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Solution equilibrium calculations were performed in this study to understand the impact of contaminant metal ions on the precipitation efficiency of selected rare earth elements (Ce3+, Nd3+, and Y3+) using oxalic acid as a precipitant. Trivalent metal ions, Al3+ and Fe3+, were found to considerably affect the precipitation efficiency of REEs. When Al3+ and Fe3+ concentrations were increased by 1 × 10−4 mol/L, in order to achieve an acceptable cerium recovery of 93% from solutions containing 1 × 10−4 mol/L Ce3+, oxalate dosage needed to increase by 1.2 × 10−4 and 1.68 × 10−4 mol/L, respectively. Such great impacts on the required oxalate dosage were also observed for Nd3+ and Y3+, which indicates that oxalic acid consumption and cost will be largely increased when the trivalent metal ions exist in REE-concentrated solutions. Effects of the divalent metal ions on the oxalate dosage is minimal. Furthermore, solution equilibrium calculation results showed that the precipitation of Fe3+ and Ca2+ (e.g., hematite and Ca(C2O4)∙H2O(s)) likely occurs during the oxalate precipitation of REEs at relatively high pH (e.g., pH 2.5), which will reduce rare earth oxalate product purity. In addition to the metal ions, anionic species, especially SO42−, were also found to negatively affect the precipitation recovery of REEs. For example, when 0.1 mol/L SO42− occurs in a solution containing 1 × 10−4 mol/L Ce3+ and 4 × 10−4 mol/L oxalate, the pH needed to be elevated from 2.0 to 3.3 to achieve the acceptable recovery. Overall, findings from this study provide guidance for the obtainment of high-purity rare earth products from solutions containing a considerable amount of contaminant metal ions by means of oxalic acid precipitation. Accepted version

A kaolinite containing 245 ppm of REEs was used as a feedstock for REE recovery. Acid leaching test results showed that less than 20% of the REEs were leached using 1.2 mol/L HCl. However, after thermal treatment at 600 °C, a recovery of over 80% was obtained with a much lower acid concentration (0.01 mol/L). Meanwhile, only 3.91% Al, 2.34% Fe, 8.63% K, 9.32% Mg, 0.08% Ti, and 5.34% Ca were simultaneously extracted. Therefore, high-purity rare earth products are likely produced from the kaolinite. To achieve a more comprehensive utilization of the kaolinite, the residual solid after REE recovery was subjected to a selective leaching process to prepare porous silica. Comparisons among three different types of mineral acids showed that H2SO4 provided the best performance. The effects of different factors were investigated using the response surface methodology. A porous silica with as high as 336.95 m2/g specific surface area was ultimately produced.

Due to the increasing criticality of rare earth elements (REEs), it has become essential to recover REEs from alternative resources. In this study, systematic REEs leaching tests were performed on the calcination product of a coal coarse refuse using hydrochloric acid and different types of organic acid as lixiviants. Experimental results show that the recovery of REEs, especially heavy REEs (HREEs) and scandium (Sc), is improved by using selected organic acids. Citric acid and dl -malic acid afford the best leaching performances; whereas, malonic acid, oxalic acid, and dl -tartaric acid are inferior to hydrochloric acid. Results of zeta potential measurements and solution chemical equilibrium calculations show that malonic acid is more likely adsorbed on the surface of the calcined material compared with citric acid and dl -malic acid. The adsorption may reduce the effective concentration of malonic species in solution and/or increase the amount of REEs adsorbed on the surface, thereby impairing the leaching recovery. Compared with light REEs (LREEs), a stronger adsorption of the HREEs on the surface is observed from electro-kinetic test results. This finding explains why organic acids impose a more positive impact on the leaching recovery of HREEs. By complexing with the HREEs, organic acids can keep the metal ions in solution and improve the leaching recovery. The adsorption of Sc3+ on the surface is the lowest compared with other REEs. Therefore, rather than complexing, the organic anionic species likely play a function of solubilizing Sc from the solid, which is similar to that of hydrogen ions.

It is critical to separate carbon from fly ash due to its increasing amount being produced per year and component requirements for secondary utilizations. Flotation is an efficient technique to fulfill this task. Effects of diesel emulsification using sodium petroleum sulfonate on unburned carbon flotation from a fly ash sample were studied. Flotation tests showed that carbon recovery was increased from 41% to 86% using the emulsified collector at the same dosage level (0.5 kg/ton) as diesel. Zeta potential measurements indicated that both the carbon and emulsified collector droplets carried negative charges within the studied pH range. For fly ash particles with high carbon content, adsorption bands of carbonyl bond (C O) were detected in the FTIR spectra, which agreed with the XPS characterization findings. XPS analysis indicated that the emulsified diesel (ED) collector interacted more intensively with the surface carbon-oxygen groups relative to the normal diesel collector. Finally, the improved efficiency of ED is proved to be related to its polar groups and reduced oil droplet size. The polar groups of ED (mainly sulfonate) interacted with the surface oxygen containing groups (e.g., C O and COOH) through hydrogen bonding and Lewis acid base interaction.

The pretreatment of middlings and coarse refuse material collected from a Pocahontas No.3 coal source using calcination was investigated for the potential of improving the leaching recovery of rare earth elements. Calcination at 600 °C significantly and preferentially improved light REE (LREE) recovery to values in the range of 80–90% using 1.2 M HCl. The positive effect was due to the increased pore diameter and volume as well as thermal decomposition of the minerals associated with the LREEs. Heavy REE recovery was not increased after calcination at 600 °C due to the need for higher temperatures to achieve thermal decomposition of HREE minerals. Furthermore, some HREEs were also associated with sulfide minerals (e.g., pyrite) which were transformed into low-solubility oxides (e.g., hematite) after roasting at 600 °C. Based on the leaching characteristics of the major elements (Fe, Al, Ca and Mg) and REEs along with SEM-EDX data, it was concluded that REEs in the two samples mainly occurred as phosphate minerals. In the middlings, some of the minerals were completely encapsulated by kaolinite, which was less apparent in the coarse refuse. In addition, it was found that a portion of the REEs (10–25%) in the samples were associated with the organic matter and micro-dispersed minerals in the organic matrix, was released after calcination and easily recovered using 0.1 M (NH4)2SO4 solution at pH 5. Scandium primarily occurred in mineral forms with minimum amounts associated with ion-adsorbed clays and organic association.

In this study, 2-phosphonobutane-1,2,4-tricarboxylic acid (PBTCA) was introduced as a depressant for the flotation separation of apatite from calcite. The micro-flotation test results showed that good floatability of both apatite and calcite were obtained at around pH 8.5 when using sodium oleate (NaOl) as collector. Mixed minerals flotation test results showed that an efficient separation of apatite from calcite was achieved using PBTCA as depressant, which presented a selective depression for the calcite. Zeta potential measurements indicated that the addition of PBTCA prior to NaOl did not interfere the interaction between NaOl and apatite while it decreased NaOl adsorption on calcite surface.

The effect of depressant 2-phosphonobutane-1,2,4-tricarboxylic acid (PBTCA) on the flotation separation of smithsonite from calcite was studied by micro-flotation and electro-kinetic tests. The micro-flotation results showed that good floatability of both smithsonite and calcite were achieved in the presence of sodium oleate. The depressant PBTCA selectively depressed calcite without interfering the floatability of smithsonite. The mixed minerals flotation results showed that efficiency separation of smithsonite from calcite was achieved using PBTCA as the depressant. Zeta potential measurements indicated that NaOl was able to adsorb on the PBTCA pre-treated smithsonite surface, but failed to adsorb on the PBTCA pre-treated calcite surface.

High-temperature pretreatment of coal-based mineral matter in an oxidizing environment significantly enhances the leaching characteristics of rare earth elements (REEs). A research study has found that the temperatures used in fluidized bed combustion (FBC) of coal to produce electricity are near optimum for pretreating the associated mineral matter prior to leaching to maximize the recovery of critical materials. Tests were performed on representative samples collected from preparation plants treating West Kentucky No. 13, Illinois No. 6, and Fire Clay coal seam sources as well as fly ash and bed ash samples from two FBC power plants. Acid leaching tests using 1.2 M HCl at 75 °C were performed on both the coal and the FBC ash samples. Prior to leaching, the coal samples were pretreated at temperatures of 600, 750, and 900 °C in an oxidizing environment to study the effect on leaching characteristics. The results showed that pretreatment at 600 °C for 2 h resulted in a significant increase in REE recovery...

Despite that the negative effects of lizardite on pentlandite flotation have been realized, the lizardite effects under different slurry pH have not been sufficiently investigated. Micro-flotation test results showed that pentlandite recovery gradually decreased with the increase in slurry pH when lizardite slime existed. Both electro-kinetic tests and turbidity measurements indicated that hetero-coagulation between the pentlandite and lizardite particles occurred in the entire pH range, which failed to explain the lower recovery at high slurry pH. Zeta potential distribution measurements and sedimentation tests suggested that self-coagulation among lizardite particles occurred at higher pH, which increased the amount of lizardite slime coated on the pentlandite surfaces and thus reducing collector adsorption and flotation recovery. The particle interaction in the pentlandite/lizardite flotation system was investigated using DLVO theory and microscope observation. The findings of the study will contribute to the improvement in pentlandite flotation by regulating the particle interactions.

Effects of sodium chloride on the sulfidization flotation of cerussite (PbCO3) were investigated by micro-flotation tests, adsorption tests and zeta potential measurements. The micro-flotation test results indicated that sodium chloride effectively improved the sulfidization flotation performance of cerussite. The electrical double layers of cerussite particles were compressed in sodium chloride solution, thus providing more lead sites on the cerussite surface and promoting the adsorption of sulfide ions. In sodium chloride solution, more lead sulfide species were formed on the cerussite surfaces and improved its sulfidization flotation performance.

A novel method to improve carboxymethyl cellulose (CMC) performance in talc flotation is described in this paper. Micro-flotation results showed that using CMC as the depressant, strong depression on talc floatability was observed when the pulp pH was adjusted from low (4.0) to high (8.5), which improved the separation performance between chalcopyrite and talc. Zeta potential measurements illustrated that talc surface charge was less negative when the pH was regulated from acid to base than the situation that the pulp pH was altered from base to acid. ICP test and XPS results indicated that more magnesium ions were dissolved from the talc lattices at pH 4 and the hydrolyzed species of magnesium cations were adsorbed onto the talc surfaces when the pulp pH was adjusted to 8.5, and thus promoting the CMC adsorption on the talc surfaces, which further decreased its floatability.

Selective depression of calcite in a monazite-calcite flotation system was systematically studied using micro-flotation and electrokinetic tests as well as X-ray photoelectron spectroscopy (XPS) characterization and solution chemistry calculations. Monazite was found to be depressed by both sodium silicate and sodium hexametaphosphate when calcium ions appeared in solution. It was hypothesized that co-adsorption of calcium species (CaHCO3+ and CaCO3(aq)) and Si ( OH ) 4 as well as SiO ( OH ) 3 - via hydrogen bonding formed a compact layer over the monazite surface, thereby reducing octanohydroxamic acid adsorption by steric hindrance. The use of citric acid or EDTA formed soluble chelates (e.g. CaCit− and CaEdta2−) with the calcium ions which negated the flotation suppression effect. An XPS study indicated that both EDTA and citric acid chemically adsorbed on the monazite surface, while the adsorption on the calcite surface was achieved mainly through hydrogen bonding. EDTA was more effective than citric acid due to two reasons: (i) citric acid itself presented stronger depression effects and (ii) citric acid has weaker chelating properties for calcium. Using 2.5 × 10−4 M octanohydroxamic acid, 0.05 g/L sodium silicate and 6 × 10−5 M EDTA, monazite recovery of more than 90% was achieved while recovering only 20% of the calcite.

In this investigation, the fine-grained bauxite ore flotation was conducted in a plate-packed flotation column. This paper evaluated the effects of packing-plates on recovering fine bauxite particles and revealed the fundamental mechanisms. Bubble coalescence and break-up behaviors in the packed and unpacked flotation columns were characterized by combining Computational Fluid Dynamics (CFD) and Population Balance Model (PBM) techniques. Flotation experiments showed that packing-plates in the collection zone of a column can improve bauxite flotation performance and increase the smaller bauxite particles recovery. Using packing-plates, the recovery of Al2O3 increased by 2.11%, and the grade of Al2O3 increased by 1.85%. The fraction of &minus, 20 &mu, m mineral particles in concentrate increased from 47.31% to 54.79%. CFD simulation results indicated that the packing-plates optimized the bubble distribution characteristics and increased the proportion of microbubbles in the flotation column, which contributed to improving the capture probability of fine bauxite particles.

An innovative self-designed medium was packed in a bench-scale flotation column to study its influence on the flotation recovery of bauxite. Computational fluid dynamics (CFD) simulation was conducted to reveal the impact of the packing medium on the turbulent characteristics of collection zone in the column. Simulation results show that multilayer packing of the medium divides the collection zone into small units having different turbulent intensities, which is more suitable for flotation separation. The packing medium decreases the turbulence kinetic energy (from 1.08 &times, 10&minus, 2 m2/s2 to 2.1 &times, 3 m2/s2), turbulence eddy dissipation (from 3.71 &times, 2 m2/s3 to 9.8 &times, 3 m2/s3) and axial fluid velocity of fluid in the column. With three layers of packing, the recovery of Al2O3 increased by 2.11% and the aluminum to silicon content ratio of the concentrate improved from 5.16 to 9.72.

Many studies have been published in recent years focusing on the recovery of rare earth elements (REEs) from coal-related materials, including coal, coal refuse, coal mine drainage, and coal combustion byproducts particularly fly ash. The scientific basis and technology development have been supported by coal geologists and extractive metallurgists, and through these efforts, the concept has progressed from feasibility assessment to pilot-scale production over the last five years. Physical beneficiation, acid leaching, ion-exchange leaching, bio-leaching, thermal treatment, alkali treatment, solvent extraction, and other recovery technologies have been evaluated with varying degrees of success depending on the feedstock properties. In general, physical beneficiation can be a suitable low-cost option for preliminary upgrading; however, most studies showed exceedingly low recovery values unless ultrafine grinding was first performed. This finding is largely attributed to the combination of small RE-bearing mineral particle size and complex REE mineralogy in coal-based resources. Alternatively, direct chemical extraction by acid was able to produce moderate recovery values, and the inclusion of leaching additives, alkaline pretreatment, and/or thermal pretreatment considerably improved the process performance. The studies reviewed in this article revealed two major pilot plants where these processes have been successfully deployed along with suitable solution purification technologies to continuously produce high-grade mixed rare earth products (as high as +95%) from coal-based resources. This article presents a systematic review of the recovery methods, testing outcomes, and separation mechanisms that are involved in REE extraction from coal-related materials. The most recent findings regarding the modes of occurrence of REEs in coal-related materials are also included.

In this paper, the effect of sodium alginate (SA) on the flotation separation of molybdenite (MoS2) from chalcopyrite using kerosene as collector was systematically investigated. The results of single-mineral micro-flotation tests indicated that SA exhibited strong depression on chalcopyrite flotation while it imposed no impact on the floatability of molybdenite. However, in the chalcopyrite–molybdenite mixed-mineral flotation system, the presence of chalcopyrite significantly increased the depressing effect of SA on molybdenite flotation, leading to a considerable reduction in the flotation selectivity. The negative impact of chalcopyrite on the performance of SA in molybdenite flotation was eliminated by adding a certain dosage of kerosene prior to SA. A concentrate containing 53.43% of molybdenum (Mo) was obtained at 76.90% of recovery using 19 mg/L kerosene and 40 mg/L SA at pH 5.4. Zeta potential measurements indicated that the adsorption of SA on chalcopyrite surfaces was stronger than that on molybdenite surfaces, which agreed with the single-mineral flotation test results. The adsorption of SA on chalcopyrite was further confirmed to be chemisorption by Fourier-transform infrared spectroscopy (FTIR) spectra analyses. When Cu2+ appeared in solution, the flotation of molybdenite was strongly depressed by SA. Mechanism analyses indicated that more active sites were generated on molybdenite surfaces after the addition of Cu2+, thus promoting the adsorption of SA. Published version

The flotation separation of scheelite from calcite using sodium humate (SH) as a potential depressant has been studied through micro-flotation experiments, zeta potential and Fourier transform infrared (FT-IR) spectra measurements. The flotation results indicate SH exhibited selective depression effect on calcite in the presence of sodium oleate, achieving the flotation separation of scheelite from calcite. Zeta potential results indicate that SH strongly absorbed on calcite surface but negligibly on scheelite surface, which prevents the adsorption of sodium oleate on calcite surface. The FT-IR results also show that SH can be adsorb on the surface of calcite. Results shown a higher Ca density on the calcite than on scheelite surface, and that calcite can accommodate the interaction of SH.

Rare earth elements (REEs) found in coal are in the form of minerals, ion-adsorbed onto clay surfaces or inner layers, or organically bound. Rare earth minerals such as monazite exist in coal and have grain sizes smaller than 5 µm. In this study, froth flotation was successful in concentrating rare earth minerals existing in a thickener underflow material derived from Fire Clay seam coal that contained around 431 ppm of total rare earth elements (TREE) on a dry ash basis. Conditioning with fatty acid followed by processing using multiple stages of conventional flotation produced a final concentrate containing 2,300 ppm TREE. Using a laboratory flotation column to limit hydraulic entrainment, the TREE content was further enriched to around 4,700 ppm, which equated to an enrichment ratio of 10:1..

The surface charge mechanisms of monazite were studied using electrokinetic tests, solution equilibrium calculations, crystal structure characterization, and electrostatic model predictions. The isoelectric point (IEP) of a natural monazite sample was found to occur at pH 6.0. To allow a theoretical study of the surface charge on monazite, fundamental values associated with cerium monazite (CePO 4 ) were utilized to quantify the isoelectric point (IEP) and point of zero charge (PZC) using solution equilibrium calculations and electrostatic models. The IEP and PZC of the cerium monazite were determined to be equal with a pH value of 7.2. Hydroxyl and hydrogen were found to be potential determining ions and preferential hydrolysis of the lattice ions played a minor role. The findings support the hypothesis that the charge of monazite surfaces in water is a result of protonation/deprotonation reactions. The difference in the IEP values obtained for the natural monazite and cerium monazite along with the wide range of values reported in literature may be partially explained by the amount of carbon dioxide dissolved in solution. When equilibrating with carbon dioxide at 10 –3.5 atm partial pressure, the IEP of the cerium monazite shifted to a solution pH value of 4.5.

The adsorption mechanism of octanohydroxamic acid (OHA) on monazite was studied using kinetic, isotherm, and thermodynamic adsorption tests as well as FTIR, titration, and micro-flotation experiments. The adsorption mechanism was described as a chemisorption/surface precipitation process. At low OHA concentrations, adsorption occurred by chemisorption as a result of the reaction between surface active sites and OHA molecules. With an increase in concentration and interaction time, surface precipitation occurred. For chemisorption, adsorption was an exothermic and entropy driven process and maximum adsorption was achieved at pH 9.0 due to more active sites. In an acidic environment, such as pH 3.0, adsorption was achieved via chemisorption and hydrophobic bonding. However, a strong basic and higher temperature environment contributes to surface precipitation of basic rare earth hydroxamate. FTIR tests showed the movement of CH 2 band position from 2924 cm − 1 to 2920 cm − 1 with increases in pH values from 3.0 to 6.0, 9.0, and 11.0, which corresponds to the status of chemisorbed and surface precipitated OHA, respectively.

The dissolution of minerals provides elevated concentrations of alkaline earth metal ions such as calcium (Ca 2+ ) in water which may have a significant influence on flotation performance. The effects of Ca 2+ ions dissolved from calcite surfaces on the flotation performance of a monazite-calcite system have been investigated using micro-flotation tests followed by fundamental collector adsorption studies. The results indicate that specially adsorbed CaOH + via hydrogen bonding competed with hydroxamic acid for the phosphate-oxygen active sites on the monazite surface and decreased its hydrophobicity due to hydration. Electrostatically adsorbed CaOH + on monazite increased the adsorption density of hydroxamic acid through a chelating interaction. At higher dosages of calcium and hydroxamic acid, bulk and surface precipitation of calcium-hydroxamate occurs in the monazite-calcium-hydroxamic acid system.

Acid mine leachate (AML) may be a significant source of rare earth and other critical elements needed for the manufacturing of advanced electronics and renewable energy technologies. A systematic study was performed on a natural leachate collected from a coal preparation plant that treated bituminous coal containing elevated amounts of pyrite. The leachate contained 1.96 ppm total REEs, 2.52 ppm Zn, 2.15 ppm Ni, 1.22 ppm Cu, 0.77 ppm Co, and 25 ppm Mn. Sequential precipitation using simple pH control generated a pre-concentrate containing 0.82% total REEs, 1.08% Zn, 0.91% Ni, 0.50% Cu, 0.34% Co, 7.1% Mn. The pre-concentrate was further treated by re-dissolution using a 10 M HNO3 solution, which resulted in a solution pH of 1.5. The addition of oxalic acid to selectively precipitate the REEs followed by roasting of the precipitate generated a product containing >98% rare earth oxides. More than 95% of the Cu and Zn were recovered from the residual liquid using Na2S at pH 2 and 3, respectively. The CuS and ZnS contents in the corresponding concentrates were 60% and 58%, respectively. The majority of the Co and Ni remained in the residual liquid. Based on these findings, a process flowsheet was developed to recover the REEs and critical metals from AML.

Three segments of a core sample collected from the Guxu coalfield (Sichuan Province, China) were evaluated to determine the rare earth element (REE) leachability and mineralogy. Elemental analysis of the three samples showed that the roof and floor strata contained as high as 2087 ppm of total REEs, a value much higher than other coal-based materials reported in the literature. For the floor samples, 47% to 65% of the light REEs (LREEs) were leached within 5 min of contact time using 1 M mineral acid, and prolonged contact, up to 120 min, increased the LREE recovery to as high as 75%. For the roof sample, only 25% of the LREEs were leached at extended contact times. Unlike the LREEs, heavy REE (HREE) recovery from both the roof and floor materials was less than 25%. The relatively high REE content of the feed materials facilitated detailed SEM-EDX and TEM-EDX analyses of the samples. These studies confirmed the presence of several rare earth carbonate minerals, including bastnaesite and parasite, as well as rare earth phosphate minerals such as florencite and apatite. Given the prevalence and rare earth content of the particles analyzed via SEM-EDX, the data suggests that the LREEs present in the floor and roof samples were primarily associated with carbonates and phosphates, respectively. Further SEM-EDX analysis of the leaching residues showed that florencite was the only LREE-bearing species remaining after prolonged acid leaching, indicating that florencite is less soluble relative to the other rare earth minerals. The leaching results corroborate the SEM-EDX and TEM-EDX findings, while providing critical fundamental information on the relationship between REE content, REE mineralogy, and REE leachability in coal refuse materials.

Lithium recovery from the density fractions of Baker (namely West Kentucky No. 13) seam coal was investigated in this study. Proximate and elemental analyses showed that lithium contents in the 1.8–2.2 SG and 2.2 SG sink fractions were 185 ppm and 150 ppm, respectively, which are significantly higher than the average content of coal sources (12 ppm) and coal ashes (66 ppm) worldwide. Moreover, due to the high lithium contents and mass distribution, nearly 90% of the lithium present in the Baker coal sample was distributed in the two density fractions. Direct leaching using a HCl solution resulted in less than 10% lithium recovery, which was due to nearly 90% of the lithium being associated with insoluble solids as determined by the results from sequential extraction tests. Calcination of both density fractions under 600 °C for two hours followed by leaching resulted in recovery increases in the range of 70% to 80%. The pyro-metallurgical pretreatment step converted most of the associated lithium minerals to more easy-to-leach forms such as carbonate and metal oxide. Based on mineralogy characterization and leaching test results, it was concluded that the positive effects of calcination on lithium leachability resulted from the dehydration and disintegration of kaolinite as well as dehydroxylation and expansion of muscovite/illite. An analysis of the leaching kinetics revealed that the leaching rate of lithium was controlled by interface transfer and diffusion across the product layer and was negatively impacted when the calcination temperature exceeded 600 °C due to sintering of the kaolinite.

The impacts of calcination on the leaching recovery of rare earth elements (REEs) and other critical metals from a Baker seam coal refuse material were investigated in this study. Acid leaching test results showed that leaching recovery of the REEs, especially the light REEs, was increased significantly from around 24% to 80–90% after calcination under 600–750 °C for two hours. By considering the major element leaching kinetics and the mineralogical changes of the dominant minerals caused by calcination, two mechanisms were proposed for the improved leaching performance: (1) decomposition of difficult-to-dissolve REE-bearing minerals into soluble forms; and (2) liberation of REE-bearing minerals encapsulated in clays after calcination due to the dehydration and disintegration of the layered clay structure. Sequential chemical extraction tests showed that a significant fraction (around 75%) of the REEs in the raw material were present in an insoluble form. After calcination under proper temperatures (600–800 °C), the majority of the insoluble light REEs (LREEs) were converted into soluble forms. Changes in the modes of occurrence of the LREEs and phosphorous based on leaching rate and sequential leaching correlated well as a function of the calcination temperature. This finding in addition to SEM-EDX results tends to indicate the potential that the LREEs occur primarily as rare earth phosphates such as monazite in the untreated coal refuse material. The acid leaching and sequential chemical extraction results show that co-extraction of REEs and other critical metals such as Co, Mn, and Li is possible using a calcination-acid leaching procedure.

Physicochemical Problems of Mineral Processing; ISSN 2084-4735

Flotation separation of scheelite from calcite is difficult due to the similarities in their surface properties. In this work, the flotation behavior of scheelite and calcite using oxidized paraffin soap(OPS), benzohydroxamic acid(BHA) and the mixed OPS/BHA collectors was investigated through micro-flotation experiments. The flotation results of single mineral experiments demonstrated a higher selectivity for the flotation of scheelite from calcite at pH 9 than individual OPS and BHA when using water glass as depressant. In order to probe the validity of the findings, mixed binary minerals experiments, contact angle and zeta potential experiments were also carried out successfully. Accepted version

The influence of conditioning on the flotation of pyrrhotite in the presence of chlorite was investigated through flotation tests, sedimentation tests, and X-ray photoelectron spectroscopy (XPS) analysis. The flotation results show that chlorite slimes dramatically impair the flotation of pyrrhotite. Sedimentation and flotation tests reveal that conditioning can effectively remove chlorite slimes from pyrrhotite surfaces, resulting in an enhanced flotation recovery of pyrrhotite. When mixed minerals were conditioned under the natural atmosphere, a faster conditioning speed and longer conditioning time decreased the flotation recovery of pyrrhotite. However, when mixed minerals were conditioned under a nitrogen atmosphere, a more intensive conditioning process provided better flotation results. XPS analyses illustrate that a faster conditioning speed and longer conditioning time under the natural atmosphere accelerates the oxidation of pyrrhotite, leading to a decrease in the flotation recovery of pyrrhotite. Published version

The activation mechanism of Pb(II) on the flotation of wolframite was systematically investigated. Micro-flotation test results showed that Pb(II) effectively improved the flotation of wolframite in the presence of benzyl hydroxamic acid (BHA). Maximum floatability of wolframite was achieved at approximately pH 8.5. Adsorption of the hydrolyzed species of Pb(II) occurred at this pH value, which provided more activated sites on the wolframite surface. As such, BHA-Pb complexes were formed on the wolframite surface, which increased the BHA adsorption. In addition, BHA-Pb complexes were also formed in solution, which replaced the Fe and interacted with the Mn on the wolframite surface. These findings were proved by the results of zeta-potential measurements, adsorption measurements, solution chemistry calculations, Fourier transform infrared spectroscopy (FTIR) analysis, and X-ray photoelectron spectroscopy (XPS) characterization. Accepted version