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By C. E. Jordan

The fundamental properties of dielectric separation were studied to provide insight into the design and operating characteristics of a dielectric separator. Force equations on mineral particles were derived and evaluated experimentally. Responses to variations in the electrical voltage and frequency, feed rate, particle size, dielectric fluid, and dielectric constant were tested theoretically and experimentally, using a 5-cm-diam dielectric separator. High-dielectric mineral recovery was proportional to the calculated effective dielectric force volume (EDFV) that surrounded each drum-electrode wire. Electrode spacing was the most sensitive design parameter; spacing of less than 0.4 cm was required for a good separation. Dielectric separation was effective for separating minerals with dielectric constants that differed by approximately 2 or more.

Jan 1, 1985

By L. W. Kelly

The primary purpose of this paper is to show the appreciable savings that were made by roof-bolting the main west entry of the Kings mine, which was to be the permanent, main haulage road, as compared to the cost of supporting the same roof with peg timbering as first planned. Additional items that might prove valuable to other mine operators are contained in this paper and detailed below. 1. A detailed cost of the system of roof support known locally as the peg-timbering method is given. It consists fundamentally of supporting steel cross bars on pegs inserted into holes drilled in the ribs close to the roof. 2. A method of supporting the roof in caved entries is shown, where such support must endure for an extended period in haulageways and air courses.

Jan 1, 1952

This Bureau of Mines publication describes the plan used to instrument two longwall gateroads at the Powhatan No. 4 Mine of the Quarto Mining Co. Vibrating wire stressmeters, flat jack U-cells, multipoint extensometer stations, and convergence stations were installed in or on the coal pillars, roof bolts, and roof strata of the gateroad entries. The instruments measure pillar stresses, bolt loadings, and strata movements as the gateroad entries are developed and the longwall panel(s) is mined. The report details only the instrumentation plan of the longwall gateroads and not the evaluation of any subsequent data. This is because a complete analy-sis of the data cannot be made until extraction of the longwall panel(s) is complete.

Jan 1, 1982

By L. C. IlsLey

"The maintenance of permissible equipment which is tested and inspected at the Pittsburgh Experiment Station of the United States Bureau of Mines is dependent to a surprisng degree upon the voltage at which the equipment is operated. The difficulty of maintaining 500-volt equipment is frequently brought to the bureau's attention, and it is quite possible that if the bureau had appreciated at the beginning the full hazard involved, it would never have undertaken the approval of equipment operating at voltages greater than 250 volts.In this connection, the following statement from State mine inspection departments and other sources are of interest as showing the general feeling in safety organizations with respect to voltages higher than 250. These statements, although obtained as a result of a questionnaire sent cut several years ago, still apply very well to present conditions.A. I have not had time to consider seriously the proposed change from every angle, but looking at it from a standpoint of hazard, feel that the danger would be greatly increased. I have always occupied the position that underground voltage on portable equipment should never exceed 250 volts.High voltage of course would mean smaller transmission lines, which would be more economical and no doubt would require in several instances smaller equipment; notwithstanding, if this statement is true, the writer is still favorable to low voltage on underground portable equipment.B. I believe 500 volts should be the limit on the best equipment ender the best conditions and that the U. S. Bureau of Mines should do all in its power to limit the voltage to 500 volts on permissible equipment, as any increase in voltage means increased hazard.C. I will refer you to a section of our mining law, entitled ""An Act to Regulate the Tine of Electricity in the Mines of the State of ----"" which reads in part, as follows: Trolley wires, or other exposed electrical wires, shall not carry a voltage above 275 volts.You may readily see by this that we are not permitted to operate mining machines under ground in -----, with more than 275 volts at the terminals, except that we are permitted to install transforming stations throughout the mines, after complying with certain regulations at the mine.However, as we have had to date, within the calendar year of 1925, six men electrocuted in the mines of the State, using a voltage of 275 volts, I am of the opinion that if this voltage was increased, the hazard would also be increased in proportion.D. I do not care to make a statement for publication: however, I am referring you to our mining laws. On page 27 and 28 you will find what the laws of -----¬require relative to the installation of electricity in the mines.""Hereafter where electricity is installed no higher than two hundred seventy-five (275) volts shall be used underground, except for transmission, or for application to transformers or other apparata, where the high-voltage circuit is stationary. High-voltage motors, and transformers, shall be installed in suitable chambers, built of fire-proof construction, and well ventilated, and shall not be placed on intake, or between intake and return air courses. All high-voltage lines lines and apparata must be clearly marked, to indicate their danger. Main and distribution switchboards for high voltage shall be made of insulating and incombustible material, such as marble or slate, and to be free of all metallic veins, or of material equivalent thereto. Same shall be installed in a dry place. All switches or other instruments used in connection with high voltage shall be installed on suitable switchboards, with at least a three-foot passageway in front and rear properly floored, and kept looked by a door at each end that can be opened from the inside without a key. With higher than medium voltage, there shall not be any live metal work on the front of switchboard, and the entry circuit must be protected by an oil breaker s"

Apr 1, 1932

By James J. Dowd

The investigation to evaluate the reserves of coking coal in the United States suitable for the manufacture of metallurgical coke 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 coals and coal blends not now widely used for metallurgical coke making. This is the fifth of a series of reports giving the results of studies by countless under part (1) of the investigation. (See appendix.) This report covers Westmoreland County, Pa., which comprises parts of New Kensington, Freeport, Elders Ridge, Pittsburgh, Greensburg, Latrobe, New Florence, Johnstown, Brownsville, Connellsville, Donegal, and Somerset quadrangles. (See fig. 1.) A short resume" of the chemical and carbonizing properties of the coals are included. Data on the coal beds in this county were obtained from mining companies, land companies, individual landowners, and other sources of authentic information. A base map for each bed in each quadrangle was made to the scale 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 available 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, 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 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 and 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 tenth of a series of reports giving the results of studies by counties under part (1) of the investigation. (See appendix.) This report covers Knott County, Ky., which comprises parts of the Salyersville, Prestonsburg, Troublesome, Hindman, Pikeville, Cornettsville, and Whitesburg quadrangles. (See fig. 1.) A base map for each bed in each quadrangle was made to the scale 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 available 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 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, 1952

Consumption and intensity of use trends for 12 metals, by industrial end use, were estimated for 1972 through 1982 by the Bureau of Mines and U.S. Department of Commerce. The trends were then forecast through 1993 using standard statistical methods. The intensity of use measure selected is the quantity of metal consumed per constant dollar output of a specific industrial sector. The metals studied were aluminum, chromium, cobalt, copper, lead, manganese, nickel, the platinum-group metals, tin, titanium, tungsten, and zinc. Average annual growth rates for the following 4 of the 12 metals increased from 1972 to 1982 as follows: aluminum, 1.00%; platinum-group metals (consisting of platinum, palladium, and iridium), 8.83%; titanium, 7.82%; and tungsten, 6.09%. Consumption for the remaining eight metals decreased at average annual rates ranging from 0.58% (copper) to 5.37% (manganese).

Jan 1, 1986

By R. M. Santmyers

Except for optical purposes and a few other special uses quartz and quartzite are to a considerable degree interchangeable. This also may be said with respect to pulverized quartz and ground silica sand, which are discussed in a separate paper of this series. All these materials, except perhaps quartz crystals of optical grade, are abundantly available in many parts of the United States. In the form of ordinary crushed stone much of it is sold at $1 a ton or less, and ordinarily

Aug 1, 1931

By Steven A. Fechner

In 1984 the Bureau of Mines cunducted mineral investigations of specific sites within the Yentna Mining District. One hundred eight placer and rock samples were taken. Reconnaissance placer samples were collected from drainages in the Talkeetna, Susitna, and Yentna River basins. The highest placer gold values were found in the historic placer gold producing creeks of the Cache-Peters Creek area. Samples taken from an outcrop in the Middle Fork of Iron Creek in the Talkeetna Mountains contained previously unreported high lead, zinc, and silver values.

Jan 1, 2011

By H. F. MILLE, Charles A. Kumke

This paper describes mining operations at the Iron King mine of the Shattuck -Denn Mining Corp. in Yavapai County , Ariz . It is one of a series being prepared by the Mining Division of the Bureau of Mines on mining practices and methods in various mining districts of the United States . The mine and surface plant of the Iron King mine are situated in secs . 15 and 21 , T. 13 N. , R. 1 E. , about 20 miles by State highway 69 east by southeast from the city of Prescott and 1 mile southwest of the mining town of Humboldt on the Prescott -Mayer branch of the Atchison , Topeka & Santa Fe Railroad (fig . 1 ) . A l-mile truck haul brings the mill concentrates (lead and zinc ) to bins on a siding of this line at Humboldt ; the silver -gold bullion is handled separately . Some of the oldest residents in the vicinity of Prescott claim that the original location for the Iron King mine was made in 1880. The earliest recorded production was in 1906-1907 . During the first World War the mine produced fluxing ore for the Humboldt smelter . In 1936 the property was purchased under tax sale , and a mill (140-ton capacity ) was erected in 1938. The capacity of this mill was increased to 225 tons the following year , and a cyanide plant was added in 1940 to treat mill tailings . In 1942 the Shattuck -Denn Mining Corp. purchased the property . A new working shaft was sunk , mine development was speeded , and mill capacity was again increased . The predominating locally formation of this district is a pre -Cambrian schist known as Yavapai schist . The Iron King deposit , which carries gold, silver , lead , zinc , and copper , occurs as a series of lenses in a sheared zone of the pre -Cambrian schist . Down to the 600 -foot level the ore was mined and drawn by the shrinkage method . The cut -and -fill method was adopted for the lower levels . Waste rock for stope fill was provided by block-caving an area on the 600 level . Main haulage drifts were driven in the firm footwall , and crosscuts were driven across the mineralized zone at 100 -foot intervals . Grab samples taken from each car of every working place and channel samples cut every 5 feet in development work are considered sufficient for estimating ore reserves .

Jan 1, 1950

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

By J. R. Cobble

Dissolution of chalcopyrite with acidified ferric sulfate produces a sulfur layer on the unreacted chalcopyrite surface that interferes with further dissolution. The U.S. Bureau of Mines has developed a process to improve the leaching kinetics of these types of diffusion-controlled systems. The system features the Bureau-developed turbomill, an attrition grinder that scrubs the reaction product coating away from the unreacted mineral grains during leaching. The system can be operated continuously at any temperature up to the boiling point of the leaching agent. This method is a hydrometallurgical alternative to chalcopyrite smelter technology and eliminates sulfur dioxide emission problems with its associated costs. Batch testing at 900 C with the chalcopyrite-ferric sulfate system resulted in copper extraction exceeding 99 pct. At an optimal energy consumption of 1,520 kW· hjst, 95 pct of the copper was extracted. The pregnant liquor was clarified and tested for typical solvent-extraction recovery of the copper. Two-stage extraction and two-stage stripping recovered 96 pct of the copper in an electrolyte suitable for copper electrowinning.

Jan 1, 1993

By R. V. Ramani

In this report, the application of the Open Pit Materials Handling Simulator (OPMHS) to a mine located in Illinois using bucket wheel excavator (BWE) and a stripping shovel in tandem for overburden removal is described. The specific objective of the study is to demonstrate the application of the simulator to analyze mine planning and scheduling problems and to generate and analyze alternatives for increased production. On the basis of the simulations it was concluded that the maximum productive capacity of the system in raw coal was around 3000 tons per shift. At this production level, the equipment combination and availability also satisfy the match requirements for the stripping shovel and the BWE, and the overburden and coal removal systems. Several alternatives requiring new equipment are proposed and analyzed for increasing production. However, no economic analyses of these proposals were done.

Jan 1, 1977

By R. C. Hickman

"INTRODUCTION The largest known deposits of ilmenite and rutile in Virginia are in Amherst and Nelson Counties. Other deposits occur in the State, but their extent is not known. One of the most promising of the latter is the deposit in Roanoke County on the T. M. Bush and the H. L. Hutchins property near Roanoke. A geophysical survey was made of the deposit, followed by core drilling. In April and May 1946, four holes were drilled, totaling 822 linear feet.ACKNOWLEDGMENTSIn its program of exploration of mineral deposits, the Bureau of Mines has as its primary objective the more effective utilization of our mineral resources to the end that they make the greatest possible contribution to national security and economy. It is the policy of the Bureau to publish the facts developed by each exploratory project as soon as practicable after its conclusion. The Mining Branch, Lowell B. Moon, chief, conducts preliminary examinations, performs the actual exploratory work, and prepares the final report. The Metallurgical Branch, Oliver C. Ralston, chief, analyzes samples and performs beneficiation tests.The project was under the immediate supervision of Mitchell H. Kline, chief, Raleigh Division, Mining Branch,. Bureau of Mines."

Aug 1, 1947

By Michael S. Miller

In 1986 and 1987, at the request of the U.S. Bureau of Land Management, the U.S. Bureau of Mines studied the 550-acre Carson-Iceberg Wilderness Study Area (NV-030-532), in order to evaluate its identified mineral resources. The Wilderness Study Area is located in Alpine County, CA, 11 mi southeast of Markleeville, CA. Tertiary volcanic and granitic rocks, as well as alluvium, underlie the area. Hydrothermal alteration and faulting have affected some rocks; argillic, silicic, and limonitic alteration are found, particularly near faults. No mineral resources were identified in the Wilderness Study Area. The area has one small pit and no mines or claims. Samples from the Wilderness Study Area and vicinity contained uneconomic amounts of gold, silver, copper, lead, zinc, mercury, arsenic, and molybdenum. Mineral occurrences are too small and low grade to be of interest within the foreseeable future.

Jan 1, 1988

By John F. Papp

Based on a review of historical domestic and international chromium metal production capacity, production, consumption, trade, and trade literature, world annual chromium metal production capacity in 1995 was estimated to have been 38,300 metric tons, a drop from 43,500 tons in 1994. World chromium metal production and consumption were estimated, on average from 1984 through 1992, to have been 21,000 tons per year. U.S. consumption was estimated, on average over the same time period, to have been 6,200 tons per year. The major world chromium metal producers between 1984 and 1995 were China, France, Germany, Japan, Russia, United Kingdom, and United States. From 1984 through 1992, world trade averaged 8,400 tons per year. Germany and the United States accounted for over 75 percent of imports on average during the same time period. U.S. chromium metal import value rose from under $1 per kg before 1964 to over $6 per kg in 1993. Over the same time period, prices rose from about $4 per kg to about $10 per kg. Chemical specifications have become more stringent. Detailed chromium metal trade, prices, value, chromium metal chemical specification, and production capacity by country are reported. A shift in trade patterns was anticipated as Japan exits production and displaces Germany as the second largest importer and as Russia increases its role in world trade.

Jan 1, 1995

By B. K. Shibler, I. S. Solet, H. W. St. Clair

"This study was undertaken to provide a more adequate knowledge of the chemistry of the sulfide volatilization process for recovery of tin from low-grade Bolivian ores. The investigation was partly supported by the Department of State under its program for scientific cooperation with South American republics.The results of this investigation demonstrate the conditions under which cassiterite (Sn02) may be converted to stannous sulfide by reaction with pyrite or sulfur vapor. Quantitative data are given on the rate of sulfidization and on the vapor pressure of stannous sulfide.INTRODUCTIONAlthough the United States is the world's largest consumer of tin, virtually no tin ores occur within our national borders. Most of the world's tin supply comes from alluvial deposits in the Far East and in Africa, The major sources of tin ores are Malaya, Indochina, Bolivia, Belgian Congo, Nigeria, Thailand, Burma, China, and the United Kingdom. Bolivia is the only important source of tin in the western hemisphere. It was the chief source of tin f or the United States during World War II. The Government-owned tin smelter at Texas City, Tex., was built primarily for smelting Bolivian ores and depends largely on Bolivian ores for its operation. Bolivia is the only major tin-producing country in which the tin ores occur as vein or lode deposits requiring underground mining. Most of the tin from the Far East and Africa is mined by dredges from alluvial or placer depos its. It is recovered as a high-grade concentrate of cassiterite relatively free from impurities. The tin minerals in the lode deposits are more intimately associated with other minerals and more difficult to separate into a clean concentrate. Furthermore, a large part of the tin in the original ore is lost during concentration. The ores as mined contain only a few percent of tin, so they must be milled to separate the gangue minerals and produce a concentrate containing 30 to 40 percent tin before they are shipped to the smelter. The losses of tin that occur during milling vary from 18 to 65 percent. The average recovery of tin from Bolivian ores is probably 50 to 60 percent."

Dec 1, 1954

By J. E. Pahlman

The Bureau of Mines performed research on a fluid-bed magnetic roasting process for nonmagnetic taconites that utilizes low-rank coals and lignites as reductants and fuels. Roasting tests on a Michigan nonmagnetic taconite were conducted at 600° to 725° C with lignite or subbituminous coal as reductants. Partial combustion of the reductant reduced the amount of heat externally applied to the reactor to maintain the desired reactor temperature. Cleaned, dried reactor top gases were recycled to recover a portion of the reducing potential in these gases without adversely affecting concentrate grade and iron recovery. Highest iron recoveries for magnetic concentrates were obtained at a roast temperature of 650° C with lignite and at a roast temperature of 725° C with subbituminous coal. Magnetic concentrates were obtained that were comparable with commercial concentrates.

Jan 1, 1976

By H. R. Babitzke

The purpose of this Bureau of Mine- investigation was to develop columbium and tantalum alloys suitable or u at elevated temperatures. Columbium and tantalum were combined with tungsten, hafnium, zirconium, molybdenum, vanadium, and titanium, and the resulting alloys were evaluated with respect to workability, hardness, electrical resistivity, strength at elevated temperatures, oxidation resistance, and weldability. Cb-33Hf-10W-10Ti, Cb-20111-5Mo, and Cb-30Hf-5Mo exhibited superior high temperature properties. Cb-33Hf-10W-10Ti had tensile strength values of 44,300 psi at 1,200° C and 16,500 psi at 1,400° C, Cb-20Hf-5Mo had values of 35,700 and 20,700 psi, respectively. The corresponding values for Cb-30Hf-5Mo were 38,900 and 17,800 psi. These alloys are fabricable by conventional methods with excellent strength at temperatures up to and including 1,400° C. An oxidation rate of 7 to 21 mg/cm2/hr was observed up to 1,000° C.

Jan 1, 1967

By Henry J. Kellner, Michael J. Brnich

This U.S. Bureau of Mines report addresses issues involved in using a limited number of training apparatus to give hands-on self-contained self-rescuer (SCSR) training to a succession of miners, and provides guidance regarding methods for minimizing the spread of infectious diseases. Procedures used for preventing the spread of infectious organisms in dentistry, cardiopulmonary resuscitation (CPR) training, and scuba training are examined. The sanitizing methods of three organizations that require individuals to insert the mouthpiece as part of the SCSR donning sequence are discussed. It was found that each of these organizations was able to render their mouthpiece assemblies microbe free simply and inexpensively.

Jan 1, 1989