Search Documents

By Arcy, D&apos Banister

The productive potential of the Nation's principal active and inactive silver mining districts has been evaluated, using 1964 economic conditions and assumed silver prices ranging up to $3 per ounce. On this basis, the United States has an estimated 5 billion ounces of unmined silver, excluding minor byproduct silver in gold deposits. Approximately 3.5 billion ounces are contained in deposits requiring higher silver prices to be economically minable, and about 1.3 billion ounces are in currently operating mines. An analysis of silver consumption trends since 1954 indicates that the free world industrial market will climb to nearly 800 million ounces annually by 1985. U.S. industrial silver consumption has been projected to increase at an annual rate of 4 percent from 1965 to 1985. Average annual increase for the remainder of the free world has been 8.1 percent since 1953; this is expected to level off with an estimated average increase of 5 percent to 1975 and 4 percent from 1976 to 1985. Foreign silver production has, in the past, exceeded consumption, but in 1960 the supply deficit was over 100 million ounces. U.S. production, also traditionally exceeding demand, was suddenly overtaken at the start of World War II. By 1964 the annual deficit, excluding coinage, was over 84 million ounces. World coinage requirements of silver are fast declining and expected to total only 25 million ounces annually by 1985. In fact, coins are already becoming major short-term source of silver to fill the large deficit between industrial demand and current production. The larger, long-term deficit (to 1985, the limit of this study) must be balanced by silver from the hoards and new production from silver deposits. Eventually all of the deficit must be met by new production or silver prices will rise until consumption is greatly curtailed. To balance free world supply and demand, it is estimated that the free world must produce an additional 85 million ounces by 1975 and 375 million ounces by 1985. The estimated proportion expected from U.S. deposits is 17 million ounces by 1975 and 74 million ounces by 1985.

Jan 1, 1969

By R. H. Oitto

To determine whether reflectors could substitute for the trip lights referred to in the Federal Mine Safety Code, the Bureau of Mines studied trip safety devices under simulated haulageway conditions. The noticeability of reflectors compared with trip lights and fluorescent panels was evaluated from distances of up to 850 feet with light source intensities of up to 1-6 foot-candles (ft c). Reflectors tested consisted of a reflective lens in a holder secured to a frame that can be attached to a trip. Reflectors were found to be suitable for use as a trip safety device. In fact, for the conditions in many underground haulageways some reflectors would be more effective than a permissible trip light.

Jan 1, 1968

By George S. Rice

"""To prevent the propagation of mine explosions, the Bureau of Mines, Department of Commerce, recommends rock-dusting all coal mines, except anthracite mines, in every part, whether in damp or dry condition. It also recom¬mends that rock-dust barriers be used to sectionalize the mine as additional defense; but these should not be regarded as a substitute for generalized rock-dusting."" (Mine safety decision, No. 5, April 7, 1927.)The purposes of this circular are to present concisely the requirements for effective rock-dusting of coal mines in order to prevent propagation of mine explosions, and to point out the principal related facts on explosibility of coal dusts and mixed coal-mine dusts, as determined by extensive testing in the Experimental Mine of the Bureau of Mines. study of these principal facts on coal-dust explosibility in conjunction with the recommendations for rock-dusting, should help to make clear to mine operators and others concerned the necessity for systematic rock-dusting in all coal mines except anthracite mines, and the futility of inadequate or sporadic rock dusting.The descriptions of nearly 1000 explosion tests in the Experimental Mine are available in a forth-coming bulletin2 and in earlier publications.3For detailed specifications as to the kind of rock-dust, amount to use, where to be applied, and method of sampling and testing, the reader is referred to Bureau of Mines Serial 26064, and American Engineering Standard Committee ""Recommended American Practice for Rock-Dusting Coal Mines"" quoted in Bureau of Mines Information Circular No. 60305.Summary of Specifications for Rock-Dusting1. Mines to be rock-dustedAll coal mines, except anthracite mines, should be rock-dusted.2. Extent of rock-dustingAll open accessible parts of a coal mine should be rock-dusted, including slopes, entries, cross,-cuts, and rooms, headings and pillar workings to within at least 50 feet of the face.3 Amount of rock-dustingAfter cleaning up the coal dust as thoroughly as practicable, sufficient rock-dust should be used to have the mixture of remaining coal-dust plus the rock-dust in any zone contain 65 per cent or more of inert or noncombustible matter. First dusting may require 3 pounds to 5 pounds per linear foot of passageway.4. When to rock-dust In any zone or section (made by dividing the mine for rock-dusting purposes) when the incombustible content falls to 55 per cent, that section should be re-rock-dusted so as to bring the incombustible content well above 65 per cent."

Jun 1, 1927

By Z. J. Fink

Ultraviolet-visible and infrared emission spectra from propagating methane-coal-dust-inhibitor mixtures have been obtained in a 0.2-by 0.2¬by 5.3-m vertical duct over a range of flame speeds from 2 to 100 m/sec. Results serve as a basis for the design and evaluation of thermal radiation detectors for triggered barriers that suppress ongoing mine explosions. Of particular significance is the large continua superimposed on the methane-air spectra when small amounts of coal dust are present and the variation of emissivity with infrared (IR) wavelength in dust flames. In conjunction with the spectral measurements, studies of flame speed, particle size, gas sampling plus thermodynamics, and dust sampling are presented and discussed in order to gain insight into the more fundamental aspects of propagating explosions. Sampling results have shown that only a near-stoichiometric amount of dust is consumed in the flames studied even though initial concentrations are five times stoichiometric. Difficulty in propagating coal-dust inhibitor flames of realistic mine characteristics in a cold-walled laboratory duct has been encountered. A simplified analysis of the flame zone utilizing spectral and sampling results indicates that the stability of slowly propagating flames is more dependent upon thermal radiation than upon convection losses to the walls. It is predicted that thermal radiation losses will make any measurements of coal dust burning velocities, "quenching diameters," or flammability limits more apparatus dependent than those corresponding to gaseous fuels.

Jan 1, 1974

By Rimas Pakalnis, Andrea M. Ouchi, Thomas M. Brady, Mary M. MacLaughlin, Cristian Caceres, Paul Hughes

A major focus of ground control research presently being conducted by the Geomechanics Group at the University of British Columbia, Canada, in conjunction with the National Institute for Occupational Safety and Health’s (NIOSH) Spokane Research Laboratory, is the development of design guidelines for underground mining within weak rock masses. The study expands upon the span design curve for man-entry operations and the stability graph for nonentry operations developed at UBC by extending the application to weak rock masses. The original database has been augmented by weak rock mass information from mines throughout the United States, Canada, Australia, Indonesia, and Europe. The common factor in all of these mines is the presence of a weak back and/or walls. This paper expands on the North American database and how the design curves have been employed at mining operations throughout the world. The definition of a weak rock mass for this study has been defined as having an RMR76 under 45% and/or a Q-value under 1.0.

Jan 5, 2007

By Annie G. Smelley

The Bureau of Mines is testing a dewatering technique for phosphatic clay waste that will recover a portion of the water lost using conventional waste disposal methods and produce solids suitable for land reclamation. The technique utilizes a flocculant, polyethylene oxide (PEO), that forms strong, stable flocs that can be dewatered on a static screen followed by further dewatering on a rotary screen. In field tests conducted in northern Florida, clay wastes containing a nominal 2.5 pct solids have been consolidated to greater than 20 pct solids. The rate at which PEO-treated material continues to dewater in a mine cut also was monitored. Preliminary results indicate that PEO--treated material will dewater to 30 pct solids in 140 days. Pretreatment of phosphatic clay waste with lime and H202 was shown to improve the dewatering technique for problem clay wastes.

Jan 1, 1985

By D. A. Elkins

This report is the second of a series of three that will evaluate, in order, the metallothermic, carbothermic, and electrolytic methods of magnesium production. The historical development of the carbothermic process is outlined. Production costs varying from 38.9 to 21.8 cents per pound of magnesium are estimated for different versions of the process. Additional suggested process modifications are listed. The technical and economic potential of the different versions are compared. Areas deemed most promising for future research on the process are delineated.

Jan 1, 1967

By Louis Schaffer

A literature review concerning dredging hazards and dredge safety practices and an examination of MSHA accident and injury statistics on dredge mining in the United States provided the background for an analysis which was completed by visiting 31 dredge operations, dredge manufacturers, and dredging experts. The principal hazards are discussed and countermeasures are suggested in several areas. Drowning is the single largest fatal injury category. Special emphasis is given to the design, care and use requirements of personal flotation devices.

Jan 1, 1984

By P. J. Remington

Field measurements utilizing a seat pad accelerometer were made on 25 underground coal mining machines in order to quantify the vibration levels experienced by the operators. Those data were combined with information on the statistics of the operating times of those machines in order to compare the operator's vibration exposure to the criteria in ISO 2631, "A Guide for the Evaluation of Human Exposure to Whole Body vibration." The comparison showed that about 30% to 40" of all underground coal mining machine operators can be expected to be exposed to vibration exceeding the fatigue decreased proficiency criterion, whereas only 10 to 15% would be expected to experience vibration exceeding the exposure limit. The primary sources of the exposure are shuttle cars and possibly scoop trams. Detailed exposure statistics are presented and the implications of the results are discussed.

Jan 1, 1983

By S. H. Lorain

Mechanized dirt-moving equipment has greatly increased the scope of prospect trenching by lowering the costs and increasing the speed of such work. Where the soil covering was more than 3 or 4 feet deep, operators were quickly discouraged and abandoned surface exploration in favor of shaft sinking, drifting, or diamond drilling; wildcat trenching was seldom undertaken, even where the soil covering was thin. Although power shovels, trench diggers, and draglines are useful under special conditions, caterpillar-mounted angledozers will usually do the same type of work and have a far wider range of adaptability. They are therefore incomparably the most generally useful equipment for the purpose. Since 199 the Bureau of Mines has been engaged in a widespread search for minerals useful to the war program; this work required a great deal of pioneer exploration in areas where known outcrops were marginal or subcommercial by pre-war standards. Engineers in charge of field operations were quick to perceive the applicability of mechanized trenching on a fairly large scale, particularly in the Northwestern States, where completely exposed vein-outcrops are rare. Probably no other organization to date has done a comparable amount of angledozer prospect trenching under such a wide range of conditions. The following comments and accompanying cost tables are based on work performed chiefly in Idaho, under a variety of physical and climatic conditions. The most important applications of the angledozer for such work may be described under three heads:

Jan 1, 1944

By Sean Gallagher, Thomas G. Bobick, William S. Marras

Twelve healthy, experienced underground cool miners performed lifting capacity tests in stooped and kneeling postures using a modified psychophysical procedure. Subjects adjusted weight in a lifting box to the maximum they could handle without undue fatigue in an asymmetric lifting task. Lifting periods were 20 min in duration and the frequency was 10 lifts/min. Tests were performed under a 48-in. roof that restricted the subject&apos;s posture. Psychophysical. physiological, and biomechanical dependent measures included the maximum acceptable weight of lift (MAWL), heart rate (HR), rate of oxygen consumption (VO2), ventilation volume rate (VF), respiratory exchange ratio (R), and integrated electromyography (EMG) of eight trunk muscles. Results indicated that the MAWL was significantly lower when kneeling than when stooped (p < 0.05), Furthermore, metabolic demands were greater in terms of HR (p < 0.005) mid VO2 (p < 0.05) when kneeling. Left and tight erectores spinae muscles exhibited increased EMG activity- in the kneeling posture (p < 0.001). It was concluded that psychophysical lifting capacity is decreased, while the metabolic stress and internal load on the spine are increased, in the kneeling posture. Results of this Bureau of Mines study indicate that it may be advisable to reduce the weight of materials that are handled repetitively in the kneeling posture.

By Michael G. Zabetakis

Vaporization of a solvent into an airstream at temperatures above the flash point of the solvent is hazardous if vapor is liberated at a rate that is sufficient to produce a flammable mixture. In may operations, the formation of such a mixture is prevented by rapid diluton of the vapors with air. The minimum air requirements for the dilution process are determined by the composition of the lower limit mixture at the temperature and pressure of interest. Because limits of flammability cannot always be calculated with sufficient accuracy for a proposed operation, they .rust often be determined experimentally. The present Bureau of Mines investigation was prompted by an application in which the lower limits of flammability of a number of solvent mixtures were required over a range of temperatures at atmospheric pressure. These limits were to be used along with available solvent-vaporization-rate data to establish minimum airflow rates needed to prevent the formation of flammable mixtures in an industrial process. Lower limits of flammability were determined for a number of mixtures containing methyl ethyl ketone, tetrahydrofuran, toluene, ethyl alcohol, and ethyl acetate in dry air-at atmospheric pressure and temperatures between 16° and 200° C. The experimental values were found to be in good agreement at with those predicted by Le Chatelier&apos;s law; variations with temperature were in fair agreement with those predicted by a modified Burgess-Wheeler law.

Jan 1, 1962

By Larry Joe Linley

The Department of the Interior, Bureau of Mines is reviewing explosion protection methods for use in gassy coal mines. This performance criteria guideline is an evaluation of three explosion protection methods of machines electrically powered with voltages up to 15 000 volts AC. A sufficient amount of basic research has been accomplished to verify that the explosion-proof and pressurized enclosure methods can provide adequate explosion protection with the present state of the art up to 15 000 volts AC. The routine application of the potted enclosure as a stand alone protection method requires further investigation or development in order to clarify performance criteria and verification-certification requirements. An extensive literature search, a series of high voltage tests, and a design evaluation of the three explosion protection methods performed at White Sands Test Facility indicates that the explosion-proof, pressurized, and potted enclosures can all be used to enclose up to 15 000 volts AC, if the performance criteria and verification-certification requirements recommended in this guideline are implemented.

Jan 1, 1982

By S. D. Hill

To help insure the continued viability of the minerals and materials sector of the U.S. economy, the Bureau of Mines is conducting research and disseminating technology that will result in more efficient minerals extraction. As part of this effort, a process is being investigated that involves an aqueous chlorine-oxygen leaching procedure to treat complex sulfide concentrates to produce ZnCl2. Subsequent fused-salt electrolysis of the ZnCI2 produces high-purity zinc metal. Previous reports described preliminary, small-scale research on the chlorine-oxygen leaching system and fused-salt electrolysis of ZnCl2. This report discusses the details of design, construction, and longterm, continuous operation of both monopolar and bipolar fused-salt cells. The effects of temperature, electrode gap, current density, electrolyte composition, and electrode configuration on current efficiency and energy consumption were determined for a ZnCl2-KC1-NaCl electrolyte system. Zinc metal was produced at the rate of about 50 lb/day in a monopolar cell which was operated continuously for 60 days with current efficiencies from 85.8 to 98.6 pct and an average power consumption of 2.1 kwhr/lb. Bipolar cells with several different electrode configurations were operated for 90 days. A current efficiency of about 70 pct and power consumption of about 1.7 kwhr/lb of zinc were achieved using a three-electrode-pair, bipolar assembly with a chlorine gas lift pump to circulate electrolyte.

Jan 1, 1981

By Thomas M. Barczak

State-of-the-art longwall roof supports provide effective strata control, but failures of these support systems still occur. To identify failure mechanisms and the impact these failures have on the safety of the support system, the U.S. Bureau of Mines has been conducting research on shield mechanics in the Bureau&apos;s unique mine roof simulator, as well as field studies of in situ support loading and investigations of longwall failures, Three types of failures are discussed: (1) inadequate support capacity, (2) structural failure, and (3) instability. Hypothetical situations with proposed courses of action and case studies of actual longwall failures are described. This information is intended to assist the Mine Safety and Health Administration (MSEIA) and industry mining personnel in safety evaluations of longwall roof support systems.

Jan 1, 1989

By B. J. Barfield

Standard erosion plots (4.6 m by 22.1 m on 9 percent slope) were constructed on the Alston Surface Mine of Peabody Coal Company in Western Kentucky to evaluate the effects of placement technique and density on the erodibility and eroded size distribution from mine spoil and reconstructed topsoil. PLots were constructed with a dozer and scraper pan from shale spoil material, topsoil (A horizon) and subsoil (B and C horizon mixture) using soils stockpiled at the Alston Surface Mine. Plots were evaluated under fallow conditions with a tilled plot surface and two compacted conditions. One compacted condition was generated by tracking the surface with a dozer and rolling with a loaded scraper pan. The second was generated by rolling the surface with a solid wheel compactor. Rainfall was supplied with the Kentucky Rainfall Simulator at a rate of 63.5 mm h-l with the plots in three different soil moisture conditions; dry, field capacity, and very wet. Runoff was evaluated for flow rates, sediment concentrations, and aggregate and primary particle size distributions. Sediment yields and the Universal Soil Loss Equation E130 indexes were calculated from the data for each plot. The erodibility was calculated from the tilled plot data using a weighted soil loss and E130 value from the dry, field capacity and wet runs, and a cover management factor calculated for every other condition. For the topsoil and subsoil plots, sediment yield was greatest for the dozer scraper pan plots and lowest for the tilled plots. For the shale plots, erosion rates from the wheeled compacter plots were the highest and tilled plots the lowest. Average erodibilities were 0.14, 0.38, and 0.41 tons ac h ft-1 tonf-1 in-1 hundred ac-1 for the shale, subsoil, and top-soil plots. Cover management factors for each plot were 1.29, 1.42, and 1.17 for shale, topsoil and subsoil plots compacted with the solid wheeled compacter and 1.29, 3.79, and 2.51 for shale, subsoil, and topsoil plots compacted with the dozer scraper pan conbination. The accuracy of the Wischmeier erodibility nomograph was evaluated on the data collected. Erodibilities were accurately predicted for the topsoil plot and poorly predicted for the sub-soil plot. The accuracy of the CREAMS procedures for predicting the composition (size distribution) of eroded particles was also evaluated. The composition was well predicted for the topsoil plots, moderately well predicted for the shale plots and poorly predicted for the subsoil plots.

Jan 1, 1983

By Zsuzsanna Zborovszky

[Certain chapters of the tentative document entitled, "Test Requirements for instruments & Apparatus to be Considered for MESA Intrinsically Safe Certification" are evaluated and supporting data are presented on certain key issues with the purpose of helping the regulatory agency in improving the above document. The subjects discussed in detail are: hot wire ignition with W and BeCu wires; the most sensitive methane-air :and propane air ratio; the ignition limits for inductive-resistive circuits in 8.3°; methane-air mixtures in the 24-OV, 100 mH-2 [ ]H range, the resistive ignition limit as a special case when L ? 0; the capacitive-resistive ignition limit representing a four ,dimensional problem is analyzed and discussed in general and the variation of the ignition limit using Vo and/or r2 as a parameter are determined experimentally. The ignition limit is also determined for the [r1 ,~ r2] = 0 case ("capacitor alone") and for the most dangerous capacitive resistive circuit combination. The ignition limits are below the limits proposed in the document. A survey of the different factors influencing the ignition is made as to their suitability to provide a constant safety factor. It is shown that a constant safety factor cannot be achieved. The best method for applying a safety factor appears to be the increase of initial pressure or use of other gas mixture. The revision of the safety factor concept and ignition limit curves is recommended.]

Jan 1, 1980

By J. L. Holman

The Bureau of Mines, at its Rolla Metallurgy Research Center, has devised methods for separating and reusing the major portion of waste plastics discarded in urban and industrial refuse. Waste plastic concentrates from a local collection drive, recycling center collections, urban refuse pilot plants, and the secondary metals industry were separated into three major thermoplastic families--polyolefins, styrenes, and vinyls--based on differences in density using a Bureau-designed sink-float-elutriation hydraulic separator. Other separation methods and systems such as air classification, jigging, liquid media, screening, and electrostatics were studied. Vinyl plastics contain up to 50 percent HC1. It was found that over 90 percent of the HC1 in vinyls can be recovered as usable hydrochloric acid by pyrolysis at 350° C. The remaining pyrolysis products which had heats of combustion in excess of 12,000 Btu/lb, could become valuable as a future source of energy. A characterization study conducted on eight chopped wire insulation wastes from secondary copper and aluminum processors showed that most wastes contained 2 to 7 percent residual metal and large percentages of vinyl. The residual metal content of some of these wastes was lowered to 0.1 percent and the vinyl portion segregated using various separation techniques. The research also suggests ways in which reclaimed thermoplastics may be fabricated into useful products.

Jan 1, 1974

By F. A. Penning

An experimental and analytical investigation was conducted on a recoverable pit bottom buffer. Work was conducted at the Colorado School of Mines using the U.S. Bureau of Mine-s free fall test tower. The four-sided butter, one of several designs by Cable Belt Ltd., was constructed of reinforced rubber slabs. The kinetic energy of falling weights, 1670 to 6000 lb, was absorbed by the flexure of rubber slabs loaded along their edge. A series of constant energy, 3000 ft-lb and maximum design velocity, 600 fpm tests were conducted. Static testing of the buffer established the load-deformation behavior. A large hysteresis loop plotted above and below the data published by the manufacturer. Creep was evident during; the loading Lend unloading cycle. Electronic instrumentation measured deceleration using your accelerometers; overall and intermediate movement was found from four displacement transducers; bending strains were recorded at four locations; and base reactions were measured by using four load cells. Data were recorded on tape and played back on a recording oscillograph. High speed movies were taken of all the tests. Symmetry of response and rebounds, when they occurred, were noted. From the displacement-time and load-time data, maximum decelerations were calculated by four methods and different assumptions. Good agreement was found from these approaches. Impact loads ranged from a low of 0.7 to a high of 6.1 gs. A dynamic computer model was written for the response of the buffer. Displacement-time, and acceleration-time plots show close correspondence with experimental data for the ten experiments. Data from accelerometers, displacement transducers, strain gages, and load cells are included. The computer program listing, and displacement-time acceleration-time plots are also presented.

Jan 1, 1980

Significantly increase the amount of fresh air that sweeps the face in front of a continuous mining ma-chine, in order to prevent methane gas buildup. Approach Water sprays are mounted along the sides of a continuous miner and behind the cutter head to help push ventilating air along the required path, and prevent its short circuiting. How It Works A continuous miner is fitted with two sets of water sprays, one of which is used at a time. The set of sprays shown in the diagram, helps move the ventilating air forward along the left side of the miner and across the cutter drum from left to right. That air is then drawn into the exhaust air passageway located along

Jan 1, 1978