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Experimental process analysis and mathematical modeling for selective gold leaching from slag through wet chlorination
Abstract
The aim of this paper consists in defining optimal conditions of controlled and selective gold leaching contained in copper - gold slag, by wet chlorination. Studies were conducted on three types of copper - gold slag, resulting as by-products from pyrometallurgical processing of non-ferrous raw materials. Due to high levels of copper in slag, an oxidative pre-treatment is necessary to separate gold from accompanying elements and also for copper recovery. After nitric acid pre-treatment, nearly 87% Cu were readily extracted, and lead and silver content in the all three types of solid samples has been reduced by 91.34% and 96.29% respectively. Selective leaching of the solid phase resulting after HNO3 pre-treatment was done by wet chlorination, using nascent chlorine (Cl2) as leaching agent, in presence of hydrochloric acid (HCl). The chlorine was obtained «in situ» by the reaction of sodium hypochlorite (NaOCl, 5% Cl2) and HCl. In this study, the effect of nitric acid pre-treatment, redox potential (Eh) given by sodium hypochlorite (NaOCl) consumption, hydrochloric acid concentration, temperature and leaching time have been studied. The potential required for a rapid leaching rate and for avoiding the reduction of gold is 950 mV (vs AgCl/Ag(KCl sat)).The 5% NaOCl consumption required for gold leaching from pre-treated samples was lower than NaOCl consumption necessary for original samples leaching. The highest gold extraction yield (98%) was obtained after 6 hours at ambient temperature, solid/liquid ratio (w/v) 1:2 and 4M HCl. In order to explain experimental results a mathematical model of the simultaneous gold, copper, lead and silver dissolving has been developed. It considers at particle level, the competition between surface reaction and the diffusion of the active species by a gel coating that covers the dissolving particle.
... Therefore, it is predictable, raising Clconcentration increases the affinity of AuCl − 4 dissolution. Olteanu et al. [33] also reported a similar result for the gold dissolution rate so that the hydrochloric acid concentrations lower than 2.74 M had an effective kinetic dissolution model. gram that enhances the Au leaching rate [34]. ...
... Based on Eq.(1), changing the pulp density changes the reactant (Cl 2 and Au) ratio, influencing the reaction rate. On the other hand, increasing pulp density affects the fluid flow through the solid particles, and mass transfer of lixiviate around the particle was decreased [33]. However, after 120 min, Au recovery was equaled with 50 and 80 g/L, approximately. ...
This study showed the kinetics of gold chlorination leaching from an anode slime complex containing copper, silver, barium, selenium, and tellurium. AgCl and BaSO4 control the rate of gold leaching, and the interface transfer and diffusion across the product was selected as the desirable kinetic model. The concentration of hydrochloric acid, chlorine gas flow rate, pulp density, and temperature have been precisely studied. The results showed that temperature and hydrochloric acid have a favorable influence on the leaching rate. However, the chlorine gas flow rate is inactive in the reaction rate. Furthermore, the leaching rate decreased with an increasing pulp density. The leaching process was controlled by the interface transfer and diffusion across the product, and the activation energy was determined to be 13.66 kJ mol–1.
... Leaching of silver with sodium hypochlorite can be demonstrated as given in the following reactions ((Eq. (4); Eq. (5)) [26,30]. ...
The mobility of gold and silver metals from refractory ores is very low due to free metal ions in their bodies that bind the transportable and free metal molecules or they are encapsulated in minerals such as sulfides, arsenosulfides, carbonaceous, and clayey materials which make it difficult to extract them. As known from the literature, the enrichment of oxide refractory gold, silver ores, Ag, and Au-rich flotation concentrated from sulfur refractory ores can be leached in the chloride-hypochlorite solutions. In this work, the leaching approach was studied to extract silver ions from refractory ores in the presence of hypochlorite media under high pressure. The mineralogical and scanning electron microscopy analysis of the sample indicated that the silver was encapsulated in barite and quartz minerals in fine sizes. In order to obtain the optimum leaching conditions for the silver ore, the effects of several parameters, namely, temperature, leaching time, NaCl and NaOCl concentration, solid ratio, and particle size on the leaching recovery were investigated in detail. The leaching experiments showed that the highest silver recovery of 62.35% was obtained under the following conditions: temperature of 150 °C, oxygen pressure of 1800 kPa, leaching time of 2 h, concentrations of 2 mol/dm3 and 0.4 mol/dm3 for NaCl and NaOCl, respectively, the solid ratio of 30%, and particle size (d80) 15 μm. The results obtained from this study indicated that the enrichment of refractory silver ores to extract the silver encapsulated in barite and quartz minerals could be possible in the chlorite-hypochlorite under high pressure, a relatively shorter leaching time compared to conventional cyanide leaching.
... Zhao et al. [8][9][10] proposed a separation technique using the different chalcophile properties between W and Mo, which had previously been revealed through research on the formation of molybdenite and tungsten oxide minerals. As another example, chlorination metallurgy [11,12] is an important method that extracts gold (Au) from ores by the coordination of Au 3+ and Cl − in [AuCl 4 ] − , which is a formation mechanism of gold deposits on earth [13][14][15]. Similarly, Au, tellurium (Te), bismuth (Bi), and arsenic (As) are highly chalcophile elements [16], and these elements follow the extraction process of copper [17]. ...
In 2003, Railsback proposed the Earth Scientist’s Periodic Table, which displays a great deal of elemental geology information in accordance with the natural environment of the earth. As an applied science, metallurgy is based on mineral composition and element behavior, that is similar to geochemistry. In this paper, connections and similarities between geology and metallurgy are identified, based on geochemical laws and numerous metallurgical cases. An obvious connection is that simple cations with high and low ionic potential are commonly extracted by hydrometallurgy, while those with intermediate ionic potential are extracted by pyrometallurgy. In addition, element affinity in geology is associated with element migration in metallurgic phases. To be specific, in pyrometallurgy, lithophile elements tend to gather in slags, chalcophile elements prefer the matte phase, siderophile elements are easily absorbed into metal melt, and atmophile elements readily enter the gas phase. Furthermore, in hydrometallurgy, the principles of hard/soft acids and bases (HSAB) offer an explanation of how precipitation and dissolution occur in different solutions, especially for fluoride and chloride. This article provides many metallurgical examples based on the principles of geochemistry to verify these similarities and connections.
... Currently, dry chlorination and wet chlorination are most used techniques for conversion of metals into metal chlorides (Cui et al., 2018a). In dry chlorination agents like NaCl, CaCl 2 , KCl, and NH 4 Cl are used for treatment while Cl 2 and HCl are being used in wet chlorination (Xia et al., 2020;Olteanu et al., 2014;Cui et al., 2018b). Wet chlorinating agents are highly corrosive in nature and are very difficult to handle with respect to dry chlorinating agent. ...
The substantial growth of electronic waste (e-waste) in recent years has become a serious threat to environment. However, there is an excellent opportunity to recover and reuse metals present in e-waste, which eventually leads to conservation of natural resources for future generation. A greener and sustainable approach for the recovery of metals from electronic waste is the need of the hour. In this study, thermal decomposition of printed circuit boards (PCBs) was carried out in presence of nitrogen for conversion of polymers into oil and combustible gases. The metal rich pyrolysis residue was roasted in presence of ammonia chloride as chlorinating agent to recover metals. The effect of roasting parameters on the metal recovery investigated in temperature range of 200 °C to 325 °C for 1 h to 5 h while the NH4Cl dosage varied from 1 g/g to 4 g/g. Under the optimized roasting conditions, around 93% Cu, 100% Ni, 100% Zn, and 100% Pb were recovered at temperature of 300 °C, time of 4 h and NH4Cl dose of 3 g/g. The present process provides an eco-friendly solution for the recovery of metals from e-waste, which are valuable and avoid pollution.
... Penelitian pengolahan emas/perak dari lumpur anoda cukup banyak yang telah dipublikasikan dengan menggunakan beberapa metode; ada yang menggunakan kombinasi teknologi pirometalurgi dan hidrometalurgi yang disebut dengan proses hybrid (Chen dkk.,2015) yaitu melalui rute proses-proses roasting oksidasi, pelarutan air (water leaching), peleburan reduksi, pemurnian oksidasi dan elektrolisa. Filcenco Olteanu et al. (2014) telah melakukan pelindian emas secara selektif dari terak peleburan tembaga melalui tahapan awal pelindian Cu, Ag, Pb menggunakan HNO3, kemudian residunya di proses dengan cara klorinasi menggunakan pelarut HCl+NaOCl yang mampu mengekstrak emas hingga 98%. Penelitian lainnya seperti yang dilakukan oleh Baba dkk. ...
Percobaan mengekstraksi emas dilakukan secara selektif dari residu pelindian timbal dalam lumpur anoda melalui proses klorinasi basah, diikuti proses ekstraksi pelarut menggunakan ekstraktan pelarut organik dibutyl carbitol (DBC). Hasilnya menunjukkan bahwa pemberaian emas oleh pelindi HCl atau gas Cl2 yang dibantu oksidator H2O2 ternyata sangat efektif, bahkan menjadi lebih efektif jika menggunakan peralatan pelindian yang kedap gas. Berikutnya, proses ekstraksi pelarut terhadap larutan mengandung emas hasil klorinasi menunjukkan terbentuknya larutan organo-emas (loaded DBC) meningkat dengan meningkatnya konsentrasi DBC hingga 40% (v/v) dengan lama waktu pengadukan yang singkat, persen ekstraksi emas yang diperoleh adalah 99,31%. Akhirnya, loaded DBC direduksi untuk menghasilkan logam emas.
... ليچينگ روش از استفاده به میاالت در بيشداري مطالعات اگرچه سنگ از هيوميکي مواد استخراج براي قليايي معدن لئون هاي ارديت کرده تاکيد روش از که دارند وجود هدچنين مطالعاتي ،اند هاي جسته بهره فيزيکي بروي را مطالعاتي ،ترکيه کشور از محییاني .اند روش با لئونارديت از غيرارگانيک مواد استخراجDebora et al., 2002;Tarhan, 2011روش و )Nan et al., 2005, Olteanu et al., 2014, Bas et al., 2015Abbruzzese et al., 1995. ...
زغال¬سنگ¬های لئوناردیت، نوعی زغال¬سنگ با درجه اکسیداسیون بالا هستند که در فرآیند کربنی¬شدن لیگنایت¬های حاوی مواد هیومیکی تشکیل می¬شوند. مواد هیومیکی، یک ترکیب ارگانیک حاوی فولویک اسید، هیومیک اسید و هیومین هستند که اکثرا به¬عنوان اصلاح¬کننده¬های خاک استفاده می¬شوند. استخراج مواد هیومیکی از زغال¬سنگ¬های لئوناردیت حاوی هیومیک اسید بیش از %50، یک موضوع برجسته مطالعاتی در سال¬های اخیر شده است. درحالی¬که استخراج مواد هیومیکی از لئوناردیت¬ها عمدتا با استفاده از تکنیک¬های حلالیت شیمیایی (فرآیند لیچینگ) در یک محیط قلیایی انجام می¬گیرند، روش¬های غنی¬سازی فیزیکی همچنین در مطالعات محدودی بکار گرفته شده¬اند. حذف ترکیبات غیرارگانیک موجود در لئوناردیت، می¬تواند مصرف واکنش حل¬کننده¬ها را کاهش و از بکارگیری تجهیزات غیرلازم ممانعت بعمل بیاورد. بدین منظور، در این پژوهش، اثر پیش-غنی-سازی فیزیکی به¬عنوان پیش¬تیمار بروی لیچینگ مواد هیومیکی مورد بررسی قرار گرفت. این تحقیق در دو مرحله انجام شد. در اولین مرحله، پارامترهای عملیاتی مانند مقدار حلال، زمان هم¬زنی، دما و نسبت جامد برای فرآیند لیچینگ مطالعه شد و شرایط بهینه فرآیند لیچینگ تعیین گردید. در مرحله دوم، اثر آزمایش-های پیش-غنی¬سازی فیزیکی بروی فرآیند لیچینگ مورد مطالعه قرار گرفت. یافته¬های این پژوهش نشان داد، درحالی¬که یک محصول حاوی %2/48 مواد هیومیکی با بازده استخراج %63/87 از طریق آزمون¬های غنی¬سازی شیمیایی بدست می¬آید، یک محصول حاوی %74/62 مواد هیومیکی (کنسانتره) با بازده استخراج %4/92 بعد از پیش-غنی¬سازی فیزیکی در ترکیب با فرآیندهای لیچینگ حاصل می¬شود.
... Fig. 7. Comparing by means of X-ray diffraction spectra humic substances obtained using different amounts of potassium hydroxide It can be observed from Fig. 6b that humic substance content carried to the solution increased along with the increase in the stirring time in the optimal stirring time experiments. These results explained that the possibility of chemical reactions providing dissolution increased as the contact time increased (Abbruzzese et al., 1995;Nan et al., 2005, Olteanu et al., 2014, Bas et al., 2015. However, the 24-h stirring time that produced best results was considered as too long in terms of industrial application. ...
Background: Leonardite mineraloid is known as a highly oxidized ore formed during carbonisation process of lignite containing humic substance. Humic substance is an organic material containing fulvic acid, humic acid, and humin, and that is mostly used as soil conditioner. Extraction of humic substances from leonardite ores that contain high amounts of humic acid (>50%) has become a prominent area of study in recent years. While humic substance extraction from leonardite is generally carried out by chemical dissolution technique (leaching) in alkali medium, physical enrichment methods were also used in limited number of studies. However, removing inorganics found in leonardites would decrease dissolving reactive consumption and would also prevent unnecessary capacity use. This study investigates the effect of physical pre-enrichment processes on humic substance leaching. This research was carried out in two stages. In the first stage, operational parameters such as amount of reactive, stirring time, temperature, and solid ratio for the leach process were studied, and the most suitable leach conditions were then determined. In the second stage, the effect of physical pre-enrichment experiments on leach process was investigated. While a product containing 48.2% humic substance was obtained with 87.63% humic substance extraction efficiency through chemical enrichment experiments, a product containing 62.74% humic substance (concentrate) was obtained with 92.4% humic substance extraction efficiency after pre-enrichment combined with leach processes.
... Therefore, it is generally combined with different acidic reagents. The chlorination is capable for higher scale processing and has already been applied in industrially significant scale [4], [29]- [31]. ...
The importance of gold recovery from low grade ores and flotation tailings is continuously increasing due to raising gold demand and thereby high gold prices. However, due to raising social awareness of the ecological impacts of cyanidic gold extraction and environ-mental specifications, the development and the implementation of alternative hydrometallurgical extraction processes have been a focus for many research institutions in last decades. Present work gives a comparison between compatible extraction reagents, with focus on less harmful processes. The target of this review is to point out the best cyanide-free processes of following methods and reagents: Bioleaching, Chlorination, Aqua Regia, Bromine, Thiocyanate, Thiosulfate and Thiourea leaching. For this propose, the gold leaching reagents are described and discussed in terms of their environmental and economical points of view. As result of this comparison, thiourea stands out as the most promising alternative gold leaching reagent to cyanide.
... Biological oxidation for this purpose is cheap and environmentally friendly, but it is inefficient and too sensitive to operating conditions (Kaksonen et al., 2014;Yang et al., 2013). Although chemical oxidation can easily corrode equipment and contaminate the environment due to the heavy use of strong oxidative reagents such as nitric acid, and chlorine gas, its performance for the liberation of encapsulated gold in some cases is satisfactory (Marsden and House, 2006;Olteanu et al., 2014;Van Weert et al., 1986;Xu, 2013). For example, permanganate in acid media can effectively liberate the encapsulated gold in ore or cyanidation tailings (Tang et al., 2013;Zhai et al., 2010). ...
A theoretical substantiation of an environmentally friendly technology for extracting gold from mineral raw materials is presented. In terms of an alternative to conventional cyanide, the authors propose chlorine-bearing derivatives of organic isocyanuric acid (C 3 H 3 N 3 O 3 ). In the studied process, this reagent combines the roles of a long-acting oxidant and complexing agent due to the resulting Cl – ion. For experimental purposes, the rotating disc technique was applied. Experimental results were evaluated by measuring the concentration of gold in solutions using the atomic adsorption method. The theoretical characteristics of gold dissolution using the proposed reagent were studied. In order to evaluate the applicability of the revealed regularities for practical purposes, comparative indicators for leaching gold from the ore of the Byn’govskoye deposit (Sverdlovsk Oblast) using the proposed reagent, cyanide and CYCAD cyanide substitute (China) were evaluated. The dependence of the gold disc dissolution rate on temperature, as well as leaching reagent and hydrochloric acid concentrations, were studied. In disc sample experiments, the exponential dependence of the dissolution rate on the concentration of the studied reagent was established. In this case, the maximum intensity of the process is achieved at a concentration higher than 50 g/dm ³ . The optimum acidity level comprises 0.3–0.4 g ion/dm ³ of hydrochloric acid. Despite an increase in temperature being shown to accelerate the dissolution process, this appears to be technologically unjustified. Under optimal conditions, the gold disc dissolution rate reaches 0.5•10 ⁻³ g ion/(cm ² •min), which is approximately one hundred times higher than during the dissolution by the cyanic solutions in conventional modes. Close gold extraction values were achieved in the solutions of cyanides and chlorine-bearing derivatives of isocyanuric acid during the leaching of gold from the ore under the comparable conditions. For CYCAD solutions, the same indicator is two times lower. The results of the performed studies demonstrate the high dissolution rate of gold using the proposed reagent along with the technological possibility of leaching gold from ores using this environmentally friendly reagent.
More than half of the world’s tin production is from secondary ore deposits. However, along with its depletion, PT Timah Tbk has explored and mined the primary deposits in Belitung, one that is expected to become tin’s new source. Tin was found liberated as cassiterite in the secondary deposits, while in the primary, it was associated with other minerals requiring more complex processing. Therefore, a suitable technology is needed to extract the tin from the primary ore deposits. This research aims to observe the extraction of tin from primary deposits (oxide and skarn ores) using HCl wet chlorination, observed at different temperature (60 – 90°C), HCl concentrations (10 – 50 % excess of stoichiometric amount), leaching time (60 – 300 minutes), and with or without the addition of oxidant (hydrogen peroxide, H 2 O 2 ). The research found that the highest percentage of tin extraction from oxide and skarn ores was 92.32% and 97.28%, respectively. Both results were achieved at the same optimum condition: temperature of 60°C, 50% excess of HCl, 240 minutes of leaching time, and at two-fold excess of the stoichiometric amount of H 2 O 2 . It was concluded that oxidative wet chlorination could be used to recover tin from its primary ore deposits.
This work proposes an efficient and simple hydrometallurgical process based on a chlorination step followed by an ion-exchange step for recycling gold (Au) from a waste printed circuit boards (WPCBs) enriched in Au resulting from a first leaching step under mild oxidizing conditions for extracting Cu and other base metals. Under optimized [3.5 mol/L HCl and 0.46 mol/L NaClO, with a liquid/solid (L/S) ratio of 40, at 40 °C for 3 h with agitation] leaching conditions, 95% Au was extracted from the residue originating a multi-metal solution containing 1.0% Au. Subsequently, Au (initial concentration: 38 µmol/L) present in the multimetal-leached solution was purified in continuous mode using two strong anionic exchange resins: DOW™ XZ-91419.00 and Purogold™ A194. Both resins were suitable in purifying Au from the multimetal-leaching solution, with at least 70% of Au recovered relative to the initial residue. When the DOW™ XZ-91419.00 resin was used, a solution containing 1.7 mmol/L Au with a purity grade of 94% was obtained, with Pb and Sn being the major contaminants (3.3 and 2.4%, respectively). For Purogold™ A194 resin, a solution containing 0.73 mmol/L Au with a purity grade of 92% was achieved; Ag, Pb and Pd were the major contaminants (1.4, 3.6 and 1.8%, respectively). In conclusion, this work demonstrates a novel hydrometallurgical strategy for recycling Au with a high grade from WPCBs, minimizing the number of leaching and purification steps and the amount of waste created.
Metallurgical plants constituting of smelters and refineries recover metals (i.e., copper) from mineral deposits. Copper production generates several waste streams of which slag, sludge and dust are generated in the largest quantities. The need to eliminate or at least reduce their adverse effects on the environment call for developing methods for recovering valuable components such as copper, zinc and iron through their selective separation from toxic components present in the waste (mainly arsenic and lead). This can be achieved through hydrometallurgical methods (leaching with organic and inorganic media), techniques facilitating mobility of elements (roasting with leaching) and biological processes (bioleaching). The valorization of metallurgical waste as a source of fertilizer micronutrients can be a sustainable and value-added direction of its management. This review presents ways of useful-metals recovery from the copper smelter and refinery wastes, including selective separation of valuable metals. The novelty of this review is a demonstration of the application potential of recovered components from metallurgical waste in the agricultural sector.
A novel process of chlorinating roasting followed by chlorinating leaching to extract platinum and gold from copper anode slimes was proposed in this research. Results of thermodynamic analysis and experimental research showed that the platinum was chlorinated into PtCl2 while the gold existed in the form of metallic Au during the roasting process. With the copper anode slime being directly leached using a traditional process, the Pt recovery rate was low and came to 80.72%. After the roasting process with sodium chloride and concentrated sulfuric acid in oxygen atmosphere, the recovery rate of Pt increased to a value around 95%. Moreover, with excessive addition of concentrated sulfuric acid, more H2O (g) was generated and the formation of Cl2 (g) decreased due to the transition from HCl (g) and Cl2 (g), as a result of which the Pt recovery rate decreased. In addition, this chlorinating roasting had little effect on the Au recovery due to its difficulty to be chlorinated.
The microbial fuel cell, which can convert the chemical energy of organic matter into electricity via the catalytic action of microorganisms, is a novel environmentally friendly technology for wastewater treatment and energy generation. The electrical energy generated by the microbial fuel cell can be used as an alternative to a traditional external power source required to extract copper via electrolytic treatment. A dual-chamber microbial fuel cell (DMFC) for the treatment of copper slag sulfuric acid leach liquor was constructed. The electrogenesis performance of the DMFC and its ability to extract copper from the copper slag leachate were investigated. The results demonstrated that the maximum voltage was 540 mV when the DMFC achieved steady-state operation. The removal rate of copper ions was greater than 80.0%, and the maximum value was 92.1%. Moreover, X-ray diffraction and scanning electron microscopy were used to characterize the cathodal products. The results showed that the product deposited onto the cathode was copper and that its morphology was similar to that of the electrolytic copper powder. The DMFC can generate electricity and recover copper from copper slag simultaneously. © 2017, University of Science and Technology Beijing and Springer-Verlag Berlin Heidelberg.
The rapid growth in the use of electronic equipments, combined with early obsolescence has contributed enormously to the generation of large quantity of electronic (e) waste. One such e-waste, the mobile phone printed circuit boards (PCBs) contain various precious metals which can be extracted by different hydrometallurgical routes. The present work deals with the recovery of gold using ammonium thiosulfate as a leaching agent from waste mobile PCBs containing 0.021% Au, 0.1% Ag, 56.68% Cu, 1.61% Ca, 1.42% Al, 1.40% Sn, 0.24% Fe, 0.22% Zn, 0.01% Pd etc.. The cutting granules of 0.5-3.0 mm PCBs were used for leaching in a 500 mL glass beaker in open atmosphere. The effect of various parameters viz. ammonium thiosulfate concentration, copper sulfate concentration, pH and pulp density was studied. A leaching of 56.7% gold was obtained under the optimum condition of 0.1M ammonium thiosulfate, 40 mM copper sulfate, pH: 10-10.5, pulp density: 10 g/L at room temperature and stirring speed of 250 rpm in 8h duration. The maximum leaching of gold in the pH range 10-10.5 may be attributed to the higher stability of the ammonium thiosulfate. The decomposition of ammonium thiosulfate in the different pH ranges was chemically analysed by iodometric method. The ammonium thiosulfate contents in the leach liquors were in agreement with the quantity of gold leached in the respective pH ranges. In this process the copper sulfate worked as a catalyst. The experiment conducted with complete PCBs scrap exhibited a maximum leaching of 78.8% gold at the above optimised condition.
The origin of a rational (scientific) approach to extraction of metal values from ores with the aid of microorganisms (bioleaching) is traced. The removal by microbiological means of ore constituents that interfere with metal extraction (biobeneficiation), an outgrowth from bioleaching, is also traced. © 2004 SDU. All rights reserved.
A series of processes has been developed to recover the gold from electronic scrap containing about 200~600 ppm Au. First, mechanical beneficiation including shredding, crushing and screening was employed. Results showed that 99 percent of gold component leaves in the fraction of under 1mm of crushed scrap and its concentration was enriched to about 800 ppm without incineration. The crushed scrap was leached in 50% aqua regia solution and gold was completely dissolved at withing 2 hours. Other valuable metals such as silver, copper, nickel and iron were also dissolved. The resulting solution was boiled to remove nitrous compounds in the leachate. Finally, a newly designed electrolyzer was tested to recover the gold metal. More than 99% of gold and silver were recovered within an hour by electrowinning process.
The more restrictive regulations for controlling the environmental pollution by cyanide bearing streams and materials have given an impetus to the researches aiming the discovery of some new, more ecological methods for gold leaching. This paper present an alternative lixiviant system for the recovery of gold from secondary sources such as gravitational concentrates from alluvial sands. The gold leaching using nascent chlorine resulted from NaClO decomposition in hydrochloric acid medium was experimented. Also, the influence of the main parameters: solid: liquid ratio, temperature, chlorine and chloride ion concentration; redox potential are presented. A mathematical model for lixiviation time computing was proposed, using as parameters the initial radius of gold particle ( Rin) and molar fraction of chlorine in solution ( yCl) :
Cyanide has been recognized for a long time as a powerful lixiviant for gold and silver, forming very stable cyano complexes with both metals. While cyanide is very effective in leaching free milling ores, there are certain classes of gold and silver ores (i.e., carbonaceous, pyritic. arsenical, manganiferous, cuperferous) that are considered refractory to conventional cyanidation dissolution. Recently there has been considerable effort directed towards new and improved reagents for leaching these difficult-to-treat ores and concentrates. A large portion of this effort has been devoted to finding alternative lixiviants that might compete with conventional cyanidation. Furthermore, there is a general interest in developing non-toxic environmentally safe substitutes for cyanide.There are a number of reagents that form stable complexes with gold and silver e.g., thiourea, thiosulfate, halides, malononitrile, acetonitrile and polysulfides. The chemistry of gold and silver dissolution using alternative lixiviants is discussed in this paper. Special emphasis is given to the application of Eh-pH diagrams to interpret the dissolution behavior.
A process for the recovery of gold from processed gold-plated printed circuit boards, gold-coated bangles and gold-coated mirrors is described. It is based on heating the scrap material along with an “eco-friendly” or “green” reagent, 20% (v/v) formic acid solution to boiling temperature to separate the inorganic component from epoxy resin in gold-coated mirrors. Dissolution of base metals is affected using a strong oxidizing agent, 20%(w/v) potassium persulphate, which is also “eco-friendly” or “green”; and heating to boiling temperature. Recovery of gold is achieved by melting. The recovery efficiency is on par with the cyanidation process. The process can be scaled up and adopted by all categories of industries.
The application of macroporous strong base resins to cyanide recovery from synthetic solutions containing ferrous cyanide, nickel cyanide and high levels of copper cyanide is discussed. Adsorption experiments show that strong base resins readily extract metal cyanocomplexes, in a sequence of affinity given by Ni>Cu>Fe, for polystyrene resins, and by Cu>Ni>Fe, with polyacrylic resins. The presence of ethyl ammonium quaternary groups results in a remarkable low iron uptake. Acid solutions are able to desorb WAD cyanides but the precipitation of Prussian blue-like compounds in the presence of ferrocyanide ion poisons the resin. This work shows that iron cyanocomplexes may be removed with concentrated sodium chloride or sodium nitrate solutions. The role of these eluants on the desorption of copper and nickel cyanocomplexes is also discussed.
The intensification of leaching of gold from a waste using mechanical activation (milling in water) and mechanochemical activation(milling in thiourea solution) were studied as the pretreatment steps. The leaching of “as-received“ sample in an acid thiourea solutionresulted in 78 % Au dissolution, after mechanical activation 98 % and mechanochemical activation up to 99 % of the gold was leachedduring 120 min. The mechanochemical activation resulted in an increase of the specific surface area of the waste from 0.6 m2g-1to a maximum value of 20.5 m2g-1. The activation was performed in an attritor using variable milling times. The physico-chemical changesin the waste as a consequence of mechanochemical activation had a pronounced influence on the subsequent gold extraction.
The more restrictive regulations for controlling the environmental pollution by cyanide bearing streams and materials have given an impetus to the researches aiming the discovery of some new, more ecological methods for gold leaching. This paper presents an alternative lixiviant system for the recovery of gold from secondary sources such as gravitational concentrates from alluvial sands. The gold leaching using nascent chlorine resulted from NaClO decomposition in hydrochloric acid medium was experimented. Also, the influence of the main parameters: solid: liquid ratio, temperature, chlorine and chloride ion concentration; redox potential is presented. A mathematical model for lixiviation time computing was proposed, using as parameters the initial radius of gold particle (R(in)) and molar fraction of chlorine in solution (y(Cl2 center dot)): tau(s) = rho(Au)/M(Au) R(in)(2)/D(Cl2) CT(ln(1-y(Cl2 center dot))
Mintek has developed a process called iGoli Mercury-free Gold Extraction Process that can be used to extract gold from 0.1 % gold concentrate to produce 99.90 percent gold product. The process uses a mixture of pool acid (dilute hydrochloric acid), bleach (sodium hypochlorite) and sodium metabisulphate to leach and recover gold. Artisanal and small-scale miners concentrate their gold using sluice boxes followed by panning as a final physical separating method. The gold produced in this manner is extracted from the concentrate by a chlorine solution produced from a mixture of pool acid and bleach. The leach liquor contains gold and other metals that are soluble in chlorine media such as base metals and ferrous iron. During leaching excess gas that is produced is redirected to a separate vessel where it is scrubbed using caustic soda to form water and sodium chloride (salt). The slurry from the leaching process is filtered to separate the gold bearing leach liquor and the solid residue. This gold pregnant solution is treated with sulphur dioxide, introduced in the form of sodium metabisulphate, to reduce gold ions in solution to a metallic gold powder. The solid waste produced from iGoli process is rich in silicate with effluent and is neutralized using lime/ limestone with an addition of apatite where necessary to destroy or precipitate all base metals and ferrometals available. The development of this process was mainly driven by the need to ensure reduction in mercury use and subsequent alleviation on the entire biophysical environment. There is an array of mitigating measures that could be used as treatment methods for the safe use of iGoli leaching process. The process wastes which are significantly less toxic are characterized as follows: → Waste solids Feed → → Waste solution (effluent) → Waste heat → Waste gases Any potentially hazardous gases released from the gold leaching process are monitored and controlled by neutralization below the legislative requirements. The tailings produced have a high percentage of silica and iron with low levels of base metals. The process is very important to artisanal and small-scale miners as it lowers the risk of mercury poisoning on the miners themselves and mercury pollution into the environment. The process also offers economic benefits because of its high recoveries and product purity coupled with its simple way of waste management. Lastly the process recovers refractory gold which is very difficult to recover using other conventional ASM extraction methods.
A review is presented summarising the leaching of gold with ammoniacal thiosulfate solutions, and evaluating the current use and development of ion exchange resins for the recovery of gold and silver from such leach liquors. Comparisons are also made with other recovery processes, including carbon adsorption, solvent extraction, electrowinning and precipitation. Thiosulfate leaching chemistry is compared with cyanide leaching, and the problems associated with obtaining a high yield of recovered gold using the former process are discussed. The present limitations of using Resin-in-Pulp (RIP) and Resin-in-Leach (RIL) systems with thiosulfate liquors are indicated and possible solutions discussed.
The thiourea leaching of silver from a tetrahedrite concentrate mechanically activated in a planetary mill or an attritor was studied. It was found that the two types of equipment gave rise to different rates of new surface formation and of crystal structure disordering. The rate of thiourea leaching of silver from tetrahedrite (Cu,Ag)lo(Zn,Fe)E(Sb,As)4S 13 is a structure-sensitive quantity, while the dependence of the rate constant of leaching on the empirical coefficient Sg/(l-R) (S A = specific surface, R = disordering of tetrahedrite structure) exhibits a linear character with equal slope for both types of mills. The results are also of prognostic character because they enable us to propose suitable equipment for intensive grinding depending on the demand for fineness or reactivity of the solid substances.
During the last 15 years about 700 t of gold have been consumed in the manufacture of low voltage electrical contactors, which,
over a period of time, have become redundant and thus the gold component has become candidate for recovery. The results of
laboratory and batch pilot experimentation, reported in this article, have demonstrated an average gold recovery rate of 95
per cent from used contactors by means of mechanical and chemical processing.
This study investigates the biooxidation of a refractory gold concentrate using a mixed culture of acidophilic mesophiles, moderate thermophiles and extreme thermophiles and their effect on the subsequent cyanidation and gold recovery. The experiments with high % solids using mixed mesophiles showed better oxidation potential compared to moderate thermophiles and extreme thermophiles. However, the extreme thermophiles performed better than mesophiles and moderate thermophiles during the biooxidation with
The effect of pyrite on gold dissolution was investigated in the ammoniacal thiosulphate leaching system using pure gold foils. Special emphasis was placed on gold leaching in association with pyrite dissolution, thiosulphate decomposition and gold leaching passivation. The presence of pyrite retarded gold dissolution and this detrimental effect became more pronounced at higher pyrite contents. Pyrite catalysed the decomposition of thiosulphate to trithionate. The addition of sulphate enhanced gold leaching by retarding the dissolution of pyrite in ammoniacal thiosulphate solutions. XPS analysis indicated the presence of iron hydroxide species at the gold surfaces in the presence of pyrite, which was likely responsible for the reduced gold dissolution. SEM analysis indicated a lesser extent of gold dissolution occurring in the presence of pyrite.
Thiosulfate leaching appears to be the most promising alternative to cyanidation for the recovery of gold from ores. The current paper presents a fundamental study on the gold oxidation half reaction in the solutions containing thiosulfate. It will be shown that this electrochemical reaction is unusual, as the presence of the heavier inert cations such as potassium, calcium, and caesium enhance the oxidation of gold in thiosulfate solutions. The use of sulfur-containing species such as thiourea (TU) and potassium ethyl xanthate (PEX) as additives was also studied. It will be shown that in the presence of thiourea, the gold oxidation half reaction is dramatically improved, with a typical Tafel-like polarisation curve being obtained in the presence of thiourea. However, the presence of low concentrations of xanthate were shown to completely block the gold surface and prevent its oxidation in thiosulfate solutions.
The thiourea leaching of silver from a tetrahedrite concentrate mechanically activated in a planetary mill or an attritor was studied. It was found that the two types of equipment gave rise to different rates of new surface formation and of crystal structure disordering. The rate of thiourea leaching of silver from tetrahedrite (Cu,Ag)10(Zn,Fe)2(Sb,As)4S13 is a structure-sensitive quantity, while the dependence of the rate constant of leaching on the empirical coefficient SA(1 − R) (SA = specific surface, R = disordering of tetrahedrite structure) exhibits a linear character with equal slope for both types of mills. The results are also of prognostic character because they enable us to propose suitable equipment for intensive grinding depending on the demand for fineness or reactivity of the solid substances.
An oxide gold ore was subjected to chloride/hypochlorite leaching at room temperature. The effects of three factors, including Ca(OCl)2 vs. NaOCl, OCl− concentration, and HCl concentration on gold leaching performance were investigated. Due to formation of CaOCl+ complex in solution and hence less reactivity, calcium hypochlorite produces a sluggish gold leaching kinetics, taking twice the time (46 h) to achieve maximum gold recovery of 58% compared to sodium hypochlorite. 10 g/L of total initial hypochlorite species in solution produces reasonable gold recoveries. The amount of added HCl and hence the initial pH was found to have a major effect on gold leaching kinetics and maximum gold recovery. A high level of 9 g/L of added HCl causes HClO to be very reactive, producing very fast kinetics, reaching 67% gold extraction in 4 h. It also causes a faster consumption of hypochlorous acid, through catalytic decomposition (by NiO and CuO) and disproportionation. Hypochlorous acid reactions with sulfide and ferrous content of ore proceed very slowly in the pH range of 4–11. Gold–chloro complexes are strongly adsorbed on quartz component of ore. To minimize this undesirable adsorption of gold–chloro species, the aging time must be limited to a few hours only.
This study was carried out on a finely disseminated gold-bearing ore from Kaymaz, Turkey. The main objective was to determine the mechanism of reaction between thiourea and Fe3+ in the thiourea leaching of gold. Thiourea consumption levels were also examined during the leaching operations for different leaching parameters, i.e., molar ratio of ferric sulfate/thiourea, pH, and temperature. The leaching using the thiourea process was compared with cyanidation. In optimum leach conditions, 85.8% gold extraction was obtained with feed of −53 μm particle size in 6 h.
The demand for metals is ever increasing with the advancement of the industrialized world. On the other hand, worldwide reserves of high grade ores are close to depletion. However, there exists a large reserve of metals in low and lean grade ores and other secondary sources. Metal recovery from low and lean grade ores using conventional techniques such as pyrometallurgy, etc. requires high energy and capital inputs which often result in the secondary environmental pollution. Thus, there is a need to utilize more efficient technologies to recover metals. Biohydrometallurgy, which exploits microbiological processes to recover metal ions, is regarded as one of the most promising and revolutionary biotechnologies. The products of such processes are dissolved in aqueous solution, thereby rendering them more amenable to containment, treatment and recovery. On top of this, biohydrometallurgy can be conducted under mild conditions, usually without the use of any toxic chemicals. Consequently, the application of biohydrometallurgy in the recovery of metals from lean grade ores and wastes has made it an eco-friendly technology for enhanced metal production. This paper reviews the current status of biohydrometallurgy of low grade ores around the world. Particular attention is focused on the bioleaching of black shale ore and its metallogenic diversity in the world. The review assesses the status of bioprocesssing of metals to evaluate promising developments. Bioleaching of metals is comprehensively reviewed with the emphasis on the contribution of microbial community, especially fungal bioleaching coupled with ultrasound treatment. In this manuscript, the principles of bioleaching, their mechanisms, and commercial applications are presented. The case studies and future technology directions are also reviewed
The use of ammonium thiosulphate as leaching agent for gold from a flotation concentrate has been studied. The gold content in the concentrate was 95 g/t, whereas the main mineralogical species were chalcopyrite, pyrite, pyrrhotite, tennantite and sphalerite. The experimental variables studied were: (i) Cu(II) concentration (0–6.5 g Cu(II) (added) per liter of leaching solution), (ii) thiosulphate concentration (0.1–0.7 M), (iii) pH (9–10) and (iv) pulp density (10–40%). Best gold dissolution (94%) was obtained at 0.05 M Cu(II), 0.3 M S2O32−, pH 10 and 10% pulp density after 15 h, whereas cyanidation gives the same yield but after about 46 h of reaction.
The extraction of gold, silver, copper, zinc, and nickel present in thiosulfate solutions with or without the presence of ammonia was studied. Results show that neither a primary amine nor the mixed extractants of a primary amine with tributyl phosphate or trialkyl amine oxide can separate gold from other metals in the absence of ammonia in aqueous solutions. In the case of ammonium thiosulfate solutions, the percent extraction of gold is increased tremendously in comparison with other metals, and thus gold can be separated from them. The pH50 of gold extraction for a mixed solvent of primary amine with amine oxide is about 9.0, which is very close to the pH of the actual leaching solutions. By using the mixed solvent to extract gold in the solutions from leaching of gold-containing sulfide concentrate, the separation factors of Au to Ag and Au to Cu could reach as high as 15 and 1695, respectively, for a pH less than 8.
Zinc exists as smithsonite and hemimorphite in the lead flotation tailings from the Dandi mineral processing plant in north western Iran. In this research, zinc-rich tailings produced in the Dandi plant were characterized mineralogically and a leaching study was carried out to assess the effect of several parameters on the kinetics of zinc dissolution. Parameters studied included: sulfuric acid concentration, reaction time, temperature and slurry density. It was found that leaching is controlled by a single rate-controlling step with an activation energy of 23.5 kJ/mol. To overcome some of the filtration problems associated with polymerization of silicic acid, lime was added as a coagulant. The optimum pH, holding time and temperature required to maximize the filtration rate were determined.
Elemental gold on non-metallic secondary raw materials is easy to dissolve by leaching with sulphuric acid solutions of thiourea in the presence of an oxidant such as iron(III) sulphate or hydrogen peroxide. Leaching time is highly dependent upon oxidant concentration and the thickness of the gold coating. The gold goes into solution as the Au[SC(NH2)2]2+ complex and can be removed almost completely using cationites, the gold residues being not greater than 6 μM. The gold-free leaching solution may be reused for leaching; the completely-loaded cationite resin (with a gold content of about 30–60 g/l) is burnt to recover the gold.A study of the adsorption of Au and Fe by cationites indicates that the adsorption of gold takes place purely by means of an ion-exchange mechanism, the bond of the gold complex being considerably stronger than that of the iron compounds.
The leaching of gold, silver and base metals from a sulfidic gold ore in the presence of an oxidant (peroxomonosulfate (HSO5−) or iron(III)) and leaching agent (thiourea, chloride, bromide or iodide) is compared in 1-butyl-3-methylimidazolium hydrogen sulfate (bmimHSO4) and chloride (bmimCl) ionic liquids, as well as in aqueous saturated K2SO4 as the solvent medium. Over 85% of gold and silver was recovered in the presence of HSO5−/ thiourea at 25–50 °C in both bmimHSO4 and bmimCl, with silver recovery significantly enhanced compared with that from aqueous sulfate medium. The leaching efficiency with HSO5− was similar to that with Fe(III) as oxidant in bmimHSO4 and was far superior in bmimCl. With HSO5− /halide ion (Cl−, Br−, I−) as leaching agent, gold and silver recovery in bmimHSO4, bmimCl or saturated aqueous K2SO4 improved from Cl− to Br− to I−, but only I− gave a high recovery in the bmimCl ionic liquid due to the particular stability of the iodo complex anion in this medium. However, recovery was significantly higher than in an aqueous medium. Negligible recovery of base metals occurred in the ionic liquid medium, making it highly selective for Au and Ag. Concentration dependence studies with respect to halide and oxidant have defined optimum conditions for gold and silver recovery.
a b s t r a c t Waste of electric–electronic equipment (WEEE) with an annual growth rate of about 3–5% is the fastest growing waste stream in municipal wastes. Notwithstanding their environmental pollution potential, waste of electrical and electronic equipment (WEEE) with their high content of base and precious metals, in particular, are regarded as a potential secondary resource when compared with ores. For the recovery of metals from WEEE, various treatment options based on conventional physical, hydrometallurgical and pyrometallurgical processes are available. These process options with particular reference to hydromet-allurgical processes were reviewed in this study. With their relatively low capital cost, reduced environ-mental impact (e.g. no hazardous gases/dusts), potential for high metal recoveries and suitability for small scale applications, hydrometallurgical processes are promising options for the treatment of WEEE. Since the metals are present in native form and/or as alloys, an oxidative leaching process is required for the effective extraction of base and precious metals of interest. A two-stage process based on oxidative acid leaching of base metals (Cu in particular) followed by leaching of precious metals using cyanide, thiosulfate, thiourea or halide as lixiviant(s) can be suitably developed for the hydrometallurgical treat-ment of WEEE. However, further research is required to develop new, cost effective and environmentally friendly processes and/or refine existing ones for leaching and, in particular, downstream processes.
In this study, competitive adsorption of metal chloro complexes on the polymer adsorbent XAD-7 has been investigated. Au(III), Fe(III) and Te(IV) dissolved in HCl and alkali metal chloride solutions were used as the model compounds. Adsorption equilibria were measured for individual elements as well as for binary pairs and the data were correlated in terms of activities of the adsorbable complexes. Finally, the equilibrium model together with the experimentally determined apparent diffusion coefficients was utilized in simulation of break-through curves of multicomponent mixtures.Adsorption of gold and iron can be explained satisfactorily by taking into account the formation equilibria of the tetrachloro complexes and the difference in their acid strengths. The uptake of the weakly acidic HAuCl4 depends only slightly on the HCl concentration and adsorption was substantially stronger in NaCl solutions. The opposite trends observed for iron were due to the lower stability and higher degree of dissociation of HFeCl4. The adsorbable complex of Te(IV), HTe(OH)Cl4, has lower affinity for the adsorbent, but at high concentrations it can reduce substantially the adsorption capacity for Au(III) and Fe(III). This effect was confirmed with the adsorption column runs, and the change in the break-through behavior of Au(III) and Fe(III) can be explained quantitatively by the simulation model.
The intensification of the thiosulfate leaching of silver, gold and bismuth from sulfide concentrates using mechanical activation and mechanochemical pretreatment step was investigated. The physico-chemical changes in a complex sulfide concentrate (Casapalca, Peru) as a consequence of mechanochemical pretreatment had a pronounced influence on the subsequent silver extraction. The optimum results from mechanochemical pretreatment and subsequent leaching of the concentrate with ammonium thiosulfate were achieved with 99% recovery of Ag after only 3 min of leaching. The leaching of gold from a mechanically activated complex sulfide concentrate (Banská Hodruša, Slovakia) using ammonium thiosulfate was studied as follows. Physico-chemical transformations in the concentrate due to mechanical activation have an influence on the rate of extraction and the recovery of gold. It was possible to achieve 99% Au recovery within 45 min for a sample mechanically activated. Mechanical activation proved to be an appropriate pretreatment for this concentrate before extraction of gold into thiosulfate leaching solution. The selective leaching of a bismuth from the lead concentrate (Atacocha, Peru) by using of sodium thiosulfate and mechanical activation as the pretreatment step was examined as the last example. It is possible to achieve more than 90% recovery of a bismuth in leachate even in three minutes for mechanically activated samples.
Thiosulphate leaching of precious metals has been developed as an alternative and non-toxic technique to the conventional cyanidation. This process has advantages over cyanide in a decreased interference from foreign cations and a lower environmental impact. Ammoniacal thiosulphate solution allows the solubilization of gold as stable anionic complex. Leaching of gold occurs at appreciable dissolution rates.The aim of this work is devoted to a preliminary evaluation of the feasibility of thiosulphate leaching for the extraction of gold from precious metals ores on a laboratory scale. The experimental work has allowed the authors to point out the influence of temperature, thiosulphate concentration, ammonia concentration and copper sulphate concentration on the gold dissolution from an ore (51.6 g/t Au) that originated from the Dominican Republic.Gold was recovered from the leach liquors by adsorption onto activated carbon or by electrowinning. With the optimisation of the process parameters about 80% gold recovery has been attained, as in the case of conventional cyanidation.
An investigation has been carried out to identify the parameters of importance in the dissolution of gold and silver values from an ore using ammoniacal thiosulfate solutions containing copper. The chemistry of the dissolution process has been analyzed through the construction of EH −pH diagrams. The experimental results show that it is possible to solubilize about 70% of silver and 90% of gold values in one hour. The leaching process is quite sensitive to thiosulfate and ammonia concentrations. Maintenance of proper Eh and pH conditions is necessary to prevent precipitation of copper as Cu2S with concomitant loss of silver values from solution.
Previous researchers have rationalized the thiosulfate oxidation by ammoniacal copper(II) on the basis of the rate equation −d[Cu(II)]/dt = k[Cu(II)][S2O32−]/[NH3] for the decomposition of the reaction intermediate Cu(NH3)3(S2O3)0, but no such detailed analysis has been made in the case of gold. This work revisits literature data to show that the oxidation of thiosulfate takes place via mixed complexes Cu(NH3)2(S2O3)n−(2n−2) (n = 1 or 2). The oxidation can be considered as a first-order decomposition of these complexes, with a rate constant kCu(1) = 4×10−4 s−1, to produce S2O3− which undergoes dimerization to S4O62−. It can also be treated as a second-order decomposition of Cu(NH3)3(S2O3)0 or Cu(NH3)2(S2O3)0 with rate constants kCu(2) = 0.2 and 0.1 L mol−1 s−1, respectively, which produce S4O62− in one step.The literature data reported for gold dissolution in the copper(II)–ammonia–thiosulfate system show first-order dependence with respect to copper(II) and thiosulfate concentration, but it is relatively less affected by ammonia and hydroxide concentrations. At higher concentrations of copper(II) and thiosulfate, the rate becomes zero order with respect to these reagents. The leaching data obey a shrinking core model for both gold powder and carbonaceous or roasted ore with an apparent rate constant ksc≈10−5 s−1, but the ksc also shows a first-order dependence with respect to the measured copper(II) concentration in solution. The copper(II) sepeciation analysis shows Cu(NH3)2(S2O3)22−, Cu(NH3)2(S2O3)0 and Cu(NH3)3(S2O3)0 as the predominant species in ammoniacal copper(II)–thiosulfate solutions. Thus, gold oxidation can be represented by a sequence of equations representing adsorption and redox reactions.
The rudimentary nature of small-scale gold mining activities often generates a legacy of extensive degradation and deplorable social conditions, both during and after activities have ceased. Small-scale mining usually involves the extraction of secondary gold from placer deposits (alluvial, colluvial or elluvial), which can be liberated and treated using gravity methods. In the Guianas, the most popular form of small-scale gold mining is referred to as “land dredging”, a combination of hydraulicking and suction dredging. This method requires application of large volumes of water for both mining and mineral processing; in most cases, there are no containment structures for the waste tailings generated. Mercury, a dangerous pollutant, is the preferred method employed by small-scale miners for gold recovery. Gold extraction using mercury is comprised of the following four stages: (1) amalgamation, (2) separation of amalgamation, (3) removal of excess mercury, and (4) burning of the remaining amalgam to produce a gold sponge. Mercury can be released into the environment at each stage, which makes the promotion of mercury-free alternatives imperative.Technical alternatives for small-scale gold mining, however, must be thoroughly evaluated, pre-tested, modified accordingly and successfully transferred. Moreover, technology must be inexpensive, relatively simple and easy to adapt, while allowing a rapid rate of return. WWF-Guianas is working with the regulatory agencies and other relevant stakeholders of the Guianas to reduce the environmental footprints caused by small-scale mining. The major aspects of this program are to develop the capacity and regulatory mechanisms within the local government, to promote mercury-free technology, and monitor mercury in the environment. This paper reviews the alternative technologies being investigated by the WWF-Guianas for use in the small-scale gold mining industry.
The leaching behaviour of three Fe-containing sphalerite minerals (0.45, 11.40, and 12.90 wt.% Fe) was investigated at pH 1 in O2-purged HClO4 from 25 to 85 °C. For all three sphalerite samples, the leach data closely fitted a rate expression which incorporates both a shrinking sphere term and rate control via diffusion through a reacted surface layer. The leach rate increased with increasing Fe content of the sphalerite. At 85 °C after 168 h of leaching, 0.89 of the Zn had been leached from the 12.90 wt.% Fe-containing sphalerite while, over the same period of time, 0.49 of the Zn had leached from the 0.45 wt.% Fe-containing sphalerite.The activation energies for leaching decreased with increasing Fe content of the sphalerite (i.e. for Zn leaching 63±6, 50±7 and 39±2 kJ mol−1 for the 0.45, 11.40 and 12.90 wt.% Fe-containing sphalerite samples, respectively). The activation energy of leaching of Fe and Zn was within experimental error for the 11.40 and 12.90 wt.% sphalerite samples. The solution Fe content was not measured reliably for the 0.45 wt.% Fe-containing sphalerite.The leaching mechanisms have been interpreted as a function of surface speciation, which was analysed using time of flight secondary ion mass spectrometry (ToF-SIMS), X-ray photoelectron spectroscopy (XPS), powder X-ray diffraction (XRD) and scanning electron microscopy (SEM) for samples leached at 85 °C. The decrease in Zn dissolution rate on leaching is attributed to the formation and growth of a polysulfide surface layer, which forms during the initial rapid leach period. The formation of elemental sulfur during the subsequent slow leach period does not further affect the leach rates. The concentrations of polysulfide and elemental sulfur that form increase with increasing Fe concentration as expected from the higher leach rates. The formation of polysulfide and elemental sulfur on the ZnS surface consumes H+ (via conversion to H2O) thus decreasing solution acidity.
The effects of the particle geometry and size, stirring speed, pH, chlorine and chloride concentrations and temperature on the kinetics of the aqueous chlorination of gold were studied. Surface examinations of gold samples leached in different conditions were carried out by SEM/EDS and XPS.The kinetics appear to be chemically controlled in conditions of full suspension of the particles. The dependence of pH and the chlorine and chloride concentrations indicate that the rates were essentially determined by the concentration of the trichloride ion. The expression obtained for the specific rate was: Rate (mol cm−2 min−1) = 4.25 105 [Cl3−] exp () A discussion of the possible mechanism is included, based on the kinetic dependence, previously reported electrochemical studies and mixed potential measurements.
A new mathematical model was developed for leaching of zinc ores containing silicates such as hemimorphite which produce a gel during leaching with sulfuric acid. This model is based on the shrinking core model in which the particle size and the reacting core shrink simultaneously. It was shown that the actual dissolution time of the ore particles is longer than the time corresponding to the dissolution of chemical zinc oxide itself. It was suggested that because of the existence of silicates in the ore, a gelatinous layer was formed around the reacting core. Since the gel product is soft, it breaks apart when the particles collide and as a result, the particles shrink. However, a thin gelatinous layer always covers the reacting core which increases the mass transfer resistance and increases the leaching time. This model was applied to leaching of a zinc-rich tailing containing hemimorphite and the thickness of the gelatinous layer as well as the diffusion coefficient in this layer was determined.
Crosslinked polyacrylamide and polyvinylpyrrolidone gels have been used to subcutaneously implant (125)I-labeled immunoglobulin, (125)I-labeled luteinizing hormone, (125)I-labeled bovine serum albumin, (125)I-labeled insulin, [(3)H]prostaglandin F(2alpha), and Na(125)I into hamsters. From the rates of absorption of the solutes, their diffusion coefficients were determined. The diffusion coefficients showed a logarithmic dependence on implant polymer concentration and solute molecular weight. Release of the solutes from gel preparations incubated 10 mM phosphate buffer (pH 7.2) at 37 degrees revealed a similar relationship between solute diffusion coefficient, molecular weight, and the concentration of polymer. A general equation was derived that gives the expected diffusion coefficient of a substance in a polymer gel from its molecular weight, diffusion coefficient in solvent, and polymer concentration of the gel.
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Gold and silver recovery from chlorous mud
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Gold recovery from computer circuit boards with acidic thiourea
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Study of the thiosulphate leaching of gold from carbonaceous ore and the quantitative determination of thiosulphate in the leached solution
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