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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 E. H. Bronson

PROPERTY of Malartic Gold Fields Ltd. is situa-ated in the great clay belt in the northern part of the Province of Quebec. This belt represents the floor of the glacial lake Barlow-Ojibway. The clay blanket has a gently rolling, almost flat surface, with here and there projecting knobs and ridges that outline the large irregularities of the surface on which the clay was deposited. At irregular intervals glacial eskers dominate the landscape. The ore-bearing country rock is covered by strata of clay, sand, and silt to a depth varying to 40 ft. There are serious structural problems when it is necessary to place loads of unequal intensities on this surface to support buildings and machinery for a modern mining plant. Earlier mining companies and their engineers paid little attention to the science of soil mechanics, partly because they believed it unnecessary and partly because they knew little of the subject. The results are evident in some of the tottering structures in older camps. Careful analysis of soil problems is well worth-while, since cost of foundations may vary from 20 to 50 pct of the final cost of a structure. Low operating and maintenance costs can be achieved only in this way. If it is definitely established that a plant will be in operation for only a few months, this limited time factor will probably influence the design of the foundation, and rightly so. If sensitive and complickted equipment is to be mounted on these foundations and the plant is to operate for several years, there should be a thorough study of foundations and a very careful design. A familiar problem is presented by the difficulty of operating a permanent plant originally designed for a few months' work. The engineer must beware of these hazards and disassociate himself from them if he cannot prevail on the promoter to spend the proper amount for a thorough foundation study. Laws of certain provinces and states hold the engineer equally responsible with the contractor for any 'damage that may be caused to building through inadequate study-ignordnce of the facts is no excuse in the eyes of the law. Types of Soil: The engineer should be familiar with such terms as silt, clay, peat, and varved clay, as well as perpafrost, especially if he is to work in the far northern reaches of the country. He should also be familiar with soil attributes such as porosity, texture, consistency, permeability, compressibility, consolidation, and shearing resistance. The design of any foundation requires an accurate knowledge of the physical properties and location of the soils involved. Older textbooks on the subject outline various empirical methods to determine the carrying power of soil. Tables are given in reference books. The modern ehgineer should be wary of these modes of thinking. He should analyze his problem carefully, calculate to what depth of soil the loads may be transmitted, and then check the bearing value of all soils to that or greater depths. If the strata are water-carrying, he must determine the possibility of this water being accidentally re-

Jan 1, 1958

By W. A. Vine

WHEN a hole is made in a stressed solid, such as rock pierced by mine openings, equilibrium of the solid is destroyed. To re-establish that equilibrium the stress condition in the rock surrounding the opening becomes rearranged. This rearrangement may or may not develop local stresses in the vicinity of the opening which exceed the elastic strength of the rock. Since stability of the opening depends, therefore, on the strength of the naturally occurring rock surrounding the opening, this rock is an engineering material of construction, and all rock affected by the associated opening or system of open- ings may be considered a structure. The term invert is chosen to describe the formation of the structure —built from the inside, as it were, and a consequence of mining rather than an entity built for a purpose. By definition, an invert structure is formed when openings are made underground and is composed of that material on the solid side of a rock-air interface wherein the stress condition is affected by the associated opening or system of openings. The shape of an invert structure is never clearly defined; the skin of the structure may be mapped, and even described, but the boundary of the structure within the solid rock mass must be arbitrarily set. The name geomechanics is proposed to describe a branch of scientific knowledge which will encompass principles and working hypotheses for behavior and design of invert structures. It is foreseen that study

Jan 1, 1956

By Robert E. Thurmond

DISCOVERY of a commercial ore deposit has resulted from a program emphasizing geophysical techniques and concentrating upon exploration of alluvial-covered areas. United Geophysical Co. selected these areas as offering the best possibility of success, for here the surface evidence of ore deposits has been hidden from the eyes of prospector and geologist alike. Pima mining district, Pima County, Arizona, was selected for initial exploration. The project was successful in that a commercial ore deposit was discovered and proved through drilling and extensive underground development. A limited region of activity was chosen and library research was begun on available economic and regional geological publications relating to the mining districts. Field reconnaissance of those districts which seemed to hold exploration possibilities provided a general picture and ascertained the existence of alluvial-covered areas of interest. At the conclusion of the reconnaissance trips all data obtained were compiled, and the district which seemed to present the best exploration possibilities was selected by statistical methods. The statistical summary was based on the following considerations: Geological Factors: 1—type of intrusive, P—presence of favorable host rocks, 3—apparent intensity of mineralization, and 4—presence of favorable structure in covered or unexplored areas. Economic Factors: 1—past production from district or mine; 2—size, shape, grade, and type of ore deposit in the district; 3—ease of acquiring mineral rights; and 4—general development facility. Exploration Favorability Factors: 1—results to be anticipated from geological methods alone; 2—application of geophysical methods, namely, topographic conditions, surface interference, depth-size relation, physical property contrasts as related to ores or controlling structural features; and 3—drilling factors. Admittedly, many of these, factors do not lend themselves to absolute evaluation, but relative evaluation, subject to inaccuracies of personal judgment, is always possible. A relative scale from 1 to 10 was established for every factor, representing varying degrees of favorability. For each district a numerical evaluation of each factor was established and a numerical total determined. The obvious advantage of this scheme is that a more concise summation can be presented by numerical means than can be achieved by descriptive means. After a district had been selected as favorable, the first step was to carry out more detailed geological and economic studies to narrow down the areas of interest. The second step was to compile all available geological data from publications and add to these data by mapping where necessary. Geological interpretation and projection of structural or other control into covered or unexplored areas was made whenever logical. To determine the possibilities of geophysical application and to select a geophysical method, physical property contrasts were related to mineral assemblages and/or structural control. If geophysics seemed applicable, trial profiles were run crossing the area normal to the strike and spaced

Jan 1, 1955

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 R. C. Mahon

THE importance of ground water control in glacial drift overlying mines is widely recognized. Adequate handling of the problem results in considerable saving in overall pumping costs, as the cost of pumping the larger quantity of water under low heads from surface wells is more than offset by the decrease in the quantity of water to be pumped from the mine under high heads. Eliminating or partly reducing seepage of water through the ledge reduces moisture content of the ore, a factor especially important in iron ore mines, where a reduction of 1 pct in moisture effects a very material saving. An important though indeterminate saving is also brought about by the elimination of water in the working places underground. The Homer and Wauseca mines of Hanna Coal & Iron Co. at Iron River, Mich., cover 480 acres. When these mines were started it was possible to operate on only one 40-acre tract, as the depth of water over the remainder of the property was too great to permit operation. Although the average depth of the ground water over the area being mined at that time was less than 50 ft it was necessary to pump 1200 gpm from the mine, the greater part from a depth of 750 ft. The slate footwall and graywacke hanging wall are almost impervious, and it was apparent that seepage into the mine was mainly through the iron formation in which the orebodies were located. Analysis of the water in the iron formation showed it to be of the same character as that in the glacial drift, while the water in both walls was heavily mineralized, also indicating that the water passed much more freely through the iron formation. The first attempt to take off the water above the ledge was by means of drainage drifts driven 15 to 20 ft below the ledge, Fig. 1. Various schemes were used to get the water into these drifts: 1—Holes 15 to 20 ft long were drilled from the back of the drifts to the ledge, but these were difficult to drill and blocked easily. 2—Sections of the drifts were blasted down. This gave a large flow for a time but eventually almost total blockup occurred. 3—Raises 4x4 ft were driven above the drifts to within 8 ft of the

Jan 1, 1955

By E. P. Pfleider, O. Rellensmann

For many years mine surveyors and exploration engineers have sought an accurate means of transferring meridian underground by using the gyro-compass. These efforts have generally failed, either because they were inaccurate or because the size and cost of the unit were prohibitive. During the past ten years, however, great strides have been made. Autonetics in California has developed a comparatively light-weight (100-lb) gyro-theodolite, and the Institute for Mine Surveying at Clausthal Mining Academy in Germany has developed an instrument that is accurate to within 20" to 30" of arc. Concurrently, Minneapolis-Honeywell has produced an astounding gyro, the size of a salt shaker, that weighs less than 0.5 lb and compares favorably in performance with the 50-lb gyros of conventional design. This unit, designed for missile guidance, should be adaptable for surveying small boreholes. Given the proper interest and approach from the engineering fraternity, excellent gyrocompasses could probably be available for accurate underground survey work at nominal cost within the next three years. Principles of the Gyro-Compass: The gyrocompass, used for 50 years as a navigational instrument in ships and aircraft, senses the turning axis of the earth and points out true polar north. It does this by employing the earth's rotation, the force of gravity, and the precession characteristics of a gyro. For further details the reader is referred to several excellent publications'"3 covering the principles and application of the gyroscope. The heart of the gyro-compass is the gyroscope, or rapidly spinning wheel, which rigidly maintains its angular orientation in inertial space. Free gyros employ universal joints to allow the wheel axle to maintain its rigid spatial orientation independent of base motion disturbances. Free gyros are independent of the earth's turning and are not north-seeking. Only as they might be initially oriented parallel to the earth's turning axis could any north indication be obtained. In practice, the free gyro tends to drift off its set position and lose its accuracy. The Sperry-Sun Well Surveying Co. instrument (5M in O. D.) is of this free gyro type. It is counterbalanced, however, against the rotational effects of the earth, so that orientation of the spinning axis will be maintained throughout the survey.

Jan 1, 1960

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 Jules E. Jenkins

IN the past quarter century the seismograph has played an increasingly important role in evaluating vibratory effects transmitted to adjacent communities by industrial blasting operations. In this period research has advanced through two major phases—from falling pin seismometers to three-component recording seismographs. The initial phase of this research is now regarded as the era of recording, during which the seismograph merely recorded occasional events. It has emerged from this stage into the era of control and analysis and in this later period has been recognized as a full-fledged industrial tool. Although the falling pin machines of earlier years indicated probable amplitudes of industrial blasting vibrations, they gave no reliable measure of such displacements, nor did they provide essential information concerning direction or frequency of motion. The first contribution of the three-component recording seismograph was the weight-for-distance formula developed by the U. S. Bureau of Mines. This empirical formula enabled users of commercial explosives to determine beforehand the probable vibratory effects of a blast when the weight of the explosives charge and the distance from the shot point were known. This formula has withstood the test of time and is still applicable for blasts detonated by instantaneous methods. The second major contribution of the three-component recording seismograph followed the introduction in 1945 of short period or millisecond delay detonation. This might be described as a two-stage contribution, the first of which revealed a 50 pct reduction in the vibration magnitude when millisecond detonation was used rather than instantaneous discharge. This second part of this double contribution led to the discovery of the vibration dip phenomena, see Fig. 1. In Fig. 1 the horizontal coordinates indicate the displacements amplitude in inches, and the vertical coordinates, the explosive weight in pounds. The top curve is that of the U. S. Bureau of Mines for instantaneous detonation. Its curvature conforms to the two-thirds power of the explosive weight as provided in the weight-for-distance formula; hence the vibratory effects increase in proportion to an increase in the explosive charge. If the other curves shown below are temporarily disregarded, it may be considered that the entire area below this top curve represents both useful and waste energy. As mentioned previously, when millisecond delay detonation was first investigated it was found that in general the U. S. Bureau of Mines curve could be decreased by about 50 pct, or to the position of the middle curve. At the outset it was assumed that this curve would likewise increase in some proportion to the explosive charge, as did the top curve, that is, the curve would indicate a continuous upswing. As seismographic data accumulated this assumption was found to be erroneous in part. It was discovered that

Jan 1, 1957

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 L. A. Panek

During the past three years, USBM has developed an apparatus and technique for direct measurement of existing pressure and change of pressure in mine rock. This relatively simple and inexpensive monitor is reliable for months after being installed. It is used to determine shift of ground pressure created by different sequences of mining, to ascertain the cause of rock failures, and to evaluate the need for artificial support. The technique has been employed to measure pressures in limestone, greywacke, concrete, diabase, and soft iron ore. When rock is subjected to a load it is deformed. Ordinarily this is observed in a mine as the displacement of one point with respect to another—the deflection of the roof, which may be observed as a convergence between roof and floor; or the extrusion of material from the rib, which may be observed as a decrease of the distance between the rib and the post of a timber set. The effect of excessive pressure may be a rockburst if the rock is strong, or it may be squeezing ground if the rock is soft. Some desirable effects of high stress (high in relation to strength) are the caving of roof in a longwall mining operation, the caving of ore in block caving, and the decrease in mechanical energy required to break down the mineral seam in a retreating pillar-robbing operation. In any case, whether the observable effect of rock load is desirable or undesirable, it is a displacement, and depends on the following four factors: 1) The structure—the size and shape of openings, pillars, and linings, the geologic bedding and jointing. 2) The mechanical properties of the rock—prin-cipally the strength, modulus of elasticity, and flow characteristics. 3) The load or applied stress—primary sources are the weight of superincumbent rock, which increases with depth, and unrelieved tectonic stresses; secondary sources are redistributed pressures caused by other nearby openings, especially large mined out zones (rock pressure depends partly on the rock structure created by mining). 4) Duration of load, related to the length of time the opening is exposed. CONTROL OF ROCK DISPLACEMENT Rock displacement can be controlled by control of these four factors. Consider now the means of exercising such control over these factors. Control of the structural features is obviously possible to a great extent, as such control is exercised largely by choosing the method of mining and the methods of natural and artificial support. Rock properties vary, even within a particular mine, but they are controllable only in the limited sense that control may be exercised by choosing the beds or zones to be mined so that rocks with undesirable properties will not occupy critical positions within the rock structure created by mining. Rock pressure is the most complex of the four factors through which ground control can be achieved because it is invisible, difficult to measure, and poorly understood. Rock pressure is controllable only to the extent that control is exercised on the rock structure created by mining. Considering openings within a particular mine, time of exposure varies, and is readily controllable because it is easily measured and easily understood — the longer an opening stands, the greater the likelihood of failure or excessive convergence. Control is exercised by choice of an appropriate sequence of driving openings of different classes, such as haul-ageways and rooms, which are required to remain well supported for different lengths of time under different conditions. Again, control is exercised through the method of mining. All controllable factors can be controlled by proper design of the mining method. The orientation and relative positions of the mine workings and the sequence of their excavation are likely to be much more important to ground control than is the design of artificial support. This implies that the major decisions in regard to ground control are made, knowingly or not, at the time the mining method is chosen. WHY MEASURE ROCK PRESSURE In addition to restrictions on the several factors, control implies the measurement of these factors in some sense, whether only qualitatively by visual observation, or by actual quantitative determination with a measuring instrument. Rock pressure is the most difficult of these factors to measure, largely because of the interaction between the measuring device and the rock. Nevertheless, the quantitative

Jan 1, 1961

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