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Solvent extraction and electrowinning (SX/EW) play an important role during the hydrometallurgical production of copper for low-grade ores. To better understand the rapid Cu extraction/strip kinetics for producing a high-purity CuSO4–H2SO4 electrolyte and elucidate the method to maximize current efficiency by preventing leach carryover to electrolyte during SX and minimizing the iron concentration, a detailed speciation study is required to quantitatively describe the iron and copper distribution in the SX/EW solutions. In the present study, by virtue of our previous modeling work, thermodynamic modeling of the SX/EW solutions is carried out to quantify the distribution of iron (in the form of ferric and ferrous), copper, and sulfuric acid. The developed model provides a mathematical tool capable of quantifying the concentrations of free ions and complexes in terms of temperature (25–60 °C) and solution composition under simulated industrial conditions. Finally, oxidation reduction potential measurements are employed to validate the model.

The process of Minera y Metalurgica del Boleo recovers cobalt from a complex oxide and sulfide ore by solvent extraction and electrowinning after the recovery of copper and zinc (as copper cathode and zinc sulfate monohydrate). As part of the process development, cobalt electrowinning was systematically studied with respect to temperature, pH, pitting behavior, deposit morphology, deposition stress and interfacial tension. Electrowinning test work was conducted in a bench scale diaphragm cell in a sulfate medium with control of the catholyte pH. Cobalt was deposited on a coated stainless-steel mandrel cathode with exposed circles to allow plating of disk shaped deposits over the surface of the cathode that could be easily stripped after 5–7 days of deposition. Limiting the area of plating prevents the detachment of the cobalt deposit from the cathode due to high deposition stress. The morphology and pitting behaviour varied for different commercial cobalt raw materials used for making cobalt sulfate solution as they contained different impurities or additives. The deposition behaviour changed after the solutions had an activation carbon treatment as some impurities (mainly organic compounds) were removed. In the temperature range from 50 to 65 °C, a higher temperature resulted in a brighter deposit with more pits and smaller lateral dendrites. The cobalt deposition current efficiency was around 98% at pH 3.0 and was not sensitive to temperature. With decreasing catholyte pH from 3.5 to 2.0, the cobalt current efficiency decreased from 99 to 88% and the growth of lateral dendrites were significantly suppressed. The formation of pits resulted from the adhesion of hydrogen bubbles on the cobalt deposit. The adhesion of hydrogen bubbles was determined by the prevailing surface and interfacial tensions. Anti-pitting agents were used to change the interfacial and surface tensions in such a way that the detachment of bubbles was facilitated, resulting in a pit-free cobalt deposit. The anti-pitting agents affected the deposit morphology and internal stress. The impact of the anti-pitting agents on the disengagement of organic and aqueous phases during cobalt solvent extraction was also investigated. The internal stress of the cobalt deposit as a function of temperature and anti-pitting agent dosage and the conductivities of the CoSO4-Na2SO4-H2SO4 system are reported.

A comparison study about the extraction and separation of Cu2+ and Ni2+ with Lix984N in a microchannel reactor and separating funnel has been conducted. The results showed that, in the microchannel reactor, the overall volumetric mass transfer coefficient of copper was 20 times that of nickel, whereas in the separating funnel, it was only 2 times that of nickel. In addition, the separation coefficient of copper and nickel in the microchannel reactor was 5 times that of the conventional one. Typically, at initial pH=2.5, contact time 1.95 s, volume fraction of extractant Lix984N 15% and within 1.9 g l−1 nickel ion concentration, the extraction rate of copper was higher than 95%, but the nickel was hardly extracted. In comparison, it needed almost 50 s to reach a Cu extraction of 95% in the separation funnel, with more than 5% Ni co-extraction rate. Although the microfluid extraction showed excellent extraction performance, there exists a need to further improve its processing capacity to apply it to industrial production. Furthermore, scaling up the microreactor has become increasingly promising under the fast developing 3D printing technology.

Quantitative data have been obtained on the distribution of sulfuric acid in ROH - H2SO4 - H2O systems in a broad acid concentration range for aliphatic alcohols of the ROH (R = C8-C10) homologous series. Alcohols had different structures. A correlation has been established between the extractive ability of the alcohols and the acid concentration, alcohol structure, temperature, time of phase contact, and phase volume ratios. The data obtained on the alcohol solubility in aqueous H2SO4 solutions allowed us to estimate the extragent loss in practical hydrometallurgical applications. Optimal conditions have been determined for the process of sulfuric acid recovery and purification from copper electrolyte waste. The conditions have been tested in continuous mode at a cascade of laboratory extractors.

The reduction of iron ions concentration from copper electrowinning electrolyte solution in copper solvent extraction and electrowinning process, is important to maintain good current efficiency. In this study both chemical precipitation (hydroxide and jarosite) and ion exchange methods were investigated using industrial copper raffinate and electrowinning solutions from Sarcheshmeh copper complex, Iran. Batch precipitation and adsorption tests were initially performed before, a fixed bed column was investigated for iron removal by ion exchange. Batch results indicated that with increasing the pH and dilution, improved the iron removal by chemical precipitation methods. Hydroxide and jarosite precipitations were performed efficiently at pHs of 3 and 2 and temperatures of 70 °C and 95 °C respectively. The iron removal efficiency was greater with jarosite precipitation (∼70%) compared to hydroxide (∼20%) for a particular pH (2). Diluting by 50% increased the removal efficiency up to 150%. Ion exchange results indicated that pH and dilution did not have a significant effect on the efficiency with this method and it practically performed at ambient temperatures. The batch equilibrium and kinetics experiments were performed with Purolite S957 resin and the results showed that the Sips isotherm (the maximum adsorption capacity was 42 mg/g) and both the pseudo-nth order (kad = 0.00083 mg1−n/g1−n·s and n = 1.92) and Langmuir-Freundlich kinetics models ( k a LF = 0.08 mL min - 1 mg - 1 , k d LF = 0.12 min - 1 and n = 0.12) were able to predict the equilibrium and kinetics data. In the fixed bed column experiments, the effects of parameters such as the volumetric flow rate and bed height on the breakthrough curves were studied. The iron removal efficiency for the industrial solution which included impurities was less than for synthetic solutions. The results showed that the resin selectively removed the iron ions but did not decrease the copper concentration and used resin was completely regenerated by hydrochloric acid (the amount of 5 mL/g of resin and concentration of 30 wt%) over several regeneration did not significantly alter the sorption capacity. Finally, a mathematical model based on chemical sorption was used to predict the experimental breakthrough curves data and results showed this model accurately predicted the experimental breakthrough curves. Overall, ion exchange was more efficient at iron removal than chemical precipitation.

Progress towards a sustainable industrial process for the recycling of unwanted batteries using a deep eutectic solvent is presented. The effect of lactate anions on the recovery of metals from chloride media by extraction by a lipophilic chloride ionic liquid has been investigated with using solvent extraction experiments with a special emphasis on the behaviour of cobalt and nickel. A new solvent extraction system for the recovery of both cobalt and nickel from acidic (lactic acid) media is presented. The relative impacts of the new extraction system and an aliquat system on both workplace air and the wider environment are compared. The new system was found to pose a smallet threat to both workers and the environment.

Copper and zinc were recovered from spent chloride brass pickle liquor using solvent extraction and precipitation-stripping. Speciation diagrams reveal that bivalent copper and zinc and their monovalent chlorocomplexes were present in the liquor. Copper was extracted by hydroxyoxime LIX84IC in kerosene while zinc was extracted using D2EHPA. 98.8% copper was extracted at an equilibrium pH of 2.5 using 30% v/v LIX 84IC whereas 98.3% zinc was extracted at an equilibrium pH of 2 using 15% v/v D2EHPA. Precipitation-stripping of loaded organic phase using oxalic acid resulted in recovery of copper and zinc as water insoluble oxalates. Solubility plots developed on theoretical considerations for copper-oxalic acid and zinc-oxalic acid system predicted the yield of precipitates in close agreement with experimental values. 97.3% of both copper and zinc present in the pickle liquor were recovered as their oxalates. The organic phase could be used for multiple cycles. Calcinations of the oxalates to their oxides did not induce any change in the ordered structure observed in oxalates. Morphology of the precipitates revealed highly agglomerated particles. Copper oxide 100% pure and zinc oxide with 99.1% purity were obtained as value added products. A process flow diagram of the proposed scheme is presented that could be implemented to generate wealth from waste.

The purification of coal fly ash (CFA) leach liquor has been carried out using an integrated solvent extraction and precipitation technique. The integrated technique was employed to primarily remove two dominant impurities, Ti and Fe, from CFA leach liquor. The solvent extraction experiments were carried out at varying factor values of contact time, aqueous-to-organic ratio, and extractant concentration. Preliminary results showed no loading of Fe2+ onto the organic phase, partial loading of Fe3+, and a high loading of Ti4+. Unable to adequately load onto the organic phase due to the high CFA leach liquor acidity at pH 0.1, Fe3+ required reduction to Fe2+ in order to bypass the solvent extraction stage. To achieve this, an integrated solvent extraction and precipitation technique was formulated and adopted. The technique involved iron reduction from Fe3+ to Fe2+ as the first treatment of CFA leach liquor, followed by solvent extraction of Ti4+ and subsequent recovery of titanium as TiO2, crystallization of Al3+ and the precipitation of Fe2+, in the form of Fe(OH)2, as the last treatment. Alumina was recovered as a precipitate product in the form of (NH4) Al(SO4)2·12H2O, and the calcined alumina product contained 99.91% Al2O3. Primene JMT was used in small amounts of ~10% in kerosene for short contact times of ~15 min using an A/O ratio of 1:1 in order to achieve high extraction efficiencies during the solvent extraction step. The results presented in this paper identify the integrated precipitation and solvent extraction route as an effective option for CFA leach liquor purification and alumina recovery.

Flotation tailings from copper production are deposits of copper and other valuable metals, such as Mo, V and U. New hydrometallurgical technologies are more economical and open up new possibilities for metal recovery. This work presents results of the study on the extraction of copper by mixed extractant consisting p-toluidine dissolved in toluene. The possibility of simultaneous liquid–liquid extraction of molybdenum and vanadium was examined. D2EHPA solutions was used as extractant, and recovery of individual elements compared for the representative samples of ore and copper flotation tailings. Radiometric methods were applied for process optimization.

Work is devoted to studying electrowinning of precious metals, in particular gold, in rich cyanide solutions on three-dimensional cathodes of electronegative metals. The object studied is production solution obtained during cyanide leaching of gravitation concentrates containing about 95 g/m3 of gold, 30 g/m3 of silver, and 6.5 g/dm3 CN–. The effect of current density, electrolyte volume consumption, and cathode material on the degree of gold extraction are studied. An operating regime is established making it possible to extract more than 99% gold with a residual concentration below 0.1 g/m3.

We studied the separation and recovery of copper(II), nickel(II), cobalt(II), zinc(II), and cadmium(II) from magnesium and calcium, using synergistic solvent extraction (SSX) in a typical hydrometallurgical waste solution. A mixture of Versatic 10 acid and Mextral 984H, diluted with Mextral DT100, was used to obtain fundamental data on pH and distribution isotherms, as well as the kinetics of extraction and stripping. We also investigated the main effects and interactions of common solvent extraction factors: the extraction pH at equilibrium, the temperature, and the extractant concentration. The synergistic effect for extracting metals was confirmed. The results showed that the addition of Mextral 984H enhanced the separation factors of copper, nickel, cobalt, zinc, and cadmium over magnesium and calcium. Compared with Versatic 10 acid alone, for a mixture of 0.5 mol·L− 1 Versatic 10 acid/0.5 mol·L− 1 Mextral 984H, ΔpH50 values of copper, nickel, cobalt, zinc, and cadmium were found to be > 2.0, 3.30, 2.85, 0.95, and 1.32 pH units, respectively. The ΔpH50(Zn–Mg) and ΔpH50(Zn–Ca) values were 3.27 and 2.25, respectively, indicating easy separation and recovery of copper, nickel, zinc, cobalt, and cadmium. The extraction and stripping of copper, cobalt, zinc, and cadmium were fast, with 90% of the metal transferred in 2 min. We next studied whether the metals could be stripped from the extracted liquid selectively in sequence, by using sulfuric acid at different concentrations. The influence of the molecular structure of the oxime and carboxylic acid components upon the synergistic effects was identified by numerical analysis. Excellent separation of copper, nickel, cobalt, and zinc over magnesium and calcium was achieved with this synergistic solvent extraction system.

A brass ash sample was characterized, presenting a high content of Zn (50.1%) and Cu (18.9%), with zincite, tenorite, quartz, willemite, and metallic Zn and Cu as major phases. It is a heterogeneous material composed of agglomerated particles of different shapes and sizes. Aiming at reaching the requirements for the subsequent Zn electrolysis step in a hydrometallurgical route to recover Zn and Cu, the initial content of Cl (0.14%) and F (0.22%) in the brass ashes was reduced by oxidative roasting and alkaline leaching. In the oxidative roasting method, the increase in roasting temperature from 400 °C to 900 °C favored the removal of Cl and F (nearly 60% average) from the brass ashes; however, a high Zn evaporation efficiency (18.4% average) was observed. The thermodynamics of the reactions involved in Cl and F removal from the brass ashes was evaluated to explain the experimental results. In contrast, remarkable selectivity in removing Cl and F was obtained by leaching the brass ashes with Na2CO3 solutions, and 100% Cl and 89% F were removed with only 0.05% Zn and 0.2% Cu co-extractions using Na2CO3 (14% w/w) after 2 crosscurrent stages, both operated at 90 °C, L/S ratio of 4 mL g−1, for 30 min, and without pH control (pHinitial ≈ pHfinal ≈ 10.8). Practically no Cl and low F contents were found in the Zn/Cu-rich sulfuric leach solution, which proved to be appropriate for the hydrometallurgical treatment of brass ashes when using subsequent operations such as solvent extraction and electrowinning.

Copper oxide minerals are normally extracted by acidic leaching followed by copper recovery with solvent extraction and electrowinning. However, copper oxide deposits often containing large amounts of acid consumable gangue leads to a very high acid consumption. In addition, if the oxide deposit also contains precious metals and iron bearing minerals, significant (lime) neutralisation costs are incurred to establish the conditions for subsequent cyanidation, with concomitant production of gypsum, potential formation of silver-locking jarosites and potential gel formation when acids interact with layer silicate minerals. In this study, an alternative aqueous alkaline glycine system has been employed to evaluate the batch leaching behaviour of copper oxide mineral specimens of the minerals azurite, chrysocolla, cuprite and malachite. The effects of glycine concentration and pH were investigated at ambient temperature and atmospheric pressure. Glycine concentration and pH both had a major effect on the copper extraction. Complete extraction of copper from azurite was achieved in

Ammoniacal leaching followed by solvent extraction is practiced for the production of copper sulfate from copper dross generated in lead refinery during the pyro metallurgical production of lead. Copper is leached from its dross as Tetra-amino-Copper (II)-carbonate complex in the solution. Copper leached solution is guided through the solvent extraction process, using LIX-54 as a solvent and sulphuric acids as stripping agent to generate copper sulfate solution of the desired grade. Due to limited availability of LIX-54, performances of other solvents having a base of cinnamate and β diketone groups are evaluated. Both the solvents and LIX-54 have been compared in fields of effective organic concentration, effective loading, extraction kinetics, extraction isotherms, stripping activity and the optimized values are 20% organic concentration, 20 g/l Cu, 35 s, one theoretical stage for extraction and stripper acidity of 120 g/l respectively. Under these conditions, > 98% of copper has been extracted.

The enrichment of copper from copper–cyanide wastewater by solvent extraction was investigated using a quaternary ammonium salt as an extractant. The influences of important parameters, e.g., organic-phase components, aqueous pH values, temperature, inorganic anion impurities, CN/Cu molar ratio, and stripping reagents, were examined systematically, and the optimal conditions were determined. The results indicated that copper was effectively concentrated from low-concentration solutions using Aliquat 336 and that the extraction efficiency increased linearly with increasing temperature. The aqueous pH value and concentrations of inorganic anion impurities only weakly affected the extraction process when varied in appropriate ranges. The CN/Cu molar ratio affected the extraction efficiency by changing the distribution of copper–cyanide complexes. The difference in gold leaching efficiency between using raffinate and fresh water was negligible.

Iron removal is a common procedure in hydrometallurgy of non-ferrous metallic ores. Chemical precipitation and solvent extraction are currently the main methods employed for separation of iron from copper leach solution. However, chemical precipitation methods have the problem of producing large amounts of iron precipitates, and/or requiring high temperature and high pressure conditions for precipitation. The precipitates are difficult to separate from the solution. Emulsification is a general problem for solvent extraction, which causes loss of extractants and contamination of the electrolyte. A common problem for solvent extraction and chelating ion exchange processes is the difficult iron stripping process. Therefore, finding an efficient and environmentally friendly method for iron removal is of great significance. In this paper, new processes of using ion exchange resins to remove iron from copper leach solution have been investigated. Salicylic acid, amino carboxylic acid, amino phosphonic acid and hydroxy-oxime resins were employed to adsorb iron from a synthetic copper leach solution containing 40 g/L of Cu2 + and 36 g/L of Fe3 +. It was found that a hydroxy-oxime resin, named Z-Fe, has excellent adsorption selectivity on Fe3 +. Desorption behavior of the resin was studied by two step elutions using dilute H2SO4 and oxalic acid solution, respectively. After optimizing the adsorption-elution processes, the extent of removal of iron from the simulated copper leach solution using resin Z-Fe was 89.83%, while copper recovery could reach 100% in closed circuit processing.

In the rhenium solvent extraction process, molalities of perrhenate after extraction were measured from 278.15 to 303.15 K at impurity ionic strength from 0.2 to 2.0 mol·kg–1 in aqueous phase. Optimization condition of extraction was determined in this sulfuric acid medium. The standard extraction constants K0 at various temperatures were obtained by method of polynomial approximation. Thermodynamic quantities for the solvent extraction process were calculated.

The typical hydrometallurgical processing of nickel and cobalt ores involves the precipitation of a mixed nickel–cobalt intermediate product. Alternatively, cobalt could be separated from nickel while in the aqueous phase using solvent extraction. However, solvent extraction has some problems including the need for considerable area for the separators, management of fire hazard, and operational issues such as contamination with organic cross-over with sequential solvent extraction unit operations. The use of an ion exchange resin can mitigate some of these issues; however, conventional cation exchange resins are not sufficiently selective for cobalt over nickel. Solvent impregnated resin offers the engineering advantages of ion exchange resin and the chemical selectivity of the solvent extraction process. This paper describes key properties of a new commercial solvent impregnated resin, Lewatit® TP272, including cobalt and nickel loading as a function of pH and cobalt loading isotherms at pH 5.5 and 5.0 which are identified as the ideal conditions to recover cobalt. The chemical degradation of this solvent impregnated resin was also determined by exposing the resin to greater than pH 6 solutions, which are conditions outside the recommended operating range for this resin. The capacity of the degraded resin was restored by reimpregnating the resin using a Cyanex® 272-ethanol–water mixture. Optical images of resin containing the blue cobalt complex reveal the porous structure.

The development of a hydrometallurgical hydrochloric acid based leaching process for the treatment of nickeliferous laterites of various types is described. Work carried out in the Department of Mining and Mineral Engineering, Leeds University, U.K. over a number of years, on the development of a route for treating all types of lateritic nickel ore, is reported in fairly general terms. This paper presents an update on the work carried out and a discussion of a possible final process. Conceptually the process comprises leaching of the ores in azeotropic HCl; counter-current washing to maximise nickel recovery; separation of iron either by precipitation as goethite or by solvent extraction; cobalt recovery by solvent extraction with a tertiary amine; recovery of nickel by solvent extraction followed by electrowinning, hydrogen pressure reduction or pyrohydrolysis and regeneration of hydrochloric acid by pyrohydrolysis with by-product recovery of magnesia some of which might be used for acidity control in the process. Attempts to leach successive batches of ore in a cyclic or counter-current manner to build up metal concentration were unsuccessful due to passivation by formation of a product layer around the ore particles. A key process step is concentration of the nickel in the leach liquor (1–5 g/L) and its separation from magnesium and/or iron to a tenor compatible with recovery methods. This can be achieved using solvent extraction with Cyanex®301, Cyanex®302 or Versatic 10, the last of which displayed the best extraction characteristics. Cyanex®301 showed excellent selectivity for nickel but stripping required high concentrations of acid whilst Cyanex®302 showed appreciable co-extraction of magnesium. Whilst some of the process steps have been thoroughly researched, others remain to be proven and it was not possible to test the process as a whole or to carry out an energy balance, so that nothing can yet be stated definitively regarding the economics of the route but some factors regarding energy consumption in pyrohydrolysis are discussed. Relevant up to date literature on chloride based hydrometallurgical processes for laterites has been briefly reviewed. The contributions of individual research workers are acknowledged at appropriate points in the text.

High concentrations of iron represent a drawback in the hydrometallurgy of copper, since this metal cation could be co-extracted or entrained to the electrowinning step causing difficulties in obtaining pure copper. In this work, a pre-treatment step based on solvent extraction of iron using a hydrophobic imidazolium-based ionic liquid is proposed to overcome this drawback. The ionic liquid [bmim][Tf2N] was used as the diluent. The β-diketone TFA was used as the extractant. The results show that the combination extractant-ionic liquid is capable to remove iron(III) selectively from a pregnant leach solution (PLS) containing mainly copper(II) and iron(III) at low pH values. Furthermore, the use of this ionic liquid as diluent shows high extraction capacities. Finally, 97% of iron(III) can be stripped from the ionic liquid using a diluted solution of sulfuric acid (2 M).

Solvent extraction and separation of indium and copper from zinc residue reductive leach liquor containing a high concentration of zinc (ca. 85 g/L) and iron (ca. 35 g/L) was studied. Di-(2-ethylhexyl) phosphoric acid (D2EHPA) and Acorga M5640 were used as extractants for indium and copper, respectively. Indium was preferentially extracted by 15%(v/v) D2EHPA with three-stage extraction using an aqueous/organic (A/O) phase ratio of 6/1. Indium was effectively stripped from the D2EHPA organic phase by 4 M HCl. Copper was subsequently extracted selectively from the indium-extracted solution using 15%(v/v) Acorga M5640 with three-stage extraction using an A/O phase ratio of 2/1. Copper in the loaded Acorga M5640 organic phase was stripped by a synthetic spent electrolyte solution containing 25 g/L Cu and 180 g/L H 2 SO 4 . Based on the experimental results, a flowsheet for the separation and recovery of indium and copper from the reductive leach solution was developed and tested. The results showed that 95.4% of indium and 95.5% of copper were extracted, and the measured separation factors β In/Zn , β In/Cu , β In/Fe , β CuZn , and β CuFe were 4979, 1975, 3719, 1719, and 2446, respectively.

As part of the ongoing development of the Next-Generation Caustic-Side Solvent Extraction (NG-CSSX) process, the thermal stability of its process solvent, the Next-Generation Caustic-Side Solvent (NGS) was investigated and shown to be adequate for industrial application. The solvent was thermally treated at 35°C over a period of 13 months whilst in dynamic contact with each of the aqueous phases of the current NG-CSSX process, namely SRS−15 (a highly caustic waste simulant), sodium hydroxide scrub solution (0.025 M), and boric acid strip solution (0.01 M). The effect of thermal treatment was evaluated by assessing batch extract/scrub/strip performance as a function of time, by monitoring the sodium extraction capacity of the solvent, and by analysis of the solvent using electrospray mass spectrometry. Current studies indicate that the NGS should be thermally robust without replenishment for a period of 7 months in the Modular Caustic-Side Solvent Extraction Unit (MCU), which has been treating waste using ...

Highlights: • We developed an ecofriendly mediated electrochemical process for copper recovery. • The recovery of copper was achieved without mechanical pretreatment of the samples. • We identified the optimal flow rate for the leaching and electrowinning of copper. • The copper content of the obtained cathodic deposits was over 99.9%. - Abstract: The present study aimed at developing an original and environmentally friendly process for the recovery of copper from waste printed circuit boards (WPCBs) by chemical dissolution with Fe{sup 3+} combined with the simultaneous electrowinning of copper and oxidant regeneration. The recovery of copper was achieved in an original set-up consisting of a three chamber electrochemical reactor (ER) connected in series with a chemical reactor (CR) equipped with a perforated rotating drum. Several experiments were performed in order to identify the optimal flow rate for the dissolution of copper in the CR and to ensure the lowest energy consumption for copper electrodeposition in the ER. The optimal hydrodynamic conditions were provided at 400 mL/min, leading to the 75% dissolution of metals and to a low specific energy consumption of 1.59 kW h/kg Cu for the electrodeposition process. In most experiments, the copper content of the obtained cathodic depositsmore » was over 99.9%.« less

The paper is focused on the study of hydrometallurgical processes for recovery of rare earths (RE) from fluid catalytic cracking catalysts (FCCC). According to the experimental results two processes were proposed: in the first one RE were recovered as double sulfates by selective precipitation with sodium hydroxide after leaching with sulfuric acid. The second approach consisted of dissolution of powder by acids, solvent extraction by (2-ethylhexyl)phosphoric acid (D2EHPA), stripping and precipitation of RE oxalates using oxalic acid. Experiments showed that solvent extraction was found to be beneficial in terms of achieving improved final products quality with greater purity (∼98%).

Electronic waste treatment within EU has focused on valuable metal recycling and has been defined by the EU directives. The difficulties for recycling are induced by the high complexity of such waste. In this research, a hydrometallurgical process was developed to recycle copper from industrially processed information and communication technology (ICT) waste. By using air as the oxidant and ammonia–ammonium carbonate leaching solution, copper could be extracted with high recovery – more than 90%, and high extraction selectivity – around 98%. In order to understand the copper extraction process and the reaction mechanisms, the effects of a range of parameters during copper leaching were comprehensively investigated, including ammonia concentration, leaching temperature, ammonium carbonate concentration, the liquid-to-solid ratio, air flow rate and mechanical stirring rate. The controlling step for the leaching kinetics was identified and the effects of different parameters were investigated. This research is potentially beneficial for further optimisation of the copper leaching process and the whole process design for copper recycling after incorporating with solvent purification and electrowinning of the copper-rich solution.

A combined pyro-hydrometallurgical treatment processes involving leaching, cementation, electrowinning, and melting have been applied to industrial brass wastes, brass ash, and flue dust to recover copper and zinc. In the first stage, the effects of leach parameters on metal recovery ratio were studied. Almost all zinc was dissolved from brass wastes in conditions of 1 M H2SO4, 120 min, 60°C, 400 rpm stirring speed, and 1/10 solid liquid ratio. Negligible amount of copper was also dissolved in the same conditions and this copper could be precipitated by cementation. Pure zinc of 99.4% and 99.5% were produced by electrowinning of purified brass ash and flue dust leach solutions, respectively. Leach residue containing high amount of copper was melted with various fluxes and more than 99% of copper recovery ratio was achieved.

This study includes solvent extraction and electrowinning of copper from leaching process. The Taguchi method was usedto determine the optimum conditions for solvent extraction process of copper ore in sulfuric acid solution. The experimentalparameters of solvent extraction were: Extractant mass of sodium diethyldithiocarbamate (SDDT), agitation time, and pHvalue. While, the electrowinning parameters were constant at 25°C temperature, reaction voltage of 2 V, 200 A/dm2current density, and 2.5 cm distance between electrodes. The following optimum solvent extraction parameters levels were found: Extractant mass of SDDT 100 mg, agitation time 30 min, pH = 8. Under these conditions, the extractionratio was 97.15%. It was observed that the extraction ratio increases with the increase of SDDT mass. A contactingtime of 30 min is considered for ensuring state of equilibrium, and the extraction ratio decreases in a more alkalinemedium (pH >8) when the cation of hydrolysis at the aqueous phase was sufficiently disturbed from heating, allowingelements to return to the aqueous phase, thereby decreasing the extraction ratio. After electrowinning is carried out, theresults for the experiments were taken to quantitative analysis using the X-ray fluorescence pectrophotometer, scanningelectron microscopy, and X-ray diffraction patterns which showed the presence of pure copper at a purity level of 100%.

The continuous growth of the electronic equipment market has led to an increased amount of scraping that it becomes necessary to recover. A hydrometallurgical method for copper and precious metal recovery from e-waste must consist of a number of steps: leaching, ion separation and subsequent electrochemical re-deposition of the target metal. Although this task is achievable in aqueous solutions, it requires strong acid or cyanide solutions. The aim of the study is to develop a new environmentally benign process by using a Deep Eutectic Solvent (DES), a form of cheap and safe ionic liquid, as an electrolyte for both leaching and electrodeposition. The experiments were conducted in a mixture of choline chloride and ethylene glycol, under 20 kHz ultrasonic agitation. In optimized temperature and sonication conditions, the leaching rate of copper attains 65 mg/h cm 2 . An interesting feature is that, in this media, copper dissolves into the form of Cu(I) complex instead of Cu(II) complex, so that subsequent electroreduction of copper ions is energetically more favorable. Copper deposits obtained from Cu I -DES solution present a sand-rose cluster morphology inducing a large developed surface. Application of ultrasound during electrodeposition allows for grain size reduction and crystallographic texture modifications, compared to silent conditions.

Chalcopyrite heap leaching is not yet a commercial reality, not just because of its lower recovery or slower kinetics, but rather, because of the economic impacts lower recovery and slower kinetics impose to the operating cost of the process. Considering the importance of the financial analysis on the economics of the operation, in this research, the economics of hydrometallurgical treatment of low-grade chalcopyrite deposit at Sarcheshmeh mine as a function of ore grade and extraction recovery were investigated. First, a brief history and theory of chalcopyrite ore leaching has been reviewed, showing that bio heap leaching at a possibly higher temperature with forced aeration is the only viable technique in treatment of low-grade chalcopyrite ores. In the economic feasibility study section, an operating cost estimate, capital cost estimate and economic analysis for a plant with 7000 tons per year copper cathode production capacity as a function of ore grades, acid consumption and heap leaching recoveries were investigated. Preliminary capital cost and operating cost estimates were made within two scenarios for a plant utilizing bio heap leaching, solvent extraction and electrowinning of low-grade chalcopyrite ore. The operating cost, for instance, was shown to be between US$5.99/kg to US$4.29/kg of copper cathode for extraction recoveries of 30% to 50% in a scenario where the construction of new SX and EW plants are required (Scenario 1). The operating cost was shown to reduce to US$5.36/kg to US$3.66/kg, for a similar recovery range, if the current SX and EW facility can absorb the bio heap leaching solution (Scenario 2). Using the operating cost data and the current copper selling price, the ore cut-off grade as a function of copper extraction efficiency was identified. The economic analysis showed that for an ore containing 0.28% copper, 7000 tons per year of copper cathode production requires a total investment of 38 million US dollar. The internal rate of return for such a plant was shown to vary between −2% to 27% for copper extraction efficiencies between 30% to 50% (Scenario 1). The sensitivity analysis of the similar plant showed that a copper selling price of US$5/kg, for instance, can even make the 50% copper extraction efficiency economically unacceptable for scenario 1.

This study was conducted as part of the development of a novel process for copper recovery from copper sulfide concentrates by chloride leaching, simultaneous cuprous oxidation and cupric solvent extraction to transfer copper to a conventional sulfate electrowinning circuit and hematite precipitation to reject iron. Four LIX reagents (LIX84-I, LIX612N-LV, XI-04003 and LIX984N) were used as extractants while kerosene was used as a diluent. Copper was effectively extracted into LIX reagent organic solution from chloride solution. However, iron co-extraction exceeded 0.5 g/L. Two-stage countercurrent solvent extraction is proposed to reduce iron extraction. In the first stage, copper-loaded organic solution was pre-scrubbed by contacting with the cuprous and ferrous chloride pregnant leach solution. The extent of iron removal increased from 78 to 91 to 95% with decreasing pH from 1.0 to 0.5 to 0.0 while the extent of copper removal increased from 1 to 17 to 36%. Although more iron was removed at a pH below 0.5, copper was removed to an unacceptable level. Therefore, it is recommended that pH should be controlled around 1.0 to remove the bulk of the loaded iron while avoiding excessive copper stripping. In the second solvent extraction stage, copper is loaded into the organic solution by simultaneous cuprous oxidation by oxygen and cupric extraction. After two-stage counter current solvent extraction, the organic solution was further scrubbed by contacting with copper sulfate (4 g/L Cu) and sulfuric acid (15 g/L) solution. The iron extraction was reduced to a satisfactory level at an A/O ratio of 1/40 to 1/20 while the copper extraction practically did not change. The Cu/Fe ratio in the copper-loaded organic solution can meet the industrial requirement.

A novel separation process for detoxifying cadmium-containing residues arising from zinc hydrometallurgical processes has been developed. The solution of sulfuric acid and ammonium citrate ((NH4)3C6H5O7) is used as the lixiviant to partially extract copper and other metals from the residue at room temperature. Copper in the leachate is recovered by solvent extraction (SX) with LIX 973. Cobalt in the raffinate from the copper SX process is recovered by precipitation with α-nitroso-β-naphthol. Zinc remaining in the solution, at high concentrations, is further separated from cadmium by solvent extraction with P204 (di(2-ethylhexly)phosphoric acid, D2EHPA). The results show that around 50% of the copper and virtually all the zinc, cadmium and cobalt in the residue can be leached under the experimental conditions used here. Virtually all dissolved copper and cobalt can be recovered in the subsequent solvent extraction and precipitation processes. The presence of (NH4)3C6H5O7 could significantly facilitate the extraction of zinc by D2EHPA. More than 97% of the dissolved zinc can be recovered through a two-stage counter current solvent extraction process. After separating copper, cobalt and zinc, cadmium can easily be recovered from the solution either by cementation with zinc powder or by electrowinning, while the purified solution can be recycled back to the leaching process.

© 2014 Balaban Desalination Publications. All rights reserved. Abstract: A laboratory-/bench-scale investigation relating to copper recovery from Chilean copper flue dust is described. Since the dust contained a high amount of copper (nearly 24%), the considered option consisted in the leaching of the metallic material with ammonium chloride solution, thus enabling the dissolution of copper via its ammoniacal complexes, and also rendering, due to the almost neutral conditions of the leaching solution, a near complete lean solution with respect to the presence of other metal accompanying in the dust. Results obtained show that with ammonium chloride solutions, nearly 90% of the copper contained in the dust can be leached at 20°C. Copper enters the solvent extraction-electrowinning circuit due to the following equilibria: (Formula presented.) (Formula presented.) regenerating both the leaching and the solvent extraction reagents. In the above equation, HR represented the active substance of the extractant (Acorga PT5050) used in the solvent extraction circuit; the first of the above equilibria is related to the extraction operation and the second to the strip stage.

Extraction of copper using halide media is more rapid than in sulfate media, and the resulting cuprous ions require less energy to electrowin than cupric ions from sulfate media. The quality of the deposit is influenced strongly by the deposit roughness, which is controlled by mass transport. Thus, coupled halide leaching of chalcopyrite and subsequent electrowinning with controlled mass transport provide an opportunity for more sustainable copper production. This paper presents new insights regarding copper extraction and recovery methods from chalcopyrite ore in aqueous chloride media using leaching and electrowinning as well as steps for deposit quality improvements. Viability of integrated leaching and electrowinning was examined using low grade copper ore and continuous leaching solution recirculation. Other sets of experiments were conducted where leaching and electrowinning were performed separately, in order to find optimum experimental conditions for coupled leaching and electrowinning. Analysis based on scientific and statistical tools are presented for comparative evaluation. Copper electrodeposit morphology improvement by use of a variety of smoothing additives as well as mass transport control was also evaluated.

The extractant M5640 that extracted and stripped copper in the ammonia leach solution of copper oxidized ores solution of were mainly studied for the experimentation examined the M5640 concentration, phase ratio, extraction time, and liquid pH. Laboratory Investigations indicated that M5640 is really an efficient copper extractant with a maximum extraction of as high as 99.64% under the conditions of 6.48g/L Cu in leach solution, the extraction agent concentration is 10%, phase ratio is 1:1, extraction time is 2 min and liquid pH was 9.58. The stripping can be conducted with the barren electrolyte as a stripping agent in which the sulfuric acid concentration remains about 150g/L, and a stripping rate of 99.80% is obtained.

The disposal of large quantities of electronic scraps generated globally is causing not only environmental pollution but also loss of valuable resources. Therefore, R&D efforts have been made to develop a suitable solvent extraction (SX) process for extraction and separation of various metals viz. copper, zinc, cadmium and nickel from the solution expected from the leaching of e-scraps. Different process parameters viz. pH, aqueous to organic ratio, extraction isotherm to establish stage requirement for extraction in continuous mode, etc. have been studied to optimize the condition for the separation of metals. The studies showed the selective extraction of copper after iron precipitation from the sulfuric acid leach solution of e-wastes containing 0.92 g/L Cu and minor impurities with 2-hydroxy-5-nonylacetophenoneoxime (LIX 84) above pH 2.0 in single stage. Further, extraction of zinc, nickel and cadmium was carried out using Cyanex 302. It was found that complete extraction of cadmium and zinc was taken place simultaneously using 2.5% Cyanex at eq. pH 2.1 leaving nickel in the aqueous raffinate, moreover Zn and Cd was selectively stripped out using 10% H 2 SO 4 and 10% HCl, respectively. Based on the above studies a complete process flow sheet has been developed for separation/recovery to Cu, Zn, Cd and Ni from the leach liquor of e-waste.

This study was conducted as part of the development of a novel process for copper recovery from copper sulfide concentrates by chloride leaching, simultaneous cuprous oxidation and cupric solvent extraction to transfer copper to a conventional sulfate electrowinning circuit, and hematite precipitation to reject iron. The kinetics of cuprous and ferrous oxidation by oxygen and cupric solvent extraction were studied as a function of reaction time, stirring rate, oxygen pressure and extractants. The concentration of Cu(I) + Fe(II) decreased linearly with reaction time while the copper extraction increased with reaction time. The rate of cuprous and ferrous oxidation increased with increasing stirring rate and it was proportional to the partial pressure of oxygen. The cuprous and ferrous oxidation by oxygen was enhanced with the addition of organic solution. During the simultaneous cuprous oxidation and cupric solvent extraction, the aqueous solution pH first decreased sharply with time, then remained constant and finally increased, which was determined by the generation of H + ions (cupric and ferric solvent extraction) and consumption of H + ions (cuprous and ferrous oxidation by oxygen). The iron extraction increased as the oxidation reaction generated more ferric ions and raised the pH to a higher value. Simultaneous cuprous oxidation and cupric solvent extraction is feasible to transfer copper to the conventional sulfate electrowinning circuit from a chloride leaching system. The optimum copper solvent extraction conditions in chloride solution were proposed.

This study was conducted as part of the development of a novel process for copper recovery from chalcopyrite by chloride leaching, simultaneous cuprous oxidation and cupric solvent extraction to transfer copper to a conventional sulfate electrowinning circuit, and hematite precipitation to reject iron. Copper solvent extraction from chloride solution has been studied using four LIX® extractants (LIX84-I, LIX612N-LV, XI-04003 and LIX984N) from BASF with respect to copper extraction as a function of pH and A/O ratio, and behavior of the impurities. At a pH below 0.5, the copper extraction increased quickly with increasing pH while at a pH above 0.5, it increased only slightly with pH. The copper extraction in organic solution was virtually not affected by the impurities. The iron extraction in organic solution increased with decreasing A/O ratio from 2:1 to 1:8 as the copper extraction decreased. Conversely, the Cu/Fe ratio in organic solution increased as copper extraction increased. The extractions of silver and lead were 1 mg/L or lower under all conditions tested. The other impurities (Zn, Ni, Cd, Cr, Hg, As and Sb) were virtually not loaded into the organic solution. The optimum copper solvent extraction conditions in chloride solution were proposed.

This study was conducted to develop a novel process for copper recovery from chalcopyrite by chloride leaching, simultaneous cuprous oxidation and cupric solvent extraction to transfer copper to a conventional sulfate electrowinning circuit, and hematite precipitation to reject iron. Copper leaching from chalcopyrite concentrate in ferric and cupric chloride system was investigated using a two-stage countercurrent leach circuit under a nitrogen atmosphere at 97 °C to minimize the concentrations of cupric and ferric ions in pregnant leach solution for subsequent copper solvent extraction while maintaining a maximum copper extraction. A high calcium chloride concentration (110–165 g/L) was used to maintain a high cuprous solubility and enhance copper leaching. With 3–4 h of leaching time for each stage, the copper extraction reached 99% or higher while that of iron was around 90%. With decreasing concentrate particle size from p80 of 26 to 15 μm, the copper extraction increased by about 0.2% while the iron extraction increased by about 2.0%. The concentration of Cu(II) + Fe(III) in the pregnant leach solution was able to be reduced to 0.04 M. When the cupric concentration fell below the above limiting value, the elemental sulfur present was reduced by cuprous ions to form copper sulfide, eventually stopping the leaching of copper. Under this condition, only iron was leached. A very small amount of sulfur (1.2–1.4%) was oxidized to sulfate, resulting in an increase from 3 to 9 g/L in HCl concentration. The extractions of trace metals (Cr, Pb, Ni, Ag and Zn) were 96–100%.

Nickel and cobalt metal able to process from low grade laterite using solvent extraction and electrowinning. One of electrowinning methods which has good performance to produce pure metal is electrometal-electrowinninge(EMEW). In this work, solventextraction was used to separate nickel and cobalt which useCyanex-Versatic Acid in toluene as an organic phase. An aqueous phase of extraction was processed using EMEW in order to deposit the nickel metal in Cathode electrode. The parameters which used in this work were batch temperature, operation time, voltage, and boric acid concentration. Those parameters were studied and optimized using the design of experiment of Taguchi. The Taguchi analysis result shows that the optimum result of EMEW was at 60°C of batch temperature, 2 Voltage, 6 hours operation and 0.5 M of boric acid.

Electrowinning of zinc and lead from alkaline solutions depends on zinc or lead concentrations, temperature, current density, NaOH concentration, electrode distance and electrode material on current efficiency, and energy consumption of zinc or lead electrowinning. The optimum conditions for zinc electrowinning are zinc concentration 30–40 g/L, current density 800–1000 A/m2, NaOH concentration 180–200 g/L, and temperature 30–50 °C, and those for lead electrowinning are lead concentration 20 g/L, 5 g/L lead concentration left in the spent electrolyte, below 50 °C, 400 A/m2 current density, 5 mol/L NaOH concentration, and stainless steel of electrode material. Impurities ions have great effect on the quality of the metallic zinc or lead powder in alkaline leaching electrowinning process. The electrolysis deteriorates when the concentration of arsenide is over 500 mg/L, but \( {\mathrm{CO}}_3^{2-} \), \( {\mathrm{SO}}_4^{2-} \), \( {\mathrm{SiO}}_3^{2-} \), and F- have no effects on zinc electrowinning. Copper will be electrodeposited prior to lead in the cathode. Tin, \( {\mathrm{C}}_4{\mathrm{H}}_4{\mathrm{O}}_6^{2-} \), \( {\mathrm{CO}}_4^{2-} \), and gluten in the solution decrease the current efficiency of the lead electrowinning process, while As, Sb, and W, as well as \( {\mathrm{SO}}_4^{2-} \), \( {\mathrm{PO}}_4^{3-} \), \( {\mathrm{SO}}_3^{2-} \), \( {\mathrm{SiO}}_3^{2-} \), SCN−, \( {\mathrm{CO}}_3^{2-} \), Cl−, and AC−, have no effects.