AQUEOUS AND LOW TEMPERATURE PROCESSING OF MINERALS, METALS AND MATERIALS

This study focuses on hydrometallurgical treatment of lateritic nickel ores obtained from Manisa - Caldag region of Turkey. 1.58 % of nickel and 0.12 % of Co containing lateritic ores characterized by both chemical analysis and x-ray diffraction. Leaching conditions of ore were carried out by determining the optimum parameters such as sulphuric acid concentration, leaching duration, stirring rate and solid / liquid ratio. Nickel and cobalt were extracted in the form of nickel sulphate and cobalt sulphate solutions after the leaching, respectively. The dissolution of iron in the ore during the leaching process was monitored. Finally several hydrometallurgical purification techniques were carried out to obtain relatively pure nickel and cobalt solution for further reduction / refining techniques.

Self-propagating high temperature synthesis processes are one of the highly productive and economically reasonable methods for the metallurgy industry. The latest stage in the production of advance ceramic powders by SHS method was the leaching process, in which the elimination of several impurities such as MgO and magnesium borates from SHS product was took place by using concentrated hydrochloric acid solutions. Leaching of SHS product resulted in the production of high purity advance ceramic powders together with Mg and B containing spent solution. In this study, removal of zirconium, magnesium and boron compounds from relevant chloride containing spent solutions was aimed. The effects of chemical reagents, pH of the solution on the precipitation efficiency were investigated in the study. Chemical analysis and the characterization of solid phases by using x-ray diffraction and chemical analysis were the latest stage.

Flotation assisted with acidic non-oxidative leaching of the carbonate feed has been presented. The untreated feed is hardly-upgradeable by standard procedures due to a very fine mineralization and severe problems in liberation of copper-bearing sulphides. The acidic leaching results in chemical effective liberation of copper and valuable met-als sulphide particles finely disseminated in a carbonate gangue. For this leaching a part of the clay-carbonate feed material was separated from the commercial flotation circuit and subjected to leaching with sulfuric acid prior to further flotation. The feed suspension, after non-oxidative leaching, was returned to the flotation circuit in order to separate freshly precipitated gypsum and was finally subjected to cleaning flotation. The combined leaching-flotation process was eventually successfully applied on the in-dustrial scale. In this paper a flotation circuit with acidic leaching of the feed was pre-sented. Flotation indices of the chemically modified feed were compared with the stan-dard system without chemical pretreatment. After several months of operation of the new technological system, a significant increase of copper and other metals recovery was observed along with an increase in concentrate grade. On the basis of technologi-cal experience and additional laboratory investigations application and development of proposed chemical pretreatment and further full hydrometallurgical processing of flota-tion concentrates for improving metals recovery is considered

Different cyanide forms exist according to their composition and stability, each one with different toxicity grade. The most toxicity is the free cyanide. The mensuration of concentration of the cyanide can be carried out by diferents methods. The free cyanide in solution can transform to less toxic forms by means of the natural degradation, precipitation, cyanide recovery, adsorption has more than enough activated carbon and oxidation being this last one the most applied. This paper presents the evaluation of degradation tests for oxidation of cyanide applied to solids and liquids effluents, and pulps of different treatments in the mill, located in the county of Segovia at Antioquia and Marmato, at Caldas - Colombia. The technological alternatives researched for at the oxidation of cyanide in effluents of solids and liquids taken in Segovia, were: Sodium hypochlorite, hydrogen peroxide and the combination of the reagents before mentioned. For the treatment with sodium hypochlorite it is required 12.5 kg NaOCl/kg CN- - for a time of 40 minutes; While for the process with hydrogen peroxide it is needed 3.5 kg H2O2/kg CN- - for a time of 2.5 h. For the combination of the methods of the sodium hypochlorite with hydrogen peroxide is required 2.5 kg H2O2/kg CN, with 5 kg NaOCl/kg CN- for a time of 20 minutes. The samples taken in Marmato, correspond to pulp mineral that leave the extractive process and that it contains high cyanide concentrations. In these samples the degradation was applied by oxidation using hydrogen peroxide and the Caro´s acid (AC). The speeds in the degradation indicated a reduction of the cyanide content in more than 93% in a time of 4 h for the use of the hydrogen peroxide and, 5 minutes for the use of the Caro´s acid, with relationships among the H2O2/CN- of 2,5 and AC/CN- of 2 and H2SO4/H2O2 of 2.

HfO2 is expected to be an alternative gate dielectrics used in ultra large scale integration (ULSI) to the conventional SiO2 thin film. It is required that the materials used in ULSI are sufficiently purified to avoid unexpected errors. Commercial Hf, however, contains > 20,000 mass ppm Zr, which is the most difficult impurity to be removed and causing flat-band voltage shift. Therefore, it is important to establish process of removing Zr from Hf. Furthermore, high purity Hf is indispensable for investigating the intrinsic properties of Hf metal. A lot of processes have been proposed for separation of Hf and Zr. In the industry, solvent extraction and chloride distillation are used. In addition, solid/liquid extraction methods such as chelating or solvent impregnated resins have been reported. Among those we focus on anion-exchange separation which is very useful to purify some transition metals. Anion-exchange behavior of Hf(IV) and Zr(IV) has been reported by a several researchers, but the reported distribution coefficients of Hf(IV) and Zr(IV) vary widely. In this work, we investigate the anion-exchange behavior of Hf(IV) and Zr(IV) by the batch experiment in the condition identical to that in the column separation as possible and determine the optimum condition for separating Zr from Hf. Strongly basic anion-exchange resin, DIAION SA10A supplied by Mitsubishi Chemical Corp., was used. The examined species were dissolved into HCl solutions in the form of chlorides. The mixture of anion-exchange resin and the solution containing the examined species was shaken in the thermostat bath for a certain time. After the resin was filtered, the concentration of the species in the initial and final solution were analyzed by inductively coupled plasma atomic emission spectroscopy, ICP-AES. As a result, adsorption isotherms of Hf(IV) and Zr(IV) from HCl solutions follow the Langmuir isotherm and the distribution coefficient of Zr(IV) is higher than that of Hf(IV) in whole examined range, 8 – 11 M HCl and at 10 – 40 °C. In the range of < 8 M HCl both Hf(IV) and Zr(IV) are not adsorbed on anion-exchange resin. It indicates that negatively charged species does not form in this hydrochloric acid concentration range. And adsorption of both Hf(IV) and Zr(IV) are not as strong as reported. The separation factor assessed from the distribution coefficients determines that the optimum separation condition is in 9.5 M HCl and at 20 °C.

Despite extensive investigation of alternative lixiviants, aqueous alkaline cyanide remains the mainstream technology for leaching gold from gold ores. When gold ores contain reactive sulfide minerals, these react with cyanide to form metal cyanide complexes and thiocyanate. High levels of cyanide are needed to ensure that enough free cyanide remains available to leach gold. The metal cyanide complexes must be removed from the circuit via a bleed stream to prevent their reaching a concentration that interferes with gold recovery. We are examining electrolytic treatment as an environmentally acceptable and economically sound approach to treating the bleed streams. Metal cyanide complexes can be reduced cathodically, albeit at reduction potentials lower than those of the uncomplexed metal ion. However, the overall feasibility of such an approach would be strongly influenced by the anodic reactions that predominate during such electrolytic treatment. Several reactions are possible, for example oxidation of water or cyanide, using sulfite as an anode depolarizer, or oxidation of thiocyanate. The latter would be particularly useful, since it removes thiocyanate impurities from solution, and returns cyanide to the solution. Here we present a theoretical, thermodynamically based analysis of the relative favorability of different possible anodic reactions; Unfortunately, thiocyanate oxidation is less favorable than the oxidation of cyanide or sulfite under most conditions. We also present experimental work done to investigate the polarization behavior of the most favorable reactions on different anode materials under different experimental conditions, with the objective of identifying substrates that are electrocatalytically active towards thiocyanate oxidation while inhibiting cyanide oxidation.

Process Research ORTECH has developed a unique hydrometallurgical process based on a mixed chloride lixiviant. This mixed chloride technology has been used to develop process flowsheets for the extraction of base metals and PGMs from sulfide ores and high purity titanium dioxide from ilmenite ores. For base metals and PGMs recovery, base metals are recovered during a first stage leach and elemental sulfur is liberated from hydrogen sulfide formed during this leaching stage. During second stage leach, precious metals are recovered using proven separation technologies. Some of the advantages of the process are the recovery of elemental sulfur from sulfide sulfur, recycle of the chloride lixiviant thereby eliminating effluent treatment steps and use of excess energy from the Claus reactor for the pyrohydrolysis step. For the synthesis of high purity titanium dioxide, ilmenite ore is first leached in a mixed chloride lixiviant at atmospheric pressure. The solution purification is carried out using solvent extraction in two stages. In the first stage solvent extraction, iron removal is carried out and in the second stage titanium is extracted. The strip liquor obtained from the second stage solvent extraction containing titanium is used for the production of high purity TiO2 by thermal precipitation. The barren solution is used for reagent recovery and the regenerated leach liquor is recycled.

Experiments on the beneficiation of multi metal pyritic sulphide ore of Copper - Lead - Zinc from Hindustan Zinc Limited, Udaipur (India) conducted using chemical auxiliaries based on guargum and its derivatives as depressant systems without any basic alteration in plant operating conditions, yielded better metallurgical results for recovery of copper values in flotation of copper concentrate. The silica depressants studied in present investigation, includes sodium silicate, guargum as well as its cationic derivative along with sodium isopropyl xananthate (IPX) as collector and methyl isobutyl Ketone (MIBK) as frother. The other operating parameters viz. Reagent, concentration and pH of the pulp were also systematically studied and optimised. The Improved metallurgical results were obtained with combination of sodium silicate and cationic guargum as silica depressant at pH 10.5.

Significant amounts of secondary zinc bearing materials are converted in C. Z. O., which is an attractive feed for the production of zinc metal in conventional processes. The ISP, one of the traditional technologies using this C. Z. O. Material, is phasing out and most of the plants based on this technology are not in operation. This is moving C. Z. O. Towards its use in zinc sulphate E. W. Unfortunately, the C. Z. O. Cannot be fed directly to the sulphate E. W. Instead, it has to be carefully washed to reduce the content of harmful elements as chlorides. Also, after the removal of these harmful elements, the presence of other elements commonly requires the blending of this material with traditional feed. The EZINEX® Process, a zinc E. W. Based on a chloride bearing electrolyte developed by Engitec Technologies, allows the direct treatment of C. Z. O., without washing, even if it is the primary feed. Engitec Technologies also successfully piloted a thermal technology, the INDUTEC® Process, which is able to convert secondary zinc bearing material to C. Z. O. The combination of these two technologies can be used to recover zinc coming from many different sources

The article offers the mechanism of nickel protoxide dissolution through its oxidation to oxide. The direct reaction of nickel protoxide dissolution by sulfuric acid: NiO + H2SO4 ® NiSO4 + H2O proceeds very slowly making practical use of this process impossible. It is known that higher nickel oxides dissolve very well; Considering this fact nickel protoxide leaching through NiO oxidation to Ni2O3 with sulfuric acid use was studied. To initiate the dissolution it is necessary to supply strong oxidizing agent into the system in the amount sufficient for the creation of redox potential in the system on the level of ~1.4B. After that the oxidizer is not requires as it regenerates the oxide dissolution:2NiO + O ® Ni2O3 Ni2O3 + 2H2SO4 ® 2NiSO4 + 2H2O + OSo in the process of this kind the oxidizing agent may be required only for its mechanical loss compensation. Beside the atomic oxygen there can be used other oxidizing agents capable to provide the creation of the required redox potential, for example, S2O82-. The process indices are compared as applied to nickel protoxide synthesized from pure reagents and produced in industrial scale in the process of sulfide nickel concentrate roasting. Synthesized nickel protoxide dissolution by the acid occurs in less than a minute after the oxidizing agent is charged into the system. The oxidizing agent is used only for the process initiation in the amount not more than 1% of the stoichiometry. With this, acid consumption equals to the stoichiometric one. Industrial product dissolution lasts 2-3 hours, it requires 15-20% surplus of the acid and the oxidizing agent consumption is 8-10% of the stoichiometry, that can be explained by admixtures availability. The process is controlled and managed by the value of the redox potential in the system.

The technology for processing of rare metal ores (Chuktukon deposits, East Siberia, Russia) was developed. Ores are composed of iron-manganese oxides and contain 3-7 % of rare-earth oxides in the phosphate form (turnerite, florencite) and 0.5-1 % of niobium oxide (chalcolamprite). These ores are refractory and do not practically undergo to concentrate, therefore various processes for their direct chemical processing are considered. A flowchart including autoclave nitro-acid leaching and extractive separation of rare-earth elements, regenerative processing of tails after leaching and extraction of Nb concentrate into non-magnetic fraction is proposed. Under this scheme nitrate solutions subject to thermal decomposition for the purpose of nitric acid regeneration. Pyrometallurgical process of ore processing with production of cast-iron and products from rare metals was suggested. The different technological problems on realization of the developed processes will be discussed.

This article presents the results of evaluating the acid leaching of ores containing black copper, by modifying the redox potential of the leaching solution through the addition of several potential modifiers. Five leaching conditions are evaluated: Traditional sulphuric acid medium and 4 with the addition of a potential modifier: Ferrous sulfate (FeSO4), white metal (Cu2S), sulfur dioxide (SO2) and ozone (O3). Two methodologies were used for leaching in the laboratory: A bottles test and use of 1m columns. The ore that is leached comes from the Lomas Bayas mining company whose characterization indicates 0,251%CuT and 0,380%Mn. Mineralogical analysis indicates the presence of black copper at a rate exceeding .... The results of the tests in bottles indicate that by decreasing the potential for reducing agents, extraction of copper was achieved in the range of 83.7%, which exceeds by 16.7% the removal obtained in conventional acid leaching. The potential increase (ozone) does not favor the removal of Cu. The extraction percentage of Cu is directly related to the Mn extraction percentage. The ratio of dissolved moles of Mn per dissolved moles of Cu is between 0.5 and 2 for the different agents used. The results of the leaching in columns show that with the white metal and SO2, it is possible to maintain the potential of the solution to values below 450 mV, obtaining the best results of copper extraction, at approximately 60%, which is 18% higher than conventional leaching and an acid consumption in the range of 11 kg/t, over a period of 45 days of leaching.

The main method for the extraction of gold and silver used in Mexico and worldwide is cyanidation. In this process some specific minerals contains species of copper which are soluble in the cyanide solution, causing difficulties in the process (increase of reagent consumption, contamination of the final metal, increase of the water consumption due to the impossibility of recycling the solution back to the process, etc.). According to the above, in this work three methods are proposed in order to eliminate the copper cyanide complexes, namely: 1) the acid precipitation of CuCN, 2) the use of amines at alkaline pH to precipitate the copper-cyanide-amine complexes and 3) the solvent extraction of the copper-cyanide complexes with amines at alkaline pH. These processes were evaluated using synthetic solutions of composition similar to those found in cyanidation plants (197-739 mg/L copper and different cyanide/copper molar ratios). For the precipitation of copper by lowering the pH, it was found that 93 – 98% of the copper can be removed as CuCN (70 % copper, 10 µm average size particle) at pH 2.5 by filtration. When the CuCN solid is formed, more than 80% of cyanide contained in the copper-cyanide complexes is liberated as CN- (free cyanide), which can eventually be recycled back into the cyanidation process. The use of quaternary amines salts (hexadecyl trimethyl ammonium chloride, tallow trimethyl ammonium chloride, and ditallow hydrogenated dimethyl ammonium chloride) were evaluated in order to remove the copper-cyanide complexes by forming a precipitate of copper-cyanide-amine. The best efficiency of copper removal (more than 93%) was obtained using the amine tallow trimethyl ammonium chloride at pH of 12 and 12.32 g amine/g of copper. For the solvent extraction technique, tallow trimethyl cmmonium chloride, methyl trialkyl ammonium chloride and tricaprylylmethylammonium chloride were studied as extractants. The experimental results showed that it is possible to obtain a copper removal of 99% when using 0.033 mol/L of the extractant Adogen 464 (organic/aqueous volume ratio (O/A) = 1) in the range of pH of 9 – 11. Up to 99% of copper can be stripped from the organic solution after three contact times (5 min each) with 50 mL of sodium hydroxide 0.5 M (O/A =1).

Traditional hydrometallurgical process for oxide ore is using sulfuric acid for copper extraction with poor dissolution of noble metals. A Halide Chlorine leachant at acidic conditions are tested to extract copper, gold and silver. Dissolution is enhanced in copper at pH 1.8-2.0 and for noble metals with fresh blend formation of chlorine and NO3 ions to form aqua regia insitu. The chlorine ions controlled the process reaction process for gold and silver dissolution. The rate of noble metals dissolution is function of chemical reaction on solid-liquid interphase and mass transfer phases of determinate reaction area. The silver rate dissolution is increased at chlorine concentration. Variables such as mineralogical composition, concentration of reagents, dissolved oxygen, reaction time, pH and temperature were evaluated.

A significant portion of the known gold reserves in the world are contained in ores that also contain copper mineralization. Most copper minerals react quantitatively with cyanide, which is the lixiviant used universally in the gold industry to leach gold from its ores and concentrates. In practice, it is difficult to leach gold without simultaneaously leaching any cyanide soluble copper minerals that may be prtesent in the ore body. This can lead to very high cyanide consumption during leaching and high reagent consumption during cyanide consumption, which in some cases will render a potentially attractive gold prospect uneconomic. There are several ways of dealing with the problem and they all boil down to one of two choices (i) either take measures to minimize the reaction between cyanide and the copper minerals, or (ii) treat the gold plant tailings to recover copper cyanide, which allows the copper to be sold as a commercail product and the cyanide to be internally recycled to leaching. The opportunities for minimizing the reaction between copper and cyanide are limited as copper is usually present in the ore at a concentration at least 10 to 100 times greater than that of gold, and their rates of leaching are similar under typical gold leaching conditions. However, the mechanisms by which copper and gold leach in cyanide are quite different and this presents an opportunity to achieve a partial separation based on kinetic differentiation. Copper leaches in a chemically controlled reaction while gold leaches in a redox reaction in which the rate of reaction is mass transfer or diffusion controlled. Some separation can be achieved by manipulating the concnetrations of cyanide and oxygen. When copper minerals react with cyanide they form water-soluble copper cyanide complexes, and several processes have been developed to treat gold plant tailings to either extract the copper cyanide (with anion exchange resins or activated carbon) for subsequent treatment or to break it down in a chemical reaction and precipiate the copper in a recoverable form. These processes will be discussed, with particualr emphasis on the SART process, which is gaining acceptance in the gold mining industry. Gold plant tailings are treated with sulphuric acid and a source of sulphide ions in the SART process, which precipitates copper as chalcocite (Cu2S) for recovery and sale as a high grade concentrate, and restores the complexed cyanide as free cyanide for recycling to leaching. Several SART plants have been built in the last few years, and are operating successfully.

For the recovery of refractory metals such as niobium and tantalum from their ores, fluoride extraction method was examined by the use of highly oxidative NO2-HF solvent. In this procedure, refractory metals reacted with the solvent at room temperature forming complex fluoride such as NONbF6. After the filtration, the complex fluoride was extracted into the pure HF solvent. The mutual separation of niobium, tantalum as well as titanium which were contained in the ore was made by using the solubility difference of these fluorides in HF solvent. Conditions for selective fluorination and separation were also discussed.

Authors: Przemyslaw Los, Michal Gieron, Slawomir Ruta, Przemyslaw ZaprzalskiNanometallurgy S. A., PolandIn our previous studies we have shown that potentiostatic (and potentiostatic pulse) electrolysis could be applied to obtain copper of high purity from copper industrial electrolytes, by-products and wastewaters of copper industry. On the basis of previous results a new, patent pending technologies of selective electrowining from complex acidic aqueous electrolytes and wastewaters have been developed. In this paper for the first time the pilot scale test results of the technology will be presented focusing on copper, silver, arsenic and nickel deposition with purity up to 99,999%. The properties of the cathodes obtained using a new process as well as some of its important parameters will be discussed. The new method can be implemented in the existing copper refineries without essential changes in the tank houses apart from the power supplies (rectifiers and software) which should fulfill certain specific requirements. There are many advantages of our process in comparison to the existing galvanostatic electrorefining and electrowining namely much higher cathode current densities can be used to obtain commercially pure (grade A) copper deposit with very high current efficiencies which can eventually contribute to factory foot print reduction and/or production scale-up. Other positive aspects of our technology result from the fact that cathodic copper is received in broad spectrum of copper content in electrolyte from 50g/liter to 1mg/liter, there is no co-deposition of Arsenic neither evolution of AsH3 gas and pure crystalline Arsenic as a commercial product can be obtained. Potential consequences to hydrometallurgical plants will be discussed.

The ECUPREX Process is part of the FLUOBORATE Technology developed in the nineties by Engitec Technologies for the electrochemical recovery of lead and copper from primary and secondary sources. We discover that chalcocite and covellite are highly leachable using this ECUPREX Process. However, chalcopyrite was not leachable in this system. We studied different approaches to leach copper from chalcopyrite. We needed to convert it into something suitable to be fed to the ECUPREX unit in order to realize a fully hydrometallurgical approach. We realized that matte and/or white metal has a composition very close to a secondary copper sulphide. Some years ago, we run a campaign in our lab pilot unit recovering copper directly from matte and white metal coming from a primary copper smelting. In this paper we will present the results of this campaign.

Due to systematically shrinking copper resources and remarkably declining copper ores grade the growing role of hydrometallurgical processes in manufacturing copper and accompanying metals can be observed. Technical, economical and ecological aspects of application of hydrometallurgy were discussed with regard to dominating smelting methods. Exceptional conditions of Polish copper industry was evaluated and indicated were important technical, mineralogical and chemical aspects of necessary application of hydrometallurgy at KGHM for more effective processing of copper ores and concentrates. Analyzed were the hydrometallurgical unit operations compulsory in unique conditions of Polish copper industry: High content of carbonates, decreasing content of Cu, increasing content of Pb, As and organic carbon. Application of hydrometallurgy in KGHM circuits was acknowledged as the only real chance for more rational and sustainable utilization of metal resources from the Polish copper deposits. The beneficial chemical and mineralogical composition of Polish copper ores and concentrates along with easy access to sulphuric acid additionally substantiate the need for comprehensive research work.

Since the first plant was built at Mount Isa, twenty ISASMELT plants have either been constructed or are under construction. A recent addition to the family of ISASMELT licensee plants is the Mufulira smelter in the Zambian Copperbelt, where the ISASMELT plant has replaced a rotary drier and an electric smelting furnace. The change has been beneficial to Mufulira smelter. The primary copper ISASMELT top submerged lance (TSL) technology continues to attract interest from potential users who seek high productivity smelting. In addition to the known benefits of productivity, and flexibility, ISASMELT operators have also achieved improved energy efficiency. Plant production statistics from the Mufulira smelter shows that power consumption per unit of production has been significantly reduced following the introduction of ISASMELT technology.

The extraction and beneficiation of synthetic rutile has traditionally been carried out by using either via sulphuric acid leaching after slag smelting of titaniferous concentrates or, via slag smelting to better than 80% TiO2. Besides acid leaching, the lesser grades of TiO2 are also used for direct chlorination for the synthesis of high grade of titanium tetrachloride. These two conventional processing techniques for titaniferous materials have relied on the supply of high grades of heavy mineral concentrates. In the past we have reported the processing of ilmenite for the extraction synthetic rutile using alkali roasting and leaching, and we have shown that the extraction of synthetic rutile of purity more than 90 weight % TiO2. In this paper we aim to demonstrate the improvement in extraction chemistry leading to recovery of whiter grades of titanium dioxide with better than 95 weight percent purity. We propose a new methodology, which utilizes the low-grade titaniferous minerals for the extraction of synthetic rutile and also significant quantities of rare-earth elements. We demonstrate the alkali roasting of lower grades of mineral concentrates and subsequent leaching in an aqueous medium for preferential separation of white grades of TiO2 during controlled atmosphere leaching of alkali titanate. The paper specially examines the leaching chemistry, changes in the crystal structure, and the influence of temperature and pH on the composition of titanium-rich mineral phases formed. The analyses of phases were carried out using the X-ray diffraction, scanning electron microscopy, and XRF. The separation and concentration techniques for the rare-earth elements in the magnetic and non-magnetic fractions of ilmenite are also examined for reclaiming rare-earth elements at the front end of unit process.

Battery paste smelting is the most problematic operation in lead acid battery recycling. For this reason, there have been considerable efforts in the last three decades to find a hydrometallurgical approach to produce lead from paste. Engitec Technologies is continuously updating the CX Technology to design a process more technologically advanced, more environmentally friendly, but also more economically viable. The development of the super-desulphurized paste technology gave us the perfect feed for an electrochemical process. So we resumed studying the possibility to electrowin lead out of this material. In this paper, we will describe our new process based on a chloride electrolyte fed. This technology uses a very unusual E. W. Cell that is working at a very high current density and doesn’t require the traditional stripping of the metal.

Electric arc furnace dust (EAFD) of ferromanganese production units, in the form of slurry, contains tar, alkalies, manganese, zinc, iron, silica, calcium, aluminum and other elements. A hydrometallurgical route based on solvent extraction technique was investigated for selective separation of manganese from the dust. Leaching of the EAFD resulted in an aqueous feed containing 4 g/L of manganese and 0.87 g/L iron. At the next stage, separation and extraction of manganese and iron from the leach liquor performed using D2EHPA, Cyanex 272, Cyanex 302 and their mixtures in various proportions. The synergistic effect of the extractants on the separation of iron and manganese with a mixture of D2EHPA and Cyanex 272 or Cyanex 302 was studied. Increasing the Cyanex 272/ 302 to D2EHPA ratio in the organic phase caused a right shifting of extraction isotherms of manganese and Iron. The highest separation factor of iron over manganese was obtained with 15:5% v/v of Cyanex 302: D2EHPA mixture. Effects of various aromatic and aliphatic diluents, such as hexane, kerosene, and carbon tetrachloride on the extraction were also investigated.

The technique of rational planning of multifactorial experiment based on the matrixes and M. M. Protodjakonov's equation is applied to research of hydrometallurgical process. The sulphuric acid leaching of the cakes from niobium electron-beam remelting sublimates with sodium carbonate are considered. The private dependences of the main leaching parameters and the generalized equation are received. The optimum condition for a division of aluminum and niobium are determined with the help of the five-factorial mathematical model.

In recent years, the application of microbiological methods to minerals processing, metal extraction and environmental remediation has gained an impetus with commercial bioleaching or biooxidation plants being operated for the extraction of copper, cobalt, gold and uranium. The subject of bio-flotation and bio-flocculation is relatively new, wherein the adhesion of microbial cells to minerals, modulates their surface chemistry to make them either hydrophobic or hydrophilic. The bio-modification of mineral surfaces is also facilitated by the extra-cellular bacterial secretions. There is tremendous potential to utilize microorganisms in remediation processes. In this paper, the role of acidophilic chemolithotrophs such as Acidithiobacillus thiooxidans and neutrophilic heterotrophs such as Paenibacillus polymyxa in the beneficiation of sulphide minerals is illustrated. The role of bacterial cells as well as metabolic products such as bioproteins and exopolysaccharides in significantly altering the surface chemistry of sulphide minerals is discussed. Case studies with respect to an Indian copper mine and mill tailings are presented. Typical studies on the bioremediation of acid mine drainage and toxic metals such as lead and chromium are highlighted. The mechanisms of microbe-mineral interactions are delineated

The aim of the paper is to describe a special monitoring of the characteristics of undesirable physical and chemical pollutants in the water that can adversely affect human health. Generally the causes of water contamination are human-made through industrial processes or because of natural causes. Among human made or industrial causes is the pollution from mineral and metal extraction plants, discharge of waste near water resources, use of fertilizers, pesticides, car washing etc. Among natural causes of water pollution are volcanoes, cyclones, natural disasters and emergencies. Floods increase the risk of water contamination in the water treatment installations and water supply networks and through sewage discharge. As a result monitoring of water pollutants like chemical elements and physical substances of natural or synthetic origin in order to assure those are below the maximum allowed values (MAV) is a very important process to protect the human health and have clean beverage production.

There is extensive literature on the investigations carried out to determine the process operating conditions for extraction of nonferrous, base and precious metals from complex sulphide resources. Successful aqueous processing of those ores depends on the composition and mineralogy of raw materials and on their leaching behaviour. Systematic studies have been made to understand the oxidation behaviour and dissolution reaction mechanisms operative while pure copper, zinc and lead sulphide minerals and mixtures are treated by oxidative ammonia leaching. The kinetic expressions obtained for the dissolution of bulk concentrates are discussed as a case study. This paper also describes how one approaches the problem of designing the experiments for economic extraction of metal values from multimineral sulphides.

Hydrometallurgical processes can be exceptionally corrosive. Operating conditions associated with metal extraction and refining require materials of construction that can withstand process acids and acid chlorides at temperatures up to 90°C. Process vessels, piping, scrubbers and ducting can be rapidly compromised in these aggressive environments if not designed with great care. Design engineers are frequently challenged to find materials of construction that can stand up to hydrometallurgical processes especially for copper, zinc, nickel, cobalt and uranium. This paper will evaluate a variety of material solutions to determine which provide attractive economics for initial procurement as well as high durability for reasonable life cycle costs. The comparative cost and durability information presented is intended to enable design engineers and material specialists to choose materials of construction that will be the most beneficial for their projects.

The copper concentrate which contain sulphides (in principal chalcopyrite and pyrite) are oxidized in water, in presence of air at pressure about 4-7 atm and temperature of 125 - 135°C. The paper presents the decisive influence of hematite content to the efficiency of sulphides oxidizing. The proposed process is clean, efficient and can be applied also to concentrates with low copper content.

Extraction of zinc from oxide ores has received great attention during the recent decade. Extraction of zinc from zinc silicates and carbonates (mainly zinc oxide) was carried out through a hydrometallurgical process using sulfuric acid as leaching agent. Effect of different parameters on the zinc production by hydrometallurgical root including, leaching, solution purification and zinc recovery have been previously studied. However, there are few studies on the modeling of the particle size distribution of the particles during the leaching process. In the present study, the population balance model was applied to sulfuric acid leaching of a mixed silicate-carbonate zinc ore and particle size distribution during the leaching process was determined by solving the population balance equation. For determination of leaching kinetics, the leaching experiments were carried out in the temperature range of 30-60 °C using different initial particle size distributions. The change of particle size distribution during the leaching process was also determined experimentally. By numerically solving the population balance differential equation, coupled with the kinetic study of leaching process, using the method of characteristics, the particle size distribution during the process was determined. Simulations were performed with a) constant initial particle size and b) variable initial particle size in the variable particle growth rate. The proposed model shows a good agreement with the experimental data and has the ability to predict particle size distribution at various operating conditions and different kinetic schemes of leaching process.

There are many demands for clean energy sources to solve environmental issues. Solar cells are drawing interests as an alternative energy source. In our group, trizinc diphosphide, Zn3P2, and zinc-tin-diphosphide, ZnSnP2, are focused on as absorbers for solar cells. They are reported to be p-type semiconductors. Therefore, an n-type semiconductor is needed for solar cell devices. We consider to use ZnS thin film on the above absorbers. In Cu (In, Ga)Se2 (CIGS) solar cells, CdS was used for a n-type layer, but recently, ZnS-based materials are investigated as a Cd-free material. In addition, ZnS has a band gap of 3.68 eV [1], while the band gap of CdS is 2.42 eV [2]. This means that ZnS has an advantage for solar transmission, which is the required characteristic for the n-type layer. In this study, chemical bath deposition (CBD) is adopted for a process of the preparation of ZnS thin film. CBD is a process under atmosphere pressure and at relatively low temperatures. We investigate the morphology, the crystal quality and properties of semiconductor of ZnS thin film under various fabrication conditions, and discuss from a view point of thermodynamics.

Presented are the results of laboratory investigations on pressure leaching of copper sulphide concentrate produced at Lubin Concentrator (KGHM Polish Copper). Elevated content of base metals (Co, Ni, Zn, As, Ag, Pb) and of organic carbon was the specific feature of the concentrate. High content of cobalt and zinc, high content of Pb and As and the concentration of organic carbon exceeding 10 % make this concentrate both unique and hard to process by means of flash smelting. Alternative, hydrometallurgical method of Lubin concentrate processing is currently seriously considered. The effect of initial temperature, concentration of sulphuric acid, and oxygen partial pressure on leaching rate and metals recovery has been investigated. Leaching was performed in a stirred autoclave at temperatures within the range of 120 - 180 oC. Oxygen at partial pressure from 5 to 12.5 at was used as an oxidizing agent. It was found, that consid-erable increase of the leaching rate and recovery of copper can be observed with in-creasing both temperature and oxygen partial pressure. The recovery of cobalt and nickel was evidently lover. SEM microscopy examinations were performed for solid residue after leaching in order to explain observed leaching kinetics.

The cyanidation of gold ores containing copper produce effluent solutions that contain soluble copper-cyanide species. Copper and cyanide may be recovered from these effluents by a variety of methods including AVR, SART and direct electrowinning. In this work, the reagents LIX 7950 and LIX 78 were studied for the direct solvent extraction of copper cyanide from the effluent solution to produce a purified, concentrated solution for further treatment.

Industrial minerals such as kaolin, are too fine feature sizes, the presence of these colloidal systems is related in some way to strong distortions of the rheological behavior and the possibility of bringing these minerals mining and metallurgical processes. Colloidal systems in the mineral kaolinite, from La Unión, Antioquia, Colombia, were analyzed by the following physico-chemical techniques: Potential Z, DTP laser, tensiometry, high-resolution optical microscopy and rheometry. The object of these trials was to determine the stability of such suspensions as a function of surface activity under the concepts of colloidal and from the thermomechanic. The analysis results show that the excess concentration of colloidal minerals and nanoclay dispersions of kaolinite, generate strong effects of mechanical energy dissipation, but when concentrations are controlled under any scheme of concentration, the energy dissipation mechanics decrease if you look at the problem from a perspective of the DLVO theory.

The copper sulphide ore of Canaã dos Carajás - PA has been benefited by flotation that allows this concentration to reach 30%. The oxidized copper ore that overlaps with sulfide is stripped and stored until they feasibility studies of the recovery of copper by hydrometallurgical route are completed. Currently, we study the application of leaching in stirred tanks, piles and under pressure in autoclaves for the recovery of copper from both the oxidized and the sulfide. Wishing to study the behavior of oxidized copper ore by leaching with sulfuric acid assays were performed on a laboratory scale using for these tests, a sample of ore with content considered high compared to the average content found in the mine. Were collected important data for process optimization, such as acid concentration, solid-liquid ratio, stirring time and the equilibrium isotherm. In presenting this paper, we show the methods used in stirred tanks tests and results. That these data will be used for future testing of the same ore leaching in columns and cells.

Kaolins laundering is a process of selective leaching of minerals such as hematite (Fe2O3) and / or goethite (a-FeO (OH) attached to the surface of the white Kaolinite affect that characterizes it. The bleaching process involves using a strong reducing agent, as is the case, the sodium dithionite (Na2S2O4) acting on the surface of kaolinite hetatitico reducing iron and / or goethite (Fe 3 +) to ferrous iron (Fe +2) increasing the rate whiteness (% Whitness) of this mineral. This paper presents thermodynamic and kinetic study of electrochemical phenomena involved in the bleaching of kaolin ore, industrial and economic importance in Colombia. Laundering was studied four samples of Kaolin from Union Township in the department of Antioquia (Colombia) considered the main variable, the percentage of reducing agent (sodium dithionite, Na2S2O4), the processing time and the particle size in suspension. In the study of the physicochemical characteristics of the test samples, using X-ray diffraction (XRD), Scanning Electron Microscopy (SEM), Atomic Force Microscopy (AFM), Atomic Absorption Spectrophotometer, Particle Size Distribution (DTP ), X-ray fluorescence (XRF) and colorimetry. In the thermodynamic study of the process, potentiometric titration was used in the measurement of redox potential and atomic absorption spectrophotometry to determine the concentration of dissolved iron, finally, species of potential contributors were placed in a Pourbaix diagram and determined the chemical and electrochemical equilibria in the system. In a kinetic approach to the study were drawn polarization curves obtained by voltammetry and potentiostatic mode according to the Tafel law were found equilibrium potentials and mixed system , the Tafel constant, the density of exchange current (I0) and the current limit (IL).

This work involves a systematic study which provides insight into the electrochemical oxidation mechanisms of a chalcopyrite crystal. This was achieved through the identification and characterization of the potentiostatically formed chemical species on the surface of carbon paste electrodes with crystal chalcopyrite powder (CPE-CP) in a medium containing 1.7 M H2SO4, in the absence of oxidizing agents. The composition of the electrolyte, resulting from the imposition of a specific potential, was determined using a mercury film electrode (MFE) and anodic stripping voltammetry (ASV). The characterization of the chemical species was accomplished using cyclic voltammetry (CV) and X-Ray photoelectron spectroscopy (XPS). In the oxidation study of chalcopyrite, it was observed that this mineral dissolves quantitatively in the potential zone between 0.715 < Eanod ≤ 1.015 V vs. SHE. The existence of covellite (CuS) and a neutral sulfur species (S8o) in this potential zone, demonstrated by XPS, shows that these products are not responsible for the passivation phenomenon. The passivation is due to the slow decomposition of transient products, denominated passive non-stoichiometric polysulfides (Cu1-rFe1-sS2-t), which form at potentials below 0.715 V vs SHE. The physical and chemical properties of the neutral sulfur species S8o are different from those of elemental sulphur, permitting diffusion of the charged species. For this reason, the case for the electrochemical formation of elemental sulfur in this potential zone is discarded.

It is well recognized the importance of nano-structured materials in the present technological stage. Due to their unique properties these materials can be used in a large number of applications. It is also well stablished that materials properties can be severely altered by giving them a nano-structured condition. In that sense, some materials can even be designed in order to fulfill the need for high performance in some specifics functions. One example is the growing interest in nanocomposites, in which a very fine dispersion of a ceramic phase in a metal matrix will significantly improve the material properties. In view of that, extensive studies have been carried out on a variety of materials such as alloys and different types of composites. Recently, our group at the Department of Material Science and Metallurgy of Rio de Janeiro Catholic University has developed an alternative chemical based synthetizing route to obtain nano-structured materials. This approach can be applied to the production of oxides, metals, composites and alloys. In the last two cases with a high level of homogeneity. Recently, the authors have developed a novel chemical and powder metallurgy method for in-situ formations of Cu-Al2O3 and Ni-Al2O3 nanoscale composites by decomposition of their mixed nitrate solutions, to coform the nano oxides, followed by preferential reduction of CuO or NiO by hydrogen at very low temperature. Studies carried out by the authors on the kinetics of reduction of such fine oxides indicated that under low partial pressure of hydrogen (0.25 atm) in argon, the oxides of Ni and Cu can be reduced completely, in a low temperature range of 523 to 623 K. The composites containing nanosize metal-metal oxide particles, have been found to be quite homogeneous in nature. In view of this, Cu-Ni alloy was produced by mixing the aqueous nitrate solutions of copper and nickel, followed by decompositions of their nitrates to their mixed oxides and subsequent low temperature hydrogen reduction. The alloy powder has been characterized and cold pressed and sintered to determine the density, hardness and workability. In that context, the purpose of the present work is to address the fundamental aspects of the synthesis procedure, emphasing the thermodynamics background of the two steps involved. Also, the work aims to illustrate the outcome, by presenting experimental conditions and providing relevants characterization of the obtained nano-materials, by means of electron microscopy and microanalytical techniques. Examples are given in terms of individual and co-formed oxides and composites.

An experimental investigation and optimization of the parameters of Sc extraction from a solution having the composition 0.024-0.034 g/l Sc(III); 8,6-11,9 g/l Fe(III); 0.11-0.19 g/l Ti(IV); 0,09-0,28 g/l Zr(IV), 0,08-0,155 g/l U(VI) and 80-130 g/l H2SO4 by means of a (0,1-0,3 mol) di(2-ethylhexyl)phosphoric acid (DEHPA) solution in kerosene together and reextraction of scandium by a 1-4% sodium fluoride solution accompanied by the production of rough scandium fluoride concentrate are conducted. It is established that the indicators of Sc extraction when performed in stages are determined by the DEHPA content and the residual concentration of Sc in the reverse organic phase. Mathematical models of the extraction and reextraction of scandium are proposed, and through use of these models nomograms are obtained that make it possible to determine the maximally possible extraction of Sc and maximally admissible phase ratio in the operations. It is shown that the extraction and content of Sc in rouch scandium fluoride concentrate are determined by the NaF concentration in the reextractant and the phase ratio in the operations. The optimal and feasible conditions for reextraction of Sc are determined in a combined review of the different mathematical models. It is established that the residual content of Sc in a NaF solution is determined by the concentration of Sc in the saturated organic phase and of U(VI) in the initial solution.

Hafnium metal is widely used in nuclear industry for fabrication of neutron absorbing elements (control rods) and in metallurgy as a component of superalloys. The magnesium metallothermic reduction of hafnium tetrachloride is the best method for the metal production in terms of process conditions, purity of the product, and economic characteristics. However, synthesis of HfCl4 is a separate task. Direct reaction of hafnium oxide with chlorine gas is thermodynamically impossible. For this reason, chlorination is performed in the presence of a reducing agent such as carbon. In the present work thermodynamic analysis of hafnium oxide chlorination by gaseous chlorine in the presence of carbon was performed at temperatures between 450 and 1000°C. It was found that the main reaction products are gaseous hafnium tetrachloride, carbon monoxide and carbon dioxide. Rising temperature leads to increasing carbon monoxide partial pressure, while that of CO2 decreases. In the above temperature range, the variations of enthalpy, entropy, and Gibbs energy change of the chlorination reaction were calculated and the composition of the gas phase under the conditions of thermodynamic equilibrium was determined. However, the content of the components in real gas–vapor mixtures may significantly differ from the calculated equilibrium values due to kinetic limitation. A series of thermogravimetic experiments was performed under controlled atmosphere using a custom-built set-up based on a Mettler AT20 balances connected to a computer, and intended for the continuous measurement of the sample weight in gases (including aggressive) or in liquids. The kinetics of chlorination of hafnium oxide in the presence of carbon was investigated at temperatures ranging from 600 to 950°C. It was found that below 700°C the process is limited by the chemical reaction on the surface of nonporous solid spherical particles resulting in the formation of a volatile product. Above 700°C, the rate of the process is determined by the mass transfer of the gaseous substances. From the dependences of the logarithm of the reaction rate on the reciprocal temperature the values of the apparent activation energy (Ea) of the chlorination process were calculated. The value of Ea at 600–700°C equals to 65±5 kJ/mol, while at higher temperatures, 700–950°C, it is several times lower and equals to 14±7 kJ/mol. This confirms that at 600-700°C the chlorination proceeds in the kinetic mode, and the process speed is limited by the chemical reaction rate. At 700–950°C the rate of the chemical reaction is significantly higher and the slowest stage is the diffusion of Cl2 and CO to the reaction interface or removal of gaseous HfCl4.

A mining company, located in the southwest of Mexico, produces gold from an ore with 5.5 g of Au/t. Besides the gold, the ore contains also 0.2 – 0.33 % copper. Although the proportion of copper oxides is not so high, the leaching solution becomes saturated with copper (the process operates with zero discharges) and it is necessary to bleed part of the solution to the tailings pond in order to maintain the efficiency of the process. Approximately 5 % of the copper dissolves in the cyanide solution, resulting in a gold precipitate with high copper content (an average of 14 kg Cu/kg Au). In order to purify and commercialize the precipitate, it is smelt, but only 40 % of the copper can be removed by this method, the remaining copper being eliminated by nitric acid digestion. The silver values dissolved during the latter stage must be further recovered by a chemical precipitation using NaCl as a precipitating agent. Today´s high prices on precious metals have allowed the economic operation of this process; Nevertheless, there exist some strong environmental issues related to this practice, among them the following could be mentioned: A) great amount of gases generated, b) corrosive and irritating atmosphere, c) long times of treatment (20 hours per lot of 800 kg) and d) high quantity of silver values in the solution that requires to be recovered.

This work presents a consideration of advanced technologies which would be employed in the mine waters treatment with an aim of removal and recovery of metals contained in mine waters and AMDs. A process simulating of AMDs generation from the residual overburden is established and tested with an aim of speeding up the self-leaching phenomenon which now occurs naturally. In addition a possibility of a hydrometallurgical process implementation is described. Considered technology should ensure an environmentally sustainable and economically viable process for the purification and neutralization of AMDs based on the metal extracted and produced from such sources. Possibility for technology implementation, with a brief economic analysis, is given based on the performed tests of leaching and solvent extraction for contained copper in AMDs.

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