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Acidic Leaching with Chlorate as Oxidising Agent to Extract Mo and Re from Molybdenite Flotation Concentrate in a Copper Plant
Separation Science and Technology
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.
... Hydrometallurgical processes for low-grade and complex ores have received more attention with the development and advancement of technology. Pressure leaching (Dresher et al., 1956;Khoshnevisan et al., 2012;Mirvaliev and Inoue, 2001;Smirnov et al., 2010), bioleaching (Abdollahi et al., 2015b;Nasernejad et al., 1999;Olson and Clark, 2008;Zamani et al., 2005), leaching using nitric acid (Peters, 1976;Vizsolyi and Peters, 1980) and HCl/HF (Kumar et al., 2007), leaching in alkaline environments such as sodium hydroxide (Zhao et al., 2009), potassium hydroxide (Dresher et al., 1956), and leaching by oxidizing agents such as hydrogen peroxide (Aracena et al., 2019;Lasheen et al., 2013;Moazemi Goodarzi et al., 2014), sodium hypochlorite (Bhappu et al., 1963;Hesami et al., 2022;Liu et al., 2011;Warren et al., 1977;Warren and Mounsey, 1983;Youcai et al., 2011), sodium chlorate (Abdollahi et al., 2015a;Li et al., 2021), sodium dichromate (Antonijević and Pacović, 1992) was used to extract molybdenum from molybdenite concentrates. These processes have problems such as the high cost of investing in the oxidative pressure process, low extraction in the bioleaching process, the difficulty of controlling the leaching process, and the high cost of leaching agents (Bhappu et al., 1963). ...
The kinetics of molybdenum and copper extraction from Sarcheshmeh molybdenite concentrate (Kerman, Iran) was studied by electroleaching process in batch mode. The effects of several critical parameters including sodium chloride concentration, pH, temperature, time, operating voltage, and the sodium carbonate addition on the molybdenum and copper extraction from the concentrate were investigated. Results showed that the molybdenum extraction significantly increased by increasing sodium chloride concentration, temperature, time, operating voltage, and sodium carbonate addition during the process. While with increasing pH, the molybdenum extraction decreased. Maximum molybdenum extraction (89%) was obtained at the sodium hypochlorite concentration of 30 g/L, the pH of 9, the temperature of 50 °C, operating voltage of 5 V, and reaction time of 4 h. Copper extraction was determined 17.5% in 30 min, which decreased to 3.9% over 8 h. It was found that carbonate ions added to the electroleaching process had a significant effect on molybdenum extraction. Also, carbonate ions might prevent hypochlorite decomposition by stabilizing copper (III) ions. The catalytic compounds of copper precipitate simultaneously with the catalytic decomposition of the hypochlorite. Kinetic studies have shown that the dissolution of molybdenite concentrate is described by a shrinking core model controlled by a chemical reaction. The activation energy for the electroleaching process of molybdenite concentrate was determined to be 29.4 kJ/mol. It seems that electroleaching is an efficient, green, and selective process for treating molybdenite concentrates.
... 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]. ...
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]. ...
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. ...
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.
Molybdenum (Mo) and rhenium (Re) are strategic key metals for the development of modernization. In nature, 99 wt.% of Re is associated with molybdenite minerals via the form of isomorphism, and the element abundance of Mo (about 1.1–1.5 ppm) and Re (about 0.0004–0.0026 ppm) is tremendous in the continental crust. These congenital characteristics make it difficult to separate Mo and Re. In this work, the Mo and Re resource features and products consumption are briefly summarized, and the pre-separation of Mo and Re from minerals or secondary alloy scraps, and the deep separation of the dissolved Mo and Re from solutions are systematically reviewed. Pyrometallurgy approaches are suitable for the treatment of high-grade Mo concentrates, however it lacks vast energy consumption and SO2 emission. Hydrometallurgical process is appropriate to deal with low-grade Mo concentrates, but leaching efficiency of Mo is low and numerous expensive oxidants are consumed. As for deep separation of Mo and Re from the solutions, selective chemical precipitation, ion exchange, and solvent extraction are commonly used in industry to separate Mo and Re. Due to separation efficiency, Mo and Re mixture solutions with low concentrations generally lower than 0.5 g/L are concomitantly generated after ion exchange and solvent extraction. Selective adsorption is a prospective method to separate Mo and Re from ultra-low concentration solutions, while the existing preparation technologies of high-selectivity adsorbents are characterized as complex and expensive. Developing facile and low-cost preparation of adsorbents is a promising research orientation for clean and deep separation of Mo and Re from solutions.
Hydrometallurgical extraction of molybdenum from sulfidic molybdenite concentrates is an important issue from technical, economic, and environmental viewpoints. The influence of mechanical activation on the hypochlorite leaching of a molybdenite concentrate (Kerman, Iran) were studied in a stirred reactor at batch mode. The effects of several critical parameters, including pH, temperature, reaction time, NaClO concentration, and liquid to solid (L/S) ratio on the rate and extent of molybdenum extraction from the concentrate were investigated. Maximum molybdenum extractions, 98% and 95%, obtained for the mechanically activated and nonactivated concentrates, respectively after 60 min at 50°C, stirring speed of 500 rpm, L/S of 333 mL/g, pH of 9, and using 40% NaClO solution. Specific surface area and particles size measurements in conjunction with FE-SEM images showed that improvement in the extraction after mechanical activation could be due to increasing the surface area and changing the crystal structure of the concentrate. Mechanical activation had a significant effect on molybdenum extraction after 2 min, (13% difference in extraction); however, after 60 min, molybdenum extractions from mechanically activated and nonactivated samples were 97% and 92%, respectively. It was found that mechanical activation has a significant effect on the leaching of molybdenite in short reaction times.
In view of the difficulty of separating molybdenum over iron from a concentrated sulfuric acid solution, a preliminary study of mixed solvent systems for the separation of Mo/Fe in sulfuric acid solutions was carried out. Three mixtures of PC88A and TBP, PC88A and TRPO, and PC88A and N235 as solvents were investigated in this work. Antagonistism was found to be present in the extraction of Fe (III) with these solvent mixtures. The results indicate that by increasing the percentage (vol%) of TBP, TRPO or N235 in the mixed extraction systems, Mo extraction increases somewhat, while the percentage of Fe extraction decreases sharply. Correspondingly, the separation factors of Mo over Fe (βMo/Fe) were improved from 5.1 for PC88A alone to 284, 2520 and 9100 for the addition of 30 vol% TBP, 20 vol% TRPO and 30 vol% N235, respectively. It is evident that the antagonistic effect is favorable for the efficient separation of Mo over Fe in acidic solutions. In addition, the mechanism of hindering the extraction of iron is simply discussed in this paper.
Water solution polymerization was adopted to prepare perrhenate ion imprinted polymers in this work, and preparation conditions, such as molar ratio of acrylic acid, N,N - methylene double acrylamide, hydrogen peroxide and ascorbic acid to N-vinyl pyrrolidone, polymerization time and temperature on adsorption separation property were investigated and optimized using adsorption quantities and separation degree as indexes. To enhance the adsorption selectivity, adsorption properties of samples prepared under different conditions were conducted in a mixture aqueous solution of ammonium perrhenate and potassium permanganate with the molar ratio of 1:1. Results showed that as the molar ratio of N,N - methylene double acrylamide, acrylic acid, hydrogen peroxide, ascorbic acid to N-vinyl pyrrolidone were 2.0, 1.9, 0.03, and 0.003 respectively, the reaction temperature and time were 35 °C and 24 h, adsorption quantities and separation degree arrived at 0.037 mmol/g and 2.31, respectively, while the adsorption capacities and separation degree of non-imprinted polymers were only 0.009 mmol/g and 0.22. Besides, study was also discovered that the prepared samples presented an exciting reusable stability. Based on samples prepared under optimal conditions, the structure was characterized by Fourier transform infrared spectroscopy, scanning electron microscope, transmission electron microscope, BET surface area and thermogravimetric analysis using non-imprinted polymer prepared under the same conditions as a control. Finally, the adsorption thermodynamics and kinetics were studied, and the kinetics documented that the adsorption of perrhenate ion imprinted polymers was a zero-order model process. The thermodynamics presented the adsorption was more suitable to be expressed by Langmuir model, and Scatchard analysis revealed that heterogeneous binding sites were formed in IIP.
The roasting of molybdenum concentrates using NaCl has been carried out. During the roasting process, Mo is present both as a gas (25–40%) and as solid residue (60–75%). Molybdenum chloride and oxychloride condensate both dissolve in the alkaline solution with molybdenum recovery up to 100%. The results of the study show that 30–32% of molybdenum (sodium molybdate) in the roasted residue is water-soluble. This paper proposes a methodology for leaching roasted residue in two stages: first by water, then by alkaline solutions. Using this methodology, molybdenum recovery from the roasted residue exceeds 99%. It was determined that up to 40% of the sulfur contained in the concentrate remains in the roasted residue after the roasting process in the form of sodium sulfate, which enables a significant reduction of SO2 generation. The results were achieved under optimal roasting conditions: a roasting temperature of 420–430 °С, NaCl ratio of 100% (from concentrate weight), and 90 min roasting time.
In this work, NaClO3 is demonstrated for the first time as an efficient and versatile oxidant in the catalytic C−H activation reaction. By using sodium chlorate as the oxidant, a highly regioselective cobalt-catalyzed oxidative annulation of aryl sulfonamides with alkynes has been developed and can be extended to the annulation of benzamide.
A novel technology characterized by higher recovery of molybdenum and production of high value added ammonium molybdate was developed to recover molybdenum from low grade Ni–Mo ore. The brief flow of the novel technology includes roasting, Na2CO3 and NaOH mixed alkali leaching, ion exchange enrichment molybdenum, purifying the desorption liquor, removal of vanadium and crystallization ammonium molybdate. The total recovery of ammonium reached 89.06%, which was higher than the current applied technology and the ammonium molybdate met standard specification.
With the vigorously growing demand of the steel industry, oil and gas industry, corrosion resistance alloys, cast iron, and catalyst industries, high-grade molybdenum ores are being exhausted gradually in the world. Thus, much attention have been drawn to the recovery of molybdenum from low-grade molybdenum ores in recent years. With the increasingly stringent environmental requirements, the shortcomings due to SO2 emission in the roasting process of traditional technology becomes obvious. This review outlines metallurgical processes for molybdenum production from various resources, particularly focusing on recent developments in direct hydrometallurgical and recovery processes to identify potential sources of molybdenum products and by-products such as uranium which can be economically produced.
Several methods have been extensively reviewed for molybdenum separation and purification from solution which are potentially applicable to leach solutions of molybdenum ores and raw materials. The main methods include solvent extraction, ion exchange, membrane-based separation, and precipitation. Solvent extraction is highly selective for recovery of molybdenum and the most promising method recommended for future research and development. Membrane-based separation is the next preferred method for selective extraction of molybdenum, purification of molybdenum solutions, or co-recovery of other valuable metals. Ion exchange offers useful means for purification and/or co-recovery of other base metal impurities, although the scale of application of ion exchange in the industry is limited.
Mixed mesophilic and extreme thermophilic bioleaching were evaluated to remove copper from the molybdenite concentrate. Bioleaching tests were carried out in shake flasks and in a 50-L bioreactor. The shake flask tests were performed with different inoculum size, solids density, pH, and temperature in order to identify optimum conditions. The highest amount of copper elimination, 75% was obtained with extreme thermophilic microorganisms (at 12% inoculation, 10% solids, 65 °C and a pH of 1.5). The highest copper elimination by mesophilic microorganisms was 55% (at 12% inoculation, 5% solids, 30 °C at pH 2). The optimum conditions in shake flask tests were applied to 7 days batch tests in a 50-L bioreactor. Extreme thermophilic experiment gave the best copper elimination of 60% (at 12% inoculation, 10% solids, 65 °C and pH 1.5). Mesophilic test removed 50% of the copper (at 12% inoculation, 10% solids, 35 °C at pH 2).
The application of the response surface methodology and the central composite design (CCD) technique for modeling and optimization of the influence of some operating variables on copper, molybdenum and rhenium recoveries in a bioleaching process was investigated. Three main bioleaching parameters, namely pH, solid concentration and inoculum percent, were changed during the bioleaching tests based on CCD. The ranges of the bioleaching process variables used in the design were as follows: pH 1.46–2.14, solid concentration 0.95%–11.05%, and inoculum percent 1.59%–18.41%. A total of 20 bioleaching tests were carried out by the CCD method according to software-based designed matrix. Empirical model equations were developed according to the copper, molybdenum and rhenium recoveries obtained with these three parameters. Model equations of responses at the base of parameters were achieved by using statistical software. The model equations were then individually optimized by using quadratic programming to maximize copper, molybdenum and rhenium recoveries individually within the experimental range. The optimum conditions for copper recovery were pH 1.68, solid concentration 0.95% and the inoculum 18.41% (v/v), while molybdenum and rhenium recoveries were 2.18% and 24.41%, respectively. The predicted values for copper, molybdenum and rhenium recoveries were found to be in good agreement with the experimental values. Also jarosite formation during bioleaching tests was also investigated.
Molybdenum is found in low quantities in copper porphyries as the mineral molybdenite. In the processing of these ores, a bulk copper-molybdenum concentrate is produced, from which molybdenite is separated by differential flotation. However, complete separation of molybdenum from copper by flotation is difficult, and molybdenum concentrates produced by this route require a further step of chemical purification to reduce their copper content. This paper is concerned with a novel process of chemical purification of molybdenite concentrate containing 3.4% Cu and 3.04% Fe where chalcopyrite is the main impurity. The process involves a sulfidation of the molybdenite concentrate with S2(g) at 380 °C followed by a leaching stage with H2SO4–NaCl–O2. The leaching of the sulfidized molybdenite concentrate for 90 min at 100 °C dissolved 96% of the copper which yielded a final concentrate with less than 0.2% Cu. Molybdenite dissolution was negligible in this leaching media, and less than 20% Fe was dissolved. The kinetics of copper dissolution from the sulfidized concentrate was analyzed using the shrinking core model for chemical reaction control. Activation energy of 52.5 kJ/mol was calculated for the temperature range of 70 °C–100 °C.
Spent hydrodesulphurisation (HDS) catalysts from petroleum refineries, mostly constituted by aluminium (Al), molybdenum (Mo), nickel (Ni) and cobalt (Co), are classified as hazardous wastes and disposal is restricted by environmental regulations; thus, the recovery of valuable metals is the most attractive option in the processing of these residues. Selective Mo leaching from two spent HDS (Ni–Mo or Co–Mo) catalysts, roasted at 500 °C, was studied using three different approaches: conventional, ultrasound- and microwave-assisted methods. For the Ni–Mo or Co–Mo catalysts, 77% and 84% of Mo extraction, respectively, was achieved when a conventional (10 g/L NaOH, 80 °C, 120 min, S/L = 50 g/L) approach was used. For both catalysts, ultrasound-assisted leaching (pulse 0.5 s and 20% amplitude) with 10 g/L NaOH and S/L = 50 g/L led to 66% Mo dissolution after 10 min. Under microwave-assisted (four cycles of 30 s) conditions, Mo leaching reached 89% and 91% for Ni–Mo and Co–Mo catalysts, respectively. The latter conditions represent the best compromise between Mo extraction and selectivity, since Al dissolution was between 6% and 9%.
A method for the extraction of both Pb and Mo from molybdenite flotation tailings associated with galena is described. The flotation tailings were pre-roasted and then alkaline leached. The effect of temperature, initial concentration of sodium hydroxide solution, alkali to total calcine mass ratio, leaching time and stirring speed on the leaching of Pb and Mo were systematically investigated by experiment. Optimum conditions were identified as: pre-roasting at 600 °C for 1 h, alkaline leaching for 1 h at 130 °C with an alkali to total calcine mass ratio of 1.2:1, initial NaOH concentration 30%, and a stirring speed of 600 rpm. Under these conditions 98% of the Pb and 98% of the Mo in the flotation tailings were extracted.
The extraction of molybdenum from high-impurity ferromolybdenum by roasting with Na2CO3 and CaO and leaching with water was studied. The results indicated that the process was an effective method to extract molybdenum from high-impurity ferromolybdenum. Under the optimum experimental conditions (CaO/Na2CO3 mole ratio of 0.76; Na2CO3 and CaO added total mole number of 0.57/100 g ferromolybdenum powder; oxidation roasting at 700 °C for 2 h; water leaching for 2 h at 98 °C with liquid/solid ratio of 3:1), the extraction of molybdenum was 97.5%, and the leached rate of impurities phosphorus, arsenic and silicon were only 16.2%, 18.1% and 41.2% respectively. The addition of CaO can effectively hold back the dissolution of the impurities in the leaching, which is due to the formation of Ca3(PO4)2, Ca3(AsO4)2 and CaSiO3.
A study was made of the United States Bureau of Mines' electro-oxidation process for the recovery of molybdenum from molybdenite. The investigation sought to determine operating conditions which would minimize by-product chlorate formation and the resulting loss in current efficiency. Such minimization would contribute to making the process economically attractive.It has been shown that a continuous flow-through graphite cell process produces less chlorate than either graphite or lead anode batch operations during the electro-oxidation of molybdenite at low temperatures, approximately 30°C.
The removal of molybdenum from a copper ore concentrate by sodium hypochlorite leaching was investigated. The results show that leaching time, liquid to solid ratio, leaching temperature, agitation speed, and sodium hypochlorite and sodium hydroxide concentrations all have a significant effect on the removal of molybdenum. The optimum process operating parameters were found to be: time, 4 h; sodium hydroxide concentration, 10%; sodium hypochlorite concentration, 8%; liquid to solid ratio, 10:1; temperature, 50 °C; and, agitation speed, 500 r/min. Under these conditions the extraction of molybdenum is greater than 99.9% and the extraction of copper is less than 0.01%. A shrinking particle model could be used to describe the leaching process. The apparent activation energy of the dissolution reaction was found to be approximately 8.8 kJ/mol.
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.
Molybdenum was bioleached from molybdenite-containing mine waste samples and from high purity molybdenite in solutions amended with ferrous sulfate. Bioleaching of Mo occurred when solution Eh exceeded 750 mV to 800 mV (s.h.e.). The rate of bioleaching of high purity molybdenite increased with decreasing particle size, averaging 3.22 mg Mo/m2/d at 24 °C. This surface area-based rate is less than 1% of the bioleaching rate of pyrite. Molybdenite bioleaching increased with temperature by a factor of approximately 2.5 for each 10 °C increase in leaching temperature up to 40 °C. Apparent activation energies were 61.2 to 73.4 kJ/mol for bioleaching of high purity molybdenite and for molybdenite from mineral processing waste streams, respectively. Thermophilic bioleaching at 65 °C did not improve the rate or extent of Mo extraction. Molybdenum extraction was about 50% from mineral process waste stream samples after the first month of bioleaching, and slowed thereafter. Ultimate Mo extraction was approximately 85% after 6 months of bioleaching. High concentrations of ferric iron in bioleach solutions allowed iron-oxidizing microorganisms to grow, oxidize iron and bioleach molybdenite at dissolved Mo concentrations as high as 4.4 g/L. There was no evidence for selection of microbial populations with increased Mo resistance.
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.
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.
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.
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.
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.
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.
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.
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.
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.
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.
0149-6395 print / 1520-5754 online DOI: 10.1080/01496395.2015.1059348 REFERENCES 1 Molybdenum metallurgy review: Hydrometallurgical routes to recovery of molybdenum from ores and mineral raw materials
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Copyright © Taylor & Francis Group, LLC
ISSN: 0149-6395 print / 1520-5754 online
DOI: 10.1080/01496395.2015.1059348
REFERENCES
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