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The atomic-level structure and electronic properties of monazite were investigated using a first-principles method based on density functional theory (DFT). First, the geometric structure of monazite was optimized, followed by calculations of its Mulliken population, electron density, and density of states, which were subsequently analyzed. The findings of this analysis suggest that monazite is highly susceptible to cleavage along the {100} plane during crushing and grinding. When SPA was utilized as the collector, the recovery rate of monazite was higher than that when LF-P8 was used. The zeta potential and adsorption energy results indicated that the zeta potential after SPA adsorption tended towards negativity, and the adsorption energy was smaller, indicating that SPA exhibited stronger adsorption performance. LF-P8 was stably adsorbed on the monazite (100) surface via mononuclear double coordination. SPA was stably adsorbed on the surface of monazite (100) via binuclear double coordination. The results of this study provide valuable insights into the adsorption of monazite by commonly used flotation collectors. These findings are of substantial importance for future endeavors in designing flotation collectors capable of achieving selective monazite flotation.

To overcome the difficulty of separating bastnaesite from fluorite through the flotation technique, the present study examined the suitability of sodium alginate (SA) as a depressant in the flotation process. The effect of SA on the flotation separation of bastnaesite and fluorite was evaluated using micro-flotation tests, zeta potential measurements, adsorption density measurements, Fourier infrared spectroscopy, and X-ray photoelectron spectroscopy. The micro-flotation results showed that SA exerted a strongly detrimental effect on fluorite flotation, while slightly affecting bastnaesite flotation. The surface chemistry results revealed that the -COO- and HO- functional groups in SA coordinated with Ca2+ on the fluorite surface, which induced hydrophilicity and hindered adsorption in the subsequent octylhydroxamic acid as a collector. However, the interaction of SA with the bastnaesite surface was marginal and did not affect the anchoring of the collector on the surface of bastnaesite. Based on these results, the present study proposes a possible model for the interaction of SA on the surfaces of the two minerals, laying a foundation for the flotation separation of bastnaesite from fluorite with SA as an environmentally benign depressant.

A novel environmentally friendly one-step decomposition strategy for mixed rare earth concentrate of Bayan Obo in sulfuric acid solution was proposed in this work. In this process, more than 84% of bastnasite and monazite were decomposed in the leaching step at a temperature lower than the boiling point of sulfuric acid solution. So, the dilapidation of sulfuric acid in this current proposed process will be reduced to a large extent. The stability region of rare earth ion in the RE(La, Ce, Nd)-F-P-SO4-H2O system at 170 °C has been proven through Eh-pH diagrams. The factors influencing decomposition of rare earth concentrate in this process were also investigated and the optimal leaching conditions were determined to be a leaching temperature of 170 °C with an ore/acid ratio of 1:5 (g/mL), a sulfuric acid concentrate of 75% and a leaching time of 80 min. The mineralogical changes occurring during the H2SO4 leaching process were investigated by X-ray diffraction and SEM-EDS. The analysis results showed that bastnasite and most of monazite had been decomposed, leaving only a small amount of monazite in the leaching residue.

The aim of this study was to characterize the riebeckite type rare earth ore found in the Bayan Obo deposit, in order to identify the distribution and occurrence of both rare earths and gangue species within the ore. Several analytical techniques were utilized to accomplish this, such as chemical analysis, quantitative XRD, a single polarizing microscope, and a mineral automatic analysis system. The analysis revealed that the primary rare earth minerals (REMs) in the ore were bastnaesite and monazite, with huanghoite, parisite, aeschynite, and fergusonite identified as secondary rare earth minerals. The main gangue species was magnetite, accompanied by smaller quantities of riebeckite and dolomite. The ore was rich in rare earth oxides, with a 3.81 wt% grade. The screen analysis of the pulverized ore indicates that the 43–100 μm fraction is the dominant size, while the finer size fractions below 43 μm contain the bastnaesite and monazite, as well as huanghoite, parisite, aeschynite, and fergusonite. Microstructural characterization showed that the REMs were both coarse-grained and fine-grained, occurring as granular aggregates and fine disseminations within the gangue. Bastnaesite and monazite were the major REMs, with dominant amounts of cerium, lanthanum, praseodymium, and neodymium, while parasite was identified as an impurity. Huanghoite and parisite contained barium and calcium as impurities, respectively. Aeschynite and fergusonite were REMs that included niobium in their composition. Bastnaesite and monazite were found to contain a much higher rare earth content than huanghoite, parisite, aeschynite, and fergusonite. Potential methods for recovering rare earths from this ore, such as magnetic separation and froth flotation, have been identified and may be applicable to similar ferruginous rare earth-bearing ores.

Abstract Collector OA, oleic acid, is widely used industrially for fluorite flotation. Low selectivity, dispersibility and collecting capability of the OA collector are always observed. In this study, compared with flotation of dolomite, a collector mixture of OA and SPE (styrylphenol polyoxyethylene ether) demonstrated significantly better performances for the fluorite. An optimal mass ratio 4 : 1 OA : SPE was found, and the recovery of fluorite was increased from over 85 % to more than 94 % compared with pure OA. Furthermore, the dosage of the collector agent was reduced from 50 mg mL−1 to 20 mg mL−1, which did not negatively impact the recovery of dolomite. The results from the contact angle tests indicated that SPE selectively increased the surface hydrophobicity of fluorite but had little effect on dolomite. Besides, zeta potential measurements and IR analyses revealed that the addition of SPE led to strong chemical adsorption on the surface of fluorite, resulting in a significant difference in the flotation performances of the two minerals. Therefore, SPE‐emulsified OA is corroborated to prompt more selectivity and collecting capability on flotation of fluorite over dolomite.

Cyanide Tailings (CTs) are known as hazardous resources that accumulate hundreds of millions of tons, threatening the ecological environment. This work proposes an eco-friendly and efficient way to recover gold and silver from CTs. The effects of calcium chloride dosage, silicon dioxide dosage, pellet moisture content, roasting temperature, and roasting time on Au and Ag chloridizing volatilization were studied. The kinetics of simultaneously recovering of gold and silver from cyanide Tailings by chlorination roasting was investigated. It was determined that the chloridizing volatilization rates increased with increasing calcium chloride dosage, temperature and decreasing silicon dioxide dosage, pellet moisture content. The chloridizing volatilization kinetics followed a shrinking core model, with inter-diffusion through gangue layer as the rate determining step. This finding is in accordance with the apparent activation energy (Ea) of 24.01 kJ·mol-1 (Au) and 24.62 kJ·mol-1 (Ag). The orders of reaction with respect to moisture content, temperature, calcium chloride dosage and silicon dioxide dosage were also achieved. The rate of reaction based on diffusion-controlled process can be expressed by semi-empirical equations. The control steps of the gold and silver recovery process were investigated by the kinetic study, which provided theoretical guidance for the optimization of the method. [ABSTRACT FROM AUTHOR]

Abstract Low‐intensity magnetic separation tailings (LIMS tailings) are a common by‐product obtained after magnetite magnetic separation. In this article, various techniques such as chemical analysis, X‐ray diffraction, ICP‐MS, and Mineral Liberation Analysis (MLA) were employed to investigate the LIMS tailings. The primary iron‐bearing mineral identified was hematite and rare earth minerals were monazite and bastnaesite. The main gangue species was fluorite with small amounts of dolomite and amphibole. Due to the weak magnetism of hematite and rare earth minerals, magnetic separation has low efficiency. However, magnetization roasting‐magnetic separation is an effective method to recover hematite. The present study focuses on the separation of iron and rare earth from LIMS tailings through magnetization roasting‐magnetic separation. The results demonstrate that with a roasting temperature of 650 °C, a roasting time of 60 min, a slurry concentration solid‐liquid ratio of 25 : 1, a rough magnetic field intensity of 0.16 T, and a selected magnetic field intensity of 0.10 T, the iron grade in the magnetic concentrate increases to 65.49 % and an iron recovery rate of 65.16 % can be achieved. The XRD patterns of magnetic separation concentrate show that the main mineral phases in concentrate are magnetite (Fe3O4) and fluorite (CaF2), which can be removed by grinding and reverse flotation fluorite to obtain a high‐grade iron concentrate. The REO grade of magnetic separation tailings is 11.98 %, and its recovery rate is 97.96 %. Consequently, rare earth can be effectively extracted and separated after the subsequent flotation‐leaching process.