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By Lawson F

Pyrite concentrate and calcine from the Nanjing Lead-Zinc-Silver Mines, China, were leached with aqueous cyanide and thiourea solutions. More than half of the contained gold and silver was able to be extracted from the roaster calcines. The very fine dissemination of precious metal grains, high roasting temperature (850 - 900°C) and residual sulphur are believed responsible for the Iow extractions.A recovery of about 70 per cent for both silver and gold was able to be obtained with cyanide and with thiourea leachants when the pyrite was roasted in the laboratory at temperatures between 600 and 650°C.Thiourea, as an alternative leachant, has received much attention recently and a series of constant potential leaching tests was conducted.When the thiourea to formamidine disulphide mole ratio was about two, the precious metal extraction was greatest. In the Nanjing calcined material, this occurred at a potential of about 210 mV relative to a standard calomel electrode.

Jan 1, 1987

By Kazutoshi Haga, Atsushi Shibayama, Altansukh Batnasan

The objectives of this study were to propose a viable alternative method for recovery of precious metals such as gold (Au), silver (Ag) and palladium (Pd) from waste printed circuit boards (WPCBs) and to regenerate the spent lixiviant solution for reuse. The proposed method generally consists of several steps as follows: 1). iodine-iodide leaching for metals extraction; 2). separation and recovery of precious metals and metal impurities from a pregnant leach solution; 3). regeneration of the spent iodine-iodide solution/lixiviant. In this method, more than 99% of Au leaching was achieved in the iodine-iodide solution from the solid residue of WPCBs that was obtained via sequential processes including incineration at 800 ?C and high pressure oxidative leaching with a diluted sulfuric acid at 120 ?C, while the leaching of Ag and Pd was less than 1%, respectively. Besides Ca, Cu, Ni, Co and Zn as main metal impurities, leaching efficiency of other metals did not exceed 1%. The precious metals and base metals as metal impurities present in the pregnant leach solution were separated and recovered by precipitation using L-AA and NaOH solutions. About 99.8% Au, 81.7% Ag and 72.8% Pd were recovered from the pregnant solution at pH of 2 with the addition of 0.1 M L-AA as precipitating agent. The remained small amounts of precious metals and metal impurities were precipitated at pH 13 using 1 M NaOH. The spent lixiviant after sequential precipitation was regenerated in the presence of hydrogen peroxide, hydrochloric acid and potassium iodide. It was revealed that over 90% of gold could be extracted from the residue of WPCBs with the regenerated lixiviant solution.

Jan 1, 2018

By K. Barani, S. M. S. Ganji

Creatine was used as a lixiviant for gold dissolution from gold ore. The effects of creatine concentration, temperature, leach time, and pH were examined for their influence on extent of gold dissolution. The results show that at a temperature of 75°C and pH 11.5, gold dissolution increased from 51% to 88% on increasing creatine concentration from 25 to 50 g/t. Further increase in creatine concentration showed a negative effect on the gold dissolution. Gold dissolution increased from 47% to 87% by increasing the temperature from 25°C to 55°C, and from 61% to 89% by increasing pH from 10 to 11.5. Under the optimum conditions of solids’ content of 30%, pH 11.5, creatine concentration of 100 g/t, temperature of 75°C, and hydrogen peroxide addition of 2%, about 90% of the gold could be dissolved from the ore after 24 h. A comparison between creatine and cyanide leaching showed that 90% gold dissolution was achieved at a cyanide concentration of 300 g/t, which is three times higher than the creatine concentration under optimal conditions.

Apr 14, 2024

By Jorge A. S. Tenorio, Pedro F. A. Prado, Denise C. R. Espinosa

There is an increasing challenge of waste management regarding the growth of photovoltaic (PV) technology. Often overlooked by its environmental benefits of greenhouse gas emission reduction, PV technology contain metals that are highly hazardous for the environment. Amorphous silicon (a-Si) PV modules employ indium-tin-oxide as transparent conductive oxide and given the high price of indium its recovery could be advantageous from an economic perspective as well. Leaching of indium was predicted by comparing the Pourbaix diagram with the values of oxidoreduction potential and pH measured on a phosphoric acid solution. The a-Si panel was subjected to thermal treatment to remove the proprietary, protective coating and allow leaching to occur. Then, indium could be leached by phosphoric acid in a jacketed reactor, reaching a concentration of 55 ppm (67% leaching rate) with a 0.2 g/mL of a-Si panel to aqueous solution ratio. The leaching of indium with phosphoric acid is novel and could open doors for many other recycling routes for solar panels.

Mar 1, 2018

By F. Milton Lewis

Several alternatives are available for treating lover grade gold ores averaging 0.04 to 0.10 ounce per ton. These include heap, vat and agitation leaching and may involve conventional zinc precipitation or carbon adsorption-electrowinning systems. Each of these processes have their inherent economic and operating advantages and disadvantages and the selection of the optimum method for a given gold ore is solely based on economic evaluation of the applicable processes. In this paper an effort is made to provide engineering feasibility estimates for the above processes under varying operational conditions such as ore grade and price of gold.

Jan 1, 1974

By Fathi Habashi

METALS COMMONLY leached may come from one of the following sources: 1) Occurring in nature in the native state, e.g., gold, silver, platinum metals, copper. 2) Produced as by-product of a metallurgical process, e.g., gold, silver, platinum metals, selenium. 3) Produced by reduction of an oxide ore, e.g., copper, nickel. GOLD AND SILVER Cyanidation According to Mellor (1923) the alchemists of the eighteenth century knew that gold was soluble in aqueous solution of potassium cyanide, and the early jewelers probably utilized this fact in gilding. He also mentions that Scheele in 1783 and later Bagration in 1843 noted the solvent action of aqueous solutions of alkali cyanides on gold. Yet it was not until 1887 that the value of this observation was recognized, when MacArthur and Forrest (1887) made use of it in leaching gold from its ores. Since then, the process has spread to every gold and silver mining district in the world.

Jan 1, 1970

By Bernard Scheiner

A number of wells were drilled at two phosphogypsum stacks, one at P-21 owned by IMC and the other at Agrico's South Pierce Stack. Each well was sampled, and composites representing 5-ft increments were leached with synthetic rainwater to determine the extent of metal and nonmetal migration. To simulate what was occurring in the stack, composite samples were placed in a column, so that the column represented the stack in miniature form. Leaching experiments showed that both metals and nonmetals moved down the column when water was added to the top of the column. Data obtained from P-21 showed that with time the concentration of metals and nonmetals reached equilibrium values. The concentration of Ca, SO4, F, Si, and PO4 were -600 ppm,~1,500 ppm, 32 ppm, 14 ppm, and 22 ppm, respectively. Column leaching of South Pierce material, representing 0 to 55 ft of stack, showed that after 355 in. of water, the stack was near equilibrium values down to the 33-ft level. With continued addition of water, the portion of the stack at equilibrium would continue to increase and eventually reach the values observed for P-21. The test showed that the sulfate concentration of the effluent from the stack would never reach the USEPA standard value of 250 ppm. During the leaching test, it was determined that the permeability of the phosphogypsum in the columns was 0.000103 cm/sec. This number was used in the modeling. The data obtained from the leaching experiments cast some doubt on the practice of entombment now being practiced for closure of phosphogypsum stacks. The data clearly shows that with time, the amount of material leaching from the stack is low in concentration, with only the calcium and sulfate at higher levels. When a stack is closed at an existing operating plant, allowing rainwater to fall on the stack and capturing the solution in the existing pond system would eliminate the need for covering the stack, since with time the closed stack would continue to leach and at some point reach the equilibrium values of the P-21 stack. Entombment means that the leachable material in the stack will always be there and will require monitoring forever. Allowing rainwater to decontaminate the stack may be the best environmental solution to the problem.

Jan 1, 1999

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 S. P. Fleming

An acid mine drainage sludge containing 0.26% nickel, 0.21 % cobalt and 4.5% manganese was leached in a one-liter reactor with sulfuric acid in the absence and presence of sodium sulfite. The effects of temperature (25°C and 52°C) and sodium sulfite concentration (up to 0.03 molar) on the leaching rate were particularly investigated. The leaching rates of nickel and cobalt with sulfuric acid have been determined to be controlled by diffusion through product layer. It has been found that effective leaching conditions to give more than 90% extraction of nickel, cobalt and manganese while keeping the iron extraction below 10% come with an elevated temperature and a low sodium sulfite concentration at pH 3; for example, 52°C and 0.01 molar sodium sulfite.

Jan 1, 1992

With shortage of sulfide nickel ores, laterite ore was paid more attention by scientists. Because its grade is commonly lower than sulfide nickel ores, laterite ore can not be utilized effectively be-coming the bottleneck of study. Leaching laterite ores from Yuanjing, Yunnan province, organic ac-ids show better selective than sulfide acid in nickel and cobalt recovery. Organic acid has low de-pendent on pH, keeping leaching capability even higher than 3. On the other side, when pH up to 2, sulfide acid can not leach laterite ores any more. After four kinds of organic acid leaching laterite ores, besides citrate, malate and succinate have good leaching and effectively selective capability. Especially, succinate has the best selective leaching capability of nickel and cobalt in four kinds of organic acids for latter anion purification and separation. Isolated from waste water with sugar, Aspergillus niger (zs1) has great capability to produce organic acids and vigorous growth. Laterite ores have been respectively processed by direct-leaching and indirect-leaching. Indirect-leaching gives a better result than direct-leaching. Extraction of nick-el by indirect-leaching reaches 44.3%, far more than that of direct-leaching. Moreover extraction of nickel at pulp density of 5% is higher than that at pulp density of 10%, reaching 63% in 11 days and extraction of cobalt also is 48.1% in the same period. Compared with nickel and cobalt, iron gives low extraction in solution, containing organic acids produced by Aspergillus niger .

Jan 1, 2014

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 Jack S. McPartland

Historically, research on thiosulfate leaching of precious metals ores has been focused on processing of difficult to treat ores. Thiosulfate leaching of ores and concentrates containing sulfide minerals, copper, manganese and preg-robbing carbon constituents has been evaluated. Cyanide leaching is generally considered an efficient, environmentally acceptable means for processing of non-refractory precious metal ores, Consequently, commercial processing of finely disseminated, oxidized precious metal ores is normally accomplished using alkaline cyanide leaching. Recent changes in regulatory climate have, however, called into question the acceptability of cyanide leaching in a number of locations worldwide. In some cases where the use of cyanide is being questioned, thiosulfate leaching may provide an acceptable alternative. Developments in thiosulfate leaching of precious metal ores are reviewed in this paper.

Jan 1, 1999

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 S. Jolivet, A. Klöfverskjöld, P. Latostenmaa, B. P. Wilson, M. Lundström, P. Halli

Doré slag is secondary product of copper electrore fining anode slime treatment. The slag consists mainly of different slag formers, such as borax and silica based minerals together with soda and traces of sealing materials. However, Doré slag also contains metals present in the anode slime, such as Ag, Cu, Bi, Sb and Te. At some plants, slag cleaning has been conducted utilizing conventional grinding and flotation circuits in order to enrich the content of precious metals in the slag [1]. Nevertheless, the routine consideration of the slag as a secondary raw material has so far not been adopted on an industrial scale. The current study presents the preliminary results of the nitric acid and chloride leaching of antimony from Doré slag. Chemical and mineralogical analyses (SEM-EDS, AAS, XRD) were performed for the raw material, leaching residue, and leaching solution in order to ascertain the metal recovery and kinetics during leaching.

Mar 1, 2017

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 Jaroslav Briancin, Katarína Gáborová, María Florencia Márquez-Zavalía, Marcela Achimovičová, Ol’ga Šestinová, Pavel Škácha

Naumannite (Ag,Se), a silver selenide mineral, was first discovered in 1828. This study investigates the structure, morphology, chemical composition, and leaching kinetics of naumannite. Mechanical activation in a planetary ball mill increased its specific surface area, enhancing the leaching rate of silver in nitric acid (HNO3). The study found that mechanical activation significantly improves silver recovery.

Jul 1, 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 C. A. Born

AMAX Inc. climax's mine-mill complex at Climax, Colorado is best known for its molybdenite production: 50,000 TPY of molybdenite flotation concentrate (95% MoS2) from mill heads containing 0.35% molybdenite. Less recognized are the byproducts that are concentrated from the tails generated in the rougher flotation circuit (see Figure 1). Climax recovers a heavy mineral concentrate by passing the rougher tailings from molybdenite flotation over Humphrey spirals (Figure 2a). This concentrate is floated using xanthate at pH 4 to remove sulfides, principally pyrite. The tails from the pyrite float are then cleaned of quartz, feldspar and light minerals by tabling (Figure 2b). Amine flotation follows at pH 1.5 and 75°C to float monazite (Figure 2c). The present circuit at Climax dries the monazite flotation tailings and utilizes high intensity cross-belt magnetic separation Figure 2d) to generate a high grade wolframite concentrate and a low-grade cassiterite product, both of which are contaminated with the mineral brannerite, (UO, TiO, UO2)TiO3. The uranium specifications have recently become difficult to meet due to the increasing percentage of brannerite reporting in the mill feed. A plant is being built utilizing hot sulfuric acid and sodium chlorate to leach this uranium from the monazite flotation tailings prior to their being fed to the wolframite- cassiterite magnetic separators. This modification is the subject of our paper.

Jan 1, 1975