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By J. P. Capp

The Bureau of Mines conducted field-scale experiments in underground gasification of coal with oxygen-enriched air. A path about 140 feet long was linked at the rate of LO feet per day through a coalbed hydraulically fractured with a mixture of starch gel and sand. During the best gasification period, which extended for 2-1/2 days, 785,000 std. c.f. per day of 124-B.t.u. gas was produced. In other periods the heating value at times was almost 300 B.t.u. per std, c.f,, but sustained production of such gas was not achieved. During gasification, a total of 50 tons of coal was gasified and 114 tons of coal was carbonized. Approximately 12 million B, t .u. per foot of nominal path was recovered at the surface.

Jan 1, 1962

By Richard E. Thill

The Bureau of Mines' Oil Shale Data Bank (OSDB) was created in 1976 to serve as a centralized source for the storage, retrieval, and dissemination of technical literature, patents, and information on research projects in oil shale. Thousands of bibliographic references, patents, and research projects relating to the mining of oil shale have been identified, abstracted, and indexed according to selected keyword topics. The resulting data bank listings are contained in three sections, each with a separate keyword breakdown for use in rapidly identifying references pertaining to a specific subject. All references cited in the OSDB were published prior to an established cut-off date of May 1, 1981, when a redirection of Bureau of Mines programs resulted in the loss of project funds and the subsequent phasing-out of data bank operations. ACKNOWLEDGMENTS The oil shale data bank listings are the result of a team effort of the project staff. The contributions of Jimaleen Jansma for keypunching, information acquisition, and data entry, and John Brandenburg and Sandra rake for keywording and information acquisition are gratefully acknowledged. Appreciation is expressed to the staff members of the Smithsonian Science Information Exchange, Washington D.C. for their cooperation in supplying data on recent research projects in oil shale for inclusion in the OSDR listings. Appreciation is extended also to the staff of the Gallegos Research Group for assisting with computer services to obtain the final printouts and transfer of files onto magnetic tape.

Jan 1, 1982

By Travis Q. Lyday

During 1983, Ecuador's gross domestic product (GDP) decreased by a reported 3.3% in real terms, the second consecutive decline in the economy after more than a decade of growth generated mainly by petroleum export revenues. The slowdown was due, in part, to the effects of the world recession, lower foreign demand for oil, depressed world oil prices, and reduced demand and prices for Ecuador's traditional agricultural exports. Inflation accelerated to an estimated 55% annual rate. The sucre was devalued in March, for the second time in less than a 12-month period, by 27% after remaining unchanged during the previous decade. In addition, continuous adjustments were made every weekday, effectively devaluing the sucre another 30% by yearend. Ecuador's production of crude petroleum increased almost 12% in 1983 compared with that of 1982. Average daily production was 237,100 barrels per day (bbl/d), up from 211,300 bbl/d in 1982.

Jan 1, 1985

By L. A. Morley

This report summarizes the work performed on Bureau of Mines Contract J0199106, "Portable Trailing Cables, Splices, and Couplers - Design and Installation Considerations," from May 1, 1979 until December 31, 1981. Splice installations and considerations for the possible use of conductive-rubber shielding in low-voltage trailing cables as well as coupler installations in high-voltage cables were the principal tasks. Flammability measurements on cable components and general testing facilities for mine electrical equipment were secondary tasks. Theoretical and empirical data indicate that conductive rubber with high resistivity values is an unsatisfactory shield design for current-sensing fault detection. Elimination of partial discharge in coupler installations requires meticulous care.

Jan 1, 1982

By Kenneth Eilerts

[Efficient and econonic production of ?etroleum requiros knowledge of the propertios of rescrvoir fluids a:$. their characterintics of flow throayh porous forclntions. The cllanga in propertie s and flow characteri s- tics of fluids when their corosition is altercd cs a rnsult of reservoir- pressure decliae is of 'mecial i~ortenct? ir, planning wall-spaAng pro.yama and production schedules. Rct~ntion of the ~eatent poscible quantity of dissolved gac in the rescrvoir liquid gnneml'y is recognized cs ess~ntisl to mxim ultimzte recoverj of oil fro3 the underground reecrvoir. Ilo:rever, pro6uction mothodn tending to mintain gis in solution in the oil and re- sulting in the optimm ocolo~ic recgvery of oil not only vary from field to fiold but also my have to be changed from time to time in a given field as it is being depleted of its recoveraklo oil. Const vestly there is need tc investigztc tY.e propertiec of resa:-voir f kids corresponding to successivc stages of reservoir depleJ&ion with the view of obtnining infomation oacential to efficient prcriuctian practices.]

Jan 1, 1939

By P. M. Daling

This report, sponsored by the U.S. Bureau of mines, presents the results of an examination of available safety assessment methods and their applicability to the mining industry. Standard safety analysis procedures were evaluated for relevance to the mining industry and mire operators' needs. The major safety concerns and needs of mine operators were identified through telephone contacts and a visit to a mine site. Safety assessment methods recommended for use by the mining Industry were also investigated for potential adaptability to a user-interactive computer system. Several methods were determined to be Immediately transferrable to the mining industry; these techniques are described in a user's handbook produced as part of this project. The handbook gives detailed instructions on appropriate application and utilization of these methods. Other techniques will require further development before they can be effectively transferred to the mining Industry. These methods, their advantages, disadvantages and resource requirements are described in the final project report. Recommended system safety analysis techniques facilitate an evaluation of total mine system safety and a determination of safety-related interactions among systems within a mine. The application of these methods to the mining industry could result In reductions in both the frequency and the consequences of mine accidents.

Jan 1, 1983

By M. P. Hardy

Copper-Nickel ores in northeastern Minnesota could establish a new industry in a region where underground mining has not been active. This study reviews geotechnical data relevant to the selection and design of underground mines in this region and by numerical structural analysis evaluates the importance of some geotechnical parameters to mine method selection. Because of the rapid development of computer modeling many new techniques have become available for modeling rock mass behavior in recent years. An overview of these methods is presented and their applicability to mine design reviewed. Methods included in this summary were the differential continuum models, such as finite Pint and finite difference, integral continuum models, such as boundary element, displacement discontinuity and boundary integral equation methods, and discontinuum methods such as the distinct block model. The displacement discontinuity method was used to evaluate the influence of geotechnical parameters on some aspects of mine design. Three problems; room stability, pillar stability on a weak inclined contact and sill pillar thickness influence on sill pillar stability, were analyzed to evaluate the relative importance of insitu stress field and joint properties on design. It was concluded that for stability analysis the initial horizontal stress is of most importance in all aspects of design. The position, orientation, persistence and strength of joints is also of primary concern. In a more general sense, however, it was concluded that detailed geotechnical parameters have much less influence on mine method selection than more general factors such as rock quality, initial stress state, orebody shape and economics.

Jan 1, 1978

By Todd M. Ruff

The majority of fatalities that occur in road construction work zones in the United States involve a worker being struck by a piece of construction equipment or other vehicle. The Spokane Research Laboratory of the National Institute for Occupational Safety and Health, in cooperation with the Washington State Department of Transportation, is evaluating methods to decrease these accidents. One such method uses devices that assist equipment operators in monitoring blind areas around the equipment to prevent collisions with workers on foot or other objects. Several cameras and sensor systems are available for this application, and a study was conducted to evaluate these systems on various trucks used in road construction and maintenance. Tests were conducted using sanding trucks during the winter months, which allowed researchers to investigate the effectiveness and limitations of various technologies under the most extreme conditions. Tests were also conducted using dump trucks during the warmer months to study the effectiveness of the systems in highway work zones. Results show that many difficulties arise when using camera and sensor systems in cold, snowy climates. And, while the operation of these systems is more reliable during the warmer months, challenges still exist in using them on equipment in crowded work areas.

Jan 1, 2004

By R. L. de Beauchamp

A preliminary investigation was made of the extraction of chromium from ferrochromium by low-temperature chlorination in the 3400 to 5250 C range. The objective was to devise a method for separating chromium from iron in high-carbon ferrochromium produced from domestic chromite concentrates. The majority of iron was removed as a volatile chloride; chromium was contained in the residue as a nonvolatile chloride, which was separated from carbon and unattacked ferrochromium by dissolution. A small rotary chlorination apparatus was designed, built, and used for the chlorination tests, and samples of commercial-grade high-and low-carbon ferrochromiums were chlorinated for comparison of chlorination characteristics. Low-temperature chlorination of three types of ferrochromium resulted in the removal of 81.9 to 91.6 percent of the iron content as sublimed FeC13. Recovery of chromium in a leach solution of the reactor product varied from 82.4 to 93.1 percent.

Jan 1, 1968

By Sam G. Demou

Drag bit cutting tests were conducted chiefly to determine bit forces encountered in rock types representing a range of compressive strength and hardness and thus enable the design engineer to estimate a continuous miner's performance in a range of rock types. Testing was carried out on the Bureau of Mines uniquely designed, instrumented rock cutting device (RCD). This device can make 3-ft-long cuts at various depths and speeds to approximate full-scale drag bit cutting by continuous miners. Point attack and cutter bits were tested on four rock types ranging in compressive strength from 7,200 psi (trona) to 29,600 psi (Valders white rock). Cutting tests were made at depths up to 1 in and speeds up to 70 in/sec. Data included normal and cutting forces, groove depth, groove width, and weight of rock broken. The effect of varying the cutting speed was determined, and specific energies were calculated for the various conditions. The bit wear was also observed. For the rocks tested, compressive strength and hardness are good indicators of forces, specific energy, and wear encountered. Results, especially bit wear observations, show the possibility of extending continuous miners, already being used in trona, to Indiana limestone (10,000 psi). Rapid bit deterioration precludes use in Tennessee marble (16,800 psi) and Valders white rock (29,600 psi).

Jan 1, 1983

By Robert H. Merrill

Several investigators have obtained solutions for the concentration of stress around circular, ovaloidal, elliptical, and rectangular holes in an elastic, isotropic, semiinfinite plate subjected to uniaxial and biaxial stress. These solutions are used to evaluate the changes in stress around openings immediately under the undercut normally used in conventional block caving operations. The evaluations are based on the assumption that a horizontal undercut removes most of the vertical component of a biaxial stress field. The location and magnitude of the significant changes in stress concentration around the opening are computed. Conditions under which the assumptions used in the evaluations may or may not be consistent with conditions around block caving mines are discussed, and examples of preferred opening shapes for a biaxial (hydrostatic) stress field are shown. This report is a summary of one phase of an investigation to establish design methods for the size and shape of mine openings used in block caving mining methods .

Jan 1, 1962

By John K. Alley

Despite the extensive use of fluorphlogopite, investigators do not agree as to its exact melting temperature. Data on other fluormicas and on related fluorsilicate systems are largely nonexistent. A knowledge of the melting temperature is necessary for the effective synthesis and use of fluorphlogo¬pite or other fluormicas. The increasing use of two- and three-mica systems to yield polycrystalline ceramics has made more urgent the need for melting temperature information concerning these other fluormicas. This report is part of a program by the Bureau of Mines to synthesize large-size or large-crystal fluorphlogopite and develop a usable reconstituted fluormica sheet. The melting temperature of fluorphlogopite, K2Mg6A12Si6020F4, was found to be 1,383° ±2° C. This temperature was found by studying single crystals of fluorphlogopite, heating behavior of glass of fluorphlogopite composition, and the actual crystallization of fluorphlogopite from a melt that had been heated above 1,430° C. The formation of forsterite or norbergite was directly proportional to the loss of volatiles during the experimentation. Many trials revealed that the only reliable seal was a thick platinum sheet with welded seams. If no volatiles were lost, neither norbergite nor forsterite were formed; fluormica was then the only product. Excellently formed crystals of forsterite with an optical sign different from ordinary forsterite were obtained during the single crystal melting experiments. These crystals resulted from a slow loss of volatiles. A congruent melting of fluorphlogopite is indicated by the data of this report; however, incongruency within a small temperature range is not precluded. A more refined method using phase equilibria-quench methods would be necessary to settle the question absolutely.

Jan 1, 1962

By W. A. McKinney

The segregation process consists of heating crushed oxidized or mixed oxide-sulfide copper are with a copper-volatilizing salt and carbonaceous material at about 750° C. to produce metallic copper which is recovered from the furnace calcines by conventional sulfide flotation. The Bureau of Mines made a number of bench-scale laboratory tests on two copper ores to determine the relative merits of using halogen salts, potassium chlorate, or various salt mixtures as volatilizing agents in segregation of the copper. Laboratory investigation of typical siliceous and calcareous oxidized ores revealed that sodium chloride was a better segregating agent than the other salts. Calcium chloride, magnesium chloride, sodium fluoride, or potassium chlorate gave nearly as good segregation as the sodium chloride on the siliceous ore. Segregation of the calcareous ore with calcium chloride was nearly as good as that obtained with sodium chloride, but inferior segregation resulted when magnesium chloride, sodium bromide, or potassium chlorate was employed. The use of sodium bromide or sodium iodide on the siliceous ore, and of sodium fluoride or sodium iodide on the calcareous are, gave appreciably poorer segregation than sodium chloride or the other salts. Other tests of the ores, using mixtures of sodium chloride and the other halogen salts or potassium chlorate, gave segregation approaching that obtained with sodium chloride.

Jan 1, 1962

By David Oyler, Craig Compton

The The National Institute for Occupational Safety and Health (NIOSH) Pittsburgh Research Laboratory (PRL) is continuing to investigate the behavior of fully grouted roof bolts in the weak roof rock of the Safety Research Coal Mine (SRCM). This paper describes the results of three studies: • A series of 24 pull tests of bolts installed fully grouted and overcored to leave 12 inches of grouted bolt. The tests compared the pull-out performance of offset-head roof bolts with that of standard 5/8-inch bolts; • A second series of 24 tests that compared the pull-out loads obtained by bolts installed with 1 ft of resin (a Short Encapsulation Pull Test or SEPT) to that of fully grouted bolts overcored to leave only 12 inches of grouted bolt, and; • An investigation into the pressures generated during the installation of fully grouted resin bolts. All of the bolts installed and pulled were overcored and removed from the roof after the pull tests to allow inspection of the resin. The first study found that there was no significant difference in pullout load between the offset-head bolts and the standard bolts. The second study, comparing the SEPT bolts and the partially-overcored fully grouted bolts, indicated that SEPT can significantly underestimate bolt anchorage grip factor, suggesting that the SEPT is a conservative measure of actual bolt performance. Significant loss of resin to cracks in the roof was observed in the bolts overcored in the SRCM. The third study explored one possible explanation for the resin loss, high pressures generated during bolt installation. These tests employed bolts installed in strain gauged steel tubes. The results confirmed that significant pressures (greater than 4,000 psi) can be generated during bolt installation. The results of the pressure tests, combined with the observations of resin loss in bolts installed in the SRCM, suggest that under some circumstances resin loss and under-encapsulation could take place and lead to degraded bolt performance. Glove fingering was observed in all recovered bolts to various degrees. The limited comparative data suggested that the effect of glove fingering ranged from slight to moderate. Complete mixing of the resin was observed on all of the 40 bolts recovered (28 fully grouted and 12 SEPT). Australian and U.S. resins were compared and the Australian resins were observed to have much smaller catalyst compartment areas and lower resin viscosities. These differences are suggested as potential explanations for occasional resin mixing problems noted in Australian resin bolt installations.

By Robert C. Bates

Variations in water percolation rates through saturated cohesionless hydraulic backfill (mill tailings) have been studied in relation to the many physical aspects of the system to determine the most important factors in causing these variations. A multivariable least-squares regression analysis of 135 separate tests shows that a reliable estimate of the water percolation rate through typical hydraulic backfill can be made if the bulk density (unit weight), specific gravity, and grain-size distribution are known or can be estimated. Equations for predicting percolation rate, coefficient of permeability, and seepage velocity are developed both for the range of data analyzed and for extrapolation. The prediction equations for use within the range of data studied are quite accurate, while the ones developed for extrapolation are less accurate but still give reasonable approximations of flow rates. The equations given are valid for cohesionless materials that are free of clays and micas, exemplified by classified mill tailings. The equation developed for predicting percolation rates is as follows: 1nP20 = 11.39147 + 2.853422 1n(VR X D10) + 0.1747436 VR X Cu -178.8039 D10 X D50 + 311.7034 (DSO )3, (1) where P20 = percolation rate, inches per hour at 20° C; correction factors for other temperatures are given in the text, VR = void ratio, dimensionless, D10 = effective size (diameter in millimeters of the 10-percent-finer grain size), Cu = coefficient of uniformity (ratio of the diameter of the 60-percent-finer grain size over D10), D50 = average size (diameter in millimeters of the 50-percent-finer grain size), and In = natural logarithm (log to the base e). The prediction equation for percolation rate is rather complicated, so for practical use a nomograph that simplifies the solution of the prediction equation was developed.

Jan 1, 1967

By Mandius C. Lundal

This report describes the work done to develop a coal mining machine to increase productivity and recovery from thin seams beyond the highwall, in addition to decreasing the miner's susceptibility to to highwall hazards First, radar transducers were adapted to measure coal thinkness and detect rock and coal interfaces; Then, the radar was incorporated into a cutterhead with steering controls. Next, visual displays were designed to show cutter position in the coal seam, and anticipated change in coal seam inclination, Finally, development of a telemetry system made remote control of the cutterhead possible. Simultaneously, a high-performance extended depth (HPxD) auger machine was designed with an automatic system to handle 600 ft of 36-in, diameter dual spiral flight conveyor sections. A two man crew controls all operations from a weather conditioned cab outside the VDZ, either at night or day time. As a primary mining tool, working 4000 hr/yr (2 shifts/day) recovering 60% of the coal 600 ft beyond the highwall, productlon to 500,000 ton/yr can be achieved. Volume I provides an overview of machine, systems and productivity.

Jan 1, 1977

By Ellisworth R. Spencer, Lynn A. Alcorn, Peter G. Kovalchik, Adam K. Smith

Noise induced hearing loss (NIHL) is the most common occupational disease among workers in the mining industry. Previous studies conducted by the National Institute for Occupational Safety and Health (NIOSH) have shown that approximately 90% of coal miners and 49% of metal/non-metal miners have a hearing impairment by age 50. The Mine Safety and Health Administration (MSHA) has determined that continuous mining machines rank first among all equipment in underground coal mining whose operators exceed 100% noise dosage. The conveying system is one of the principal noise sources on continuous mining machines, due to metal on metal impacts that occur between chain flights and the conveyor deck. A highly durable polyurethane coating has been developed for the chain flights to decrease noise generated by these impacts. A continuous mining machine retro-fitted with coated flight bars has achieved overall sound level reductions of 5-7 dB in a laboratory setting. This paper evaluates the effectiveness of this engineering noise control in reducing the noise exposure of continuous mining machine operators in an underground coal mine environment. The results show that the operator of a continuous mining machine utilizing the coated flights receives a predicted 35% reduction in total work shift noise dose.

By L. H. McGuire, L. B. Berger

"During the period October 31 to November 4, 1944, tests were conducted by the Bureau of Mines in the Cascade Tunnel of the Great Northern Railway Cc. to determine the effects upon the tunnel atmosphere of the use of a Diesel freight locomotive through the tunnel. These tests were conducted at the request of the company to ascertain whether undesirable conditions might be caused by the discharge of Diesel engine exhaust gas in the tunnel, which was constructed for an electric haulage system and is not provided with equipment for mechanical ventilation. The use of Diesel locomotives on the division that includes the Cascade Tunnel was under consideration as a means of augmenting haulage facilities for the abnormal freight tonnage that must be handled during wartime.As the Bureau of Mines has received requests for information on the extent of contamination of the air of railroad tunnels by Diesel locomotives, this report was prepared to make available for general information the data obtained during the tests in the Cascade Tunnel. As the test locomotive is of a type used on other railroads, and as the bore and grade of the tunnel and the tonnages moved are comparable to such factors on other roads, the data should be sufficiently representative to be of general interest. The exceptional length of the Cascade Tunnel made it an excellent location for obtaining information on the composition of the tunnel air and on the movement of air induced in the tunnel by the passage of trains.The writers are indebted to C. O. Jenks, vice president of the Great Northern Railway Co., who arranged for the tests, and to the following personnel of the railroad who cooperated in the conduct of the tests: I. E. Clary, superintendent, Spokane Division, J. L. Robson, superintendent of motive power, oil and gas equipment, J. F. N. Gaynor, superintendent of electric operations, William Lowney, master mechanic, Spokane Division, and D. F. Marston and Verne Nelson, assistant superintendents of electric operations.The Cascade Tunnel is on the Spokane Division of the Great Northern Railway and pierces the Cascade Mountains some 49 miles west of Wenatchee, Wash. The west end of this division is electrified from Appleyard Station, 2 miles east of Wenatchee, to Skykomish, Wash., a distance of 73 miles. Until the described tests, only electric haulage had been employed through the tunnel since it was first opened to traffic in January 1929.The Cascade is the longest railroad tunnel in the United States - 7.79 miles. The tunnel is single-track, perfectly straight, concrete-lined, and is on an essentially uniform grade of 1.56 percent, ascending from west to east. The cross-sectional area of the tunnel is 314 square feet, the dimensions being shown in figure 1. Elevation of the east portal is 2,880 feet and of the west portal, 2,240 feet."

Jun 1, 1946

By J. E. Chilton

A technique for use in underground coal mines measures the response times of machine-mounted methane monitors. This technique uses a calibration cup to deliver gas to the sensor head to measure response times. For comparison, in the current study, response times were measured by placing methane sensor heads in a specially designed test box where they were exposed to an atmosphere containing a constant concentration of methane. Response time readings measured in the test box were shorter than the readings obtained using the calibration cup. The response times measured in the test box were comparable for sensors from three different manufacturers. The effects of dust cap and flame arrestor design on response times were also evaluated using the test box.

By John E. Kelley

A method for measuring the spark-shower radiance of metals when abraded on a grinding wheel was developed by the Bureau of Mines for evaluating the sparking tendency of metals. The radiance was measured in units of emissive power per solid angle per area of the detector. Over 100 metals were tested and rated relative to commercially pure iron. The metals tested included elemental metals, iron binary and ternary alloys, iron-carbon ternary alloys, commercial steels, nonferrous alloys, and a few hard-metal compounds. Some alloys were synthesized; others were obtained commercially. Alloys that possessed good strength properties were emphasized for selection. The results indicate that the spark-shower radiance of plain carbon steels and carbon alloy steels was greater than that of pure iron. Certain elements, particularly vanadium and chromium, reduced the sparking tendency of iron. Stainless steels and particularly nickel alloys displayed low spark-shower radiance. Some metallic elements had much lower radiance than iron; others, much greater. Plain carbon steels, heat treated to a variety of hardnesses, gave radiance values that generally increased with increasing hardness. The effect on sparking tendency of an increase in hardness of an alloy steel was studied. The sparking tendency increased with an increase in hardness to a maximum, then decreased with a further increase in hardness.

Jan 1, 1974