The U. S. Bureau of Mines has been investigating the application of thiosulfate to recover gold directly from low grade refractory carbonaceous ores for the past four years. Statistical experimental methods were used to determine the effects of various leaching parameters on both gold extraction and thiosulfate consumption. Seven variables were investigated, including thiosulfate, copper, ammonia, sulfite, and sulfate concentrations, leaching time, and leaching atmosphere. This information was used in face-centered cubic (FCC) surface response experiments to define the optimum leaching conditions for a low grade ore. Two models were generated from the data to predict gold extraction and thiosulfate consumption as a function of the leaching parameters. The models predicted a gold extraction of 67 pct and reagent consumption of 2 kg S2O3²-/mt ore at 0.25 M S2O3²-, 0.0014 M Cu²+, 0.27 M NH4OH, 0.00625 M SO3²-, and leaching for 16 h. Actual tests resulted in 62 pct gold extraction with 0.93 kg S2O3²- consumed/mt ore. Subsequent column leaching tests have further demonstrated the viability of applying this technology to recover gold from carbonaceous ores without first pretreating the material. Gold extractions in the column tests ranged from 69 to 78 pct with thiosulfate consumption ranging from 0.4 to 5.4
The Bureau of Mines conducted research on the flotation hydrodynamics of fine particle flotation to improve the flotation efficiency in our domestic minerals industry. The turbulent fluid velocity profile was measured in various flotation cells at different levels of dissipation energy. The importance of dissipation energy and bubble size within the flotation cells for the flotation response of a chalcopyrite ore was evaluated. Both increased turbulent fluid flow and fine size bubbles improved the flotation rate of fine size chalcopyrite particles.
The importance of surface chemistry in the flotation process has long been recognized as recorded by Fuerstenau (1962, 1976). Research to improve the chemistry of the flotation process is continuing. Arbiter (1988) commented that the flotation machine for carrying out flotation has changed little in the last 50 years. Bigger cells have been produced but the basic concept remained the same. Considerable research into the use of column flotation device has been conducted and column flotation has found a place in the mineral processing industry. However, the bulk of minerals recovered by flotation is done in standard flotation machines. The hydrodynamic effects in these cells have received little attention. Early pioneers, like Arbiter (1969), recognized that the hydrodynamic effects during
The U.S. Bureau of Mines examined the influence of turbulence on the fine bubble flotation of fine-sized (minus 10 micron) galena in order to improve flotation efficiency. Fine bubble flotation was carried out in a spinning-disk flotation cell and was compared to results obtained from flotation in a conventional cell at equivalent agitation levels. An automated electrochemical technique was used to monitor the turbulent velocity distributions of the fluid within the cell as a function of time and position. The velocity profiles were used to increase the precision of the fine particle, turbulent flotation model. While improved flotation response was obtained with intense agitation, the effect of bubble size was small. These results matched the model’s predicted response for fine particle flotation.
Fine particle beneficiation depends heavily upon froth flotation. Conventional froth flotation is generally very effective for particles between 300 and 20 µm in diameter. For smaller particles, flotation efficiency drops off. Slower flotation kinetics have been used by Nonaka (1982) to explain this inefficiency.
Schulze (1981) and others have expressed the rate of flotation of a single particle as the product of several probabilities:
Pf = Pc x Pa x Ps………………………………………………….(1)
where Pc is the probability of collision, Pa is
Investigations were conducted by the U.S. Bureau of Mines to develop flowsheets for the recovery of tungsten and gold from four Alaskan scheelite-bearing ores. Basic gravity concentration and cyanide leach tests were performed on all samples. About 46 pct of the tungsten and 30 pct of the gold from one sample were recovered in a gravity concentrate containing 24 pct WO3 (tungstic oxide) and 119 troy ounces per short ton (tr oz/st) Au. Froth flotation of the gravity concentrate resulted in separate tungsten and gold products with grades of 38 pct WO3 and 2,720 tr oz/st Au and recoveries of 60 and 97 pct, respectively. Additional gold was recovered from the gravity and flotation tailings by cyanidation. Overall recoveries of tungsten and gold from this sample were 28 and 94 pct, respectively.
A flowsheet with mill design parameters was developed that consisted of (1) free gold and scheelite recovery by gravity methods and froth flotation and (2) fine-grained gold recovery by cyanide leaching of reground gravity and flotation tailings, followed by a carbon-in-pulp (CIP) circuit. Data were developed from the Bond mill work index, gold and tungsten flotation, conventional and high-rate thickening, cyanide leaching, and CIP adsorption tests.
The United States is
We evaluated synthetic fluorspar as a substitute for natural fluorspar flux in basic cupola ironmaking. The synthetic fluorspars tested were prepared from waste fluosilicic acid generated during the processing of fluorapatite ore. The cupola trials showed the synthetic products to be the equivalent of natural fluorspar from an operational standpoint. However, wet scrubber sampling showed that substantial losses of volatilized fluorine occurred from the cupola during operation. The losses during synthetic fluorspar trials considerably exceeded those from natural fluorspar trials. Laboratory viscosity testing, using a high-temperature rotational viscometer, confirmed this finding and indicated that fluorspar volatility in slags is related to slag basicity. Synthetic fluorspar is accordingly recommended more for highly basic cupola operation (>1.3 basicity) than for neutral or acid operation, because of the increased potential for pollution and baghouse filter deterioration.
This research is a continuation of work on synthetic fluorspars applied as a flux for iron and steelmaking operations. This report covers synthetic fluorspar testing in cupola ironmaking operations and concludes the Bureau of Mines research on the metallurgical application of this material.
Fluorspar (CaF2) is the predominant auxiliary flux employed in U.S. ironmelting and steelmaking operations. It is of concern to the Bureau of Mines because of its
We treated mixed and contaminated superalloy scrap by pyrometallurgical and hydrometallurgical methods to separate and recover metal values. Best results were obtained by leaching Zn-treated or atomized scrap with HCl-O2 at 95° C and 50 psig O2. This resulted in dissolving approximately 98% of the Al, Co, Cr, Cu, Fe, Mn, Ni, and Zn while rejecting over 98% of the Mo, Nb, Ta, Ti, W, and Zr as an insoluble refractory residue. Chlorine was successfully substituted for HCl to leach Zn-treated scrap but was unsuccessful for leaching atomized scrap. The leaching solution was treated by pH adjustment and hydrothermal precipitation at 200° C for 4 h to remove Al, Cr, Fe, and other contaminants as a filterable precipitate. Recovery of Co and Ni would be accomplished by solvent extraction and electrowinning. Chromium recovery as a ferroalloy was demonstrated.
The Bureau of Mines is investigating techniques to recycle contaminated bulk superalloy scrap by sequential pyrometallurgical and hydrometallurgical methods. Cobalt- and nickel-based superalloys have critical applications in the aerospace and power-generation industries; they are used in turbines and other high-temperature equipment that require oxidation resistance and high strength at elevated temperatures. The United States must import a large percentage of the metals that
We investigated anhydrous chlorination and caustic-acid leaching processes to recover Ni, Co, Mo, and W from waste hydroprocessing catalysts. In batch laboratory-scale equipment, the chlorination process extracted 61 to 99 pct of the metals. Final recovery was 65 to 99 pct. The caustic-acid leaching process extracted 81 to 98 pct of the metals. Final recovery was 36 to 99 pct. Chlorination processing included roasting, chlorination, hydrolysis of metal chlorides for recovering Mo or W, water leaching of Ni or Co from the spent charge, purification, and solvent extraction or precipitation of Ni or Co from the purified solutions. Caustic-acid leach processing included NaOH leaching, solvent extraction of Mo or W from the NaOH leach liquor, H2SO4 leaching of the NaOH leach residue, purification, and solvent extraction or precipitation of Ni or Co from the purified solution.
An objective of the U.S. Bureau of Mines is to provide the technology that will help to assure an adequate supply of critical metals for the United States. This objective requires meeting U.S. economic and strategic needs and reducing or avoiding total U.S. dependence on foreign supplies. Achieving these goals necessitates evaluation of the potential recovery of metals from secondary sources as well as from
We investigated a hydrothermal sulfidation (HTS) and chloride leaching procedure to recover lead and silver from plumbojarosite. The plumbojarosite, generated during pressure-oxidative leaching of zinc concentrate in a commercial operation, contained, in percent, 25.4 Pb, 18.8 Fe, 1.4 Zn, 29.5 sulfate (SO4²-), 3.2 elemental sulfur (S°), 13.4 total S, and 7.4 tr oz/st Ag. The HTS was conducted in an autoclave and converted the plumbojarosite into amorphous iron oxide and liberated the Pb and Ag as sulfides (S²). The best operating conditions were 50 g of plumbojarosite and 7 pct S° (2.04 g S° added) in 500 ml of 2.0-mol/L NH3 at 150° C for 1 h. The sulfidation product contained, in percent, 37 Pb, 28 Fe, 2 SO4²-, 7 total S, and 11.2 tr oz/st Ag. Flotation of the sulfidation product was conducted at pH 3 with 15-pct pulp density and 2.4 lb/st of isopropyl xanthate as collector. The flotation concentrate contained 43 pct Pb, 25 pct Fe, and 13 tr oz/st Ag. FeCl3 leaching of the sulfidation product in a resin kettle extracted 98 pct Pb and 97 pct Ag from the suifidation product, and HCl-O2 leaching in a shaker glass bottle extracted 99 pct Pb and 98
We determined the effect of common sulfide minerals on the corrosion rates of various types of ferrous alloy grinding media. Data obtained from this study will aid in determining the contribution of any electrochemical reactions between sulfide minerals and grinding media to the total grinding media consumption. Common sulfide minerals used in this study were chalcopyrite, galena, and sphalerite. In the presence of oxygen, chalcopyrite was found to increase the corrosion rate, galena was found to decrease the corrosion rate, and the effect of sphalerite was dependent upon the type of grinding media. Possible electrochemical reactions of these minerals in the presence of grinding media are suggested.
Grinding requires a large capital investment for the minerals processing industry and frequently is the area of maximum use of power and wear-resistant materials. A typical copper producer processing 27,000 st/d of ore, may spend $25,000 per day for replacement of grinding balls. A National Academy of Sciences report estimates that the domestic copper industry consumed 205,000 st of grinding balls in 1978.
Total wear of grinding media is due to corrosion, mechanical wear, and interaction between corrosion and wear. The contributions and interactions of these processes are poorly understood, and much disagreement exists in
We investigated the role of imported natural flake graphites in dolomite-carbon refractories used in steelmaking processes and evaluated carbon fibers as a potential substitute. Varying quantities (1.5-30 pct) of natural flake graphite and carbon fibers were added to test samples. The effect of the additions on modulus of rupture (75°-2,750° F), deformation under load (2,750° F), and air-slag-metal resistance (3,000° F) was studied.
Carbon purity of natural flake graphite additions did not influence hot strength, deformation under load, or air-slag-metal resistance. When the quantity of 90-pct-carbon graphite addition varied between 0 and 30 wt pct, hot strength was highest, deformation under load lowest, and air-slag-metal wear the least at 10 wt pct. As test temperature increased from 500° to 2,750° F, the hot strength difference became less.
Carbon fiber additions were limited to 1.5 pct in dolomite-carbon brick. At this level, physical properties were generally comparable to those obtained with natural flake graphite, but were below values obtained with 10-pct-flake graphite additions. Carbon fibers are not considered a satisfactory substitute for natural flake graphite.
Continuous casting of steel combined with increased operating temperatures and demands for longer refractory life have focused attention on carbon-containing refractories. Carbon, in the form of natural flake graphite,