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By D. Harrington

The mining industry of the United States has for many years possessed the dubious aistinction of having the poorest accident rate of all of the major lines of industrial endeavor in this country, and until recently there ras little or nothing to show that any material improvement might be expected. However, there are now numerous trends indicating that the corner has ceen turned at last, and it is strongly hoped that the mining industry is likely to follow the railroads, the cement and other industries in operating with occurrence of only a reasonable amount of loss of life and limb through accidents. The complete accident figures for metal and nonmetallic mineral mining for 1931 are not yet available, but fragmentary data shor that 1831 vill un- questionably have very low accident rates, probably the lowest in the history of our metal mines. One underground metal mine has produced over 14,000,000 tons of copper ore rithout a fatality; another metal mining company with an electric slovel operation has handled upwards of 75,000,000 tons of material, copper ore and waste, without the occurrence of a fatality; an iron-ore compan, which orns six mines (three underground, three open pit), has operated every one for a year or more without a lost-time accident with a con- siderable number of employees and a large production; and alot.er iron-ore producer with 30 mines and about 3,000 employees as an average of nearly 10 months of operation in 1931 without a lost-time accident. umerous other wonderful safety records of metal mines in 1931 indicate unmistaimbly tat the record of the noncoal mining industry of the United States as a role in 1931 will undoubtedly be found to be excellent. According to the latest relatively complete figures (those of 1930) available for the coal industry, fewer than 650,000 persons are employed in and around the coal mines of the United States, the figures indicating at a little over 493,000 people are engaged in bituminous mining and a little under 151,CCO in anthracite mining. Bureau of Mines complete figures suc that 2,053 persons were killed and 103,821 non-fatally injured in our coal

Dec 1, 1932

By D. G. Foot

The Bureau of Mines investigated energy-efficient methods for recovering potash values from process and waste brines. Laboratory pan evaporation of four chloride brines produced crude salts containing predominantly sylvite, halite, and carnallite. Six sulfate-chloride brines produced crude salts containing primarily schoenite, kainite, leonite, sylvite, carnallite, and halite. Potash grades ranged from 7.2 to 22.2 pct K20, and recoveries from 84 to 99 pct. Sylvite flotation from chloride evaporites, with amine collector, recovered 90 to 97 pct of the potash in a concentrate containing 54.3 to 60.3 pct K20. Fatty acid flotation of the high-sulfate evaporite recovered 78 pct of the sulfate minerals in a 27.8-pct-K20 concentrate. Flotation of the chloride minerals with amine collector recovered 80 pct of the potash in a 59.7-pct-K20 concentrate. Solar evaporation of 10,000 gal of brine recovered 99 pct of the potash in a crude evaporite containing 24.5 pct schoenite and 20 pct sylvite. Continuous flotation in a 100-lb/h process research unit recovered over 95 pct of the potash in schoenite and sylvite concentrates containing 28.0 and 62.3 pct K20, respectively. An economic evaluation suggested a rate of return of 3 pct for a new facility and 9 pct if. the process is adapted to fit an existing plant.

Jan 1, 1984

By R. C. Buehl

A pyrometallurgical process for recovering manganese from open-hearth furnace slags or low-grade manganiferous iron ores is described. The slag or ore is smelted in a blast furnace to produce spiegeleisen; the speigeleisen is then blown with air in a basic converter to give an oxide slag. This slag product contains 60 to 75 percent manganous oxide and may be considered a synthetic manganese ore. In the smelting operations coke rates for five test periods ranged from 3,255 to 4,130 lb per ton of spiegeleisen. The recovery of manganese in the spiegeleisen varied from 60 to 76 percent. Recovery increased to 83 percent when the open-hearth slag feed to the blast furnace was diluted with a high-? grade iron ore sinter to facilitate smelting. In blowing the spiegeleisen a cyclic procedure was developed that gave nearly 90 percent recovery of manganese in the slag product (synthetic ore) yet prevented excessive contamination of the slag by phosphorus and iron. The overall recovery of manganese was about 75 percent when iron ore sinter was used with the open-hearth slag in the smelting step. The metal remaining after oxidizing the manganese from the spiegeleisen contained approximately 3.5 percent phosphorus. It was refined in a basic converter by blowing with air in the presence of lime" A high--grade iron and a basic phosphate slag were produced.

Jan 1, 1965

By E. J. Lintner

This paper is one of a series being prepared by the United States Bureau of Mines describing cloy reining, crushing, and grinding methods and costs at various operations throughout the United States. These Toners are designed to disseminate technical information regarding methods used. The cost tabulations represent local operating expenditures only and not total production casts. Although the publication of total production cants might not be advisable in sore instances, yet a knowledge of operating costs is essential to the technical discussion end study of the methods employed. The attention of the reader is specifically coiled to this differentiation in order that no misunderstanding of the scope of the cost tabulations shall ensue. The following report deals specifically with the methods employed by the Clay City Fire Co., Uhrichsville, Tuscarawas County, Ohio.

Jan 1, 1936

By Jon K. Ahlness

The Bureau of Mines conducted laboratory and field experiments to determine tne amount of permeability reduction in uranium sandstone after its exposure to different drilling fluids. Seven polymer and two bentonite fluids were laboratory-tested in their clean condition, and six polymer fluids were tested with simulated drill cuttings added. Sandstone cores cut from samples collected at an open pit uranium mine were the test medium. The clean fluid that resulted in the least permeability reduction was an hydroxyethyl cellulose polymer fluid. The greatest permeability reduction of the clean polymers came from a shale-inhibiting synthetic polymer. Six polymer fluids were tested with simu-lated drill cuttings added to represent field use. The least permeability reduction was obtained from a multipolymer blend fluid. A field experiment was performed to compare how two polymer fluids affect formation permeability when used for drilling in situ uranium leaching wells. For this test, the polymer fluid with the best laboratory results (multipolymer blend) was compared with a commonly used polymer fluid (guar gum) that gave poorer laboratory results. When fluid injection rates for the four wells drilled with the guar gum were compared with those for the four drilled with the multipolymer blend, no statistically significant difference was found.

Jan 1, 1984

This Bureau of Mines report is a compilation of information on computer models of mineral commodity markets in the Federal Government, The directory was compiled by the Analytic System Working Group of the Interagency Minerals Information Coordinating Committee to support the formulation of policies related to nonfuel minerals.

Jan 1, 1984

By Peter Grandone

"INTRODUCTION The Bureau of Mines, recognizing the importance of knowing the evapor¬ation losses of aviation gasoline as larger quantities of this material must be stored, has made studies at a large refinery in the Texas Panhandle area during the past 13 months to determine the magnitude of the losses, by evaporation, under different storage conditions. Proper storage of aviation gasoline is very important, not only because large supplies are being stored for military use but because the specifications to which his fuel must conform are very rigid. Aviation gasolines that conform present-day specifications are blended from base-gasoline stocks (in some instances relatively pure hydrocarbons), each of which has different physical and chemical characteristics. Consequently, partial loss of any one of the more volatile components from the blended fuel results not only in material loss of the blended fuel but also may affect the characteristics or performance of the gasoline. One characteristic that has considerable influence or the vaporization tendency of gasoline is vapor pressure; and although the vapor pressure of aviation gasoline usually is lower than that of average commercial motor fuel, the quantity of certain components having relatively higher vapor pressure is greater in aviation gasoline, so that under prolonged storage the evaporation loss of these components may be appreciable."

May 1, 1943

By Forrest T. Moyer

THIS CHAPTER of the Minerals Yearbook (Volume III) contains overall injury experience for coal mines, both anthracite and bituminous coal, coke plants, petroleum and natural gas, peat, and asphalt and related bitumens (native); metal mines, metallurgical plants, including ore-dressing plants and primary nonferrous reduction plants and refineries; nonmetal mines; nonmetal mills; sand and gravel operations; slag (iron blast-furnace) plants; and stone quarries and their related plants. Volume I of the yearbook contains a chapter showing injury experience and employment data treated separately by categories: Metal mines, nonmetal mines, quarries and their related plants, metallurgical plants (ore-dressing plants and primary nonferrous reduction plants and refineries are shown separately), nonmetal mills, sand and gravel plants, and iron blast-furnace slag plants. Volume II contains injury and employment experience in the fuel industries (anthracite, bituminous coal including lignite, coke, petroleum and natural gas, native asphalt and related bitumens, and peat). Injury and employment data were collected from coal producers on the regular mandatory basis as required by the Federal Coal Mine Safety Act (30 U.S.C., sec. 455). Similar data for 1962 from metal, nonmetal (except well and brine operations), stone, and sand and gravel producers were collected by a. mandatory reporting system under the provisions of the Act of September 26, 1961, Public Law 87-300 (75 Stat. 649). Before 1962, these nonfuel producers reported voluntarily to the Bureau of Mines. Producers of all other minerals (fuel and nonfuel) reported the requested injury and employment data voluntarily to the Bureau for 1962, the same as in preceding years.

Jan 1, 1963

By R. F. Tenney

This report has been prepared because little detailed information has been published on the production and use of anthracite producer gas for industrial-process and brick-kiln heating. To obtain this information, an investigation was conducted by the Federal Bureau of Mines in cooperation with a building-brick manufacturer (firing both periodic and tunnel kilns with anthracite producer gas) and with a manufacturer of anthracite-gas producers. In addition to a study of available plant records, 2 weeks was spent in systematic observation and measurement during regular operation of the Hazleton Brick Co. plant, Hazleton, Pa., in July 1953. This report presents the results of this investigation and a description of producer and periodic-kiln equipment, operating procedures, and labor requirements. A summary of the conditions of producer and kiln operation is given for the 2-week observation and for the 11 preceding months. The two 10-foot standard anthracite-gas producers consumed 301 tons of Rice and Buckwheat No. 1 anthracite, as delivered, averaging over 13,000 B.t.u. per dry pound during the 2-week period, with a producer hot-gas efficiency of 95 percent, at an average gas offtake temperature of 5400 F. The yield of gas5/ of 153 gross B.t.u. per cubic foot amounted to 140,000 cubic feet per ton of fuel as fired, or 73,500 cubic feet per kiln heating hour, which was sufficient for kiln requirements.

Jan 1, 1960

By James J. Dowd

The investigation to evaluate the reserves of coking coal is being made by the Bureau of Mines in three parts: (1) To estimate known (measured plus indicated) recoverable reserves of all coking coal, (2) to upgrade marginal coals through effective preparation, and (3) to study the carbonizing properties of coal and coal blends not now widely used for metallurgical coke making. This is the seventh of a series of reports giving the results of studies by counties under part (1) of the investigation. (See appendix.) This report covers Floyd County, KY., which comprises parts of the Paintsville, Prestonsburg, Harold, Hindman, and Pikeville quadrangles. (See fig. 1.) A base map for each bed in each quadrangle was made to the scale of 1 inch equals 1,200 feet. Maps of mines, locations of drill holes, bed and total coal thicknesses, and the outcrop of the bed were plotted on the maps. With all avail-able data plotted, isopach lines were drawn to limit areas of known unmined reserves in beds up to 14 inches thick, 14 to 28 inches thick, 28 to 42 inches thick, and over 42 inches thick. These areas of coal reserves also were divided into "measured" and "indicated" categories. All areas in each thickness range and in each category, mined-out areas, areas excluded from the estimate but which may contain reserves based only on geologic inference, and areas outside of the outcrop were measured by planimeter on the base maps. Estimates of total reserves 14 inches and more thick and maps for individual beds were prepared from these data.

Jan 1, 1951

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

Studies on continuous - circuit leach - aluminum precipitation and leachelectrodeposition of mercury were made by the Bureau of Mines to determine optimum conditions and cost estimates for recovering mercury from mercury sulfide flotation concentrates . With 150 grams of sodium sulfide and 50 grams of sodium hydroxide per liter , from 98 to 99.8 percent of the mercury was extracted from flotation concentrates , with about 0.5 pound of sodium hydroxide used per pound of mercury produced . Precipitation of mercury from the caustic - sulfide solution with minus 8 - mesh aluminum shot required about 0.2 pound of aluminum per pound of mercury produced at a cost of $ 0.05 per pound of mercury . Electrodeposition cost with 7 mill power , based on 50 percent current efficiency , was below 2 mills per pound of metal deposited . Feed solutions ranged from 10 to 130 grams of mercury per liter , and current densities , from 2 to 70 amperes per square foot . The cost of open- pit mining 250 tons per day was estimated at $ 1.75 per ton , and flotation - leachingelectrodeposition of the ore , at $ 3.54 , resulting in a cutoff grade of about 0.1 percent mercury . The cost of underground mining 50 tons per day was estimated at $ 20.00 per ton , and the cost of milling the ore was estimated at $ 13.33 , which resulted in a cutoff grade of 0.6 percent mercury .

Jan 1, 1964

By F. P. Haver

Because State and Federal agencies have limited the amount of sulfur oxides that can be discharged to the atmosphere, most copper smelters plan to convert the S02 in offgases to sulfuric acid. In some cases, however, it might be preferable to use a nonsmelting method to fix the sulfur in the form of a discardable product. Research by the Federal Bureau of Mines has shown that sulfur in copper concentrate can be converted into insoluble anhydrite (CaS04) by roasting the concentrate with lime (CaO). After leaching the calcine to recover metal values, the residue may be discarded without harm to the environment. The process can use a rougher flotation product, thus reducing separation costs and saving valuable minerals (for example, molybdenite) that are now lost in making cleaner concentrate. In tests, rougher concentrate was pelletized with the theoretical amount of lime and roasted for 1 hour in a 6-1/2-inch-ID by 11-foot-long rotary kiln at 700° C. Sulfur retention was about 97 percent. Leaching the calcine with 20 percent RCI extracted approximately 99 percent of the copper, 73 percent of the iron, and 91 percent of the molybdenum. After the molybdenum was removed from the leach solution by adsorption on activated carbon, copper with a purity in excess of 99 percent was obtained by cementation on sponge iron. The strip solution then could be spray-roasted at 500° C to recover the RCI for recycle and obtain F2O3 as a byproduct.

Jan 1, 1975

By F. P. Haver

A method to recover copper and elemental sulfur from chalcopyrite concen¬trate, using a ferric chloride leach, CuFeS2 + 3FeCl3 ? CuCl + 4FeCl2 + 2S, was reported by the Federal Bureau of Mines in 1971 (Report of Investigations 7474). Since that time, additional work has been done to improve certain steps in the procedure; that is, the recovery of sulfur from the leach residue, the recovery of copper and iron from the leach solution, and the regeneration of the leach solution. It was determined that over 99 pct of the elemental sulfur in the leach residue could be extracted at ambient temperature with an ammonium sulfide solution. Heating the pregnant solution precipitated crystals of rhombic sulfur (99.5 pct purity) and liberated ammonia and hydrogen sulfide for recycle. Copper with a purity in excess of 99 pct was obtained from the leach solution by electrolysis, using a diaphragm cell, with an energy requirement of less than 0.7 kW-hr/lb. Metal was deposited as a weakly adherent powder that could be easily dislodged from the cathode and withdrawn continuously from the cell in the form of a slurry. During electrolysis, part of the Fe(II) was oxidized to Fe (III) at the anode so there was no chlorine evolution. The leach solution was regenerated by direct oxidation of the spent electrolyte with air or oxygen in a turboaerator at 80° C. This precipitated the iron dissolved during leaching in an easily filterable form, Fe203.xH20, and converted the remaining Fe(II) to Fe(III).

Jan 1, 1975

By Seth C. Schaefer

The Federal Bureau of Mines studied the Gibbs free energies of formation of chromous fluoride (CrF2), chromic fluoride (CrF3), and chromium (II,III) fluorides (CrF2.40 and CrF2.45) with high-temperature galvanic cells employing calcium fluoride doped with 1 mole-pct yttrium fluoride as the electrolyte. Free-energy changes were obtained from open-circuit potential measurements for the following cells: Cr, CrF2 // CaF2 // NiF2, Ni (a) NiO, NiF2 // CaF2 // CrF3 , Cr203 (b) CrF2 , CrF2.40 // CaF2 // NiF2, Ni (c) and CrF2 ?45 , CrF3 // CaF2 // NiF2, Ni (d) with the following overall reactions Cr(s) + NiF2 (s) = Ni(s) + CrF2 (s) (1) 3NiO(s) + 2CrF3 (s) = 3NiF2 (s) + Cr203 (s) (2) 5CrF2 (s) + NiF2 (s) = 5CrF2.4O (s) + Ni(s) (3) and 3.64CrF2.45 + NiF2 (s) = 3.64CrF3 (s) + Ni (s) . (4) These results were combined with the best available literature values of ?Gf° of NiF2, NiO, and Cr203 to deduce Gibbs free energies of formation for the chromium fluorides expressed as follows: ?Gf°(CrF2) = -179.31 + 24.33 X 10-3T ± 0.43 kcal-mole. (5) (858.1 - 964.9 K) ?Gf° (CrF3) _ -226.16 + 1.82 X 10-3T ± 0.79 kcal/mole (6) (880.8 - 957.8 K) ?Gf°(CrF2.40) _ -213.65 + 32.14 X 10-3T ± 0.44 kcal/mole (7) (0053.9 - 936.8 K) and ?Gf°(CrF2.45) _ -176.17 - 12.67 X 10-3T ± 0.80 kcal/mole (8) (875.1 - 937.6 K). Free energies and equilibrium fluorine pressures for three univariant and one bivariant equilibria are illustrated.

Jan 1, 1976

By D. E. Wolfson

As part of its overall program of coal-carbonization research, the Bureau of Mines is studying the coking properties of coals from various fields of this country. Few of the coals included in this survey are used for the manufacture of metallurgical coke, although some well-known coking coals necessarily are tested to provide comparative data, particularly on blending properties. The overall objectives of the study were given in detail in a previous report.3/ The inherent suitability of coal for metallurgical purposes, as distinguished from economic considerations, depends primarily on (1) the available reserves, (2) chemical composition, (3) preparation characteristics, and (4) carbonization behavior. These items are considered in selecting coal for carbonization tests.

Jan 1, 1956

By Danton L. Paulson

Elevated temperature electrical resistivity measurements were made by the Bureau of Mines on arc-melted, spin-cast, and hot-pressed columbium carbide¬carbon (CbC-C) and zirconium carbide-carbon (ZrC-C) materials up to 2,000° C. The CbC-C materials displayed electrical resistivity values ranging from 35.20 microhm cm at 23° C for hot-pressed material containing 12 wt pct carbon to 242.64 microhm cm at 1,622° C for arc-cast material containing 17.0 wt pct carbon. Electrical resistivity values ranged from 29 microhm cm at 20° C for arc-cast ZrC-C containing 16 wt pct carbon to 279 microhm cm at 1,951° C for hot-pressed ZrC-C containing 21 wt pct carbon. Data from a previous report were included to establish comparison between the electrical resistivities of the CbC-C, ZrC-C, and HfC-C systems.

Jan 1, 1969

By J. B. Zink

The Bureau of Mines developed a spectrochemical method to determine six constituents in coal ash in the following concentration ranges: silica 9 to 63 percent, alumina 7.5 to 35 percent, ferric oxide 2 to 30 percent, calcium oxide 1 to 29 percent, magnesium oxide 0.5 to 6 percent, and titanium oxide 0.5 to 4 percent. Samples of coal ash are fused with a mixture of lithium tetraborate and vanadium pentoxide. The resulting beads are crushed, mixed with powdered graphite, and briquetted. The pellets are analyzed spectrographically, using controlled high-voltage spark excitation. Vanadium serves as the internal standard. Effects due to matrix variations are minimized by fusion with lithium tetraborate. Intensity ratios of selected analytical lines are determined photometrically, Concentrations are read from analytical curves, prepared from synthetic standards, relating log intensity ratios to log concentration. A combustion procedure is used to determine sulfur tri-oxide. The method is most useful when large numbers of coal ashes must be analyzed but a high degree of accuracy is not essential.

Jan 1, 1967

By Charles Prasky

The Bureau of Mines investigated the conversion of nonmagnetic iron minerals to magnetic form through reduction Toasting operations. The primary objective was to obtain a concentrate that could be used as a direct shipping material for blast furnace consumption. Crude iron materials from the Steep Rock Lakes, Ontario, area were prepared into a blend of 8 parts siliceous ore, 3 parts pyritic ore, and 1 part tailings. Experimental work included laboratory reduction roasting tests, two continuous 10-day campaigns, and one 12-hour test all in a 34-inch by 36¬-foot rotary kiln. The objective of the project was achieved in producing an iron concentrate containing about 66 percent iron with over 90 percent iron recovery. Desirable operating conditions were developed in the kiln but optimum process variables were not completely established.

Jan 1, 1969

By Jack E. Fraley

Because of their noncombustibility, silicate and magnesium oxycements are desirable for sealing shale in underground coal mines. In this Bureau of Mines investigation, cement samples were cured and tested (nonstandard tests) for submersion to note resistance to dissolution (breaking into parts), which is an important property for a coating. Nonappreciable dissolution during 1 week of submersion was set for passing the test. Over half of the coatings passed the test.

Jan 1, 1977

By Agnes Y. Lee

The Bureau of Mines has developed an effective hydrometallurgical method to recover high-purity Pb metal and elemental S from concentrates. This low-temperature process eliminates S gases and Pb emissions, in contrast to the current high-temperature smelting technology. The method consists of (1) oxidative leaching with H202, Pb02, and waste fluosilicic acid at 95° C to produce a solution of PbSiF6 and a residue containing elemental S, (2) electrowinning the PbSiF6 solution at 35° C to produce 99.99 pct Pb metal and H2SiF6, and (3) solvent extraction to recover S, leaving a residue containing Cu, Ag, and other metal values. Spent electrolyte was recycled repeatedly, with impurity buildup controlled by controlling the leach parameters.

Jan 1, 1986