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By S. H. Ash

More than 100 minerals are mined and processed in the U. S. Management and labor have negotiated wage-scale agreements that have balanced wages and affected cost of labor in such a manner that comparisons can be made of transportation costs. Transportation of materials underground has become of vital importance. This report describes underground haulage methods and equipment, presents cost analyses, and discusses future plans for underground transportation in metal mines. It does not cover hoisting in shafts or slopes by cage, skip, or cars. Advancements have been made by the metal mining industry in lowering transportation costs in several mines, making it possible to mine orebodies or portions of orebodies not considered economically recoverable a few years ago. Numerous reports have been written on transportation at metal mines; this report covers recent haulage at some representative mines throughout the U.S. The term metal mine is used to include all noncoal mines from which up-to-date information was obtainable.

Sep 1, 1956

By J. Wayne Erickson

Hundreds of millions of dollars have been spent to discover or purchase uranium resources that cannot, or should not, be mined with conventional methods if human, natural and financial resources are to be conserved. Three years ago, an industry-supported research stud was initiated to determine whether a safer and more efficient (in terms of cost and environment) uranium mining method could be designed to replace existing techniques. As a result of this study, the "underground in situ" method was conceived and developed to the stage where it is now feasible. This approach involves: (1) sinking a shaft; (2) driving a drift on the ore horizon the length of the ore body; (3) drilling almost-horizontal long holes the width of the ore body; and (4) oxidizing and leaching the uranium drawn from these holes. The method is unique in that the uranium is removed with a minimum of disturbance to the natural hydrology of host or surrounding sands.

Jan 11, 1978

By O. N. Wampler

UNDERGROUND lighting in the zinc mines of the Tri-State district can be separated into four di-visions. 1. Foot of shaft. 2. Pumps, fans. incline hoists and other stationary machinery. 3. Haulage ways. 4. Headings and stopes. One of the main improvements in lighting around the bottom of the shaft is in the use of skidoo flash lights, which have replaced the old electric skidoo bells. These bells were used as signals from the hoisterman to the tub hooker. In the event that a can bumped the side of shaft or a rock spilled from the can when dumping; the hoisterman rang the skidoo bells as a signal for all to leave the vicinity of the shaft bottom on a dead run. These bells were much better than no signal at all, but many of them were often out of order; the worst feature being that when out of order this fact was not immediately indicated. The electric flash-light signals work on the opposite principle and are burning at all times except when the hoisterman breaks the circuit when signaling danger. Therefore should a break or short occur in this arrangement, the shaft bottom is immediately thrown in darkness. Since the reaction from the eye is quicker than from ear, there is a slight saving of time- to the men escaping from the point of danger. Even this brief period of time may save a life. One other, point, to be emphasized, is that all lights near the shaft bottom should be on the signal circuit so that the entire area is momentarily darkened. I. have inspected mines where only one or two lights are used as signals, with the remaining lights around the shaft on another fixed circuit. In this case the flashing of the signal lights is almost drowned out by the stationary lights. These flash signals must be as startling as possible, which can only be done by flash-ing all the lights.

Jan 6, 1928

By E. D. Gardner

The tide of increasing demand for metals turned during early 1949, with resultant lower prices. The production of the principal metals was less than in 1948. The problem of high unit costs, however, was held over from 1948. With the lower prices for metals, greater efforts are being made to reduce operating costs. The labor supply was adequate during the latter half of the year in most districts. The force has been reduced at most mines; some mines were idle during part of the year because of strikes, and others ceased operations on their own account. In the main, labor efficiency increased during the year in underground metal mines. Efforts are being made generally to improve labor relations. Green men hired in the time of the scarcity of skilled miners are becoming better trained. The supervisory personnel of most mines is cost conscious, and the planning of operations has been improved. Overhead costs are being scrutinized. The trend toward mechanization underground continued strong. Mining technology rapidly is becoming a problem of handling materials. Surface miners, underground miners and miners of metals, nonmetallics, and coal can learn much Gardner from each other. A mine, however, may become overmechanized. The capital charges against new equipment may exceed the saving in labor costs where sufficient daily tonnage is not handled.

Jan 1, 1950

By John Dilworth

THE word "development," as used in connection with .mining, is a rather general term and in most instances must be qualified or explained before the exact thought in the mind of the user is made clear. For instance, by the development of a mining district is often meant its entire facilities of mineral production, such as railroads, towns, mine plants, etc. In speaking of the development of a property its prospecting by test pits or diamond drill borings may be referred to; or it may be the mines, strip pits or quarries from which the coal, ore or rock is won. But, as a rule, when the development of an individual mine is spoken of the idea in mind is the system of underground workings by means of which the mineral is recovered and transported to the surface. It is this character f development that is considered in this paper. The discussion will be confined to the ordinary room-and-pillar, system of mining. practically horizontal coal seams that is used in most bituminous coal mines in this country, and to only two phases of this subject; viz., what is properly included under the head of "Development"' as it appears on the books of a coal company or the inventories of an examining engineer, and how its "value" should be determined -by the accountant or engineer. No absolute rule or definite agreement covers these points, yet these questions are far more than academic. Development is often an important part of a mining company's assets, affecting the amount of its invested capital (and hence its rate of taxation under present Federal income tax laws) as well as the value of its total assets and the amounts chargeable to cost of coal and to depreciation.

Jan 7, 1921

WITHOUT in any way detracting from the credit due those engineer-miners of copper who operate with power shovels, it may be said that compared with block-caving underground their work is simplicity itself. True enough, ingenuity and foresight of high order are required to lay out an open-cut operation to the best advantage; good organization and intelligent care and use of machinery are required to achieve minimum costs. But really successful undercut block-caving demands all this; and, in addition, it requires a conception and application of unseen natural forces that reflect a type of genius that is rare. Assuming that preparatory work has gained access to the orebody, mining of whatever kind consists essentially of three operations: 1. Breaking the ore "in place," as Nature left it, into pieces small enough for convenient handling. 2. Loading these pieces into some kind of a vehicle on wheels. 3. Transporting the broken ore to some central plant on the surface where the process of extracting the valuable metals can be commenced. In shovel mining, dynamite and steam or electric power are used-with remarkable efficiency, be it said-to accomplish each of these three processes. In underground-caving methods the old force of gravity is called upon to accomplish, in large part, processes 1 and 2. To do this with success is an artistic achievement of a sort. Put it another way. The earlier methods of underground mining constituted a translation of surface procedure to an underground theater. The ore was blasted and dug out in small quantities; in its place were put props, usually of timber, to prevent the caving of the rock above; and men toiled with

Jan 1, 1933

By George S. Rice

The subject of violent bumps in coal mines has been again brought to attention by a recent succession of such occurrences in the coal mines of the Cumberland field of eastern Kentucky and southern Virginia, which have caused fatalities and injuries to underground employees, a report on which is being issued by the United States Bureau of Mines1. The A.I.M.E. has had two valuable papers on bumps in the No. 2 Mine, Springhill, Nova Scotia, each followed by general discussion on bumps and their causes2. This paper gives a brief review of suggested causes and remedial measures given in English, Canadian and United States reports on blimps in coal mines, and then presents suggestions arising from the writer's recent investigation of the bumps in the Cumberland mines, made at the request of the mine operators, to determine the causes and remedies. These bumps, which occur under the simplest natural conditions, provide better opportunity for analysis of causes and effects than afforded under more complex conditions. The simple conditions in this field consist of relatively thin and level coal beds and uniform roof strata free from folding and faulting. Definition of Bumps Bumps in coal mines, as defined in this paper, are sudden violent bursts of coal that usually occur without warning in the extraction of pillars and may or may not be accompanied incidentally by falls of draw slate, upheaval of bottom and smashing of timbers. They are not known to occur where the advancing longwall method of mining, accompanied by

Jan 1, 1936

By George S. Rice

The subject of violent bumps in coal mines has been again brought to attention by a recent succession of such occurrences in the coal mines of the Cumberland field of eastern Kentucky and southern Virginia, which have caused fatalities and injuries to underground employees, a report on which is being issued by the United States Bureau of Mines1. The A.I.M.E. has had two valuable papers on bumps in the No. 2 Mine, Springhill, Nova Scotia, each followed by general discussion on bumps and their causes2. This paper gives a brief review of suggested causes and remedial measures given in English, Canadian and United States reports on blimps in coal mines, and then presents suggestions arising from the writer's recent investigation of the bumps in the Cumberland mines, made at the request of the mine operators, to determine the causes and remedies. These bumps, which occur under the simplest natural conditions, provide better opportunity for analysis of causes and effects than afforded under more complex conditions. The simple conditions in this field consist of relatively thin and level coal beds and uniform roof strata free from folding and faulting. Definition of Bumps Bumps in coal mines, as defined in this paper, are sudden violent bursts of coal that usually occur without warning in the extraction of pillars and may or may not be accompanied incidentally by falls of draw slate, upheaval of bottom and smashing of timbers. They are not known to occur where the advancing longwall method of mining, accompanied by

Jan 1, 1936

By L. W. Saperstein, Y. J. Wang

The development of computer programs to solve complex ventilation networks has reached a point of refinement where these programs become a necessary tool of the ventilating engineer. Such a program is presented. This program has an increased efficiency over its predecessors; it also includes the important facility of fixed quantity branches. This is in addition to the earlier capabilities of free-splitting, internal or external fans, and natural ventilation pressures. The program is written in Fortran IV with double-precision arithmetic and is listed here. Sample problems with input and output are also given. The recent, and presumably continuing, official interest in mine safety is bound to result in more stringent control requirements for ventilation. Anticipating this requirement, the Dept. of Mining, The Pennsylvania State University, has developed a computer program for the solution of complex mine ventilation networks. It is the opinion of this department that the program has sufficient merit to warrant its use in normal industrial operating conditions. Although the undergraduate mining engineering students of this department are routinely trained to use the program, it was felt that this training did not present a fast enough means of dissemination; consequently, this paper was written. The conception of the program is not new;1; however, the present version incorporates features that changes it from a scientific endeavor to a useful operating tool. These features include those of earlier programs, such as free-splitting, external or internal fans, and natural ventilation processes plus the important ability of handling fixed quantity branches, and of increased efficiency giving reduced cost. The facility for fixed quantities will be discussed in more detail. The solution of any network depends upon satisfying Kirchchoff's laws. These state that the sum of flow at any junction is zero, and that the sum of the pressure drop in branches totals the pressure drop in the mesh constituted by those branches, that is that the pressure drop around any mesh is zero. The terminology is derived from Synge," and is, in fact, that used by Topologists. A branch is any airway of uniform characteristics; a junction is where two or more airways meet (the minimum number of two airways is useful for describing the point where a low resistance airway changes into a high resistance one due to some physical constriction); a mesh is a path, along some branches, that returns on to itself, but does not need to traverse the same branch twice in order to do so; and a tree describes all those branches in a network which, while connected through all the junctions, do not form any meshes. Obviously the branches-in-tree are open-ended and can be shown to be one less than the number of junction (J-1). The branches-out-of-tree, also called basic branches, are equal to the number of basic meshes and are M = N — J + 1, where M is the number of meshes and N is the number of branches. The major difference between this network and an electrical one is that the law attributed to Atkinson for pressure drop is utilized rather than Ohm's law. Atkinson stated that the pressure drop (potential) is equal to a resistance factor multiplied by the quantity squared. This is written H = RlQlQ where H is the head loss. Q? is written in this factored fashion so that H will always have the same sign as Q. A negative sign is used to indicate that flow in a branch is opposite to that of its containing mesh. A mesh takes the direction of its basic branch. Utilizing Atkinson's law in Kirchchoff's second law and bearing in mind that the first law must remain satisfied, the program will determine all quantities, or head losses as they are related, for the branches of the network. The program uses a Geuss-Seidel form of in-teration, starting from an arbitrary Q and using a correction factor of the type first suggested by Professor Hardy Cross' that will handle the nonlinear equation of Atkinson's law. A fuller description of the interative process is included in Wang and Hartman.' Iteration continues until the correction factor is less than or equal to a preset error (E) or until a maximum number of iterations (MAXIT) have been reached. Utilizing an E of 50 cfm, sample problems have had rapid solutions. Rapidity of solution is also ensured by the method of selecting meshes and basic branches. This is done by the computer, which chooses the highest resistance branches, fixed quantity branches, and those containing fans, as basic branches. Natural ventilation pressures are handled as fans of constant pressure and may be assigned to any branch. The computer will generate a fan characteristic by polynominal fitting if a few operating points from the desired fan are input. A fan may be placed in any branch of the network, except those with fixed quantities. The strength of the present program over its earlier versions is its ability to handle fixed quantity branches. This means that certain branches can have their quantity determined, or preset, by the ventilating engineer. Operating under this constraint, the program will analyze the network and output the pressure change necessary to achieve this desired quantity. A call for positive head loss would require an auxiliary fan; negative head loss would require a regulator. This represents a powerful tool for the ventilating engineer who must insure that certain quantities of air pass across the operating faces or through working stopes. McPherson6 suggested that a ventilation program would be useful for determining in advance which branches were receiving inadequate air, and that subsequent calculations would then indicate the necessary corrections. The present program makes these corrections. Thus the mine manager can know immediately the consequences of changes to his system;

Jan 1, 1971

By N. G. W. Cook

The conventional cyclic system of deep-level mining by drilling and blasting gives rise to an inadequate degree of stope sorting when mining thin reefs. This results in poor utilization of the capital facilities of a mine in the form of shafts, haulages, airways, and associated equipment. Continuous and controlled removal of the thin gold-bearing portion of the reefs would permit better stope sorting and hence greater utilization of capital facilities. Results of experiments to develop hard-rock cutting machines for mining are reported and the benefits which could be derived from their use are discussed. Mining, from exploration through refining, is essentially a process of sorting in which the payable mineral, or metal, is progressively separated from the other constituents of the earth's crust with which it was originally associated. This takes place more distinctly at each step of the operation. What is it that determines the degree to which sorting should be carried out at each of the several steps comprising a whole mining operation? The formal answer is that degree of sorting at each step which results in the lowest overall cost for the complete separation. In practice, individual steps are chosen from those available in current technology, each of which effects a degree of sorting such that the quantity of material which must be sorted in the succeeding step is economically acceptable. It follows that any new technological development has repercussions throughout the whole mining operation and, more important, that the solution to excessive costs in any one step of the operation may lie not in improving the costly operation itself so much as in increasing the degree of sorting preceeding that operation. This concept, particularly in relation to deep-level mining of thin, tabular gold-bearing reefs in South Africa, is discussed here, and the most recent results achieved in the development of hard-rock cutting machines for stoping more selectively than is possible with explosives are presented. Deep-Level Mining Deep-level mining involves operations which are either not encountered, or are of only trivial importance, in near-surface mining. Near-surface, the major operations are those of rock breaking, transport, and milling. In deep-level mining, hoisting, environmental control, and strata control assume major importance. Some idea of the relative magnitude of these operations may be gained by comparing the separate amounts of energy which are required, or which must be controlled, to effect the various operations when, say, mining a tabular deposit 40 in. thick at 8000 ft below surface, Table 1. It is true that the costs of handling a given quantity of energy are not the same for each operation. Nevertheless, Table 1 does emphasize the fact that the operations of hoisting, strata control, and environmental control are of unique and major significance in deep-level mining. In particular, hoisting and environmental control place a heavy load on the reticulation system of the mine—the shafts, haulages, and airways. Typically, a new, deep gold mine with an annual revenue of about $35 million requires a total capital expenditure of about $140 million of which some $100 million is invested in developing and equipping this reticulation system. The ratio between annual turnover and capital invested of about one-quarter is exceptionally low, and it typifies the poor utilization of capital by the current technology of mining hard rock at depth. The average thickness of the reefs in the new South African goldfields varies from 10 to 30 in.,l and even in the thicker reefs the gold is often confined within a small fraction of the nominal thickness. Nevertheless, it is universal practice to mine these reefs at a stope width of about 40 in. or more, so that the quantity of rock broken in the stopes and hoisted out of the mine is between two and ten times the quantity of rock actually carrying a significant amount of gold. The reason for the adoption of such excessive stope widths is to be found in the method of rock breaking by drilling and blasting. The only free surface to which a blast hole can break is the stope face. It follows2 that each hole cannot have a burden in excess of the height of the free face if it is to break satisfactorily. To

Jan 1, 1971

By C. E. Tsoutrelis

A new method for determining rock drillability in diamond drilling is discussed; the method takes into consideration both penetration rate and bit wear. The method is based on drilling a rock specimen under controlled laboratory conditions using a model bit. The technique used for determining the experimental variables is extremely simple, quick, and reliable. Drillability is then determined by the mathematics of drilling. In considering the different factors that affect diamond drilling performance, the nature of the rock to be drilled is of outmost importance since it affects significantly the drilling costs and such other variables as bit type and design, drilling thrust, and bit rotary speed. Many attempts have been made to study this effect by correlating actual drilling performances either to certain physical properties of the rock being drilled1-? or to test drilling data obtained under laboratory conditions.7-13 These attempts were aimed at providing a reliable method of predicting by simple means the expected rock behavior in actual drilling, thus giving the engineer a tool to use in estimating drilling performances and costs in different types of rock. The purpose of this paper is to describe such a method by which rock drillability (a term used in the technical literature to describe rock behavior in drilling) could be determined in diamond drilling. It is believed that the proposed simple and reliable method will cover the need of the mining industry for a workable method of measuring the drillability of rocks. It should be emphasized, however, that since drill-ability depends on the physical properties of rock and each drilling process (diamond, percussive, rotary) is affected by different or partly different rock properties,14-l6 the proposed method of determining rock drillability cannot be extended to the other drilling processes. The results presented in this paper form part of an extensive three-year research program carried out by the author in the laboratories of the Greek Institute of Geology and Subsurface Research. During this period the effects of the physical properties of rocks and of such operational variables as drilling thrust and bit rotary speed in diamond drilling were investigated in detail. DRILLABILITY CONCEPT The literature is not devoid of drillability studies. While there are a number of investigators1,3,5-7,9-0,12-13,17 who have attempted to establish by direct methods (i.e., drilling tests under laboratory conditions) or indirect (i.e., through a physical property of rock) an index from which the drilling performance in a given rock may be estimated, very few6-7,9,12, of the proposed methods seem to be of much practical value to the diamond drilling engineer and none to date has been universally accepted. Commenting on the proposed methods for assessing rock drillability, Fish14 remarks that "for a measure of drillability to be accepted it is essential that penetration rate at a given thrust and bit life are elucidated as otherwise the method is of little value." This statement should be examined in more detail by making use of the penetration rate-drilling time diagram obtained in drilling a rock under constant operational conditions. Furthermore, the merits of using this diagram to describe rock drillability will be pointed out. At the same time reference will be made to this diagram when discussing some previously proposed methods. Fig. 1 illustrates such a diagram for three rocks,A, B, and C, which have been diamond drilled under identical conditions. It is assumed here that rocks A and B have the same initial penetration rate, i.e., VOA = Vog, but since rock B is more abrasive than A, rapid bit wear occurs and as a result the fall of its penetration rate with respect to time is more vigorous than in rock A. This is shown graphically by a steeper V = f(t) (0 curve in this rock than in rock A. Rock C has a lower initial penetration rate, due to higher strength properties16 but since it is not very abrasive, only a slight fall of its penetration rate occurs during drilling (in this category are some limestone and marbles with compressive strength above 1000 kg per sq cm). It follows from the foregoing considerations that the characteristic for each rock curve (I) is a function of (i), the penetration rate of the rock Vo recorded at the instant of commencing drilling, which determines the starting point of the curve (1) on the y-axis and (ii), the abrasive rock properties which determine the rate of fall of Vo with respect to time. Thus, curve (I) provides an actual picture of the rock behavior in drilling for given operational conditions, and it can be used with complete satisfaction to assess rock drillability. It can be seen clearly from Fig. I that proposed methods for assessing rock drillability by measuring the

Jan 1, 1970

By R. L. Ash, R. R. Rollins, C. J. Konya

An investigation was performed to determine conditions for optimizing the spacing of simultaneously initiated multiple explosive columns. This was done by using models of mortar, dolomite, and Plexiglas with 10-grain mild detonating fuse as the explosive charge. It was desired to simulate blastholes with multiple primers initiated by detonating fuse or when high-velocity explosives are used in low-velocity materials. It was found that optimum spacing between multiple charges was strongly influenced by charge length. At less than optimum charge length, the spacing at which complete shearing was possible between adjacent charges decreased exponentially with a subsequent loss of broken material volume. For charges fired simultaneously, larger burdens and spacings were possible as compared to those necessary for single-crater charges. For each material studied, there was a characteristic optimum charge length and a maximum attainable spacing at any given burden. Proper selection of the spacing distance between charges is fundamental to successful blasting. Its value directly affects the cost of drilling and explosives used per unit of broken material. In addition, the choice of a spacing that is Compatible with a given set of blasting conditions aids in the control of fragmentation sizing, ground vibrations, overbreak, and throw which in turn, influence other production costs. For example, normally loaded blastholes that are spaced too closely invariably promote overbreak and usually give coarse fragmentation. Unless care is taken, airblast and violent flyrock will occur and under certain conditions cutoffs and misfires may result. Too large a spacing, on the other hand, frequently leads to conditions that form bootlegs or toes. The choice of a particular spacicg to use, however, is largely a matter of individual experience and judgment, usually based on trial and error. Very little is known or can be found in the literature with regard to how the spacing between charges is related to field conditions and charge geometry. As a general rule, the firing time sequence of adjacent charges and properties of a material are thought to have the most significant influence on the spacing distance best suited for any given field condition. For example, delayed initiation of adjacent charges usually always requires a closer spacing than when charges are fired at the same time. This should be expected if one considers that the energy normally dissipated and lost in the surrounding ground from charges fired independently would be captured and utilized for breaking material between charges when they are initiated together. Spacing can be extended also when charges are aligned with structural planes of a material, such as jointing, along which shearing is relatively easy. It is customary to relate the spacing (S) between charges to their common burden (B) in the form of a spacing ratio, or SIB. The burden normally is considered as the optimum depth or distance from any single charge perpendicular to the nearest free or open face at which the desired fragmentation and maximum crater yield are obtained. For production blasting, value of the ratio is generally considered to vary from 1 to 2, depending on conditions.1-6 When adjacent charges are fired independent of one another, the value varies from 1 to about 1.4, the closer amount being employed to square corners or produce craters having the ideal 90" apex angle. The larger ratio is the geometric balance value for craters having an apex angle of 135". The basic ideal crater forms in the plane of the charge diameter for charges fired independently are shown in Fig. 1. In the event charges are fired simultaneously, geometric balance in the plane of their charge diameters suggests that a spacing ratio near 2 would be appropriate, as illustrated by Fig. 2. In practice, however, some compromise ratio value must be selected to conform with the specific ground conditions. An example would be where the jointing planes tend to produce 60° or 120° crater angles, the appropriate geometrically balanced charge arrangement being given by Fig. 3. In this condition, the spacing ratio is 1.15, not 1 or 1.4 as suggested for the 90° cratering of independently fired adjacent charges. In view of the foregoing, it would seem logical to assume that whenever charges all having the same burden are fired at the same time, spacing distances always can be greater than those permitted by charges fired independently. In practice this is not the case, however.

Jan 1, 1970

By F. A. Jerabek, H. L. Hartman

This study presents results of investigations in a small model of segregation in discharged fill slurry, the quantity of solids escaping in the drain liquid, and the relationship of solids to slurry concentration and discharge rate. Compressibilities of fine fill material, in dry and wet states and individual size fractions, are compared. Two terms are frequently used in describing the hydraulic placement of fill material underground: hydraulic stowing is mostly associated with the underground flushing of coarse-sized fill material, such as coal washery refuse; while hydraulic backfilling refers to placement of small-sized fill material, e.g., sand or classified mill tailings, mainly in metal mines. Characteristically, in backfilling, only a few hours may elapse between the time the material is hoisted as ore and redeposited as fill underground. Hydraulic filling was first used about 100 years ago to control subsidence in coal mines. It was subsequently adopted in many mining operations all over the world, but its basic concepts have never been adequately studied and analyzed. This is not surprising, when one considers the number of variables involved in the hydraulic backfilling operation. Recently, much work has been done on the problems of transportation of slurries in pipes. Comparatively less research work has been done concerning the physical properties of fill material (compressibility, cementation, permeability, etc.). Very little work has been done on the problems of fill placing, distribution, and segregation of fill material discharged in the mine void. It is the purpose of this paper to explore two of the basic, lesser known aspects of filling: segregation and compressibility in discharged fill slurry. Difficulties in observing the behavior of hydraulic fill slurry during the pouring operation underground, and insufficient knowledge to conduct any kind of meaningful theoretical analysis, suggest experimental model studies. The tests described in the following sections were conducted in the Rock Mechanics Laboratory of the Department of Mining at Penn State.

Jan 1, 1965

By J. F. Bryson

Instantaneous outbursts of coal in underground workings have occurred frequently in various coal fields in Nova Scotia; British Columbia; Canada; South Staffordshire, England; and the states of Washington, Virginia, West Virginia and eastern Kentucky in the United States. Such outbursts in eastern Kentucky have been called "mountain bumps " because they are more likely to occur under mountain cover than elsewhere, but recently they have become generally known as "coal bumps." The eastern Kentucky coal field has been troubled by these outbursts since 1923. Many men have been injured and killed, and the morale of the workmen in the vicinity of such outbursts has been affected to such an extent that mine production has suffered for a considerable period following each outburst. Such phenonlenal occurrence in any coal field naturally has led to discussion as to the cause and means of prevention. In the eastern Kentucky field, local mine officials, workmen, mining engineers, and many others have offered many ideas as to the cause and many methods of prevention have been suggested, and tried without success. Some of the plans were either practically or economically unsound and in general impossible to follow. To enumerate all the suggestions would require a large volume. One cause mentioned was that they might be due to gas under high pressure, as in Lower Silesian mines, and the remedy reeommcnded was to drain the gas by drilling holes into a suspected area. A hole 30 ft. long was drilled into a 60-ft. pillar to allow the gas to drain off, hut the pillar bumped the next day. Realizing the importance of obtaining the best information available on coal bumps, the writer, with the permission of the Harlan County Coal Operators' Assoeiation, invited Mr. George S. Rire, Chief Mining Engineer of the U.S. Bureau of Mines, who had investigated bumps in several countries, to conduct an investigation of coal bumps in this field. Mr. Rice, by permission of the Director of the G.S. Bureau of Mines, accepted the invitation, and it was the writer's privilege, in company with

Jan 1, 1936

By J. F. Bryson

Instantaneous outbursts of coal in underground workings have occurred frequently in various coal fields in Nova Scotia; British Columbia; Canada; South Staffordshire, England; and the states of Washington, Virginia, West Virginia and eastern Kentucky in the United States. Such outbursts in eastern Kentucky have been called "mountain bumps " because they are more likely to occur under mountain cover than elsewhere, but recently they have become generally known as "coal bumps." The eastern Kentucky coal field has been troubled by these outbursts since 1923. Many men have been injured and killed, and the morale of the workmen in the vicinity of such outbursts has been affected to such an extent that mine production has suffered for a considerable period following each outburst. Such phenonlenal occurrence in any coal field naturally has led to discussion as to the cause and means of prevention. In the eastern Kentucky field, local mine officials, workmen, mining engineers, and many others have offered many ideas as to the cause and many methods of prevention have been suggested, and tried without success. Some of the plans were either practically or economically unsound and in general impossible to follow. To enumerate all the suggestions would require a large volume. One cause mentioned was that they might be due to gas under high pressure, as in Lower Silesian mines, and the remedy reeommcnded was to drain the gas by drilling holes into a suspected area. A hole 30 ft. long was drilled into a 60-ft. pillar to allow the gas to drain off, hut the pillar bumped the next day. Realizing the importance of obtaining the best information available on coal bumps, the writer, with the permission of the Harlan County Coal Operators' Assoeiation, invited Mr. George S. Rire, Chief Mining Engineer of the U.S. Bureau of Mines, who had investigated bumps in several countries, to conduct an investigation of coal bumps in this field. Mr. Rice, by permission of the Director of the G.S. Bureau of Mines, accepted the invitation, and it was the writer's privilege, in company with

Jan 1, 1936

By R. S. Montgomery

Laboratory experiments simulating rock drilling were conducted on a number of commercial grades of tungsten carbide in order to determine their wear rates when they are used as inserts in percussive rock drill bits. The test rock was Barre granite and tests were made at conditions similar to those of the carbide buttons near the periphery of a large, down-hole, button bit. In addition, some wear rate measurements were made at a low blow energy. Wear rates of these carbide buttons were an extremely sensitive function of their cobalt contents and hardnesses and were proportional to e-°.57Hup to a hardness of about 91 RA. The point of minimum wear rate, which occurs at about 92 RA, corresponds to the point of maximum compressive strength. Modern percussive rock drill bits have inserts of cemented tungsten carbide on their wear faces. These inserts have drastically reduced bit wear in comparison to that of earlier all steel bits. The most common cemented carbide grade used in percussive rock drill bits contains about ten weight percent cobalt, although grades containing more or less cobalt are also used. Contrary to the case with carbide in metal-cutting tools, alloying materials are rarely used in mining grades. There are a number of designs of rock bits in current use. The most common form is an X bit which has four chisel-form radial wings. The button bit is another important design and one which has been recently developed. This design uses a flat steel face studded with relatively small hemispherically tipped carbide buttons. These types are illustrated in Fig. 1. Studies have been made of the sliding wear of cemented tungsten carbide on rock 1-4 and multiple-impact fracture', 5-6, b~t there has been comparatively little published on the percussive wear of this material although the wear of the impact face is the most important. There is some available data on the rate of dulling of rock bit inserts in actual drilling 7 but no laboratory studies. Sliding wear studies are useful for understanding the wear of coal and mineral cutting tools and gauge wear of percussive drill bits. but the

Jan 1, 1970

By Roy R. Jones, R. R. Sawyers

It has long been comnlon knowledge that workers in anthracite are prone to develop a disabling disease of the lungs. Some of the earliest scientific contributors dealing with anthracosis were: Pearson1, in his report in 1813, on the coloring matter of the bronchial glands and black spots in the lungs; Stratton2, in 1838, who first termed the condition anthracosis; Greenhow3, who reported to the pathological society in 1865, on the Black Lung of the Collier, and Pick4 in 1870, who reported on the chemical composition of the ash of a collier's lungs. Visconti5, in 1870, is accredited with being the first to use the term "silicosis" to denote the pathological condition of the lungs resulting from the inhalation of silica. Rovida6 confirmed the reports of Pick and Visconti in 1871, in his report on a case of silicosis and chemical examination of the lungs for silica content. The relationship of pulmonary infection, as associated with disabling pneumoconiosis affecting the coal miner, has frequently been discussed. While most reports have emphasized the role of infection, there has been some disagreement regarding the prevalence of tuberculosis among those suffering from anthracosis. The report by Wainwright and Nichols7 in 1905 has frequently been cited to show that the incidence of pulmonary tuberculosis among coal miners was less than that among the general population. It was their opinion that pulmonary tuberculosis was more prevalent among the miners' families than among the miners themselves. Collis and Gil-christ, in 1928 in their study of the Effects of Dust upon Coal Trimmers, concluded that among such workers there was an excess of deaths from bronchitis and pneumonia but not from pulmonary tuberculosis. They

Jan 1, 1936

By Roy R. Jones, R. R. Sawyers

It has long been comnlon knowledge that workers in anthracite are prone to develop a disabling disease of the lungs. Some of the earliest scientific contributors dealing with anthracosis were: Pearson1, in his report in 1813, on the coloring matter of the bronchial glands and black spots in the lungs; Stratton2, in 1838, who first termed the condition anthracosis; Greenhow3, who reported to the pathological society in 1865, on the Black Lung of the Collier, and Pick4 in 1870, who reported on the chemical composition of the ash of a collier's lungs. Visconti5, in 1870, is accredited with being the first to use the term "silicosis" to denote the pathological condition of the lungs resulting from the inhalation of silica. Rovida6 confirmed the reports of Pick and Visconti in 1871, in his report on a case of silicosis and chemical examination of the lungs for silica content. The relationship of pulmonary infection, as associated with disabling pneumoconiosis affecting the coal miner, has frequently been discussed. While most reports have emphasized the role of infection, there has been some disagreement regarding the prevalence of tuberculosis among those suffering from anthracosis. The report by Wainwright and Nichols7 in 1905 has frequently been cited to show that the incidence of pulmonary tuberculosis among coal miners was less than that among the general population. It was their opinion that pulmonary tuberculosis was more prevalent among the miners' families than among the miners themselves. Collis and Gil-christ, in 1928 in their study of the Effects of Dust upon Coal Trimmers, concluded that among such workers there was an excess of deaths from bronchitis and pneumonia but not from pulmonary tuberculosis. They

Jan 1, 1936

By H. Landssberg

Subsidence caused by mining operations has been a matter of interest for the mining engineer for just 111 years, since the Belgian committee for study of subsidence in the city of Liege submitted its first report in 1825. This interest is fairly expressed in the enormous amount of literature that has since appeared on this subject. A bibliography, probably far from being complete, which preceded this study as a preliminary survey, comprises more than 800 titles. In spite of this fact the problem remains today as acute as it has ever been. The questions of as good a recovery as possible of the national resources, the safety of the operations, and the protection of private and public property are the important angles from which to look at the problem. The Problem To systematize an investigation regarding the effects of mine excavations two different aspects have to be dealt with: (1) the subsidence of the immediate roof above the workings, affecting the safety of the workers and the progress and efficiency of the operations; (2) the subsidence of the surface, which is also, from the point of time, the second event. The immense importance of surface subsidence as regards the impairment of surface structures of every kind in a mining area is self-evident. Both effects result from reactions of the strata overlying the excavations and they depend mainly on the extent of the excavations and the structure and geological condition of the strata. It has been stressed repeatedly, in this country principally by G. S. Ricel,† that for the understanding of the phenomenon of subsidence accurate surveys are the most urgent necessity. In the past, however, at least in the normal course of mining operations, these surveys have been very crude. The control of the roof in the mine is left to the observation of the behavior of the timber, warning cracks and sounds in roof and walls and to the general judgment and experience of the worker and his supervisors. In any given case this may or may not be sufficient to prevent fatalities but certainly it does not give any accurate information as to what is happening in roof and strata. On the surface, principally in endangered areas, surveys have been conducted more frequently, but also in this case usually weeks or even

Jan 1, 1936

By H. Landssberg

Subsidence caused by mining operations has been a matter of interest for the mining engineer for just 111 years, since the Belgian committee for study of subsidence in the city of Liege submitted its first report in 1825. This interest is fairly expressed in the enormous amount of literature that has since appeared on this subject. A bibliography, probably far from being complete, which preceded this study as a preliminary survey, comprises more than 800 titles. In spite of this fact the problem remains today as acute as it has ever been. The questions of as good a recovery as possible of the national resources, the safety of the operations, and the protection of private and public property are the important angles from which to look at the problem. The Problem To systematize an investigation regarding the effects of mine excavations two different aspects have to be dealt with: (1) the subsidence of the immediate roof above the workings, affecting the safety of the workers and the progress and efficiency of the operations; (2) the subsidence of the surface, which is also, from the point of time, the second event. The immense importance of surface subsidence as regards the impairment of surface structures of every kind in a mining area is self-evident. Both effects result from reactions of the strata overlying the excavations and they depend mainly on the extent of the excavations and the structure and geological condition of the strata. It has been stressed repeatedly, in this country principally by G. S. Ricel,† that for the understanding of the phenomenon of subsidence accurate surveys are the most urgent necessity. In the past, however, at least in the normal course of mining operations, these surveys have been very crude. The control of the roof in the mine is left to the observation of the behavior of the timber, warning cracks and sounds in roof and walls and to the general judgment and experience of the worker and his supervisors. In any given case this may or may not be sufficient to prevent fatalities but certainly it does not give any accurate information as to what is happening in roof and strata. On the surface, principally in endangered areas, surveys have been conducted more frequently, but also in this case usually weeks or even

Jan 1, 1936