Search Documents

By Gale D. C, Giles D. E, James Parker A

A zinc-lead-iron sulphide concentrate from McArthur River can be leached at atmospheric pressure in aqueous solutions with HCI, FeCl3, CuCI2, HCl/02 HC1/02/ CuCI2, Cl2, and CI2/CuCI2 to extract >98 per cent zinc, >90 per cent silver, and <15 per cent iron in 1-7 h at 95°. Galena is more reactive than sphalerite and is converted to lead chloride. Except for leaching with HCI, elemental sulphur is the the main by-product of the leach with little or no sulphate ion detected. The rates and products of these reactions are compared. It is recommended that aqueous slurries of the concentrate (50 per cent w/v) are leached at atmospheric pressure with Cl2 in the presence of copper(II) chloride, which accelerates the extraction of zinc. Hydrochloric acid is a by-product which is consumed bYJeaction with fresh concentrate in air at 60°. Copper is subsequently stripped from the leach solution and recycled by reaction with fresh concentrate at 100°. The separation, recovery, and electrolysis of zinc and lead chloride is outlined since this process would produce Cl2 gas which can be recycled and used directly in the leaching stage.

Jan 1, 1976

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 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

By S Moradi

The aim of this study was to investigate the possibility of leaching zinc from the tailing (-2 mm) of Angouran flotation plant using sulphuric acid. The tails were first calcined at 800¦C for one hour. The effect of factors such as stirring rate, temperature, initial acid concentration and solid/liquid ratio, on zinc extraction were studied. The results obtained show that at the best conditions, about 91 per cent of zinc were leached using a stirring rate of 800 rpm, a reaction temperature of 50¦C, a 30 minute reaction time, a 100 g/L initial sulphuric acid concentration and solid:liquid ratio 1:6 g/ml. The P85 of the feed material was about 150 ¦m and acid consumption was 600 kg/t.

Jan 1, 2002

By Kyoungkeun Yoo

A citric acid leaching test after heavy medium separation process was performed to remove the heavy metal contaminat such as Pb, Cu, and Zn from contaminated soil. The lead component was removed from soil by heavy medium separation whereas zinc component shows 0 % - 37.9 % removal ratio with particle size. The citric acid leaching was conducted to remove the remained zinc component. The removal ratio of Zn reached to more than 95 % under the conditions; 10 % in pulp density, 200 rpm in agitation speed, 50 °C, 2 mol/L in citric acid concentration at 2 hrs in leaching time. These results suggest that the heavy medium separation method followed by citric acid leaching is effective for treatment of soil contaminated with various chemical forms of heavy metal. Keywords: metal contaminated soil, citrate, heavy medium separation, zinc

Sep 1, 2012

By E. LaRocque

"Denison Mines Limited has been processing uranium ore in the Elliot Lake area since May of 1957. The ore body is located i11 the Huronion sediments on the north limb of the Quirke Lake syncline. Minerals of metallurgical significance are brannerite, uraninite and monazite. Gangue minerals include quartz, chlorite, sericite, pyrite, pyrrhotite and rutile. Silica content of the ore is greater than 75 % and pyrite-pyrrhotite content around 7 %. Specific gravity of the ore is 2.75.The Denison flowsheet is based on sulphuric acid leaching to extract the uranium values from the ore followed by liquid-solid separation to wash the uranium-bearing solution from the barren ore. The resultant pregnant solution is concentrated and purified by ion exchange followed by two-stage precipitation with ammonia to produce an ammonium yellow-cake.Elliot Lake ores require high acid concentrations and high temperatures to extract the uranium from brannerite and monazite. Uraninite will leach at low acid concentrations as long as oxiclizing conditions exist. The whole ore must be treated so that these conditions do exist.This paper provides a description of the leaching process equipment as originally installed and major changes in that equipment to date. Process control refinements to economically improve extraction are also discussed."

Jan 1, 1979

By C. Arthur Lefler

The hydrometallurgical processes for recovery of gold and silver in use at Round Mountain, Nevada and the practical aspects of the operation are discussed. Included are heap leaching by cyanidation, upgrading of solutions by adsorption of values on activated carbon and desorption (stripping) for electrowinning and subsequent refining steps to produce dore metal.

Jan 1, 1981

By C. Arthur Lefler

The Smoky Valley Mine is jointly owned by Louisiana Land and Exploration Company, Felmont Oil Corporation and Case Pomeroy Company. Copper Range Company, a subsidiary of L.L. & E. is the operator of the mine. The mine is located about 60 miles (100 km) north of Tonopah, Nevada on Highway 376 (formerly 8A) in the Big Smoky Valley on the western flank of the Toquima Mountain Range In the shadow of Mount Jefferson. The mine is at an elevation of about 6100 ft. (1860m) and is situated on the southwestern flank of Round Mountain, the hill for which the town of Round Mountain is named. The ore is a rhyolitic iglmbrite with coarse gold occurring along fractures and with fine microscopic gold disseminated in the rock mass. Minor silver is associated nth gold. Annual mining is planned at 5.8 million tons (5.3 million tonnes) per year with 2.2 million tons (2 million tonnes) as ore (7,000 tons or 6,400 tonnes per day).

Jan 1, 1981

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 Lavers G

Factors governing the dissolution of nickel and iron in pentlandite and pyrrhotite concentrates from Mt. Windarra were studied for aqueous solutions at normal temperatures and pressures. The effects of dissolved dioxygen, carbon dioxide, hydrogen ions, and chloride ions are described. The direct reaction of aerated water wiili pyrrhotite is much faster than with pentlandite, and reaction rates are proportional to the dissolved dioxygen concentration. In both cases divalent metal sulphates are produced initially. The subsequent oxidation of iron(II) to iron(III) and its oxolation produces acid solutions which rapidly attack both minerals. The presence of carbon dioxide or chloride ion has little effect on these rates.

Jan 1, 1976

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 H. W. Morse

THE CHAIRMAN (H. W. MORSE).-Gentlemen, for the first time in the history of the American Institute of Mining Engineers, we have a full session on the subject of leaching-especially on the leaching of copper ores.. This branch of metallurgy is rather in its infancy. We have had one big plant running for a year or more, and pretty successfully; and another one which will be started in February or March, the New Cornelia. We have another one at Garfield for the Utah company, which is just designed and construction begun; and from several other points come reports of interesting test work. The first paper on the program this evening is one of some of the test work on the New Cornelia ore by myself and Mr. H. A. Tobelmann. LAWRENCE ADDICKS, New York, N. Y. (written discussion).-The Ajo proposition has had, as I understand it, two great advantages: it has not been obliged to measure leaching against flotation or some other metallurgical process, and it deals with a copper content high enough

Jan 12, 1916

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

By F. Letowski

A column reactor was applied for leaching with simultaneous flotation a complex sulfide ore in ferric sulfate-solution under oxygen pressure. Elemental sulfur and the flotable, nonreacted sulfides were collected in the froth and separated from the insoluble, nonflotable gangue minerals and solid particles formed during pressure leaching. The two-stage laboratory unit, built from Corning-glass elements is described. Results of the continuous leaching/flotation experiments conducted under 20-40 psig of oxygen pressure are presented, as are supplementary batch leaching experiments with iron precipitation under 150 psig oxygen pressure. High zinc extraction to solution and near total accumulation of lead sulfate in the nonflotable fraction are easily attainable. Relatively poor flotation of pyrite is probably caused by the partial oxidation of its sulfidic sulfur to sulfate.

Jan 1, 1988

By Franciszek Letowski

"Recovery of precious metals associated with a sulphidic matrix has an important economic effect on complex sulphide ore treatment. Because a significant portion of the silver and gold associated with pyrite are rejected to tailings during the selective flotation of pyritic Zn-Pb-Cu ore, other ore processing routes have been investigated.Direct leaching of the ore with ferric sulphate solution followed by separation of different components of the solid suspension by flotation was applied. Such a process allows for: production of zinc sulphate solution; transformation of galena PbS to insoluble lead sulphate PbS04 and its accumulation in a non-floatable fraction together with precipitated geothite FeOOH and other insoluble gangue minerals; separation of sulphur and non-decomposed sulphides in the floatable fraction.Laboratory research carried out on the ore samples from an Alaskan pyritic deposit from the Delta region indicates that gold and silver are mostly concentrated in the floatable fraction.Total recovery of zinc from the ore attains 96.5% to 98.5% after ferric sulphate leaching. About 85% of copper was extracted from the ore during the ferric sulphate and subsequent ferric chloride leaching, and about 6% of the copper remained in the non-floatable fraction and mixed fraction during the leaching-flotation step. Extraction of lead from the nonfloatable fraction during leaching with calcium chloride solution attained 98% to 99%. Extraction of silver from the floatable fraction during leaching with ferric chloride solution reached 98%. Extraction of gold from the floatable fraction during cyanide or thiourea leaching reached 71% to 89%. The results of the laboratory research are presented."

Jan 1, 1987

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