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By Masami Hayashi

Cost evaluations were made for five processes for producing 100,000 tons of electrorefined copper per year from a chalcopyrite concentrate. The processes were (1) conventional smelting with all furnace gases vented to the atmosphere; (2) conventional smelting with converter gases routed to an acid plant and reverberatory gases vented to the atmosphere; (3) conventional smelting with converter gases routed to an acid plant and reverberatory gases to a citrate system plant; (4) conventional smelting with all furnace gases routed to a citrate system plant; and (5) a modified smelting process in which the concentrate is roasted before being charged to the reverberatory furnace, with roaster and converter gases routed to an acid plant. The total fixed capital costs of the plants were estimated as (1) $56.7 million, (2) $77.4 million, (3) $87.2 million, (4) $73.9 million, and (5) $77.0 million, respectively. Total S02 removal from stack gases as sulfuric acid or sulfur for each process was (1) none, (2) 66 percent, (3) 99 percent, (4) 98 percent, and (5) 78 percent. Assuming no credit for byproduct sulfur or sulfuric acid, total operating costs per pound of electro-refined copper produced were estimated as (1) $0.081, (2) $0.103, (3) $0.119, (4) $0.111, and (5) $0.103, respectively.

Jan 1, 1974

By Robert T. MacMillan

An essentially chemical technique has been developed on a small scale for the efficient utilization of certain domestic titaniferous iron ores. Three main steps are involved: (1) Sintering the ore at 1,050° C. with carbon and soda ash to reduce the iron to metal powder, retaining the titania in the slag, (2) separating the powdered iron from the slag, after wet grinding, by magnetic or gravity methods, and (3) sulfuric-acid decomposition of the titania fraction to produce pigment and certain byproducts. The technique was first applied to the tailings from the Bureau of Mines bauxite-concentrating plant at Bauxite, Ark. A good grade of powdered iron was produced from this material. Subsequently, the work was diverted to the study of the titaniferous magnetite from the MacIntyre Mine at Tahawus, N.Y., because of the potentially higher recoveries of iron, titanium and vanadium. Ninety percent of the iron in this ore has been recovered as an iron powder analyzing 90 percent metallic iron and containing 2 to 3 percent TiO2. The remaining titania was found in the slag and was recovered as pigment-grade TiO2 by application of a suitable technique. The vanadium was distributed between the TiO2 fraction and the alkaline liquor resulting from the wet-grinding and magnetic-separation operations, with only a minor proportion retained by the metallic iron.

Jan 1, 1950

By M. R. Peterson

The U.S. Bureau of Mines investigated column flotation for recovery of a high-grade fluorite (CaF2) concentrate and byproduct concentrates from the Fish Creek fluorite deposit in Eureka County, NY. The recovery scheme consisted of (1) grinding the ore to minus 48 mesh, (2) fluorite rougher and cleaner flotation, (3) desliming the rougher fluorite flotation tailings at 20 Jjm, (4) mica flotation, and (5) silicate rougher flotation. Acid-grade fluorspar, mica, free silica (Si02) sand, and low-grade beryl (Be3Al2S~018) concentrate were produced in a 100 lb/h continuous column flotation unit (CCFU). The highest results achieved were %.6-pct CaF2 recovery in a fIrst cleaner concentrate containing 99.7 pct CaF2. Beryl recovery was as high as 92 pct at a grade of 5.3 pct BeO. The silica sand, which assayed 98 pct Si02, was recovered as the rougher beryl flotation tailings. Collector reagent consumption was reduced 45 pct for fluorite and 58 pet for silicate flotation in comparison with bench-scale conventional flotation. All products were signillcantly improved in both grade and recovery over conventional flotation results, and the recovery scheme was simplifIed by removal of several stages of flotation in the fluorite and silicate flotation steps.

Jan 1, 1990

By H. H. Fields, Joseph Cervik, G. N. Aul

Coal is a soft and brittle material. Drilling rates in the Pittsburgh coalbed using a drag bit exceed 3 ft/min at 2,500-pound thrust. However, maintaining the bit on a horizontal trajectory or parallel to bedding planes to attain lengths of 1,000 feet is difficult. This Bureau of Mines report presents a drill string configuration that can be guided through the coalbed. The angle of the borehole is measured periodically and by proper combinations of bit thrust and rotational speed (r/min), bit trajectory can be lifted or dropped to keep the bit in the coalbed. An example of a hole drilled to 2,126 feet is presented to show levels of thrust and rotational speed used to lift or drop the bit, the seemingly unexplainable events that occur, and the strategy used during drilling. Thrust levels during drilling range from 600 to 2,500 pounds. Hydraulic motors powered by a 30-hp, 440-volt motor provide ample power for drilling horizontal holes 2,000 feet long.

Jan 1, 1975

By Glenn A. Hazen

This report presents a stress investigation of four coal pillars located under two depths of cover. Each coal pillar was instrumented with ten load cells designed to measure coal pressure during initial and secondary mining. The problem considered was one of uniform triaxial stress as a large coal pillar was loaded under a thick limestone bed. Special attention was given to pillar design procedure. Unconfined uniaxial compression tests were conducted for coal specimens 7/8 in., 2-1/8 in., 3 in., and 4 in. square with length to square side ratios of 1/2, 1, 1-1/2, and 2. Stress-strain curves were recorded. Six specimens were tested at each size. Results show that a coal pillar is composed of an inner zone which is highly loaded plus a region around its perimeter which has lost its elastic strength. The bearing capacity of a coal pillar consists of elastic and crushed coal response: The failure strengths have been determined and will be useful in coal recovery, mine design, and mine safety.

Jan 1, 1975

By W. K. Collins

Engineered scavenger compounds (ESC's) developed by the U.S. Bureau of Mines are a novel class of compounds that selectively can recover a desired element from a solid or molten alloy. Lithium titanate (L2Ti307 or Li2O?3Ti02) is used as an ESC to recover lithium (Li) from aluminum-lithium (Al-Li) alloys. X-ray diffraction measurements have shown that Li2Ti307 undergoes a phase change during scavenging from an orthorhombic structure to a hexagonal structure. This change is due to the incorporation of lithium in the matrix of the material and the effect of temperature. Although both phases are metastable, the hexagonal phase that forms during the scavenging of lithium from Al-Li alloys appears to be the more stable phase. Recovering lithium from the ESC by electrodeposition does not cause the structure to revert to the orthorhombic phase. The orthorhombic and the hexagonal structures of Li2Ti307 have similar scavenging capacities for lithium. This report proposes a new mechanism for the phase transformation.

Jan 1, 2010

By Joseph Cervik

Coal is a soft and brittle material. Drilling rates in the Pittsburgh coalbed using a drag bit exceed 3 ft/min at 2,500-pound thrust. However, maintaining the bit on a horizontal trajectory or parallel to bedding planes to attain lengths of 1,000 feet is difficult. This Bureau of Mines report presents a drill string configuration that can be guided through the coalbed. The angle of the borehole is measured periodically and by proper combinations of bit thrust and rotational speed (r/min), bit trajectory can be lifted or dropped to keep the bit in the coalbed. An example of a hole drilled to 2,126 feet is presented to show levels of thrust and rotational speed used to lift or drop the bit, the seemingly unexplainable events that occur, and the strategy used during drilling. Thrust levels during drilling range from 600 to 2,500 pounds. Hydraulic motors powered by a 30-hp, 440-volt motor provide ample power for drilling horizontal holes 2,000 feet long.

Jan 1, 1975

By H. R. Babitzke, H. Kato, M. D. Carver

A number of columbium and tantalum alloys were investigated to determine their suitability for use at elevated temperatures . Columbium-vanadium , tantalum-vanadium , and tantalum-hafnium alloys , and also columbium- 10 wt pct titanium alloys modified with additions of 0.1 to 2 mole pct ZrO2 , TiO2 , ZrB₂ , TiB , ZrC, or TiC were prepared and examined . Workability , tensile properties , and oxidation resistance were tested . Of the columbium alloys studied , the columbium- 12 at . pct vanadium exhibited the best properties ; it compared closely with the columbium- 10 at . pct vanadium alloy studied in a previous investigation.

Jan 1, 1964

By A. Dale Marcy, Pamela L. Drake

Workers who perform routine welding tasks are potentially exposed to fume that may contain manganese. Manganese may cause respiratory problems and is implicated in causing the occurrence of Parkinson-like symptoms. In this study, a field colorimetric method for extracting and measuring manganese in welding fume was developed. The method uses ultrasonic extraction with an acidic hydrogen peroxide solution to extract welding fume collected on polyvinyl chloride filters. Commercially available pre-packaged reagents are used to produce a colored solution, created by a reaction of manganese(II) with 1-(2-pyridylazo)-2-naphthol. Absorbance measurements are then made using a portable spectrophotometer. The method detection limit and limit of quantification (LOQ) were 5.2 mg filter-1 and 17 mg filter-1, respectively, with a dynamic range up to 400 mg filter-1. When the results are above the LOQ for the colorimetric method, the manganese masses are equivalent to those measured by the International Organization for Standardization Method 15202-2, which employs a strong acid digestion and analysis using inductively coupled plasma-optical emission spectrometry.

By R. August Heindl

[A proposes process for recovering manganese from open-hearth steel-plant slogs is indicated by the flow diagram shown in figure 1. Investigation of the process, the laboratory development of which was reported previously,4/ was expanded to a larger scale. Mixtures of flush slags and limestone were proportioned to have molar [SiO2] ratios of 1.5 to 2.2 and were fired at temperatures up to 2,400° F. In an oxidizing atmosphere. This produced clinker consisting principally me various combinations of the oxides of manganese, iron, aluminum, and magnesium aloe with a dicalcium silicate phase containing the phosphorus. Some tests were made with a continuous sintering machine of the traveling-grate type, but a rotary kiln was the most satisfactory apparatus for this operation. Material was fed at 150 to 200 pounds per hour after first being agglomerated either on the sintering or else binder extruding and drying finely ground mixture of the sing, limestone, binder, and water. Tests with slag; from tree different source indicated overkill manganese recoveries of 65 to 80 percent by ammoniacal-ammonium-carbonate leaching.]

Jan 1, 1955

By E. G. Davis

As a result of interest by the U.S. Department of Energy, the Bureau of Mines conducted studies on comminution of bituminous coal to the 2-um particle size range using the Bureau-developed turbomill. Both "plant grind" size (80 pct minus 75 um) and microsize (minus 10 um) coal can be produced in a single step with the Bureau's turbomill. In 15 min, minus 2.4-mm coal milled in water with steel shot was reduced tc 65 pct minus 75 m and 26 pct minus 2 m, with an energy requirement of 139 kW^h/mt coal. Plant-grind size coal milled in water, with Ottawa sand as the milling medium, was reduced to more than 45 pct minus 2 m in 15 min, at 175 kW?h/mt. Plant-grind coal milled with steel shot was reduced to 57 pct minus 2 um in 15 min, at 138 kW?h/mt. When diesel fuel was substituted for water, the milling of the coal was less effective and less energy efficient. Keeping the pulp dispersed was critical. Without an effective dispersant, the slurry became very viscous with increased milling time, resulting in higher energy requirements, low heat dissipation, and poor grinding efficiency.

Jan 1, 1987

By F. W. Benn

The Bureau of Mines investigated flotation methods for recovering fluorite and byproducts from two samples of the Fish Creek deposit in Eureka County, NV. The preferred method for tre.ating a fluorite-beryllium sample included (1) fluorite rougher and cleaner flotation, (2) desliming at 20 um, (3) muscovite flotation, (4) silicate flotation, (5) sodium hypochlorite wash of silicate concentrate, and (6) beryllium rougher and cleaner flotation from the silicate concen-trate. Laboratory results showed that over 94 pct of the fluorite was recovered in a recleaner concentrate containing 98 pct CaF2, Beryllium flotation produced concentrates containing in excess of 5 pct BeO with recoveries over 70 pet. The preferred method used for a fluorite-mica sample consisted of (1) fluorite rougher and cleaner flotation, (2) desliming at 20 um, and (3) mica flotation using an anionic-cationic collector system. Laboratory results showed that 84 pct of the fluorite was recovered in a recleaner concentrate containing 96 pct CaF2, Mica flotation produced a concentrate containing 95 pct mica at 85-pct recovery.

Jan 1, 1986

By M. Hertzberg

The Bureau of Mines measured the energy requirements for the spark ig-nition in air of Pittsburgh seam bituminous coal dust, lycopodium spores, and polyethylene powder with a 1.2-L furnace and 8-L and 20-L chambers. Thermal autoignition temperatures of the same dusts were mea-sured in the 1.2-L furnace. Electrical ignition requirements are given in terms of both effective spark gap energies, [ ] and stored circuit energies, 1/2 CE2. The measured order of electrical ignitability for the three dusts is consistent with the data of other researchers; however, the absolute values are systematically higher, probably because of higher flow and turbulence levels in the chambers used and lower electrical efficiency in the circuit used here. The temperature dependence of the lean limit of flammability for lyco-podium was measured wi th the 1.2-L sys tem, and those measurements con--firm the applicability of the modified Burgess-Wheeler law to a dust. Due to experimental complexities, the minimum ignition energies for dusts may not reflect intrinsic flammability behavior. However, some valuable information may be obtained from the relative ignition energies of various dusts at ambient and elevated temperatures. In addition, the concept of minimum electrical ignition energies for homogeneous gas mixtures is reevaluated theoretically.

Jan 1, 1985

By S. Akhtar

Interstitial compounds of iron, cobalt, and nickel, and their binary and ternary mixtures with silver, gold, palladium, and platinum were prepared and tested as fuel cell catalysts. The first step in the preparation of the interstitial compounds consisted of preparing finely divided metals. This was done either by the Raney method or by reduction of precipitated hydroxides of the metals. The finely divided reduced metals were then treated with carbon-and/or nitrogen-containing gases to prepare their carbides, nitrides, and nitrocarbides. The materials were tested for cathodic activity towards O2 and for anodic activity towards H2, C3H8, NH3, CO, and N2H4 in acid and alkaline electrolytes at 25° to 90° C in half-cell assemblies. Samples of carbided and nitrocarbided INi-lAg-1Au were about one fifth as active as Pt-black in the reduction of O2 in 35 percent KOH solution, whereas the activities of several interstitial preparations exceeded the activity of Pt-black in the reduction of O2 in 12M KOH containing N2H4. Nitrocarbides of x-Fe2C, 3Ni-1Au, and 1Ni-3Au were effective catalysts for oxidizing N2H4 in liquid NH3 containing NH4SCN. Corrosion of the catalysts by electrolytes was severe in many cases.

Jan 1, 1970

By S. D. Hill

To help insure the continued viability of the minerals and materials sector of the U.S. economy, the Bureau of Mines is conducting research and disseminating technology that will result in more efficient minerals extraction. As part of this effort, a process is being investigated that involves an aqueous chlorine-oxygen leaching procedure to treat complex sulfide concentrates to produce ZnCl2. Subsequent fused-salt electrolysis of the ZnCI2 produces high-purity zinc metal. Previous reports described preliminary, small-scale research on the chlorine-oxygen leaching system and fused-salt electrolysis of ZnCl2. This report discusses the details of design, construction, and longterm, continuous operation of both monopolar and bipolar fused-salt cells. The effects of temperature, electrode gap, current density, electrolyte composition, and electrode configuration on current efficiency and energy consumption were determined for a ZnCl2-KC1-NaCl electrolyte system. Zinc metal was produced at the rate of about 50 lb/day in a monopolar cell which was operated continuously for 60 days with current efficiencies from 85.8 to 98.6 pct and an average power consumption of 2.1 kwhr/lb. Bipolar cells with several different electrode configurations were operated for 90 days. A current efficiency of about 70 pct and power consumption of about 1.7 kwhr/lb of zinc were achieved using a three-electrode-pair, bipolar assembly with a chlorine gas lift pump to circulate electrolyte.

Jan 1, 1981

By E. G. King

As low-temperature heat-capacity and entropy data were available for only one sesquisulfide (cerium sesquisulfide), it appeared desirable to augment the information for this class of compounds by measuring the sulfides of antimony and indium. In addition a pressing need existed for data on indium sulfide to supplement vaporization studies in progress in the Department of Mineral Technology, University of California. This Bureau of Mines paper reports low-temperature heat-capacity measurements and entropy evaluations at 298.15° R. of antimony sesquisulfide of stoiehiometric composition (Sb2S3) and of indium sesquisulfide which was somewhat deficient in sulfur (In2S2.93). The data for the latter substance were adjusted to correspond to the stoichiome tric composition (In2S3 ) . Indium sesquisulfide has a lower entropy at 298.15' K, than either antimony sesquisulfide or cerium sesquisulfide, in line with the higher atomic density of the indium compound.

Jan 1, 1962

Illumination hardware developed in conjunction with previous Contracts H024202l, H0242049, and H0242023 had not been adequately tested in actual underground work situations. The purpose of this contract was to install and monitor this equipment on mining machines and in actual working places of coal mines to gain experience with the use of the lamps and support equipment in the abusive and critical environments for which they were intended. The equipment that was to be tested consisted of the following: a. Lamps -A variety of sizes of permissible fluorescent lamps which operate from 12 volts DC power. These include the following sizes: Model 8, (9 watts) 2" diameter; Model 15/2 (14 watts) 2" diameter; Model 15/3 (21 watts) 3" diameter; Model 20 (25 watts) 3" diameter; Model 30, (28 watts) 3" diameter; and Model 40, (39 watts) 311 diameter. Representative photographs of these lamps are shown in Figure 1-1 and 1-2.

Jan 1, 1976

By J. P. Coughlin

Gypsum plaster is calcium sulfate hemihydrate (CaS04-1/2H2O). There are two varieties, commonly designated as a and ß. The hydration setting) of this plaster involves the chemical reaction: CaS04?1/2H20(c) + 3/2H20(l) = CaSO4?2H20(c); that is, the hemihydrate combines with water to form the dihydrate. Many substances, when present as impurities, either in the hemihydrate or the water have influenced the rate of the hydration reaction. The objective of the Bureau of Mines work described in this report was to investigate the accelerating or retarding effects of small amounts of dissolved substances and to attempt to classify these effects upon a chemical basis. To eliminate competing effects, the work was conducted separately on the two varieties of hemihydrate; both were in highly pure form. (Commercial plasters usually are a mixture of the hemihydrates, with other substances present as impurities.) As the hydration reaction of each variety of hemihydrate evolves heat in amounts known from previous work, 5,6/ the authors decided to use calorimetric methods to follow the extent of the hydration reaction as a function of time. Two calorimetric methods were employed. One was designed to study the hydration reaction with a large water to plaster ratio and the other with a small water to plaster ratio (to simulate commercial practice in utilizing plaster).

Jan 1, 1959

By E. O. Binyone

"https://www.youtube.com/watch?v=PBPyGXjlBHIhttps://www.youtube.com/watch?v=3AbtiKOYg7Y&feature=push-fr&attr_tag=uKtVW3yq2ZngptnE%3A6https://www.youtube.com/watch?v=PR2THayf-N8https://www.youtube.com/watch?v=mF-BXiWocqIhttps://www.youtube.com/watch?v=TuesIw-OZA4https://www.youtube.com/watch?v=v2HP7sTIZgEhttps://www.youtube.com/watch?v=v2HP7sTIZgEINTRODUCTIONIn 1937 the Eureka Corporation, Ltd., of Canada obtained a lease on the mining property of the Richmond-Euroka Mining Co. in the Eureka mining district, Eureka County, Nevada. In 1910 the Richmond-Eureka Co. suspended all operations except small-scale leasing, but company and other geologists contended that the El Dorado dolomite would also carry ore in its continuation from fault block north and northeast of the productive old working. The company began to explore the fault block in 1923 by diamond drilling from the,900 level of the Locan shaft, but for various reasons the hole was drilled to a depth of only 2 feet and was then abandoned. Eureka Corporation, Ltd., planned an extensive diamond-drilling program, which included four deep diamond-drill holes, one designated ""A"" fig. if it is to be drilled from the surfaced, and three designated ""E"", ""C"" and ""D"" figs land to be drilled from the 900 level of the Locan shaft. Of these only B and C were completed by the company. Hole C was a contribution,of the hole started by the, started by the Richmond Eureka Co. The site for hole D (figs. 1 and 2) was prepared, but the hole was not drilled. Both B and C holes penetrated sulfide ore carrying good percentages of zinc, lead, gold, and silver. The ore bodies were 40 feet thick and were found approximately 1,400 feet below the 900 level of Locan shaft.On the basis of the very favorable results obtained from the drilling program, the Eureka Corporation planned extensive development of the property and in 1942 commenced sinking the Fad shaft (figs. 1 and 2). The shortage of equipment and materials brought on by the war and difficulties with water in the shaft greatly hindered progress. However, early in 1943 the Eureka Cor¬poration, Ltd., of Canada submitted to the War Production Board a proposal for developing the mine and milling 500 tons of zinc-lead ore a day. The proposal involved an expenditure of $2,890,750."

Dec 1, 1946

By R. L. Hopkins, H. T. Rall

A continuous method has been developed for preparing thiolacetic acid . Acetic anhydride and hydrogen sulfide are passed concurrently over an amine ion- exchange resin , under 20 - psi pressure . The thiolacetic acid , a product of the reaction , is flashed continuously from the crude reaction mixture and is purified by fractional distillation . The yield of thiolacetic acid is in excess of 80 percent based on the acetic anhydride charged . INTRODUCTION In the synthesis of reference compounds needed in the identification of sulfur compounds in petroleum , thiolacetic acid has become an important reagent because of its use in the synthesis of thiols by addition to appropriate olefins : R - CH = CH₂ + CH3 - COSH R- CH2 - CH2 - S - COCH R - CH2 - CH₂ - SH + CH3COONa NaOH (1 ) The commonly used procedures for its preparation , as described by Clark and Hartman (1 )4 and by Ellingboe (2 ) , are based on the batchwise reaction of acetic anhydride with hydrogen sulfide :

Jan 1, 1964