Minerals Engineering最新文献

Pyrometallurgical recycling of lithium-ion batteries (LIB) has established itself as a robust process in industrial practice due to its good scalability. A major drawback of this approach is the slagging of lithium, which limits its recovery and usually requires thermal energy and large amounts of leaching reagents using a hydrometallurgical recovery process in order to accomplish a return into the material cycle. To counteract this disadvantage, the present study investigates the fine grinding behaviour of battery slags in a stirred media mill and its possibility to increase the solubility of lithium containing slag phases. As the main influencing factors, the grinding media stress energy, processing time, and the pH value were investigated. The results show that by selecting suitable fine grinding process parameters, the specific surface area of the battery slag can be increased significantly from 0.2 m²/g to 55 m²/g and a lithium dissolution efficiency of up to 30 % can be reached in an aqueous environment. The variation of pH value during fine grinding enables a further process improvement with a dissolution efficiency of up to 90 % at pH=4. Particularly in the context of sustainable recycling process design, fine grinding offers a notable benefit in decreasing the quantity of leaching reagents required by as much as 76 % compared to standard leaching processes.

The interaction of surfactants at minerals-water interfaces in flotation system plays an important role in minerals separation. Although the increasing number of reports have been published on the sodium oleate (NaOL) adsorption at the mineral–water interface, there has been little development in describing and predicting its adsorption behaviors from a quantitative molecular insight. In this study, based the adsorption experiments, the adsorption characteristics of NaOL on the hematite/quartz-water interfaces were quantitatively depicted using the surface complexation model (SCM). There was a monodentate binding form between the function group on NaOL and the hematite and quartz surface sites during SCM fitting. The binding constants (logK) of NaOL are 10.72 (≡Fe-OL) and 8.03 (≡Si-OL), respectively. Notably, there are more positive ≡FeOH₂⁺ site on the hematite surface, and it has strong adsorption capacity with anionic surfactant NaOL. Moreover, the minerals surface potentials, NaOL adsorption capacity and flotation recovery in mixed ore systems were successfully predicted by the model. This study provides a credible evaluation of the adsorption characteristics of NaOL under a broad scope of pH and concentration conditions. Meanwhile, quantitative analysis of surfactants adsorption at minerals-water interfaces is beneficial for the intelligence of mineral processing technology and efficient separation of minerals.

Generating a flotation reagent scheme for a complex base-metal sulphide ore using reagents that are commonly found in mineral processing can be a challenging, complex process. Ensuring that optimum operating conditions including particle size, pH, types of reagents, and dosages of reagents are found is critical in establishing a foundation for future research. A modified xanthate reagent scheme was chosen using potassium amyl xanthate (PAX) as the collector, methyl isobutyl carbinol (MIBC) as the frother, copper sulphate as activator, carboxy methyl cellulose (CMC) and zinc sulphate as depressants. The operating conditions that generated the highest recoveries and grades for copper and zinc were found using a grinding time of 3 min and a slightly acidic medium of pH 6 achieving maximum recoveries of 84.0% and 83.8% for chalcopyrite and sphalerite, and grades of 3.9 and 21.0, respectively. These results allow for comparisons to industrial results with ore of a similar composition while simultaneously providing a baseline in which to evaluate the results of future, innovative research.

Copper oxide minerals often require fine grinding to liberate them from gangue minerals, but overgrinding during industrial milling can result in their loss during flotation. This study investigates the use of triethanolamine (TEA) as a cost-effective and low-polluting modifier to enhance the recovery of fine malachite. The flotation tests of an oxide copper ore indicated that TEA modification increased the total copper recovery of concentrate by 17.69 % and the copper recovery in the fraction with particle sizes <30 μm by 6.59 %. The modified fine malachite (<38 μm) achieved a maximum recovery of over 82 % in terms of malachite micro-flotation. Zeta potential measurements and adsorption tests confirmed that TEA increased the surface potential and the number of active sites, enhancing the S ion chemisorption. Solution chemistry, X-ray photoelectron spectroscopy, and scanning electron microscopy analysis revealed that the predominant Cu − TEA complex was Cu(TEA)(OH)⁺, which exhibited high reactivity with S ions and increased the amounts of Cu(I) sulfides and polysulfides in the form of flakes and micro-globular precipitates on the fine malachite surfaces. These precipitates increased the distribution density of hydrophobic Cu(I)-xanthate upon xanthate addition, based on a microscopic Fourier-transform infrared spectroscopy investigation. The findings suggest that TEA exhibits significant potential for enhancing the surface sulfidization and xanthate flotation of fine copper oxide ores for industrial applications.

Investigate the effects of Ionic Liquid and metal ions on the swelling and pyrolysis characteristics. The simulation results show that IL can weaken intermolecular hydrogen bond of coal, Magnetic Ionic Liquid has a stronger modification effect after adding Fe ions, and hydrogen bond weakening leads to the improvement of coal swelling degree. After [Bmim]Cl modification, the swelling degree is increased by 17.37% compared with raw coal, and after [Bmim]FeCl₄ treatment, the swelling degree is further improved. Ionic liquid swelling also has a better effect on subsequent pyrolysis of coal samples, the modification effect of [Bmim]FeCl₄ is more significant, which increases by 6.77%. The swelling of MIL significantly reduces the pyrolysis activation energy of coal and thermochemical reaction temperature. On the one hand, the content of heavy groups in pyrolysis products is reduced, and on the other hand, the target of carbon emission reduction can be achieved.

Beneficiation of the tailings from Iron Oxide Apatite (IOA) ore has become an important topic in the field of mineral processing as phosphate rock is considered as critical raw material by the European Union. Driven by the strong call for sustainability and green technology, this paper introduces the application of novel and bio-based organosolv lignin particles (OLP) as a reagent for apatite flotation. In the artificial mineral mixture flotation tests, OLP addition or replacement to tall oil fatty acid-based collector (TOFA) was shown to improve flotation kinetics and recovery. In this study, it was demonstrated that one of the widely used commercial TOFA collectors could be replaced with OLP by 70 %. The replacement led to an increase in recovery (+2%) and only a minimal decrease in P grade (−0.3 %) for the rougher-cleaner flotation tests in one of the two feed types tested. The influence of OLP and other reagents on apatite floatability has been investigated through Hallimond tube tests and laboratory scale batch flotation tests as well as zeta potential measurements and spectroscopy tests to further understand the possible mechanism and synergism of reagents in the apatite flotation system.

In this study, a novel collector, named O-methylisobutyl-N-allyl thionocarbamate (MIBATC) was synthesized and characterized by Fourier Transform Infrared Spectrometer (FTIR), ¹³C NMR, ¹H NMR analysis. The flotation behavior of MIBATC was investigated by flotation experiments and interaction mechanism was conducted by adsorption amount, zeta potential, FTIR, X-ray photoelectron spectroscopy (XPS) measurements and density functional theory (DFT) calculation. Flotation results demonstrated that MIBATC appeared powerful collecting ability to chalcopyrite against sphalerite and pyrite. The recovery of chalcopyrite was more than 80 % with pH of 9.5–10 and MIBATC concentrate of 30–40 mg/L. Adsorption amount and zeta potential measurements showed a tendency that MIBATC exhibited more affinities for the chalcopyrite compared with sphalerite and pyrite. FTIR and XPS analysis provided the evidence that MIBATC chemisorbed on chalcopyrite surface of Cu site, whereas physisorbed on sphalerite and pyrite surface. Specially, DFT reconfirmed the fact that S atom in MIBATC donated electrons to Cu on chalcopyrite surface and the adsorption energy was −21.67 kcal/mol, regarding as an exothermic reaction. These findings provided evidence for superior separating chalcopyrite from sphalerite and pyrite in the presence of MIBATC.

In the flotation of water-soluble minerals (e.g., NaCl and KCl crystals), the high ionic-strength environments alter the typical interfacial interactions and challenge the applicability of conventional flotation theories. Here, we report an intriguing effect of Pb(II) ions on the flotation of NaCl and KCl crystals in brines using sodium/potassium laurate as collectors. We revealed contrary effects of Pb(II) ions on flotation recoveries of NaCl crystals and KCl crystals. Pb(II) ions strongly enhance the NaCl crystals recovery but show no considerable effect on KCl crystals recovery. Interestingly, such activation is observed exclusively for Pb(II), not for Ca²⁺, Mg²⁺, Ba²⁺, and Cu²⁺. Our analysis using XPS and XRD rules out the adsorption of lead on crystal surfaces but reveals a significant presence of lead on the collector colloids via ion substitution. In particular, we unveil a crucial role of interfacial hydrogen bonding between Pb(II) ions in the collector colloids and the outward-pointing OH groups of surface-bound water on NaCl crystals. These hydrogen bonds sustain a strong bubble-particle attraction for high recovery of NaCl crystals. No similar hydrogen bonding is possible for KCl crystals which causes a poor recovery of this salt. These insights provide hints to tailoring the flotation conditions to enable efficient separation of water-soluble minerals, which is the core interest of the potash industry.

Driven by the explosive development of the photovoltaic (PV) industry, the treatment of the quartz crucible waste ash (QCWA) from monocrystalline silicon rod production must be considered to combat the shortage of silicon materials and promote sustainable development. In particular, the loss of grade 4 N high-purity silicon in QCWA is a frustrating fact for the silicon supply chain. In this work, an electrical separation process is proposed for the recovery of silicon and quartz from QCWA to realize waste resource reutilization. The charging processes and forces in the feed QCWA are first analyzed. Then, the electric field distribution during electrical separation is simulated to clarify the movement models of silicon and quartz particles and formulate reasonable electrical separation parameters. After systematic theoretical analysis, calculation, and simulation, electrical separation experiments were conducted. The results prove that the content of silicon in the concentrate and the corresponding content of quartz in the tailing respectively exceeded 93 % and 61 % under a particle size of 80 ∼ 120 mesh, a voltage of 40 kV, and a roll speed of 75 r/min. This work demonstrates that electrical separation is a sustainable process that could be recommended for silicon and quartz recovery from QCWA.