. On large dam jobs, as the number of these projects increased, the gantry crane or highline came into use. Today several methods of handling concrete are employed on smaller surface construction jobs, for example, transit-mix trucks or dumpcrete trucks, which have crawler cranes with buckets for placing concrete into forms. In 1944, during early stages of developing Mather mine A shaft, several large underground concrete jobs were necessary. At this time the Cleveland-Cliffs Iron Co, purchased the first pump-crete machine, introduced by the Chain Belt Co. of Milwaukee. The machine was used to pour approximately 200 cu yd of concrete for a dam, or bulkhead, located 400 ft from the shaft. Concrete was mixed on surface, lowered down the shaft 1000 ft in a 2-cu yd bucket hung under one skip, spouted into the bowl of the pumpcrete machine from the bucket, and pumped directly into the forms. Since the day of the first pipeline concrete in 1944 to the present time, other equipment and other methods have been developed to permit transportation of concrete by pipeline through vertical and horizontal distances totaling 1 mile from mixer to forms. Much of the efficiency in present handling of underground concrete can be credited to the Bethlehem Cornwall mines, where concrete was transported through 6-in. pipe for great distances down an inclined shaft and along levels into forms.&apos; During initial development of Mather mine B shaft, with concrete work under way on two or more levels at one time, the pneumatic concrete placer, Fig. 1, was selected as best adapted for underground concrete transportation. The 3/4-cu yd pneumatic placer is a small machine readily moved from one location in the mine to another. It can be equipped with two sets of mine car wheels, which will permit moving on regular mine tracks. It is therefore possible to send concrete through the pipe at great velocity; the pipeline is clean after each shot except for the film of cement adhering to the inside. With the proper slump in the concrete, it is possible to shoot concrete 2000 ft with this machine, using the mine supply of compressed air at 95 psi. This equipment was first used at Mather mine B shaft to concrete slusher drifts, Figs. 2 and 3, and finger raises located about 2000 ft from the shaft. In several instances there were bends into crosscuts and up vertical distances into the forms. For the first pours two placers were used. The first was located near the shaft where the concrete could be spouted into it from a 2-cu yd concrete bucket on the cage. The second was set on the side of the drift at a point approximately 1500 ft from the shaft. The concrete was shot directly into the second placer from the first unit and from the second machine directly into the forms. After completion of several pours with the two machines, a trial pour with only one placer located at the shaft proved that the second placer could be eliminated. Since then all pours have been successfully completed with only one placer underground. As production of iron ore from the mine increased and the development program expanded, use of the cage for handling mine supplies and concrete became a major problem. This brought about the first attempt at shooting concrete vertically down the shaft for 2600 ft. A 6-in. pipeline with victaulic couplings installed during shaft sinking was used for the trial. One placer was set on surface 250 ft from the collar of the shaft so concrete could be spouted directly into it from the mixer. This machine shot the concrete 250 ft horizontally on surface to the shaft, 2600 ft vertically down the shaft, and 100 ft horizontally into the second placer located near the rib of the shaft station or plat. The second machine shot the batch into the forms, about 2000 ft. Total distance horizontally and vertically was 4800 ft. The entire time cycle for a ¾-cu yd batch of concrete from the mixer on surface to the forms underground totaled about 5 min. During the past two years the two-placer method from the mixer on surface to the forms underground has proved a very efficient means of transporting underground concrete. Advantages of using pipeline concrete are as follows: 1—Interference with normal mining operation is eliminated. When the cage, skips, mine cars, or mine openings are used for transporting concrete and materials used for making concrete, mine operation suffers in one way or another.

Jan 1, 1955

By N. A. Elmslie

This subject covers much ground, therefore it must be treated in a general way rather than in detail in this paper. Personnel To approach the measure of a mine, it is, of course, essential that the personnel be measured, and this measuring must be done in a circumspect manner and conclusions must be reached slowly. There may be one unit in the personnel of an organization that may be out of step with the remainder of the organization, not always because of lack of ability, but because of misplacement or lack of proper measurement of his capabilities. Too much time can scarcely be spent in the study or measure of the personnel ; in fact, a manager or a superintendent of a mine never completes his work along this line. Mine or Plant Just as no two men are alike, so no two mines or plants are alike; and yet there are many similarities. Find out what the similarities are, and this will help to make clear the dissimilarities. Try to visualize details of the difficulties, not because you will have to deal with them yourself, but to consider them in later studies of general problems. To outline all the steps to be taken in familiarizing oneself with a plant or operation would be impossible, but the following points may be of interest in any investigation. The study of a mine is generally made backwards-—that is to say, one sees the finished product first-—-but this is immaterial, because each step must be considered separately in any event. As a rule, there are certain heads under which the different features of mining can be placed, as follows: Ventilation.—Ventilation is usually the most important item in power cost and safety, especially in gassy mines. Quantity of air, water gage and power consumption, and finally analysis of the return air in full return and separate splits should be considered. Type of fan and type of power unit should enter into this study because it is important to view expenditures which are going on 24 hr. a day and 36.5 days in the

Jan 1, 1931

By Eugene McAuliffe

The term "mechanical mining" carries an ambiguity which justifies a preliminary word of explanation. . All mining activity conducted in this day is more or less mechanical; that is to say, power expressed in the form of electricity, steam or compressed air is used in some portion of practically every mining activity of consequence. "Mechanical loading" is more nearly correct than "mechanical mining" in reference to present practice. The first substantial effort made in the direction of mechanical mining, subsequent to the application of steam for hoisting and underground pumping purposes, was that of substituting for hand labor undercutting and shearing by power-driven machines. The undercutting of coal by hand represented at one time the most arduous task confronting the miner. Occupying a kneeling position in the thicker beds, or lying on his side in the thinner coal, the task was tedious and exhausting. The first attempt to substitute crude air-driven, and later electrically driven, "punchers," as the percussion type of machine was called, met with much the same opposition on the part of the workers as was expressed by workers in the other industries when any innovation, whether labor-saving or otherwise, was introduced. We should not be too critical of this attitude of mind shown by the mine workers. The peculiarly isolated character of the work carried on in semidarkness by a class of men who to a large extent lived and held themselves aloof from the other branches of society led long ago to the development of a craft consciousness which they have found difficult to surrender. It was not until 1891 that a definite record of the results obtained by the use of undercutting machines of all types was made available, the quantity undercut in that year totaling 6,211,732 tons of bituminous coal from all American mines. The volume of bituminous coal undercut by machines in 1928 totaled 369,687,007 tons, or 73.8 per cent. of the nation&apos;s production. In the light of this experience, we can properly feel that the work of introducing the coal-loading machine as a substitute for hand shoveling is making substantial progress.

Jan 1, 1931

By L. Adler

Longwall caving, one of the most economical and attractive mining methods, is yet one of the most difficult and hazardous.1 This dualism is inherent in a method which manipulates the mine supports themselves to aid in excavation. The difficulty is further compounded by the lack of an analytical method to predict and control results. This paper is intended to meet this need by providing a first approximation to the mechanics involved. The analysis will be limited to materials which are linear-elastic, homogeneous and isotropic. The rock structure is assumed to consist of beds of uniform thickness with no interbed bonding.2 Consequently, the mine roof can be likened to a beam, in which the action characteristic of longwall caving is the controlled deflection of the barrier supports. This deflection causes load to be intentionally transferred from these barrier props to the working face.3 Two of the requirements of this method, shared in common with other mining methods, are the provision of (1) adequate and (2) safe working space. The third requirement is the one characteristic of longwall caving and is somewhat in conflict with the second; (3) the development of near-failure static loads at the face. Requirement (1) is handled by spacing the barrier props sufficiently far from the working face for efficient operation. Requirement (2) limits the static stresses in the structure to predetermined allowable values. This means that it is not desirable to induce a maximum load on the working face as requirement (3) may imply. Such a condition could produce collapse in the whole structure. It is the purpose of the undercut to limit the depth, and thereby control the failure of the face. Therefore, requirement (3) really means the development of an optimum, not maximum, load which, in conjunction with undercutting, produces failure at the working face, with minimum powder consumption. Stress analysis requires an exact delineation of the structure and loads. The structure will be defined as the immediate roof and its supports. Since the roof is severed at the caving line, the supports reduce to the barrier props and the working face, as shown in Fig. 1. The mechanics of failure within the working face itself are complex and beyond the scope of this study, since the face represents a typical mine invert structure.4 However, this lack can be made good by the judicious collection of field data. The primary load on the structure is its own weight or body load which can be considered uniformly distributed over its upper surface. Additional loading occurs due to interference of overbeds with the structure. Wherever this occurs, the effect can be approximated by a load uniformly distributed on the upper surface of the structure.5 When this interference is intermittent, as with a much thicker overbed, the elastic curve for the overbed is superimposed on that for the structure and the extent of interference noted, as Fig. 2 shows. However, for simplicity, this study will ignore such intermittent loads without impairing the generality of approach. Consequently, the free-body diagram is that of a uniformly loaded cantilever beam with an inter-mediate support, as shown in Fig. 3. For convenience of analysis, the diagram will be divided just to the

Jan 1, 1961

By D. H. Trollope

In engineering in general, close agreement between theoretical predictions and structural performance is rare—this is particularly true in rock slopes. Since the complexity of natural arrangements makes exact theoretical solutions of associated structural problems almost impossible, the role of theory should be that of developing guides or markers to which observations and experience can be related. The author in this paper has outlined the results of a theoretical study of highly idealized models which represent both granular materials and jointed rock. Some important practical considerations result—especially in regard to horizontal stresses—when theory is compared with full-scale experience. In the field of engineering activity, specific examples of close agreement between theoretical prediction and structural performance are rare. Particularly is this so for the case of rock slopes, where, owing to the complexity of the natural arrangements, it is quixotic to expect exact theoretical solutions to the associated structural problems. The role of theory is rather that of developing signposts or guide rails to which observations and experience can be related. It is with this in mind that the present paper has been prepared; the author&apos;s aim has been to set out briefly the results of a theoretical study of highly idealized models which represent both granular materials and jointed rock. Despite the idealization of the problem, it will be seen that the theoretical conclusions are in accord with full-scale experience, particularly as far as horizontal stresses are concerned, and they point to some important practical considerations. CLASSIFICATION OF ROCK SLOPES At the risk of over-simplification, an attempt has been made to classify the problem of rock slopes into three broad categories with respect to internal stress systems. These are illustrated in Fig. 1. It is tacitly assumed in this classification that the strength of individual particles is very much greater than the interparticle strength whether the particles are the crystals of an igneous rock (Case I) or the blocks of Cases 2 and 3. The Homogeneous Monolith: This case is representative of uniform unjointed rock. Under relatively low stress conditions the internal stress system may approximate to that of an elastic body as given by Terzaghi.&apos; It is more usual, however, for overall stability, to consider the material as a rigid plastic which yields in shear when S =C + c tan 4> where S is the shear yield stress, o is the normal stress on the surface of failure, and C and d are empirical constants according to the Mohr-Coulomb criterion of failure. Sokolovsky has recently developed, for the first time, a series of mathematical solutions to problems involvinga C, mass subject to self-weight. He demonstrates that for such a case a semi-arch (Fig. 2) is stable. A corollary of this argument is that for a true C, d material a curved overhanging slope is possible and such an observation may well aid field identification of these materials. The shape of the curved slope face is however directly dependent on the C-value, and from the long-term point of view this is a most difficult value to establish with any degree of confidence. It is customary to measure C and 0 in some form of shear test—usually a triaxial compression test—on core specimens sampled from the rock, and as far as the author is aware the present case is the only one in which, as far as overall stability is concerned, such tests are

Jan 1, 1961

By George A. Morrison

A deposit Of limestone Was known to exist at a depth of 2000 ft under the property of the Columbia Chemical Division of the Pittsburgh Plate Glass Co. at Barber-ton, Ohio. A 662-acre site was selected for the projected mine, some two miles northwest of the plant and near the main line of the Erie Railway. An electric railway 8800 ft in length connects the mine and the plant. The deposit is overlain by 18 ft of un-consolidated (sand and gravel), 100 ft of water-bearing sandstone, and finally 2080 ft of shale into which thin layers of sandstone are embedded. The limestone bed itself is 345 ft thick and overlies 200 ft of gypsum. A 300-ton-per-hour capacity mine was decided upon. Two shafts were deemed advisable, one for production and the second for service; and these were sunk 550 ft apart, one to a depth of 2323 ft and the other to 2258 ft. Each is 16 X , ft in the clear and timbered with 6-in. steel H-sets on 8-ft centers, using 3/8 X 4 X 4-in. angles as studdles. Twelve inch I-beams were used at 160 ft intervals as bearers set in deep hitches and concreted to a vertical depth of 8 ft. Wooden sheathing was first employed but sloughing of the shale created a pressure on this lining and finally concrete was deemed necessary. Water was encountered only within the first 315 ft, flowing at 60 gpm and by grouting this was reduced to 15 gpm. Temporary wooden sinking headframes were installed with 300 cu ft capacity storage pockets and with doors automatically controlled by air cylinders by hoistmen. Each shaft was operated on four 6-hr shifts. The usual V-cut was drilled at the center of the shaft and the round consisted of 34 holes- A 40 pet gelatin dynamite was employed with some 60 pet in the cut holes. Drilling crews consisted of four to six men. Shale drilled easily but became more bed in depth and sandstone wore away the gauge Of the bits rapidly. A room-and-pillar mining system was adopted whereby rooms turned off on 80-ft centers, each room 40-ft wide and 50-ft high. Rib pillars between rooms were cut into 100 ft lengths by frequent holes. Ventilation is achieved through the two shafts, the down-draft through the service shaft, and the return through the pr0duction shaft. The hoist has a a Of 300 tons per hour and is equipped with 10-ton balanced skips requiring. a 2-min cycle Or 30 trips Per hour- Hoisting is designed to Operate by automatic loading and no hoistman is required for normal Operation. The 48 X 60 in. jaw primary crusher is underground. It is fed through chute and chain feeders and is flood-oil lubricated and cooled. An inclined belt conveyor returns oversized material to the crusher. Tramming is by Diesel-powered trucks.

Jan 1, 1948

By George A. Morrison

A deposit Of limestone Was known to exist at a depth of 2000 ft under the property of the Columbia Chemical Division of the Pittsburgh Plate Glass Co. at Barber-ton, Ohio. A 662-acre site was selected for the projected mine, some two miles northwest of the plant and near the main line of the Erie Railway. An electric railway 8800 ft in length connects the mine and the plant. The deposit is overlain by 18 ft of un-consolidated (sand and gravel), 100 ft of water-bearing sandstone, and finally 2080 ft of shale into which thin layers of sandstone are embedded. The limestone bed itself is 345 ft thick and overlies 200 ft of gypsum. A 300-ton-per-hour capacity mine was decided upon. Two shafts were deemed advisable, one for production and the second for service; and these were sunk 550 ft apart, one to a depth of 2323 ft and the other to 2258 ft. Each is 16 X , ft in the clear and timbered with 6-in. steel H-sets on 8-ft centers, using 3/8 X 4 X 4-in. angles as studdles. Twelve inch I-beams were used at 160 ft intervals as bearers set in deep hitches and concreted to a vertical depth of 8 ft. Wooden sheathing was first employed but sloughing of the shale created a pressure on this lining and finally concrete was deemed necessary. Water was encountered only within the first 315 ft, flowing at 60 gpm and by grouting this was reduced to 15 gpm. Temporary wooden sinking headframes were installed with 300 cu ft capacity storage pockets and with doors automatically controlled by air cylinders by hoistmen. Each shaft was operated on four 6-hr shifts. The usual V-cut was drilled at the center of the shaft and the round consisted of 34 holes- A 40 pet gelatin dynamite was employed with some 60 pet in the cut holes. Drilling crews consisted of four to six men. Shale drilled easily but became more bed in depth and sandstone wore away the gauge Of the bits rapidly. A room-and-pillar mining system was adopted whereby rooms turned off on 80-ft centers, each room 40-ft wide and 50-ft high. Rib pillars between rooms were cut into 100 ft lengths by frequent holes. Ventilation is achieved through the two shafts, the down-draft through the service shaft, and the return through the pr0duction shaft. The hoist has a a Of 300 tons per hour and is equipped with 10-ton balanced skips requiring. a 2-min cycle Or 30 trips Per hour- Hoisting is designed to Operate by automatic loading and no hoistman is required for normal Operation. The 48 X 60 in. jaw primary crusher is underground. It is fed through chute and chain feeders and is flood-oil lubricated and cooled. An inclined belt conveyor returns oversized material to the crusher. Tramming is by Diesel-powered trucks.

Jan 1, 1948