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By Thomas M. Barczak, Joe Bower, Jinsheng Chen

Pumpable roof supports provide an alternative longwall tailgate roof support and have grown in usage during the past few years. Heintzmann Corporation has been installing pumpable roof supports at the RAG Resources Emerald Mine in western PA since 1998, where they have provided effective roof control in their longwall tailgates. Despite the success of these supports in this application, questions remain regarding critical design issues for optimizing the use of this support technology. The support loading profile is characterized by a high initial stiffness with peak loading occurring at less than one inch of convergence, followed by significant load shedding with a post failure capacity comparable to that of wood cribbing. Therefore, the key to optimizing the support utilization is to provide sufficient load density to prevent convergence from occurring beyond the peak loading capacity. This requires an understanding of how the supports interact with the ground conditions, hence measurement of the ground reaction curve. In order to obtain this information, pumpable roof supports were instrumented to measure support loading. Roof deformation and roof-to-floor convergence measurements in the vicinity of the instrumented supports were also made. The experimental parameters for the installation were the support spacing and water-to-solids ratio of the grout, which controls the grout strength and ultimately the maximum capacity of the support. The study clearly shows that at this mine a 24-inch diameter support is fully capable of providing adequate ground control under depths of cover of 750 ft as only 50 pct of the available support capacity was utilized outby the longwall face. It was also shown that the 2.00 to 1 water-solids ratio, despite providing a slightly softer support response, is sufficient for maintaining the same degree of roof control provided by the traditional 1.75 to 1 grout mix. The 10-ft spacing of the supports did not cause any ground control problems outby the face. However, inby the face the performance of the support is degraded once the peak capacity of the supports is exceeded. It appears that the large load shedding behavior, which is characteristic of this support following peak loading, allows the immediate roof to separate. When this happens, the support is unable to regain control of the roof in time to prevent failure of the immediate roof beam.

By C. H. Larsen

The results of a comprehensive investigation of conveyance arrestment devices (safety catches) in United States coal mines are presented. Availability and use, performance, inspection, maintenance, testing, procedures, and regulatory requirements were investigated. Information was obtained from equipment manufacturers, mine operators, and regulatory-agency personnel, as well as from studies of domestic and foreign mining safety regulations. Traditional mine hoists and commercial elevators for mine use were investigated. Results of this and an earlier study of wire-rope practice in United States coal mines show a continued need for conveyance arrestment devices. A wide variety of these devices are in use. Only two types have been proved reliable for stopping a conveyance without injuring passengers or equipment. These are elevator overspeed-actuated wedge-clamp (Type B) safeties and broken-rope-actuated devices for wooden guides using a single-tooth-dog design developed in Ontario as a braking element. Elevators are widely used in U. S. coal mines. When maintained according to American National Standards Institute (ANSI) Specifications A17. I and A17. 2, they appear to present no major arrestment problems. Of 42 coal-mine hoists used for man transport that were inspected, 30 are in vertical and 12 in inclined shafts. Twenty-four vertical shafts use wood guides; 6 use steel. The 12 inclined-shaft systems use flanged-wheel T-rail guidance. Two wood-guided vertical-shaft systems use the single-tooth Ontario-type dog; the other 22 use multi-toothed wedges, cams, or dogs that have been proved highly unreliable. The arrestment devices in the six steel-guided vertical-shaft systems and in six of the inclined-shaft systems cannot be evaluated because performance data are lacking. The remaining six inclined-shaft systems have no arrestment devices at all. Of the 36 hoists with arrestment devices, the devices of 16 are never tested, the devices of 10 are actuated periodically with the conveyance at rest, and the devices of the other 10 are actuated periodically with a slow-moving conveyance. Although high speed is as probable as low speed in an emergency and would provide the best test of arrestment capability, no tests are performed in the United States which demonstrate the ability of an arrestment device to stop a loaded conveyance from high speed. Mine operators indicate little confidence in their conveyance arrestment devices, and manufacturers recognize a need to improve their products. Neither state nor Federal regulations provide a guide to this improvement. Mine operators, manufacturers, and state and Federal regulatory agency personnel would welcome specific guidance. In view of this, a set of application, design, actuation, stopping-performance, and testing and inspection criteria is presented as a guide to eliminating the inadequacies of these devices. Recommendations are included for improving hoist controls and machinery directly affecting conveyance safety.

Jan 1, 1972

By Branch of Editorial Services

"The Bureau of Mines was established in the public interest to conduct inquiries and scientific and technologic investigations concerning mining and the preparation, treatment, and utilization of mineral substances; to promote health and safety in the mineral industries; to conserve material resources and prevent their waste; to further economic development; to increase efficiency in the mining, metallurgical, quarrying, and other mineral industries; and to inquire into the economic conditions affecting those industries. The organic act of the Bureau, as amended by Congress and approved February 25, 1913, made it the province and duty of the Bureau of Mines to ""disseminate information concerning these subjects in such manner as will best carry out the purposes of this Act.""In accordance with this directive, the Bureau reports the findings of its research and investigations in its own series of publications and also in articles that appear in scientific, technical, and trade journals; in proceedings of conventions and seminars; in reference books; and in other non-Bureau publications. The number of these reports, the wide range of subjects they cover, and the variety of mediums in which they appear make the kind of list and index presented in this special publication both necessary and valuable. This issue of the Bureau of Mines List of Publications and Articles describes reports and articles published during the period January 1, 1980, to December 31, 1984. It supplements the 50-year list of Bureau publications issued from July 1, 1910, to January 1, 1960;1 the 50-year list of articles by Bureau authors published outside the Bureau from July 1, 1910, to January 1, 1960;2 the 5-year lists of publications and articles from January 1, 1960, to December 31, 1964,3 from January 1, 1965, to December 31, 1969,4 from January 1, 1970, to December 31, 1974,5 and from January 1, 1975, to December 31, 1979;6 and the annual lists of Bureau publications and articles from January 1 to December 31, 1980,7 from January 1 to December 31, 1981,8 from January 1 to December 31, 1982,9 and from January 1 to December 31, 1983. 10"

Jan 1, 1985

By Dennis R. Dolinar

In general, the direction of the maximum horizontal stress in the eastern United States is fairly well defined. However, the variation of the magnitudes of the horizontal stresses is not very well understood. Because the horizontal stresses cause severe ground control problems in underground coal and limestone mines throughout the eastern United States a more complete understanding of how the magnitudes vary would be useful for developing mine design strategies to combat horizontal stress related ground control problems. Therefore, in this National Institute for Occupational Safety and Health (NIOSH) study, the variation of the magnitude of the horizon stresses in sedimentary deposits in the eastern and Midwestern United States are examined with respect to two factors, the elastic modulus of the rock and the site depth. Stress measurements from thirty-seven sites are used in the evaluation. Examining the applied excess strains indicates that the eastern United States can be, separated into high and low strain zones. For most of the eastern United States the maximum applied excess or tectonic strain ranges from only 300 to 550 micro strains. However, there is one area, a portion of the Beckley seam in the central Appalachian region where the strains are significantly higher than the other regions. In this higher strain zone, the maximum applied tectonic strains range from 700 to 1,000 micro strains. Regression models for each zone based on the elastic modulus can explain between 83 to 85% of the variation of the maximum horizontal stress. Because one region, the northern Appalachian district, has strains that are about 20% higher than the other regions in the low strain zone, multiple strain models based on geographic regions were developed for the low strain zone that can explain 87 to 91% of the maximum horizontal stress variation with the elastic modulus. Depth was found not to be a significant causal factor in any increase in the horizontal stress even though the site depths ranged from 275 to 2,500 ft. Beyond a theoretical increase, based on Poisson’s effect and gravity, no other increase in the horizontal stress with depth can be justified with this data. The most significant factor controlling the variation of the maximum horizontal stress is the elastic modulus of the rock, not the overburden depth.

By R. M. Strong, Lauson Stone

Under the terms of the act establishing the Bureau of Mines, this bureau was authorized to carryon the work of testing and analyzing fuels which had been conducted by the technologic branch of the United States Geological Survey. That work included in its scope an investigation of the availability and uses ofliquid as well as solid fuels, for the original outline of the fuel-testing investigations contemplated, as soon as the funds would be available, a study of the liquid-fuel resources of the country and the making of related researches to determine how these resources could be utilized with greatest efficiency. Owing to the fact that many difficulties were being encountered in the adaptation of the heavier fuel oils for convenient use in internalcombustion engines, it was deemed best to begin the investigation of liquid fuels with tests of gasoline, a fuel in more or less general use. When this investigation began, the extensive introduction, especially by foreign powers, of liquid fuels for small naval craft had awakened much interest. However, the quality of gasoline was reported to vary materially in different countries and the quantity available was said to be rapidly decreasing, with the probability of a prohibitive increase in price. At the same time the claim was made that denatured alcohol, of fairly uniform quality, could be procured in all parts of the world, that unlimited quantities could be readily produced at a low cost, and that this fuel could be used much more efficiently than gasoline in internal-combustion engines. Such statements naturally led to a widespread belief that the time was near at hand when denatured alcohol would entirely displace gasoline as engine fuel. Therefore, the first investigations of the liquid mineral fuels logically embraced a careful series of comparative tests of gasoline and denatured alcohol in engines. A series of over 2,000 such tests was conducted at the Government fuel-testing plants at St. Louis, Mo., and Norfolk, Va., details of which are given in the following pages. The report is published by the Bureau of Mines because of the transfer of the fuel-testing investigations to this bureau.

Jan 1, 1912

Objective To reduce dust generated from soiled floors within enclosed operator cabs. Background Enclosed operator cabs are widely used on mobile excavation equipment in the mining and construction industries. They protect the operator from bad weather, noise, and airborne dusts. The inside cab environment usually must be kept at very low dust concentrations because of quartz dust generated from the excavation of silica-bearing rock. For cabs to be effective in controlling airborne respirable dust, they must have an efficient air filtration system, while providing a positive pressurization of the cab interior. Ideally, the recirculated and exterior makeup air filtration should provide at least 99% capture efficiency for dust particles as small as 0.3-µmaerodynamic diameters. Many of these enclosed cabs use heating units mounted or directed along the floor (see figure 1), which can cause dust entrainment problems. Prior studies by the National Institute for Occupational Safety and Health (NIOSH) of enclosed cabs have shown that these floor heaters can generate notable amounts of dust within the cab enclosure (see Technology News 486). Multiple-shift dust levels measured inside a drill cab increased on average from0.04 mg/m3 during the summer months to 0.68 mg/m3 during the winter months. Additional dust level measurements, made with optical particle counters during a nondrilling time period, verified that dust levels inside the cab increased from 0.03 to 0.26 mg/m3when the floor heater/fan was turned on. Dust levels within the cab also became greater than the ambient air dust levels outside of the cab when the floor heater was turned on during the nondrilling time period. This shows that the floor heater was a notable dust source problem. Cab floors are commonly soiled from operators tracking dirt inside the cab upon entering from the mine or construction site. Interior cab dust levels can be increased by (1) airflow disturbance of the soiled floor or (2) operator disturbance of the soiled floor. Approach During prior NIOSH surveys of surface coal mine drills, several drills with very low cab dust levels (#0.1 mg/m3) were observed to have floor sweeping compound spread out over the cab floor during the working shift. This initial observation led to a more controlled A-B study of the effect of the sweeping compound on the interior cab dust levels. This study involved dust sampling with and without sweeping compound inside the cabs of a bulldozer and two drills. Sweeping compounds commonly use sawdust or cellulose as the main bulk material, with oil or wax added for dust adhesion purposes. Sand is sometimes added to increase bulk density. Sweeping compounds with sand are commonly used to sweep up concrete floors. Those without sand (usually referred to as "gritless") are used to sweep up smooth finished floors. Most of the oils and waxes used for the adhesive ingredient are petroleum-base and can have an irritating odor. People sensitized to petroleum distillates could have allergic reactions to these sweeping compounds. A few companies offer non petroleum-base sweeping compounds, using either a natural oil or chemical additive for dust adhesion. During this study, it was decided that only the gritless sweeping compound (without sand) would be used, eliminating the addition of another potential silica source inside the cab. Since the sweeping compound would be applied on the enclosed cab floor for the duration of the shift, it was decided to use a canola oil-base sweeping compound to eliminate irritating odor from the petroleum-base sweeping compounds. The natural canola oil-base sweeping compound had a slight woody scent. The three pieces of equipment used in this study had notably different cab layouts. The bulldozer had a small cab with vents discharging air-conditioned or heated air from adjustable louvers at operator waist level and floor level. The bulldozer also had a cloth fabric seat and a carpeted floor. Drill A had a large cab with the heat directed from adjustable louvers at operator knee level and the air-conditioned air directed from cabin ceiling louvers. This drill had a vinyl fabric seat and a steel plate floor. Drill B had a small cabin with the heat directed across the floor (see figure 1) and the air-conditioned air directed down from the ceiling. It had a vinyl fabric seat and a worn rubber mat on a steel floor.

May 1, 2001

By E. S. Chalpin

On 25 August 1971, the U.S. Bureau of Mines contracted with the Aerojet Liquid Rocket Company to perform an engineering research program to be completed on 25 May 1972. The program objective was to evaluate the advantages and disadvantages of the underground use of several power sources including but not limited to a closed cycle diesel, electric cable/motor, rechargeable battery/ motor and exhausting diesel for mine shuttle cars, with emphasis on health and safety. This report, Research in Advanced Power Systems for Mining Health and Safety consists of four (4) main parts: (1) Hazards and Safety Analysis of Mine Power Systems, (2) Evaluation of Closed Cycle Diesel Power System for Mine Applications, (3) Closed Cycle Diesel Power System Consumables Compatibility, and (4) Concluding Remarks and Recommendations. The hazards introduced to a mine by shuttle car power system failure modes were evaluated and reported. A "fault tree" analysis was generated for each power system and the probability of death or injury during an eight (8) hour shift and for a week of operation (105 hours) were estimated. Advantages and disadvantages were also reported for each system studied. Aerojet's assessment, based on a "forced decisions analysis," of the four shuttle car power systems as to their relative health and safety is that the exhausting or open cycle diesel is preferred, followed by the closed cycle diesel, the electric cable/motor and the rechargeable battery/motor. The assessment was based on ten parameters pertinent to the health and safety aspects of shuttle cars. Drawing layouts of two typical mine shuttle cars incorporating closed cycle diesel powerplant concepts are included in this report. The performance parameters established for the two concepts are tabulated. A study was made of the logistics involved in transporting and replenishing the consumables needed to operate a shuttle car equipped with a closed cycle diesel. The closed cycle diesel powerplant consumables -oxygen, potassium hydroxide, and diesel fuel - were analyzed as to their compatibility with materials found normally in a gassy coal mine. Spill tests simulating inadvertent leakage of liquid oxygen in a coal mine were performed. A critical evaluation of existing oxygen sensitivity data and the relationship to the use of oxygen on a shuttle car in a coal mine environment is presented.

Jan 1, 1972

By Rita D. Sylvester

"This compilation lists and summarizes publications by Bureau authors published in the regular Bureau of Mines series, in scientific, technical, or trade journals, or in other media; those available from the Bureau of Mines are indicated. Patents issued to Bureau personnel are also listed, and instructions are given on how to apply for permission to use them. One of the outstanding features of this Special Publication is an exhausive subject and author index. IntroductionThe Bureau of Mines was established in the public interest to conduct inquiries and scientific and technologic investigations concerning mining and the preparation, treatment and utilization of mineral substances; to promote health and safety in the mineral industries; to conserve material resources and present their waste; to further economic development; to increase efficiency in the mining, metallurgical, quarrying, and other mineral industries; and to inquire into the economic conditions affecting those industries, The organic act of the Bureau, as amended by Congress and approved February 25, 1913, made it the province and duty of the Bureau to ""disseminate information concerning these subjects in such manner as will best carry out the purposes of this Act.""In accordance with this directive, the Bureau reports the findings of its research and investigations in its own series of publications and also in articles that appear in scientific, technical, and trade journals; in proceedings of conventions and seminar's; in reference books; and in other non-Bureau publications. The number of these reports, the wide range of subjects they cover, and the variety of mediums in which they appear make the kind of list and index presented in this special publication both necessary and valuable. This issue describes reports and articles published during the period January 1, 1970, to December 31, 1974. It supplements the 50-year list of Bureau publications issued from July 1, 1910, to January 1, 1960;2 the 50-year list of articles by Bureau authors published outside the Bureau from July 1, 1910, to January 1, 1960;3 and the 5-year lists of Bureau publications and articles from January I, 1960, to December 31, 1964,4 and from January I, 1965, to December 31, 1969.5 It includes all the material in the four annual lists of Bureau publications and articles from January 1, 1970, to December 31, 1973."

Jan 1, 1975

By R. K. Guthrie, H. B. Hill

"INTRODUCTION The Rodessa oil field in Marion and Cass Counties, Tex., Miller County, Ark., and Caddo Parish, La., was discovered in August 1930, when a well drilled to a depth of 5,506 feet in Caddo Parish, close to the surface trace of the Rodessa fault, produced gas in commercial quantities. During the next 5 years 21 gas wells were drilled, and a large reserve of natural gas was developed in the high part of the Rodessa structure underlying the Massive Anhydrite member of the Glen Rose formation.In July 1935 oil was discovered on the flank of the structure when a well drilled to 6,048 feet also in Caddo Parish, was completed as an oil well. The discovery of oil in the Rodessa structure initiated art intensive development program of more than usual interest to geologists, engineers, and operators in the petroleum and natural-gas industry. The complex nature of the different reservoirs, the large volume of ""free"" gas under high pressures and temperatures in the upper part of the reservoirs, and the high gas:oil ratios of some of the wells during the early production of oil presented many problems the solution of which taxed the ingenuity of those responsible for the development and exploitation of the Rodessa field.Early waste of gas and resultant dissipation of reservoir energy led to State intervention, which brought about effective regulation of the quantity of gas blown to the air. Oil production from wells producing more than 2.000 cubic feet of gas per barrel of oil was restricted, and oil allowables generally were based equally upon acreage and upon reservoir pressures. Consequently, oil-production statistics of the field are closely associated with gas:oil ratios and reservoir pressures.The primary object of this report is to present factual and statistical data on well performance in the Rodessa field and to show established trends under certain operating methods and practices in high-pressure gaseous reservoirs, which have been carefully developed and outlined by modern technique and exploited under close engineering supervision.The large size of the Rodessa field, it’s fairly well defined gas caps, and a somewhat definite water table in a part- of the oil- and gas-producing region are of special interest and importance in a study of the fluid and energy relationships in reservoirs. Consequently, the results of much of the authors' work in collecting and compiling date are presented in the form of tables, maps and cross sections to show the relative thicknesses of the different oil-and gas-producing members and their permeability and porosity relationships, and, when possible, the variations in stratigraphy are exhibited in cross sections. Results of a detailed analysis of cores from a number of wells also are presented in tabular and graphic form."

Aug 1, 1943

By Michael J. Sapko, Kenneth L. Cashdollar

In This Chapter [Methane ignition as initiation source for much larger secondary coal dust explosions Rock dusting requirements to prevent coal dust explosions Dangers of hybrid mixtures of methane and coal dust] Although methane explosions are dangerous, those that involve coal dust are even more so. If exploding methane disperses and ignites the coal dust that has accumulated on the mine ribs and floor, the burning coal dust immeasurably increases the strength of the explosion. Such methane-dust explosions are prevented by inerting the coal dust in a way that prevents the exploding methane from igniting it. This chapter discusses the dust hazard and how it is pre-vented in U.S. coal mines. METHANE IGNITION AS INITIATION SOURCE FOR MUCH LARGER SECONDARY COAL DUST EXPLOSIONS The typical scenario for coal mine explosions starts with the ignition of a flammable methane-air atmosphere near the face. The turbulent winds from the primary methane explosion then disperse the coal dust. If there is insufficient rock dust (usually limestone), a secondary coal dust explosion then propagates throughout large sections of the mine. These scenarios have been studied extensively at the Bruceton Experimental Mine (BEM) and the Lake Lynn Experimental Mine (LLEM) of the NIOSH Pittsburgh Research Laboratory. The minimum quantity of methane required to initiate a coal dust explosion was studied in 1930 in the BEM [Rice et al. 1933; Nagy 1981] and then later in the LLEM, whose cross-sectional area (130 ft2) is over twice that of the BEM [Sapko et al. 1987a]. Studies conducted in the BEM closely simulated conditions that existed in operating mines in the early 20th century. The later tests in the 20-ft-wide entries of the LLEM simulated the geometries of modern mines with advanced roof support technology. The data from the BEM tests show that 13 ft3 was the minimum quantity of methane at the face that, when ignited, would disperse and ignite coal dust. In the BEM, this amount of methane was mixed with air to form a total flammable volume of about 140 ft3 of a 9% methane-in-air mix-ture. In the wider entries of the LLEM, about 37 ft3 of methane was required to disperse pure coal dust and start a self-sustained coal dust explosion. This amount of methane was mixed with

Jan 6, 2006

By C. P. Haller, S. J. Aresco

"INTRODUCTION The Bureau of Mines has been active in promoting the purchase of coal for Government use under specifications that define the requirements in terms of heat¬ing value of the coal, expressed in British thermal units, and the composition as shown by proximate analyses. To these, when required, are added the fusing temperature of ash and the free-swelling index. Under most of these specification contracts the bidders guarantee the quality of the coal, and that guaranteed by the successful bidder becomes the standard of his contract. The deliveries are sampled in accordance with instructions./ issued by the Bureau; the samples are analyzed in the Bureau's laboratory to determine whether the coal is of the quality guaranteed by the contractor and, if it is not, a price adjustment is made.Analyses of the delivered coal and ""tipple"" samples (samples collected at mine tipples as coal is being loaded into railroad cars or trucks) provide valuable data for use in evaluating future bids. Moreover, the continuous sampling of coal as delivered is a check on the practical results obtained in burning the coal.The Government purchased approximately 5-1/2 million tons of coal in fiscal year 1951. In connection with these purchases, the Bureau analyzed 10,700 samples. These are published for the use of Government officials and the public. A series of publications s been completed on coal analyses by States, or, where production is small in any State, by groups of adjacent States. In addition to analytical data showing composition and quality of coals of the States, the publications contain brief descriptions of the geological structure of the coal basins, typical mining conditions in the various districts, the principal economic data of the industry, and the relationship of mine or channel samples to delivered and tipple samples. The present publication covers many States but deals only with the analytical data of the coals.The technical papers published by States, with their numbers are: Alabama, 347; Alaska, 682; Arizona, California, Idaho, Nevada, and Oregon, 696; Arkansas, 416; Colorado, 574; Illinois, 641; Indiana, 417; Iowa, 269 and 706; Kansas, 455; Kentucky, 308 and 652; Maryland, 465; Michigan, North Dakota, South Dakota, and Texas, 700; Missouri, 366; Montana, 529; New Mexico, 569; Ohio, 344 (and Bulletin 499); Oklahoma, 411; Pennsylvania Bituminous, 590 (Supplement, 645); Pennsylvania Anthracitic, 659; Tennessee, 356 and 671; Utah, 345; Virginia, 365 and 656; Washington, 491 (Supple¬ment, 618); West Virginia, 405 and 626; and Wyoming, 484."

Feb 1, 1953

By Sumner M. Anderson

ATHOUGH scattered signs of increased activity in the mining industry accompanied an incipient improvement in the overall economic situation during the latter months of 1963, production levels for the year as a whole responded to the general 1962-63 business recession with a 4.2 percent decline. The petroleum industry, active since 1907, did not become a dynamic force in the economy until exploration and drilling contracts were concluded with private companies in 1958 and 1959; on November 15, 1963, having virtually achieved self-sufficiency in petroleum in the preceding year, those contracts were unilaterally annulled. In the field of metallics, the only approach to internal sufficiency-aside from the iron and steel industry that has depended heavily on the importing of basic raw materials-has been in mining, smelting, and refining lead and zinc although iron ore, silver, beryllium, tantalite, and tungsten have been mine production items of economic significance. Reserves of lead, zinc, silver, bismuth, vanadium, and low-grade and metallurgically difficult iron ore-as well as those of beryllium, tantalite, and tungsten which are produced primarily for export-appear to be fairly abundant. The exploration of deposits of copper, cobalt, tin, lithium, gold, manganese and uranium has been undertaken on a relatively small scale, and little precise information is available on the extent of reserves. Argentina is dependent on foreign sources for part of its required iron and steel and, except for lead and zinc, most of its nonferrous metals. Nonmetallies and solid fuels present an exceptionally varied array of production items among which cement had the second largest 1963 output in South America. Borates, bentonite, common clay, diatomite, fluorite, granite, lime, limestone, obsidian, onyx, rhodochrosite, salt, and spodumene ordinarily are produced for export either exclusively or in excess of domestic consumption requirements. Barite, cement, chalk, gypsum, marble, and mica are both exported and imported. Argentina depends entirely or partly on imports for coal and coke, certain special types of clay, mineral abrasives, graphite, magnesite, mineral fertilizers, sulfur, talc, and minerals for construction, the last group being chiefly attributable to convenience of river transport from Uruguay.

Jan 1, 1964

This NIOSH Toolbox Training program contains 52 different modules designed to stimulate safety discussions among your employees. Each module is intended to be completed in roughly 10 or 15 minutes, so a Toolbox Training session can be conducted once a week for the entire year. Each session counts toward your employee’s 8-hour annual MSHA refresher training, thus saving an entire day of work.. The modules are used to refresh employees on the procedures, policies, and practices of working safely at a specific mine site and can easily be tailored to fit conditions and situations found on your property. Every module follows the same format, so that as the weeks pass, miners will start to approach every safety situation in the same way. A typical Toolbox Training session contains five basic steps that cover all the information in the module. 1. Identification of the hazard and discussions about the types of injuries typically associated with the hazard. 2. Discussions of some of the causes of the hazard. The MSHA example included with each module should be referenced. 3. After the hazard and accident has been explained and understood, discussions about best practices and tips for avoiding a similar accident. 4. Review of similar conditions or near misses from your mine site using the leading questions on the form to start a discussion. Specific examples of equipment or structures on the site that could lead to the same accident should be discussed. The blank space on the module is used to add your own remarks and questions to supplement the given information. 5. Miners fill out the training record on the back of the form with name, initials, and date of training session. The modules are designed to be kept in a standard three-ring binder and easily photocopied. This way, the same set of modules can be used for years to come. A blank form is included that can be used to create more Toolbox Training modules. The key component of Toolbox Training is to start a discussion among employees that will allow them to see the hazards around them every day and respond to these hazards in a safe and consistent manner. To start a discussion, ask questions that need more than a one-word response. Get the miners to talk about similar situations from their own experiences and give suggestions about how to fix the hazards around them. Toolbox Training is a quick and effective way to cover specific mining hazards, but it only works if miners are involved in the discussion.

Jan 1, 2004

By Clifford H. Summers

This report describes an experimental program which we conducted on behalf of the U. S. Bureau of Mines, to select from currently available materials two novel systems of construction for air-restraining bulkheads to be used in control of radon-polluted air in underground uranium mines. The program included both laboratory and field tests, the latter conducted in a portion of an operating mine owned and operated by Union Carbide Corporation at LaSal, Utah. The test site was a worked-out, double entry stope having a Volume of about 2000 cubic meters, and was located immediately adjacent to an exhaust airway. The bulkheads were constructed of 15 cm x 1.5 cm timber uprights, 5 cm x 15 cm horizontal members supporting a close-fitted 13 mm plywood curtain-wall., which was covered with a low-permeability membrane to seal in the radon-polluted air. The two membrane materials selected for testing were chosen from a list of twenty-two potentially acceptable candidates using criteria such as strength, flammability, ease of application, adhesion, radon permeability, and elasticity. Thirteen of the candidates were tested in our laboratories for adhesion to plywood and to both wet and dry concrete (simulating wet and dry mine rock surfaces), and for their general physical properties, including radon permeability. The final selections made by the Technical Project Officer, and based on our recommendation, were a pre-formed ethylene propylene diene monomeric rubber (EPDM) membrane 1.12 mm thick, normally used as a roof sealant in industrial construction, and available from Carlisle Tire and Rubber Co., Division of Carlisle Corp., Carlisle, PA, and Aquafas 48-00, a water-based mastic used in a wide variety of commercial applications and available from H. B. Fuller Co., Foster Products Division, Spring House, PA. The EPDM membrane was applied as a single sheet, laminated (dry) between two layers of plywood, and the Aquafas was applied both by troweling and spraying onto a plywood surface. In the field test, radon concentrations were continuously measured before, after, and in the trapped air behind each bulkhead. Temperature, pressure, and velocity of the air in the adjacent airway were also measured and recorded continuously. Radon fluxes at the rock/air interface and through the bulkheads themselves were measured manually on a frequent basis. All continuously acquired raw data were recorded automatically on two USBM Data Aquisition Systems (DAS), and the raw data were converted into engineering units and recorded on magnetic disks by a specially programmed microcomputer that was interfaced with the DAS's. Relatively clean (i.e. low radon concentration) air was pumped into a point near the face of one test bulkhead from a location 170 meters away through vent bag. Several air curtains and two air-restraining bulkheads were erected in passageways adjacent to the test area in an effort to control both velocity and radon concentration of the air flowing past the bulkhead faces. An average concentration of approximately 500 pCi/liter at a flow of 550 m3/minute was attained as the background concentration against which leakage through the bulkheads was measured. The entire field test period extended from October, 1981 to March, 1982, a total of six months, and was subdivided into seven successive time periods: 1) site preparation 2) background measurements prior to bulkhead construction 3) passive restraint, with bulkheads in place and totally sealed 4) open bleeder period of natural convection ventilation in and out of the bulkheaded area through a 10 cm (diameter) opening, with connecting tubing; leading into an Exhaust airway 5) open bleeder period, having forced exhaust through a 1.0 cm opening using a servo-controlled fan 6) charcoal trapping period, using the servo-controlled fan to exhaust air through the charcoal traps, to test the effectiveness of such traps in removal of radon from the exhausted air 7) site decommissioning Each of the two systems tested, one incorporating the pro-formed EPDM-rubber membrane and the other utilizing Aquafas 48-00, was assessed for effectiveness in controlling radon-polluted air by observing the change in the radon concentration in the ventilation air flowing continuously over the surface of each bulkhead. The sensitivity of the observed measurements was limited by the volume flow-rate of the ventilation air and the radon concentration in that air. The data would indicate that the observed increments in radon concentration in the ventilation air moving across bulkheads B-l and B-2 were essentially zero within the limits of sensitivity of the measurements. Data from radon flux measurement taken both on the sealed ribs adjacent to the bulkheads and on the bulkheads themselves did not differ significantly from blank data.

Jan 1, 1982

By GEORGE S. POPE

This bulletin is a revision of Bulletin 63 and is published by the Bureau of Mines in order that purchasers of coal for Government, State, municipal, or private use may be informed regarding advances or refinements in sampling coal deliveries and preparing specifications for the purchase of coal. The purchase of coal by the Government under specifications depending on the heating value of the coal, its content of ash and of moisture, and other considerations, rather than upon the reputation or trade name of the coal, was based on the fuel investigations begun by the Technologic Branch of the United States Geological Survey in 1904. The plan was first adopted by the Treasury Department in 1906. Since then the plan, variously modified in form, but the same in principle, has been gradually adopted by other departments until, at present, of the coal used by the Government, the total value of which approximates $8,000,000 annually, more than half is purchased under specifications. The Government publications dealing with the adoption by the Government of the specification plan, the number of contracts awarded on that basis, and the quality of the coal delivered under such contracts in the several fiscal years covered by the reports are given in publioations a listed at the end of this bulletin. Under the authority of acts of Congress making appropriations for analyzing and testing fuels belonging to or for the use of the United States Government, a laboratory is maintained at the headquarters of the Bureau of Mines, Washington, D. C., where samples representing deliveries of coal purchased under specifications for Government use are analyzed and tested. At this laboratory more than 1,200 samples have been analyzed and tested in one month. With the complete equipment and the efficient force employed, the analytical work has been so perfected that the analyses and heating-value tests of the samples received are as accurate as may be reasonably expected in laboratory work of this nature. One of the serious drawbacks to the general adoption of the specification method for the purchase of coal is the difficulty of obtaining at reasonable cost samples of coal that can be considered fairly representative of the commercial product delivered in wagons, railroad cars, or ships. Therefore the method of taking and preparing samples for shipment to the laboratory has been given fully as much care as the making of the analyses and tests, and a general plan of collecting samples fairly representative of the delivered coal has been evolved through various modifications and improvements based on experience and increased knowledge of the physical and chemical characteristics of the various coals that are purchased by the Government. The method that is in general use by the Government is described in the following pages. In connection with studies of the coal deposits of the country, of the best methods of preventing waste in mining, and of increasing efficiency in the utilization of coal belonging to or for the use of the Government, geologists and engineers of the United States Geological Survey and of the Bureau of Mines have visited more than 2,500 coal mines scattered through all of the coal-producing States and Territories. From each mine 2 to 8 or more samples were taken, the number depending upon the size of the mine and its output. The analyses and descriptions of the samples collected up to the end of the fiscal year 1913 are published in Bureau of Mines bulletins 22 a and 85,b and those of samples collected in the fiscal years 1914 and 1915 are presented in a bulletin that will soon be published by the bureau. A study of the analyses of these samples and of the samples taken from cars shipped from a number of the same mines shows that the mine samples are as a rule of higher grade than the average commercial shipments, particularly with respect to containing a lower percentage of ash and having a correspondingly higher heating value. This difference is due to the fact that the Government inspector, proceeding on the basis of what can be done by a good miner, usually removes more of the partings of bone, slate, and other extraneous matter from his mine samples than the average miner does in ordinary practice, the aim of the miner being to get the maximum number of tons past the tipple inspection, as his earnings are based on his output. The average miner, therefore, does not always take as much care as he might in rejecting the impurities. The difference may be caused, too, by the presence of pieces of the mine roof and floor in commercial shipments. The Government inspector can readily

Jan 1, 1916

By Forrest T. Moyer

The overall injury experience, excluding officeworkers, at metal mines and mills during the period 1961-63, reported by operators, was 174 fatal and 11,357 nonfatal disabling work injuries during an aggregate worktime of 451 million man-hours. These injuries occurred at a frequency rate of 25.55 per million man-hours, comprising 0.39 for the fatal and 25.17 for the nonfatal injuries. The combined severity rate of the disabilities was 3,779 days lost or charged per million man- hours, consisting of 2,314 for the fatal and 1,465 for the nonfatal injuries. The average severity of all disabilities during 1961-63, was 148 days per injury. The number of days lost and/or charged, ex- cluding officeworkers, at metal mines and mills during the period 1961-63, was 1,440,000 days charged to fatal injuries and 661,290 days lost from nonfatal injuries. An average of 70,966 men were at work daily. The outstanding cause for fatal injuries was the fall of rock or ore from the back, while most of the nonfatal injuries were caused by handling materials, slips or falls of persons, machinery, falls of rock or ore from back and side, and haulage. Injury experience at primary nonferrous smelters, refineries, and reduction plants during 1961-63, excluding officeworkers, was 23 fatal and 3,214 nonfatal disabling injuries during a total work- time of 277 million man-hours. The injuries occurred at frequency rates of 0.08 per million man-hours for the fatalities and 11.61 for the nonfatal disabilities. The severity rate for all injuries of 1,085 days lost or charged per million man-hours comprised 498 for fatal and 587 for nonfatal injuries. The number of days lost, excluding officeworkers, at primary nonferrous smelters, refineries, and reduction plants during 1961-63 was 138,000 from fatal injuries and 162,385 from nonfatal injuries, An average of 34,938 men were at work daily an active plant days during the three year period. The injury experience for officeworkers at metal mines and mills during the period 1961-63 was 20 nonfatal disabling work injuries, (there were no fatal injuries reported), during an aggregate work-time of 28,694,244 man-hours. These injuries occurred at a frequency rate of 0.70 per million man-hours. There were no fatal injuries reported for officeworkers at non-ferrous smelters, refineries, and reduction plants during 1961-63, but seven nonfatal injuries were reported during an aggregate work-time of 30,814,772 man-hours. The frequency rate for these injuries was 0.23 per million man-hours for this period.

Jan 1, 1967

By A. T. Iannacchione

Synopsis Underground stone mining Is an emerging sector of the U.S. mining industry. As this expansion takes mines under deeper cover, and as more efficient mining methods are utilized, effective stone pillar design methods will become even more Important. Current design practices are examined and a discussion of safe mine layouts is presented as a first approach towards weighing the demands for increased production against increased risk. Risks to underground stone-mine workers include rib Instabilities, pillar failures and roof falls. Seventy-two stone-mine pillar designs were examined. Pillars with width to height ratios of less than 1.5 appear more likely to fail when subjected to excessive stress levels. When width to height ratios fall below 1.0 defects in the pillars, such as through-running discon¬tinuities, can have a significant influence on stability. Discontinuity persistence, dip, material properties and orientation are also important determining factors in pillar strength. During the past three years the number of active underground stone mines in the U.S.A. has ranged between 90 and 100. This number is expected to increase as the crushed stone industry responds to growing demands for its products.1 As more of the industry moves to benefit from the advantages of underground mining production from underground stone mines is expected to increase above its current level of approximately 66 000 000 r/year. Parker2 identified four advantages of underground mining operations: (1) surface developments, zoning laws and environmental concerns are often less of an issue; (2) stripping and restoration requirements are eliminated; (3) additional reserves are often available beneath the quarry floor, under pit slopes or under an adjoining property; and (4) space is created for secondary use. Underground mines enjoy the added benefits of work in a constant underground climate, rather than in the variable surface climate, minimization of community concerns by placement of the crushing, sizing and stowing operations underground and reduction of surface vibration concerns through smaller blasts. The drawbacks of underground mining include added health and safety hazards for the stone miners, whose health can suffer from increased exposure to falls of ground, airborne contaminants and fog in large underground openings. Injuries from falls of ground in stone mines have occasionally exceeded the incidence rates for other mineral resources mined underground. 3 Existing underground stone operations, by comparison with stone operations a decade ago, mine more stone at a faster rate and with larger equipment. Because of high demand mines face increased pressure to yield more stone per production blast. A majority of the miners use the V-cut blasting pattern, which limits the depth or pull for each shot to about 4 m. Therefore, to produce more stone, either the mines must work more faces or existing faces must be enlarged. The enlargement of underground openings poses problems in the maintenance o€ strata stability, First, the widths of the mine entries partially control the amount of sag or deflection that any given roof beam can withstand before failure. This deflection can take pi ace quickly after an opening is excavated or much later as weathering processes form additional, thinner beams. Because increased deflection increases the potential for roof beam failure, there is a limit to how wide a room can safely be made. These limits depend on local geological and stress conditions and, to some extent, the cost effectiveness of roof-bolting. Given these room width limitations, underground stone mines have placed more attention on expansion of production through benching where mining thicknesses permit. The heights of rooms partially control the strength characteristics of adjacent pillars: as the room height increases for a given pillar width the pillar width-to-height ratio (w/h) decreases. In general, stone pillars are very strong, but when the width-to height ratio diminishes the potential for pillar rib instabilities and pillar failures increases (see Fig. 1). [ ] In the underground environment the desire to utilize wider rooms, higher benches and multi-level mining results in the potential for unstable ground conditions. Although the number of failed pillars is currently very low, the potential for additional failures will grow if more slender pillars are developed, especially under deeper cover, with wider mining sections and in multi-level operations. It should also be noted that the passage of time acts to decrease the strength of many existing pillars. An examination of current design practices

By William W. Adams

The quarrying and related industries of the United States improved their safety record in 1938 by reducing their accident-frequency rate to a lower acid more favorable position than it has occupied at any time except 1935 since comparable accident statistics became available in 1916. The rates for 1935 and 1938 were identical. The number of men employed in 1938, although smaller than in 1936 and 1937, was larger than in any other year since 1930. Volume of employment, as measured by the number of man-hours worked in the industry as a; whole, likewise was larger than in any year since 1930 except 1936 and 1937. Reports from operating companies to the Bureau of Mines, United States Department of the Interior, showed that 77,497 men were employed in 1938, a decrease of 6,597 compared with 1937. This figure represents the average number of employees for the period during which the quarries and plants were active during the year, which in 1938 averaged 223 workdays per employee compared with 241 in 1937. A total of 17.3 million man-days or 133.8 million man-hours of work was shown for the industry as a whole, which indicated a weighted average of 7.75 hours per man-shift. Reports for 1937 showed a total of 158.3 million man-hours for the entire industry and an average workday of 7.81 hours. Accidents resulted in the death of 82 employees, 5 more than in 1937, and in 5,027 nonfatal injuries involving disability for more than the remainder of the day on which the injuries were received. The number of nonfatal injuries was 1,321 less than in 1937. As the number of deaths and injuries was reduced more, proportionately, than man-hours of employment, the accident-frequency rate per million man-hours of work declined, the rate for 1938 being 38.19 compared with 40.59 for 1937. In other words, accident occurrence to the average employee was 6 percent more favorable in 1938 than in 1937 on the basis of actual number of accidents and actual number of man-hours worked during the 2 years.

Jan 1, 1940

By M. D. Heresy, H. F. Coward

"Whenever a manometer has been used for recording explosion pressures it has been calibrated under a series of static pressures. Whether the readings taken during the rapidly changing pressures of an explosion (dynamic pressures) truly represent the actual pressures therefore is an open question. The authors had hoped to solve this problem by conducting the calibration with the aid of known pressures, applied at about the ammo rate as in an explosion, for example, by means of a cylinder and piston, the piston being driven in by an eccentric cam at a rate which would be determined accurately by the shape and rate of revolution of the cam. If the rate of compression were great enough, then the pressure in the cylinder (to which the manometer is attached) could be calculated from the laws of adiabatic changes of pressure and volume. It seems highly probably, however, from H. T. Tizard's observations on pressure changes in a cylinder during rapid compression (55) 4/ that there is much heat loss, therefore pressure loss, in the very small interval of time during compression; although Tizard was able to estimate the amount thereof, he did so only by relying on his nanometer, which is exactly the procedure that the authors challenge on the grounds that the manometer was not standardized under the conditions of its use. Hence the authors were driven from this scheme of calibration to another, namely, comparing records of ono and the same explosion taken simultaneously on several instruments of as varied a type as possible. If 2 of a number of instruments agree, perhaps both are accurate; if 3 agree, probably each of the 3 is accurate; if 4 agree, it is almost certain each of the 4 is accurate.There was a second objective in the present work. If the Bureau of Mines and the B.C.D. (British Coal Dust) manometers differ in their records, then from a sufficiently comprehensive series of experiments it may be possible to coordinate the extensive results already obtained at the American and British stations.Thirdly, it was hoped to learn which is the best available instrument for observing the pressures of coal-dust and as explosions and, if possible, to indicate means for improvement in manometry."

May 1, 1935

By David C. Byrne

PURPOSE: Evaluate practical technological advances in level-dependent hearing protectors and inexpensive personal acoustic monitors. RESEARCH SUMMARY: The results of previous NIOSH research show that miners develop hearing loss much more quickly than nonoccupationally noise-exposed individuals and that miners experience a greater severity of hearing loss than would be expected for nonoccupational noise-exposed persons of the same age and gender. Using hearing thresholds at 4,000 Hz as an indicator, coal miners experience hearing loss 2.5 to 3 times greater than would be expected for persons not exposed to occupational noise. At age 50, 90% of coal miners and 49% of metal/nonmetal miners were found to have a hearing impairment. In comparison, only 9% of a nonoccupationally exposed control group had a hearing impairment at age 50. There are numerous barriers to the prevention of hearing loss among miners. In an effort to overcome some of these obstacles, this work was initiated to evaluate practical technological advances in level-dependent hearing protectors and inexpensive personal acoustic monitors. Experiments will involve analysis of this new technology under both laboratory and field conditions. Laboratory components will consist of controlled studies using an anthropometrically correct test fixture as well as human subjects to evaluate the adequacy/effectiveness of various hearing protective devices. Electroacoustical performance testing of new noise monitors will also be conducted in a laboratory setting. Subsequent field studies will determine the actual usefulness and practicality of the personal protective/monitoring equipment as used in different mining situations. It is anticipated that increased on-the-spot awareness of hazardous noise levels and the availability of better-suited hearing protectors will provide a degree of protection beyond what is currently attained in different mining environments. Therefore, the intent of this project is to identify and adapt new acoustical technologies and products that have a direct application to individual mine workers. With increased knowledge of these technological advances, a greater demand for similar devices will likely occur, which, in turn, will lead to a better understanding and awareness of hearing loss prevention methods throughout the mining community. The ultimate goal is for widespread integration of useful hearing loss prevention products and devices into the daily routines of miners

Jan 2, 2000