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By R. G. Knickerbocker, David Schlain, F. K. Shelton, J. H. Jacobs

"Many manganese ores contain from 0.1 to several percent cobalt. If such ores arc used for the production of ferromanganese, the cobalt is lost; on the other hand, if they are used for the production of electrolytic manganese, the cobalt can be recovered during the purification of the electrolyte.The production of cobalt in the United States is insignificant compared to the amount consumed, most of which is imported from the Belgian Congo. Cobalt-bearing manganese ores may therefore constitute an important reserve of cobalt in times of emergency. Moreover, the combined value of manganese and cobalt in some ores may be sufficient to warrant the recovery of both metals by the electrolytic manganese process, whereas the recovery of either manganese or cobalt alone might not be economically feasible.In view of the potential importance of manganese ores as a source of cobalt, a survey has boon made of the cobalt content of manganese ores. Most of the data have been obtained from analyses made in the Bureau of Mines, but valuable data have also been obtained from publications of the Geological Survey, various State publications, and other sources. Unfortunately, most published analyses of manganese ores do not include cobalt determinations; when such analyses were fairly detailed, but did not include cobalt, it was presumed that cobalt was not present in significant quantity.Information on the cobalt content of domestic manganese ores is given in table 1. Ores containing 0.1 percent or more cobalt are listed in table 2, and the districts in which such ores occur are shown in figure 1. Most of the known cobalt-bearing manganese ores occur in the southeastern part of the United States. A group of deposits in central Arizona and a few scattered deposits in other States also' contain cobalt."

Apr 1, 1946

By T. H. Jeffers

The Bureau of Mines conducted laboratory tests to devise a procedure for extracting copper from low-grade discarded mill tailings. A solid extractant consisting of an organic solvent extraction reagent adsorbed on granular activated carbon was investigated and proved to be satisfactory for extracting copper directly from leach pulps. When this material was used in a cyclic leaching, stripping, and electrowinning pro-cess, 95 pct of the acid-soluble copper was recovered from oxidized mill tailings. The weathered tailings contained 0.24 pct Cu (total) and 0.20 pct acid-soluble Cu. The copper was recovered as cathodes containing, in parts per million, 1.6 Pb, 3.6 Fe, 0.5 Sb, 1.5 AI, 1.9 Mn, and <1 Mg.

Jan 1, 1985

By R. E. Head

"The microscopic work described below was engaged in primarily into ascertain the approximate proportions in which the copper sulphides existed in the ore examined, and by a comparison of the results obtained from a study of both head are tailing samples to supply data which might be of value in treating the ore by flotation. Incidentally, the physical characteristics and association of the sulphides were noted in order to determine the possible relation or bearing of these features on the present conditions of treatment. Since the copper in the ore occurred as three distinct sulphide minerals - chalcopyrite, chalcocite, and bornite each containing copper in different proportions, there appeared to be little or no possibility of obtaining the desired information by chemical methods of determination.Type of material examinedThe material examined was in the crushed condition and was said to be representative of the flotation feed or heads and the resulting tailing. No hand samples of the ore were examined nor any account taken of the physical condition and association of the minerals in place, because the possible value of the analytical results depended on conditions as they existed in the crushed ore. As the problems under consideration dealt only with the sulphide minerals, especially those of copper, the products from screen sizing were concentrated by hand panning and the gangue discarded. The concentrate obtained consisted almost entirely of the sulphide minerals, since the panning had been very carefully done. The object of this procedure was to obtain a sufficient quantity of high-grade sulphide material, from the sample of each screen size selected for examination, to permit its being so prepared for examination as to be representative of the physical characteristics exhibited by a large number of sulphide grains from each respective size, In the tests and investigations carried on by the company' treating the ore the following sizes had been used in tabulating data and basing conclusions:+48, +65, +100, +150, +200, and -200. These same sizes were used in the microscopic investigation, since the results obtained thereby could be most readily utilized for comparison with the results obtained in previous experimental work, together with deductions based on actual concentration of the ore on a commercial scale."

Jun 1, 1921

By R. S. Dean

This report, the third of a series of annual reports of the Metallurgical Division to be released in this form, covers the fiscal year July 1, 1936, to June 30, 1937. The technological work of the division is performed by sections. Thus, an efficient organization is provided to prosecute the various problems under investigation, and a logical system is set up for presenting the material contained in this report.

Jan 1, 1937

By J. E. Conley

Available statistics show that the demands of the American chemical industry for sodium sulphate have been increasing slightly in recent years. The supply from domestic sources, particularly In the form of byproduct salt cake from the acid manufacturers, has been tending to decrease due to the growing manufacture of synthetic acid. Both hydrochloric and nitric acids are being synthesized from their elements. The natural result has been that a smaller amount of salt cake is available. Simultaneously, the expansion of the Kraft paper industry, especially in the South, has brought about greater demand for sodium sulphate in that section. This dean is being met by domestic byproduct salt cake, by imported German sodium sulphate made from potash waste liquors, and by favorable situated domestic deposits of Glauber salts and tile anhydrous salt, thenardite. No shortage of natural sodium sulphate exists in the United States, but usually such deposits occur in semiarid regions far removed from transportation facilities and from markets. Moreover, the major portion of the natural deposits is Glauber salt or mirabilite (Na2SO4 ? 10H20), which must be dried or dehydrated in order to meet consumer requirements. The water of hydration represents, roughly, 55 percent of tae total weight of the crude salts and therefore has a very important influence on transportation costs. In marry instances the cost of getting, even the dehydrated product to market is the deciding economic factor. Therefore, the recovery oil production of anhydrous sodium sulphate from plant-.waste liquors or from natural brines of suitable composition might be economically feasible. The U.S. Bureau of Mines, at its nonmetallic Minerals Experiment Station, New Brunswick, N. J., has, for a number of rears, been investigating the commercial possibilities of treating the complex saline mineral polyhalite (K2SO4?NgSO4?2CaSO4?2H20) for the recovery of potash and magnesia. Certain proposed processes (1, 2, 3, 4)4/ yield waste liquors that contain the bulk of the magnesium and sulphate originally combined as magnesium sulphate in the polyhalite molecule. Sodium chloride In limited concentrations is also present, as extensive beds of salt occur in strata overlying the polyhalite deposits. A typical waste liquor would contain:

Jan 1, 1936

By J. E. Conley

[Available statistics show that the demands of the American chemical industry for sodium sulphate have been increasing slightly in recent years. The supply from domestic sources, particularly in the form of byproduct salt cake from the acid manufacturers, has been tending to decease due to the growing manufacture of synthetic acid. Both hydrochloric and nitric acids are being synthesized from their elements. The natural result has been that a smaller amount of salt cake is available. Simultaneously, the expansion of the Kraft-paper industry, especially in Lie South, has brought about greater demand for sodium sulphate in that section. This demand is being met by domestic byproduct salt cake, by imported German sodium sulphate made from potash waste liquors, and by favorably situated domestic deposits of Glauber salts and tile anhydrous salt, thenardite. No shortage of natural sodium sulphate exists in the United States, but usually such deposits occur in semiarid regions far re¬moved from transportation facilities and from markets. Moreover, the major portion of the natural deposits is Glauber salt or mirabilite (Na2SO4?10H20), which must be dried or dehydrated in order to most consumer requirements. The waters of hydration represents, roughly, 55 percent of tae total weight of the crude salts and therefore has a very important influence on transportation costs. In marry instances the cost of getting even the dehydrated product to market is the deciding economic factor. Therefore, the recovery or production of anhydrous sodium sulphate from plant-waste liquors or from natural brines of suitable composition might be economically feasible.]

Jan 1, 1936

By D. J. Bauer

The Bureau of Mines demonstrated the feasibility of fractionating lanthanide nitrates with a tertiary amine on lanthanum-cerium and thulium-ytterbium pairs. Optimum values were determined for pH, phase concentration, and percent chelation of metal ions in the aqueous phase. Percent extraction is markedly dependent on the concentration of lithium nitrate in the aqueous phase. Selective masking of the heavier element in each extraction system with DTPA or EDTA aqueous-phase chelating agents resulted in lanthanum-cerium separation factors of 8.5 and thulium-ytterbium separation factors of 3.4.

Jan 1, 1966

By N. S. Dalal, X. Shi

"Electron spin resonance (ESR) spectroscopy has been employed to· examine the role of tetraperoxochromate( V) ions (Cro:-) and other Cr(V) species in the generation of hydroxyl (OH) radicals in the reaction of Cr(VI) with H2O 2 in biological media. In contrast to earlier suggestions, the present ESR studies using crystalline K3CrO8 as a source of Cro:- show that decomposition of Cro:- in water or in H2O 2 does not generate significant amounts of OH radicals. Addition of NADH to a solution containing Cro:- yields a Cr(V)-NADH complex, which readily reacts with H2O 2 to generate OH radicals. Similar results obtained from several other biological reductants, including vitamin B2 , indicate that a vacant coordination site on a Cr(V) complex facilitates its reaction with H2O 2 to generate ""OH radicals. We thus suggest that in biological media, reaction [3] instead of [1] or [2] is the major pathway for the OH radical generation: While the carcinogenic potential of Cr(Vl)-based compounds has been well established via epidemiologic and animal studies (1), the underlying biochemical mechanism is still not well understood (2). In a recent development, however, Kawanishi et al. (3) have reported that hydroxyl (OH)2 radicals might be the key species in the mechanism of Cr(VI) toxicity. As to the mechanism of the ""OH radical formation they suggest that Cr(VI) reacts with cellular hydrogen peroxide (H2O 2) to form the tetraperoxochromate(V) (Crog-) ions, which decompose to produce hydroxyl radicals. It was also suggested (3) that the ""OH radicals thus produced cause genotoxicity via DNA base modification and deoxyribose phosphate backbone alterations. More recently Aiyar et al. (4) reported the detection of OH radicals in the reaction of Cr(VI) with H2O 2 in biological media, including the reaction of ""OH radicals with deoxyguanosine to generate 8-hydroxydeoxyguanosine adducts, confirming that ""OH radicals could indeed play a significant role in the mechanism of Cr(VI) toxicity. However, the mechanism for the ""OH generation was not clarified, although the role of Crog- ions as possible precursors was mentioned, somewhat along the lines suggested by Kawanishi et al. (3). The possibility of the participation of Crog- in the ""OH generation has also been indicated by Sugiyama et al. (5), since they mention that vitamin B2 enhances the formation of""OH radicals in the reaction of Cr(VI) with H2O 2 because of the generation of Crog- in the reaction."

Jan 1, 1990

By E. Morrice

This report of the Federal Bureau of Mines presents a laboratory procedure, with the requisite equipment, for electrowinning ingot (liquid) cerium metal from a fused-fluoride bath, the metal purity being about 99.9 percent. Physicochemical studies, including bath properties and electrochemistry, are included, with a speculative model of the mechanisms involved. An historical review of the methods used for electrowinning misch metals and yttrim-group metals, as well as individual rare-earth metals, is presented. The electrolytic preparation of metal from fused salt and nonaqueous media is summarized, along with the unsuccessful attempts at depositing the metals from aqueous solutions.

Jan 1, 1961

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 M. J. Ferrante

The Bureau of Mines recovered rhenium powder from scrap tungsten alloy containing approximately 25 percent rhenium. Oxidation of the scrap at 950° C separated volatile rhenium heptoxide from tungsten trioxide. Minor amounts of tungsten, iron, and nickel were removed by selective precipitation from a rhenium solution from which was recovered crystalline ammonium perrhenate. Experimentation showed that hydrogen reduction of minus 325-mesh ammonium perrhenate at 1,000° C produced rhenium powder averaging 2.1 microns in size (Fisher Sub-Sieve method). The powder contained <50 ppm of total metallic impurities and 1,070 ppm of total interstitial impurities. Sintered compacts had a density of 95 percent of theoretical. Approximately 96 percent of the rhenium contained in the scrap alloy was recovered.

Jan 1, 1969

By M. B. Shirts

Numerous alternative processes have been proposed as being potentially more economical or less polluting than the traditional smelting of copper sulfide concentrates. One such process as recently developed in the laboratory by the Bureau of Mines is composed of a two-stage fluid-bed, oxidation-reduction roasting followed by ammoniacal leaching, solvent extraction, and electrowinning. A copper-leaching recovery of 97.4 percent was obtained, and electrolysis of the acid strip solution deposited metal assaying 99.99 percent copper. The roaster gases contain over 95 percent of the sulfur at a concentration of 12 percent which is better suited for acid making or elemental sulfur production than conventional smelter gas. When considering byproduct recovery, a preliminary estimate indicates copper production costs are com-parable to the conventional smelters.

Jan 1, 1975

By R. E. Coleman

The molecular species occurring at the mineral-reagent interface of the galena-aqueous potassium ethyl xanthate system was studied by developing a technique for using an attenuated reflection method in conjunction with standard infrared spectroscopy procedures. Lead ethyl xanthate was detected from the infrared spectra of all galena surfaces treated in the pH range of 7 to 11. At pH 8.5, lead ethyl xanthate formed as a film chemically bonded to the galena surface. A solvent extraction sampling method was employed with infrared spectroscopy to obtain quantitative data on the lead ethyl xanthate removed from the galena surface. The data showed that between pH 7 and 11 thy quantity of lead ethyl xanthate physically adsorbed decreased as the pH was decreased; in the same pH range, the quantity of free lad ethyl xanthate in the aqueous medium remained constant, and the amount of unreacted xanthate varied.

Jan 1, 1966

By W. W. Simpson

During cyanidation of mercury-bearing gold-silver ores, significant amounts of mercury are extracted. The presence of mercury decreases. gold loading and increases mercury stripping time on activated carbon, complicates fire refining of the gold cathodes, and creates a potential health hazard. Addition of calcium sulfide (CaS) to the cyanide leach slurry precipitates the solubilized mercury but also some silver. The Bureau of Mines investigated silver loss during CaS precipitation of mercury. Addition of 0.1 lb CaS per short ton ore to a mercury-bearing gold-silver ore precipitated both mercury and silver. However, if copper was added with CaS, or if the ore contained copper, much of the sil-ver loss was prevented. Addition of 235 ppm Cu with 0.1 lb/st CaS, prior to carbon-in-pulp adsorption, resulted in 89-to 97-pct, 80-to 91-pct, and O-pct adsorption of the solubilized gold, silver, and mer-cury, respectively, on the activated carbon.

Jan 1, 1986

By J. A. Rowland

Manganese-copper alloys containing 60 percent or more electrolytic manganese exhibit an unusual combination of high vibration damping capacity with good strength properties. Properties of the pure alloys were first studied in the laboratories of the Federal Bureau of Mines and have been described in a series of technical papers1/ published from 1941-49. The unusual characteristics of these manganese-copper alloys attracted the attention of research metallurgists and several additional investigations have also been reported in the literature.2/ The engineer, also, was attracted by the combination of high strength and high vibrational damping capacity possessed by these alloys. A wide field of application in structural and machine design was obvious. However, serious difficulties encountered in commercial fabrication of structural or machine components from the alloys retarded attempts at their practical utilization. For this reason the Federal Bureau of Mines, in cooperation with the United States Navy, Bureau of Ships, undertook a program designed to develop practical methods for fabricating a selected group of the high-damping alloys. It is the purpose of this paper to summarize a part of that investigation in a form suitable for use by the metallurgist interested in commercial fabrication of the alloys.

Jan 1, 1955

By E. D. Gardner

Open-cut mining of copper ores in the western United States is a relatively new industry. At first the mining methods were mainly adapted from those used at open-cut iron mines and at quarries, but now efficient practices that are more suitable for the copper deposits have been developed. The pioneer in this field is the Utah Copper Co., of Bingham, Utah, which deserves great credit for working out much of the early procedure. Stripping began at that mine in 1906, and the large Garfield concentrator was put in commission in 1908. Next in order of development were the Nevada Consolidated Copper Co., at Ely, Nev. ; the Chino mine of the Ray Consolidated Copper Co., at Santa Rita, N. Mex. ; the New Cornelia Copper Co., at Ajo, Ariz. ; the Sacramento Hill mine of the Copper Queen Branch, Phelps Dodge Corporation, at Bisbee, Ariz.; and the open cut of the United Verde Copper Co., at Jerome, Ariz. The Sacramento Hill mine began to produce in 1923, and the United Verde open-cut mine has produced ore mainly in connection with stripping. The writer visited the Utah Copper mine in September, 1923; the Nevada Consolidated in August, 1923; the Chino in July, 1924; and the New Cornelia, the Sacramento Hill, and the United Verde open cut at various times during 1922, 1923, 1924, and 1925, mostly in connection with other work.

Jan 1, 1927

By L. A. Neumeier

A refinement of the tungsten-cobalt phase diagram is presented for compositions to 85 at. pct cobalt and for temperatures to the solidus. Preparation of alloys by powder metallurgy methods is described, and the results of studies made by a combination of techniques involving microscopy, X-ray diffraction, thermal analysis, and dilatometry are discussed. The existence of u-W6Co7 is confirmed (rhombohedral D85 structure). The W6Co7 phase forms periectically at 1,685&apos;115° C by reaction of liquid with primary W; the peritectic melt is at 71 at. pct cobalt. The homogeneity range of u-W6Co7 extends to both sides of the ideal stoichiometric ratio, through a compositional range of about 5 at. pct. The compositional limits for u-W6Co7 vary slightly below 1,471°±5° C, the temperature of the eutectic reaction forming u-W6Co7 plus a. The eutectic composition is at 79 at, pct cobalt. The existence of WCo3 is confirmed (hexagonal D019 structure). WCo3 forms peritectoidally at 1,093°±5° C by reaction between a and u-W6Co7 and exists over a narrow homogeneity range of 0.5 to 0.6 at. pct. The solid solubility of cobalt in tungsten is low (maximum 0,9 at, pct cobalt) and results in contraction of the tungsten lattice. Tungsten is extensively soluble in a-Co at higher temperatures; the a-Co lattice is expanded by the presence of tungsten in solid solution. There was no indication of the formation of a o-phase nor of any intermediate phases other than WeCo7 and WCo3. The WeCo7 and WCo3 phases are compared with similar phases in other systems.

Jan 1, 1967

By J. L. Eakin

The Bureau of Mines determined agents that are effective in foaming heavy brines and oil from gas wells in severe concentrations of sodium chloride, in severe concentrations of calcium chloride, or in moderate oil-brine mixtures. Dynamic foam tests were designed to select the more effective foaming agents. Numerous foaming agents were tested in concentrated simulated brines, in oilfield brines having varied concentrations of ions, in heavy natural brines, and in a commercial brine generally used as a reservoir pressure-balancing fluid. Several agents were detected that could perform efficiently in the field under the most severe saline conditions. Agents that promote foaming in oil-brine mixtures were also determined.

Jan 1, 1965

By A. Y. Lee

Prior work at the Bureau of Mines resulted in the successful development of a bench-scale, combination electrorefining-electrowinning method for recycling the lead from scrap batteries using waste fluosilicic acid (H2SiF6) as electrolyte. This paper describes larger scale experiments. Anodes cast from scrap battery lead were electrorefined in a 20-L multielectrode cell for 3 to 7 days. The anodes, containing 2 to 2.5 pct antimony, were ideal for obtaining firm and adherent slime blankets. Cathode deposits, assaying 99.99 pct Pb, were obtained, with current efficiencies near 99 pct. Sludge leaching was done in 100-L tanks followed by filtering in 61 cm square vacuum pan filtersu Lead of greater than 99.99-pct purity was recovered from the filtrate by electrowinning in a 20-L multielectrode cell using insoluble Pb02-Ti (lead dioxide-coated titanium) anodes and pure lead cathodes. The fluo-silicic acid electrolyte depleted in lead was repeatedly recycled to leach more sludge, and there was no problem with impurity buildup. The amount of Pb02 formed on the anodes during electrowinning was less than 1 pct of the total lead deposited on the cathodes, as long as the phosphorus concentration in the electrolyte was greater than 1.3 giL. Emissions of lead in air were less than 10 ~g/m3, well below the proposed OSHA limit of 50 ~g/m3.

Jan 1, 1984

By Allan M. Johnson

Case histories were used to document and evaluate deficiencies in underground mine closures which have resulted in post-operation problems such as subsidence shaft failures, acid or toxic water drainage and others. Three different underground mining districts in the United States were investigated; 1) an iron mining district in Michigan 2) coal mining districts in eastern Ohio and southern Illinois and 3) three metal mining districts in central Colorado. Previous work in the Michigan district on problems with subsidence and acid water drainage provided the background from which the concept for this project was developed Student investigators working towards their MS degrees in Mining Engineering did the work on the eastern coal and western metal districts. The combined studies show that problems from acid water drainage, subsidence and inadequately protected mine shafts are the most severe and widespread. Suggestions of a technical nature are presented as partial solutions to some of these problems. It was determined that pre-closure evaluations could result in the identification of some potential problems that may occur after the mine is closed. Early recognition of these problems would be an advantage for at this time more options are available to the mine operator to either prevent them or to plan for their occurrence. Mine closure and reclamation should be considered as a part of mining and be incorporated in the mining plan. This would result in reduced environmental impact and would be achieved at a lower overall cost to the mining companies and to society. The role of water in the occurrence of several types of subsidence during mine flooding was significant at some mines in the Michigan district. These mechanisms which involve, friction reduction, gravity loading, overburden transport and hydraulic forces in a flooded mine probably have a broader effect than is currently recognized. Control of these processes may offer some interesting options during mine closure. It is recognized that technical solutions alone cannot provide all of the answers to problems associated with underground mines. Education, modified attitudes on the part of regulators and operators, and a more realistic perception of the importance and value of mining by the general public are also needed.

Jan 1, 1983