Search

Iron oxides (hematite, Fe2O3, and magnetite, Fe3O4), previously used as electron mediators in the galvanic system with zero-valent aluminum (ZVAl), have been shown to recover Au upon cementation in Au–Cu ammoniacal thiosulfate media selectively, and this warrants further investigation. This research is focused on investigating the role of the semiconductive properties of metal oxides by performing a cementation experiment by mixing 0.15 g of electron mediators (Fe3O4, Fe2O3, TiO2 (anatase and rutile)) and 0.15 g of zero-valent aluminum powder as an electron donor in various electrochemical experiments. The results revealed that upon the cementation experiment, synthetic Fe2O3 and Fe3O4 were consistently able to selectively recover Au at around 90% and Cu at around 20%. Compared to activated carbon (AC), TiO2, in anatase and rutile forms, obtained selective recovery of gold, but the recovery was utterly insignificant compared to that of iron oxides, obtaining an average of 93% Au and 63% Cu recovery. The electrochemical and surface analysis supports the results obtained upon the cementation process, where TiO2, upon cyclic voltammetry (CV), obtained two reduction peaks centered at −1.0 V and −0.5 V assigned to reducing Au and Cu ions, respectively. Furthermore, various electrochemical impedance spectroscopic analyses revealed that the flat band potential obtained in the Mott–Schottky plot is around −1.0 V and −0.2 V for iron oxides and titanium oxides, respectively, suggesting that the electrons travel from semiconductor interface to electrolyte interface, and electrons are accessible only to Au ions in the electrolyte interface (reduction band edge around −1.0 V). The determination of this selective cementation mechanism is one of a kind. It has been proposed that the semiconductive properties of Fe2O3, Fe3O4, and, by configuring their relative energy band diagram, the travel of electrons from the iron oxide–electrolyte interface facilitate the selective cementation towards Au(S2O3)23+ ions in gold–copper ammoniacal thiosulfate solutions.

Microbially induced carbonate precipitation (MICP), a widespread phenomenon in nature, is gaining attention as a low-carbon alternative to ordinary Portland cement (OPC) in geotechnical engineering and the construction industry for sustainable development. In the Philippines, however, very few works have been conducted to isolate and identify indigenous, urease-producing (ureolytic) bacteria suitable for MICP. In this study, we isolated seven, ureolytic and potentially useful bacteria for MICP from marine sediments in Iligan City. DNA barcoding using 16s rDNA identified six of them as Pseudomonas stutzeri, Pseudomonas pseudoalcaligenes, Bacillus paralicheniformis, Bacillus altitudinis, Bacillus aryabhattai, and Stutzerimonas stutzeri but the seventh was not identified since it was a bacterial consortium. Bio-cementation assay experiments showed negligible precipitation in the control (without bacteria) at pH 7, 8, and 9. However, precipitates were formed in all seven bacterial isolates, especially between pH 7 and 8 (0.7–4 g). Among the six identified bacterial species, more extensive precipitation (2.3–4 g) and higher final pH were observed in S. stutzeri, and B. aryabhattai, which indicate better urease production and decomposition, higher CO2 generation, and more favorable CaCO3 formation. Characterization of the precipitates by scanning electron microscopy with energy dispersive X-ray spectroscopy (SEM-EDS) and attenuated total reflectance Fourier transform spectroscopy (ATR-FTIR) confirmed the formation of three carbonate minerals: calcite, aragonite, and vaterite. Based on these results, all six identified indigenous, ureolytic bacterial species from Iligan City are suitable for MICP provided that the pH is controlled between 7 and 8. To the best of our knowledge, this is the first report of the urease-producing ability and potential for MICP of P. stutzeri, P. pseudoalcaligenes, S. stutzeri, and B. aryabhattai.

Rare earth elements (REEs) are crucial for green energy applications due to their unique properties, but their extraction poses sustainability challenges because the global supply of REEs is concentrated in a few countries, particularly China, which produces 70% of the world’s REEs. To address this, the study investigated TK221, a modified extraction chromatographic resin featuring diglycolamide (DGA) and carbamoyl methyl phosphine oxide (CMPO), as a promising adsorbent for REE recovery. The elemental composition and functional groups of DGA and CMPO on the polystyrene-divinylbenzene (PS-DVB) support of TK221 were confirmed using scanning electron microscopy with energy dispersive X-ray spectroscopy (SEM-EDX), attenuated total reflectance Fourier transform infrared spectroscopy (ATR-FTIR), and X-ray photoelectron spectroscopy (XPS). The adsorption kinetics of neodymium (Nd), yttrium (Y), cerium (Ce), and erbium (Er) followed the pseudo-second-order kinetic model and Langmuir isotherm, indicating monolayer chemisorption. Furthermore, iron (Fe) adsorption reached apparent equilibrium after 360 min, with consistent Fe adsorption observed at both 360 min and 1440 min. The inclusion of Fe in the study is due to its common presence as an impurity in most REE leachate solutions. The Fe adsorption isotherm results are better fitted with the Langmuir isotherm, implying chemisorption. Maximum adsorption capacities (qmax) of the resin were determined as follows: Nd (45.3 mg/g), Ce (43.1 mg/g), Er (35.1 mg/g), Y (15.6 mg/g), and Fe (12.3 mg/g). ATR-FTIR analysis after adsorption suggested that both C=O and P=O bands shifted from 1679 cm−1 to 1618 cm−1 and 1107 cm−1 to 1142 cm−1 for Y, and from 1679 cm−1 to 1607 cm−1 and 1107 cm−1 to 1135 cm−1 for Ce, implying possible coordination with REEs. These results suggest that TK221 has a huge potential as an alternative adsorbent for REE recovery, thus contributing to sustainable REE supply diversification.

“Affordable and clean energy” is enshrined in the UN Sustainable Development Goals (SDGs; #7) because of its importance in supporting the sustainable development of society. As an energy source, coal is widely used because it is abundant and its utilization for electricity and heat generation do not require complex infrastructures and technologies, which makes it ideal for the energy needs of low-income and developing countries. Coal is also essential in steel making (as coke) and cement production and will continue to be on high demand for the foreseeable future. However, coal is naturally found with impurities or gangue minerals like pyrite and quartz that could create by-products (e.g., ash) and various pollutants (e.g., CO2, NOX, SOX). To reduce the environmental impacts of coal during combustion, coal cleaning—a kind of pre-combustion clean coal technology—is essential. Gravity separation, a technique that separates particles based on their differences in density, is widely used in coal cleaning due to the simplicity of its operation, low cost, and high efficiency. In this paper, recent studies (from 2011 to 2020) related to gravity separation for coal cleaning were systematically reviewed using the Preferred Reporting Items for Systematic Reviews and Meta-Analyses (PRISMA) guidelines. A total of 1864 articles were screened after removing duplicates, and after a thorough evaluation 189 articles were reviewed and summarized. Among of conventional separation techniques, dense medium separator (DMS), particularly dense medium cyclone (DMC), is the most popular technologies studied, which could be attributed to the growing challenges of cleaning/processing fine coal-bearing materials. In recent years, most of works focused on the development of dry-type gravity technologies for coal cleaning. Finally, gravity separation challenges and future applications to address problems in environmental pollution and mitigation, waste recycling and reprocessing, circular economy, and mineral processing are discussed.

We developed a technique called the reverse hybrid jig, an advanced physical separation technique that combines the principles of jig and flotation to separate floating plastics. This technique is a promising green technology that is more economical and environmentally friendly compared with the conventional flotation. Although the applicability of this technique to separate PP/PE have been reported, the index to illustrate the possibility of separation for the reverse hybrid jig is still not available. In this study, a reverse apparent concentration criterion (CCRA) is proposed to estimate reverse hybrid jig separation efficiency. This modified concentration criterion can be calculated using the specific gravity (SG) of particle with attached bubbles called the apparent specific gravity (SGA). To determine the volume of attached bubbles on plastic surfaces under water pulsation, a laser-assisted apparatus was used under various conditions, including plastic type, air flow rate, dosage, and type of wetting agent. The results of attached bubble volume measurements were used to calculate the SGA and CCRA. The estimated values were then compared with the results of reverse hybrid jig separation. It was found that higher CCRA resulted in better separation efficiency. In addition, an empirical linear equation for estimating the reverse hybrid jig separation efficiency is proposed.

The use of thiosulfate to extract gold from refractory ores is promising because of its non-toxicity and high selectivity. Sulfide minerals (i.e., pyrite, arsenopyrite, chalcopyrite), major gold carriers in refractory gold ores, however, hinder gold extraction due to the high consumption of a lixiviant. In this study, a new method to improve gold extraction from sulfide bearing gold ores is proposed based on the model experiments using a mixture of gold powder and arsenopyrite-bearing sulfide (HAsBS) ore. The effects of HAsBS ore on gold leaching in ammonium thiosulfate solutions were investigated, and it was found that gold extraction in the presence of HAsBS ore was suppressed because of the unwanted decomposition of thiosulfate on the surface of sulfide minerals. To improve gold extraction in the presence of the sulfide minerals, this study investigated the effects of the pretreatment of HAsBS ore using ammonium solutions containing cupric ions and confirmed that HAsBS ore was oxidized in the pretreatment and its surface was covered by the oxidation products. As a result, thiosulfate consumption was minimized in the subsequent gold leaching step using ammonium thiosulfate, resulting in an improvement in gold extraction from 10% to 79%.

Ammonium thiosulfate leaching is a promising alternative to the conventional cyanide method for extracting gold from ores. However, strategies for recovering gold from the leachate are less commercially used due to its low affinity to gold. The present study investigated the recovery of gold from the leachate using iron oxides (hematite, Fe2O3 or magnetite, Fe3O4). Cementation experiments were conducted by mixing 0.15 g of aluminum powder as an electron donor and 0.15 g of an electron mediator (activated carbon, hematite, or magnetite) in 10 mL of ammonium thiosulfate leachate containing 100 mg/L gold ions and 10 mM cupric ions for 24 h at 25 °C. The results of the solution analysis showed that when activated carbon (AC) was used, the gold was recovered together with copper (recoveries were 99.99% for gold and copper). However, selective gold recovery was observed when iron oxides were used, where the gold and copper recoveries were 89.7% and 21% for hematite and 85.9% and 15.4% for magnetite, respectively. An electrochemical experiment was also conducted to determine the galvanic interaction between the electron donor and electron mediator in a conventional electrochemical setup (hematite/magnetite–Al as the working electrode, Pt as the counter electrode, Ag/AgCl as the reference electrode) in a gold–thiosulfate medium. Cyclic voltammetry showed a gold reduction “shoulder-like” peak at −1.0 V using hematite/Al and magnetite/Al electrodes. Chronoamperometry was conducted and operated at a constant voltage (−1.0 V) determined during cyclic voltammetry and further analyzed using SEM-EDX. The results of the SEM-EDX analysis for the cementation products and electrochemical experiments confirmed that the gold was selectively deposited on the iron oxide surface as an electron mediator.

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.

Rare earth minerals (REMs) contain rare earth elements (REEs) that are important in modern technologies due to their unique magnetic, phosphorescent, and catalytic properties. However, REMs are not only non-renewable resources but also non-uniformly distributed on the Earth’s crust, so the processing of REE-bearing secondary resources via recycling is one potential route to ensure the long-term sustainability of REE supply. Flotation—a method that separates materials based on differences in their surface wettability—is a process applied for both mineral processing and recycling of REEs, especially when the particles are fine and/or a high-purity product is required. In this review, studies about rare earth flotation from 2012 to 2021 were systematically reviewed using the PRISMA guideline. It was found that most REM flotation research works focused on finding better collectors and depressants while, for recycling, studies on advanced flotation techniques like froth flotation, ion flotation, solvent sublation, electroflotation, and adsorbing colloid flotation with an emphasis on the recovery of dissolved REEs from aqueous solutions dominated.

In this study, the microencapsulation-hybrid jig separation technique was developed to improve the separation efficiency of pyrite and coal in the particle size range of 1–4 mm where conventional jig separation becomes inefficient. A hybrid jig is a gravity concentrator combining the concepts of jig separation and flotation to stratify particles based on their apparent specific gravity. Meanwhile, microencapsulation—a technique that encapsulates target materials with a protective coating—was applied to render pyrite hydrophilic and improve its separation from hydrophobic coal. The results showed that the required time for separation in the hybrid jig (0.5 min) was shorter than in conventional jig (2 min). Moreover, the effects of particle size on separation efficiency were reduced when a hybrid jig is used. However, the separation efficiency of hybrid jig separation was lower than that of the conventional jig because attachment of bubbles occurred to both pyrite and coal, which are hydrophobic. Using the microencapsulation-hybrid jig separation technique, the separation of coal and pyrite was significantly improved (~100%) because of the formation of hydrophilic iron phosphate coatings on pyrite that limited bubble attachment. This means that microencapsulation-hybrid jig separation is a promising clean coal technology that not only enhances the separation efficiency of the hybrid jig but also passivates pyrite and limits AMD formation in the tailings/rejects.

In 2015, 193 governments agreed to act on climate change by drastically reducing carbon dioxide (CO2) emissions as envisaged in the sustainable development goal (SDG) number 13 [...]

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.

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: 2+ due to the galvanic interaction with chalcopyrite, which resulted in the elimination of the effectiveness of zinc sulfate in depressing sphalerite floatability (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 presence of anglesite (PbSO4) in complex sulfide ores negatively affects the separation of Cu-Pb sulfides and sphalerite (ZnS) due to lead activation, and PbSO4 rejected to tailings dams contaminates the surrounding environment with lead. To address these problems, this study investigated the application of ethylene diamine tetra acetic acid (EDTA) pretreatment extracting PbSO4 to ZnS flotation and the recovery of the extracted Pb2+ as zero-valent Pb by cementation using zero-valent iron (ZVI). The application of EDTA pretreatment could extract ~99.8% of PbSO4, thus depressing ZnS floatability from 82% to 30%. In addition, cementation using ZVI could recover ~99.7% of Pb2+ from the leachate of EDTA pretreatment.

The continuous accumulation of artisanal and small-scale gold mining (ASGM) tailings in the Philippines without adequate storage and disposal facility could lead to human health and environmental disasters in the long run. In this study, ASGM tailings was simultaneously stabilized and repurposed as construction material via geopolymerization using coal fly ash, palm oil fuel ash and a powder-based alkali activator. Scanning electron microscopy with energy dispersive X-ray spectroscopy (SEM-EDS) identified iron sulfides in the tailings containing arsenic (As), cadmium (Cd), copper (Cu), lead (Pb) and zinc (Zn), which could be released via weathering. The average unconfined compressive strengths (UCS) of tailings-based geopolymers at 14 days curing were 7.58 MPa and 7.7 MPa with fly ash and palm oil fuel ash, respectively. The tailings-based geopolymers with palm oil fuel ash had higher UCS most likely due to CASH reaction product formation that improved strength formation. The toxicity characteristic leaching procedure (TCLP) results showed very low leachabilities of As, Pb and Fe in the geopolymer materials suggesting ASGM tailings was effectively encapsulated within the geopolymer matrix. Overall, the geopolymerization of ASGM tailings is a viable and promising solution to simultaneously stabilize mining and industrial wastes and repurpose them into construction materials.

Flotation is a common mineral processing method used to upgrade copper sulfide ores; in this method, copper sulfide mineral particles are concentrated in froth, and associated gangue minerals are separated as tailings. However, a significant amount of copper is lost into tailings during the processing; therefore, tailings can be considered secondary resources or future deposits of copper. Particle–bubble collision efficiency and particle–bubble aggregate stability determines the recovery of target particles; this attachment efficiency plays a vital role in the selectivity process. The presence of fine particles in the flotation circuit is because of excessive grinding, which is to achieve a higher degree of liberation. Complex sulfide ores of markedly low grade further necessitate excessive grinding to achieve the maximum degree of liberation. In the flotation process, fine particles due to their small mass and momentum are unable to collide with rising bubbles, and their rate of flotation is very slow, further lowering the recovery of target minerals. This collision efficiency mainly depends on the particle–bubble size ratio and the concentration of particles present in the pulp. To overcome this problem and to maintain a favorable particle–bubble size ratio, different techniques have been employed by researchers to enhance particle–bubble collision efficiency either by increasing particle size or by decreasing bubble size. In this article, the mechanism of tailing loss is discussed in detail. In addition, flotation methods for fine particles recovery such as microbubble flotation, column flotation, nanobubble flotation, polymer flocculation, shear flocculation, oil agglomeration, and carrier flotation are reviewed, and their applications and limitations are discussed in detail.

Zinc plant leach residues (ZPLRs), particularly those produced using old technologies, have both economic importance as secondary raw materials and have environmental impacts because they contain hazardous heavy metals that pose risks to human health and the environment. Therefore, the extraction and recovery of these metals from ZPLRs has both economic and environmental benefits. In this study, we investigated the removal of lead (Pb) and zinc (Zn) from ZPLRs by alkaline (NaOH) leaching and the concurrent cementation of dissolved Pb and Zn using aluminum (Al) metal powder. The effects of the leaching time, NaOH concentration, solid-to-liquid ratio (S/L), and dosage of Al metal powder on the extraction of Pb and Zn were investigated. Pb and Zn removal efficiencies increased with increasing NaOH concentrations and decreasing S/Ls. The Pb and Zn removal efficiencies were 62.2% and 27.1%, respectively, when 2.5 g/50 mL (S/L) of ZPLRs were leached in a 3 M NaOH solution for 30 min. The extraction of Pb and Zn could be attributed to the partitioning of these metals in relatively more mobile phases—water-soluble, exchangeable, and carbonate phases—in ZPLRs. Around 100% of dissolved Pb and less than 2% of dissolved Zn were cemented in leaching pulp when Al metal powder was added. Minerals in the solid residues, particularly iron oxides minerals, were found to suppress the cementation of extracted Zn in leaching pulp, and when they were removed by filtration, Zn was recovered by Al metal powder via cementation.

Mining activities often generate large amounts of sulfide-rich wastes containing arsenopyrite (FeAsS), which when dissolved releases toxic arsenic (As) and generates acid mine drainage (AMD) that are both disastrous to the environment. To suppress arsenopyrite dissolution, a technique that selectively coats sulfide minerals with a protective layer of Al-oxyhydroxide called Al-based carrier-microencapsulation (CME) was developed. Although a previous study of the authors showed that Al-based CME could significantly limit arsenopyrite dissolution, nature of the coating formed on arsenopyrite, including its electrochemical properties, is still not well understood. Moreover, stability of the coating once exposed to weathering conditions remains unclear. Better understanding of these important issues would greatly improve Al-based CME especially in its application to real mine wastes. In this study, nature of the coating formed by Al-based CME was investigated using SEM-EDX, DRIFTS and XPS while the electrochemical properties of the coating were evaluated by cyclic voltammetry and chronoamperometry. Meanwhile, stability of the coating was elucidated using consecutive batch leaching experiments and weathering cell tests.SEM-EDX, DRIFTS and XPS results indicate that the protective coating formed on arsenopyrite by Al-based CME was mainly composed of bayerite (α-Al(OH)3), gibbsite (γ-Al(OH)3), and boehmite (γ-AlO(OH)). These Al-based coatings, which have insulating properties, made arsenopyrite less electrochemically active. The coatings also limited the extent of both the anodic and cathodic half-cell reactions of arsenopyrite oxidation that suppressed As release and acid generation. Weathering cell tests indicated that the oxidation of CME-treated arsenopyrite was effectively limited until about 15 days but after this, it started to gradually progress with time due to the increasing acidity of the system where Al-based coatings became unstable. Nonetheless, CME-treated arsenopyrite was less oxidized based on the released amounts of Fe, As and S suppressed by 80, 60 and 70%, respectively, compared with the one treated with control.

Metals are a finite resource that are necessary to maintain living standards in modern society, due to their countless applications, such as transportation vehicles, building and construction, household appliances, electronic devices, etc. [...]

The enhanced cementation technique by galvanic interaction of aluminum (Al; electron donor) and activated carbon (AC; electron mediator) to recover gold (Au) ions from the ammonium thiosulfate solution is a promising technique to eliminate the challenges of poor recovery in the system. This study presents the kinetics of Au ion cementation in an ammonium thiosulfate lixiviant as functions of initial Au concentration, size/amount of Al and AC, temperature, and shaking speed. The recovery results basically followed first order kinetics and showed that the cementation rate increased with a higher initial concentration of Au, smaller electron donor size, greater both electron donor and mediator quantity, decrease in temperature, and higher shaking speed in the system, while size of electron mediator did not significantly affect Au recovery.

Non-availability of rich primary resources of rare earth metals (REMs) and the generation of huge amounts of discarded magnets containing REMs, compelled the researchers to explore the possibilities for the recovery of REMs from discarded magnets. Therefore, the present paper reports the recovery of REMs (Nd, Pr, and Dy) from discarded Nd-Fe-B magnets. The process consists of demagnetization, pre-treatment, and hydrometallurgical processing to recover REMs as salt. Leaching studies indicate that 95.5% Nd, 99.9% Pr, and 99.9% Dy were found to be dissolved at the optimized experimental condition i.e., acid concentration 2 M H2SO4, temperature 75 °C, pulp density 100 g/L, and mixing time 60 min. Solvent extraction technique was tried for the selective extraction/separation of REMs and Fe. The result indicates that 99.1% (24.42 g/L) of Nd along with 90% (1.08 g/L) of Pr and total Fe were co-extracted using 35% Cyanex 272 at organic to aqueous (O/A) ratio 1/1, eq. pH 3.5 in 10 min of mixing time. It requires multistage separation and therefore, not feasible in view of economics. Thus, direct precipitation of REMs salt and iron oxide as pigment was studied using two stages of precipitation at different pH. The obtained precipitate of REMs and Fe hydroxides were dried separately to remove the moisture and further treated at elevated temperature to get pure REMs oxide and red oxide.

The global demand for rare earth elements (REEs) is expected to increase significantly because of their importance in renewable energy and clean storage technologies, which are critical for drastic carbon dioxide emission reduction to achieve a carbon-neutral society. REE ore deposits around the world are scarce and those that have been identified but remain unexploited need to be developed to supply future demands. In this study, the Khalzan Buregtei deposit located in western Mongolia was studied with the aim of upgrading low-grade REE ore via magnetic separation techniques. The total REE content in this ore was ~6720 ppm (~3540 ppm light REE (LREE) + ~3180 ppm heavy REE (HREE)) with bastnaesite, pyrochlore, synchysite, and columbite-(Fe) identified as the main REE-bearing minerals. As the particle size fraction decreased from −4.0 + 2.0 mm to −0.5 + 0.1 mm, the recovery by dry high-intensity magnetic separation (DHIMS) increased from 20% to 70% of total rare earth oxide (TREO) while the enrichment ratio reached 2.8 from 1.3. Although effective, gangue minerals such as quartz and aluminosilicates were recovered (~22%) due most likely to insufficient liberation. Meanwhile, the wet high-intensity magnetic separation (WHIMS) could produce a magnetic concentrate with TREO recovery of ~80% and enrichment ratio of 5.5 under the following conditions: particle size fraction, −106 + 75 μm; feed flow rate, 3.2 L/min; magnetic induction, 0.8 T. These results indicate that combining DHIMS and WHIMS to upgrade the low-grade REE ore from the Khalzan Buregtei deposit is an effective approach.

Nickel metal hydride (NiMH) batteries are extensively used in the manufacturing of portable electronic devices as well as electric vehicles due to their specific properties including high energy density, precise volume, resistance to overcharge, etc. These NiMH batteries contain significant amounts of rare earth metals (REMs) along with Co and Ni which are discarded due to illegal dumping and improper recycling practices. In view of their strategic, economic, and industrial importance, and to mitigate the demand and supply gap of REMs and the limited availability of natural resources, it is necessary to explore secondary resources of REMs. Therefore, the present paper reports a feasible hydrometallurgical process flowsheet for the recovery of REMs and valuable metals from spent NiMH batteries. More than 90% dissolution of REMs (Nd, Ce and La) was achieved using 2 M H2SO4 at 75 °C in 60 min in the presence of 10% H2O2 (v/v). From the obtained leach liquor, the REMs, such as Nd and Ce, were recovered using 10% PC88A diluted in kerosene at eq. pH 1.5 and O/A ratio 1/1 in two stages of counter current extraction. La of 99% purity was selectively precipitated from the leach liquor in the pH range of 1.5 to 2.0, leaving Cu, Ni and Co in the filtrate. Further, Cu and Ni were extracted with LIX 84 at equilibrium pH 2.5 and 5, leaving Co in the raffinate. The developed process flow sheet is feasible and has potential for industrial exploitation after scale-up/pilot trails.

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.

Porphyry-type deposits are the major sources of copper and molybdenum, and flotation has been adopted to recover them separately. The conventional reagents used for depressing copper minerals, such as NaHS, Na2S, and Nokes reagent, have the potential to emit toxic H2S gas when pulp pH was not properly controlled. Thus, in this study the applicability of microencapsulation (ME) using ferrous and phosphate ions as an alternative process to depress the floatability of chalcopyrite was investigated. During ME treatment, the use of high concentrations of ferrous and phosphate ions together with air introduction increased the amount of FePO4 coating formed on the chalcopyrite surface, which was proportional to the degree of depression of its floatability. Although ME treatment also reduced the floatability of molybdenite, ~92% Mo could be recovered by utilizing emulsified kerosene. Flotation of chalcopyrite/molybdenite mixture confirmed that the separation efficiency was greatly improved from 10.9% to 66.8% by employing ME treatment as a conditioning process for Cu-Mo flotation separation.

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.

The purpose of this study is to propose the flotation procedure of seafloor massive sulfide (SMS) ores to separate chalcopyrite and galena as froth and sphalerite, pyrite, and other gangue minerals as tailings, which is currently facing difficulties due to the presence of water-soluble compounds. The obtained SMS ore sample contains CuFeS2, ZnS, FeS2, SiO2, and BaSO4 in addition to PbS and PbSO4 as Pb minerals. Soluble compounds releasing Pb, Zn2+, Pb2+, and Fe2+/3+ are also contained. When anglesite co-exists, lead activation of sphalerite occurred, and thus sphalerite was recovered together with chalcopyrite as froth. To remove soluble compounds (e.g., anglesite) that have detrimental effects on the separation efficiency of chalcopyrite and sphalerite, surface cleaning pretreatment using ethylene diamine tetra acetic acid (EDTA) was applied before flotation. Although most of anglesite were removed and the recovery of chalcopyrite was improved from 19% to 81% at 20 g/t potassium amyl xanthate (KAX) after EDTA washing, the floatability of sphalerite was not suppressed. When zinc sulfate was used as a depressant for sphalerite after EDTA washing, the separation efficiency of chalcopyrite and sphalerite was improved due to deactivation of lead-activated sphalerite by zinc sulfate. The proposed flotation procedure of SMS ores—a combination of surface cleaning with EDTA to remove anglesite and the depression of lead-activated sphalerite by using zinc sulfate—could achieve the highest separation efficiency of chalcopyrite and sphalerite; that is, at 200 g/t KAX, the recoveries of chalcopyrite and sphalerite were 86% and 17%, respectively.

Porphyry Cu-Mo deposits, which are the most important sources of copper and molybdenum, are typically processed by flotation. In order to separate Cu and Mo minerals (mostly chalcopyrite and molybdenite), the strategy of depressing chalcopyrite while floating molybdenite has been widely adopted by using chalcopyrite depressants, such as NaHS, Na2S, and Nokes reagent. However, these depressants are potentially toxic due to their possibility to emit H2S gas. Thus, this study aims at developing a new concept for selectively depressing chalcopyrite via microencapsulation while using Fe2+ and PO43− forming Fe(III)PO4 coating. The cyclic voltammetry results indicated that Fe2+ can be oxidized to Fe3+ on the chalcopyrite surface, but not on the molybdenite surface, which arises from their different electrical properties. As a result of microencapsulation treatment using 1 mmol/L Fe2+ and 1 mmol/L PO43−, chalcopyrite was much more coated with FePO4 than molybdenite, which indicated that selective depression of chalcopyrite by the microencapsulation technique is highly achievable.

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.

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 μm) 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–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–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.

A new concept for copper (Cu) metal leaching by the simultaneous use of cupric ions (Cu2+) and oxygen (O2) was proposed to improve Cu metal leaching in sulfuric acid. According to this concept, Cu2+ oxidizes Cu metal into cuprous ion (Cu+), and O2 oxidizes Cu+ into Cu2+. The improvement in Cu leaching efficiency from Cu metal was investigated experimentally in the sulfuric acid solution using Cu2+ and O2 simultaneously. Furthermore, the result was compared with that for the sulfuric acid solution containing neither Cu2+ nor O2 and with the sulfuric acid solution without Cu2+ and O2. When both Cu2+ and O2 were used in the leaching solution, the leaching rate of Cu from Cu metal powder was higher than at other leaching conditions, and the leaching efficiency of Cu increased to more than 99.9% in 1 mol/L sulfuric acid solution at 400 rpm and 50 °C with ≤75 µm Cu metal powder, 1% pulp density, 10,000 mg/L initial Cu2+ concentration, and 100 cc/min O2 introduction. These results indicated that the leaching of Cu from Cu metal could be accelerated by adding Cu2+ and O2 in the sulfuric acid solution.

Zinc plant leach residues (ZPLRs) contain significant amounts of metal compounds of lead (Pb), zinc (Zn), iron (Fe), etc., hence, they are considered as a secondary source of metals. On the other hand, ZPLRs are regarded as hazardous materials because they contain heavy metals that pollute the environment. Resources and environmental concerns of ZPLRs were addressed in this study by removing/recovering Pb and Zn using a concurrent dissolution and cementation technique. To cement the dissolved Pb and Zn in leaching pulp, zero-valent aluminum (ZVAl) was added during ZPLRs leaching in the hydrochloric (HCl)–sodium chloride (NaCl) solution. The resulting cemented metals were agglomerated and separated by sieving. Lead removal increased with increasing both NaCl and HCl concentrations. However, when ZVAl was added, significant Pb removal was achieved at a low concentration. Zinc was not cemented out of the pulp using ZVAl and its recovery from ZPLRs was dependent on the HCl concentration only. By applying a concurrent dissolution and cementation technique, both Pb and Zn were removed using a low concentration of NaCl, and most importantly Pb—the most toxic metal in ZPLRs—was captured and separated before the solid-liquid separation, hence, eliminating the need for extensive washing of the generated residues to remove the inherent residual solution.

In flotation, the size of mineral particles is one of the most important parameters: when the size becomes fine, collision efficiency of the particles and air bubbles becomes low, causing low flotation recovery. To improve the collision efficiency and flotation kinetics, agglomeration using the emulsified oil of finely ground chalcopyrite (D50 = 3.5 μm) was carried out before flotation. In this study, the effects of agitation strength during agglomeration, kerosene dosage and potassium amyl xanthate (KAX) dosage on the flotation were investigated. Agglomeration using emulsified oil improved Cu recovery because the median diameter of agglomerate increased. With increasing agitation strength, KAX and kerosene dosages, Cu recovery was further increased. Agglomeration-flotation of a mixture containing chalcopyrite and quartz with 1:1 ratio (w/w, weight by weight) showed that Si recovery in froth was low and did not change with varying conditions (agitation strength, KAX and kerosene dosages); however, Cu recovery was significantly improved with increasing agitation strength, KAX and kerosene dosages, and thus the separation efficiency was improved.

The pretreatment processes consisting of grinding followed by roasting were investigated to improve the selective leaching of Nd and Dy from neodymium-iron-boron (NdFeB) magnet scraps. The peaks of Nd(OH)3 and Fe were observed in XRD results after grinding with NaOH as the amount of water addition increased to 5 cm3. These results indicate that the components of Nd and Fe in NdFeB magnet could be changed successfully into Nd(OH)3 and Fe, respectively. In the roasting tests using the ground product, with increasing roasting temperature to 500 °C, the peaks of Nd(OH)3 and Fe disappeared while those of Nd2O3 and Fe2O3 were shown. The peaks of NdFeO3 in the sample roasted at 600 °C were observed in the XRD pattern. Consequently, 94.2%, 93.1%, 1.0% of Nd, Dy, Fe were leached at 400 rpm and 90 °C in 1 kmol·m−3 acetic acid solution with 1% pulp density using a sample prepared under the following conditions: 15 in stoichiometric molar ratio of NaOH:Nd, 550 rpm in rotational grinding speed, 5 cm3 in water addition, 30 min in grinding time, 400 °C and 2 h in roasting temperature and time. The results indicate that the selective leaching of Nd and Dy from NdFeB magnet could be achieved successfully by grinding and then roasting treatments.

Electrical cell-substrate impedance sensing is increasingly being used for label-free and real-time monitoring of changes in cell morphology and number during cell growth, drug screening, and differentiation. In this study, we evaluated the feasibility of using ECIS to monitor C2C12 myoblast differentiation using a fabricated indium tin oxide (ITO) electrode-based chip. C2C12 myoblast differentiation on the ITO electrode was validated based on decreases in the mRNA level of MyoD and increases in the mRNA levels of myogenin and myosin heavy chain (MHC). Additionally, MHC expression and morphological changes in myoblasts differentiated on the ITO electrode were comparable to those in cells in the control culture dish. From the monitoring the integration of the resistance change at 21.5 kHz, the cell differentiation was label-free and real-time detectable in 30 h of differentiation (p < 0.05).