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Carrier flotation is a technique that can recover fine particles by using coarse carrier particles during the flotation process. In heterogeneous carrier flotation, coarse mineral particles of different minerals are used as carriers to recover fine mineral particles. By using Cu2+-treated pyrite particles as carriers, fine chalcopyrite particles recovery could be improved. However, a disadvantage of this heterogeneous carrier flotation is that it requires a post-flotation separation process to improve the grade of the final Cu concentrate. This study tested mechanical and chemical treatments to detach finely ground chalcopyrite (D50~3.5 µm) particles from Cu2+-treated coarse pyrite particles (−125 + 106 µm) after flotation. The results showed that the ultrasonic treatment was not effective to detach chalcopyrite fines from Cu2+-treated pyrite particles. However, acid treatment was effective to detach chalcopyrite fines from coarse pyrite particles. At pH 2, approximately 96% of chalcopyrite fines were detached from coarse Cu2+-treated pyrite particles. The acid treatment of flotation froth (mixture of chalcopyrite fines and Cu2+-treated pyrite particles) decomposed the collector KAX (potassium amyl xanthate) and dissolved the Cu precipitates adsorbed on the pyrite surface. This weakened the hydrophobic attraction force between the chalcopyrite fines and coarse pyrite particles, thus promoting the detachment of chalcopyrite fines from Cu2+-treated coarse pyrite particles.

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. 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 the 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.

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.

Historical tailings storage facilities (TSFs) are either abandoned or sparsely rehabilitated promoting acid mine drainage (AMD) formation and heavy metal release. To sustainably manage these sites, a geochemical audit coupled with numerical simulation to predict AMD flow paths and heavy metal migration are valuable. In this study, a 40-year-old TSF in Hokkaido, Japan was investigated. Tailings in this historical TSF contain pyrite (FeS

Lead (Pb) contaminated soils pose serious risks to human health and the environment. One remediation technique that permanently removes Pb is chemical soil washing. This study investigated a modified chemical soil washing technique whereby Pb extraction in citrate solution and recovery of dissolved Pb by reductive precipitation (cementation) were simultaneously carried out in a single reactor by the addition of micro-sized zero-valent iron (mZVI) to decontaminate Pb-contaminated soil from a shooting range. Pb-loaded mZVI were then recovered by magnetic separation. The effects of mZVI dosage, citrate concentration, pH, and hydrodynamics on Pb removal were evaluated. Under acidic pH, Pb removal increased with increasing citrate concentration. The addition of mZVI suppressed Pb removal from 59.6 to 30% in 0.1 M citrate solution, which was attributed to dissolved Fe ions competing for free citrate ions to form soluble Fe-citrate complexes. Increasing the citrate concentration and shaking speed enhanced Pb extraction from the soil and subsequent recovery via cementation onto mZVI. Pb removal and recovery onto mZVI of up to 99.0% were achieved at citrate concentration of 0.7 M, pH of 5, and shaking speed of 165 strokes/min. The addition of mZVI during chemical soil washing in citrate solution not only extracted and recovered Pb from the contaminated soil but also eliminated the need to wash soil residues obtained after treatment because the bulk of dissolved Pb was sequestered from the solution.

The use of galvanic interactions between zero-valent aluminum (ZVAl) and activated carbon (AC) to recover gold (Au) ions is a promising technique to overcome the challenges due to the poor recovery in ammonium thiosulfate systems, but the applicability to practical Au ore processing remains elusive so far. The present study describes (1) the recovery of Au ions from low Au concentrations, which are typical concentrations used in Au ore processing, and (2) an investigation into the effects of various coexisting base metal ions that can be present in pregnant ore-leached solutions. The results showed that high Au recovery (i.e., over 85%) was obtained even at low Au concentrations under the following conditions: 1:1 of 0.15 g of ZVAl and AC with 10 mL of ammonium thiosulfate solution containing 5 mg/L of Au ions at 25 °C for 1 h in an anoxic atmosphere. Selected coexisting metal ions (i.e., copper, iron, cobalt, nickel, and zinc) were studied to establish their effects on Au recovery, and the results showed that the Au recovery was enhanced (about 90%) when copper ions coexist in the solution with minimal effects from other competing base metal ions.

As technology advances, consumers are looking to buy new phones. This phenomenon leads to decrease in life-cycle of mobile phones (ca. 2–3 years), resulting in increase of waste mobile phones. In waste mobile phones, it contains various components including display unit, battery, cases, and printed circuit boards (PCBs). PCBs alone contain variety of materials such as gold (Au), copper (Cu), and iron (Fe) as well as aluminum (Al) with plastics, which causes difficulty in recycling. The coexistence of Cu and Al in crushed mobile phones has been recently reported to cause very low Au extraction efficiency in ammonium thiosulfate medium due primarily to enhanced re-deposition of extracted Au ions via galvanically induced cementation. To limit this antagonistic effect of Cu and Al coexistence on Au extraction, Au-containing components were separated from Cu- and/or Al-bearing parts using a two-step crushing approach in tandem with various physical separation techniques (i.e. jig, hybrid jig, magnetic separation, flotation, and sink-float separation) that were applied on appropriate size fractions. The results showed that over 96% of PCBs, the primary Au-containing material in mobile phones, were recovered after jig separation. These concentrated PCBs were collected and further crushed to liberate Au-containing parts and then treated by diverse physical separation techniques such as magnetic separation, jig separation, hybrid jig separation, flotation, and sink-float separation depending on the crushed particle sizes. Ammonium thiosulfate leaching of the mixed Au-concentrated products from various physical pretreatment techniques showed that Au extraction improved by about 15-fold, indicating that the physical separation scheme proposed in this study is very promising.

Acid mine drainage (AMD) with toxic arsenic (As) is commonly generated from the tailings storage facilities (TSFs) of sulfide mines due to the presence of As-bearing sulfide minerals (e.g., arsenopyrite, realgar, orpiment, etc.). To suppress As contamination to the nearby environments, As immobilization by Ca-Fe-AsO

Acid mine drainage (AMD) formation is mainly caused by the oxidation of pyrite. Carrier-microencapsulation (CME) using metal-catecholate complexes has been proposed to passivate sulfide minerals by forming surfaceprotective coatings on their surfaces. Among the various metal-catecholate complexes, Ti-catecholate formed stable coatings having superior acid-resistance, but a thick enough passivating film required considerable time (ca. 14 days) to grow. Meanwhile, Fe-catecholates can form Fe-oxyhydroxide coatings within 2 days, however, they are less stable than Ti-based coating. To address these drawbacks of using a single metal-complex, this study investigated the concurrent use of Fe-catechol and Ti-catechol complexes for accelerating the formation of stable passivating coating on pyrite & nbsp;Compared with a single metal-complex system, the coating formation was significantly accelerated in mixed system. Linear sweep voltammetry showed the simultaneous decomposition of [Fe(cat)]+ and [Ti(cat)3]2 & ndash; as the main reason for improved coating formation. Electrochemical properties of coatings formed by single and mixed complex systems, confirmed by electrochemical impedance spectroscopy and cyclic voltammetry, indicated the coating formed in the mixed system had higher resistance and more electrochemically inert than the other cases. The simultaneous use of Fe-catechol and Ti-catechol complexes enhanced pyrite passivation by accelerating metal-complex decomposition and forming more stable coating composed of Fe2TiO5.

Porphyry copper deposits (PCDs) are some of the most important sources of copper (Cu) and molybdenum (Mo). Typically, the separation and recovery of chalcopyrite (CuFeS2) and molybdenite (MoS2), the major Cu and Mo minerals, respectively, in PCDs are achieved by two-step flotation involving (1) bulk flotation to separate Cu–Mo concentrates and tailings (e.g., pyrite, silicate, and aluminosilicate minerals) and (2) Cu–Mo flotation to separate chalcopyrite and molybdenite. In Cu–Mo flotation, chalcopyrite is depressed using Cu depressants, such as NaHS, Na2S, Nokes reagent (P2S5 + NaOH), and NaCN, meaning that it is recovered as tailings, while molybdenite is floated and recovered as froth product. Although conventionally used depressants are effective in the separation of Cu and Mo, they have the potential to emit toxic and deadly gases such as H2S and HCN when operating conditions are not properly controlled. To address these problems caused by the use of conventional depressants, many studies aimed to develop alternative methods of depressing either chalcopyrite or molybdenite. In this review, recent advances in chalcopyrite and molybdenite depressions for Cu–Mo flotation separation are reviewed, including alternative organic and inorganic depressants for Cu or Mo, as well as oxidation-treatment technologies, such as ozone (O3), plasma, hydrogen peroxide (H2O2), and electrolysis, which create hydrophilic coatings on the mineral surface.

Ammonium thiosulfate solution is an ideal lixiviant to extract gold (Au) from electronic wastes (E-wastes) because it is non-toxic, less corrosive, and more selective than conventional cyanide or halide solutions. It was reported recently, however, that Au leaching efficiency in ammonium thiosulfate medium dramatically decreased at high solid-to-liquid ratios (S/L), even though the amounts of reagents used were in excess. To understand how this occurred, leaching experiments were conducted using printed circuit boards (PCBs) from waste mobile phones, and Au distribution in the leaching residues was examined by scanning electron microscopy with energy dispersive X-ray spectroscopy (SEM-EDX). Significant amounts of Au were found together with copper (Cu) and aluminum (Al), implying that extracted Au ions were likely re-deposited during leaching onto Cu and Al found in PCBs via cementation (i.e., reductive deposition). A more detailed elucidation of this phenomenon by cementation experiments using pure Cu and/or Al powders indicates that these metals could only recover Au ions alone via cementation at very high amounts, however, this process became more extensive when Cu and Al powders were suspended together in solution even though the amounts of the individual metals were much lower. Electrochemical experiments (chronoamperometry) in ammonium thiosulfate solutions containing Au ions using an Al working electrode also showed that Au ion cementation was dramatically enhanced when Cu powder was present in solution, and the bulk of Au was cemented on Cu powder rather than on the Al electrode. These results suggest that coexistence of Cu and Al interfered with the extraction of Au in ammonium thiosulfate medium at high S/L because of the enhanced re-deposition of extracted Au via galvanic interaction.

The zinc leach residues (ZLRs) generated by Zn hydrometallurgical process are disposed of in landfills due to a lack of proper treatment methods although they contain toxic heavy metals like Cd and Zn. To remove these heavy metals, this study investigated cementation using aluminum (Al) together with magnetite (Fe3O4) from sulfate solutions. Cementation using only Al could not remove Cd2+ and Zn2+ because the insulating Al oxide (Al2O3) layer on Al surface was not removed at high pH (5.0) from sulfate solution, and inhibiting electron transfer from Al to metal ions. The addition of Fe3O4, however, enhanced the cementation of Cd2+ and Zn2+ from 0% to 83% and 92%, respectively. The results of SEM-EDS analysis for the cementation products confirmed that Cd and Zn were deposited on Fe3O4 attached on Al surface. Electrochemical impedance spectra for Al electrode in sulfate solutions confirmed that charge transfer resistance decreased when Fe3O4 powder was attached on the electrode surface, suggesting that Fe3O4 acts as a mediator to enhance electron transfer from Al to metal ions in solutions. By applying magnetic separation for the cementation products, the majority of cemented Cd and Zn were recovered together with Fe3O4 in magnetic fraction, and they were separated from Al in non-magnetic fraction.

Bimetallic materials—well-known for their strong reducing properties—are promising tools to recover battery/critical elements from waste streams for resources conservation and decontamination. To date, however, their synthesis has been limited to the use of pure monometallic elements. In this study, we synthesized magnetic bimetallic materials from aluminum wastes (Al-scrap) using a two-stage mechanical-chemical method composed of (1) polishing to remove protective coatings, and (2) etching-cementation for zero-valent iron (ZVI) deposition. Argon-etching using X-ray photoelectron spectroscopy (XPS) showed the effective removal of anodized/polymeric coatings on Al-scrap but the exposed Al0 was rapidly passivated by an Al-oxyhydroxide/oxide film composed of corundum, bayerite, and boehmite. In the etching-cementation stage, the Al-oxyhydroxide/oxide film was removed using NaCl–HCl and Fe3+ was deposited in corrosion pits on Al-scrap. Strong “magnetic susceptibility” was observed in 1.0 M Fe3+/3.5 M Cl− synthesized products but not in 0.5 M Fe3+/2 M Cl−. This difference could be attributed to the prevalence of ZVI formation in the former but akaganeite precipitation in the latter. The results also showed “targeted” recovery of copper dissolved from Al-scrap on ZVI regions of synthesized product. Finally, two pathways are proposed for Fe3+ reduction: (i) direct cementation (Fe3+→ZVI), and (ii) sequential reduction (Fe3+→Fe2+→ZVI).

Acid mine drainage (AMD) with toxic arsenic (As) is commonly generated from the tailings storage facilities (TSFs) of sulfide mines due to the presence of As-bearing sulfide minerals (e.g., arsenopyrite, realgar, orpiment, etc.). To suppress As contamination to the nearby environments, As immobilization by Ca–Fe–AsO4 compounds is considered one of the most promising techniques; however, this technique is only applicable when As concentration is high enough (>1 g/L). To immobilize As from wastewater with low As concentration (~10 mg/L), this study investigated a two-step process consisting of concentration of dilute As solution by sorption/desorption using schwertmannite (Fe8O8(OH)8–2x(SO4)x; where (1 ≤ x ≤ 1.75)) and formation of Ca–Fe–AsO4 compounds. Arsenic sorption tests indicated that As(V) was well adsorbed onto schwertmannite at pH 3 (Qmax = 116.3 mg/g), but its sorption was limited at pH 13 (Qmax = 16.1 mg/g). A dilute As solution (~11.2 mg/L As) could be concentrated by sorption with large volume of dilute As solution at pH 3 followed by desorption with small volume of eluent of which pH was 13. The formation of Ca–Fe–AsO4 compounds from As concentrate solution (2 g/L As(V)) was strongly affected by temperature and pH. At low temperature (25–50 °C), amorphous ferric arsenate was formed, while at high temperature (95 °C), yukonite (Ca2Fe3–5(AsO4)3(OH)4–10·xH2O; where x = 2–11) and johnbaumite (Ca5(AsO4)3OH) were formed at pH 8 and 12, respectively. Among the synthesized products, johnbaumite showed strongest As retention ability even under acidic (pH 9) conditions.

As high-grade and easily exploitable copper ores have been getting depleted, the development of flotation techniques for copper ores with low-grade and small grain size is becoming an important issue. To process these kinds of ores, fine grinding is essential to achieve sufficient liberation; however, flotation rate of finely ground minerals ( The addition of anionic emulsifiers (SDS, KAX) could stabilize the oil–water emulsion and produce smaller oil droplets, which improve Cu recovery in the agglomeration-flotation of chalcopyrite and quartz from ~68% (without emulsifier) to >97% (with emulsifier). When Ca2+ and quartz co-existed, the amounts of SDS used as emulsifier were reduced due to its adsorption onto the surface of Ca2+ activated quartz, which caused the coalescence of oil droplets resulting in low Cu recovery (i.e., 81% with quartz only; 44% with quartz and Ca2+). Meanwhile, emulsified oil stabilized by KAX was not affected by the presence of Ca2+ and quartz, and high Cu recovery (~83%) could be obtained.

Batch and continuous jig experiments were compared using plastic mixtures with different light and heavy particle ratios. The purity of bottom layer products of continuous jig was lower than those of batch jig because of light particle entrainment induced by the screw-type extractor. To minimize entrainment, a new discharge system consisting of a vertical restraining wall and a screw-type extractor was developed. The restraining wall was installed close to the “product-end” to separate the jig chamber into two and allow particles to transfer from one part to the other via the gap under the wall. The experimental results showed that this proposed discharge system could improve the purity of heavy particles in bottom layer products because of two reasons: (1) higher material mass ratio of heavy plastic close to product-end, and (2) changes in the flow of water and particles. Moreover, the purity of bottom layer products during continuous jig separation was influenced by the difference of material mass ratio in feed; that is, purity was higher when the ratio of heavy plastic was high. In addition, an estimation procedure based on multi-step treatment using separation curves is proposed to achieve the target purity of bottom layer products.

Porphyry ores and E-wastes/WEEE are two of the most important copper-bearing materials on the planet. Over 60% of world copper output comes from porphyry copper ores while E-waste(s) is globally the largest copper-bearing waste category since the 1980s. They also contain critical elements for low-carbon technologies essential in the clean energy transition's success. In this review, a critical analysis of ore distribution/processing, metal extraction, E-waste generation and E-waste recycling is presented, focusing on identifying challenges and how to address them with emerging technologies and sustainable socio-environmental strategies. Access to ore deposits is a major hurdle for mine development while the absence of a consistent E-waste classification and legislation, including poor collection rates, remains serious problems in E-waste recycling. As lower grade porphyry ores are exploited, difficulties in processing/extraction due to mineralogical complexities, very fine particles and the generation of “dirty” concentrates will become more prevalent. For E-wastes recycling, current trends are to develop smaller, more mobile, and eco-friendly hydrometallurgical alternatives to pyrometallurgy that can handle localised compositional and feed variabilities. Finally, more sustainable mine waste management strategies, including better LCIA tools with spatial and temporal dimensions, are needed to limit socio-environmental impacts of resources exploitation and maintain the sector's SLO.

A recycling process consisting of swelling, ammonia leaching and hydrochloric acid leaching was developed to recover Sn, Bi, Cu, and resin from Pb-free Sn-Bi-Cu solder paste. The resin was separated from metal powder by swelling the solder paste in methyl ethyl ketone at 30 °C, 200 rpm and 5% pulp density. The addition of distilled water precipitated the resin from methyl ethyl ketone. The recovered resin was confirmed to be epoxy resin by FT-IR analysis. Ammonia leaching followed by hydrochloric acid leaching was performed to separate Cu, Sn, Bi sequentially from metal powder. Copper was solubilized easily with leaching efficiency reaching 100% within 15 min under the following conditions: 5 M NH 3 solution, 1 M (NH 4 ) 2 CO 3 , 0.1 M CuCO 3 , 50 °C, and 1% pulp density. Tin and bismuth were recovered as leach residue in the ammonia leaching. Tin was selectively leached in 0.5 M HCl solution with 10,000 mg/L Sn 4 + and 1% pulp density at 50 °C and 400 rpm, while Bi was not detected in the leach solution. The recycling process has successfully separated Sn, Cu, Bi and resin.

Cadmium (Cd), cobalt (Co), nickel (Ni), and zinc (Zn) are important components of batteries and renewable energy technologies but their recovery from pregnant leach solutions or removal from polluted soil systems and water bodies remain challenging. In this study, we report the removal/recovery of heavy metal ions (Cd2+, Co2+ and Ni2+, and Zn2+) from sulfate solutions by a novel cementation technique using aluminum (Al) and activated carbon (AC) called “AC/Al-cementation”. The results showed ~0% removal/recovery efficiencies with Al only for the four heavy metals but with “AC/Al-cementation”, these values increased to 96%, 61%, 57%, and 45% for Zn2+, Ni2+, Co2+, and Cd2+, respectively. Detailed residue characterizations using SEM-EDX, XPS, and AES showed substantial deposition of heavy metal ions in their elemental forms, indicating that the primary mechanism of removal/recovery was cementation. Moreover, the results identified the critical role of AC as an electron pathway from Al (electron donor) to metal ions (electron acceptors) even though Al was covered with Al2O3. Finally, hydrogen overpotential and the subsequent passivation of cemented heavy metals by Al(OH)3 were important limiting factors and processes affecting the efficiency of “AC/Al-cementation” technique.

Mineral processing, pyro- and hydrometallurgical processes of auriferous sulfide ores and porphyry copper deposits (PCDs) generate arsenopyrite-rich wastes. These wastes are disposed of into the tailings storage facilities (TSF) in which toxic arsenic (As) is leached out and acid mine drainage (AMD) is generated due to the oxidation of arsenopyrite (FeAsS). To suppress arsenopyrite oxidation, this study investigated the passivation of arsenopyrite by forming ferric phosphate (FePO4) coating on its surface using ferric-catecholate complexes and phosphate simultaneously. Ferric iron (Fe3+) and catechol form three types of complexes (mono-, bis-, and triscatecholate complexes) depending on the pH, but monocatecholate complex (i.e.,[Fe(cat)](+)) became unstable in the presence of phosphate because the chemical affinity of Fe3+-PO43- is most probably stronger than that of Fe3+-catechol in [Fe(cat)](+). When two or more catechol molecules were coordinated with Fe3+ (i.e., [Fe(cat)(2)](-) and [Fe(cat)(3)](3-)), however, these complexes were stable irrespective of the presence of phosphate. The treatment of arsenopyrite with [Fe(cat)(2)](-) and phosphate could suppress its oxidation due to the formation of FePO4 coating, evidenced by SEM-EDX and XPS analyses. The mechanism of FePO4 coating formation by [Fe(cat)(2)](-) and phosphate was confirmed by linear sweep voltammetry (LSV): (1) [Fe(cat)(2)](-) was oxidatively decomposed and (2) the resultant product (i.e., [Fe(cat)](+)) reacts with phosphate, resulting in the formation of FePO4.

The flotation of ultrafine particles is marked with low recovery due to low collision efficiency between particles and bubbles. Carrier-flotation is a method to enhance the recovery by the attachment of fine particles onto larger carrier particles. In this study, the effects of coarse chalcopyrite particles as a carrier on flotation behavior of fine chalcopyrite was investigated by the flotation experiments using 20 g fine chalcopyrite (particle size D50 = 2.3 µm) with a varied amount of carrier particles of different sizes (–75 + 38 µm and –106 + 75 µm) using a 400 mL-mechanically agitated flotation cell. The addition of carrier particles improved the recovery of ultrafine particles into the froth from around 25% (without carrier) to around 80% (with 20 g of carrier, –75 + 38 µm size). Recovery of fine particles was higher with smaller carrier size (–75 + 38 µm) compared with larger carrier size (–106 + 75 µm), though the difference was not significant. Theoretical calculations showed that hydrophobic interaction played a more dominant role compared with van der Waals force in the attachment of fine chalcopyrite particles on the carrier (larger chalcopyrite particles).

Cobalt and nickel have become important strategic resources because they are widely used for renewable energy technologies and rechargeable battery production. Cementation, an electrochemical deposition of noble metal ions using a less noble metal as an electron donor, is an important option to recover Co and Ni from dilute aqueous solutions of these metal ions. In this study, cementation experiments for recovering Co2+ and Ni2+ from sulfate and chloride solutions (pH = 4) were conducted at 298 K using Al powder as electron donor, and the effects of additives such as activated carbon (AC), TiO2, and SiO2 powders on the cementation efficiency were investigated. Without additives, cementation efficiencies of Co2+ and Ni2+ were almost zero in both sulfate and chloride solutions, mainly because of the presence of an aluminum oxide layer (Al2O3) on an Al surface, which inhibits electron transfer from Al to the metal ions. Addition of nonconductor (SiO2) did not affect the cementation efficiencies of Co2+ and Ni2+ using Al as electron donor, while addition of (semi)conductors such as AC or TiO2 enhanced the cementation efficiencies significantly. The results of surface analysis (Auger electron spectroscopy) for the cementation products when using TiO2/Al mixture showed that Co and Ni were deposited on TiO2 particles attached on the Al surface. This result suggests that conductors such as TiO2 act as an electron pathway from Al to Co2+ and Ni2+, even when an Al oxide layer covered on an Al surface.

Zinc plant leach residues (ZPLRs) are hazardous solid wastes generated from zinc metal production owing to their substantial contents of lead (Pb), a toxic heavy metal. This study investigated the detoxification of historic ZPLRs from Kabwe, Zambia by removing Pb using a coupled extraction-cementation method in chloride media. For the coupled extraction-cementation method, micro-scale zero-valent iron (mZVI) was added during ZPLRs leaching in acidified chloride solution. Cemented Pb on the surface of mZVI was recovered easily from the leaching pulp by magnetic separation. Pb removal was evaluated in different solution compositions (NaCl:1–5.13 M, HCl: 0–0.1 M) with and without the addition of mZVI. The addition of mZVI during ZPLRs leaching (i.e., coupled extraction-cementation) increased Pb removal from 3% to 24%, 1.3% to 27.5%, 5.2% to 34.9%, and 6.5% to 55.8% when NaCl concentration was fixed at 0.86 M and HCl concentrations were 0 M, 0.01 M, 0.05 M and 0.1 M, respectively. When NaCl concentration was increased above 3.42 M and HCl maintained at 0.1 M HCl, Pb removal increased to 80%. Analysis of the Pb-loaded mZVI (magnetic fraction) by scanning electron microscopy with energy dispersive X-ray spectroscopy (SEM-EDX) and X-ray photoelectron spectroscopy (XPS) revealed that Pb was recovered during leaching via cementation as Pb°. The toxicity characteristic leaching procedure (TCLP) and in vitro solubility and bioavailability research consortium gastric phase (SBRC-G) tests for Pb of ZPLRs before and after treatment decreased drastically from 11.3 to 3.5 mg/L (below 5 mg/L threshold) and 12 300 to 2 840 mg/Kg, respectively.

Flotation is the conventional method for processing porphyry copper deposits, one of the most economically important sources of copper (Cu) worldwide. The rapidly decreasing grade of this type of Cu ore in recent years, however, presents serious problems with fine particle recovery using conventional flotation circuits. This low recovery could be attributed to the low collision efficiency of fine particles and air bubbles during flotation. To improve collision efficiency and flotation recovery, agglomeration of finely ground chalcopyrite (CuFeS2) (D50 = 3.5 &mu, m) using emulsified oil stabilized by emulsifiers was elucidated in this study. Specifically, the effects of various types of anionic (sodium dodecyl sulfate (SDS), potassium amyl xanthate (KAX)), cationic (dodecyl amine acetate (DAA)), and non-ionic (polysorbate 20 (Tween 20)) emulsifiers on emulsified oil stability and agglomeration&ndash, flotation efficiency were investigated. When emulsifiers were added, the average size of agglomerates increased, resulting in higher Cu recovery during flotation. This dramatic improvement in flotation efficiency could be attributed to the smaller oil droplet size in emulsified oil and their higher stability in the presence of emulsifiers. The utilization of emulsifiers during agglomeration&ndash, flotation not only lowered the required agitation strength for agglomeration but also shortened the agglomeration time, both of which made the process easier to incorporate in existing flotation circuits.

The redox potential dependence of chalcopyrite leaching was investigated in acidic ferric chloride solutions using a specially designed batch leaching experimental setup that accurately determined and monitored changes of solution redox potential with time. The concentration ratio of ferric (Fe3+) to ferrous (Fe2+) ions and leaching rate-redox potential relationship were then determined from measured redox potential data assuming chalcopyrite oxidation with Fe3+ occurred stoichiometrically. A distinct peak was observed in the leaching rate vs. redox potential plot; that is, the leaching rate increased with increasing redox potential, reaching a (local) maximum (i.e., the peak rate) at a certain redox potential (i.e., the peak redox potential) followed by the gradual decrease in leaching rate at higher redox potentials. For example, in the experiment performed at [Cu2+] = [Fe2+] = 0.01 kmol m−3 in HCl (0.1 kmol m−3) at 343 K, the rate continuously increased with increasing redox potential, reaching the fastest rate of 9.7 × 10−8 mol s−1 g−1 (peak rate) at 0.442 V (peak redox potential). After passing this redox potential peak, the rate started to decrease, slowing down to 2.9 × 10−8 mol s−1 g−1 at 0.487 V. Using this setup, the effects of temperature and solution composition (Cu2+, Fe3+ and HCl concentrations) on the peak rate and peak redox potential were investigated. The results showed that the peak rate increased with increasing temperature, decreased with increasing HCl and Fe3+ concentrations but was not strongly affected by Cu2+ concentration. In comparison, the peak redox potential was less dependent on the temperature and concentrations of HCl and Fe3+. Although the role of Cu2+ was apparently not crucial in the occurrence of the peak rate, it exerted a relatively strong influence on the peak redox potential; that is, the peak redox potential increased with increasing Cu2+ concentration. Based on these results, an empirical equation for the relationship between peak redox potential and Cu2+ concentration was derived.

Complex sulfide ores are typically mixtures of various sulfide minerals like sphalerite (ZnS), chalcopyrite (CuFeS2), galena (PbS), pyrite (FeS2), and barite (BaSO4) and processed by selective flotation to recover target-sulfide minerals like chalcopyrite and sphalerite. Some complex sulfide ores, however, contain anglesite (PbSO4), a mineral with relatively high solubility, that complicates selective flotation because its dissolution releases Pb2+, which ‘activates’ co-existing sulfide minerals (e.g., ZnS). Because both target and non-target sulfide minerals are recovered when flotation is non-selective, another flotation stage to recover target minerals in froth products is required. Aside from anglesite, co-existing gangue sulfide minerals like pyrite also complicate selective flotation because of their strong effects on the floatability of target-sulfide minerals via electrochemical interactions. In this study, the effects of pyrite on the floatability of lead-activated sphalerite were investigated using flotation tests and electrochemical techniques coupled with contact angle measurements. Moreover, factors important to lead-activated sphalerite floatability were elucidated in detail using dissolution experiments and X-ray photoelectron spectroscopy (XPS). Finally, a mechanism for lead-activated sphalerite depression by pyrite is proposed.