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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.

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.

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.

Leaching recovery of rare earth elements (REEs) from two natural kaolinite samples were investigated in this study. Using standard leach conditions without pretreatment, only 10% and 56% of the total REEs were leached from the raw samples, respectively. However, pretreatment by calcination at temperatures between 500 °C and 900 °C increased the leaching recovery of total REEs to values as high as 93% for both samples. Mineralogical characterization and acid leaching results suggested that the decomposition of the kaolinite structure occurs as a function of calcination temperature. Comparisons between the REEs and the major constituent elements of kaolinite showed that their leaching characteristics contrasted. Moreover, sequential chemical extraction results suggested that the fractions of REEs occurring in an ion-adsorption form were negligible. Therefore, the REEs were primarily associated with REE-bearing minerals of which, a portion was encapsulated in kaolinite aggregates while the remaining existed as liberated particles. These association characteristics were proven through SEM-EDX analysis. Based on the mineralogy characterization, acid leaching, and sequential chemical extraction results, it was concluded that the REEs have complicated modes of occurrence and association characteristics with kaolinite. The positive impacts of calcination on the leaching recovery of REEs are likely due to two mechanisms, i.e., (1) liberation of the encapsulated REE-bearing minerals through expansion of the kaolinite aggregates, and (2) conversion of the REE-bearing minerals into more soluble forms.

The majority of rare earth elements (REEs) existing in the feed to coal preparation plants report to the coarse refuse streams which are transported for permanent storage in contained piles. In this study, an integrated flowsheet was developed based on laboratory test data which combines physical separation, pyrite bio-oxidization, heap leaching, selective precipitation and solvent extraction processes. The test data was obtained from (1) characterization of a number of natural leachate and solid samples collected from different preparation plants which process coals from a number of coal seams and (2) laboratory acid leaching and selective precipitation tests results. The highly-valued critical REEs (i.e., Y, Nd, Eu, Tb and Dy) were selectively leached from the refuse samples in the natural environment due to the acid generated by pyrite oxidization. The leachate samples were evaporated to remove water and obtain residual solids (i.e., dissolved solids in the leachates). The total REE content in the dissolved solids from a given leachate sample was 380 ppm, which was higher than the REE content of the coarse refuse material that generated the leachate (322 ppm). Acid leaching tests recovered as much as 80% of the total REEs from the coarse refuse samples using a 1.2 M sulfuric acid solution. Afterwards, the pH of the leachate was increased in a step-wise fashion which resulted in the production of precipitates containing 0.3–1.1% total REEs. A significant amount of contaminants, such as Fe, Al, and Ca, were eliminated in the sequential precipitation process, which allowed further upgrading using oxalic acid precipitation and/or solvent extraction. In the proposed flowsheet, the coarse refuse is arranged in heap leach pads and the acid needed for REE leaching is primarily produced from pyrite bio-oxidization, which enhances the selectivity of REE recovery and significantly reduces the cost. The successful application of the flowsheet would result in significant benefits to both the coal and rare earth industries.

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.

The modes of occurrence associated with rare earth elements (REEs) in bituminous coal sources were evaluated along with their potential recovery by calcination treatment at 600 ℃ followed by sequential extraction. The coal samples were collected from three different resources, i.e., Fire Clay, West Kentucky No. 13, and Illinois No. 6 seams. Sequential extraction tests indicated that the majority of REEs in the calcined samples occurred as metal oxides especially for the West Kentucky No. 13 calcined material (54% of total REEs), which is distinct from untreated coal and combustion byproducts from pulverized coal boilers reported in literature. In addition, heavy REEs were more likely associated with easily dissolvable forms (i.e., ion-exchangeable, carbonates, and metal oxides). The calcined samples were leached under weak acidic conditions with and without adding ammonium sulfate to recover REEs. About 13%, 24%, and 20% of the REEs were extracted from the Fire Clay, West Kentucky No. 13, and Illinois No. 6 materials, respectively, by leaching at pH 4.0. The addition of 1 M ammonium sulfate further increased the recovery values to 18%, 45%, and 32%, respectively. In addition, about 64% of HREEs were extracted from the West Kentucky No. 13 calcined material under this condition. As such, a significant portion of the REEs can be easily extracted from the samples obtained by calcining the coals at 600 ℃.

Getting magnetite concentrate with extremely high Fe grade using reverse flotation method has become increasingly important due to the enhanced requirement of iron making industry. The coagulation in magnetite-concentrate flotation system may impose negative effect on flotation performance. This paper combined the DLVO calculation and microscope observation to study the coagulation behavior in the system. The result shows that the coagulation between fine magnetite and coarse silica particle is very likely to occur with energy barrier 1.5 × 10 −18 J at pH 10.0, which is much smaller than 1.6 × 10 −17 J and 3.3 × 10 −17 J for fine magnetite and fine magnetite-coarse magnetite particles. The coating of fine magnetite particle on silica surface was observed under microscope. Mechanical scrubbing increases the concentrate Fe grade and recovery by 1.5% and 8.5% respectively. Magnetite concentrate of 67.9% Fe grade and 71.5% Fe recovery was finally obtained.

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