Leaching kinetics of calcification roasting calcinate from multimetallic sulfide copper concentrate containing high content of lead and iron

Rigid-Flexible Combined Impeller Enhancement in Leaching of Phosphate Rock: a Kinetics Study

Conventional rigid impellers are frequently used in the leaching process of phosphate rock, which often form a symmetrical flow field in the reactor, leading to a reduction in the leaching efficiency. In this work, a rigid-flexible combined impeller was applied to the leaching process of phosphate rock to increase the leaching efficiency. The effects of the reaction temperature (T), sulfuric acid excess coefficient (ε), liquid-solid ratio (L/S), agitation speed (N), and leaching time (t) on the leaching of phosphate rock were investigated, and based on this, the leaching kinetics was studied. The results indicated that under the optimum parameters of a reaction temperature of 353 K, a sulfuric acid excess coefficient of 1.15, a liquid-solid ratio of 4.0 mL/g, an agitation speed of 280 rpm, and a leaching time of 120 min, the leaching rate of phosphate rock using the rigid-flexible combined impeller reached 89.1%, which was 7.1% higher than that of the conventional rigid impeller under the same electric energy consumption. The leaching process complied with the unreacted core shrinking model, and the reaction rate was controlled by product layer diffusion. The apparent rate equation of the leaching process was 1 - 2X/3 - (1 - X)^2/3 = 2.06 × 10^-3[ε]^1.375[L/S]^1.273[N]^0.748 exp(-19.03 × 10³/RT)·t, and the apparent activation energy was 19.03 kJ/mol. The numerical simulation and analysis of the leaching residue showed that the system temperature in the rigid-flexible combined impeller system was homogenized, and the mixing effect of reactants was enhanced through the multiposition movement of the flexible connection piece in the axial direction, so that the reactants participated in the chemical reaction more efficiently.

Extraction of Lead and Zinc from a Rotary Kiln Oxidizing Roasting Cinder

In this study, sulfuric acid leaching and gravity shaking-table separation by shaking a table are used to extract lead and zinc from a Pb-Zn oxidizing roasting cinder. The oxidizing roasting cinder—containing 16.9% Pb, 30.5% Zn, 10.3% Fe and 25.1% S—was obtained from a Pb-Zn sulfide ore in the Hanyuan area of China by a flotation-rotary kiln oxidizing roasting process. Anglesite and lead oxide were the main Pb-bearing minerals, while zinc sulfate, zinc oxide and zinc ferrite were the main Zn-bearing minerals. The results show that a part of lead contained in lead oxide is transformed to anglesite, and a 3PbO·PbSO4·H2O-dominated new lead mineral phase after acid leaching. A zinc leaching efficiency of 96.7% was obtained under the leaching conditions used: a leaching temperature of 55 °C; a leaching time of 90 min; a sulfuric acid dosage of 20%; a sulfurous acid dosage of 4%; a cinder particle size of <0.3 mm; and a solid-liquid ratio of R = 1:4. After the gravity shaking-table separation, a lead concentrate with 50.2% Pb, 2.33% Zn and lead recovery of 86.0% was produced. The main chemical compounds in leaching residue are anglesite, 3PbO·PbSO4·H2O, SiO2 and ZnFe2O4, while the main chemical compounds in lead concentrate are anglesite, 3PbO·PbSO4·H2O and SiO2.

Fractal kinetic characteristics of hard-rock uranium leaching with sulfuric acid

In order to study the fractal dynamic properties of uranium leach mining and discuss the influence of ore crushing on the dynamics of leach mining, uranium mine ore rocks in southern China were selected as the research object and an acid leaching experiment was performed on the ore samples with different fractal dimensions of 1.1, 1.4, 1.7, 2.0, 2.3 and 2.6. In the column leaching experiment, a PVC pipe with an inner diameter of 112 mm and a height of 1500 mm was used. The uranium content was determined by using titanium trioxide that was placed into a 0.1 mg ml^-1 standard uranium solution, and a sampling rate of once daily with a 5 ml volume of leaching solution was adopted after 8 h drenching time. The results show that the flow rate of the leaching solution depends on the distribution of the ore's particle size, that is, a larger fractal dimension results in a smaller flow rate. The concentration of the uranium leaching solution reaches a maximum value which subsequently decreases with time on the third day of the experiment, and it seems that the changes in the uranium concentration tend to be stable at around 15 days. Moreover, the concentration seems to increase with the increasing fractal dimension, and the fractal dimension of the ore particle size has a significant impact on the leaching kinetics. When the fractal dimension is between 1.1 and 2.6, the uranium dissolution rate, K, increases with the increasing fractal dimension. The kinetic reaction of the uranium leaching is a liquid-solid one, which is controlled by chemical reactions in the earlier phase. While the middle reaction phase is mainly chemical-diffusion reaction coupling, and the latter part of the reaction is controlled by diffusion. As the fractal dimension increases, the liquid-solid reaction controlled by diffusion appears at earlier phases. When the fractal dimension is greater than 2.0, the time of entering the diffusion control phase only changed little with the increasing of the fractal dimension. At last, a fractal dimension of 2.0 is suggested for the acid leaching of uranium ore crushing.

Recovery of value-added products from cathode and anode material of spent lithium-ion batteries

Herein we report a low cost and eco-friendly approach for the recovery of metals from cathode and anode materials of mobile phone spent lithium-ion batteries (LIBs). Li-based metal oxide and graphite were efficiently separated from their respective foils and used for lixiviation. Acetic acid (CH₃COOH) and water were used as lixiviants for the recovery of metals from cathode and anode materials respectively. It was found that with 3 M Acetic acid and 7.5 vol% H₂O₂ as reducing agent 99.9% Li, 98.7% Co, and 99.5% Mn were leached out from cathode material in 40 min at 70 °C and a pulp density of 20 g/L. Besides the cathode leaching, Li was also extracted from anodic material graphite using water as a solvent and further recovered as solid Li₂CO₃ (99.7% Li). The kinetic evaluation of the cathode lixiviate process was studied using three different shrinking-core kinetic Models and established that the reaction follows the product layer diffusion controlled mechanism. From the cathode leach liquor, 99% Co was recovered as metal sulfide by controlled sulfide precipitation with 99.2% purity, and subsequently, MnCO₃ and Li₂CO₃ were obtained with the purity of 98.7% and 99.4%, respectively. The purity of the salts revealed that these products recovered from spent LIBs might be utilized in the electrochemical energy-storage applications. In addition, this recycling process would promote the sustainable development of the battery industry.

A perspective of stepwise utilisation of Bayer red mud: Step two--Extracting and recovering Ti from Ti-enriched tailing with acid leaching and precipitate flotation

The extraction and recovery of Ti from Ti-enriched tailing with acid leaching and precipitate flotation, as one of the critical steps, was proposed for the stepwise utilization of red mud. The factors influencing acid leaching and precipitate flotation were examined by factorial design. The leaching thermodynamics, kinetics of Ti(4+), Al(3+) and Fe(3+), and the mechanism of selectively Fe(3+) removal using [Hbet][Tf2N] as precipitating reagent were discussed. The extracting of Ti(4+), Al(3+) and Fe(3+) in concentrated H2SO4 is controlled by diffusion reactions, depending mainly upon leaching time and temperature. The maximum extracting efficiency of Ti(4+) is approximately 92.3%, whereas Al(3+) and Fe(3+) leaching are respectively 75.8% and 84.2%. [Hbet][Tf2N], as a precipitating reagent, operates through a coordination mechanism in flotation. The pH value is the key factor influencing the flotation recovery of Ti(4+), whereas the dosage of precipitating reagent is that for Al(3+) recovery. The maximum flotation recovery of Ti(4+) is 92.7%, whereas the maximum Al(3+) recovery is 93.5%. The total recovery rate for extracting and recovering titanium is 85.5%. The liquor with Ti(4+) of 15.5g/L, Al(3+) of 30.4g/L and Fe(3+) of 0.48g/L was obtained for the following hydrolysis step in the integrated process for red mud utilisation.