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Apatite concentrate is an important source of phosphate and contains significant amounts of rare earth elements (REEs). In this study, the possibility of selective phosphoric acid leaching of phosphate and calcium from the Esfordi (Yazd province, Iran) apatite concentrate, containing 16.3% phosphorus, 37.1% calcium, 2.9% iron and 1.2% REEs was investigated. REEs concentrated in the leaching residue were processed using acid baking and bicarbonate precipitation methods. In the phosphoric acid leaching process, almost all the Ca and phosphate of the concentrate were leached from the concentrate, whereas only about 1% of the REEs were leached. By leaching the concentrate with 25% phosphoric acid at 90 °C and 1200 rpm, a residue containing 17.5% REEs, 4.8% phosphorus and 28.5% iron was obtained. The residue was acid baked, water leached and processed for removal of iron and recovery of REEs. Acid baking at 270 °C with acid to solid ratio of 2 (w/w) for 5 h led to 95% leaching of REEs. By increasing the pH of the leachate to 3.9, all the iron and phosphate were precipitated within 2 h. Finally, REEs were precipitated from the iron-free solution by using the bicarbonate precipitation method, and a product containing 39% of total REEs was produced. The results demonstrate that the three-stage process is an elegant method for processing of the Esfordi concentrate to produce phosphoric acid and an iron-free REE carbonate concentrate.

Pre-concentration of lead and zinc ore in Calcimin (public Co) which is done using heavy-medium cyclones (HMC) undergoes lots of inefficiencies and not in the desired manner. Considering th...

The present study attempts to investigate the chloride removal from industrial soils and zinc slag in zinc production companies. The presence of chloride in factory’s samples causes some problems such as the anode erosion in electrolysis cells, negative effects on the quality of zinc sheets, and reducing the quality of product. Having said that, decreasing the amount of chloride ion will make it possible to produce zinc sheets with SHG quality. There are different methods that can be used to decrease the amount of chloride ion, and one of them has been practiced in this study. A hydrometallurgy method uses sodium metabisulfite and copper(II) sulfate in which Cl− reacts with Cu+ to form CuCl precipitate and removes the chloride content. The parameters pH, temperature, time, and the consumed ratio of copper(II) sulfate to sodium metabisulfite were investigated. Once optimal values of these parameters were calculated, the efficiency of chloride removal from NaCl solution with a value of 57% and the obtained solution from slag leach with a value of 2.6% were evaluated as inefficient.

The present study is focused on the leaching and recovery of phosphate and Rare Earth Elements (REEs) from the Esfordi phosphate concentrate (Yazd province, Iran), which contains 16.31% P, 37.13% Ca, 2.91% Fe, 3% F, 0.3% Cl, and the TREE content is 1.2%. Cerium, lanthanum and neodymium account for 82% of the total rare earth content. Fluorapatite, magnetite and calcite-dolomite, with 88.7%, 4.8% and 2.3%, respectively, are the predominant minerals in the concentrate. Monazite (1.2%), apatite and xenotime (0.2%) are the main REE-bearing minerals. A process was designed for the simultaneous recovery of phosphate and REEs form flourapatite concentrate. Through direct sulphuric acid baking of the concentrate, maximum Ce, La, Nd and Y recoveries of 85.5 ± 3.0%, 85.2 ± 2.4%, 83.8 ± 2.1% and 84.8 ± 2.0%, were achieved, respectively, at an acid to ore ratio of 2.9, temperature of 270 °C and digestion time of 110 min. The phosphate precipitation method was used as an effective method to separate iron and phosphate from rare earths. By increasing the Fe 3 + P O 4 3 − ratio from 0.11 to 2 and gradually increasing the pH to 3.5, 98% of iron, 45% of Ca and all the Th precipitated out of solution. Subsequently, a total of 97% of rare earth elements were precipitated from solution by conventional bicarbonate precipitation to obtain a mixed concentrate containing 22.4% rare earth elements. The carbonate concentrate thus produced by using the mentioned novel process contained considerable amounts of calcium and sulfur. For this reason, a selective leaching stage is proposed to leach the calcium and phosphate from the concentrate prior to the acid baking process. This has been investigated in the part II of this research.

Germanium transport from a solution containing tartaric acid by a flat sheet supported liquid membrane (FSSLM) using trioctylamine (TOA) as a carrier and polytetrafluoroethylene (PTFE) as a membrane was investigated. A mass transfer model was developed to monitor the transport process based on experimental results. The effect of parameters such as feed solution pH, TOA concentration, initial germanium concentration, and strip hydrochloric acid concentration on the germanium flux and the transport percentage were studied. A high permeation was observed at a feed solution pH of 3.00, 40%v/v TOA and 5 mg/dm3 Ge4+. At HCl concentrations of 1–3 mol/dm3, the germanium transport was complete. Finally, based on the mass transfer model, the aqueous and organic resistance values were 11,802 and 860.85 h/cm, respectively. The validity of the model was investigated by fitting the model and experimental data. The correlation coefficient of 0.99 showed the validity of the model.

In this research, both the kinetics and the mechanisms of the roasting of a sphalerite concentrate from the Bafgh mining complex in Iran were investigated. The oxidation process was performed in a muffle furnace in air and the effects of time and temperature on the degree of oxidation of the zinc sulfide sample were quantitatively studied. The experimental data were fitted to the shrinking core model. In the temperature range of 650–800 °C, the rate-controlling step was the chemical reaction between the zinc sulfide and oxygen with an activation energy of 103 kJ mol−1. On the other hand, in the temperature range of 850–950 °C, the rate-controlling step was oxygen diffusion with an activation energy of 50 kJ mol−1. Also, the roasting process was studied using thermogravimetric (TGA) and derivative thermogravimetric (DTGA) techniques.

In the current study, a novel procedure for studying the zinc solvent extraction kinetics in the Lewis cell is presented. In this study, three steps were followed. In the first step, using the response surface methodology (RSM), the conductivity relation as a function of the pH and the zinc concentration was found. With respect to this model, knowing the pH and the conductivity of a zinc solution, the zinc concentration of the solution is obtained. In the next step, with respect to the RSM design of the experiment, an optimum condition, in which the high extraction of zinc occurred in the Lewis cell, was found to be the pH of 2.88 and the initial zinc concentration of 14.96 g/L. In the final step, the kinetic study was carried out on the solution with the aforementioned condition. Taking into account the mathematical kinetic models obtained by the Lewis cell, the extraction of zinc(II) from a zinc sulfate solution with D2EHPA was limited both by the diffusional and chemical regimes.

A systematic study for the separation of zinc from solutions containing zinc, manganese, iron, cadmium and magnesium as sulfates was developed after a batch countercurrent extraction, followed by scrubbing and stripping. In the first step, batch countercurrent experiments were carried out on a synthetic solution containing 14.60 g/L Zn(II), 6.4 g/L Mn(II), 2 g/L Mg(II), 819 mg/L Ca(II) and 60 mg/L Fe(II). The results of this step showed that at a three-stage extraction and six cycles under the conditions being O/A ratio of 1:1 and temperature of 40 °C, 13.06 g/L equal to 89.50% of zinc was extracted by D2EHPA whereas only 83 mg/L of manganese was loaded by the extractant. As a result, the low extraction percentage of manganese is an attractive advantage of this process. In comparison to the literature, the high separation factor of zinc and manganese (i.e. 643.79) of this study was obtained without any pH adjustment or using any synergetic mixture. To remove the trace concentrations of cadmium, calcium, manganese and iron from the loaded organic solution, a series of scrubbing experiments was conducted. With regard to the scrubbing results, a solution with a composition containing 6 g/L Zn and 18 g/L H2SO4 could scrub trace concentrations of metals from the loaded organic solution. Finally, the McCabe–Thiele diagram of the stripping stage predicted that at an O/A phase ratio of 2:1, two countercurrent stages are sufficient to transfer zinc from the organic solution to the aqueous solution. The obtained aqueous solution contained 88.8 g/L of Zn and the impurities were under the limited concentrations of electrowinning.

The assessment of heavy metal transition from Ni–Cd zinc plant residue (filtercake), as concentrations of zinc, nickel, cadmium and lead, requires quantification and a mathematical model that predicts their relative concentrations. Variability of filtercake characteristics may change the availability of heavy metals to the leachate or environment. In this study, a novel artificial neural network (ANN) model was constructed to predict Zn, Ni, Cd and Pb concentration leached from Ni–Cd filtercake in the leaching column. A three-layer backpropagation neural network was optimized and developed based on the Bayesian training algorithm. The inputs of this network are pH, flow rate of acidic influent, particle size and time. The geometry of the network giving the minimized mean square error (MSE) and sum of squared error (SSE) was a three-layer network having 18 neurons in the hidden layer (4:18:4) with a tangent sigmoid transfer function (tansig) at the hidden layer and linear transfer function (purelin) at the output layer. The fitting, regression, error and histogram plots for each response illustrate that there is a good agreement between the experimental data and the predicted values. Finally, a generalization of the developed model was carried out as 3D plots to evaluate the interactions of the input parameters on the transition of heavy metals to the leachate. With respect to these results, the effect of particle size on concentrations of zinc, nickel and lead are less (

The objective of the study was to determine the type and distribution of rare earth elements (REEs) in the phosphate flotation concentrate from the Esfordi flotation plant in central Iran, in order to understand their leaching behaviour. Samples of the concentrate were prepared and analysed using XRF, XRD, ICP, and EMPA. The Ce, Nd, La, and Y contents were 5608, 2227, 1959, and 679.7 mg/kg, respectively. The main REE-bearing minerals in the concentrate are fluorapatite, monazite, and xenotime. Apatite is the chief mineral of the concentrate and the ratio of apatite to monazite is greater than 60. From EMPA analysis, monazite contains a much higher proportion of REEs than apatite, but more than 50% of the REE content of the concentrate is in apatite, as this is the main mineral constituent of the concentrate. During the leaching of REEs in the phosphoric acid production process, more than 99% of the total rare earth values report to the large amounts of phosphogypsum (CaSO(4)2H2O) formed during the dissolution reaction. Accordingly, two process options are suggested for treating the Esfordi concentrate. In the first option, the concentrate is treated with 98% H2SO4 at 200-300°C. After digestion and water leaching, phosphorus and REEs are precipitated out of solution. In the second option, the concentrate is treated in a pre-leach stage. The process residue from the pre-leach stage is then subjected to acid digestion at elevated temperature for phosphoric acid production.

Artificial neural network was used to predict the synergetic effect of LIX 984N and D2EHPA on separation of iron from zinc solution. The aim was to predict iron and zinc extraction as a function of pH, temperature, and various organic compositions. Optimum number of hidden layers and nodes in each layer were determined. A multilayer network, with two hidden layers (4:9:5:1) was applied to predict zinc and iron extraction. Effect of pH, temperature, extractant composition, and interaction of them on extraction percent was also investigated using 3D plots. The modeling results showed that there is an excellent agreement between the experimental data and the predicted values.

A step by step hydrometallurgical process for the production of copper cathode was developed after a two-step precipitation from leaching solution of copper oxidized ore, followed by copper concentrate leaching and electrowinning. The copper oxidized ore was primarily comminuted to a size below 100 μ, followed by acidic leaching at 25 °C for 40 min in H2SO4 solution, in which recovery of copper and iron were 95.95% and 12.63%, respectively. To remove iron impurity, at the first step of precipitation, NaOH was added to increase pH from about 1.5 to the optimum pH of 3.8 at 60 °C for 60 min; thus iron precipitation with recovery of over 80% was achieved. Copper precipitate as concentrate was obtained in the same method from iron-removed solution. The optimum condition of copper precipitation was found to be pH of 5.5, 25 °C and 45 min with 98.69% recovery. One of the advantages of this process was production of Na2SO4 with 99.1% purity after vaporization of the remaining solution from two-step precipitation. The obtained copper concentrate was leached at approximately the same condition of the first leaching step, and then the provided pregnant solution proceeded to an electrowinning cell with lead alloy anode contained antimony and steel sheet cathode under the following condition: temperature of 50 °C, reaction voltage of 2 V and current density of 300 Am− 2. Finally, a scale-up experiment was carried out and the copper cathode with 99.99% purity produced.

The application of D2EHPA in zinc solvent extraction has extensive background. To utilize more effectively, response surface methodology was used to optimize the concentration condition of zinc pregnant solution (ZPL) extracted by D2EHPA. In the current research, zinc, iron and manganese extraction along with separation factor of zinc-iron (Sf (Zn-Fe)) and zinc-manganese (Sf (Zn-Mn)) were considered as the response values. The optimal ZPL conditions extracted with 30% D2EHPA as the extraction solvent were as follows: Zn 21.96 g/L, Fe 382.57 ppm, Mn 1 g/L, Sf (Zn-Fe) 8.26 and Sf (Zn-Mn) 1529.82. In addition, it was found that the iron and manganese concentration were the most effective factors affecting the zinc and manganese extraction, respectively.

The sulfuric acid leaching of zinc plant residues was studied in an attempt to find a suitable hydrometallurgical method for zinc recovery. The parameters evaluated consist of reaction time, Solid-to-liquid-ratio, reaction temperature, agitation rate and pH. The results of kinetic analysis of the leaching data under various experimental conditions indicated that there is a reaction controlled by the solution transport of protons through the porous product layer with activation energy of about 1 kJ/mol for different constant solid to liquid ratios. Based on the shrinking core model (SCM), the following semiempirical rate equation was established:1-3(1-a)2/3+2(1-a)=0.001187×[H+]0.016×[(S/L)]-1.34×exp(-1/RT) ×t. On the other hand, activation energy was obtained from a model-free method using isothermal measurements. Values for activation energy were calculated as a result of the conversion function with an average of 2.9 kJ/mol. This value is close to that determined previously, using shrinking core model (SCM).

A three-stage methane gas (CH4) reduction of manganese ore, dissolution, and precipitation from solution procedure was conducted to synthesize MnO2 nanorods. Methane gas reduction was carried out at 850, 875, 900, 925, and 950 °C for 120 min. Precipitation of the α-MnO2 nanorods was performed in the temperature range of 25–90 °C with a constant reaction time of 90 min. The morphology and particle size of the products were determined from scanning electron microscope (SEM) images and X-ray diffraction (XRD) patterns. The BET and BJH of the products were found out by the surface area analyzer. Reduction results indicated that MnO-rich phase is significantly formed at 950 °C as MnO2 phase disappears. Precipitation results also showed an average diameter size of ∼ 50 nm for the embedding α-MnO2 nanorods with BET surface area of 174 m2/g.

Annually, a great amount of Cold Filter Cake (CFC) residue is produced in lead and zinc plants in Iran. CFC, the Ni–Cd purification step residue, with elemental analysis of 45–55% Zn, 11–16% Cd, 2.5–4% Ni and 0.2–1.5% Cu can be used as a secondary source of zinc, nickel and cadmium. Unfortunately, despite its hazardous content, the waste is treated in such a way that most of the hazardous constituents get easily exposed to the environment. In the present study, treatment of CFC for separation of zinc and recovery of cadmium has been conducted employing alkaline leaching and electowinning processes respectively. Affecting factors, such as alkaline concentration, leaching temperature, holding time, and liquid:solid ratio (volume/weight, L/S), and stirring rate were studied on zinc leaching from CFC with NaOH solution. Results showed that under two-stage alkaline leaching, when CFC was leached for 1 h at 75 °C in the presence of 8 mol/l sodium hydroxide and liquid:solid ratio of 10:1, the leaching rate of Zn was 98% while Ni and Cd remained intact in residue. Thus, zinc was separated from Ni and Cd selectively. In the next step, sulfuric acid leaching of alkaline leach residue, with 1 mol/l H2SO4 at 70 °C and liquid/solid ratio of 10:1, provided a stock sulfate solution containing 49.92 g/l Cd, 2.77 g/l Zn, 12.76 g/l Ni, and 1.18 g/l Cu. The Cd electrowinning experiments and the effects of current density (A/m2) and electrolysis time (h) were studied on stock sulfate solution in the presence of Ni and Zn. The electrowinning was carried out at various current density (100, 180, 250 and 400 A/m2) and electrolysis time (5, 8, and 18 h). Concentrations of Cd2 +, Ni2 + and Zn2 + were determined every hour. Cadmium recovery of 9.8% at 100 A/m2 increased sharply to 87% at 400 A/m2 in 5 h. Maximum total recovery of cadmium from solution was 99.9% achieved at 25 °C and 400 A/m2 in 18 h. Final product was cadmium sponge of high purity 98% and copper was considered the main impurity in cadmium sponge deposit (1.7%). A complete flow sheet of process for the selective separation of zinc and cadmium recovery is presented.

The leaching kinetics of the cobalt present in zinc plant purification residues (cobalt filtercakes) was investigated in the presence of the aromatic reducing agent — phenol — without any pretreatments. As such, the effects of particle size (− 75 + 53 to − 180 + 150 μm), reaction temperature (25 to 85 °C), phenol dosage (2 to15%) and sulfuric acid concentration (0.5 to 2.0 M) on the dissolution rate of cobalt and manganese were investigated. The experimental data conformed well to the shrinking core model, and the dissolution rate of cobalt was found to be controlled by diffusion through the product layer. Due to the low percentage of cobalt (< 1%) in the filtercake, SEM and XRD studies did not show any cobalt-containing phases. It was demonstrated that sulfuric acid with phenol addition can effectively dissolve cobalt and manganese present in the filtercake. The activation energy of the reaction was calculated as 22.92 kJ/mol. The order of the reaction with respect to H2SO4 concentration, particle size, and phenol dosage was found to be 0.61, − 1.88 and 0.70, respectively. The optimum conditions for the leaching of cobalt and manganese were found to be as 2.0 M H2SO4, 85 °C, − 75 + 53 μm, and 10% phenol at the stirring speed of 600 rpm, which correspond to more than 97% cobalt and ~ 100% manganese recoveries. The rate of cobalt leaching based on shrinking core model can be expressed by a semi-empirical equation as: 1−23X−(1−X)2/3=k0.[H2SO4]0.61⋅r0−188.[phenol%]0.70.exp(−22.92RT)⋅t.

The electrowinning of cadmium in the presence of zinc has been investigated. As such, the effects of initial cadmium concentration (40–160 g/L), current density (50–150 A/m2), free acid concentration (0–196 g/L), initial zinc concentration (20–120 g/L), and temperature (25–70 °C) on current efficiency (CE) and specific power consumption (SPC) in electrowinning of cadmium have been examined. Under the optimum working conditions (initial cadmium concentration 120 g/L, current density 80 A/m2, free acid concentration 98 g/L, and temperature 40 °C), CE and SPC were 98% and 1.21 kWh/kg, respectively. The presence of zinc (20–120 g/L) had no systematic effect on CE and SPC, but decreased the purity of the deposited cadmium down to 99.92%.

The cementation of lead with aluminum powder from the brine-leaching filtrate of lead-bearing zinc plant residues (ZPRs) has been investigated. The influence of several parameters on the course of the reaction such as aluminum powder quantity, time, aluminum powder particle size and temperature was examined. The optimum cementation conditions were found to be as aluminum powder quantity = 1.5 times of stoichiometric value, time = 90 min, aluminum powder fraction with a particle size − 125 + 88 µm and temperature = 70 °C. Cementation of lead was shown to be a feasible process to achieve a high degree of lead removal (~ 95%) within a fairly reasonable contact time. Also, the kinetics of the cementation was studied over a range of experimental parameters. The cementation was shown to be a diffusion-controlled reaction in the range of 50–70 °C with activation energy of 23.6 kJ/mol. However, the process was found to be highly temperature sensitive in the range 40–50 °C, suggesting a chemical reaction control model for that range. This finding was in concordance with a severe slope change in Arrhenius curve in the range 40–50 °C.

As the scope of the present work, zinc which is engaged in Cd–Ni zinc plant residues has been successfully recovered. While zinc contained in Cd–Ni filtercake has a considerable economic value, cadmium present in the material is one of the major sources of cadmium contamination of the environment. A novel hydrometallurgical technique was utilized for the separation of zinc from cadmium and nickel. Moreover, high purity cadmium and nickel were advantageously separated in successive cementation stages which in turn served as an environmental breakthrough. The process was composed of the collective leaching of the filtercake in order to bring as much metal as possible into the solution followed by selective cementation of cadmium and nickel, respectively. Therefore, a stock solution was provided under the optimum leaching conditions being as t = 150 min, T = 45 °C, stirring speed = 500 rpm, S/L = 1/13, pH = 1.3–1.5 and then cadmium was separated through cementation with zinc plates while nickel was cemented using zinc powder. The optimum working conditions for cadmium cementation were found to be as T = 25 °C, t = 180 min, pH = 2–2.5, stirring speed = 20 rpm and active plate area = 100 cm2/L, while those of nickel were as T = 80 °C, t = 15 min, pH = 3.9–4.1, stirring speed = 50 rpm and zinc powder amount equal to stoichiometric amount needed for the cementation of nickel. Under such conditions, 98.95% of cadmium and 99.98% of nickel were recovered from the solution. Finally, a simple and effective conceptual flow diagram was proposed for the process.

Orthogonal array design (OAD) has been used to arrange the experimental runs in order to maximize lead recovery and also minimize zinc dissolution from an Iranian zinc plant residue (ZPR). The residue assaying 9.21% Zn, 12.37% Pb and 5.31% Fe was used in successive processes involving sulfuric acid (H 2 SO 4 )–water leaching to bring zinc into solution and therefore enrich the lead in the residue and then brine leaching for lead recovery. An L 25 orthogonal array (OA, five factors in five levels) was employed to evaluate the effects of NaCl concentration ( C brine = 50, 150, 200, 250, 300 g/L), time ( t = 10, 20, 30, 40 and 50 min), pulp density ( S / L = 25, 50, 100, 150 and 200 g/L), pH (1, 2, 3, 4, 5) and stirring speed ( R = 100, 200, 300, 400 and 500 rpm) on extraction of the individual metal ions (Pb, Zn, Fe). Statistical analysis, ANOVA, was also employed to determine the relationship between experimental conditions and yield levels. The results showed that increasing pulp density and initial pH reduced mean responses for lead. However, increasing brine concentration promoted the recovery of lead. The experimental results for selective leaching showed that under optimal leaching conditions ( S / L ) 1 : 25 g/L, (pH) 1: 1, t 3 :30 min, ( C brine ) 5 : 300 g/L and R 4 : 400 rpm, the extraction of lead and zinc were 89.43% and 3.84%, respectively; giving a solution containing 2.99 g/L Pb, 0.02 g/L Zn and negligible amount of iron.

This research is part of a continuing effort to reduce environmental conflicts and occupational hazards of cadmium bearing zinc plant residues (ZPR's) and to break through this problem and recover valuable constituents of the wastes. In this paper, effects of influential factors on extraction efficiencies of Cd, Zn, Ni, Pb and Cu from Cd–Ni filtercake as a major ZPR were investigated. In the view of above, the systematical and analytical evaluation method of Taguchi quality engineering has been applied for the leaching of the Cd–Ni filtercake to evaluate the optimal experimental conditions and hence to achieve the highest leaching performance and the best robustness of quantitation from the least number of trials in a batch laboratory scale. An L25 orthogonal array (OA, five factors in five levels) was employed to evaluate the effects of temperature (T = 25, 35, 45, 55 and 65 °C), acid concentration (C = 0.31, 0.51, 0.82, 1.22 and 1.84 M), time (t = 15, 30, 45, 60 and 90 min), pulp density (S/L = 125, 142.86, 166.67, 200 and 250 g/l) and stirring speed (R = 200, 300, 400, 500 and 800 rpm) on extraction percent of the individual metals. Statistical analysis, ANOVA, was also employed to determine the relationship between experimental conditions and yield levels. The results showed that increasing temperature reduced performance characteristics. Two approaches were taken into consideration for the experiments, i.e., selective and collective leaching, and then the optimum conditions were sought for each considered approach. The experimental results for selective leaching showed that under optimal leaching conditions (T: 45 °C, C: 0.82 M, t: 60 min, (S/L): 250 g/l, R: 400 rpm), the extraction of cadmium, zinc, nickel, lead and copper were 90.29, 96.86, 8.66, 0.11 and 0.01, respectively. However, the same results for optimal collective leaching conditions (T: 25 °C, C: 1.22 M, t: 60 min, (S/L): 142.86 g/l, R: 200 rpm) showed that the extraction of Cd, Zn and Ni could be over 95%, with negligible dissolution of lead and copper.

The orthogonal array design has been used to determine the optimum conditions for the cementation of cadmium from a synthetic zinc sulfate solution and hence to achieve the highest cementation efficiency and the best robustness of the quantitation from the least number of trials in a laboratory scale. Cementation was performed using zinc powder. The orthogonal array L 9 (3 4 ) that comprises four parameters at three levels was chosen. The parameters and their levels were as: zinc powder quantity ( Z ): 1, 1.2 and 1.5 times of stoichiometric quantity of cadmium; time ( t ): 4, 8 and 12 min; temperature ( T ): 25, 35 and 45 °C and zinc powder particle size distribution ( R ): nos. 1–3. The ultimate optimum cementation conditions were found to be Z 1 (equal to stoichiometric quantity of cadmium), T 1 (25 °C), R 3 (no. 3) and t 1 (4 min). Under these conditions, cementation percentages for cadmium and nickel were 95.83% and 6.63%, respectively.

Cadmium is a toxic metal, which is primarily produced as a by-product from mining, smelting and refining of sulphide ore concentrates of zinc. Secondary cadmium is recovered from spent Ni–Cd batteries. Some cadmium is also recovered from zinc sinter plant fume, EAF dust, cadmium containing alloys, cadmium containing fluorescent materials, etc. In all cases cadmium is associated with some other impurities depending on the source. The hydrometallurgical processing is very effective for treating such materials because it can control the different levels of impurities. The most common lixiviant used to dissolve the desired metals is sulphuric acid. In the present paper, the hydrometallurgical processes have been described for the recovery of cadmium from various resources using sulphuric acid as the main lixiviant. The leach solution obtained has been purified using cementation or solvent extraction methods. The metal is then produced from the purified solution by electrowinning or precipitation with zinc dust and melting.

Metal pH50 and ΔpH50 values were determined and compared for a number of organic systems to select a synergistic system for the separation of Cd from Zn. The largest ΔpH50(Zn–Cd) value was obtained with HRJ-4277 (nonylsalicylic acid)/Cyanex 471X (triisobutyl phosphine sulphide or TIBPS) system and reached 1.53 pH units, indicating excellent separation of Cd from Zn with this system. Shellsol 2046 was chosen as the diluent of the system due to its larger ΔpH50(Zn–Cd) and excellent phase separation compared with other diluents. Distribution isotherms and subsequent McCabe–Thiele diagrams indicate that 99% Cd could be extracted in two stages at pH 3.80 and A / O ratio of 1.5 : 1. The extraction kinetics and stripping kinetics of Cd and Zn were very fast with > 97% metal transferred within 30 s. In a three-stage batch continuous extraction test, > 99% Cd and 0.4% Zn were extracted. The loaded organic solution contained 3.87 g/L Cd and only 12 ppm Zn. In a two-stage batch continuous scrubbing test, all the co-extracted Zn was removed with the scrubbed organic containing

Simultaneous extraction of zinc and cadmium by a mixture of di-2-ethyl hexyl phosphoric acid (DEHPA) and mono-2-ethyl hexyl phosphoric acid (MEHPA), and the synergistic effect of MEHPA on co-extraction of zinc and cadmium with DEHPA have been investigated. It was shown that the extraction of zinc and cadmium by DEHPA and/or MEHPA can be increased by an increase of pH. The results also illustrate that the pH 0.5 of zinc, by increase in MEHPA from 0.1 to 8 vol%, did not vary significantly while that of cadmium varies from 0.63 to 2.4. In other words MEHPA prevents selective extraction of zinc from aqueous media containing both zinc and cadmium. Studies on the selective separation parameter β Zn/Cd show that it was reduced 50 to 13 times by an increase in MEHPA from 0.1 to 8.0 vol%. In this paper, the effects of DEHPA and MEHPA concentration on co-extraction of zinc and cadmium have been investigated.

Addressed is the effect of heat treatment on the dissolution behavior of zinc, cadmium and nickel from Cd–Ni zinc plant residues. The proposal was considered for possible positive effect of heat treatment on the separation of nickel present in the residue, due to some phase transformations. The leaching experiments were carried out with the heat-treated residues at identical leaching conditions. The residues were heat treated at 100–1000 °C for 1 h and water-leached after subsequent cooling down to the room temperature. The results obviously showed that the heat treatment necessarily reduces the dissolution of zinc but does not affect the dissolution of cadmium. However, the recovery of nickel was increased by 83.80% upon the temperature increase from 25 to 800 °C, whereas at temperatures above 800 °C, the recovery of nickel did not change. This work offers a possible vehicle for the separation of nickel from the residues. The increase in nickel dissolution was attributed to the formation of water-soluble nickel sulfate which was further verified by the XRD detection.

A hydrometallurgical treatment involving solvent extraction of zinc using di-2-ethylhexyl phosphoric acid (D2EHPA) has been investigated to recover zinc from an industrial leach residue. The residue was leached with sulfuric acid producing leach liquor which was subjected to solvent extraction for enrichment of zinc and removal of impurities. Operating variables, such as pH, D2EHPA concentration, temperature, aqueous/organic (A/O) phase ratio, tri-butyl phosphate (TBP) concentration and sodium sulfate (Na 2 SO 4 ) concentration in aqueous phase were studied. Practically, all zinc was extracted from the aqueous solution at pH 2.5 with 20% w/w D2EHPA in kerosene. Increasing either TBP concentration up to 5%, or Na 2 SO 4 concentration up to 0.2 M, increased the zinc extraction. Zinc could be extracted at one theoretical stage at A/O of 1/1, as calculated by McCabe–Thiele method.

Cd-Ni filtercakes are produced continuously at the third purification step in the electrolytic production of zinc in the National Iranian Lead and Zinc Company (NILZ) in northwestern Iran. In this research, the dissolution kinetics of cadmium from Cd-Ni residues produced in NILZ plant has been investigated. Hence, the effects of temperature, sulfuric acid concentration, particle size and stirring speed on the kinetics of cadmium dissolution in sulfuric acid were studied. The dissolution kinetics at 25-55 degrees C and t