Search Documents

By C. A. Sosa-Blanco

Minera Sabinas in central Mexico treats a complex sulfide ore containing galena, chalcopyrite, sphalerite and argentite as valuable minerals, and in some cases important amounts of non-magnetic pyrrhotite. For high pyrrhotite ores, results obtained by selective flotation showed good grades and recoveries in the lead-copper circuit, but a commercial zinc concentrate could not be produced due to the high floatability of pyrrhotite. When this is the case, an alternative method was proposed to produce a bulk concentrate by flotation which is then leached to recover copper and zinc from liquor in a SX-EW circuit. Leaching was conducted in an autoclave in order to study the effect of temperature and pressure, but bioleaching tests were also conducted. Leaching lead-silver residues were treated by cyanidation in order to recover silver values. The results obtained by selective flotation and by bulk flotation with leaching of the bulk concentrate showed that the latter alternative increased zinc recovery by 30 % with similar recoveries of the other metals.

Jan 1, 2007

By Maurizio Marezio Bertini

Ammonia - ammonium sulfate solutions were sparged in a glass reactor with a SO2/O2 gas mixture (2% SO2 and 98% O2) to leach copper from pure chalcopyrite and chalcopyrite concentrate. The SO2/O2 mixture produces higher and faster extractions than O2 by itself. The effect of temperature, particle size and free ammonia concentration was investigated. The fastest and highest recoveries were obtained at 80 oC from a chalcopyrite concentrate with a d80equal to 13 m, with about 83 gpl free ammonia, an average pH equal to 8.75 and a stirring speed of 500 rpm. The extraction process was controlled by diffusion through a product layer, according to the shrinking core model, with an activation energy of 41 kJ/mole.

Jan 1, 2003

By R. B. Shu

A moderate thermophilic microorganisms, has been isolated (45?) from a typical sulphide mine in the region of Mo Jiang located in the southwest of China. The culture has been used to extract copper from chalcopyrite concentrates obtained from De Xing copper mine. The majority of experiments were carried out in 300mL Erlenmeyer flasks containing 150mL volumes of mineral suspension. Effects of some variable parameters such as solids concentration, pH and the concentrations of inorganic nutrients on the bioleaching of chalcopyrite obtained from De Xing copper mine were investigated. For a given concentrate the most important factor affecting the fractional copper extraction was the size of the mineral particles. The overall rate of leaching was found to be limited by the area of mineral surface. The attack started in cracks and genetic defects and finally reached the whole chalcopyrite surface. As a consequence of this, the ore showed strongly cracks. The additive of ferrous sulfate increases the bacteria activity, but did not improve chalcopyrite bioleaching. The results indicate that the nature and relative proportions of the various minerals present have been shown to be important factors.

Jan 1, 2014

By Eun-young Kim

Electrogenerated chlorine leaching of waste printed circuit boards was investigated in hydrochloric acid solutions. The pulverized printed circuit board contained about 37.6% of metal component, in which copper was about 83.6%. The leaching rate of copper was greatly affected by current density and agitation speed. The leaching of copper up to 98% was achieved at 20 mA/cm2, 50?, 180 minutes, and 600 rpm in 1M HCl solution. Increasing agitation and lowering current density enhanced utilization efficiency of electrogenerated chlorine. Leaching of copper was suppressed at initial stage, while the minor metal elements, such as aluminum, lead, and tin, were dominantly leached out.

Jan 1, 2006

By Xiong Tong, Jiushuai Deng, Yingbo Mao, Lingyun Huang

Ammonium carbamate has been used as a reagent for the leaching of copper from the copper oxide ore found in the Honghe state, Yunnan province, China. In order to find the optimal conditions for the leaching of copper from this copper oxide ore,the effects of several processing parameters, including the concentration, time, temperature, particle size and stirring intensity have been investigated. The results of these optimization experiments showed that the leaching rate of copper reached up to 72.76% under the following conditions: liquid-to-solid ratio of 10:1, ammonium carbamate concentration of 2.5mol/L, leaching temperature of 45 °C, stirring speed of 300rpm, a reaction time of 1.5 h and 85% by weight thereof passes through a sieve having 0.074 mm.

Jan 1, 2015

By G. Deschênes

In an earlier study (Deschênes et al., 2009), gold and silver were shown to be efficiently extracted from samples taken from the Kupol Mine in Russia using lead nitrate in a cyanide leach system. In the current study, a com-parison of the technical and economic merits of a new technology called CANMET enhanced leaching process (CELP) was made with the cyanidation/acidification-volatilization-recycling (AVR) process using eight metal-lurgical variability samples, as well as a large composite sample, which was to represent the initial two years of Kupol mill feed. The composite sample leach feed assayed 33.4 g/t Au and 289 g/t Ag. The metallurgical results from the large composite sample demonstrated that CELP produced very similar gold and slightly higher silver leach extractions compared to the conventional cyanidation/AVR leach conditions at 1/2 the cyanide concentra-tion (1,000 ppm NaCN in CELP vs. 2,000 ppm for cyanidation/AVR leach conditions). It was estimated that the total cost per ton was approximately the same for the two technologies (total cost of CELP with the use of Merrill-Crowe at $46.48/t, with cyanidation/AVR at $46.13/t). However, the implementation of CELP would eliminate a capital expenditure of approximately $5 million for an AVR system based on a 2005 feasibility level cost estimate. Eight variability samples used to confirm the composite sample results showed that, on average, the gold leach extractions with CELP and the conventional leach conditions were very similar, but the average silver leach extraction was 2.1% higher for CELP. Furthermore, CELP silver leach extraction was 4.6% higher on average compared to conventional cyanidation for the four mid-grade variability samples evaluated. The decision was made to remove AVR cyanide recovery from the flowsheet and implement CELP at Kupol. The ability to use the conventional leach equipment with CELP and simplify the flowsheet was also a factor, given the remote location of the Kupol Mine.

Jan 1, 2010

By F. Xie

Low gold and silver leaching kinetics has been commonly observed in traditional gold-silver cyanidation process, especially in those heap-leaching operations. The reactivity and availability of oxidant was believed to be the critical limitation for gold and silver dissolution. The occurrence of silver sulfide such as acanthite (Ag2S) in the ore may also result in low overall silver extraction. Leaching of gold and silver sulfide in the ferricyanide-cyanide solution has been investigated through rotating disc tests and the assistant effect of ferricyanide on increasing the dissolution rate of gold and silver was noticed. The potential use of ferricyanide as an auxiliary oxidant in gold/silver cyanidation process was proposed based on the results of leaching of gold-silver ores.

Jan 1, 2014

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 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 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 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 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 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

By P. Raffinot, C. H. G. Bushell, S. Hodgson, E. F. Milner, H. J. Bosshardt, N. Draper, D. E. Pickett, Edwin T. McKnight, J. Prieto

Zinc ranks third and lead fourth in world nonferrous metal conumption. In 1972 the free world consumed 3,333,000 mt of refined lead and 4,422,000 mt of refined zinc. The lead tonnage was divided into the following end uses: storage batteries (42%), lead allows (7%), pigments and chemicals (11%), cable sheathing (9%), gasoline additives (13%), pipes and sheets (8%), and miscellaneous uses (10%). The corresponding distribution of zinc uses in 1972 was galvanizing (39%), die casting (26%), brass (15%), rolled zinc (8%), oxide excluding American Process zinc made directly from concentrate (6%), and other uses (6%). The pattern of usage of lead and zinc varies considerably from country to country. During 1972 in the USA 49% of refined lead was consumed in storage batteries, 19% in gasoline additives, and only 3% in cable sheathing, whereas in the United Kingdom 28% of the lead was consumed in batteries, 14% in cable sheathing, and 14% in gasoline additives. For zinc the major US uses were die casting (41%) and galvanizing (37%), with about 3% consumed as rolled zinc; corresponding usages in the same applications in France were 12, 28 and 28% and in Australia 12, 67, and 3%, respectively. The outstanding resistance of lead to corrosion, its density, and

Jan 1, 1985

By T. W. Bingay

The production of both lead and zinc in Canada has steadily increased during recent years until today Canada ranks among the world's leading producers of both of these metals. The high prices which they have commanded have led to an energetic search for new deposits and also to an investigation of abandoned properties which, under the present conditions of better prices and more efficient methods of treatment, offer chances of commercial success. It is confidently expected that during the next few years success in both these lines of activity will further increase Canada's lead and zinc output, giving her a still higher ranking among the world's producers. In 1926, approximately 140,000 tons of metallic lead was produced from ores mined in Canada, and of this, 125,000 tons was refined in Canada at Tadanac, B.C. During the same year the production of metallic zinc from Canadian ores amounted to 68,000 tons, of which 62,000 tons was refined at Tadanac. History Of Lead and Zinc in Canada The production of lead in Canada is much older than that of zinc. The chief source has always been the interior of British Columbia. As earl y as the 'eighteen twenties', Indians and Hudson's Bay Company's trappers had discovered an outcrop of lead ore on the site of the present Blue Bell mine on the east shore of Kootenay lake, but it was not until1865 that active prospecting for lode deposits began. In the early nineties the activity of the silver-lead-zinc region of the Slocan reached its culmination and a steady production has been maintained down to the present time. Up to the end of 1924 the Slocan area produced over 38 million ounces of silver, nearly 300 million pounds of lead, and almost 122 million pounds of zinc.

Jan 1, 1927

By FREDERICK J. ALCOCK

THE high prices which lead and zinc have commanded during recent months have given a great impetus to search for workable deposits of these metals, and there has accordingly risen a demand for information about deposits which have been worked in the past and which might possibly be opened up anew under these more favorable conditions. Production of zinc and lead in Canada has been on the increase during the last few years. In 1924, 175,- 485,499 lb. of lead with a computed value of $14,221,- 345 were mined. Of this, 96 per cent. was produced in British Columbia, 3 per cent. in Ontario, while Quebec and the Mayo district of the Yukon accounted for the remaining one per cent. During the same year the production of zinc amounted to 98,909,077 lb.. valued at $6,274,791. Of this 96,000,069 lb. were produced in British Columbia and the remaining 2,909,008 in Quebec. The large producer of zinc and also of lead is the Sullivan Mine at Kimberley, B. C. In comparison with the Sullivan, the deposits which are known to occur in eastern Canada are relatively small. Two, however, the Kingdoll lead mine at Galetta, Ontario, and the Tetreault zinc property at Notre Dame des Anges, Quebec, are steady producers. The object of this paper is to describe briefly these and other deposits of eastern Canada, to show their relation to geological events and to point out what are the chances of any of those which are idle at present developing into active producers or what are the possibilities of new deposits being discovered.

Jan 1, 1926

By C. A. Summers

The U.S. Bureau of Mines applied process mineralogy to investigate a lead and zinc deposit located near Story Creek, Colville Mining District, Alaska. Mineralization is hosted by the Kayak Shale in veins, veinlets, and breccia. Sphalerite and galena, with supergene anglesite, occur in symmetrical quartz veins and between breccia clasts. The study sample analyzed 2.11 percent Pb, 5.37 percent Zn, 0.014 percent Cu, and 1.2 oz/ton Ag. Analyses of crushed products showed that 33 percent of the sample weight, and more than 50 percent of the Pb and 65 percent of the Zn, are in the minus 850 um fraction. Liberated grains of sphalerite and galena were found in crushed products finer than 300 um. Beneficiation tests indicate the material is amenable to processing by traditional gravity and flotation techniques.

Jan 1, 1991