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In pursuit of carbon neutrality, the demand for metals that are necessary for the development of clean energy technologies is rapidly increasing. Metallurgical waste, such as slag, presents a promising secondary source of these key metals. This research aims to develop an eco-friendly hydrometallurgical process to recover Cu, Ni, and Co from discarded copper/nickel slag. The high-pressure acid leaching (HPAL) was used to selectively leach Ni, Cu, and Co from the fayalite slag, yielding high leaching efficiencies of 99.9%, 89.4%, and 99.9%, respectively, with low Fe and Si tenors to the pregnant leach solution (PLS). The solvent extraction (SX) technique utilizing LIX 984N was used to selectively extract and enrich copper from the dilute PLS to about 23 g/L Cu with a very low Fe concentration of 0.05 g/L. Potassium amyl xanthate (PAX) solution was used to form Ni and Co xanthate complexes from the raffinate solution. Nickel was selectively recovered using ammonia solution, while cobalt xanthate complex was thermally decomposed and recovered as cobalt oxide solids of about 25 wt.% Co. A comprehensive process flowsheet is presented, furthermore, to realize the real application of the developed slag cleaning process, a preliminary economic evaluation was performed.

Porphyry copper deposits are important sources of copper and typically processed by flotation to produce copper concentrates. As mining areas become deeper, the amounts of impurities, such as sphalerite, can be increased in copper ores, so the appropriate depression of sphalerite floatability should be achieved to obtain saleable copper concentrates. In this study, the flotation behaviors of chalcopyrite and sphalerite in model samples mimicking copper ores with high Cu/Zn ratios (i.e., the ratio of chalcopyrite/sphalerite = 13:1) were investigated with zinc sulfate as a depressant for sphalerite. In addition, the effect of pyrite—a major gangue mineral in copper ores—on the depression of sphalerite floatability with zinc sulfate was examined. When sphalerite and chalcopyrite coexisted, the floatability of the former was effectively depressed by zinc sulfate (Zn recovery: 90%). Despite the presence of much higher amounts of chalcopyrite and pyrite than sphalerite, the application of nitrogen (N2) gas limiting the galvanic interaction between pyrite and chalcopyrite by reducing the dissolved oxygen (DO) concentration in the system effectively depressed the floatability of sphalerite (Zn recovery

The increase of impurities and complexity of copper ores are among the recent challenges in the mining industry. Complex carbonaceous sulfide ores are extremely difficult to treat due to their mineralogical complexity and impurities of organic carbon and carbonates. This study focuses on the development of a hydrometallurgical process for efficient copper extraction from complex carbonaceous sulfide ore which contains chalcopyrite, carbonates (dolomite and calcite), and carbonaceous gangue minerals. Characterization of the ore sample and leach residues was conducted using XRD and EPMA analysis, while ICP-OES was used for the determination of total dissolved metals in solution. High-pressure leaching of complex carbonaceous sulfide ore in oxygenated sulfuric acid solution was performed and the influence of leaching parameters such as sulfuric acid concentration, temperature, total pressure, and pulp density was studied. The extraction of copper increased with increasing temperature, sulfuric acid concentration, and total pressure. On the other hand, an increase in pulp density resulted in a decline in copper extraction due to an increased slurry viscosity and resistance in the diffusive mass transfer of reactants. Selective dissolution of copper from iron can be achieved by controlling free acidity in the pregnant leach solution (PLS). Under these leaching conditions: 100 g/L, 1 M H2SO4, 160 °C, 1.0 MPa total pressure, the highest copper and iron extractions achieved were 97.55% and 95.37%, respectively. Precipitation of copper from the PLS by NaSH sulfidization was investigated and more than 99.9% of copper was recovered at a Cu: NaSH molar ratio of 1:1.8.

In mining operations, an ore is separated into its constituents through mineral processing methods, such as flotation. Identifying the type of minerals contained in the ore in advance aids greatly in performing faster and more efficient mineral processing. The human eye can recognize visual information in three wavelength regions: red, green, and blue. With hyperspectral imaging, high resolution spectral data that contains information from the visible light wavelength region to the near infrared region can be obtained. Using deep learning, the features of the hyperspectral data can be extracted and learned, and the spectral pattern that is unique to each mineral can be identified and analyzed. In this paper, we propose an automatic mineral identification system that can identify mineral types before the mineral processing stage by combining hyperspectral imaging and deep learning. By using this technique, it is possible to quickly identify the types of minerals contained in rocks using a non-destructive method. As a result of experimentation, the identification accuracy of the minerals that underwent deep learning on the red, green, and blue (RGB) image of the mineral was approximately 30%, while the result of the hyperspectral data analysis using deep learning identified the mineral species with a high accuracy of over 90%.

Apatite is one of the important host minerals of the rare earth elements (REEs). Separation of REEs from apatite has not been studied extensively compared with other main minerals such as monazite, bastnasite and xenotime. Therefore, it is essential to develop an effective method for recovery of REEs from apatite-based ores. This paper presents the results of laboratory studies for the recovery of rare earth elements from an apatite ore sample by using sulphuric acid leaching, solvent extraction, and precipitation processes. Results showed that the most effective dissolutions of light REEs (>85%) and heavy REEs (>89%) from the apatite ore sample were achieved with a dilute sulphuric acid solution (1 M H2SO4) at a temperature of 20 °C in an hour. The heavy REEs present in the pregnant leach solution were almost entirely extracted using 1.2 M di-(2-ethylhexyl) phosphoric acid (D2EHPA) as a cation exchanger, while the most light REEs remained in the raffinate. The heavy REEs loaded into the organic phase were stripped completely by 3 M H2SO4 solution. REEs remained in raffinate and stripped into the aqueous solution were precipitated as RE oxalates by the addition of 0.08 M oxalic acid.

Sulfide copper mineral, typically Chalcopyrite (CuFeS2), is one of the most common minerals for producing metallic copper via the pyrometallurgical process. Generally, flotation tailings are produced as a byproduct of flotation and still consist of un‒recovered copper. In addition, it is expected that more tailings will be produced in the coming years due to the increased exploration of low‒grade copper ores. Therefore, this research aims to develop a copper recovery process from flotation tailings using high‒pressure leaching (HPL) followed by solvent extraction. Over 94.4% copper was dissolved from the sample (CuFeS2 as main copper mineral) by HPL in a H2O media in the presence of pyrite, whereas the iron was co‒dissolved with copper according to an equation given as CCu = 38.40 × CFe. To avoid co‒dissolved iron giving a negative effect on the subsequent process of electrowinning, solvent extraction was conducted on the pregnant leach solution for improving copper concentration. The result showed that 91.3% copper was recovered in a stripped solution and 98.6% iron was removed under the optimal extraction conditions. As a result, 86.2% of copper was recovered from the concentrate of flotation tailings by a proposed HPL‒solvent extraction process.

Different separation techniques such as solvent extraction, ion exchange, and precipitation are often used for recovery of rare earth elements (REEs) from pregnant leach solutions obtained from acid leaching. Solvent extraction is generally accepted as the most appropriate commercial technology for separating REEs due to the need to be able to handle larger volumes of diluted pregnant solutions. This study focused on the development of selective separation of light and heavy REEs from the pregnant leach solution obtained from leaching of apatite ore in 1 M sulfuric acid (H2SO4) using solvent extraction. Three different commercial organophosphorus extractants (di-(2-ethylhexyl) phosphoric acid (D2EHPA), 2-ethylhexyl phosphonic acid mono-2-ethylhexyl ester (PC88A) and tributyl phosphate (TBP)), and the influences of experimental parameters such as extractant concentration, organic/aqueous phase ratio, diluent type, pH, extraction time and stripping agent concentration were examined. Results showed that light REEs (LREEs) and heavy REEs (HREEs) in the pregnant leach solution were selectively separated with D2EHPA via a two-stage extraction process. In the first-stage of solvent extraction, >90% of (0.05 g/L) HREEs was extracted with 1.8 M D2EHPA in kerosene while the vast majority (>95%) of LREEs was remained in raffinate. In the second-stage, >93% (1.01 g/L) of LREEs was extracted from the raffinate with 1.8 M D2EHPA dissolved in kerosene at pH 1.6. HREEs (>95%) and LREEs (>90%) loaded with D2EHPA after the first and second-stage of extraction were stripped by 4 M H2SO4 and 1 M H2SO4 solutions, respectively. Distribution of middle rare earth elements (MREEs) was discussed through the extraction processes in this study.

This study investigates the removal of silica and alumina as impurities from hematite based low-grade iron ore containing 34.18 mass% iron, 31.10 mass% of silica and 7.65 mass% alumina. Wet high-intensity magnetic separation (WHIMS) and reverse flotation (RF) were investigated. In WHIMS process, 93.08% of iron was recovered with a grade of 53.22 mass% at an optimum magnetic density of 10,000 mT, and pulp density of 2% used the L-4 machine. In RF experiments, optimal results showed 95.95% of iron recovered with 51.64 mass% grade using 1 kg/t of 1% alkaline starch as iron depressant and 1:1 mixture ratio of 0.75 kg/t DAA and NaOL as silica and alumina collectors. The designed multi-stage process involving feeding the concentrate from WHIMS into RF process reduced silica to 2.02 mass%, alumina to 1.04 mass% whilst recovering 81.94% of the iron with 67.27 mass% grade. As a result of this research, a process to produce high quality iron concentrate from hematite based low-grade iron ore with high iron recovery rate was constructed.

This paper considers the issue of recycling of waste printed circuit boards (WPCBs) containing precious and base metals in appreciable amounts. High-pressure oxidative leaching (HPOL) with dilute sulfuric acid resulted in removal of a significant amount of base metals from a WPCB ash sample obtained by incineration at 800°C. The parameters investigated in the precious metal leaching from WPCB residue after HPOL included the sulfuric acid concentration, thiourea concentration, oxidant concentration, leaching temperature, and leaching time. Recovery of gold, silver, and palladium of 100%, 81%, and 13% from the WPCB residue sample was achieved by thiourea leaching under optimized conditions. The results show that the efficiency of precious metal dissolution from the WPCB sample using thiourea solution depended strongly on the concentration of both thiourea and oxidant.

The leaching of copper from chalcopyrite in H2SO4 solution under pressure-oxidative conditions and its kinetics were investigated in this study. Leaching variables that affect the rate of copper dissolution from chalcopyrite are agitation speed (300–900 rpm), total pressure (0.8–2.0 MPa), temperature (160–180 °C), and sulfuric acid concentration (0.1–2.0 M). Results showed that dissolution of chalcopyrite increases with increasing agitation speed, total pressure, and temperature, whereas it decreases with the increasing sulfuric acid concentration. Under the optimal conditions, copper extraction of 94.5% was achieved after 90-min leaching, while a dissolution of iron at 4.2% was obtained. The kinetic study showed that the dissolution of chalcopyrite is represented by a shrinking core model with chemical reaction controlling mechanism given as (1 − (1 − α)1/3). The activation energy (E a) for the leaching reaction was calculated to be 42.4 kJ/mol. The reaction order with respect to total pressure was about 8.0, which indicates that total pressure, i.e., oxygen partial pressure, in an autoclave is the most important factor in controlling the dissolution of chalcopyrite in H2SO4 solution under pressure-oxidative leaching conditions. The effect of Fe/Cu mole ratio (1–20 mol/mol, adjusted by addition of pyrite) on chalcopyrite leaching from a copper ore was investigated. The results show that the sulfuric acid produced during pyrite oxidation promotes the chalcopyrite dissolution.

The present study focused on the re-processing of copper and nickel from mine tailings. In this work, recovery of copper and nickel from mine tailing by combined process of flotation and high pressure oxidative leaching were considered. In the first stage, effects of flotation parameters including collector type, collector dosage, and pH and pulp density were examined. The results showed that over 80% copper recovery was achieved under the optimized flotation conditions while nickel recovery was lower than 30% due to its co-ex-istence with gangue minerals of pyrrhotite, pyrite and other clay minerals. In the second stage, key parameters, particularly concentration of sulfuric acid, temperature, pressure and leaching time were investigated to test the leaching efficiency of copper and nickel from the flotation concentrate with high pressure oxidative leaching (HPOL). A comparison was made between the leaching efficiencies of copper and nickel from flotation concentrates and mine tailing.

Iodine-iodide leaching and activated carbon adsorption processes for recovery of gold (Au) from waste printed circuit boards (WPCBs) were discussed in this paper. A pressure oxidative acid leaching (POAC) was carried out in an autoclave using diluted sulfuric acid (1M H 2 SO 4 ) solution in order to remove high concentrations of some base metals especially copper (Cu), aluminum (Al), iron (Fe) and zinc (Zn) due to their negative effect on gold dissolution from WPCBs. The factors affecting the performance and efficiency of the iodine-iodide leaching process; such as iodine/iodide concentration, pulp density, leaching time and leaching temperature were optimized in order to maximize the gold dissolution efficiency from the WPCBs in the iodine-iodide solution. Results indicated that the vast majority (> 99%) of gold was dissolved in the solution from the WPCBs under the optimized leaching conditions. Adsorption tests were conducted on leach liquor solutions resulting from the iodine-iodide leaching using activated carbon. Nearly 98% of gold was adsorbed from the liquor solution onto the carbon under the condition optimized in this study. The results obtained revealed that gold can be successfully recovered from this secondary resources, where the percent recovery amounts to nearly 97% for gold.

The surface properties of slightly oxidized chalcocite, bornite, and chalcopyrite with sodium hydrosulfide (NaHS) treatment were investigated by solid analyses of X-ray photoelectron spectroscopy, solution analyses, and zeta potential measurements. NaHS is known as a flotation activator for copper-oxide minerals by sulfurizing their surfaces. This study focused on the surface properties before/after sulfurization by the NaHS at several NaHS dosages and alkaline pHs. Iron leaching and precipitates of copper–iron-sulfide minerals have also been investigated. The analysis results showed that NaHS sulfurized the oxidized copper species, CuO and CuSO4, to copper sulfide, but oxidized iron species, Fe2O3 and Fe(OH)3, were not sulfurized by NaHS. These oxidized iron species dissolved into solutions as complex ion, Fe(HS)20, then Fe(HS)20 was decomposed by dissolved oxygen and reprecipitated as Fe2O3. The results of chalcocite showed that Cu2S was generated on the surfaces with NaHS treatment at any pH. The bornite and chalcopyrite results indicated that Fe2O3 was produced on bornite surfaces with an increase in pH, whereas the chalcopyrite surfaces were covered as the pH decreased. A flotation test was conducted using a Chilean copper ore that contained chalcocite, bornite, and chalcopyrite, and the flotation recovery of each mineral was consistent with the surface sulfurization at several NaHS dosages and pHs.

This paper deals with the recovery of gold from waste printed circuit boards (WPCBs) ash by high-pressure oxidative leaching (HPOL) pre-treatment and iodide leaching followed by reduction precipitation. Base metals present in WPCB ash were removed via HPOL using a diluted sulfuric acid solution at elevated temperatures. Effects of potassium iodide concentration, hydrogen peroxide concentration, sulfuric acid concentration, leaching temperature, and leaching time on gold extraction from pure gold chips with KI&ndash, H2O2&ndash, H2SO4 were investigated. The applicability of the optimized iodide leaching process for the extraction of gold from the leach residue obtained after HPOL were examined at different pulp densities ranging from 50 g/t to 200 g/t. Results show that the removal efficiency was 99% for Cu, 95.7% for Zn, 91% for Ni, 87.3% for Al, 82% for Co, and 70% for Fe under defined conditions. Under the optimal conditions, the percentage of gold extraction from the gold chips and the residue of WPCBs was 99% and 95%, respectively. About 99% of the gold was selectively precipitated from the pregnant leach solution by sequential precipitation with sodium hydroxide and L-ascorbic acid. Finally, more than 93% of gold recovery was achieved from WPCB ash by overall combined processes.

Although Rh is an industrially important and the most expensive platinum group metal (PGM), the selective and preferential separation of Rh from PGM mixtures still remains as a big challenge. In this work, the separation of Rh (III) from Pd (II) and Pt (IV) in a hydrochloric acid (HCl) solution was studied using a m-phenylene diamine-containing precipitant (m-PDA). At high HCl concentrations (6.0–8.0 M), most of the Rh (III) (>90%) was precipitated, and Pd (II) and Pt (IV) were hardly precipitated (m-PDA enabled selective and preferential precipitation of Rh (III) at high HCl concentrations. XPS and TG analyses revealed that the Rh-containing precipitate is an ion-pair complex composed of one [RhCl6]3− anion and three m-PDA cations. The Rh desorption from the precipitate as well as the recovery of m-PDA was successfully achieved using an NH4OH solution. This method is a promising practical approach to Rh recovery.

An investigation on improving copper and iron recovery by removal of organic carbon and carbonaceous gangue mineral through flotation was carried out. The ore contains chalcopyrite and bornite as valuable minerals, and dolomite and calcite as carbonaceous gangue minerals. The chemical composition of the ore indicates that grade of 2.08 and 5.37 mass% for copper and iron respectively, while carbon was 3.70 mass%. The total organic carbon in the feed was recorded as 2.02 mass%. The presence of these gangue minerals is detrimental to copper and iron recovery during flotation. Therefore a two stage flotation study was carried out to remove hydrophobic carbon and carbonaceous material from ore and to improve copper and iron recovery. The objective of the study was to find effective and selective flotation conditions which can eliminate gangue minerals from valuable minerals and improve quality (grade) and recovery of copper and iron. The obtained results could be used to develop a flotation circuit for recovering copper and iron from carbonaceous sulphide ores.

This study presents a viable approach for recovery of precious metals such as gold (Au), silver (Ag), palladium (Pd), and base metals, including copper (Cu), nickel (Ni), cobalt (Co), lead (Pb) and zinc (Zn) from waste printed circuit boards (WPCBs) via iodine-iodide leaching and precipitation. The behaviours of dissolution and precipitation of precious and base metals during iodine-iodide leaching and precipitation processes were discussed. Sodium hydroxide (NaOH) was used to remove base metal impurities exist in the pregnant leach solution under alkaline conditions. Precious metals remained in the resulting solution from NaOH precipitation were recovered by reduction using ascorbic acid (L-AA) solution. Results show that under optimum leaching conditions, almost all (> 99%) of Au was dissolved in an iodine-iodide solution when the dissolution efficiencies of other precious metals (Ag, Pd) and base metals, besides calcium (leaching of 25%) were less than 1 and 6%, respectively. The study revealed that more than 95% of Cu, Ni, Pb, Zn, Fe and Mn were initially removed from the pregnant leach solution at pH of 9.3 with addition of 0.1 M NaOH. Then 99.8% Au, 81.7% Ag and 74% Pd were precipitated from the obtained solution after NaOH precipitation while L-AA dose was 0.6 ml/ml at the condition. It can be concluded that the precious and base metals could be recovered selectively and economically from WPCBs via iodine-iodide leaching followed by precipitation using NaOH and L-AA.

The aim of this research was to develop a copper recovery process from mine tailings (0.34%Cu) using flotation followed by high-pressure oxidative leaching (HPOL) and solvent extraction. The results of HPOL using the concentrate of mine tailings obtained by flotation under the optimal conditions of the previous study shown that an efficient copper dissolution of 94.4% was achieved in an H2O media, while the copper concentration of PLS reached to be 2.9 g/L. The solvent extraction of PLS obtained from the optimal HPOL showed that 91.3% copper was recovered in stripped solution under the determined optimum conditions, in which the copper concentration achieved to be 44.8 g/L. Finally, a proposed copper recovery process from the concentrate of mine tailings was developed by combination of HPOL and solvent extraction, while a total copper recovery of 86% was achieved.

A process to co-treat electronic and pyrite (FeS2) wastes in an autoclave (high pressure and temperature) under oxygen gas atmosphere for extraction of metals from electronic waste was investigated in this study. The electronic waste sample contained high Cu (13–20%), Au (230–360 g/t), Ag (480–1020 g/t), including Pt, Pd and Ni among many other metals. Pyrite samples obtained as mine tailings for the study contained 69.1% FeS2 (32.5% Fe and 36.6% S) and was used as sulfur source for the generation of H2SO4 during autoclave oxidation to subsequently extract Cu and other base metals in the electronic waste. Reaction temperature and time, autoclave pressure, pulp density and H2SO4 concentration were investigated to evaluate metals extraction. At temperatures ranging from 120 to 180 °C and autoclave pressure of 2 MPa, pyrite was completely oxidized. Subsequently, Cu extraction reached over 99% and very high extraction of Zn, Ni and Cr metals was achieved within 1 h. The solid products from autoclave oxidation containing precious metals (Au, Ag, Pt, Pd) can be treated separately for their recovery.

SCS pincer ligands 1–4 were synthesised, and their ability to extract Pd(II) from HCl and HNO3 media was studied. The Pd(II) extraction properties of 1–4 were compared with those of commercial extractants (DOS and LIX®84-I) in kerosene. 1 and 2 showed superior Pd(II) extractability (E% = 99.9) relative to DOS and LIX®84-I from 0.1–8.0 M HCl and to DOS from 0.1–8.0 M HNO3 and mixed HCl + HNO3 media. The Pd(II) extraction rate, acid durability, the most suitable organic/aqueous (O/A) phase ratio, and Pd(II) loading capacity of extractants 1, 2, and DOS were evaluated. 1 and 2 exhibited a greater Pd(II) extraction rate and Pd(II) loading capacity than DOS. 1 was very stable in acid media (HCl and HCl + HNO3), whereas 2 and DOS deteriorated in HCl + HNO3. Selective extraction of Pd(II) by 1 and 2 was achieved from a mixed solution containing Pd, Pt, Rh, rare metals, and base metal ions that simulated the leach liquors of automotive catalysts. The back extraction of Pd(II) and reusability of extractants 1 and 2 were studied. The Pd(II) extraction mechanism of 1–4 was investigated using FT-IR, UV-visible, and NMR spectroscopy.

This paper describes a hydrometallurgical process for recovering neodymium (Nd) and dysprosium (Dy) from a magnetic waste sludge generated from the Nd–Fe–B(–Dy) manufacturing process. Phase analysis by XRD study revealed Nd(OH)3 and Fe2O3 as main mineral phases, and chemical analysis by ICP showed the contents of 35.1 wt% Nd, 29.5 wt% Fe, 1.1 wt% Dy and 0.5 wt% B. A solution of 1 M HNO3 + 0.3 M H2O2 was used to dissolve up to 98 % Nd and 81 % Dy, while keeping Fe dissolution below 15 % within 10 min. Fe dissolved in solution was completely removed as Fe(OH)3 at pH 3 followed by precipitation of Nd and Dy with oxalic acid (H2C2O4) and recovered 91.5 % of Nd and 81.8 % of Dy from solution. The precipitate containing Nd and Dy was calcined at 800 °C to obtain Nd2O3 as final product with 68 % purity, and final recovery of 69.7 % Nd and 51 % of Dy was reported in this process.

A flotation pre-treatment study for the separation of enargite (Cu3AsS4) from chalcopyrite (CuFeS2) ores of different origins was investigated in this work. The copper ore bearing enargite mineral contained 5.87mass% As and 16.50mass% Cu while the chalcopyrite bearing ore contained 0.32mass% As and 21.63mass% Cu. The two ore samples were mixed at 7 : 3 (enargite : chalcopyrite) by weight ratio to prepare a mixed ore sample with As content at 3.16 and 18.25mass% Cu for the flotation study. Effect of particle size, slurry pH, flotation time, collector type, collector addition or dosage and depressants were investigated to evaluate efficiency of enargite separation from chalcopyrite and recovery of both minerals as separate concentrates. For enargite single ore flotation, the 38­75µm size fraction showed that over 98% of enargite was selectively recovered within 5min at slurry pH of 4 and As content in the final tailings was reduced to 0.22mass%. In mix ore (enargite + chalcopyrite) flotation, 97% of enargite was first removed at pH 4 followed by chalcopyrite flotation at pH 8, and over 95% recovery was achieved in 15min flotation time. The As content in the final tailings was reduced to 0.1mass%. [doi:10.2320/matertrans.M2011354]

Application of flotation for upgrading and recovery of copper (Cu) and molybdenum (Mo) contained in flotation tailings from a copper/molybdenum mine was investigated in this study. The sample contained mainly pyrite (FeS2) and quartz (SiO2) as dominant phases with Cu and Mo contents at 0.38% and 0.23% respectively. A flotation system employing use of collector/frother mixture [kerosene + sodium dodecyl sulphate (SDS) + polyethylene alkyl ethyl] with tannic acid and sodium silicate as FeS2 and SiO2 depressants, compared against potassium amyl xanthate (PAX) and methyl isobutyl carbinol (MIBC) additions were investigated to evaluate Cu and Mo recovery. According to the results, the grades of Cu and Mo in the concentrate did not increase to above 1% when PAX and MIBC were used as collector and frother, even at increased PAX addition. Also grades for tests performed at varying pH rangig from pH 3 to 11 showed no significant improvement. However, with the use of the collector/frother mixture and tannic acid, Cu and Mo grades increased to 1.4 and 2.4% respectively. The grade of Mo was further increased to 8.0% at 52% recovery when sodium silicate was used with the collector/frother mixture. These results indicate that Cu and Mo lost in such tailings can be upgraded and offered as additional resource for possible recovery.

Generally, trace precious metals remaining in wastewaters generated from the refining process of precious metals are not recovered, due to a relatively high processing cost as well as various technical problems. Recovery of precious metals from wastewaters is very important for the conservation of resources and the protection of environment. However, wastewaters containing a large amount of ammonium ion (NH4 þ ) cannot be treated by general neutralization operation, due to formation of metal ammine complexes with increasing pH. In this study, the possibility of recovering precious metals and other valuable metals from wastewaters by various traditional metallurgical processes such as cementation, neutralization and reduction, were investigated. A recovery of 99% Copper (Cu), 96% Palladium (Pd), and 85% Gold (Au) by cementation using Iron (Fe) powder, and 99.6% Cu, 99.5% Pd by cementation using Aluminum (Al) powder was achieved. However, complete recovery of all valuable metals by a one-step cementation process was not possible. On the other hand, precious metals and other valuable metals including Copper and Indium, etc., were precipitated by combining neutralization, deammoniation and reduction processes. Results showed that the recovery of Platinum (Pt) in the reduction process was improved by adding deammoniation step. Finally, precious metals are concentrated in the crude copper metal by fusion process. The recovery of Au, Ag, Pd was more than 91%, and that of Pt was about 71%. [doi:10.2320/matertrans.M2010432]