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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 G. R. Hopkins

One of the largest items in the expense of operating a refinery is the cost of the fuel used in distillation and for miscellaneous purposes such as in the heating of buildings and in the generation of electricity, On a B. t. u. basis, the average refinery consumes approximately one-tenth as much fuel as it produces. Theoretically, the fuel consumption at refineries should be much smaller than this, but furnaces can riot be designed to be 100 per cent efficient' and, in addition, a large part of the waste heat is low-temperature heat which would not pay for the expense of recovering it. Annual surveys of fuel consumption at refineries have been made by the United States Bureau of Mines, beginning with 1925. During the 5-year period, 1925-1929, the refineries made steady progress in fuel economy. This is evidenced by the Computed figures of heat requirement Per barrel of crude oil refined, which show a decline from 829,000 B. t. u. in 1925 to 639,000 B. t. u. in 1929 (see fig. 1). In 1930, however, the trend in fuel economy was reversed and the quantity of heat utilized in refining the average barrel of crude oil rose to 672,000 B. t. u. There were undoubtedly a number of reasons for the apparent decrease in fuel economy at refineries in 1930, chief of which were the tendency to carry the distillation farther instead of merely topping the crude and a gain in the consumption of waste oils, such as acid sludge, which in actual practice yield less heat per unit consumed than has been theoretically assigned to oil. The consumption of natural as fuel at refineries has declined since 1927, but this has been more than compensated by a rapid increase in the use of refinery or still gas. The total consumption of oil as refinery fuel in 1930 was not much below the 1925 total, but the consumption of coal has declined rapidly, so that in 1930 liquid and solid fuels contributed only 56 per cent of the total heat units- utilized at refineries as compared with 77 per cent in 1925 (see fig. 2).

Jan 1, 1931

By A. J. Farstad, R. G. Olsen

The Coal Mine Health and Safety Act of 1969, passed by Congress in the wake of the disaster at Farmington, West Virginia, led to an extensive study by the National Academy of Engineering of the problems involved in coal mine safety. One of the results of this report and subsequent research conducted by the U. S. Bureau of Mines has been the investigation of electromagnetic techniques for the location of trapped miners after a mine emergency. The Westinghouse Georesearch Laboratory has been involved in both the 2 theoretical and field studies relating to this problem. A compact personal beacon transmitter has been developed which can be attached to and powered by a miner's lamp battery. A compatible lightweight loop antenna package has also been developed which can be easily carried on a miner's belt. This antenna can be quickly deployed on the mine tunnel floor. For the purpose of calculating its electromagnetic fields on the earth's surface, the loop can be considered to be a magnetic dipole buried in the earth. (The earth is assumed to be homogeneous and to have a conductivity of s mhos/m.) The surface of the earth is assumed to be uniformily sloping and the transmitter dipole oriented as in Figure 1. The fields on the earth's surface can be computed by using a superposition of the fields of a 4 vertical magnetic dipole3 and of a horizontal magnetic dipole in proportions which depend on the slope of the terrain. In the case of level terrain a null in the horizontal magnetic 5 field occurs on the surface directly above the dipole. This null is designated as the apparent source location and is the criterion by which the source is located. The criterion was tested at two hardrock mines and two coal mines in the fall of 1972. The apparent source locations were compared to actual locations determined by conventional surveying techniques. It was found that due to nonlevel terrain the apparent and actual source locations differed by up to 13 meters. This discrepancy can be resolved by an examination of the fields of a source beneath uniformly sloping terrain. Figure 2 shows the experimental and theoretical results for a

Jan 1, 1973

By George H. Ashley

Character The bituminous coals of Pennsylvania range from low volatile to high volatile, but are all low moisture coals The coals are of three types caking of coking, noncaking or splint, and cannel The percentage of volatile matter in all kinds changes from low at the east to high at the west This change is thought to be due to the pies sure and folding to which the coals along with the other rocks have been subjected In the Broad Top field, and in Somerset, Cambric, Clearfield, Centre, and other counties to the northeast, most coals contain from 3 to 5 times as much fixed carbon as volatile matter, the fixed carbon ranging from 67 to 75 per cent, and the volatile mattes from 16 to 24 per cent These coals are classed as lovol coals In the line of counties just west of those listed the coals are of the medium volatile class (midvols) This includes the famous Connellsville coking coal Still farther west, covering most of Washing ton and Westmoreland counties, and the counties to the north of them are high volatile coals of gas coals (hivols) This is discussed in Part I on the geology of the coals of Pennsylvania, where is given in isocaib map Cannel coals ale found in many par is of the bituminous coal field of Pennsylvania, but in particular in Beaver Butler, Armstrong, Indiana, and Clear field counties The cannels of Beaver and Butler. counties are typical high volatile cannels, many of which have been used in times past for oil distillation The cannels of Armstrong and Indiana counties are medium volatile cannels, and while main taming the physical character of cannels, have no more volatile neat ten than the coals around them The cannels of Clearfield County tie low volatile Splint coals of noncakmg coals are not abundant in Pennsylvania. The Sharon coal, formerly mined in Mercer County, was a noncaking or splint or block coal, and was formerly used in the raw state in non smelting Uses STEAM COAL Practically all of the coals of Pennsylvania are steam coals, though many coals in the Commonwealth, because of high ash, are not good steam coals The coal of the Pittsburgh district has long been famous

Jan 1, 1928

By C. J. Tsai, K. Olson, R. Caldow, B. Cantrell, D. Y. H. Pui

"The amount of new surface generated during fracture of coal and rock has been found to be directly proportional to the amount of energy associated with the fracture. To confirm this and study the relationship between the new surface and dust aerosol formed during fracture the U.S. Bureau of Mines has conducted experiments simulating fracture as it occurs during mining operations. In addition to fracture energy measurements, fragment size distributions were determined over a size range of 0.5 microns to 1 inch. The size measurements accounted for the mass of the original sample. Material studied was Illinois No. 6 coal, coal from the Taggart seam of Virginia, and Berea Sandstone. The experiment employed a chamber to confine sample fracture fragments and a combination of particle measuring systems. Measurement methods included aerodynamic sizing of entrained dust aerosol with optical counting and inertial impaction techniques, physical sizing of non-airborne fragments by sieving, and optical sizing of fragments using optical counting and scanning electron microscopy.Results of the size distribution measurements indicate that a single size distribution law does not adequately describe the entire fragment size distribution. A combination of two size distribution laws, such as the Rosin-Rammler and log-normal distributions, is necessary to describe this size distribution. New surface created by fracture is primarily found 1n particulate between 2 and 50 microns with a mean size of 12 microns. It was also found that the specific energy of fracture can be linearly related to the .concentration of aerosol in this size range.Only a .small fraction, between 0.02 and 0.16 pct, of the dust particles .generated during fracture become airborne. The remainder adhere to the new surface created during fracture. An entrainment fraction, defined for a given particle size is the ratio of the mass of airborne particles to that of the total particles generated in that size range. This experimental entrainment fraction is related to the Van der Waals adhesion force and the electrostatic forces operating during fracture and is used to derive empirical parameters to model these two processes."

Jan 1, 1989

By John C. Edwards

Several experimental programs were conducted at the Pittsburgh Research Center to investigate the level of mine fire detection and alarm capability possible using state-of-the-art technology. These programs involved comparison of the response and alarm time of optical and ionization type smoke sensors to smoldering and flaming coal combustion in a smoke chamber. Coal combustion experiments conducted in the smoke chamber demonstrated that a CO concentration of 5 ppm above ambient corresponded to an optical density of 0.022m1. For the four smoke sensors for which a continuous analog signal was available, a smoke sensor alarm was defined as the average background signal plus ten times the peak-to-peak noise. This alarm criterion resulted in the association of the alarm for three of the four sensors with a smoke optical density of 0.022m-1. and 0.033m-1 for the fourth smoke sensor. The smoke sensors used in the smoke chamber studies were incorporated into large scale diesel fuel fire experiments conducted under normal ventilation conditions in the Safety Research Coal Mine (SRCM) at the Pittsburgh Research Center. These experiments showed that a diffusion mode smoke sensor alarmed earlier than a diffusion mode CO sensor; and, based upon the measurement of the optical transmission at the location of a pump mode ionization type smoke sensor that a pump mode ionization smoke sensor alarmed at an average smoke optical density of 0.021m" for the twelve experiments conducted. A series of 30 and 330 kW diesel fuel fire intensity experiments were conducted under zero airflow conditions in the SRCM. It was determined that :(I) thermal sensors are inadequate at 30 m from a 300 kW fire;(2) a diffusion mode ionization smoke sensor can be more effective for mine fire detection than a diffusion mode CO sensor;(3) recommendations can be made for sensor spacing in a mine entry based upon the measured CO buoyancy induced spread rates along the mine roof and the time for a developing coal fire to ignite a conveyor belt.

By Steven J. Page, John A. Organiscak, Thomas Mal

Monitoring the miner respirable quartz dust exposure for compliance with U.S. federal regulations is mandatory to protect the health of coal miners. However, there is no technique available for the direct measurement of the respirable quartz dust concentration in the mining environment. The only means to estimate the respirable quartz dust concentration is to perform an analysis of the respirable dust sample collected, which contains both coal and quartz dust as well as other mineral constituents. For samples collected at coal mines, the U.S. Mine Safety and Health Administration (MSHA) utilizes the P7 method for respirable quartz analysis. The objectives of this work are to 1) evaluate the performance of the P7 method on field samples of low quartz mass loadings, 2) evaluate the magnitude of dust mass spatial variability within an enclosed cab environment, and 3) estimate the subsequent propagation of error in calculating the per. centage of quartz over a broad range of mass and quartz filter loadings at a surface coal mine. These objectives were achieved by obtaining surface coal mine dust samples over a 9-month time period in enclosed cabs of three drills and three bulldozers. By measuring weighing precision and using reasonable estimates for sampler variability, the dust mass spatial variability component was calculated by difference. The results show that 1) the P7 method can be precise and linear with quartz mass loadings as low as approximately 5 µmg of quartz, 2) the dust mass spatial variability, even within an enclosed cab environment, is the dominant factor of variability in propagation of error in estimating the quartz content of the obtained dust samples on a percentage basis, and 3) weighing precision is not a significant factor in error propagation. In addition to the enclosed cab samples, samples from eight surface coal mine drills were collected at various locations exterior to the drill cab. Significant data scatter at these outside locations strongly suggests that the ambient dust mass spatial variability has a significant effect on the propagation of error in estimating quartz content of the dust samples on a percentage basis.

By Frederick L. Hoffman

Miners' nystagmus is the result of a peculiar form of eyestrain, which often results in impairment of vision and a consequent diminu- tion of the wage-earning capacity of the workman. Historically it has been pointed out by Shufflebotham that "This disease was first recognized by Gillett, of Sheffield, about 1854, although no case was mentioned in medical literature until 1861, when the symptoms were described by De Conde, a Belgian physician. Cases were reported by Pepp Müller as early as 1860, by Von Graefe in 1873, and by Bell Taylor, of Nottingham, and by Nieden in 1874. In this country (England) we are indebted to a very large extent to the observations and writings of the late Simeon Snell and for his pioneer work on this many-sided disease; in France Dransart occupies a similar position." In the treatise on "The Effects of Arts, Trades, and Professions," by C. Turner Thackrah, published in London in 1832, it is stated that- The complexion of the colliers, even after the removal of the dirt, is generally sallow and unhealthy. Their eyes, from the swelling of the lids, appear small, are affected with chronic inflammation and intolerant of full light. * Sickness and vomiting sometimes affect persons at their commencing the employ; and many, after a few years' trial, are obliged, by the injury which their health has sustained, and especially by the weakness of their eyes, to leave the mine. According to Greer, an authority on industrial diseases and acci- dents, miners' nystagmus occurs in 5 per cent of all miners, but it is quite probable that the reference is to European coal miners rather than to men employed in all branches of the mineral industry. According to the same authority, miners' nystagmus "is associated with a twitching of the eyelids and with considerable diminution of the range of movement of the eyes, and is particularly liable to occur in patients whose vision is subnormal owing to errors of refraction, especially astigmatism. It causes much inconvenience to those suf- fering from it and prevents them from following their usual occupa- tion as, owing to the movements of the eyes, everything at which they look appears to be in a state of perpetual motion, and in conse-

Jan 1, 1916

By Jeanne A. Zimmer, Charles D. Taylor, Edward D. Thimons

On a continuous miner coal face methane concentrations are affected by the quantity of fresh intake air reaching the face. When mining the box cut on a 12.2-m (40-ft) two-pass extended cut face, the continuous miner is always at the point of deepest penetration, and operation of the machine-mounted dust scrubber and water spray system improve the flow of fresh air to the end of the box cut. However, when the continuous miner leaves the box cut to begin cutting the 40-ft slab, little is known about how much ventilation air reaches the end of the box cut. The Pittsburgh Research Laboratory (PRL) of the National Institute for Occupational Safety and Health conducted a full-scale study to answer this question. Tests were run to determine how much ventilation air reaches the end of the box cut with the continuous miner at three locations in the 40-fttwo-pass extended cut mining sequence; at the end of the 40-ft box cut, at the start of the 40-ft slab cut, and 6.1 m (20 ft) into the slab cut. During these tests, methane gas was uniformly released at the end of the 40-ft box cut. Methane concentrations measured0.3 m (1 ft) from the roof and face of the box cut were used to estimate face airflow quantities. Fresh air flow rates of 1.89and 4.72 m/sec (4,000 and 10.000 cfm), scrubber flow rates of 0, 4,000 and 10,000 cfm, water spray pressures of 0 and3827 kPa (0 and 120 psi), and a blowing curtain setback position of 15.2m (50 ft) were used during these tests. To improve ventilation to the end of the 40-ft box cut when the continuous miner was starting the 40-ft slab cut and 20 ft into the slab cut, the blowing curtain was advanced from its 50-ft setback to locations 40-ft and then 8.5 m (28 ft) from the face. KEYWORDS: Face ventilation, Extended cut, Methane, Box cut, Continuous miner, Scrubber

By F. M. Goodwin, John Campbell, G. W. Jones

"INTRODUCTION In previous reports 5 results were given of a manhole investigation carried out over a period from June 1929 to July 193C in cooperation with the Edison Electric Illuminating Co. of Boston. Mass., and during 1931 in cooperation with the Edison Electric Illuminating Co. of Boston and the Boston Consolidated Gas Co., to determine the general hazards of combustible gases and vapors in manholes and other underground openings owned by the Edison Electric Illuminating Co. and to develop methods of eliminating these hazards. These investigations have demonstrated that the detection, control, and elimination of combustible hazards in manholes can best be obtained by a close cooperation of the interested parties through a systematic and continuous inspection of manholes, under proper supervision.The investigation carried out during 1932 has been along lines similar to those conducted during 1929, 1930, and 1931 test crew consisting of two men, one an employee of the Consolidated Gas Co. and the other an employee of the Edison Electric Illuminating Co., was continuously employed during the year in the testing of gases in manholes served by the Edison system. The test crew was under the direct supervision of John Campbell, superintendent of the special service department of the Edison Electric Illuminating Co., and F. M. Goodwin, vice president of the Consolidated Gas Co.As in the previous investigations, all tests were numbered. The test-data sheets, on which the information regarding the tests was recorded, were prepared in triplicate, ono copy going to the Edison Co., one to the gas company, and the third to the Bureau of Mines at Pittsburgh, Pa. These test-data sheets were then filed (1) according to the town, (2) the street location, and (3) the manhole number; in this way a complete record of each manhole tested was immediately available. A typical test-data sheet was given in the 1931 report. These data sheets were found to entirely satisfactory and no changes were made during the 1932 investigation. The procedure of sampling manhole atmospheres, shipment of the samples to Pittsburgh, and method of analysis has been the same as that described in the 1931 report and need not be repeated here. Likewise, there has been no change in the method of calculation of explosibility of the samples analyzed or the use of charcoal filter tests given in appendix A."

Jun 1, 1933

By X. Shi, N. S. Dala, V. Vallaythan

"Cellular damage induced by silica dust is known to be associated with increased fragility and destabilization or destruction of the cell membranes. The mechanism by which silica dust effects biological membranes is still unknown. Using the ESR spin-trap methodology, we find that hydroxyl radicals are generated in aqueous suspensions of freshly ground silica. The amount of the radicals generated depends on the extent of grinding and ""freshness"" of the dust. Using linoleic acid as a model lipid, we show that the silica dust causes lipid peroxidation via ·OH, 02-, or linoleic acid related radicals as intermediates. Catalase and superoxide dismutase inhibit the formation of these radicals. The results suggest that free radicals associated with freshly ground silica and the related lipid peroxidation of the memberanes may be the starting point of the silicotic processes.INTRODUCTIONDespite considerable-effort over the years, the mechanism by which the quartz particles exert their toxic action on cells and the process(es) by which these actions progress to fibrosis are still not fully understood (1,2). It is generally thought, nevertheless, that the interaction of the quartz particles with the cell membranes is the starting point of the silicotic process (3). We felt that the mechanism of the membrane damage by quartz might involve oxygenated free radicals because (a) a suspension of quartz particles in contact with alveolar macrophages has been reported.(4,5) to initiate an enhancement of lipid peroxidation, defined broadly as the ;oxidative deterioration of polyunsaturated components of lipids, and (b) hydroxyl (·OH) radicals are known to be capable of peroxidation by abstracting hydrogen atoms from cell-membrane lipids (6) and initiating lipid peroxidation in lysosomal membranes (7). Moreover it is known that exposure of cell membranes, fatty acids and unsaturated food oils to ionizing radiation, which generates ·OH. radicals, causes rapid peroxidation (6). Earlier studies o~ the aqueous chemistry of quartz suspensions have reported detection of HzOz (8), implicating the formation of ·OH radicals as transient species, but, we are not aware of any report of the detection of ·OH radicals in quartz suspensions and this provided the motivation for the present undertaking. Since it is known that, because of their high reactivity (hence short life time) in aqueous media, the ·OH radicals cannot be detected via electron spin resonance (ESR) directly (9,10), we have used ESR combined with the spin-trap methodology (9) for studying the •OH formation."

Jan 1, 1988

By H. S. Chiang, S. S. Peng

"SummaryAir velocity distribution surveys on four mechanized longwall faces are presented and discussed. Although there are discernible patterns in the velocity distribution, there are wide differences in distribution in the cases examined. This can be attributed to differences in mining and face geometry. An important factor is air leakage into the goaf.IntroductionOver the past two years several intensive air velocity surveys have been carried out in West Virginia coal mines as part of a programme aimed at understanding dust transport mechanisms on mechanized longwall faces. These are presented here and some of the patterns of velocity distribution are discussed.Ventilation survey methodologyThe air velocity vector consists of three mutually perpendicular subvectors; the vertical subvector is directed toward the roof, one of the horizontal subvectors is parallel to the coal face and the other pointing to the gob. The air velocity which was measured in this paper is the horizontal subvector parallel to the coal face.To record the velocity distributions at the longwall faces, the surveying stations and the surveying points were established as follows.(1) The first station usually was set at the third support from the headentry T-junction. Within the first few supports, the air coming from the intake airway encounters a sharp right-angle turn at the T-junction, and forms a turbulent region close to the coal face. The intake air in this region is also influenced by the motors that drive the face conveyer. But when the air reaches the different. But for the same face the number was the same. The main air-travelling area was divided into six to 12 equally sized sub-areas, and each surveying point was situated approximately at the centre of the sub-area.The velocity survey was carried out by two persons, one reading the meter and the other taking down the readings. A complete survey at one station took about 5-8 min. Usually, the dimension of the cross-sectional area was measured during that time.A third person used the smoke tubes to observe the following: (I) the air movement patterns at different locations; (2) the extent and the direction of the air leakage; and (3) the machines' influence on the main airflow. At the same time he observed and recorded the roof conditions and the compactness of the roof-caving in the gob."

Mar 1, 1989

By W. J. Fene, H. B. Humphrey, J. J. Forbes

"INTRODUCTION Loss of life from coal-mine explosions in the United States in 1949 was the lowest in any year since before 1870. Although fatality reports from coal nines before 1910 were incomplete,' the explosion fatalities from single disasters or individual States exceeded the 1949 total in each year as far back as 1870, when the records become fragmentary. Since 1890, the only other year in which fewer than 35 were killed was 1946, with a total of 32. Figure 1 shows the total yearly fatalities from explosions in all coal mines and those in anthracite mines alone since 1870. The fatalities in bituminous coal mines were not completely reported before 1910, but enough reports were issued by major coal-producing States to compile reasonably representative figures back to 1885. Any corrections would be additions to these totals.The reason for the moderate total of 46 explosion fatalities in 1948 and the low total of 8 killed in 1949 is thought to be the serious and careful attention devoted to elimination of coal-mine explosion hazards after the disastrous explosions in 1947. The standards now being observed generally with respect to ventilation, rock-dusting, and inspections for gas are better than those found earlier in even the mines having the best reputations for safety. Exceptions are still numerous, but they are fewer, month by month, because of the pressure of Federal and State inspections and a changing conception of the hazards that should not be permitted underground. The prevailing precautions have not eliminated all ignitions of gas or dust, but they have limited propagation to small explosions and have prevented serious loss of life.The number of ignitions (14) reported for the year was less than the average (24) for the 5 years 1944-48 and only half the average for the fiscal years 1929-44. Another favorable factor is that Bureau of Mines representatives were probably in closer contact with both small and large mines in nearly all coal-mining districts. Inspectors now assigned to individual districts are better able to keep informed on the happenings and the progress of safety in their districts. In 1948, one-fourth of the explosions were in mines that employed fewer than 10 men; this condition was duplicated in 1949."

Jul 1, 1950

By L. C. IlsLey

One of the details which the engineers observe closely in conducting tests on explosion-proof type equipment at the Pittsburgh experiment Station of the Bureau of Mines is the construction of covers and their means of fastening. These covers are of particular concern from a safety standpoint, because they are very liable to be frequently disturbed for the purpose of making inspection or repairs to internal parts of motors, controllers, etc. When depended upon to permit frequent inspections it is desirable from a practical operating standpoint that the cover be easily and quickly removed and replaced. The position that been taken by the Bureau of Mines is that the questio of ease with which covers will permit inspection should be determined by the mamu- facturer and the user. If the scheme used requires considerable effort, as in the removal of several bolts, it has been assumed that the one going to all this trouble to get the covers off could be trusted to put the cover back in place properly, especially as the approval plate generally calls specific attention to this point. Where a cover is simply screwed in place, fastened by a single bolt, or secured by a hing in or other similar arrangement, the bureau has not been willing to assume that the, cover would in all cases be properly fastened and has required further security by means of a lock or seal. In the course of its investigations the bureau has had occasion to inspect and test five types of covers as follows: (1) Covers having two plane surfaces bolted together. (2) Covers having two plane surfaces held together by lugs, clamps, or other means besides bolts. (3) Covers which have screw thread and fit a similar thread in the compartment. (4). Covers having plane surfaces with the addition of a tongue-and-groove joint at about the center of the plane surface and in general secured by hinge pins or similar means of fastening. (5) Covers in the form of an annular ring accurately fitted to a motor shell which can be rotated around the shell, and thus make easy access to inspection openings in the shell and rotated back again to effectively seal these openings. Tests. have also been made of large bolted covers having a smaller and more accessible cover within the large one, which could be opened for inspection purposes.

May 1, 1930

Objective A cable bolt can be a strain-measuring device as well as a supporting device in underground mines. Placing instruments on a cable bolt allows certain characteristics of a rock mass, such as high strain and load, to he evaluated during mining and provides a means for mining engineers and mine inspectors to forecast a potential roof fall. Background Thousands of cable bolts are installed in U.S. mines each year. Studies performed by NIOSH researchers show that support loads can exceed the yield point of the steel, which can result in unstable ground conditions that pose direct safety hazards for underground workers. Rock falls can also create secondary hazards, such as caved escape ways and blocked ventilation passages. Development of methods and tools to assist in the evaluation and selection of rock supports could reduce the number of unplanned rock falls significantly. Approach Instrumented cable bolts developed at NIOSH’s Spokane Research Laboratory were used along with existing ground control systems to monitor rock mass loads and movement Axial and shear loads on the cables are determined as these loads are transferred to strain gages embedded in the king wire. In longwall, room-and-pillar, or underhand cut-and-fill mines, rock mass loads and movements can be detected down to microstrains of stress or millimeters of movement. By detecting such small amounts of ground movement and support load, early intervention to prevent a roof fall is possible, and miner safety and health are increased. How it Works Up to 10 strain gages are embedded in the king wire at whatever locations a mine ground control engineer may specify. The gages m either be set in pairs in the same plane to detect shear or set at 90° angles to detect loading in three dimensions. The cables are up to 5 meters long if they arc to be grouted with resin or up to 15 meters: if they will be grouted with cement. They arc capable of holding m ultimate lend of 214 kN and produce reliable readings to 178 kN, as shown in the calibration curve. The cables are fitted with an instrument plug for monitoring the gages and a special head for insertion into the rock. A data acquisition system, such as a Vishay strain indicator box, Campbell Scientific, or Omnidata system, is required to obtain the data for load and displacement calculations.

Nov 1, 1998

By Peter Kovalchik, Ed Kwait, T. Michael Durr

1. BACKGROUND Many research projects on engineering controls for noise reduction in mining were conducted by the former U.S. Bureau of Mines, mostly under contracts from 1972 through 1986.1 Contracts were awarded to determine which mining equipment was responsible for worker noise exposure over the permitted Federal Regulation level of 90 dBA for 8 hours. Dominate noise sources, for each piece of equipment was also determined. The final goal was to develop engineering noise controls for those pieces of equipment and then evaluate them. This was the general approach taken in each research project or contract. This paper deals with what is known about the control of noise from continuous miners, one of the most commonly used underground mining machines. A literature review on continuous miners revealed that the relevant noise sources were divided into four categories: conveyor, cutting head, drive train, and motor noise. The dominate noise sources reported were the conveyor noise and the cutting noise.2 Conveyor noise consists of impacts and scraping at various locations on the conveyor structure. These impacts are associated with discontinuities on the top and bottom decks, impacts between the chain and the tail roller, impacts on the sidewalls at path points of discontinuity under horizontal pivot, and impacts on the flex plate.3 Scraping occurs along the entire length of the chain conveyor. The other dominate source, the cutting of coal, has also been addressed in past research.4, 5, 6 The research presented in this paper only addresses the controls associated with the conveyor. The literature suggests several effective treatments for noise control: retrofitting the chain conveyor with a urethane coating or a urethane sleeve on the chain flights, using a urethane coating or urethane sleeve on the tail roller, constrained layer damping of the decks, and the use of a modified take up plate.7, 8 However, the in-mine durability of these concepts has yet to be proven for the conveyor system. This paper analyzes extensive noise and urethane durability testing on a Joy 14CM9 continuous miner’s conveyor assembly, tail roller, and bottom plate conducted at the Pittsburgh Research Lab facilities. It will also discuss the initial underground testing results for the urethane-coated conveyor flights and tail roller.

By Linda L. Chasko, Ronald S. Conti, John D. Cool

Mine operators often rely on mine rescue teams to save lives during an underground emergency such as an underground fire, explosion, roof fall or water innundation. It is extremely important that team members are provided with adequate exploration equipment and that training simulations are conducted in a realistic manner. A series of mine rescue training exercises was developed, conducted, and evaluated by the National Institute for Occupational Safety and Health (NIOSH) in cooperation with the Pennsylvania Department of Environmental Protection, Bureau of Deep Mine Safety and several mining companies. The exercises were conducted at NIOSH’s Lake Lynn Laboratory and operating mines during 1995 to 1999 and resulted in improved technology and training for mine rescue teams. For example, existing technologies were identified to help rescue teams during exploration. These included various chemical light shapes and laser pointers to identify team members and mark crosscuts and mine materials, strobe lights for mapping out escapeways and marking crosscuts, and thermal imaging systems to see through smoke. A hands-free communication system showed potential for enhanced communications between team members and the fresh air base. Anew team lifeline that allows for flexibility of movement between team members was developed. A positive-pressure inflatable escape device (IED) was used to isolate the “hazardous” environment from fresh air and allow rescue team members to traverse through. An inflatable feed-tube partition that can rapidly block large openings such as underground passageways and simultaneously provide a feed-tube for high-expansion foam generators was also deployed for several simulations. During the simulations, mine rescue team members donned self-contained breathing apparatuses (SCBA) and traversed more than 400-m of mine passageways filled with nontoxic smoke. Visibility ranged from 0.3 to 0.9-m. In one area of the mine, a kerosene heater was used to simulate a fire and reduce the oxygen concentration. In another area, a 1.5 pct methane zone was established. The team members searched for "victims,” administered first aid, mapped the passageways, built temporary stoppings, and ventilated the smoke-filled entries. Advanced training included the extinguishment of a conveyor belt fire in a surface fire gallery. Two teams, each carrying charged waterlines, were escorted into the fire gallery through both doors. These teams controlled the rollback smoke and extinguished 7 to 9-m of burning conveyor belt. Team members indicated that the exercises were extremely beneficial due to the realism of thetraining simulations.

This book has employed miners' narratives to illustrate basic concepts about the escape process. One of the most powerful means by which people make sense of their experiences is through the telling and internalization of stories [Bruner 1990]. By couching one's own and others' motives and actions in terms of a coherent narrative, a person is able to learn from mistakes and plan future behaviors that may help ensure survival. A growing body of research suggests that decision-making skills used to deal with emergency situations can be taught and assessed by simulations based on narratives from the real world [Bransford et al. 1986; Brecke 1982; Brener 1984; Connolly et al. 1989; Halff et al. 1986; Jones and Keith 1983; Lacefield and Cole 1986]. Such techniques have been used to address the decision-making of medical personnel, civil and military flight crews, and even people involved in broader life events such as political and military situations [Babbott and Halter 1983; Dugdale et al. 1982; Farrand et al. 1982; Gilbert 1975; McGuire 1985; McGuire et al. 1976; Flathers et al. 1982; Giffin and Rockwell 1984; Jensen 1982; Janis and Mann 1987]. Given the validity of this method of study and the promise it holds for helping people improve the quality of their responses to nonroutine occurrences, it is perhaps surprising that there have been no studies of emergency decision-making among blue-collar workers prior to those conducted by the present authors and their colleagues. The purpose of this chapter is to describe underground coal miners' decision-making performance on a table-top simulation whose problem structure is derived from interviews with a group of eight miners who escaped from the 5 Left section at Brownfield Mine. The exercise was constructed by a panel of domain experts (mine safety and rescue personnel) with the assistance of an educational psychologist. The simulation includes actual predicaments with wise and unwise decision alternatives that, in the opinion of these domain experts, are characteristic of such escapes. Results reported in this chapter are the scores of a sample of experienced mine workers who completed the simulation. Because the exercise is a series of objective performance tasks coupled with detailed and immediate feedback, this simulation can be used to teach and refresh critical escape skills, as well as to provide data concerning the proficiency of miners at the time of exercise administration.

Jan 1, 2000

By Charles A. Kumke, GROVER J. DUFF

This paper describes mining and milling practices at the San Xavier mine of the Eagle- Picher Mining & Smelting Co. in Pima County , Ariz . It is one of a series being prepared by the Bureau of Mines on mining methods in various mining districts of the United States . The mine and the mine surface buildings are situated in sec . 11 , T. 17 S. , R. 12 E. , in the Sierrita Mountains in Pima County , 20 miles south by southwest from Tucson via the San Xavier Mission and Twin Buttes road ( fig . 1). A well-maintained dirt road continues southward toward Twin Buttes for 4 miles and then branches off northeasterly for 5-1/2 miles to Sahuarita , where the concentrator of the Eagle - Picher Mining & Smelting Co. is situated on the west slope of the Santa Cruz River Valley . Tho mill is connected by a railroad branch to the main line of the Southorn Pacific Railroad between Tucson and Nogales . It is generally assumed that the San Xavier silver - lead deposits became known to the Jesuits and Spaniards in the early part of the eighteenth century . However , no extensive mining operations were conducted by the Jesuits in what is now Arizona before the middle of the eighteenth century . The completion of the Southern Pacific Railroad from Los Angeles to El Paso in 1881 marked the beginning of intense activity in mining throughout the southern part of the State . The San Xavier mine produced 6,000,000 pounds of lead , considerable zinc ( exact figures not available ) and $200,000 in silver from 1897 to 1918 . The Eagle- Picher Mining & Smolting Co. re- opened the San Xavier mine in 1943 and built a 175-ton concentrator at Sahuarita . The capacity of the plant was increased in 1944 to 500 tons in order to take in custom ore . The surface rocks at the San Xavier mine are , in great part , of sedimentary origin . Limestones of various ages comprise most of the exposed sedimentaries , some quartzites and shales also being present . The sedimentaries rest upon coarse - textured , intrusive , igneous rocks comprised mainly of granite and granodiorite . The San Xavier deposit is composed of galena , sphalerite , chalcopyrite , and pyrite . Generally the individual ore bodies are small , discontinuous , and occur as irregularly spaced shoots or lenses . The present mine workings extend horizontally a distance of 2,700 feet and vertically 900 feet . Owing to ground weight, all mine openings are timbered on the levels and all stopes are square- setted and waste -filled .

Sep 1, 1950

In the U.S. coal mining industry, a large cohort of miners is at or nearing retirement age. As these miners retire, there will be a need to hire and train thousands of new and most likely younger workers. Moreover, a large portion of safety trainers will also require replacement due to retirement. In a speech before the U.S. House of Representatives (2004), Bruce Watzman, Vice President for Safety, Health and Human Resources for the National Mining Association said, “[W]e will need to replace a major portion, approximately 50%, of the underground coal mining workforce within the next 5–7 years.” The large influx of new mine workers, along with the deficit in safety trainers, creates the challenge to provide consistent and effective safety training. Given the expected high turnover of the mining workforce, it is very likely that incoming workers will have much less experience than those leaving the workforce. Unfortunately, inexperienced miners have been shown to incur the highest rate of injuries. The piece of machinery most associated with these injuries is the roof bolting machine, which accounts for 56% of machinery-related injuries in underground coal mining. The National Institute for Occupational Safety and Health (NIOSH) has released a new report entitled “Roof Bolting Machine Operators Skills Training for a Walk-Thru Roof Bolter: Trainer’s Guide” (IC 9489). It is designed to help safety trainers develop struc-tured training for new operators of walk-thru bolting machines. This resource will help trainees learn, understand, and apply knowledge and skills. Included in the trainer’s guide are the following tools: [ ] a six-part DVD video series, a skills check survey, talking points, and a complete job training analysis. Also included is an exercise to help trainees learn more about their mine’s roof control plan, as well as a list of supplemental training and reading materials. On-site trainers can modify this guide to fit their particular mine conditions, machines and equipment, and work procedures. The goal of every coal mine safety professional is to increase safety awareness and make mines a safer place to work. The Walk-Thru Bolting Machine Trainer’s Guide will help teach new miners skills to create a safer working environment that will ultimately prevent or reduce on-the-job injuries. Safety trainers can also use the ideas and concepts from this guide to develop their own training skills for the safe operation of other pieces of underground coal mining equipment.

Jan 6, 2007