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By A. D. Davis, C. J. Webb

Brohm Mining Co.'s Gilt Edge Mine in the northern Black Hills was a surface heap-leach operation that was active from the 1980s until 1999. Financing problems and acid drainage concerns at the mine were reported in the late 1990s. In February 1999 the Governor of South Dakota announced plans to request the US Environmental Protection Agency to help clean up acidic water at the mine and reclaim the site. In July 1999 the parent company of Brohm, Dakota Mining, filed for bankruptcy in Canada. Because of the bankruptcy and environmental problems, the State of South Dakota intervened and began water treatment. The mine was declared a Superfund site in 2000. Hydrologic issues include the potential for acidic drainage from the site. The mine is at the headwaters of Strawberry Creek, which drains into Bear Butte Creek. About four miles downstream, Bear Butte Creek loses its discharge to swallow holes in the karstic Madison Limestone aquifer, a regionally extensive and valuable groundwater reservoir. Other concerns at the site include treatment of acidic water in mine pits and a waste rock dump.

Jan 1, 2003

By A. D. Davis, C. J. Webb

Brohm Mining Company's Gilt Edge Mine in the northern Black Hills was a surface heap-leach operation that was active from the 1980s until 1999. Financing problems and acid drainage concerns at the mine were reported in the late 1990s. In February of 1999 the Governor of South Dakota announced plans to request the U.S. Environmental Protection Agency to help clean up acidic water at the mine and reclaim the site. In July 1999 the parent company of Brohm, Dakota Mining, filed for bankruptcy in Canada. Because of the bankruptcy and environmental problems, the State of South Dakota intervened and began water treatment. The mine was declared a Superfund site in 2000. Hydrologic issues include the potential for acidic drainage from the site. The mine is at the headwaters of Strawberry Creek, which drains into Bear Butte Creek. About four miles downstream, Bear Butte Creek loses its discharge to swallow holes in the karstic Madison Limestone aquifer, a regionally extensive and valuable ground-water reservoir. Other concerns at the site include treatment of acidic water in mine pits and a waste rock dump.

Jan 1, 2002

By John D. Nelson

Unprotected uranium mill tailings impoundments can pose a significant risk to nearby inhabitants and the surrounding environment in the event of a prolonged or catastrophic release of contaminants contained in the tailings. As a result of this potential, reclamation plans for such facilities must include engineering designs to protect against disruption of the tailings and the potential release of radioactive materials. The goals for such engineering designs should be to provide overall site stability for long-term periods with no need for planned ongoing maintenance and to provide a repository for the tailings that will not place a burden on future generations. Although the goal is to provide a maintenance-free system that does not require monitoring, it must be recognized that some period of surveillance would be reasonable. The surveillance period could be relatively short and only needs to assure with reasonable confidence that the reclamation scheme that was implemented is functioning adequately. An investigation was conducted by Nelson, Volpe, Wardwell, Schumm and Staub (1983) entitled Design Considerations for Long-Term Stabilization of Uranium Mill Tailings Impoundments (NUREG/CR-3397). This report assesses the degree of confidence with which long-term reclamation plans can be expected to remain effective over specific periods of time, and to assess the additional effort required to extend the design period from one long-term period to a longer one. This paper presents a summary of that investigation and discusses some of the important conclusions.

Jan 1, 1984

The epithermal Au-Ag deposits of the Hauraki Goldfield have formed in the convergent plate margin setting of the Coromandel Volcanic Zone (CVZ). The deposits form as siliceous lodes in the upper parts of geothermal systems associated with andesitic and rhyolitic volcanism. Two types of low sulphidation epithermal mineralisation are recognised in these volcanic terranes. Vein deposits within the rhyolitic terrane in the east and south of the CVZ (Whitianga Province) are typified by Chalcedonic Silica-Adularia-Illite type mineralisation. The veins are hosted by both rhyolite-dacite volcanic rocks and the underlying andesitic basement. The mineralisation formed by episodic violent decompressional boiling of deeply circulating meteoric water, with some magmatic fluid contribution; this boiling was associated with hydraulic fracturing and hydrothermal eruptions in geothermal systems where the piezometric surface of the deep geothermal reservoir lay near-surface. Such episodic boiling provides chalcedonic veining, banding and brecciation textures, adularia stability, bladed calcite, and extreme loss of H2S gas from the fluid to cause major gold deposition.   Crystalline Quartz-Illite-(Adularia-Kaolinite) type mineralisation is widespread in the andesitic terrane forming the northern and western portions of the CVZ (Kuaotunu Province). There is a spatial association with weak mineralisation of the high sulphidation (Vuggy Silica-Alunite-Kaolinite-Illite) epithermal type and porphyry eu-Au type. Crystalline Quartz mineralisation is probably formed by the mixing of shallow acidic groundwaters with the ascending deep geothermal fluid in geothermal systems beneath andesitic stratovolcanoes. The piezometric surface to the deep geothermal reservoir in these systems is likely to lie at considerable depth, with acidic gases rising from the geothermal reservoir condensing within the overlying vadose zone. The cooling effect, oxidation and pH decrease accompanying mixing apparently overcomes the dilution effect, and results in quartz deposition, illite (and locally kaolinite) alteration, and precipitation of gold. A high level of base-metal sulphide precipitation also enhances gold deposition. Violent boiling of the mixing fluids may occur at local structural dilation sites, producing high grade "bonanza" gold mineralisation.

Jan 1, 1991

By D. L. Whiting

Dump leaching and ,in-situ solution mining have received considerable attention during the past few years due to their apparent economic viability and minimal environmental impact. Solution mining actually dates back to the 17th century when leaching of copper ore was first reported. However-, modern day leaching didn't begin on a wide scale until the mid-1900's. In-situ mining for copper, nickel, salts, and oil shale is currently looking more attractive as the technology advances. Presently, about 15 percent of the copper production in the United States is from leaching operations at approximately 14 active sites, the largest of which are the waste dumps near the Bingham Copper Mine, Utah. To establish a common background, the discussion is centered around state-of-the-art conditions. For dump leaching, complications inherent to the physical conditions and chemical reactions such as channeling, porosity and permeability, sorption, air flow, rock alteration and pH control are discussed. Emphasis is made on those conditions which render solution recovery rates and solution mineral concentrations unpredictable. By illustrating typical and actual field conditions, the generally accepted flow theories for both dump leaching and in-situ solution mining, including rubblization, are presented. For in-situ mining, hydrogeologic techniques such as geophysical well logging, aquifer testing and digital computer modeling are also discussed. The importance of artificial fracturing and well stimulating to solution injection and recovery is made with regard to project development. Finally, the subject of water pollution control techniques and the potential impact of the

Jan 1, 1977

Reliable predictions of the long-term water quality and water level fluctuations in a final void are vital if mine operators propose to decommission final voids in open cut mining operations. The catchments feeding final voids typically comprise a large proportion of coal placed or re-shaped spoil piles. Spoil piles from open cut coal mining operations have distinctive hydrological characteristics, which result in unique run-off behaviour. There is a need to understand the nature of spoil hydrology processes in order to predict the long-term hydrological behaviour and water quality responses of such systems. Preliminary estimates indicate that it may take 20 or more years for a void to flood. It is impractical for mine owners and operators to wait until long-term hydrological and water quality conditions have been established in a void, before reaching agreement with regulators about mine closure conditions. Reliable simulation of the long-term water quality and water level fluctuations in a void are therefore. The adoption of an agreed modelling approach and sharing of results from field data programs could result in the establishment of a database that could be used, with regulatory approval, for assigning parameter values, without the need for a specific data collection program for every individual void. Industry and regulators should agree on the model adopted for predicting post-closure behaviour of spoil/void systems. This will require reliable, robust methods for field data collection. This paper describes the field monitoring, data collection, model preparation and simulations undertaken over a three year period for Australian Coal Association Research Projects (ACARP) Project C7007. It presents the findings of that research, and provides comment on aspects of the project that were undertaken well and on aspects that could have been done better. Limited data is presented in this paper due to confidentiality arrangements with the relevant mine sites.

Jan 1, 2003

By Douglas W. Barr

Water is a necessary part of our environment. It is a biological necessity, but more than that, it is a necessary part of our life style. Those of us who live in the north woods expect water to be part of the scene. There are few natives of this region who would be content to eliminate the streams, the lakes, the marshes, or our forests. Throughout the world, water has been so much a part of man's environment and man's adaptation to life has been so closely tied to water that our way of life and ways of accomplishing our work include water as an essential tool. The production of taconite is an example of our dependence on water. Other methods could have been developed to process taconite ore but the process methods chosen utilize water. Water is used for grinding and separation and for trans- porting the ore from one part of the process to another and for dust control. A typical large plant producing 10 million tons of taconite pellets per year will have on the order of 130,000 gallons per minute (8.202 m 3/51 of water in circulation in the plant. About 120,000 gal. per min. (7.571 m 31s) will pass to tailing thickeners where 110,000 gal. per min. (6.940 m 31s) is separated from tailing and returned to the process. About 15,000 gal. per min. (.9464 m 31s) goes with thickened tailing to tailing basin. My point is that there are great volumes of water in motion in a taconite plant,

Jan 1, 1980

By S. Castro

Hydrolysis of metallic ions depends on pH and the products of hydrolysis either activate flotation or can lead to inadvertent activation of gangue particles, but can also depress flotation of valuable components. One of the most recently studied topics is the use of seawater in flotation. Seawater is a concentrated NaCl solution (about 0.6 mol/L) that also contains considerable amount of Mg2+ and Ca2+ ions. The hydrolysis products of Mg2+ strongly depress flotation of molybdenite over the pH ranges over which they form in the process water. This effect can be eliminated by removal of hydrolysing ions from the process water or by carrying out the flotation at the pH over which such a hydrolysis can be avoided. The former can be exemplified by prior precipitation of the harmful hydrolysing ions as, for example, in the University of Concepcion Patent. The latter can be exemplified by the use of metabisulfite to depress pyrite in the flotation of Cu-Mo sulfide ores in seawater, the process that is carried out at much lower pH, which avoids formation of Mg2+ hydrolysis products.

Jan 1, 2017

By J Liu

In situ coal gasification is a method of extracting a high proportion of the energy value from coal without the need for mining. In situ conversion of coal to combustible gases is achieved by drilling wells into the seam, injecting gasification agents and initiating combustion. The product gases are then extracted to surface using production wells. These gases are clean energy sources for power generation, heating or synthetic fuel production. In situ coal gasification is of particular interest with respect to its lower environmental impact, potential application to otherwise inaccessible energy resources (ie offshore, deep deposits or beneath protected lands) and low capital intensity. Practical application of this clean technology is restricted primarily by lack of satisfactory quantitative descriptions of the processes involved, in particular the hydromechanical processes.In this study, three characteristic physical zones of an in situ gasifier are identified as the rubble zone, combustion front and virgin coal zone. Coupled mechanical processes of coal consumption and overburden collapse are used to describe the permeability profile in the rubble zone. Additionally, the redistribution of overburden load associated with coal consumption is accounted for with the formulation of a stress (compressive) dependent permeability profile, which is applied to the virgin coal zone. The model was implemented into and solved by FEMLAB, an advanced computer simulation tool. The flow process, according to DarcyÆs Law, is subsequently described in terms of permeability dependent injectivity. Injectivity is an indicator of flow velocity, which has been identified as a crucial factor in determining the commercial value of product gases from in situ coal gasification. The stress alteration to coal permeability was found to have a dominating influence on the flow process. This finding implies that a greater understanding of the in situ coal burning induced deformation of surrounding rocks would be of significant value in assessing the economic viability of applying this clean technology to potential sites.

Jan 1, 2004

By Ronald Molnar

"High demand for metals is driving the exploitation of lower grade and more complex orebodies. These projects are often the initiatives of junior mining companies with limited financial resources. In addition, existing operations seek to enhance revenue from their existing metal recovery operations, through the production of value-added by-product metals and chemicals. Typically, hydrometallurgical circuits, often with novel chemistry, are required to exploit these new ores, to produce the multiplicity of products that will make a marginal project economical, or to increase the product range of an existing operation. Often, the mineral processing engineer is given the task of coordinating the development of a back-end hydrometallurgical flowsheet for their mineral processing project. This paper discusses the objectives that companies have when they commission the operation of a hydrometallurgical pilot plant. The paper goes on to enumerate some of the principal differences between mineral processing and hydrometallurgical pilot plants and the effect these have on the design, operability and potential pilot plant outcomes. Important factors include working with chemically reactive solutions and pulps, elevated temperatures operation, changing rheological properties of slurries, the complexity of integrated circuit flowsheets with long hold-up times and multiple recycle streams, and the need in some cases for extreme attention to purity and cleanliness. Examples will be drawn from the extensive experience base of SGS Minerals Services at the Lakefield site, in piloting both mineral processing and hydrometallurgical process flowsheets.INTRODUCTIONWe are currently experiencing a period of strong growth in the demand for metals. This in turn is turning the spotlight on a great many smaller mining companies and mining promotion ventures trying to tap into the rising metal demand. Some of these companies want to develop their properties and recover metals while others want to develop to the point that a major company will take an interest, or in some cases buy them out completely.With the heightened interest in metals, many orebodies that have been known about for years are being reconsidered for exploitation. Often, the reason they have not been developed to date is because of difficult metallurgy and marginal grades leading to marginal economic viability. In both cases the key that unlocks the resources is often the application of a novel metallurgical flowsheet, often a hydrometallurgical flowsheet."

Jan 1, 2006

By John L. Shafer

A process for the conversion of stibnite (Sb2Sj) concentrates, crude antimony oxide, or metal to a pure cubic antimonous oxide (senarmontite) has been developed on a laboratory scale. The conversion of stibnite to senarmontite has been piloted. The high purity of the final product is the result of forming antimony trichloride as a low temperature distillable intermediate. To obtain the antimony trichloride the stibnite, or crude oxide is slurried in molten antimony trichloride and reacted with gaseous hydrogen chloride. Antimony metal is oxidized to the trichloride with chlorine. The hydrogen sulfide produced from the hydro- chlorination of stibnite can be upgraded to a strength of 95% or greater for further treatment. The crude antimony trichloride is separated from the gangue by a preliminary evaporation and then is fractionally distilled and the trichloride distillate has less than 50ppm of arsenic and lead combined (the most troublesome impurities in antimony processing). The purified antimony trichloride is then hydrolyzed with sodium or ammonium hydroxide to produce the cubic form of antimonous oxide (senarmontite) to the exclusion of the rhombic form (valentinite) which is photosensitive. The advantage of this process is the ability to treat a wide range of antimony ores, concentrates, and metal to yield a high purity oxide in a controllable range of particle sizes.

Jan 1, 1975

By G. Gao

Flyash from municipal solid waste (MSW) incinerators is frequently classified as a characteristic hazardous waste due to its lead and cadmium content. Extensive investigation has therefore been carried out to detoxity the flyash by leaching and to recover lead, cadmium and zinc. Detoxification can be accomplished by leaching in HCI/NaCI solutions. Optimization of leaching conditions for effective detoxification and subsequent metal recovery can be achieved by controlling the acidity of the lixiviant and flyash to lixivaint ratio. Lead and cadmium In the leaching solution is recovered by cementation with zinc dust. Leaching and cementation experiments have shown that zinc concentrations suitable for electrowinning, > 15 g/liter, can be achieved. Experiments have also been performed to determine the effects of pH, particle size of the zinc dust used, quantity of zinc added and dissolved impurities on the cementation kinetics. After recovery of lead and cadmium, electrowinning or electrowinning coupled with solvent extraction can be used to recover zinc. A process flowsheet is presented.

Jan 1, 1991

By Niels Verbaan, Ralph Fitch, Charlotte Forstner, David Dreisinger

A process has been developed to extract gold, silver, indium, gallium, lead, copper, zinc, and other metals from a large resource using an oxidizing acid chloride heap leaching technology. The ore is crushed and/or ground to improve reactivity and then contacted with a solution containing NaCl-HCl-NaOCl. The value metals are leached into solution as chloro-complexes. The leachate is treated in a series of separation and recovery processes to produce separate saleable or disposable products. The paper will provide a summary of process development and application to a silver-indium ore from Bolivia. The possible application of the process to other ore types will be presented for discussion.

Jan 1, 2015

By David G. Dixon

Although the kinetics of leaching processes are of paramount importance in the design and analysis of nearly all hydrometallurgical flowsheets, these concepts remain poorly understood and, as a result, are only rarely applied to mass and heat balance models. In this second of two papers, we demonstrate how simple yet powerful models of leaching kinetics and other rate processes may be added to mass and heat balance models, and apply these concepts using a spreadsheet model of a simplified pyrite autoclaving flowsheet. In particular, we focus on the interpretation and scale-up of batch leaching data, and on simple means for modeling other phenomena commonly associated with leaching, such as reagent depletion, gas-liquid mixing, and the precipitation of oxides and basic salts.

Jan 1, 2002

By M. Muravyov, T. Kondrat’eva, A. Bulaev

Metallurgical treatments of sulfide ores containing copper and zinc are characterized by significant losses of base metals, particularly formation of the flotation tailings and slags during the dressing and pyrometallurgical treatment respectively. In addition, SO2 gas obtained during roasting and converting may be as raw material for sulfuric acid production. A treatment of the industrial metallurgical wastes has been investigated. Two samples of old pyrite tailings from two concentrators that contained 0.36% and 0.26% of copper and 0.23% and 0.22% of zinc were leached with 10% sulfuric acid in the column. Recovery of copper reached 47% and 38% while recovery of zinc was 47% and 41% from the first and second samples, respectively. The pregnant leach solutions contained 15.9 g/l and 10.5 g/l of ferric iron. Ferric iron is the strong oxidant and ferric containing pregnant leach solutions can be used for ferric leaching of different types of raw materials containing sulfide minerals. Therefore, ferric leaching of copper converter slag flotation tailings contained 0.53% of copper and 2.77% of zinc with the first pregnant leach solution was conducted. The effect of various experimental parameters on the leaching dynamics of copper, zinc, and iron under batch conditions was investigated. Under the best conditions (temperature 70oC, pulp density 30%) the recovery of copper and zinc reached 80% and 42%, respectively. Thus, it has been shown that acid leaching of base metals from old pyrite flotation tailings with using of the pregnant leach solution for the ferric leaching of copper converter slag flotation tailings is a prospective technique for the complex treatment of mining and metallurgical wastes. Investigated processes could be introduced into the technological flow-sheet of typical pyrometallurgical plants.

Jan 1, 2014

By K. K. Mamyrbayeva, G. Z. Guseinova, V. A. Luganov

This paper presents the results of theoretical and technological studies on leaching ore of Bozshakol deposit. The major oxidized copper mineral of Bozshakol ore is atacamite, and the content of malachite, azurite, cuprite, copper sulfide minerals is insignificant. The chemical composition of the test sample, with mass, %: 0.9 Cu; 55.00 SiO2; 22.40 AI2O3; 9.00 Fe2O3; 2.24 CaO; 3.56 MgO. Technological studies on leaching the ore were carried out with sulfuric acid 5, 15, 25 and 50 g/dm3 in one and multistage (up to 4 stages) countercurrents. Size of ore 1 mm, the ratio S : L 1: 4. It was found that with increasing of acid concentration from 5 to 50 at one-mode leaching the copper extraction reaches 55 %. At multistage leaching the copper recovery reaches 73 %, a productive solution containing g/dm3: Cu - 4; Fe – 1.2; Al – 1.1 at a pH of 2.8. When leaching the cinders obtained by ore roasting in an atmosphere with limited value of oxygen, with increasing of the roasting temperature from 650 up to 750 °? the extraction of copper from the Bozshakol ores reduced from 31.5% down to 24.4%. The rate of filtration is markedly increased compared with the crude ore. When leaching the cinders obtained by roasting in a nitrogen atmosphere a marked increase up to 87.2-88 % of copper extraction from Bozshakol ores is observed. As an extractant for the extraction of copper from the pregnant solution used LIX984N extractant in kerosene. On the basis of the calculations it was built extraction isotherm and modeled processes of copper recovery with the circuit 2E + 1P, allowing up to 93 and 96% extraction. The through extraction of copper from the ore in the electrolyte was 67 %. Technology is recommended for semi test.

Jan 1, 2016

By A. N. Zagorodnyaya, T. N. Bukurov, Z. S. Abisheva, A. S. Sharipova

"The paper describes the results of stripping of copper production dusts by leaching with aqueous soda, nitric acid solutions and recovery from solutions resulting in the production of lead sulphate, tribasic lead sulfate (TBLS), ammonium perrhenate, zinc, copper-cadmium cake.The combination of two stripping operations under chosen optimal conditions allows the recovery of 98% rhenium, 82% lead, 92 % copper, and 96% zinc and cadmium each. The major amount of rhenium (~ 80%) is recovered at the carbonization stage, and other metals - at the nitric acid leaching stage.Lead sulfate was precipitated by mixing solutions resulting from carbonization and leaching. Tribasic lead sulfate was produced by dissolving lead sulfate in sodium hydroxide solutions. Rhenium was recovered from mother liquor resulting from lead sulfate precipitation by liquid-liquid extraction and zinc was recovered from raffinates. Trialkylamine (TAA) was used as an extractant for rhenium recovery and di-2- ethylhexylphosphoric acid was used for zinc recovery. Optimal liquid-liquid extraction parameters have been defined allowing the recovery of 96% rhenium in crude ammonium perrhenate and 94% zinc in sulfuric acid solution. Based on the experimental data, the flowchart for dust processing to produce salable products - lead, rhenium, zinc salts has been suggested."

Jan 1, 2003

By D. Lunt

Telfer bulk concentrates typically contain 8% chalcopyrite, 8% chalcocite, 45% pyrite and 80-120 g/t gold. Although very high copper and gold recoveries can be obtained by subjecting the concentrate to a "conventional" high temperature pressure oxidation step followed by copper recovery by solvent extraction and electrowinning, and cyanidation of the leach residue for gold recovery, the cost of lime required for neutralisation and power for oxygen generation was such that the process was uneconomic in terms of operating costs. The challenge therefore was to develop an alternative flowsheet that maximised copper and gold recovery in which the reagent costs were commercially attractive. An additional objective was to reduce the capital cost so that the overall NPV and IRR values would enhance the economic viability of the total project. Following a series of batch and locked cycle testwork programs, supported by Metsim modelling, it was established that the same high levels of overall copper and gold recoveries can be obtained by use of a novel two-stage split leaching circuit in which the chalcocite component is leached in a low temperature (90-110°C) circuit with the chalcopyrite component treated in a high temperature (220-250°C) circuit. This split circuit has a number of advantages, including minimising the required capacity of the high temperature autoclave, as well as reducing the overall lime consumption and simplifying the total process water balance. The paper outlines the basis for the development of the flowsheet and discusses the integration of the leaching stages into the overall circuit.

Jan 1, 2004

By Jean Goldschmidt

Recycling of fine oxidized particles of non-ferrous metals usually cannot be carried out by pyrometallurgical methods. 'Hydrometallurgy provides an alternative technology to reintroduce? the resulting products as secondary raw materials into the industrial cycle, thereby saving resources and energy while keeping the environment a little cleaner. The paper summarizes the methods employed by Hydrometal to achieve these objectives and? stigmatizes the latest counterproductive international regulations which create ever more obstacles to environmentally sound recycling.

Jan 1, 1995

By J. -F. Blais, M. -O. Simonnot, J. Mocellin, G. Mercier, J. -L. Morel

Large quantities of ferromanganese sludge are generated as waste material by blast furnace during the manufacturing of ferromanganese. These slags are very rich in manganese, zinc and lead (5 to 40 %). These residues have been deposited in a pond in periphery of the steel industry. Considering the market value of the metals, these brownfield sites can now be considered as secondary resources. We have developed a hydrometallurgical process for selectively leaching Zn, Mn and Pb in order to produce compounds of Mn and Zn pure enough to be economically recoverable and a residue rich in Pb. Batch laboratory experiments were carried out to determine appropriate leaching conditions to maximize zinc extraction (100 %) and manganese extraction (90.0 %) with sulfuric acid and to generate a residue containing 15 to 30% of lead. This Pb residue is recyclable by another process.

Jan 1, 2015