The hydration mechanism between the leaching agent and ore surface during the leaching process of ionic rare earth ore is complicated, and the inter-particle bridge cementation is prone to fracture due to the existence of multiple forces and dispersion during ion adsorption and exchange, resulting in migration and rearrangement of microfine particles, and precipitation at the pore throat, producing blockage phenomenon and affecting the leaching efficiency of ionic rare earths. In order to reveal the migration law of microfine particles during in situ leaching of ionic rare earth ores and to find suitable regulation methods, this paper investigates the effects of leaching agent mass concentration, viscosity, flow rate, hydraulic gradient, ore body height, and ore body water content on the migration of microfine particles. We compared ionic rare earth ores as raw ores and rare earth ores with particle sizes ranging from 0.075 to 0.09 mm using the laboratory column leaching method. The results showed that the migration of microfine particles during ionic rare earth ore leaching was an important factor affecting leaching efficiency. Under the action of external forces, the microfine particles tended to migrate with the leaching agent during the leaching process.

The characteristics of Baiyun Obo niobium-bearing minerals are complex physicochemical properties that make the beneficiation of niobium minerals extremely difficult. In this paper, X-ray diffraction, X-ray fluorescence and mineral liberation analyzer (MLA) systems were used to study the niobium occurrence state and distribution of niobium-bearing minerals in the samples from Baiyun Obo. The results show that the chemical and mineral compositions of the sample are complex, with a Nb₂O₅ grading of 0.24%. There are many kinds of niobium minerals, including ilmenorutile, nioboaeschynite-Nd, baotite, latrappite, euxenite-Y, fergusonite and columbite-Mn, and the highest mass fraction of 0.55% is achieved with Nb in nioboaeschynite-Nd, followed by the mass fraction of ilmenorutile (0.33%). All of the niobium-containing minerals demonstrate a low degree of dissociation. Flotation experiments explored the optimal flotation conditions for HOBA (1-hydroxyoctyl-1,1-bisphosphonic acid) as a flotation collector for Baiyun Obo niobium minerals, which is able to increase the grade of Nb₂O₅ in the concentrate to 1.31%. The optimal use conditions of the reagent are pH 3.5–4.5, and the amount of the collector is 1000 g/t. By further optimizing the beneficiation process and reagent system, ilmenorutile and nioboaeschynite-Nd were significantly enriched in the concentrate, which suggested that HOBA can efficiently increase the grade of Nb₂O₅ in the concentrate.

As Bayan Obo rare-earth tailings, which are generated after the production of mineral products with the raw ore from different mining areas, are considered secondary resources rich in valuable elements such as F, Fe, REE, and Nb, an effective method is urgently needed to recover such valuable elements for resource recycling and environmental conservation. A mineralogical analysis can enable process diagnosis, design, and optimization and is the key to comprehensively utilizing valuable elements. Hence, detailed mineralogical characterization is necessary as a starting point to develop a feasible processing flowsheet. In this study, various detection methods, namely inductively coupled plasma-optical emission spectrometry (ICP), X-ray Fluorescence Spectrometer (XRF), X-ray powder diffractometer (XRD), scanning electron microscopy system with an energy dispersive spectrometer (SEM-EDS), mineral liberation analysis (MLA), and electron probe microanalysis (EPMA), were applied to conduct detailed mineralogical characterization of Bayan Obo rare-earth tailings, and the occurrence state of Sc in the main Sc-bearing minerals was studied using density functional theory (DFT). The results showed that Fe mainly occurs in hematite, riebeckite, ankerite, siderite, and pyrite, with contents of 50.15 wt%, 27.94 wt%, 8.34 wt%, 4.92 wt%, and 5.59 wt%, respectively. Nearly all F occurs in 26.8 wt% fluorite. The main rare-earth minerals are bastnasite, apatite, and monazite (La), with contents of 5.0%, 5.0%, and 1.6% in Bayan Obo rare-earth tailings, respectively. Notably, 48.47%, 21.70%, 10.34%, and 10.28% of niobium element occurs in nioboaeschynite, pyrochlore, dingdaohengite, and ilmenorutile, respectively. Scandium was detected in five minerals, namely aegirine, riebeckite, monazite, ilmenorutile, and niobite, with average contents of 0.04 wt%, 0.22 wt%, 0.06 wt%, 0.06 wt%, and 1.58 wt%, respectively. According to the DFT analysis, the state of Sc in aegirine is different from that in riebeckite. Scandium in aegirine mainly substitutes Fe or enters the interstitial lattice site, while Sc in riebeckite tends to replace Fe. Based on these results, a process for recovering valuable elements from tailings is proposed.

The lead–zinc ore is a typical low-grade complex polymetallic sulphide ore. Its typical mineralogy is characterized by a high content of cryptocrystalline graphite, low content of lead, zinc, and copper, and fine grain size. However, the aggregation size of sulfide mineral aggregations is coarse, and the symbiotic relationship is relatively simple. Conventional process treatment requires the fine grinding of ore into mineral monomer dissociations followed by preferential flotation, but this method has a poor sorting index and high production cost. In this paper, the grain size of the sulfide mineral aggregates determined by mineralogical studies is used to determine the fineness of the grinding, so that the liberation degree of the sulfide ore in the coarse grinding product reaches 70%, and each flotation process is used to collect rough concentrate. In the first step of flotation, the carbon in lead–zinc sulfide ore was removed by adding #2 oil to the flotation tank at a dosage of 40 g/t. In the second flotation process, the pH was adjusted to 7.2, the dosage of isopropyl xanthate was 20 g/t, the dosage of #2 oil was 30 g/t; the flotation reagent in the third step was isopropyl xanthate, and the dosage was 7.5 g/t, with a #2 oil of 15 g/t. Each flotation process concentrate is collected. Finally, the grades of lead and zinc in the final concentrate were 1.6% and 5.71%, respectively. In addition, the recoveries of lead and zinc were 91.78% and 92.07%, respectively. The yield of tailings was 50.6%. By the flotation of sulfide aggregates, a large number of gangues are avoided to participate in fine grinding, which helps to reduce the energy consumption of the mill.

Production of rare earth enrichment from the in situ leaching solution of ion-adsorbed rare earth ores greatly decreases the treatment scale and significantly reduces production energy consumption and cost. However, the generated rare earth enrichment has a high content of impurities. Further purification of rare earth is necessary. Therefore, a process comprised of sulfuric acid leaching and removing aluminum by the neutralization precipitation method to obtain the purified rare earth solution was proposed. The results of acid leaching revealed that at a sulfuric acid concentration of 1.75 mol/L, a temperature of 60 °C, a liquid–solid ratio of 5:1 mL/g, a leaching time of 0.5 h, and a stirring rate of 300 r/min, leaching efficiency of rare earth and magnesium reached 99.11% and 97.39%, respectively, while the leaching efficiencies of aluminum and silicon reached 72.91% and 55.26%, respectively. The comparison of different precipitants during the neutralization precipitation process showed that MgO was the best precipitant for the efficient removal of aluminum and low loss of rare earth. The results of removing aluminum revealed that when the final pH of the rare earth leaching solution was controlled to be 4.7, the reaction temperature was 25 °C, the slurry concentration of MgO was 0.3 mol/L, and the feeding rate of the MgO slurry was 0.5 mL/min, the removal rate of aluminum was 99.49%, and the loss rate of rare earth was 13.93%. The obtained purified rare earth solution contained 19 g/L of rare earth and less than 0.01 g/L of aluminum. Kinetic studies showed that the apparent activation energy of the aluminum removal process was 6.8 kJ/mol, indicating that the precipitation process was controlled by a mass transfer diffusion reaction.

Fine particle content significantly affects the in situ leaching of ion-adsorption-type rare earth ores. This study investigated the effect of fine particle content on solution flow and mass transfer in leaching. The results showed that with the increase in fine particle content, the peak concentration and peak time of rare earth increased. When the fine particle content exceeded 20%, all ion-exchangeable-phase rare earth ions could be replaced with a low dosage of the leaching solution. The leachate flow rate exhibited multi-stage variation, influenced by solution permeation, ion exchange, and fluctuations in accumulated liquid height. A mass transfer analysis showed that a higher fine particle content corresponded to a smaller plate height and a larger plate number of theoretical plates. As fine particle content increased, the final rising height of capillary water decreased, with rising rates varying across different stages for the samples. Moreover, an increase in fine particle content from 5% to 20% resulted in a 94% decrease in the samples’ permeability coefficients. A mechanism analysis showed that when the fine particle content was higher, the fine particles were embedded in the gaps between coarse particles, and the ore particles in the sample were arranged continuously, resulting in a lower permeability coefficient. Then, the leaching solution could penetrate uniformly, which was beneficial for reducing leaching blind spots and improving leaching efficiency. However, excessive fine particle content might have detrimental effects. Based on these results and considering actual mining conditions, the optimal fine particle content for rare earth leaching is 20%.