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By R. J. Morrell

The Bureau of Mines conducted a series of laboratory experiments to test a cutting technique for hard rock using large-scale drag cutters. Cutting experiments were performed on rocks ranging in compressive strength from 10,000 to 27,000 Ib/in2, using large drag cutters from 3 to 6 in. in width. The tests were conducted on a special test apparatus, called a ripper tester, which made a curvilinear cut across the rock sample while measuring the cutting force acting on the drag cutter. The cutting method that evolved from these experiments is called ripper cutting, and its energy efficiency appears superior to all other large-scale mechanical fragmentation techniques. In addition, the method cre-ated very little dust, and the wear on the bits was negligible. A fragmentation system based on ripper cutting was then devised, and a large-scale test device was fabricated to allow full-scale laboratory testing of the system. The results of these tests will be published in future Bureau reports.

Jan 1, 1986

By F. P. Haver

The Federal Bureau of Mines developed a 10,000-ampere cell to determine the economic feasibility of producing high-purity electrolytic titanium from mill scrap or offgrade sponge. Operating variables were investigated, matetial handling procedures were established, and the cost of electrorefining titanium was determined. The prototype cell produced premium-grade metal (<100 Brinell hardness number (B.h.n.)) from very impure Kroll sponge at a processing cost of less than $0.60 per pound. Refined metal was deposited at the cathode in the form of coarse (85 percent plus 34-mesh), hexagonal crystals of a titanium when the cell was operated an 850° C. A cathode current density of 1,000 amperes per square foot and a molten sodium chloride electrolyte containing approximately 5 percent by weight of soluble titanium were used. In continuous operation, at the optimum number of ampere-hours, (40,000), the current efficiency averaged slightly less than 60 percent. It was possible to supply large enough amounts of electric current for electrolysis to maintain the temperature of the bath, thus reducing overall power consumption to less than 5 kilowatt-hours per pound of metal refined.

Jan 1, 1961

By D. E. Kirby

In line with the Bureau of Mines&apos; goal of providing technology to maintain a supply of minerals adequate to meet national economic and strategic needs, a method for recovering chromite from lateritic materials was investigated. Laterites are a large resource of Ni, Co, and Cr, critical metals of which the United States has insufficient domestic sources. A roast-leach method was developed by the Bureau to extract nickel and cobalt from domestic laterites. A concurrent effort was made to recover chromium from either unprocessed laterites or the leach residues. This report characterizes laterites and laterite leach residues. It describes a chromite recovery procedure using conventional industrial beneficiation equipment for sizing, low-intensity magnetic separation, and gravity concentration. Concentrates containing 20.0 to 22.4 pct Cr (29.0 to 32.7 pct Cr203) and with a 1.6:1 to 1.9:1 Cr-Fe ratio were recovered from laterites and residues containing 1.3 to 2.5 pct Cr with recoveries exceeding 50 pct. Because of their low grade and fine size, these concentrates would be suitable for preparing chromium chemicals. The direct beneficiation of laterites for chromite recovery resulted in the minerals containing nickel and cobalt concentrating in slime products. No iron oxide coproducts of acceptable grade were recovered from either laterites or residues.

Jan 1, 1982

By Christopher Mark

Australia and the United States both have large, modern underground coal mining industries. Many companies have invested in both countries, and there is growing awareness that technological developments in one country can rapidly affect the other. Yet the ability of technology to pass between the countries depends on comparability of ground conditions. Many Australian observers believe that their ground conditions are significantly poorer than typical U.S. conditions, but some U.S. professionals are not convinced. The author recently visited 10 Australian longwall mines and two room-and-pillar mines. Each mine visit included determination of the Coal Mine Roof Rating, roof support installed, pillar design, and an in-depth discussion of ground control experience. The Australian conditions observed are compared with an extensive data base from U.S. mines. The paper also offers observations on geotechnical data collection, monitoring, and the roles of the mine inspectorate and research community.

By Joseph Cervik

Gas occurs in coalbeds in an adsorbed and a free gas state, Adsorbed gas is stored in the micropore structure and its transport is governed by Fick&apos;s law. The free gas occurs in the. fracture system and flows according to Darcy&apos;s law. These two modes of mass transport are interdependent. Production decline curves of coal-gas wells are of the constant percentage decline type and, thus, show no indications of flow characteristics peculiar to coal-gas reservoirs. The effectiveness of surface boreholes as a degasification scheme depends upon both good fracture permeability and a high fracture density. Conventional methods of reservoir engineering analysis are not applicable to coalbeds.

Jan 1, 1969

By Thomas A. Phillips

The Bureau of Mines presents an economic evaluation of its leach- electrolysis process for recovering lead from galena concentrate. In this process, galena concentrate is leached at 200° F with a ferric chloride-brine solution to extract lead while leaving sulfur in the residue, which is discarded. The solution after leaching is concentrated and cooled to 100° F by vacuum evaporation to crystallize lead chloride from which metallic lead and chlorine are produced by fused-salt electrolysis. Chlorine is used to regenerate the spent leach solution before it is recycled. A cost estimate was prepared for a plant capable of producing 300 tons per day of lead, 330 days per year. The fixed capital cost on a third-quarter 1977 basis Fs $27,500,000, and the estimated operating cost without a charge for the concentrate is 4.9 cents per pound of lead product. Assuming a selling price for corroding-grade lead of 31 cents per pound and a charge for the concentrate of $332 per ton, the interest rate of return on investment after taxes is 16 percent. The leach-electrolysis process, therefore, appears to be economically attractive under the assumed market conditions.

Jan 1, 1978

By Thomas M. Barczak

Roof support systems are necessary to provide stable mine openings and much research has been conducted to design a variety of roof support systems that will function in various manners to ensure that stable ground conditions are achieved. Despite these advancements in technology, mistakes continue to be made in the evaluation and/or installation that significantly degrade the support capability or lead to erroneous determinations of support expectations. The purpose of this paper is to discuss misconceptions about how roof supports perform and factors that impact their performance. The goal is to present practical information that will assist mine operators and engineers in selecting, installing, and evaluating roof support systems properly, and help them to avoid mistakes that can lead to erroneous expectations and potentially catastrophic results that may lead to roof falls. The paper is limited to a discussion of secondary roof support systems and powered roof supports such as longwall shields.

Mining officials, as well as members of the Bureau of Mines staff, in their efforts to make mining safer are confronted from time to time pith complicated problems concerning solution of which there is a diversity of opinion. To reduce these differences of opinion as much as possible and to obtain a generally acceptable decision on any mine-safety question, the Director of the U.S. Bureau of Mines established the Mine Safety Board on June 1, 1924- *** to consider questions arising within any divisions of the Bureau that require a definition of tie Bureaus s collective opinion as to safety practices, safety devices, or safety methods for underground operations or open-pit mining. The approved2/decisions shall form the basis of teaching and policy for the Bureau.

Jan 1, 1935

The Bureau of Mines, U.S. Department of the Interior, devised a flotation technique for upgrading chromite in chromium ores and gravity concentrates. The technique consisted of cationic flotation of chromite from a slime-free pulp acidified to pH 2.5 or lower using H2S04 modifier. Flotation results showed that ore containing serpentine as the primary gangue mineral was more amenable to concentration by this technique than ore containing the readily floatable olivine as the primary gangue mineral. Rougher flotation of a chromite ore containing abundant serpertine increased the grade form 19 to 41 percent Cra03 at a chromite recovery of 89.6 percent. Rougher flotation of ore containing abundant olivine upgraded the chromite from 23.4 to 29.6 percent Cr203 at a recovery of 89.4 percent. The technique was also applied to chromite gravity concentrates to further upgrade chromite and to reject silicates. Flotation of a gravity concentrate containing 40.9 percent Cr203 and 4.4 percent SiO2 recovered 90 percent of the chromite in cleaner products containing 44.9 percent Cr203 and 0.63 percent Si02. The combined recovery by gravity and flotation was 87 percent. An economic evaluation of this technique will be undertaken to deter-mine its feasibility for commercial use.

Jan 1, 1980

By J. C. Franklin

An experiment was conducted by the U.S. Bureau of Mines and the Canada Centre for Mineral and Energy Technology (CANMET) to measure the emanation rate of radon from cemented unclassified mill tailings in two dewatered stopes of a uranium mine at Elliot Lake, Ontario, Canada. The radon was monitored upstream and downstream and in both stopes to determine the concentration in the stopes and the amount exhausted from both stopes. The emanation rate was determined to be 0.55 pCi/cm2 sec-1 from the mill tailings, which is approximately 1,000 times higher than the average flux in the mine before backfilling.

Jan 1, 1982

By James F. Devine

The radial, vertical, and transverse components of particle velocity were recorded by Bureau of Mines investigators along gage arrays extending in one or two directions for 145 to 3,170 feet at five quarries. Of the 39 quarry blasts, 12 were instantaneous blasts, 5 were of the one hole per delay type, using millisecond delayed caps, and 22 were multiple hole per delay type employing millisecond delay detonating fuse connectors. Charge weight per hole ranged from 8 to 1,500 pounds, and the charge weight per delay interval, including the instantaneous blasts, ranged from 25 to 4,620 pounds.

Jan 1, 1966

By J. E. Spurr

"Zinc Industry in BelgiumBy March F. Chase, Consulting Metallurgist, Bureau of MinesBelgium for many years was the principal spelter producing country of the world. Gradually its production was surpassed by that of Germany, and Germany in turn lost first place by the production in the 77ti.tad States. For the ten years previous to 1914, 21.6 per cent of the world's spelter was produced in Belgium. Of this production approximately 40 per cent was exported. A large percentage of the balance was rolled and of the shoots thus produced a considerable quantity was also exported.From an economic standpoint even before the war, the Belgian industry was in a unique, if not praoarious, position. Dependent upon foreign countries for ore and foreign markets for sale of products, smeltor was at a distinct disadvantage when compared with some of those of other localities. These disadvantages were partly offset by cheap coal and labor and by whatever there is of value in experience and what we may term for want of a better expression smelting tradition. However intangible this latter may be, it did have and has today an influence on commercial operations.One must not overlook the effect it will have on the revival of the industry. It may not be the determining factor but the fact that for so many years this industry flourished in Belgium, and played such an important role in the industrial life of the country will decide many operators to attempt to re-establish their operations.In 1913 the Belgian spelter plants produced 204,220 long tons of spelter or approximately 20 per cent of the world's production. During this year 500,000 to of ore were treated. The actual value of all of the products derived from treating sine ore amounted to 225 million frames."

May 1, 1919

By Stephen R. Wilson

This paper is one of a series published by the Bureau of Mines on the mineral resources of the Nation. It describes briefly the occurrence of sodium sulfate in Salt Lake and Tooele Counties, Utah, presents data on sampling, and describes ex-traction and refining raw material and a mining research project that has been conducted on the deposits. ACKNOWLEDGEMENTS Acknowledgments are given to the Tintic Standard Mining Co., Salt Lake City, Utah, and Bartlett Chemicals, Inc., New Orleans, La., for permission to use data obtained by the companies. The authors also acknowledge the courtesy of J. A. Marsh and H. K. Thurber, geologists and mining engineers, for permission to use information in reports prepared by them.

Jan 1, 1957

By Nicholas Kyriazi, John P. Shubilla

A joint effort by the U.S. Bureau of Mines and the U.S. Mine Safety and Health Administration (MSHA) was undertaken to determine how well self-contained self-rescuers (SCSR&apos;s), deployed in accordance with Federal regulations (30 CFR 75.1714), survived the underground environment with regard to both impact damage and aging. This report presents findings regarding laboratory-tested SCSR&apos;s from 1982 through 1990. The SCSR&apos;s were tested on human subjects and on a breathing and metabolic simulator (BMS). These results indicate that most of the apparatus, if they pass their inspection criteria, perform as expected except for units with manufacturing defects or design deficiencies. However, when the apparatus are carried in and out of the mine daily and stored at the working section, they may suffer abuse. Physical signs of abuse, unless extremely obvious, are frequently not detected by the miners or mine operators. This poses a potential danger to a user in an emergency. Recommendations include improved training in inspection procedures.

Jan 1, 1992

By L. C. George

The Bureau of Mines measured the vapor pressures of selected mixtures in the binary systems CdCl2 ?SnCl2, CdCl2 -ZnC12, PbCl2 ?SnCl2, and PbCl2 -ZnCl2 by the static pressure method. These data were evaluated by the second-law (sigma-plot) method. Equations expressing the vapor pressures and the heats and free energies of vaporization as functions of temperature are given.

Jan 1, 1969

By Frank D. Lamb

Pyrophyllite, a hydrous aluminum silicate (Al2 Si4O10 (OH)2 ) has been mined in North Carolina and marketed without beneficiation for the past 100 years, but recent demands for higher-grade material indicate that an economical concentration method is needed for at least a part of the pyrophyllite production. Past production has depended upon selective mining and hand sorting, which is often unsatisfactory because of the finely disseminated character of the ores. The major impurity in North Carolina pyrophyllite ores is quartz, which accounts for the abrasiveness of the commercial product. Since pyrophyllite is only slightly heavier than quartz, those ores cannot be beneficiated by simple gravity-concentration methods. Some separation can be attained by differential grinding and air classification because of the difference in hardness of the two minerals) but the separation is not sharp. Froth flotation of pyrophyllite from Quartz has never been attempted commercially, but if the demand for high-grade material is sufficient, it would seem that this process offers the most premise of being economically feasible.

Jan 1, 1950

By Lloyd A. Morley, Alan M. Christman

Presented below is a review of how the objectives, as outlined in the statement of the problem, were achieved. 1. Methods of evaluating the performance of grounding systems must be described. These methods are the subject of Chapter III. 2. Field trips to operating mines should be undertaken, and a cri¬tique of their grounding systems made upon the basis of obser¬vation and the application of the above techniques. An analysis of field trip data, along with practical recommendations, is in¬cluded in Chapter IV, while a complete review of all field trips is given in Appendix I. 3. Because of their thoroughness, present coal mine regulations should be evaluated, and their applicability to open-pit environ¬ments appraised. This has been done in consultation with governmental agency, manufacturing, operating mine, and consult¬ing engineers. As stated in Chapter IV, insight gained through field observations has served to amplify and extend the base of knowledge from which the recommendations were made. 4. The results of these investigations should lead to suggested Federal standards. Such recommendations comprise Appendix II. In addition, several important factors, which are discussed in Chapter IV, were brought to light as a direct result of the field research. 5. Seasonal variations in the earth resistance of ground mats, as illustrated by Table 12, reinforce the need for periodic

Jan 1, 1976

By G. B. Atkinson

The Bureau of Mines, in cooperation with McDermitt Mine, investigated a method for recovering mercury metal from mercury sulfide flotation concentrate by a hydrometallurgical technique. This preliminary report shows the feasibility of leaching mercury with a cupric chloride solution and electrowinning mercury from the solution. Proposed chemical reactions are shown in equations 1 through 3: 2CuC12 + HgS + 2CuCl + HgC12 + S°; (1) electrolysis HgC12 ?Hg + C12; (2) 2CuC1 + C12 ? 2CuC12. (3) A bench-scale leaching-electrowinning apparatus capable of treating 2.36 kg of sulfide concentrate in 11 days was operated continuously. An acid chloride solution containing, in grams per liter, 70 Ca, 60 Cu, 30 Fe, and 80 Hg was used as both leachant and electrolyte; 99.5 pct of the mercury was solubilized from the sulfide concentrate. Electrolysis of the leaching solution recovered mercury metal at an average current efficiency of 32 pct and an average energy consumption of 1.2 kwhr/lb Hg. Results from a preliminary economic evaluation indicated that a plant producing 31,500 flasks of mercury per year would have a capital cost of $3.7 million and an operating cost of $62 per flask of mercury.

Jan 1, 1983

These proceedings consist of an introduction and four descriptive papers. The first paper is an update on fire protection for large haulage vehicles, the second paper is concerned with improved visibility systems for large haul- age vehicles, the third paper is on improved ladders for large haulage vehicles, and the fourth paper is concerned with large haul-age vehicle operator training systems.

Jan 1, 1980

By R. K. Clifford

Chalcopyrite (CuFeSz) concentrates produced on the Lead Belt of Missouri typically contain 28 pct Cu and from 0.5 to 1.5 pct each of nickel and cobalt. Most of the nickel and cobalt are present as the mineral siegenite, (Ni,Co)3S4, almost none of which is recovered presently. The Bureau of Mines, U.S. Department of the Interior, has been conducting research on the chalcopyrite to produce a marketable fraction enriched in cobalt and nickel. Ferric chloride leaching of various chalcopyrite (copper) concentrates has yielded residues containing from 10.2 to 16.2 pct Ni and 8.3 to 11.4 pct Co with recoveries of approximately 79 pct. In another approach, laboratory froth-flotation tests have been conducted to investigate the effects of regrinding, roasting, baking, and steaming prior to flotation, with a variety of reagents. The best laboratory results were obtained after regrinding the copper concentrate to 94 pct minus 400 mesh and floating the chalcopyrite at natural pH (6.5) with 0.48 pound of sodium diethyldithiophosphate and 0.01 pound of sodium cyanide per ton of concentrate. This procedure yielded a tailings assaying 20.4 pct Cu, 5.2 pct Ni, and 3.45 pct Co. The tailings contained 67.2 pct of the nickel and cobalt and 15.9 pct of the copper originally in the copper concentrate; the remainder of the values reported to the flotation concentrate, which analyzed 29.6 pct Cu. To verify the laboratory flotation results, a 60 kg/hr process demonstration unit (PDU) was constructed in the Cominco American Magmont mill, Bixby, Mo.

Jan 1, 1978