Characterization and Structural Insights of the Reaction Products by Direct Leaching of the Noble Metals Au, Pd and Cu with N,N′-Dimethyl-piperazine-2,3-dithione/I2 Mixtures

In the context of new efficient and safe leaching agents for noble metals, this paper describes the capability of the Me2pipdt/I2 mixture (where Me2pipdt = N,N′-dimethyl-piperazine-2,3-dithione) in organic solutions to quantitatively dissolve Au, Pd, and Cu metal powders in mild conditions (room temperature and pressure) and short times (within 1 h in the reported conditions). A focus on the structural insights of the obtained coordination compounds is shown, namely [AuI2(Me2pipdt)]I3 (1), [Pd(Me2pipdt)2]I2 (2a) and [Cu(Me2pipdt)2]I3 (3), where the metals are found, respectively, in 3+, 2+ and 1+ oxidation states, and of [Cu(Me2pipdt)2]BF4 (4) and [Cu(Me2dazdt)2]I3 (5) (Me2dazdt = N,N′-dimethyl-perhydrodizepine-2,3-dithione) compared with 3. Au(III) and Pd(II) (d8 configuration) form square–planar complexes, whereas Cu(I) (d10) forms tetrahedral complexes. Density functional theory calculations performed on the cationic species of 1–5 help to highlight the nature of the bonding in the different complexes. Finally, the valorization of the noble metals-rich leachates is assessed. Specifically, gold metal is quantitatively recovered from the solution besides the ligands, showing the potential of these systems to promote metal recycling processes.

Sº and jarosite behavior during recovery of values from the direct leaching residue of sphalerite using cyanide and glycine

This paper reports the effect of the components of a direct leaching residue (jarosite and elemental sulfur), on the recovery of valuable metals such as gold and silver. Leaching media such as cyanide and mixtures of cyanide with glycine were used to recover the gold and silver from the residue; however, a low recovery of these metals was obtained. The above due to the negative effect of its components which cause problems in the extraction process such as encapsulation of silver (due to jarosite) and the formation of thiocyanate and re-precipitation of silver (due to sulfur). Various treatments prior to leaching were tested, finding that when the residue is desulfurized with perchlorethylene and subjected to an oxidizing alkaline hydrothermal treatment, the gold extraction increased from 39.73 to 88% and the silver extraction of 64.76 to 94.29%. Additionally, it was determined that when cyanide is assisted by glycine, the latter decreases the cyanide consumption by inhibition of the dissolution of iron and sulfur in cyanide.

A Combined Pyro- and Hydrometallurgical Approach to Recycle Pyrolyzed Lithium-Ion Battery Black Mass Part 2: Lithium Recovery from Li Enriched Slag—Thermodynamic Study, Kinetic Study, and Dry Digestion

Due to the increasing demand for battery raw materials, such as cobalt, nickel, manganese, and lithium, the extraction of these metals, not only from primary, but also from secondary sources, is becoming increasingly important. Spent lithium-ion batteries (LIBs) represent a potential source of raw materials. One possible approach for an optimized recovery of valuable metals from spent LIBs is a combined pyro- and hydrometallurgical process. The generation of mixed cobalt, nickel, and copper alloy and lithium slag as intermediate products in an electric arc furnace is investigated in part 1. Hydrometallurgical recovery of lithium from the Li slag is investigated in part 2 of this article. Kinetic study has shown that the leaching of slag in H2SO4 takes place according to the 3-dimensional diffusion model and the activation energy is 22–24 kJ/mol. Leaching of the silicon from slag is causing formation of gels, which complicates filtration and further recovery of lithium from solutions. The thermodynamic study presented in the work describes the reasons for the formation of gels and the possibilities of their prevention by SiO2 precipitation. Based on these findings, the Li slag was treated by the dry digestion (DD) method followed by dissolution in water. The silicon leaching efficiency was significantly reduced from 50% in the direct leaching experiment to 5% in the DD experiment followed by dissolution, while the high leaching efficiency of lithium was maintained. The study takes into account the preparation of solutions for the future trouble-free acquisition of marketable products from solutions.

Uranium removal from tin smelter slag by bisulfate fusion and acidic leaching

Slag as secondary product (waste) of tin smelter still contains not only valuable elements e.g. Ti, Nb, Ta, Zr, Hf and rare earth elements, but also radioactives such as Th and U, wich are accumulated in the slag phase during the smelting. Due to valuable element content, the slag becomes major of interest in mineral processing industries, hence the slag needs to be decontaminated before it could be processed further.Common approach to reduce U content from the slag using leaching process is considered ineffective due to association of U with refractory elements e.g. Si and Ti in the slag. To break down the refractory phases, the fusion approach by using fusing agent is required in order to release U so that they could be leached out using mild lixiviant.In this research, potassium hydrogen sulfate (KHSO4) and sulfuric acid was used as fusing agent and lixiviant, respectively. The parameters studied includes molar ratio between fusing agent and refractory elements in slag, fusion temperature, fusion time, sulfuric acid concentration in lixiviant and pulp density during leaching stage. The studies so far demonstrated that optimum condition in U removal occurred at fusion temperature 400 °C, fusion time 2 hours, molar ratio of potassium hydrogen sulfate to tin slag 5, sulfuric acid concentration 2 M and pulp density 15 ml/gr. The maximum recovery of U was 85.6%, which was significant compared to the results using direct leaching without fusion (0.1%).

Direct Acid Leaching of Sphalerite: An Approach Comparative and Kinetics Analysis

The present work reports the direct leaching of zinc from a sphalerite concentrate in acid media. Lab-scale and pilot-scale experiments were conducted in atmospheric-pressure and low-pressure reactors, respectively. Leaching of zinc and precipitation of iron was achieved in the same stage using different reagents like Fe3+, O2, O3, and Fe2+ (which is continuously oxidized in the leaching solution by H2O2 and O2). The highest percentage of zinc extraction (96%) was obtained in pilot-scale experiments using H2SO4, Fe2+, and O2. Experimental results were compared with those of other researchers to provide a better understanding of the factors influencing the dissolution of zinc. In the first instance, it was determined from analysis of variance that leaching time and the use of an oxidant agent (O2 or O3) were the most influential factors during the direct leaching of zinc from the sphalerite concentrate. Kinetic models were also evaluated to determine the rate-limiting step of the sphalerite leaching; it was concluded that the type of the sulfur layer formed in the residue (porous or non-porous) depends on the type of the oxidant used in the leaching media, which determines the dissolution kinetics of zinc.

Leaching of lead from sphalerite concentrate direct leaching residue by hydrochloric acid and sodium chloride for preparation of lead oxides

The purpose of this study was to select and propose an applicable method for extracting lead from sphalerite concentrate direct leaching residue. A large number of experiments were conducted to extract lead from sphalerite concentrate direct leaching residue by hydrochloric acid and sodium chloride solution as leachates. The main optimum parameters were determined, such as a liquid-solid ratio of 17.5-1, a reaction temperature of 85?C, an initial hydrochloric acid concentration of 1.3 mol/L, an initial sodium chloride concentration of 300 g/L, and a reaction time of 60 min. Ninety-five percent of the zinc, 96.0% of the iron, and 93.7% of the lead were extracted into leachate at the optimum conditions. The lead in the leachate was in the form of [PbCl4]2-. After the leachate was purified to remove impurities, it was converted into lead oxalate by sodium oxalate as a precipitant. Finally, lead oxalate was decomposed to obtain lead oxide powders via a high-temperature calcination process.