Froth Product Research Articles

Mixed surfactants with solubilization behaviors: Separation of feldspar and quartz by self-assembly flotation

This study designed a mixed surfactant with solubilization behavior for selective flotation separation of feldspar and quartz. By optimizing the compounding ratio and dosage of reagents, the optimal mixed surfactants, dodecylamine (DDA), sodium dodecyl sulfate (SDS), and dodecane with the concentrations of 1.75, 0.58, and 3.5 mmol/L, used in flotation separation of feldspar and quartz has the maximum flotation difference of 91.36 % at pH 3 for single mineral. The result of flotation experiments of artificial-mixed-ore shows that the Al2O3 grade was maintained at 14.23 %, and Al2O3 recovery was achieved at 69.21 % for the froth product with a selectivity index (SI) of 1.59. The bench-scale flotation tests confirmed a better effect by using mixed surfactants with Al2O3 grades of 14.72 % and the Al2O3 recovery of 66.62 %. The separation mechanism of quartz and feldspar was proposed through analyses of critical micelle concentration (CMC), fourier transform infrared spectrometer (FTIR), contact angle, zeta potential, microscopy image, and X-ray photoelectron spectroscopy (XPS) measurements, and indicated that the flotation difference between quartz and feldspar mainly originated from the selective adsorption of anions on feldspar and the aggregation and subsidence of quartz in mixed surfactants.

Влияние состава гидроксамовых кислот на показатели флотации перовскита.

On a sample of perovskite-bearing ore of the testing of reagents-collectors from the class of hydroxamic acids, differing in the structure of the hydrocarbon radical, was carried out. The reagents based on octanoic acid and acids of vegetable oils — coconut and tall oil — were considered. It is shown that in the alkaline pH range the rough flotation is more active, and a considerable amount of gangue minerals is transferred to the froth product. Strong attachment of the collector on all minerals determines low efficiency of re-cleaning operations in alkaline conditions. The optimum pH of flotation is pH = 6.2–6.5. Increase in the length of hydrocarbon radical above 10 carbon atoms leads to a significant decrease in the selectivity of the collector action. Reagents based on coconut and tallic acids actively float both perovskite and gangue minerals. Of the reagents considered, octylhydroxamic acid is characterized by the most optimal flotation properties and provides obtaining perovskite concentrate with TiO2 ~49–50 %.

Pre-concentration of REE values from nepheline syenite rocks of western Odisha, India

ABSTRACT The rare-earth elements (REEs) have gained enormous economic and scientific attention due to their distinct properties and new applications. Nepheline syenite rock is an important primary source of rare earth minerals. The nepheline syenite rocks from the Rairakhol area, western Odisha, have been characterized for their rare earth mineralogical and textural characteristics. Preliminary beneficiation studies of these rocks were carried out to upgrade the REE values using falcon gravity separation and floatation technique. Hornblende, biotite, K-feldspar, albite, and nepheline are the major mineral constituents of the nepheline syenite where REE-bearing minerals occur as accessory minerals. The REE mineral phases present are zircon, sphene, apatite, allanite, britholite, and REE phosphate, with grain sizes ranging from 10 to 50 microns. Allanite and britholite occupy the intergranular spaces of major mineral phases and occur as thin lines or veins whereas sphene and zircon occur as inclusions within major mineral phases. The flotation study shows that the total REE content of the froth product could be enriched to 1696 ppm, which is three times more than the total REE content (563 ppm) of the feed sample.

Rhenium extraction from dilute solution by precipitation flotation and oxidative volatilization techniques

The low mass transfer effect of dilute solution results in low recovery efficiency of critical metal Re ions from ultralow-concentration metallurgical industrial effluents. Bubble driving was an effective measure to enhance the mass transfer in the dilute solution, and bubble flotation technology was adapted to enrich and recover metal ions. In this work, the extraction of critical metal Re from dilute solutions by a precipitation flotation process was investigated. The precipitation flotation process includes ion precipitation by CTAB and ammonium salt, flocculation by PAM and air bubble flotation separation. The effects of technical parameters on Re recovery were systematically investigated. In the precipitation stage, NH4+ with a positive charge firstly reacts with the rhenate (ReO4-) with a negative charge via electrostatic action and then further reacts with CTAB to form fine CTAB(NH4ReO4) precipitate through complexation coordination. In the flocculation procedure, the particle size of precipitate flocs is elevated due to the physical polymerization of PAM polymer. In the flotation procedure, the long carbon chains on the surface of the flocs have a foaming effect due to their strong hydrophobicity, and they are trapped by the upfloated bubbles and rapidly enriched in the foams. Under the optimized conditions, the recovery efficiency of Re is 99.5% and qualified circulating water with a turbidity of 9.8 is obtained. Eventually, highly purified Re oxides (99.8%) are prepared from the Re froth product via oxidizing volatilization since the utilized organic CTAB and PAM can be burned completely during the oxidizing calcination process.

A phenomenological kinetic flotation model: Distinct Time-Variant floatability distributions for the pulp and froth materials

A simple and easy-to-use phenomenological kinetic flotation model, strongly connected with the physics of the process, is proposed in this paper. The model explicitly contains the cell volume, aeration rate, volumetric hold-up, mean bubble size, and particle density as input variables. It can be employed to characterize the floatability distributions of the particles in the pulp and the froth separately any time during the flotation process. Two new time-dependent kinetic parameters, the bubble loading factor φ(t) and the maximum cell mass transfer capacity Ṁmax(t) also appear in the model expression. φ(t) is a measure of the degree of crowding of the bubble surfaces and accounts for the deviations from the first-order rate equation. Ṁmax(t) describes the maximum amount of mass that can be transported to the froth phase by the bubble population in the cell. Screen fractionation of each froth product collected at different time intervals during a single kinetic flotation test is sufficient to generate the data required by the model for analysis. Application of the model to this data yields directly time-dependent functions for the floatability of the particles reporting to froth Kf(t) or remaining in the cell Kp(t) for each size fraction separately, without the need for any empirical parameters. The test of the model was carried out using published kinetic flotation data from the literature.

Effects of a cyclonic microbubble flotation column operating parameters on coal process responses

ABSTRACT Processing fine and ultrafine coal particles from primary resources is essential for sustainable development. As one of the most recently developed enrichment equipment, cyclonic microbubble flotation columns (FCMC) showed high efficiency in upgrading various fine minerals. However, understanding the effect of FCMC operating variables on coal upgrading process responses remains a black box and needs fundamental assessments. To fill the gap, this study examined the influence of circulating pump pressure and froth height on the recovery of different coal particle size ranges to assess and explore fundamental FCMC performance. The experimental results established that increased froth height would simultaneously reduce the collection and froth recovery. Coarse particles in the high froth height were detached due to bubble coalescence or rupture, while the liquid drainage reduced fine particle pollution of froth products. Increasing circulating pump pressure would increase the collection zone’s turbulence, lead to coarse particle detachment, and reduce collection recovery. Meanwhile, increasing circulating pump pressure could promote froth stability and decrease froth detachment in the froth zone. The calculated first-order rate constant (k c) confirmed that the flotation rate constants of coarse (−0.5 + 0.25 mm) and fine (−0.074 mm) particles were low, while the intermediate particles (−0.125 + 0.074 mm) had faster flotation kinetics.

Effect of the Molecular Weight of Carboxymethyl Cellulose on the Flotation of Chlorite.

The present study aimed to investigate the influence mechanism of carboxymethyl cellulose (CMC) on the flotation of fine chlorite. To this end, a series of flotation tests, sedimentation tests, and microscope analyses were conducted. Flotation tests revealed an inverse relationship between particle size and the recovery of chlorite, indicating that finer particles exhibited higher recovery rates. Moreover, it was observed that the recovery of fine chlorite was significantly associated with the water recovery (proportion of water entering the floated product to the weight of water in the initial flotation suspension) and a variety of frother types. Based on these findings, it can be inferred that froth entrainment may constitute a crucial component of the recovery mechanism underlying fine chlorite. Thus, reducing froth entrainment (the phenomenon of hydrophilic minerals entering floated products through foam water) is the key to depress chlorite flotation. Flotation tests indicate that fine chlorite recovered into froth products can be depressed effectively by CMC with a high molecular weight. The results of sedimentation tests and microscope analyses in the presence of CMC prove that CMC with a high molecular weight generates flocculation on fine chlorite particles while that with a low molecular weight does not. It is suggested that the depression of chlorite flotation may be attributed to the reduction in the entrainment resulting from the flocculation induced by CMC.

Obtaining copper concentrate during iron ore processing

The data on the complex processing of iron ore from one of the deposits of the Republic of Kazakhstan, which involves several operations of wet magnetic separation with re-grinding of raw products and their subsequent refining to produce a conditioned iron concentrate with 65–66 % iron containing 79–80 % Fe and 2.2–2.5 % Si, are presented. It was found that during the magnetic enrichment of the ore under study, the copper minerals concentrate in the magnetic separation tailings and the copper content in them increases from 0.093 to 0.2 %. A scheme and reagent system have been developed for the recovery of conditioned copper concentrate from magnetically enriched tailings. To obtain copper concentrate, magnetic separation tailings are subjected to regrinding in a lime medium to a fineness of 75 % of the –0.071 mm grade. After two operations of the main copper flotation with the use of water glass, butyl xanthate and frother MIBK, waste tailings are obtained. The froth product of the first basal flotation is cleaned twice. The result is a copper concentrate containing 15.2 % copper, 26.5 % iron, 17.5 % sulfur, 3.47 % silicon, 1.4 % aluminum and 8.5 % zinc, which corresponds to the KM-7 grade according to GOST R 52998-2008. Waste tailings contain: copper 0.08 %, iron 20.1 %, sulfur 0.25 %, silicon 16.2 %, aluminum 6.4 % and zinc 0.045 %. The influence of xanthates with different length and structure of hydrocarbon radical as well as hostaflots and amyl aeroflots on the process of copper flotation is studied. The high efficiency of butyl xanthate in the flotation of copper minerals has been confirmed.

Selective separation mechanism of hematite from quartz by anionic reverse flotation: Implications from surface hydroxylation

Reverse flotation desilication with anionic collector sodium oleate (OL) and activator Ca2+ is a common beneficiation method to purify hematite. Herein, the separation mechanism of hematite and quartz were systematically investigated by experiments and DFT calculations with the effect of surface hydroxylation fully considered. Flotation results showed that the success of anionic reverse flotation of hematite was inseparable from the activation of quartz by Ca2+. In addition, the use of sodium humate (HM), which selectively inhibited hematite, was also very critical. The Zeta potential revealed that the activator Ca2+ ions were highly selective for quartz and depression HM was highly selective for hematite. SEM clearly showed that Ca species were adsorbed on the quartz surface, which was confirmed as chemisorption by XPS. DFT calculations further confirmed that the adsorption of HM on the surface of hematite was more favorable than that of quartz, while Ca2+ on the surface of quartz was more favorable than that of hematite, and the chemisorbed Ca2+ ions on the quartz surface provided targeted reaction sites for the subsequent adsorption of OL, which enabled the activated quartz to overflow into the froth products, while the inhibited hematite to sink and concentrate in the flotation cell.

A Review on Froth Washing in Flotation

In the attempt to process lower-grade ores, mineral flotation has taken centre stage as the preferred recovery route. However, in many instances, the froth product does not have a high grade due to the entrainment of gangue minerals. Industry has solved this challenge by introducing froth washing mechanisms. Clean wash water is introduced into or on top of the froth to reduce the amount of entrained gangue in the final concentrate. This article reviews froth-washing systems in detail and highlights the advantages and disadvantages of each wash-water delivery mechanism. Comments on industrial uptake are provided. The indications are that froth washing improves the grade of the concentrate and influences froth stability and mobility. Other researchers have reported an improvement in recovery—especially of coarse particles—with wash water being added, while others have reported a reduction in recovery, especially with composite particles. Froth washing is generally applied in mechanical flotation cells by washing at the lip. In column flotation cells and Jameson cells, wash water is added to the entire froth surface. The literature also indicates that the wash-water rate, wash-water quality, type of wash-water delivery/ distribution mechanism and the area covered by wash water are critical parameters that dictate the efficacy of the washing system. Further research is necessary on the impact of wash-water quality on the froth phase sub-processes including froth rheology.

Optimization of cell depth and concentration zone height in a countercurrent flotation machine

Due to the growing complexity of ore processing, equipment and technologies with higher separation efficiency are currently in demand. Flotation machines with a countercurrent flow of slurry and bubbles are increasingly being used. Yet still, certain issues related to optimizing flows in a flotation cell by adjusting the cell height and feed point depth remain insufficiently studied. Pilot tests of a countercurrent flotation machine with a jet-ejector aerator in rougher flotation at a facility of Phosphorit Industrial Group LLC have demonstrated an increase in separation efficiency after a chamber height adjustment from 3.5 to 5.0 m with a 1.5 m deeper feed point. The mass fraction of P2O5 in the froth product increased from 20–22 to 25–27 % with a decrease in froth yield from 32–35 to 21–24 %. The flotation tail grades correspond to those of the dump waste product (1.00 ± 0.39% P2O5 according to the data on 116 samples); the load of cleaner operations has also been normalized. An original sampler designed by B. S. Chertilin was used to test the mineral load of bubbles at different depth points in the cell. The results indicate the extreme nature of the load distribution with a monotonous increase in P2O5. This confirms the main role of the process of predominant shedding of waste rock particles from the bubbles in the concentration zone (above the feed point) with the higher separation efficiency achieved.

Application of competitive adsorption of ethylenediamine and polyetheramine in direct float of quartz from quartz-feldspar mixed minerals under neutral pH conditions

Flotation is the most effective way to separate quartz and feldspar, but conventional flotation is usually achieved under acidic or alkaline conditions, which causes pollution to the environment. This study proposes a new method for a direct float of quartz from quartz-feldspar mixed minerals under neutral pH conditions. The method achieves the separation of minerals by competitive adsorption of ethylenediamine (EDA) and polyetheramine (PEA) on the particle surface. The effect of EDA and PEA on the separation of the quartz and the feldspar was investigated by flotation experiments. Using Fourier transform infrared spectroscopy (FTIR), X-ray photoelectron spectroscopy (XPS), X-ray powder diffractometry (XRD), Inductively coupled plasma-optical emission spectrometry (ICP-OES), and X-Ray Fluorescence Spectrometer (XRF) tests, the flotation behavior, and adsorption mechanism were analyzed. The result of the flotation experiments of a single mineral shows that EDA can effectively depress the flotation of the feldspar and enlarge the difference between the flotation recovery of the quartz and the feldspar, up to Δ = 77%. The result of flotation experiments of the mixed minerals shows that the combination of EDA and PEA can effectively separate the quartz and the feldspar. For the froth products, the quartz content of 2-F-Qtz reaches 97.42% and the SiO2 grade is 99.12% at a quartz flotation recovery of 70.20%. For the sink products, the feldspar content of 4-S-Fsp reaches 94.77% and the SiO2 grade is 69.50% at a feldspar flotation recovery of 87.45%. A competitive-adsorption model is proposed to explain the mechanism of the direct float of the quartz from the quartz-feldspar mixed minerals under neutral pH conditions.

Flotation separation of molybdenite from chalcopyrite using mechanically degraded polyacrylamide as a novel depressant

In this work, cationic polyacrylamide (CPAM) which was degraded by mechanical shear was used as a chalcopyrite potential selective depressant to separate molybdenite from chalcopyrite by flotation while using kerosene as a collector. Micro-flotation experiments of single minerals show that CPAM was the most suitable to be used as the chalcopyrite depressant among the three different types of polyacrylamide. The molecular weight of CPAM was reduced from 12 million to 1.8 million and the depression performance of CPAM on chalcopyrite was increased after mechanical rotational shear degradation for 6 h. When the dosage of CPAM which was degraded by mechanical shear for 6 h (CPAM-6 h) was 6.0 mg/L, good separation results were achieved in the artificially mixed minerals experiment. In froth products, the recovery of molybdenite was 87.96 %, the grade of molybdenum was 54.60 %, and the grade of copper was only 2.58 %. The depression mechanism of CPAM-6 h on chalcopyrite and molybdenite was systematically investigated by contact angle measurement, zeta potential measurements, FT-IR spectra analysis, and X-ray photoelectron spectroscopy (XPS) analysis. The analysis results show that CPAM-6 h has not only electrostatic adsorption but also chemical adsorption on the surface of chalcopyrite while only electrostatic adsorption on molybdenite, then the addition of kerosene can destroy the hydrogen bonding of CPAM-6 h on the surface of molybdenite and make it hydrophobic again, but kerosene cannot remove the copper-ammonium complex formed by CPAM-6 h on the surface of chalcopyrite, therefore, chalcopyrite remains hydrophilic.

Reverse Column Flotation of Ultrafine Magnetite Mixture with Fine Glass Beads Enhanced by Fine Bubbles

Magnetite ores are among the most important sources of iron, which is in high demand in the global economy. Metallurgical properties of the magnetite concentrate significantly depend on impurities of silicon- and aluminum-bearing minerals. These impurities have to be separated from ore by magnetic separation and flotation techniques. Reverse column flotation is one of the methods applied for reducing the content of impurities in magnetite concentrate. This method allows recovering impurities from ore in the froth product (flotation tailings). However, the efficiency of this method significantly decreases with the decrease in particle size. As previously demonstrated, the effectiveness of fine particle column flotation can be increased if, before feeding the pulp into a column, fine bubbles are introduced into the pulp and the pulp is then passed through a tubular reactor. The major purpose of this study was to define the effectiveness of the reverse column flotation performance of ultrafine magnetite from the mixture with fine glass beads (ballotini) when, before the pulp is fed into the flotation column, it is mixed with fine air-in-water dispersion, and the mixture is then passed through a tubular flotation reactor (TFR). The obtained experimental findings allowed the definition of the optimal conditions of the mixture treatment in TFR that ensured high concentrate grade and iron recovery for the initial iron content in the mixture of 63.76%. These conditions were defined as follows: treatment time of the mixture in the TFR—7.5 s; average flow shear rate inside the TFR—1000 s−1; volume dosage of fine bubbles per solid mass unit—0.032 or 0.21 mL/g. At the fine bubble dosage of 0.032 mL/g, the iron recovery and the concentrate grade were, respectively, 88.1% and 68.3% Fe, and at the dosage of 0.21 mL/g, the iron recovery reached 89.4% for the concentrate grade of 68.7% Fe.

Improving the flotation recovery of Cu from flash smelting slags by utilizing cellulose-based frother formulationss

• First published study using polymer-surfactant mixes as frothers in Cu slag flotation. • Using the PS-mix improved overall Cu recovery compared to commercial frother. • PS-mix significantly increased productivity, i.e. similar froth product in less time. • PS-mix demonstrated superior tolerance in performance when reducing xanthate dosage. Froth flotation is an operation currently used for the recovery of Cu from pyrometallurgical smelting slags due its simple operation and relatively low operational costs. While it has been utilized for many decades, it may be necessary to develop strategies that improve Cu recovery from slags, as the unprecedented increase in Cu demand has made industrial waste and side streams economically relevant sources of metals. This study presents a novel chemistry-based approach for the flotation of Cu from smelting slags by utilizing an amphiphilic cellulose-surfactant mixture as a flotation frother, hereby referred to as CellFroth. The present work pursues two main objectives: i) to improve the performance of flotation for the treatment of smelting slags and ii) to accomplish a first proof of concept for frothers produced from renewable sources in the context of industrial waste treatment. The flotation performance of CellFroth was compared with a commercially available additive (i.e., DowFroth250) and hydroxypropyl methyl cellulose (HPMC). The results showed significant advantages in performance for CellFroth. At the highest collector concentration, Cu recovery was higher with CellFroth compared to DowFroth 250 and pure HPMC (73%; 62% and 65% respectively). A 75% reduction in xanthate concentrations had negligible effects on Cu recovery for CellFroth (70%), while recoveries decreased drastically with DowFroth250 (47%) and pure HPMC (50 %). Finally, CellFroth consistently reached comparable concentrate productivity (grade and recovery) as DF250 within significantly shorter flotation times. The use of CellFroth thus offers the possibility of increasing the Cu recovery from flash smelting slags, while reducing the environmental impact of the process by using a frother produced from sustainable sources and simultaneously reducing xanthate consumption.

Recovery of Graphite and Cathode Active Materials from Spent Lithium-Ion Batteries by Applying Two Pretreatment Methods and Flotation Combined with a Rapid Analysis Technique

This work investigates the comprehensive recycling of graphite and cathode active materials (LiNi0.6Mn0.2Co0.2O2, abbreviated as NMC) from spent lithium-ion batteries via pretreatment and flotation. Specific analytical methods (SPME-GC-MS and Py-GC-MS) were utilized to identify and trace the relevant influencing factors. Two different pretreatment methods, which are Fenton oxidation and roasting, were investigated with respect to their influence on the flotation effectiveness. As a result, for NMC cathode active materials, a recovery of 90% and a maximum grade of 83% were obtained by the optimized roasting and flotation. Meanwhile, a graphite grade of 77% in the froth product was achieved, with a graphite recovery of 75%. By using SPME-GC-MS and Py-GC-MS analyses, it could be shown that, in an optimized process, an effective destruction/removal of the electrolyte and binder residues can be reached. The applied analytical tools could be integrated into the workflow, which enabled process control in terms of the pretreatment sufficiency and achievable separation in the subsequent flotation.

Beneficiation properties of ash-and-slag dumps

Purpose. Development of a mathematical model for the indices of TPP ash-and-slag waste separation in closed flotation circuit when applying recleaning operations of enriched product. Methodology. The methods for calculating separation indices using the separation characteristics of separation processes and calculating the indices of the release of valuable mineral using statistical relationships between the particle size of the inclusion, ash content and particle size distribution function were used. Findings. Separation characteristics of rough and recleaning flotation operations were preliminarily determined experimentally, relations that allow calculating the indices of the release of coal particles in ash-and-slag dumps were defined, based on which the functions of particle distribution by their content of valuable mineral fractional composition were found. The separation characteristics of the connections of the separation blocks, including several operations of cleaning of the enriched froth product, were determined. Calculations to determine the quantitative and qualitative separation indices were performed. Originality. Analytical relationships to determine the indices of the release of coal inclusions in the process of material from ash-and-slag dumps grinding were obtained. Practical value. The established dependencies make it possible to justify the required grinding coarseness to achieve the specified parameters of the enrichment process based on the values of the beneficiation properties of ash-and-slag dumps.

Recovery of Catapleiite and Eudialyte from Non-Magnetic Fraction of Eudialyte ore Processing of Norra Kärr Deposit

Eudialyte ores from Norra Kärr (Sweden) and Kringlerne (Greenland) are considered a potential source of rare-earth elements (REE) for the development of a sustainable REE industry outside China. Magnetic separation is successfully applicated to recover eudialyte as a magnetic fraction. In the case of the Norra Kärr deposit, up to 20% of the REE and up to 40% of the Zr are lost during mineral processing in the non-magnetic fraction. Zr and REE are associated with non-magnetic minerals such as catapleiite, low- or non-magnetic eudialyte species, and both their intergrowths. Besides zirconosilicates such as catapleiite and eudialyte, the non-magnetic fraction has valuable and already-liberated minerals such as alkali feldspars and nepheline, which should not be considered as tailings. In this investigation, a possible way to recover REE bearing zirconosilicates from the non-magnetic fraction using flotation is presented. First, a low-grade eudialyte concentrate (1.8% Zr, 0.94% REE) from ground ore was obtained using magnetic separation. The non-magnetic fraction was then treated using froth flotation, and a Zr-REE bearing product (9% Zr, 1.5% REE) was obtained as froth product. For this purpose, phosphoric acid esters were used as selective collectors for zirconosilicates at a pH between 3.5 and 4.5. The reagent regime could be proposed not only to recover Zr- and REE-bearing minerals, but also simultaneously to remove Fe, Ti, and other colored impurities from the nepheline-feldspar product and to minimize the tailings volume.

Recycling of electrode materials from spent lithium-ion battery by pyrolysis-assisted flotation

In this study, the flotation technology is used for the separation and purification of cathode and anode material after removing organics by pyrolysis. Effects of surface microscopic properties on the flotation efficiency of spent electrode materials are investigated, and on this basis, surface analysis combined with flotation foundation tests are conducted to reveal pyrolysis-assisted surface modification mechanism. Residual electrolyte and organic binders decrease the difference in wettability of cathode and anode material, and hinder the interaction between electrode particles and collecting agent. Organics and their pyrolysis products can be adequately removed at the optimum pyrolysis temperature of 550 °C. After pyrolysis, the change of Zeta potential and surface free energy demonstrates that hydrophilic components of cathode material increase while the hydrophobic components of the anode material increase. The induction time of cathode material rises from 190 ms to 650 ms while the induction time of the anode material decrease from 145 ms to 37 ms. Collector adsorption capacity of anode material is obviously improved and anode material is easy to incorporate with bubbles and will be collected in the froth product. After one stage flotation, the recovery rate of cathode material is 83.75% with a high grade of 94.72%.

Direct Recycling of Blended Cathode Materials by Froth Flotation

Direct Li‐ion battery recycling involves separating cathode active materials in the solid phase while preserving their electrochemical performance. To reuse the recycled ones in new batteries, it is necessary to separate individual cathode components, which would typically occur prior to the rejuvenation and any compositional reformulation processes. Herein, a froth flotation process is developed to separate pristine lithium nickel–manganese–cobalt oxide (LiNi0.333Mn0.333Co0.333O2, NMC111) and lithium manganese oxide (LMO) materials. The flotation results show that with multiple stages of the separation processes, a satisfactory separation is achieved with 95% grade or above of NMC111 in the froth product and 95% grade of LMO in the tailing product. The electrochemistry results show that the flotation chemicals have negligible impact on the electrochemical performance of recycled active materials, whereas the NMC111 exhibits a minor capacity fade after being in contact with deionized water at 2% solid concentration. An improvement in both the rate and cycling performance is achieved at 10% solid concentration or higher. It is found that there is little compositional and structural change to the aqueous recycled NMC111 powders. An effective and low‐cost separation method for direct recycling mixed cathode compositions is demonstrated.