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By Oren C. Baptist

[Several new oilfields have been developed in northwestern Wyoming during the past 10 years, production is obtained from discontinuous, permeable sands enclosed within impermeable shales. The energy to drive the oil to producing wells in supplied by gas in solution in the oil. Most of the wells had high-initial potentials, but rates of oil production have steadily declined, indicating a dissipation of the driving energy and the possibility that large volumes of oil will remain unrecovered in the reservoirs. The studies reported in this paper were undertaken to develop laboratory data of the type most likely to be needed by persons studying the susceptibility of the Newcastle sand reservoir to secondary-recovery methods.]

Jan 1, 1957

By Paul S. Goodman

This report examines the effectiveness of bonus plans in the coal industry. The analysis uses theoretical, literature review and empirical approaches to determine the impact of bonus plans on safety and productivity in underground coal mines. The basic finding is that the modal bonus plan in underground coal mining does not have a major impact on productivity and safety indicators. The design of the plan, or organizational factors, and the process of introducing the plan all contribute to its relative ineffectiveness. Strategies and options for designing effective plans are presented.

Jan 1, 1987

By J. A. Ruppert

The Bureau of Mines investigated methods of refining iron-contaminated zinc to marketable quality and developed a laboratory-scale process in which aluminum, added as a refining agent, reacted with the iron, forming solid Fe2Al which floated to the surface of the melt. Solids were separated by both filtration and centrifugation, using a unique type of "dipping" centrifuge devised during the work. Typical zinc recovery was 95 percent, and iron content of the metal was lowered from 0.37 to 0.02 percent.

Jan 1, 1967

By Elio Ferreira

This report describes the design, and development work performed to reduce the volume and weight of the miners caplamp battery, as well as a headpiece with a swivel mechanism containing a halogen bulb with two filaments. Work was also conducted in the development of a maintenance free battery. The design concept is based on non-sintered Ni-Cd technology, with an energy density of 22 Wh/lb. The cap lamp system developed has been approved by the Mine Safety and Health Administration (MSHA) and can be manufactured for mine use.

Jan 1, 1988

By Federick W. Miller

The Bureau of Mines has investigated the availability of columbium from 19 deposits (3 producers and 16 nonproducers) in 7 market economy countries (MEC?s). Brazil has the largest recoverable columbium resource, posessing approximately 2.39 million metric tons Wmt), or 69 pet of the MEC total (3.47 Mmt). The Western Hemisphere countries of Brazil, Canada and the United States contain 3.14 Mrnt of recoverable columbium or 90 pnt of the MEC total. Columbium recovered as a byproduct of tin mining is not included as part of MEC resources in this study. Currently producing mines are Araxa and Catalao in Brazil, and Niobec in Canada, These three mines have 2.42 Mmt of recoverable resources or approximately 69 pct of the MEC total. It is estimated that these mines, operating at full capacity, could satisfy cumulative MEC demand (based on an annual growth rate of 5.1 pct) for the next 20 yr. The United States is totally dependent upon foreign sources for its columbium supply. The majority of U.S. columbium is imported from Brazil in the form of steelmaking- grade ferrocolumbium. Canada has been the sole source of pyrochlore concentrates since Brazil discontinued exports in 1981. The United States also imports small (but important) quantities of columbium as columbite concentrates from Nigeria and Brazil, as columbite contained in tin slags from Thailand and Malaysia, and as synthetic concentrates from the Federal Republic of Germany.

Jan 1, 1986

By F. P. Haver

Because of increasing concern with air pollution, the Bureau of Mines has developed a hydrometallurgical procedure to obtain elemental sulfur from chalcopyrite (CuFeS2) by means of the following reaction: CuFeS2 + 3FeCl3 ? CuCl + 4FeCl2 + 2S. Under optimum conditions, 99.6 percent of the copper and 71.6 percent of the iron were extracted in 2 hours from a typical mill concentrate containing 75 percent CuFeS2 and 15 percent pyrite (FeS2). At the same time, 68.2 percent of the sulfur in the concentrate was oxidized to the elemental form. The leach was highly selective; gangue minerals, including pyrite, were not attacked, thus accounting for the relatively low recoveries of iron and sulfur. To obtain these results, the concentrate was ground to 98.9 percent minus 325 mesh, and 300 grams was treated with 1 liter of boiling ferric chloride (FeCl3) solution. The FeC13/CuFeS2 ratio was 2.7. Cement copper containing 99.9 percent copper was recovered from the leach solution by cementation on sponge iron. Excess iron, dissolved during leaching and cementation, was removed from the stripped solution by crystallization. Ferrous chloride (FeCl2?4H2O), obtained as a byproduct, was converted to high-purity ferric oxide (Fe203) and HCI by an oxidizing roast. The hydrochloric acid was then electrolyzed to obtain chlorine, which was used to regenerate the leach solution. The leach residue, which contained ~50 weight-percent elemental sulfur, was treated with an organic solvent to recover the sulfur and cyanided to reclaim gold and silver.

Jan 1, 1971

By A. E. Raddatz

Previous Bureau of Mines research has shown that tungsten ores containing as little as 40 pct W032-can be processed by a high-temperature molten-salt extraction technique to produce a tungstate-bearing sodium chloride phase with 99 pct W recovery and little impurity contamination. The objective of this work was to demonstrate that the tungstate-bearing sodium chloride phase can be a suitable feed material for preparing am-monium paratungstate (APT) by a modification to the present industrial solvent extraction process. A combined crossflow-count·ercurrent flow solvent extraction technique to extract tungsten is presented. APT produced by this new technique contained 88.6 pct W032-and the following impurities, in percent: Mn02_ 0.0007, Mo02 0.017, P205 <0.010, and Si02 0.012. Implementation of this processing technique would improve tungsten recovery from concentrates and simplify processing operations.

Jan 1, 1988

By V. E. Edlund

This report presents the results obtained in a study of the technical feasibility of recovering a usable pig iron by pyrometallurgical treatment of typical copper reverberatory slags . The iron in the slag was readily reduced with carbon , but the copper and sulfur content of the metal was best controlled by a three - stage procedure comprising ( 1 ) melting the slag with pyrite to remove copper as a low- grade matte , ( 2 ) smelting the partly decopperized slag with fluxes and carbon to reduce the iron , and ( 3 ) refining the impure iron alloy from step 2 with a flush slag to remove sulfur and residual copper . Laboratory - scale tests demonstrated that washing with 8 wt pct of pyrite lowered the copper content of a typical reverberatory slag from 0.58 to 0.12 Subsequent smelting of a mixture composed of 70 pct decopperized slag , 25 pct lime , and 5 pct coke reduced 84 pct of the iron to metal containing 2.27 pct sulfur . This metal then was successfully desulfurized by treatment with a basic flush slag to produce a finished metal that contained , in percent , 93.5 Fe , 0.02 S , 0.26 Cu , and 4.9 C. Overall recovery of iron under the conditions of trial was about 80 pct . This could be increased if higher tolerances for copper could be accepted . Commercial processing would be accomplished in three furnaces with molten materials passing from one to another . Currently produced smelter slag could be treated fairly inexpensively , because the molten product would require only a little additional heat to produce an iron byproduct that undoubtedly would be attractive to local foundries .

Jan 1, 1964

By W. H. Oppelt

A lignite from west-central North Dakota (Mercer County) was dried by the Bureau of Mines by the Fleissner process at saturated steam pressures to 1,500 pounds per square inch absolute (p.s.i.a.). Normally, the process is operated at pressures of 200 to 370 p.s.i.a. At the highest pressure the temperature of the saturated stearn is 596° F. The purpose of the investigation was to evaluate the influence of very high stearn pressure and temperature on moisture removal, steam consumption, and physical and chemical characteristics of the dried lignite. Two drying cycles were used: A normal cycle, corresponding to previously established standard time requirements, and an accelerated cycle, wherein time requirements were determined by boiler capacity and handling speed. Blank tests to determine stearn requirements for heating the empty autoclave were performed for each cycle and each test condition. Moisture removal, final moisture content, net steam consumption, and size stability of the dried lignite were determined.

Jan 1, 1959

By O. C. Blade

This report on the properties of motor fuels sold through service stations in the United States was made in accordance with a cooperative agreement between the American Petroleum Institute and the Bureau of Mines of the United States Department of the Interior. This paper is a continuation of a series made in cooperation with the Coordinating Fuel Research (CPR) Committee of the Coordinating Research Council (CRC), Inc. The first of this series of reports presented data on motor gasolines sold during the winter of 1935-36, and succeeding reports described motor gasolines sold during the summers and winters to and including the summer of 1941. No reports were made on motor gasolines sold during the winter of 1941-42 and during the summer of 1942. However, starting with the winter of 1942-43, the semiannual reports were resumed.

Jan 1, 1949

By L. Crocker

The Bureau of Mines investigated processes for recovering a marketable lithium product from the montmorillonite-type clays of the McDermitt caldera. The clay deposit, located on the Nevada-Oregon border, is estimated to contain over 3 million short tons of lithium. Limestone-gypsum roasting and selective chlorination proved most successful for extracting lithium from the clays; 80-pct Li recovery was achieved using either technique. With both processes, lithium silicate in the clay was converted to a water-soluble compound-either lithium sulfate or lithium chloride. The lithium was then recovered as a carbonate by water leaching the calcine, concentrating the leach solution by evaporation, and precipitating the lithium with soda ash. The operating costs for the limestone-gypsum roast process were estimated at $2.02/lb Li2C03; selective chlorination operating costs were estimated at $3.85/1b Li2C03. The current selling price of Li2C03 is about $1.50/lb. A detailed flowsheet, material balance, and process evaluation are presented for the limestone-gypsum roast process. In addition, lithium reserves and resources are discussed and the McDermitt clays are described.

Jan 1, 1988

By Ellis S. Hertzog, James R. Cudworth

Tremendous interest has been evidenced in the subject of dewatering and drying ooal for the past 10 years, and yet little effort has been made to summarize the literature, at least so far as this country is concerned. Only a few published articles have dealt solely with dewatering and drying coal. In most cases the subject has been discussed briefly, either in general terms or as to specific installations. It is desirable that a compendium be prepared collecting in one place information on the dewatering and drying of coal, the needs for this process in various fields, the available equipment, and the current American practice in regard to it. This work was underta..lten by the :Bureau of Mines in cooperation with the University of Alabama. New coal-washing plants in which various types of dewatering and drying equipment were in operation were visited in-Pennsylvania; Ohio, India...~a~ and Illinois. The dewatering·and drying of coel has .been studied in this coimtry by the Surface Preparation ·Committee of the ·coal Operato~s• Committee of the .American Mining ·Oongress. In GermaIJY .it has received the special attention of the Reichskohlenr·at, and in Great Britain of the Utilization of Fuel Commi-t-tee of the Insti ti.ite of Min"ing Engineers.

Apr 1, 1938

By Ludwig Schmidt

"IntroductionThe evaporation losses of gasoline in the refinery, can be reduced by proper methods of refinery construction and operation. When gasoline evaporates slowly, as when stored in tanks that are not vapor tight, the most volatile part evaporates first, so that the remaining liquid in not only less in volume, but is also lower in quality. In some,evaporation changes the characteristics of the product so much that it must be reblended to meet trade specifications.Refiners are, therefore, greatly interested in the various appliances and methods used to reduce these losses. In the past this interest has been directed largely to storage tanks, probably because evaporation from tanks is more readily observed than losses during other refinery operations. As tests by the Bureau of Mines on evaporation of gasoline from small storage tanks had indicated that the losses during the various stages of refining might be equal to or greater than those of standing storage, a survey was during the spring and early summer months of 1924 in a large number of refineries in the Mid-Continent field. In order to enable the refiner to select the results of those tests that conform most nearly to the conditions that he has to meet, tests were conducted at various stages of refining, as follows: (1) Randown tanks, (2) (agitators), (3) filling tank cars, (4) storage of refined products, and (3) miscellaneous handling. The results are presented in this paper."

Nov 1, 1925

By L. E. Green

For shortwall mining to be profitable, the additional capital investment of about $1 million for shortwall roof supports must be offset by decreased total operating costs and/or increased coal production. The Bureau of Mines therefore conducted a demonstration project to compare shortwall and room-and-pillar mining. Daily two-shift production from the bureau of Mines-Beth-Elkhorn Corp. shortwall demonstration n project averaged 975 tons of raw coal, compared with 905 tons from a room-and-pillar unit operating undo r similar conditions. The estimated direct operating costs for the shortwall and room-and-pillar unite were $4.21 and $4.03 pier ton, respectively. The profitability of the shortwall system was found to depend not only on the increased depreciation cost of the powered roof supports and the decreased cost of supplies and materials but also on such factors as the fixed and variable indirect costs and the selling price of the coal. Shortwall mining was demonstrated to be a viable alternative to room-and-pillar mining.

Jan 1, 1977

By I. AITKENS

Turquoise is prized solely for its attractive color ; it is dull and opaque , wholly lacking the brilliant luster that is the chief attraction of transparent gems . The best quality of turquoise is blue , of a soft and pleasing tone , like the color of clear sky but distinctive , as indicated by use of the term " turquoise blue . " Often its value is lessened by a greenish cast ; in many stones the green predominates . A large proportion of the turquoise that is mined is utterly worthless , so inferior and unattractive is its color . The finest grade of turquoise comes from Persia , notable from Nishapur . Large quantities have been mined in the Southwestern States of the United States , and less important deposits have been found in a number of the other States .

Sep 1, 1931

By John W. Walkiewicz

The U.S. Bureau of Mines has conducted studies to utilize rapid microwave heating to stress fracture ore samples. Iron ores containing hematite, magnetite, and goethite were subjected to microwave energy in batch operations at 3 kW and heated to average maximum temperatures between 840 and 940 °C. Standard Bond grindability tests showed that microwave heating reduced the work index of iron ores by 10 to 24 pct. In a microwave chamber designed to simulate a continuous throughput operation at 3 kW, the grindability of a taconite ore was improved by 13 pct at a bulk temperature of 197 °C. Because stress cracking occurred at a lower temperature, less energy was consumed. To further improve the economics of microwave fracturing, higher powers up to 16 kW were used to rapidly heat samples to relatively low temperatures in a continuous, belt-fed applicator. A significant improvement of grindability was obtained with a larger rod mill feed size in comparison to a minus 6-mesh Bond feed.

Jan 1, 2010

By Robert C. Dyni

This report details the investigation of multiple-seam longwall subsidence conducted by the U.S. Bureau of Mines from 1978 to 1989. A field investigation monitored ground surface movements over four upper-seam longwall panels and six lower-seam panels. The characteristics of the subsidence occurring as a result of mining these panels are examined; in particular, the angle of draw, subsidence development, total magnitude and areal extent, and critical width are evaluated and discussed. Comparisons are also made between the characteristics of multiple-seam longwall subsidence and single-seam longwall subsidence that occurred at the same site.

Jan 1, 1991

By G. A. Apell

The presence of considerable mineralization at the surface together with other favorable conditions suggested inclusion of the Nigger Jim diggings area in a comprehensive study of shallow lead deposits in southwestern Wisconsin. The area, which was the site of considerable shallow development some 60 to 70 years ago, offered the possibility of substantial quantities of lead at near-surface horizons. The investigation included completion of 30 churn-drill holes totaling 1,808 feet of drilling, 2 diamond-drill holes totaling 164.5 feet of drilling, and 25 feet of test pitting in two excavations. The degree of mineralization was determined in an area l,500 feet long and 150 to 300 feet wide along a mineralized trend that has not been investigated beyond the limits of the zone drilled by the Bureau of Mines.

Jan 1, 1948

By N. S. Dalal, V. Castranova, K. Van Dyke, V. Vallyathan, J. H. Kang

"We have reported previously that crushing or grinding crystalline silica results in the generation of silica-based radicals on the particulate surface and that these radicals can generate hydroxyl radicals in aqueous solution. Data in the present study indicate that freshly ground silica is more cytotoxic and is a more potent activator of alveolar macrophages than comparably sized aged silica. That is, compared to aged silica, fresh silica is 4.2-fold more potent in decreasing the membrane integrity of macro· phages; is 50% more potent in activating hydrogen peroxide secretion by macro· phages; and is 4.frfold more potent in stimulating cellular chemiluminescence. Prosil 28, an organosilane material, is an effective coating agent for fresh silica. It decreases the cytotoxicity of fresh silica by as much as 78% and decreases the ability of fresh silica to induce chemiluminescence from alveolar macrophages by 58%. The data suggest that surface radicals associated with freshly cleaved dust may be an important factor in the induction of pulmonary disease. Furthermore, treating dust with coating agents may substantially decrease toxicity. INTRODUCTIONThe U.S. Department of Labor estimates that over 1.25 million American workers may be exposed to crystalline silica and that nearly 60,000 may be at risk of developing some degree of occupational lung disease (Parkes, 1982; Peters, 1986). Occupations of concern include miners, stone quarry workers, sand and gravel workers, sandblasters, foundry workers, and glass workers. Acute silicosis is characterized by inflammation, alveolar lipoproteinosis, and the rapid development of respiratory disability (Banks, 1986). In contrast, chronic silicosis develops 20-40 yr after initial exposure and is characterized by the appearance of concentric hyalinized nodular lesions and the progressive development of dyspnea (Parkes, 1982; Ziskind et al., 1976)."

Mar 1, 1992

By W. A. Stickney, J. W. Town

Batch and continuous - circuit flotation , leaching , precipitation , and electrodeposition studies were made by the Bureau of Mines to determine optimum conditions for treatment of complex mercury - antimony sulfide and mercury- arsenic sulfide ores and concentrates . On the Red Devil ore sample over 95 percent of both the mercury and antimony sulfides were recovered in a flotation concentrate containing 35.6 percent mercury and 21.0 percent antimony . By selective flotation a mercury concentrate containing 71.0 percent mercury and 2.9 percent antimony was obtained with a mercury recovery of 90.4 percent . Leaching results show that , with 150 grams per liter of sodium sulfide and 50 grams per liter of sodium hydroxide , 95 to 99 percent of the mercury , antimony , and arsenic sulfides could be extracted from the concentrates treated . Sodium hydroxide reagent consumption was based on the sulfides generated in the dissolution of the sulfide minerals and on the method of metal recovery used . Selective leaching of the antimony and arsenic sulfides with potassium sulfide dissolved more than 95 percent of either of these two sulfides and 0.5 to 5 percent of the mercury sulfide present . Methods for recovering mercury and antimony from solution were electrodeposition of mercury followed by electrodeposition of antimony , aluminum precipitation of both mercury and antimony , or antimony precipitation of mercury followed by antimony electrodeposition . The method used in recovering the mercury depended on the mercury- antimony content of the solution being treated .Batch and continuous - circuit flotation , leaching , precipitation , and electrodeposition studies were made by the Bureau of Mines to determine optimum conditions for treatment of complex mercury - antimony sulfide and mercury- arsenic sulfide ores and concentrates . On the Red Devil ore sample over 95 percent of both the mercury and antimony sulfides were recovered in a flotation concentrate containing 35.6 percent mercury and 21.0 percent antimony . By selective flotation a mercury concentrate containing 71.0 percent mercury and 2.9 percent antimony was obtained with a mercury recovery of 90.4 percent . Leaching results show that , with 150 grams per liter of sodium sulfide and 50 grams per liter of sodium hydroxide , 95 to 99 percent of the mercury , antimony , and arsenic sulfides could be extracted from the concentrates treated . Sodium hydroxide reagent consumption was based on the sulfides generated in the dissolution of the sulfide minerals and on the method of metal recovery used . Selective leaching of the antimony and arsenic sulfides with potassium sulfide dissolved more than 95 percent of either of these two sulfides and 0.5 to 5 percent of the mercury sulfide present . Methods for recovering mercury and antimony from solution were electrodeposition of mercury followed by electrodeposition of antimony , aluminum precipitation of both mercury and antimony , or antimony precipitation of mercury followed by antimony electrodeposition . The method used in recovering the mercury depended on the mercury- antimony content of the solution being treated .

Jan 1, 1964