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Acidic Leaching with Chlorate as Oxidising Agent to Extract Mo and Re from Molybdenite Flotation Concentrate in a Copper Plant

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Abstract

The technology of molybdenum extraction from molybdenite concentrate by using potassium chlorate (KClO3) or sodium chlorate (NaClO3) has been investigated. The results show that leaching time, leaching temperature, agitation speed, oxidizer type, potassium or sodium chlorate, and hydrochloric acid concentration have significant effect on the molybdenum extraction efficiency. Optimum process operating parameters were established as follows: 4 hrs, hydrochloric acid concentration: 35%, solids ratio: 5%, temperature: 65-70°C, agitation speed: 600 rpm, the mass of potassium chlorate and sodium chlorate: 25 g. Under these experimental conditions, the extraction of molybdenum and rhenium were obtained about 85% and 100%, respectively.

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... Here, the leaching method was applied to separate Mo and Cu in bulk concentrates from the Geumeum mine as an alternative method. Since the 1970s, MoS 2 could be reportedly oxidized in both acidic and alkaline solution by hypochlorite and chlorate, so their application to selective leaching of Mo from Mo/Cu complex ores has been much studied [16][17][18][19][20][21]. An electro-oxidation method has also been developed and reported to be effective in selective leaching Mo from Mo concentrates [22][23][24][25]. ...
Article
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This paper proposes selective leaching of molybdenum from Mo/Cu complex bulk concentrates in a 5 M NaCl solution using the electro-oxidation method. Here, the effects of several factors such as pH, pulp density, current density, and temperatures were investigated. A higher leaching yield of Mo increased with increasing pH from 5 to 9 and decreased with increasing pulp density from 1 to 10%. A rise in current density did not help enhance Mo, and the elevating temperature did not always result in a higher leaching yield. Application of ultrasonic led to higher leaching yield of Mo. Ninety-two percent of leaching yield was obtained upon leaching of Mo in 5 M NaCl at 25 °C, pulp density of 5%, and the current density of 0.292 A/g under ultrasonic irradiation with a power of 27 kW. The resultant residue mainly consisted of chalcopyrite.
... Since the disadvantages of oxidizing roasting are well known (e.g., release of sulfurous gases into the atmosphere), the latest research has been focused on using the hydrometallurgical techniques for processing molybdenite concentrates. Such techniques include atmospheric [6] and autoclave [7] leaching with nitric acid, oxygen leaching under pressure [8,9], sodium chlorate [10,11] and hypochlorite [12] leaching, as well as bioleaching [13][14][15]. ...
Article
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The paper discusses methods of processing copper-containing molybdenite concentrates. A process flow diagram based on roasting with sodium chloride (or carbonate) and subsequent water leaching is presented. The chemistry of sulfide interaction with salt additives is described. The effect of the roasting temperature on copper and molybdenum distribution among the process products was studied. It was found that during roasting, 75 to 80% of molybdenum is converted from the sulfide into trioxide form, while copper predominantly forms water-soluble compounds and, thus, enables extraction by water leaching. It was established that the cake obtained as a result of water leaching meets the requirements for raw materials used for ferromolybdenum smelting.
... So the influence of the agitation speed on molybdenum leaching rate has relationship with the leaching temperature, indicating that the leaching process was mainly controlled by chemical reaction rate and mass transfer rate. This rule is different from some previous studies, where the leaching process was controlled by diffusion process due to molybdenite concentrate was leached directly without roasting process [5,34]. The optimization value of agitation speed was maintained at 500 rpm at 343.15 K. ...
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A technique for leaching molybdenum from low-grade roasted molybdenite concentrate was proposed by the aqueous solution containing sodium chlorate and sodium carbonate. The effect of molar ratio of sodium chlorate to molybdenum, leaching time, liquid–solid ratio, leaching temperature, sodium carbonate concentration and agitating speed on leaching rate of molybdenum was studied. The experimental results showed that the temperature and concentration of sodium carbonate are key factors to influence the leaching efficiency of molybdenum, and leaching rate achieves above 98% when the molar ratio of sodium chlorate to molybdenum is up 13.5, the temperature is 343.15 K, the agitating speed is 500 rpm, the liquid–solid ratio is 10:1 and the concentration of sodium carbonate is above 10 g/L. The leaching process was mainly controlled by chemical reaction and mass transfer. The leaching time is shorter and the heavy metal content in leaching aqueous solution is lower in basicity than those in acid situations.
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Kinetics of atmospheric leaching molybdenum from metalliferous black shales by air oxidation in alkali solution was investigated. The effects of temperature and sodium hydroxide concentration on the rate of molybdenum leaching were studied. The results of the kinetic analysis of the leaching data for various experimental conditions indicated that the reaction is controlled by diffusion with the activation energy of 15 kJ/mol higher than 65 °C, and controlled by chemical reaction with the activation energy of 57 kJ/ mol lower than 65 °C and molybdenum recovery of about 90% can be reached in about 30 min. In addition, increasing concentration of sodium hydroxide has a positive effect on the dissolution of black shales and reaction order with respect to sodium hydroxide concentration were approximately1.9 lower than 65 °C and 1.8 higher than 65 °C, respectively.
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Article
The technology of molybdenum extraction from molybdenite concentrate by using sodium chlorate has been investigated. The results show that leaching time, liquid-to-solid ratio, leaching temperature, agitation speed, sodium chlorate and hydrochloric acid concentration have significant effect; optimum process operating parameters were established as follows: time: 4 h; hydrochloric acid concentration: 20%; liquid-to-solid ratio: 10:1; temperature: 70 °C; agitation speed: 500 rpm; and the molar ratio of NaClO3 to MoS2: 3.21. Under these experimental conditions, the extraction of molybdenum is greater than 98%. The mixture of 30% triotylamine and 20% sec-Caprylic alcohol diluted kerosene oil is an effective extractant of molybdenum; the molybdenum extraction efficiencies in the examined conditions were about 99.6% under conditions of O: a ratio of 1:1, hydrochloric acid 60 g/L for 10 min at room temperature. Stripping of molybdenum to aqueous phase was efficient when 18% ammonia liquor were applied. The molybdenum stripping efficiencies in the examined conditions were about 99.5% under conditions of O: a ratio of 1:1 for 10 min at 40 °C.
Article
This study aims to refine an indigenous off-grade molybdenite [41.5% Mo] to one of high-grade MoS2 for industrial applications. Investigations were carried out on the removal of the oxide/silicate gangue and the base metal sulphide associations by their selective dissolution in two acids, namely, HCl or HF deployed singly, sequentially or in the mixed mode. Under optimum conditions practically all the oxide and silicate gangue and 90% of the metallic impurities were removed. Starting from the low-grade molybdenite concentrate a refined molybdenite (97.8% MoS2) was made whose composition compares well with that of a technical-grade MoS2 of the Climax Molybdenum Company.
Article
A novel technology characterized by higher recovery of vanadium and which was environmentally-friendly was developed to recover vanadium from stone coal. Vanadium in stone coal could be leached by NaOH solution after roasting stone coal at 850°C for 3h. H2SO4, Mg(NO3)2 and ammonia were employed, respectively, in two steps to remove the impurities of Si and Al from the leach liquor. After extracting vanadium from the leach liquor with 10vol% N235, 20vol% secondary octyl alcohol and 70vol% sulfonated kerosene, 1.5mol/L NaOH was used as a stripping agent to strip vanadium from extracting solution. Adding 80g/L NH4NO3 to the stripping solution at 30–40°C and pH 7.5, vanadium could be crystallized as ammonium metavanadate. Roasting ammonium metavanadate at 540°C for 1h, the purity of V2O5 met the standard specification. The total recovery of vanadium reached 67.39%, which was higher than the classical technology.
Article
The oxidation of molybdenite by sodium dichromate in sulphuric acid was investigated. The effects of dichromate ion concentration, sulphuric acid concentration, temperature and the duration of the treatment on the stoichiometry of the reaction, were examined. It was found that under all experimental conditions applied the reaction proceeded according to the same stoichiometry. The kinetics of this reaction were also investigated, whereby the effects of temperature, reactant concentrations, particle size and agitation speed on the rate of the oxidation of molybdenite by sodium dichromate, were investigated. It was established that molybdenite oxidation by sodium dichromate is a chemically controlled reaction (Ea = 68 kJ/mol). In addition, the effect of pyrite and chalcopyrite on kinetics of this reaction was investigated.
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A new Cu(II) ion-selective PVC membrane sensor based on 6-methyl-4-(1-phenylmethylidene)amino-3-thioxo-1,2,4-triazin-5-one (MATTO) as an excellent sensing material was developed. The electrode exhibits a Nernstian slope of 29.2±0.4 mV per decade over a very wide concentration range between 1.0×10−1 and 1.0×10−6 M, with a detection limit of 4.8×10−7 M (30.5 ng/mL). The sensor possesses the advantages of short conditioning time, fast response time (<10 s), and especially, very good selectivity towards transition and heavy metal, and some mono, di and trivalent cations. The proposed electrode was successfully applied to the determination of copper in wastewater of copper electroplating samples and as an indicator electrode in potentiometric titration of Cu(II) ions with EDTA.
Article
Extraction of molybdenum and vanadium from ammonia leaching residue (main chemical composition: 2.05% Mo, 0.42% V, 65.6% Al2O3 and 10.7% SiO2) of spent catalyst was investigated by roasting the residue with soda carbonate, followed by hydrometallurgical treatment of the roasted products. In the roasting process, over 91.3% of molybdenum and 90.1% of vanadium could be extracted when a charge containing a sodium carbonate to spent catalyst ratio of 0.15 was roasted at 750 °C for 45 min and the roasted mass was leached with water (liquid to solid ratio of 2) at 80–90 °C for 15 min. After the purification of leach liquor, an extraction solvent consisting of 20 vol.% trialkylamine (N235, commercialized in China) and 10 vol.% secondary octyl alcohol (phase modifier) dissolved in sulfonated kerosene was used to extract molybdenum and vanadium in leach liquor. 10 wt.% ammonia water was used as stripping agent. Adding 30 g/l NH4NO3 to the stripping solution and adjusting the pH to 7–8.5, over 99% of vanadium can be crystallized as ammonium metavanadate. Over 98% of molybdenum can be crystallized as ammonium polymolybdate when pH is between 1.5 and 2.5 (pH is adjusted by HNO3). Ammonium metavanadate and ammonium polymolybdate were calcinated at 500–550 °C, the purity of MoO3 and V2O5 was 99.08% and 98.06% respectively. In the whole process, 88.2% of molybdenum and 87.1% of vanadium could be achieved. The proposed roasting, leaching and separation steps give a feasible alternative for the processing of ammonia leaching residue of spent catalyst and can be applied in the comprehensive utilization of low grade molybdenum ores.
Article
This study evaluates different bioleaching treatments of a molybdenite concentrate using mesophilic and thermophilic bacterial cultures. Further studies on the chemical leaching and the electrochemical behavior of the MoS2 concentrate were carried out. Bioleaching tests showed a progressive removal of chalcopyrite from the molybdenite concentrate with an increase in temperature. Chemical leaching tests support the idea of an indirect attack of the concentrate. Electrochemical tests indicate that chalcopyrite dissolution is favored when molybdenite is present. Therefore, this type of bioleaching treatment could be applied to purify molybdenite flotation concentrates by selectively dissolving chalcopyrite.
Article
The leaching kinetics of a low grade-calcareous sphalerite concentrate containing 38% ankerite and assaying 32% Zn, 7% Pb and 2.2% Fe was studied in HCl–FeCl3 solution. An L16 (five factors in four levels) standard orthogonal array was employed to evaluate the effect of Fe(III) and HCl concentration, reaction temperature, solid-to-liquid ratio and particle size on the reaction rate of sphalerite. Statistical techniques were used to determine that pulp density and Fe(III) concentration were the most significant factors affecting the leaching kinetics and to determine the optimum conditions for dissolution. The kinetic data were analyzed with the shrinking particle and shrinking core models. A new variant of the shrinking core model (SCM) best fitted the kinetic data in which both the interfacial transfer and diffusion across the product layer affect the reaction rate. The orders of reaction with respect to (CFe3+), (CHCl), and (S/L) were 0.86, 0.21 and − 1.54, respectively. The activation energy for the dissolution was found to be 49.2 kJ/mol and a semi-empirical rate equation was derived to describe the process. Similar kinetic behavior was observed during sphalerite dissolution in acidic ferric sulphate and ferric chloride solutions, but the reaction rate constants obtained by leaching in chloride solutions were about tenfold higher than those in sulphate solutions.
Article
Extraction of molybdenum and rhenium values from low grade Indian molybdenite concentrate was investigated by roasting the concentrate in the presence of slaked lime and soda ash, followed by hydrometallurgical treatment of the roasted products. In the lime roasting process, molybdenum recoveries of around 99% were achieved when a charge containing a slaked lime to concentrate ratio of 0.875 was roasted at 550°C for 1 h and the calcine was leached twice with 1 M H2SO4 at 80–90°C for 2 h. In the soda ash roasting process, over 99% of the molybdenum could be extracted when a charge containing a sodium carbonate to concentrate ratio of 1.05 was roasted at 650°C for 1 h and the roasted mass was leached with water at 80–90°C for 2 h. A carbon adsorption technique, involving selective adsorption of molybdenum on activated charcoal from the leach solution (at a pH of 2.0), followed by desorption with ammonia, was adopted to prepare high purity MoO3 product. Rhenium recoveries of the order of 74% were obtained by water leaching of the lime-roasted calcine at 80–90°C for 1 h.
Article
Sodium chloride-water vapour roasting of calcined catalyst for 2 h at 850°C (water vapour pressure 0.253 bar, N2 atmosphere) and subsequent leaching of the roasted catalyst with water at its boiling point for 1 h dissolved 81.85% V and 81.78% Mo and minor amounts of other constituents. The separation of V from the leach solution has been carried out efficiently by precipitation (as NH4VO3), liquid-liquid extraction using di(2-ethylhexyl)phosphoric acid (D2EHPA) and tri-n-octylamine (TOA) and subsequent strippings and precipitations. The overall recoveries of V and Mo from the waste catalyst are 75.5 and 77%, respectively.
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Copyright © Taylor & Francis Group, LLC ISSN: 0149-6395 print / 1520-5754 online DOI: 10.1080/01496395.2015.1059348 REFERENCES 1. Lasheen, T. A.; El-Ahmady, M. E.; Hassib, H. B.; Helal, A. S. (2015) Molybdenum metallurgy review: Hydrometallurgical routes to recovery of molybdenum from ores and mineral raw materials. Mineral Processing and Extractive Metallurgy Review: An International Journal, 36 (3): 145– 173.
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Electrometallurgy: Past, Present, and Future
  • F Habashi
Habashi, F. (1997) Electrometallurgy: Past, Present, and Future; Proceedings of the 1997 TMS Annual Meeting: Orlando, FL, USA, pp. 351-366.