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By Guiqing Zhang, Liansheng Xiao, Zuoying Cao, Li Zeng

A new process of direct solvent extraction of nickel from sulfuric acid leach solutions of low grade and complicated nickel resources using a novel extractant HBL110 was proposed. The basic performance of new organic system was investigated and the continuous pilot plant and the industrial test were conducted. The results indicated that the extractant of HBL110 can be used to directly extract nickel selectively over iron, aluminum, manganese, magnesium, calcium and chromium from sulfuric acid leach solutions. The loaded organic solution can be easily stripped using dilute H2SO4 solutions. Both kinetics of extraction and stripping of nickel were fast. The continuous pilot plant test with the sulfuric acid leach solution of laterite containing 6.22g/L Ni2+, 14.1g/L Fe3+, 0.73g/L Fe2+, 33.9g/L Mg2+, 1.58g/L Al3+, 0.495g/L Ca2+, 0.97g/L Mn2+ and 0.5g/L Cr3+ as feed in mixer-settlers showed that the process had run stably with 98% recovery of nickel. The concentration of nickel in strip liquors reached as high as 50 g/L while the removals of Fe3+, Fe2+, Mg2+, Al3+, Ca2+, Mn2+ and Cr3+ were 98.5%, 91.5%, 99.95%, 99.8%, 99.1%, 98.6% and 99.5%, respectively. The industrial test with the sulfuric acid leach solution of electroplating sludge containing 11.5g/L Ni2+, 2.68g/L Fe3+, 7.54g/L Cr3+, 0.61g/L Mg2+, 2.09g/L Al3+, 0.68g/L Ca2+, 0.022g/L Mn2+ and 0.154g/L SiO2 as feed in mixer-settlers also showed that the process had operated successfully with 99.5% recovery of nickel. The concentration of nickel in strip liquors reached 42.7 g/L while the removals of Fe3+, Cr3+, Mg2+, Al3+, Ca2+, Mn2+ and SiO2 were 99.2%, 99.77%, 99.73%, 99.56%, 99.57%, 98.51% and 98.75%, respectively. The new process exhibited big advantages of short flow sheet, high recovery of nickel, high removal of impurities, low chemical reagent consumption and low cost. Application of this process in industry is expected to make important impact on the extraction of nickel from sulfuric acid leach solutions and to deliver good environmental and economical benefits.

Jan 1, 2016

By Broggi Andrea

Condensates produced by SiO(g) and CO(g) have a crucial importance in silicon production. Condensation will follow the reactions 3 SiOðgÞ+COðgÞ→ 2 SiO2 + SiC and 2 SiO ðgÞ→Si+SiO2. Solid compounds generated during silicon production contain Si, SiO2 and SiC in different microstructures and amounts. In industrial systems, condensates limit the SiO(g) escape to the off-gas system. On the other hand, they can block the gas flowing from the high temperature zone, hence the necessity of studying an optimized condensation process. This review discusses SiO(g) condensation and its interaction with CO(g). The report gives an overview of the condensates found both at laboratory and at industrial scale. The two main compounds are the Si–SiO2 and the SiOx–SiC mixtures. The hypothesis on the mechanism of formation of SiC–SiOx condensates is presented through a wide selection of performed works. The advantages and disadvantages of the former experimental procedures are discussed.

A study of iron-removal from copper leachate by limestone precipitation followed by direct eltctrowinning of copper has been conducted. The influence of the process parameters: amount of precipitant, contact time, temperature, agitation intensity, solution acidity, and initial iron concentration on iron-removal was investigated. The effects of electrowinning time, cell voltage, solution acidity and temperature on electrowinning of copper were also investigated. The leachate contained 1.20 g/l cu2+ and 7.41 g/l total iron at pH 2.27. After contacting with 14 g/l precipitant for 2 hours, 97.5 % of the iron was attenuated and 98.4 % of the copper was recovered. With the following electrowinning conditions: 3 lead anodes, 3 stainless steel cathodes, 2.5 V, 39.41 A/m2 and 10.7 g/l H2S04, copper recovery-was 98.4 %. Purity of copper powder produced was greater than 99.0 %.

Jan 1, 1999

By J. A. Eisele

There is currently much interest in how extraction of gold and silver from ores with acidic thiourea solution compares with extraction using alkaline cyanide solution. Agitation leaching tests were performed by the Bureau of Mines on 14 precious metal ores using sodium cyanide (?NaCN) and thiourea [(NH2)2CS) as extractants. The objective was to compare the results of the two extractants. NaCN was used at a concentration of 2 g/L with and without H202 as an oxidant. (NH2)2CS was used at concentrations of 2 and 20 gjL, and the solution potential was controlled at 390 to 420 mV standard hydrogen electrode (Eh) with Fe2(S04)3. With both reagents a.t 2 g/L extractant levels, NaCN extracted more gold and silver than (NH2liCS from all of the ores. At 20 g/L (NH2)2CS compared with 2 g/L NaCN, (NH2)2CS extracted more gold than NaCN from two ores, the same amount from one ore, and more silver from one ore. NaCN extracted more gold than (NH2 )2CS from 10 ores and more silver from 4 ores.

Jan 1, 1989

By Chuan-fu Zhang

Li(Nil/3Col/3- xMnl/3)MxO2(M=Ti, Mg) cathode materials were prepared from LiOH?H2O and oxalate precursor. The physical and electrochemical properties were studied by XRD, SEM, cyclic volt-ampere (CV), AC impedance and constant current charge-discharge. The results show that crystal volume of Ti4+ or Mg2+ doped sample is increasing and the electrochemical reaction resistant Rct is decreased at high rate, improving the cyclic performance and rate capability. The effect of Ti4+ doped is better than Mg2+. The sample is well crystallized and simple pure phase with a-NaFeO2 layered structure when doping quantity x=0.025. The second specific discharge capacity of Li(Nil/3Col/3- 0.025Mnl/3)Ti0.025O2 is 143.2mAh/g at 1C,128.0mAh/g at 2C in the voltage of 2.75-4.3V,and still has 140.7mAh/g and 121.7mAh/g after 20 cycles, respectively, keeping 98.25%, 95.07% capacity.

Jan 1, 2009

By Yuanbo Zhang

Zinc leach residues produced by hydrometallurgical zinc plants contain many valuable elements (zinc, iron, lead, gallium, indium silver, etc), while they have not been utilized effectively hitherto. In this paper, a new process of flotation-reduction roasting-magnetic separation to recover these valuable elements from zinc leach residues is developed on a basis of mineralogy investigations of them. The results show that silver concentrate with grade of 4321.6 g/t and recovery of 80.69% have been obtained, using 1750 g/t Na2S as conditioner, 750 g/t ammonium dibutyl dithiophosphate and SN-9 (assistant collector) as combined collector and 100 g/t terpenic oil as frother, if zinc leach residues with 523 g/t silver are floated at pH = 5.0. When the flotation tailings are subjected to reduction roasting at 1100? for 2 h, the volatilizations of zinc, lead and indium are over 96%. After roasted products are treated by wet-grinding and magnetic separation, the grade of gallium and gallium in the concentrates achieved respectively reaches 1805 g/t and 1536g/t, and the recovery of them is high to 94.67% and 93.20%, respectively.

Jan 1, 2006

By Yuanbo Zhang

Keywords: Complex Iron Ores, Roasting, Tin, Zinc, Pellets, Comprehensive Recovery. The complex iron ores containing tin, zinc are typical refractory ores. Large reserves of the ores are found in China and have not been utilized efficiently so far. A new process of tin and zinc recovery and pellet preparation from iron concentrate containing tin and zinc by controlled atmosphere roasting is developed in this investigation. The results of 2573 N/P compression strength of product pellets, the 91.49% and 82.22% volatilization of tin and zinc respectively, are achieved for a concentrate containing 0.39% tin and 0.28% zinc. The strength and chemical compositions of the pellets meet the requirement of blast furnace ironmaking. Volatilization mechanism of tin and zinc are investigated. Thermodynamic study shows that the reduction balance curves of SnO2(s) and SnO(g) are proximate in the presence of carbon, and the standard formation free enthalpy of FeO and SnO2 is also approachable. Therefore, it is difficult to realize the complete volatilization of tin in the form of SnO(g) by selective reduction. But SnO2(s) and SnO(g) will form SnCl4(g) or SnCl2(g) in the presence of chlorides and can be easily volatilized from the ores.

Jan 1, 2005

By Wang Dawan

During the process of arsenic bearing gold concentration biooxidation, the content of chloride ion in water is a very important factor to the activity of bacteria and the oxidation of sulfide minerals. In this paper, the influence of underground mine water with different chloride ion content to the biooxidation of arsenic bearing gold concentration and the limit of chloride ion content to bacteria were summarized. These results indicated if the content of chloride ion was less than 1.2g/L, the bacteria?s ability to oxidize sulfide minerals was little affected, and if the content of chloride ion were increased to 2g/L or more, the activity of bacteria was seriously restrained, and if the content of chloride ion were increased to 4g/L, the activity of bacteria was absolutely restrained. So, as the content of chloride ion in water was cumulated to 2g/L or more in biooxidation plant, the water must be refreshed to maintain normal production.

Jan 1, 2008

By Hojat Naderi

In this research assisted leaching of chalcopyrite using sodium peroxodisulfate was studied. Electrochemical techniques such as Tafel analysis and electrochemical impedance spectroscopy (EIS) were used to investigate the surface reactions of a chalcopyrite electrode in the presence and absence of Na2S2O8. According to the results, addition of sodium peroxodisulfate increases the current exchange of the corrosion process and prevents the formation of a passive layer on the surface of the electrode. EIS tests show the trans-passive dissolution of the passive layer on the chalcopyrite electrode in presence of Na2S2O8. Equivalent electrochemical circuits were modeled for Na2S2O8-free and -containing solutions and the model parameters were compared to determine the effect of S2O8 2− on chalcopyrite dissolution. Copper dissolution from pure chalcopyrite mineral and chalcopyrite concentrate was carried out using 15 g/L sulfuric acid solution with and without Na2S2O8. Simultaneous presence of 10 g/L ferric ion and 40 g/L peroxodisulfate salt results in about 40 and 18% of Cu extraction from chalcopyrite concentrate and pure chalcopyrite mineral, after 96 h, respectively.

By G. G. Krishna Murthy

Several experiments were carried out to simulate the bath smelting reactions where iron oxide particles were dropped on to an iron-carbon bath containing molten CaO-SiO2-Al2O3 slag layer on its surface. The composition of the slags used in the experiments was %CaO:46.1; %SiO2:41.7; and %A1203:12.2. Experiments were con¬ducted with carbon saturated iron melts contained in either a pure graphite crucible whose walls were covered with a pure alumina tube, or a pure flat bottom alumina crucible, at 1723-1823 K. Degree of reduction of a pure iron oxide particle was linear with time in the initial stages (up to about 20% oxygen of the original pellet was reduced). The rate of reduction of smaller oxide pellets was smaller than that of a bigger pellet - the time for complete reduction of a 5.6 g wustite pellet was smaller than that for a 1.2 g wustite pellet. It was also established that the interaction rates obtained by adding a molten pure oxide particle of given weight to the system were lower than that determined by using a solid oxide particle of same weight. The overall rates of reduction decreased with decreasing the bath temperature. The experimentally determined rates of reduction based on the time for 70% reduction were influenced substantially by addition of slag up to about 50 g on to a bath of 78.8 mm in diameter, and remained constant when slag present on the bath surface was between 50 g and 200 g.

Jan 1, 1993

By D. A. R. Kay

Phase stability diagrams are used to predict the abilities of lanthanum and cerium oxides to desulfurize coal gasification products in the temperature range 800-1000 C. Results of desulfurization studies in laboratory fixed bed reactors illustrate the effects of i) sorbent preparation, ii) input gas quality and iii) temperature, on the desulfurization reaction. 2CeO(2-x)(s) + H2S(g) + (1-2x)H2 = Ce2O2S(s) + 2(1x)H2O(g) The results of desulfurization/oxidation regeneration cycles are also reported.

Jan 1, 1989

By Gang Qi

The solubility of nitrogen in pure molybdenum was measured over the temperature range 2173 to 2473 K and up to one atmosphere (101 kPa) of nitrogen gas pressure. Both thermochemical and kinetic aspects of this phenomenon were considered in this study. The solution of nitrogen in molybdenum may be expressed by the reaction 1/2N2(g) = {N} (Mo,s), for which the standard Gibbs energy is given by the expression {A}G° = 78,500 + 27.36T Joules/g atom N. The formation of molybdenum nitrides, specifically Mo2N, was considered. Attention was also focused on the mechanism of nitrogen solution in molybdenum. Estimates of the diffusivity of nitrogen atoms in solid molybdenum were derived from the kinetic data.

Jan 1, 1991

By Ali Hasham

Reduction rates of FeO in CaO-SiO2-Al2O3 slags by Fe-C melts have been investigated experimentally at 1723 K. The rates were estimated using experimentally determined CO(g) evolution rates. Flash X-ray images were employed to visualize the behavior of the melt-slag interface during the reactions. The influence of melt interface was examined by conducting two types of experiments - (i) dropping one (or several) Fe-C droplet(s) weighing 2 g into the molten slag bath, and (ii) dropping a small amount (10-30 g) of slag onto a Fe-C melt weighing 400 g. The rates from both types of experiments were compared and analyzed to determine the rate limiting steps. It was established that neither the carbon transport in the Fe-C melt, nor the transport of gas species in the gas film adjacent to Fe-C melt was responsible in limiting the reaction rate. It is possible that the slag phase mass transport may be controlling the initial reduction rate in the droplet experiments. However, a, different mechanism was responsible in limiting the reduction reaction when the overall %C in the droplets fell to about 2-3.

Jan 1, 1992

By Sheng-hai Yang

The ultrafine, monodispersed, spherical cobalt powder has been fabricated by reducing cobalt hydroxide with glycol. It were investigated that the effect of NaOH concentration on the reaction rate and the effects of the concentration of cobalt hydroxide and surfactant on the morphology and particle size. The particles were characterized by the X-ray diffraction(XRD), scanning electron microscope(SEM), laser diffraction particle size analyzer and BET specific surface analyzer respectively. The XRD shows that the cobalt powders exist mostly in face centered cubic (fcc) structure, a little in simple hexagonal. The cobalt powders prepared at Co(OH)2150g/L, NaOH 40 g/L and surfactant 30g/L, its average particle size, specific surface area and cobalt content, were 0.88µm, 2.705m2/g. and more than 99.5% respectively.

Jan 1, 2006

By Wenjun Duan, Kun Wang, Qingbo Yu, Qin Qin

"A novel oxygen production technology noted chemical looping air separation (CLAS) was introduced and properties of Cu-based oxygen carriers used were investigated. The oxygen carriers were prepared by mechanical mixing. Al2O3, Fe2O3, ZrO2, TiO2 and SiO2 were used as binders during the preparation. XRD patterns show that spine! oxides are formed using Al2O3 and Fe2O3 as binders. Phases of CuO and binders keep stable using ZrO2, TiO2 and SiO2 as binders. BET values of oxygen carriers increase in the order of ZrO2>TiO2>SiO2>Fe2O3>Al2O3. The effects of temperatures, binders, particle sizes and flow rates on reduction/oxidation reaction rates were studied in TG experiments. The reduction reaction rate increases in the order of ZrO2>TiO2>SiO2. The effect of binders on the reactivity of oxidation is not obvious. The reduction and oxidation reaction rates both have the tendencies of increasing with the temperature increasing and with gas flow decreasing. However, the effects of gas flow and particle size are little.IntroductionChemical looping air separation (CLAS) is a novel oxygen production technology. The oxygen carriers used in CLAS release oxygen in reduction reactor and regenerate in the oxidation reactor. Selecting suitable oxygen carriers is the key for CLAS. Oxides of CuO, Mn2O3, Co3O4, Pb3O4, CrO2, SrO2, PdO2 and CeO2 have the property of releasing oxygen at suitable temperature and oxygen partial pressure [1]. Among these oxides, CuO, Mn2O3, Co3O4 are the most common compositions used to prepared oxygen carriers for chemical looping combustion (CLC) or chemical looping with oxygen uncoupling (CLOU) [2-5]. Moreover, thermodynamic calculations and preliminary experiments were carried out on oxides of Cu (CuO/Cu2O), Mn (Mn2O3/Mn3O4) and Co (Co3OJCoO) by Moghtaderi to ascertain their feasibility for CLAS process [6]. CuO has the highest oxygen transport capacity (0.1 g 02/g CuO compared to 0.03 g 02/g Mn2O3 and 0.06 g 02/g Co3O4). However, agglomeration is the biggest issue for copper oxides used at high temperature because of its low melting point (l 450°C for CuO, 1235°C for Cu2O and 1075°C for the mixture of CuO and Cu2O) [7]. In order to improve the capability of anti-sintering and stability ofCuO, binders (ZrO2, MgAl2O4, TiO2, SiO2, Fe2O3, Al2O3, cement) are added during the preparation of Cu-based oxygen carriers [8]. In this paper, the Cu-based oxygen carriers were prepared by mechanical mixing method. TiO2, Al2O3, Fe2O3, ZrO2 and SiO2 were selected as binder. XRD, BET and SEM were used to analyze the phase, surface area and surface morphology of oxygen carrier prepared. The reactivity of oxygen reduction and oxidation of oxygen carriers prepared were investigated in STA409PC thermal analyzer."

Jan 1, 2014

By C. A. Nogueira, F. Pedrosa, F. Margarido, R. N. G. Guerra

"A hydrometallurgical process for recycling spent Zn-MnO2 batteries was developed, involving leaching with sulfuric acid, purification by precipitation and metals separation for further recovery. Leaching of zinc oxide was easily attained while for manganese oxide was rather difficult depending on temperature and acid concentration. At 90°C and with the liquid/solid ratio of 20 L/kg, more than 95% of zinc is recovered in 30 minutes with 0.5M H2SO4. To attain similar recovery for manganese, higher levels of acid concentration and time are needed (e.g 0.7 M and 2 hours). After leaching a purification step is necessary to remove iron co-dissolved through Fe(III) precipitation. Separation of zinc from manganese by solvent extraction with 1M DEHPA follows in the process route. The countercurrent multistage separation diagram developed allows the production of a zinc electrolyte with 120 g/L Zn and 0.005 g/L Mn, and a raffinate with 16 g/L Mn and 0.013 g/L Zn.IntroductionThe recycling of spent Zn-MnO2 batteries is carried out by dedicated recyclers using several process alternatives [1]. Most of the running processes use pyrometallurgical technologies [2] based on zinc distillation. Alternative treatments by hydrometallurgy have been reported [3-5] but its application is scarce.In order to contribute to the development of alternative solutions, a process for recycling batteries based on Zn-MnO2 systems has been studied. The process is based on the hydrometallurgical treatment of spent batteries in sulfuric acid media (Figure 1). The process starts with the physical processing by shredding using cutting grinders, where the batteries are opened and size reduced, allowing the liberation of the interior electrode particles or pastes. The fines can be then separated and the coarse material containing the steel scrap can be treated to remove the remaining aggregate electrode. The final scrap also contains some plastics and other non-metal materials and can be further valorized."

Jan 1, 2008

By Chunhua Liao, Chao Peng, Guangsheng Huo

To get the separation of tungsten and molybdenum, we adjusted the pH of the solution to 7, at which tungstate (WO4 2-) will be polymerized into HW6O21 5-, while molybdate will still exist in the form of MoO4 2-. Two strongly basic anion exchange resins (D201, D213) were experimented for the separation of tungsten and molybdenum. The results of condition experiments indicate that D213 resin performs better in the treatment of lower concentration solution A (WO3 10g/L, Mo 4.12g/L) with a separation factor as high as 10.64. While D201 resin is more suitable for the high concentration solution B (WO3 70g/L, Mo 28.97g/L) with a separation factor of 6.9. The results of column testing performed on D213 resin show that the solution after loading process with a mass ratio of Mo/WO3 = 50 and the desorption solution with a mass ratio of WO3/Mo = 27 can be obtained

Jan 1, 2014

By R. Padilla

In most of the Chilean copper and gold concentrates, arsenic is present mainly as enargite, Cu3AsS4, which at the conventional roasting or smelting temperatures decomposes liberating arsenic as gaseous compounds. The objective of this investigation was to study the decomposition reaction of enargite in an inert atmosphere in the range of 500 °C - 750°C. The results obtained indicated that the decomposition begins at about 525 °C. Two distinct regions in the temperature dependence of the decomposition rate were observed: a low temperature region between 525 °C to 650°C and a high temperature region between 650 °C to 725 °C. The X-ray diffraction analysis suggests that the decomposition of enargite occurs through the formation of intermediate tenantite, according to the following sequential reactions: 4Cu3AsS4(s) ? Cu12 As4S13(s)+ 3/2S2(g) Cu12 AS4S13(s) ? 6Cu2S(s) + 2As2S3(g) + 1/2S2(g) The decomposition kinetics was analyzed by using the model for a topochemical reaction of spherical particles: 1-(1-X) 1/3 = kt, which represented well the data. Apparent activation energies of 113 kJ/mol and 60.5 kJ/mol were found for the low and high temperature regions, respectively.

Jan 1, 1999

By Fatma Arslan

Alada (Kayseri-Turkey) lead-zinc sulfide-oxide ores contain 10.17% Pb, 10.98 %Zn, 164 g/t Cd and 57 g/t Ag. During the laboratory scale flotation tests of these ores, it was found that almost all zinc was remained in the tailings that contain 12.31% Zn, 1.35%Pb, 198.3 g/t Cd, and 22.3%Fe. Three different methods were studied such as basic (NaOH) leaching, roasting+acidic (H2SO4) leaching and direct acidic (H2SO4) leaching for recovering metals that these tailings contain. In the direct acidic leaching, which gave the best results, dissolution efficiencies of 91.5% zinc, 78.2% cadmium, and 14.3% iron were reached at the optimum leaching conditions while Pb remained in the leach residues and the acid consumption was 422.3 kg/ton of tailings.

Jan 1, 2006

By M. Omneya El-Hussaini, M. Galal Abdel Wahab, A. Mohamed Mahdy, M. Fathi El Shahat

"Lead-zinc mineralization is distributed in Egypt along the Red Sea coast between El Qusier and Ras Banas. A representative sample from Um Gheig was collected and ground to -60 mesh for the mineralogical study. It contains hemimorphite (Zn2SiO4.H2O), sphalerite (ZnS), galena (PbS), calcite (CaCO3) and dolomite (CaMg(CO3)2) beside the gangue minerals gypsum (CaSO4.2H2O) and hematite (Fe2O3). Cadmium seems to be incorporated in the crystal structure of the zinc-lead minerals. The chemical analysis of Um-Gheig representative sample revealed that, it contains 16% zinc combined with 0.1% cadmium, 2.3% lead and 0.1% boron beside other elements.Pug and agitation leaching were performed on samples ground to -200 mesh to achieve the maximum dissolution efficiency of cadmium with respect to other elements. Cadmium leaching efficiency reached 90% by using 40% H2SO4 at S/L ratio 1/2 and stirring time 2 hours at 80o C. Cadmium was then separated as yellow precipitate of CdS from the sulfate leach liquor containing 0.9 g/l Cd, 145 g/l Zn, 0.15 g/l B and 0.01 g/l Pb. Separation of cadmium was carried out by direct precipitation using Na2S as well as H2S. The recovery of 94% cadmium free of zinc was achieved.IntroductionLead-zinc ores occur in Egypt in a few localities distributed along the Red Sea coast between El Qusier and Ras Banas. Um-Gheig Pb-Zn ore is a replacement deposit in limestone and calcareous sandstone in the basal series of the Middle Miocene rocks. The reserve of the ore deposits has been estimated by Soviet and German Experts to be about 900,000 Ton, Bishai [1]. As has been mentioned, zinc is a common element and cadmium is comparatively rare in almost zinc minerals. It was estimated that zinc is 200 times as abundant as cadmium. According to Bodas [2] about 2.1% of cadmium was carried by the Zn mineral hemimorphite (15.3% Zn) in Thailand. Also, Onal et al., [3] found that 164 g/t Cd was carried by Pb–Zn ore sample containing 10.17% Pb and 10.98% Zn from Aladag in Turkey. Cadmium metal has gained great importance due to its wide applications in civil and navy purposes as well as in nuclear field. It is characterized by its ability to absorb thermal neutrons during the chain reaction inside the nuclear reactors. It is also used in the manufacture of sheets which directed the chain reactions [4]."

Jan 1, 2008