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By R. J. Roman

The results of column leaching tests on 3 different mixed oxide/sulfide ores from a Chilean deposit are presented. The 3 samples contained different ratios of oxide copper minerals to chalcocite and chalcopyrite. The results of the column leach tests are interpreted in light of the shrinking core model for coarse particle leaching. The copper recovery from the oxide minerals and the sulfide minerals each conforms to the shrinking core model with the total recovery equal to the weighted average of the individual minerals. The use of column test results to predict field heap leaching results using the multi-mineral shrinking core model is discussed.

Jan 1, 1993

By P. L. Sibrell

The U.S. Bureau of Mines has tested cyanide leaching for recovery of palladium (Pd) from spent petroleum processing catalysts. Three different catalyst samples were supplied by a spent-catalyst processor. These catalysts consisted of a zeolite base and contained 0.4 to 0.7 pct Pd. During alkaline cyanide leaching, the catalysts exhibited ion-exchange properties due to their zeolite matrices. Hydrogen ions were released from the zeolite in exchange for sodium ions in solution, resulting in a significant decrease in solution pH. This could present a safety hazard because of the potential for release of toxic hydrogen cyanide gas. A pretreatment .step where the catalysts were contacted with a 1.0M sodium hydroxide solution was found to prevent the pH shift from occurring. Following the sodium hydroxide pretreatment, two stages of leaching at 160 °C with solution containing 1 pct sodium cyanide and 0.1M sodium hydroxide gave at least 75 and up to 95 pct Pd recovery. The Pd was quantitatively recovered from the leach solution by thermal decomposition in an autoclave at 250 °C for 1 h. The Pd content of the precipitate was over 50 pct. Thermal decomposition also decreased the total cyanide content of the barren solution to less than 0.2 mg/L. The catalyst leach residues passed the Federal Toxicity Characteristic Leaching Procedure and the California Waste Extraction Test, indicating that landfill disposal of the leach residues would be acceptable.

Jan 1, 1995

By D. L. Leach

This report covers laboratory experiments simulating in situ copper recovery from primary sulfide ores in sulfuric acid systems pressurized with oxygen. Copper extraction and acid consumption data are correlated with temperature in the range 50-150°C and quenched pH in the range 0.25-2. Adverse changes in the chemical composition of the leach solution and the formation of gangue-mineral alteration products and secondary sulfides can severely limit the extraction of copper above 90°C. Keeping the pH below the self-buffered value of about 2 and the temperature below about 110°C produces the maximum extraction rates. Forcing the pH too low not only results in excessively high acid consumption but may also limit the extraction of copper at temperatures above 100°C.

Jan 1, 1975

By M. Walawalkar, C. K. Nichol, G. Azimi

Rare earth elements (REEs) are essential to the production of technologically advanced products, including wind turbines, electric and hybrid cars, automotive catalytic convertors, hand held electronic devices, and military defense systems. Their application in environmentally beneficial technologies puts them at a high priority for ameliorating environmental concerns. Because of their increasingly high demand and limiting resources, it is highly beneficial to find new sources for them. The present study is focused on leaching REEs from a phosphate based fertilizer by-product, known as phosphogypsum (PG). Global estimated world production of PG is about 240 million tonnes that is stored in the form of wet stacks adjacent to fertilizer production facilities, occupying vast amounts of land. PG also contains trace metals, which are detrimental to the environment, and hence it is crucial to monitor these stacks. We obtained our PG sample from Agrium Inc. (a fertilizer manufacturing company in Canada). We performed a comprehensive characterization of the PG sample and found that it predominantly contains Y, La, Ce, Nd, and Sm. For leaching purposes, we used hydrometallurgical acid leaching techniques. We investigated three acids over a range of operating conditions such as acid concentration, temperature, solid-to-liquid ratio, and residence time. We achieved an average leaching of 81% with hydrochloric acid as leachant using operating conditions that can be easily replicated in industrial settings. If our process is implemented, not only will it reduce the size of the existing PG stacks (mitigating their environmental risks), but also enable the development of new innovative routes for extracting REEs, making countries like Canada less vulnerable to supply disruptions.

Jan 1, 2016

By Fathi Habashi

SELENIUM AND tellurium are by-products of electrolytic copper refining. Copper anodes contain 0.01-0.2 % Se and up to 0.04 % Te; during electrolysis these deposit as a slime at the bottom of the cell. Such slimes contain 2-15% Se and 0.5-8 % Te, as selenides and tellurides of gold, silver, and copper. The following compounds have been identified in the slimes: (Ag, Au)Te2, Ag2Se, Ag2Te, CuAgSe, Cu2Se, Cu2Te. Anodic slimes from copper electro¬lytic refining are a grayish black powder, minus 200 mesh. When this complex mixture is decomposed in an acid medium, elemental selenium and tellurium can be obtained, while in an alkaline medium, both metals go into solution. Acid leaching The slime is charged in acid digesters of 800-900 lb capacity, heated to 260°C with 78 % H2SO4, for 8-12 hr. During digestion the following reactions take place:

Jan 1, 1970

By Katarína Gáborová, Marcela Achimovičová, María Florencia Márquez‑Zavalía, Jaroslav Briančin, Oľga Šestinová, Pavel Škácha

In this paper, the structure, morphology, chemical composition and leaching of naumannite ( Ag2Se) were studied. Mechanical activation of Ag2Se in a planetary ball mill resulted in increasing its specific surface area. The influence of milling time on the particle size was studied by particle size distribution analysis and scanning electron microscopy. The leaching kinetics of non-activated and mechanically activated samples in 3.5–7.9 mol/L HNO3 solutions were examined. Subsequently, the effect of leaching temperature in the range of 25–50°C on the recovery of silver from mechanically activated naumannite was documented. The maximum yield of silver 94% was achieved after 120 min of leaching in 7.9 mol/L HNO3 at 50°C for the sample mechanically activated for 30 min. Based on calculated activation energy values from the Arrhenius equation, a mixed leaching mechanism consisting of diffusion and surface chemical reaction as the rate-controlling steps of the solidliquid reaction was assessed.

Mar 4, 2023

By Pelin Altinkaya, Maria Leikola, Jussi Liipo, Mika Haapalainen, Mari Lundström, Eero Kolehmainen

With decreasing ore grades, the tailings of mining operations are becoming of increasing interest as metal-containing secondary raw materials. The objective of the current work was to investigate chloride leaching of gold, copper, cobalt, nickel, and zinc present in the flotation tailings. In the current study, the effect of cupric ion as an oxidant (0–50 g/L) and NaCl (150–250 g/L) on metals extraction was investigated. The other parameters, such as pH (1.8), temperature (95 °C), solid/liquid ratio (25%), oxygen feed rate (1200 mL/min), leaching time (24 h), and stirring rate (950 rpm), were kept constant. Gold dissolution rate was shown to increase with increase in cupric ion concentrations up to 50 g/L. Also, increase in NaCl concentration up to 250 g/L increased gold extraction. Majority of the copper present in the flotation tailings could be dissolved with all the all solutions investigated. The other base metals, cobalt, nickel, and zinc extractions were also shown to increase with increase in the cupric ion and sodium chloride concentration. However, even in the absence of cupric ion addition (t = 72 h, 250 g/L NaCl), the final extraction of Cu, Ni, Zn, Co, and Fe increased up to 98, 93, 83, 76, and 80%, respectively. This shows the power of inherently originating oxidants present in the tailings. Furthermore, the solid analysis of the leach residue indicated that leaching the flotation tailings in pure NaCl may result in partial gold dissolution. The results demonstrate that significant amount of metals present in the tailings could be extracted even with only NaCl as added chemical in the presence of oxygen feed, oxidizing agents originating directly from the raw material. This can provide an advantageous cyanide-free method for extraction of metals from very low-grade tailings with low chemical consumption.

By Hai-Peng Guo, Lu Jiang, Xiao-Man Yan, Bing Xie, Hong-Yi Li

Traditional, calcium compounds cannot be used in vanadium extraction processing due to the limitation of sodium roasting technology, which results in unsatisfied dephosphorization for vanadium-containing hot metal in converter and thus increasing the dephosphorization burden for subsequent steelmaking. The calcium roasting technology is an alternative for vanadium extraction and allows adding lime into vanadium slag in converter. In the present study, vanadium was leached by soda from roasted vanadium slag with high calcium content. Effects of roasting temperature and time, leaching temperature, Na2CO3 concentration andliquid to solid ratio on vanadium leaching rate were studied. The results show that 94% of vanadium was extracted by roasting the vanadium slag at 950℃ for 2.5h and then leaching in 160g/L Na2CO3 liquid at 95℃ for 1h with liquid to solid ratio of 10:1 ml/g. Impurities of Si and P reacted with calcium and formed insoluble Ca2SiO4 and Ca3(PO4)2 respectively.

Jan 1, 2015

By Zhigao Xu, Ming Wu, Jun Qu, Rong Zhang, Wenrui Nie, Ruan Chi, Zhenyue Zhang, Zhengyan He

The leaching processes and mechanisms of weathered crust elution-deposited rare earth (RE) ores with ammonium sulfate were investigated for their high efficiency and low consumption exploitation. The leaching process is a typical reversible heterogeneous ion exchange reaction in liquid–solid systems. The raw ore and leached tailings were analyzed by SEM and FTIR. Results indicated that no dissolution or structural changes of the ores occurred during the leaching process. The migrations of RE3+, Al3+, and NHþ 4 in the leaching process were studied under various liquid–solid ratios. With increasing liquid–solid ratio, the content of NHþ 4 adsorbed on clay mineral by physical effect decreased, NHþ 4 adsorbed by ion-exchange increased, and the content of exchanged RE and Al increased until reaching equilibrium. The adsorption capacity of the RE ores for NHþ 4 remained about 0.61–0.63 mg/g. It was difficult to desorb NHþ 4 from RE ores with washing water alone. The leaching mechanism of Al was the same as RE, but the leaching of Al slightly lagged behind RE. At the optimal liquid–solid ratio of 0.4:1, the leaching efficiencies of RE and Al were 95.51% and 90.99%, respectively, and 56.65% of NHþ 4 remained in the leachate. After impurity removal and RE precipitation, the leachate can be reused by replenishing fresh leaching agent. The aim of this study is to enhance the leaching process of RE, restrain the leaching of impurities, and decrease the consumption of leaching agents.

By N Guresin, R Dunne, S Blanco

This paper presents the results of research work conducted to assess if aqueous solutions of sulfur dioxide (SO2) can be used to treat roaster calcine leach residues such that recovery of gold in this material can be achieved with subsequent application of the thiocyanate (SCN-) leaching system. The project is mainly aimed at assessing an alternative use for sulfur dioxide gas generated during flotation concentrate roasting. The calcine leach residue sample used in the study contained 2.56 g/t gold. The reason for this high level of residual gold is thought to be due to gold surface passivation by thin iron oxide films, presence of gold in solid solution with refractory iron oxides and encapsulation of gold in the structure of these iron oxides.The study was conducted in three stages. The first stage was the characterisation of a sample from a roaster calcine leach residue. The second stage in the project was a set of leaching tests to assess the feasibility of using aqueous solutions of sulfur dioxide to reductively dissolve refractory iron oxide mineral phases contained in the calcine. The final objective was to assess the efficiency of thiocyanate, a gold complexing ligand stable under acidic and oxidative conditions, to extract gold liberated during sulfur dioxide leaching.The study showed that the reductive sulfur dioxide treatment of the calcine leach residues followed by oxidative thiocyanate leaching to extract gold was not a particularly efficient process as only around 20 per cent gold could be extracted. Leaching the calcine residues by any of these systems separately also extracted the same amount of gold (12 - 14 per cent) implying that irrespective of the solution potential, the acidity helps dissolution of gold encapsulating minerals, mostly hematite, but at a very slow rate. The findings on the topic in the literature and during this study show the requirement for further technical investigations and economical analyses on two cases. The first case is to roast gold concentrates to produce magnetite rather than hematite followed by sulfur dioxide and thiocyanate leaching of magnetite containing calcines. The second case is to apply a magnetic concentration process to conventional roast-cyanidation calcine residues to produce a gold-rich magnetite concentrate that can be further treated with sulfur dioxide followed by thiocyanate leaching.For both cases, the best evaluation of the sulfur dioxide leach system in extracting residual gold from calcine leach residues will require a detailed economic analysis to find out the balance between the value of the gold recovered and the loss of revenue from acid production. There are a number of roasters that do not produce acid. In these cases, a portion of the unwanted sulfur dioxide gas can be used to recover gold in magnetite containing calcines as an economic incentive.CITATION:Guresin, N, Blanco, S and Dunne, R, 2013. Leaching roaster calcine leach residues with aqueous solutions of sulfur dioxide and thiocyanate, in Proceedings World Gold 2013 , pp 129-136 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 26, 2013

By Alex Luyima, Honglan Shi, Lifeng Zhang, Yongqing Jiang

"Leaching process has been suggested to recover most metals from waste PWBs. In this study, the leaching behavior of most metals present in waste PWBs is evaluated before and thermal treatment. Different leaching acids: hydrochloric acid (HCl), nitric acid (HNO3) and aqua regia are compared. The effects of acid concentration, particle size of sample, leaching time and temperature are examined. Results revealed that most metals are present in small particle size and sequential leaching of PWBs with a combination of HCl acid and aqua regia is capable of dissolving most of the metals content of PWBs. Leaching of thermally treated PWBs produced the highest leaching percentages for similar leaching conditions.IntroductionApproximately 20–50 million tonnes of electronic waste (e-waste) is produced globally per year. Waste Electrical and Electronic Equipment (WEEE) or E-waste is one of the fastest growing waste streams in the world. In developed countries, it equals 1% of total solid waste on an average [1]. The growing stockpile of used and obsolete consumer electronics has been called the “largest toxic waste problem of the 21st century [2]. Printed wiring boards (PWBs) is an essential part of almost all WEEE, and accounts approximately 3% of all WEEE by weight [3]. PWBs are a platform upon which microelectronic components are mounted. PWBs contain metals (~40 wt. %), ceramics (~30 wt. %) and plastics (~30 wt. %) [4]. The metallic part includes; Cu, Pb, Zn, Fe, Cd, Be, Cr, Au, Pt, Pd, Ag, etc. The plastic part of the PWBs is mostly epoxy resins. These plastic resins also contain very toxic halogenated flame retardants (HFRs) [5]."

Jan 1, 2012

By Manis Kumar Jha

Leaching studies were carried out for the recovery/ recycling of metals from PCBs containing 17.05% Cu, 0.74% Ni, 1.74% Fe, 4.35% Pb and 8.32% Sn using various acids such as H2SO4, HCl and HNO3. While sulfuric acid was not a suitable leachant for the dissolution of metals from PCBs, hydrochloric acid selectively dissolved tin. The nitric acid was found to be effective lixiviant with the recovery of 99.99% Cu, Fe, Ni and 36.7% Pb in 6M HNO3 at S/L ratio of 100 g/L and 90oC. The kinetic studies carried out with 2M and 4M HNO3 at 90oC showed ?Ash diffusion control dense constant size-spherical particles? model.

Jan 1, 2010

By R. Dugdale, F. Habashi

Studies on hand-picked samples of chrysocolla from Arizona showed that heating in an inert atmosphere enhances greatly copper extraction by ammoniacal-ammonium carbonate solutions. Heating in hydrogen improves the extraction of copper even further and maximum extraction by ammoniacal-ammonium carbonate and by sodium cyanide solutions is achieved for samples heated at 350°C. Heating in SO2 atmosphere results in the simultaneous formation of CuSO4 and Cu2O; maximum reaction rate is also achieved at 350°C. At temperatures higher than 350°C, CuCO4 starts to decompose. Chrysocolla heated above 600°C either in argon, H2, or SO2, is rendered less amenable to leaching, supposedly due to the formation of insoluble copper silicates.

Jan 1, 1974

By Luísa Maria Abrantes

Environmental control is becoming more and more stringent. As conventional processes for zinc production are pollutant routes, alternative methods have become an important subject of research. In this regard, the leaching processes have aroused a considerable interest, namely the electrochemically induced approach. The behaviour of zinc concentrate ore in the leaching processesses is influenced by several parameters. Chloride concentration and leaching solution pH, time and temperature of reaction were studied in order to evaluate their influence on the zinc recovery. The efficiency of the leaching process for zinc extraction is performed by factorial design. The experiments were carried out by 23 and 22 factorial designs of the factors. Results show that all factors, except for the time of reaction have a significant influence on the optimisation process, within each experimental region.

Jan 1, 2001

By W. A. McKinney

The U.S. Bureau of Mines, Kennecott Copper Corp., and New Mexico Institute of Mining and Technology are cooperating on a research project to improve extraction of copper from previously leached waste dumps that are in part compacted and impervious to leach solutions. The field work is being done at the "J" leach dump, Chino Mines Division, Kennecott Copper Corp., Santa Rita, N. Mex. The laboratory percolation leaching tests were made at the U.S. Bureau of Mines Salt Lake City Metallurgy Research Center on drill core samples from the "J" dump to determine how much copper might be leached if permeability could be restored by explosive heaving or other means. The core samples from nine drill holes on "J" dump contained from 0.06 to 0.14 percent copper. Copper extractions ranging from 15 to 3 percent were obtained in 75 days by percolation leaching of minus-1-inch material with dilute sulfuric acid. The tests results indicate that about 2'7 percent of the copper remaining in the "J" dump could be extracted under ideal leaching conditions.

Jan 1, 1970

By Renfeng Song, Shaoyan Wang, Zhiqiang Zhang, Ke Guo

Leaching titaniferous magnetite concentrate with alkali solution of high concentration under high temperature and high pressure was utilized to improve the grade of iron in iron concentrate and the grade of TiO2 in titanium tailings. The titaniferous magnetite concentrate in use contained 12.67% TiO2 and 54.01% Fe. The thermodynamics of the possible reactions and the kinetics of leaching process were analyzed. It was found that decomposing FeTiO3 with NaOH aqueous solution could be carried out spontaneously and the reaction rate was mainly controlled by internal diffusion. The effects of water usage, alkali concentration, reaction time, and temperature on the leaching procedure were inspected, and the products were characterized by X-ray diffraction, scanning electron microscope, and energy dispersive spectroscopy, respectively. After NaOH leaching and magnetic separation, the concentrate, with Fe purity of 65.98% and Fe recovery of 82.46%, and the tailings, with TiO2 purity of 32.09% and TiO2 recovery of 80.79%, were obtained, respectively.

Jun 16, 2021

A fundamental component of grade control at Tarmoola Mine has been the monitoring of assaying through quality assurance programs. The programs have concentrated on issues related to accuracy and precision of leachwell and fire assay techniques, sample preparation and grade variability within a sample. An outcome of monitoring has been identification of shortcomings in the application of assaying techniques to Tarmoola ore types that contain coarse particulate gold. Leachwell has shown satisfactory precision since introduction as the primary assay technique in October 1998 but leachwell accuracy and recovery through comparison with fire assay techniques has identified bias. A bias toward higher fire assays in grades above 2 g/t invoked an investigation into the cause of the discrepancy between leachwell and fire assay. Testwork evaluated Leachwell assay recovery by varying leach time duration, which showed that by increasing the leach time improved recovery significantly. Monitoring of leachwell against fire assay had shown that a leachwell (200 g) with a one hour leach has under-called grade when compared with fire assay by up to 18 per cent, on average, for Tarmoola ore types. Alternatively, optimisation of the leachwell assaying technique using a 200 g pulp and a two hour leach has returned, on average, higher assays than fire assay by up to eight per cent. Testwork on screen fire assays, duplicate assays and fire assay to extinction has demonstrated assay variability associated with coarse, particulate gold. The variability and distribution of gold in Tarmoola ore types has emphasised the importance of high quality sample preparation for leachwell assaying to ensure liberation of gold to leaching solutions. Assay quality assurance protocols are vital to assay data integrity and enable confidence in delineation of ore-waste boundaries and provide early recognition of problematic assaying periods and isolation of grade control practices, which depart from performance standards critical to the operation.

Jan 1, 2000

By G. L. Oldright, Henry de Rycker, S. F. Ravitz

The upper limit of richness of concentrates that can be smelted by means of the blast furnace without added diluents is fixed by the operation of sintering. A sinter feed with normal gangue constituents, containing more than about 50 per cent of lead, tends to fuse; as a result, air does not come sufficiently into contact with the sulphides at roasting temperatures, and sulphur elimination is poor. Since most lead concentrates now contain 55 to 80 per cent lead, it is necessary to add diluents to lower the grade of such sinter feed to about 50 per cent or less. This not only entails the expense of providing barren diluents when adequate diluents containing precious metals are not available but also appreciably lowers the capacity of the blast furnace to produce lead and increases slag losses. Attempts to desulphurize rich lead concentrates by methods other than sintering have not proved successful. Flash roasting, for example, converts much of the lead sulphide to lead sulphate rather than to oxide. However, this suggested the possibility that deliberate flash roasting of the concentrates to sulphate, followed by sinter roasting of the resulting product, might prove promising, particularly inasmuch as the elimination of sulphur from lead sulphate requires heat alone and not contact with atmospheric oxygen. The sintering of lead sulphate products has additional interest because important amounts of such products are now treated by lead smelters. Considerable quantities of lead sulphate are present in rich lead flue dusts and other by-products; lead-chamber sulphuric acid plants and other

Jan 1, 1944

By U. C. Tainton

FOR the last 15 years lead has been the standard material for anodes in electrolytic zinc production and it has been generally accepted that this lead should be as free as possible from impurities. Laist1 says, "Next in importance to the purity of solution is the purity of the electrodes and tank lining material. The purest lead obtainable should be used for anodes." The same conclusion is expressed by Ingalls2 and by two German investigators who say that experiments have demonstrated that the purer the lead, the greater is its resistance to corrosion during electrolysis.3 Pure lead, however, while the best material hitherto available for anodes, is not entirely free from objection. The plain lead anode gradually disintegrates under electrolysis, so that the life of an anode, say ¼ in. thick, is frequently not more than two years. Part of the lead from the anode finds its way to the cathode, lowering the purity of the deposited zinc and decreasing the hydrogen overvoltage. The rest of the lead that comes from the disintegration of the anode goes into the manganese dioxide which is precipitated in the cells, rendering unsaleable what might otherwise be a valuable by-product. Yet a further disadvantage of plain lead anodes comes from their tendency to bend or buckle during electrolysis, as a result of intercrystalline oxidation. This makes it necessary to use a fairly wide spacing between anode and cathode to avoid short circuits, and thus leads to a higher power consumption than would otherwise be necessary. The power consumption is also affected by the high decomposition voltage at a lead peroxide surface.

Jan 1, 1929

By Felix Edgar Wormser

JUST as the ancients used the products of their crude mining endeavors to fashion tools with which to make digging easier, so today mining enterprises are dependent upon the very metals they mine for the efficient conduct of mining and metallurgical operations. In fact, the outlet afforded by the production of minerals in the United States is no mean market for the nonferrous metals. I believe that mine managers frequently lose sight of the opportunity they have to select from competitive products those which will help the market for their own metal output. For example, some years ago a large lead mine used storage-battery locomotives for its haulage, common practice, of course, throughout the United States. These particular locomotives, however, were powered not with the ordinary lead storage battery but by a competitive battery product made from an imported metal. After studying this situation, the manager saw how he could help himself by purchasing a product made from his own mineral output, and in his next order turned tip the use of the lead storage battery.

Jan 1, 1935