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By M. J. Dailey

This paper is one of a series being prepared by the United States Bureau of Mines on mining practices, methods, and costs in the various mining districts of the United States. The property of the Silver King Coalition Mines Co. is situated in the western past of the Park City mining district, Utah, about 1 ½ miles southwest of the town of Park City, The elevation at the collar of the main shaft is 8,137 feet above sea-level.

Jan 1, 1930

By A. Wahab Khair, Richard D. Begley

"This paper presents results of destructive tests performed on rectangular coal specimen under the influence of two vertical and simultaneous wedge indentation forces. Several wedge designs and spacings were utilized in compression tests. The objective of these experiments was to qualitatively monitor the effect of different wedge designs and spacings on the stress distribution within the coal specimen.Coal samples from the Waynesburg coal seam were cut and trimmed to cubical dimensions of approximately 5 X 10 X 20 cm. These specimen were subjected to parallel compressive forces through two vertically applied wedge indentors. Loading was achieved with a hydraulic hand pump in order to facilitate interferogram construction. The load increments were limited due to the sensitivity of the holographic interferometric technique (holometry) in measuring small displacements; therefore, interferograms were taken only during similar increments in order to allow for a qualitative comparison between different wedge design and spacings. In some cases, interferograms were taken immediately before and after failure of the specimen.Introduction Failure Modes Three distinct modes of failure occur when geologic materials are subjected to wedge indentation forces. As shown in Figure 1, a crushed zone is located immediately adjacent to the indentor while chipping can occur if the fractures initiated in the crushed zone extend upwards to the free surface.. The third failure mode is a zone of intricate frac¬tures located between the crushed zone and the chipping zone. Several experiments have been conducted over the years to better understand crater formation and stress distribution within geologic materials under the influence of an indentor. Generally speaking, crack initiation and propogation observed in these experi¬ments correlate fairly well with wedge indentation theory; however, the nature of the .fractures is greatly influenced by different material properties such as grain size and cleavage planes. In other words, a wide range of behavior has been found in different rocks subjected to wedge penetration. For example, it has been observed that some rocks are merely crushed and indented while others are more prone to cracking and chip formation. Therefore, the development or nondevelopment of chips depends mostly on wedge geometry, rock type and depth of penetration. This work is part of the study associated with opti¬mization of cutting parameters, which is best modeled by using brittle materials and subjecting them to rigid wedge penetration tests."

Jun 1, 1988

By L. O. Bacon

This Bureau of Mines report summarizes the results of laboratory work in the design of instrumentation for obtaining the dynamic tensile strength of rock at minimum loading. Minimum loading is defined as the lowest tensile loading by pulse reflection to cause fracturing. The instrumentation for obtaining dynamic tensile strengths of rock cores evolved into the use of a compressed-air pellet gun to impact one end of a rock core, suspended by wires, at increasing increments of air pressure. Maximum strain was measured on most or all of the impacts, and the strain at fracture of the core was obtained by extrapolating the pellet velocity-strain curve to the pellet velocity at fracture. Stress at tensile fracture was obtained from the maximum s train, and the dynamic Young's modulus was calculated from the rod velocity by using simple rod theory.

Jan 1, 1962

By Russell R. Trengove

Core samples from an ore horizon 2,000 feet below the surface were successfully obtained by rotary drilling at the property of the Eureka Corp., Eureka, Nev. The ore was soft and friable; cores were not obtained previously with diamond drills. Whipstocking was used to get samples in a zone 200 feet from a vertical hole. A technique in shipstocking, by which the direction and deviation of the holes were controlled by means of nonfriction-type survey instruments, was successfully employed.

Jan 1, 1956

By Rudolf E. Greuer, Linneas W. Laage, Xinton Chang

MFIRE is a computer simulation program that performs normal ventilation network planning calculations and dynamic transient-state simulation of ventilation networks under a variety of conditions The program is useful for the analysis of ventilation networks under the influence of natural ventilation, fans, fires, or any combination of these MFIRE simulates a mine's ventilation system and its response to altered ventilation parameters such as the development of new mine workings or changes in ventilation control structures, external influences (varying outside temperatures), and internal influences (fires) Extensive output enables detailed, quantitative analysis of the effects that the alteration will cause to the ventilation system The program was developed at Michigan Technological University (MTU), Houghton, MI, with support from the Bureau of Mines A personal computer (PC) version of the program was completed under contract for the Bureau by MTU The purpose of this report is to document the installation and use of the PC version of the program MFIRE

Jan 1, 1990

By T. H. Ueng, S. D. Thompson, Y. J. Wang

"INTRODUCTION There is considerable interest among mining companies and government agencies in finding new ways to better control the generation and dispersion of respirable coal dust. Mine face ventilation generally has been considered the most effective tool for lowering the concentrations of respirable dust. Unfortunately, only a limited amount of research has been undertaken to determine the relationships between the various face ventilation parameters and the dust dispersion patterns. If these relationships are better understood, more efficient ways of controlling dust generation and dispersion may be evident. A practical and economical way to gain more in-depth understanding of these relationships is through scale model studies. Past research (1, 2, 3) has substantiated that scale models can be successfully used in studying ventilation parameters provided that geometric, kinematic, and dynamic similitues are maintained.To better understand face ventilation and dust dispersion, a 1/5 scale model of a room and pillar mine face was constructed. This model was then used to conduct experimental investigation on two ventilation parameters. The parameters which were examined included the line brattice distance from the face and the air quantity. The values of these parameters were varied in numerous experiments so that their effects on dust dispersion could be measured.The experiments of this study were unique in two ways. First, two different substances were used to simulate the respirable coal dust. Experiments were conducted with a tracer gas (methane) serving as the dust simulation substance. They were then repeated with an actual dust (Arizona Road Dust) replacing methane as the dust simulating substance. In previous scale model studies, the dust simulation substance has been limited to a tracer gas such as methane. The validity of using a tracer gas to simulate dust is not clearly known yet. The other unique thing about these experiments was the number of points measured for methane (or dust) concentration. Twenty five points in a grid arrangement were measured during each experiment of this study. The number of points measured for concentration in other studies has generally been very small (often 2 to 4). It is difficult to obtain an accurate assessment of how the ventilation parameters are effecting dust dispersion when only 2 to 4 points are measured."

Mar 1, 1989

By C. B. Manula, R. L. Sanford, R. A. Rivell

A - Contains the coordinate pair information for deriving the co¬efficients of the 3rd degree polynomial which approximates the loading station equation. C(N) - Coefficients for third degree approximating polynomial. I - General DO LOOP variable. K - Position in BRANCH of next character. L - Format specification in DEFINE FILE statement. X - Uniform displacement for random uniform number. Y - Load increment At. AO-A3 - Coefficients of approximating polynomial for loading station equation. GR - Grade direction, or grade resistance. IQ - Counts quadruples of distance grade speed and end of route for determining total distance. IT - Intersection counter in XXXDEF. Il - Station number trip is leaving. Jl - Station number for trip destination. J2-J8 - Motor assignments and functions for given trip. KK - Location of starting grade and distance in DISTGR KK - Position of decimal point in real number. KZ - DO LOOP variable for motor assignment check. LL - Location of ending grade and distance in DISTGR. Ll - Position in IECB of station satisfied. L3 - Next available position for IECB sort. MM - Flag for printing the failure of sanding to overcome slippage. NN - Trip destination in MOTSER RR - Rolling Resistance. XK - Station level or type.

Jan 1, 1974

By Keith G. Wallace

Coal mine ventilation systems are often subject to high leakage rates. As a result, changes in airflow resistance will strongly affect the efficiency with which air is delivered to the working place. One major source of airflow resistance is the line brattice used to direct air from the last open crosscut to the working face. Because it is a great distance from the fan, the resistance of the line brattice can result in more overall leakage than an equivalent resistance closer to the fan. Substituting auxiliary fans for brattice eliminates this source of resistance, with improvements in system efficiency and cost. Also, in some mines, the leakage that leads to spontaneous combustion can be reduced. The benefits obtained by several different brattice and fan substitutions have been studied by Mine Ventilation Services, Inc., Lafayette, CA, under contract to the U.S. Bureau of Mines. Changes in main fan duty were compared with additional costs associated with auxiliary fans. Results indicate that in some circumstances, considerable leakage reductions and cost savings are possible.

Jan 1, 1990

By Murray Jacobson

The respirable dust standard in the Federal Coal Mine Health and Safety Act of 1969 is designed to prevent disability and death from coal workers' pneumoconiosis. Beginning June 30, 1970, the operators of coal mines were required to maintain the average concentration of respirable dust in the active workings at or below 3.0 milligrams of dust per cubic meter of air. The standard is reduced to 2.0 milligrams per cubic meter after Dec. 30, 1972. This paper presents the equipment and procedures used by the coal mine operators in implementing the respirable dust sampling program established by the Secretary of the Interior and the Secretary of Health, Education, and Welfare and shows the effect of the law, during the first year, on the reduction of dust levels in selected mines. The system employed by the Bureau of Mines for analyzing and processing the dust samples is also presented.

Jan 1, 1971

NIOSH recommends that exposures to respirable coal mine dust be limited to 1 mg/m3 as a TWA concentration for up to 10 hr/day during a 40-hr workweek, measured according to current MSHA methods (see Section 5.1 and Appendix J). NIOSH recommends that sampling be conducted with a device that operates in accordance with the NIOSH accuracy criteria [Busch 1977; Busch and Taylor 198 11 and the international definition of respirable dust [ACGIH 1994; CEN 1993; IS01993; Soderholm 1991a,b; 1989].* The REL represents the upper limit of exposure for each worker during each work shift. For single, full-shift samples used to determine noncompliance, NIOSH recommends that MSHA make no upward adjustment of the REL to account for measurement uncertainties [NIOSH 1994c] (see also Section 5.6.2). NIOSH further recommends that all reasonable efforts be made to reduce exposures to respirable coal mine dust below the REL through the use of engineering controls and work practices.

Jan 1, 1997

By E. J. Gleim

To make a thorough survey of a coal mine in order to eliminate hazards, the competent safety engineer must be familiar with mere than one branch of engineering. It is essential that he know the properties of mine gases and the methods of detecting them, as well as hew to prevent dangerous as accumulations. It is also essential that he understand the action of overlying strata when the coal is removed, and the proper methods of protection by timbering or other suitable means of support. He must know the value of protective clothing in reducing injuries and mast be informed as to the correct handling and use of explosives to avoid premature blasts, misfires, and ether failures in blasting practice. He must understand hew to operate locomotives and cutting, loading, and various other types of machines to be able to recommend effective safeguards for protecting the men who use them in addition to being informed as 'to these and many ether activities connected with the mining of coal, he must net overlook electrical engineering if he wishes to combat the fire, shook, and ignition hazards attendant upon the use of electrical machinery and circuits in and around the mine intelligently. There are many points relating to safely in the selection, maintenance, and use of electrical circuits and equipment that are perfectly obvious to the electrical engineer but perhaps are not se evident to those not trained in this particular branch of engineering. Engineers of the U.S. Bureau of Mines who have made a special laboratory study of electricity as related to safety in minus haw e also made inspections of electrical installations at a large number of bituminous mines. These inspections have revealed practices and conditions that not only do not contribute to safety but that add nothing to efficiency and economy in the operation of the mine.

Jan 1, 1936

By L. H. Johnson

"INTRODUCTION The Pennsylvania anthracite industry produced 53 million net tons of anthracite during 1945. Fifty-one million tons was prepared for the commercial market (3), 3/ and 2 million tons (consisting of fine sizes) was used for colliery power consumption. Of the number of tons prepared for market, 17 million tons (33 1/3 percent) was prepared by the use of acid mine water treated with lime.Mine water from underground workings is pumped, drained, and stored in huge quantities. The mine-drainage systems in the anthracite region handle over 200 billion gallons of water annually, of which 150 billion gallons is pumped to the surface. More than 69 billion gallons of water is impounded in underground pools (1). Much of this water is employed for breaker use, dust-control installations; hydraulic backfilling, combatting mine fires; and aiding in the transportation of anthracite in gently dipping places where sheet iron is used Roos(4) estimates that 1,100 gallons of mine water is utilized in the preparation of anthracite mined."

Nov 1, 1946

By R. E. Head

[Surveys of the manganese situation conducted by the Geological Survey any other agencies have definitely established the, fact that the tonnage of high-grade manganese or in the United States is relatively small. It is therefore obvious that large-scale production of manganese is possible only through beneficiation of low-grade material. Deposits of low-grade manganese ore are being examined by engineers of the Mining Division of the Bureau of Mines to determine their size and possibilities as sources of manganese on a tonnage basis. The Metallurgical Division of the Bureau is developing and perfecting beneficiation processes with the objective of utilizing such low-grade ores. Microscopic study of these ores represents one phase of the research.]

Jan 1, 1941

By R. E. Head

[Surveys of the manganese situation conducted by the Geological Survey and other agencies have distinctly established the fact that the tonnage of high- grade manganese ore in the United States is relatively small. It is therefore obvious that large-scale production of manganese is possible only through beneficiation of low-grade material. ]

Jan 1, 1941

By James Westfield, S H. Ash

Because of the increasing shortage of fuel in the United States, especially of residential fuel in the Eastern and New England States, it was essential that the maximum production of anthracite be obtained to help win the war and to prevent human distress. The nine railroads that depend on the anthracite region for a substantial portion of their business were not congested unduly with other traffic, because the region was not participating in industrial activities to the same degree as in many other sections of the United States; consequently, transportation of anthracite to the New England States, New York, New Jersey, and eastern Pennsylvania over a shorter rail haul than for any other fuel was desirable. The prevention of floods that would cause a loss of anthracite production was important to the war program. Every shutdown deprives the northeastern section of the United States of anthracite that is needed badly. Failure to prevent inundation of anthracite mines halts production, endangers the lives of workmen, damages, sometimes beyond repair, vital underground equipment, and threatens the permanent loss of a large portion of the Nation’s anthracite reserves. Funds were made available to the Department of the Interior, Bureau of Mines, for the fiscal year ended June 30, 1945, through a supplemental appropriation for anthracite investigations, one of which was to conduct inquiries and scientific and technologic investigations that have a bearing on the anthracite mine-flood problem. As part of the investigation, the Bureau proposed to obtain the cooperation of the anthracite-mining companies and the State in effecting a program to prevent the inundation of mines in the anthracite region. No large engineering projects were undertaken by the Bureau during the fiscal year ended June 30, 1945, on the anthracite mine-flood problem. Small projects were chosen that could be completed within 1 year to prevent loss of anthracite production at a critical time. It was believed that such projects would lessen the chance that flood water would inundate active mine workings.

Mar 1, 1946

By T. J. Ballard

Corundum is a natural alumina (A120 ) and is next to the diamond in hardness . Except for the gem varieties , it is used principally as an abrasive , chiefly in grinding , dressing , and polishing metals and optical glass . Natural corundum is preferred to artificial abrasives for certain specialized jobs . Most of the corundum ( excluding emery ) mined in the United States came from the Corundum Hill mine in North Carolina and the Laurel Creek mine in Georgia . The Corundum Hill mine provided over half of the domestic production of corundum prior to 1900. Since 1900 , the importation of foreign corundum and the manufacture of artificial abrasives caused the decline of corundum mining in the United States .

Mar 1, 1947

By Oliver Bowles, J. E. Crawshow

A line of cordeau detonating fuse was placed in hole No. 2, no electric detonators being used in the hole. After 12 or 13 cases of 40 per cent nitro starch powder had been poured into the hole in loose form, it was found that a space of only 17 ft. remained for stemming. It was decided that space for an additional case of powder might be obtained by tamping the charge. The tamping was done with a heavy plunger, about 3 or 3 inches in diameter and 10 inches long, made of lead with an iron core with an iron eye in the top to which a 3/4 inch rope was attached, and weighing between 30 and 40 pounds. This weight was intended for use in sinking explosive in wet holes, and not for tamping. It is estimated that the tamping had continued at least 10 The quarry superintendent had gone a considerable distance for a box of powder, returned, and reached a point about 10 feet from the hole, minutes.

May 1, 1920

By Wm. O. Vanderburg

This paper describing the mining practices at the Block P mine of the St. Joseph Lead Co. at Hughesville, Mont., is ore of a series of similar papers being prepared by the Bureau of Mines on mining practices, methods, and costs in the various miner of the United States. ACKNOWLEDGMENTS In the preparation of this paper the writer acknowledges the assistance of R. H. Willcomb, general manager, William Thorson, mine superintendent, and B. P. Bellport, mining engineer, of the Hughesville division of the St. Joseph Lead Co.; and of E. D. Gardner; supervising engineer, U. S. Bureau of Mines, Southwest Experiment Station, Tucson, Ariz.

Jan 1, 1931

Develop an environmentally feasible method of mining shallow oil sands without removing overburden. Approach Oil sands are extracted through a single borehole by cutting into the sands around the borehole with a high pressure water jet, and pumping the resulting slurry to the surface. Although oil can be extracted from shallow fields by open-pit mining methods, this type of mining may create significant environmental impacts, including; (1) disruption of the surface, (2) accumulation of waste rock piles, (3) accumulation of tailings, (4) damage to the ground water quality, and (5) surface water pollution. The borehole mining system developed and successfully field tested by the Bureau of Mines offers a method for extracting oil from oil sand with minimal disturbance to environmental quality. (For examples of other applications of borehole mining, see Technology News Nos. 95, 63 56, and 48.) How It Works Prior to the start of mining operations, baseline conditions for ground subsidence and ground water quality are established at the test site so that the degree of subsidence and changes in ground water quality can later be determined. Oil sands mining is conducted using the Bureau of Mines borehole mining system. The borehole mining tool consists of a 12-inch-diameter pipe string and accessories capped with a three passage swivel terminated with a mining section. The mining section has an auger at the bottom. Four slurry inlet holes and an eductor are located above the auger, and a single cutting nozzle is located near the top of the mining section. The tool generates a high velocity water jet that erodes and slurrifies ores. The slurry is drawn into the inlet of the eductor which lifts the slurry to the surface where it is deposited into a slurry discharge sump. To detect evidence of subsidence, surveys of the land surface are made at regular intervals during, and until 30 days after, the termination of the mining operations. Samples of ground water from the borehole and two adjacent monitor holes are taken on alternate days during the mining operations, and are later analyzed to determine if any ground water contamination has occurred.

Jan 1, 1981

By Richard H. Oitto

The Bureau of Mines and Kaiser Steel Corp. are cooperatively demonstrating the shield-type roof support in the longwall mining of a 10-1/2-foot-thick coalbed at the York Canyon mine near Raton, N. Mex. As part of the demonstration, the shield supports have been successfully and safely moved from a completed longwall face to a new longwall face. When these shield supports were moved in 1976 they were the largest and heaviest longwall support system in use in U.S. coal mines. The shield supports, preparations for moving them, the actual move, and installation at the new face are discussed.

Jan 1, 1977