C. A. Nogueira

Leaching behaviour of electrode materials of spent nickel–cadmium batteries in sulphuric acid media

Abstract The improving awareness of environmental problems associated with the toxicity of heavy metals keeps the recycling of spent nickel–cadmium batteries an important assignment due to the presence of cadmium, nickel and cobalt on the electrode material. While cadmium from batteries is one of the major sources of cadmium contamination of the environment, the other heavy metals contained on the electrode material have a considerable economic value. Therefore a complete and valorising solution to the management of this type of residues is not possible through the existing pyrometallurgical processes because the treatment of complex materials in order to recycle all materials is difficult. An integrated process based on physical and hydroelectrometallurgical operations seems to be more efficient because it is possible to recover the three metals, Cd, Ni and Co, present on the electrode material. The study here presented deals with the first chemical stage of an integrated process, the leaching of spent nickel–cadmium electrodes with sulphuric acid. The electrode materials essentially composed of Ni, Cd and Co hydroxides were readily solubilised in 0.5 h with low acid concentrations (pH∼1) at ambient temperature. At higher pH values the solubilisation of metal hydroxides was inefficient, except when using long residence times. The leaching of nickel present in the metallic form, in the electrodes, was more difficult due to kinetic constraints, applying high temperature (e.g. 95 °C) and acid concentration (e.g. 2.5 M H 2 SO 4 ) in order to obtain complete conversions in acceptable time (∼4 h).

Recovery of Platinum and Palladium from Chloride Solutions by a Thiodiglycolamide Derivative

The liquid-liquid extraction of platinum(IV) and palladium(II) from hydrochloric acid media was carried out using N,N’-dimethyl-N,N’-dicyclohexylthiodiglycolamide (DMDCHTDGA) in 1,2-dichloroethane (1,2-DCE). Pt(IV) is efficiently extracted from 5 M HCl onwards (%E ≥ 97%), whereas Pd(II) is quantitatively recovered from 1 to 8 M HCl solutions. Both Pt(IV) and Pd(II) can be successfully stripped from the loaded organic phases, the former with a 1 M HCl solution, the latter with 0.1 M thiourea in 1 M HCl. The maximum loading capacity of DMDCHTDGA for Pt(IV) could not be determined but it is high, since molar ratios extractant:Pt(IV) within 2 and 3 have been achieved. Data obtained from successive extraction-stripping cycles suggest a good stability profile of DMDCHTDGA towards Pt(IV) recovery. Attempts to replace 1,2-DCE by more environmentally-friendly diluents showed, in general, comparable %E for Pt(IV). The study of the influence of acidity, as well as chloride ion and DMDCHTDGA concentrations, allows a proposal for the composition of the Pt(IV) species formed upon extraction. Results obtained with binary metal ion solutions point out that Pt(IV) and Pd(II) can be efficiently separated from DMDCHTDGA loaded organic phases through sequential selective stripping.

The Solvent Extraction Performance of N,N’-Dimethyl-N,N’-Dibutylmalonamide Towards Platinum and Palladium in Chloride Media

The solvent extraction performance of N,N’-dimethyl-N,N’-dibutylmalonamide (DMDBMA) in 1,2-dichloroethane (1,2-DCE) towards platinum(IV) and palladium(II) in hydrochloric acid media was systematically evaluated. Pt(IV) extraction (%E) increases with the HCl concentration in the aqueous phases, being always higher than 72%, whereas Pd(II) extraction decreases from 65% at 1 M HCl to 22% at 8 M HCl. Several stripping agents for the two metals were tested: Pt(IV) is successfully recovered by a 1 M sodium thiosulfate solution, whereas the best result for Pd(II) was achieved with 0.1 M thiourea in 1 M HCl. The loading capacity of DMDBMA for Pt(IV) is high, and data obtained from successive extraction-stripping cycles suggest a good DMDBMA stability pattern. Attempts to replace 1,2-DCE by more environmentally-friendly diluents showed, in general, worse %E for Pt(IV). The dependence of Pt(IV) distribution coefficients on DMDBMA and chloride ion concentrations, as well as on acidity, are the basis of a proposal for the composition of Pt(IV) extracted species.

Oxidative leaching process with cupric ion in hydrochloric acid media for recovery of Pd and Rh from spent catalytic converters

The recycling of platinum-group metals from wastes such as autocatalytic converters is getting growing attention due to the scarcity of these precious metals and the market pressure originated by increase of demand in current and emerging applications. Hydrometallurgical treatment of such wastes is an alternative way to the most usual pyrometallurgical processes based on smelter operations. This paper focuses on the development of a leaching process using cupric chloride as oxidising agent, in HCl media, for recovery of palladium and rhodium from a spent catalyst. The chloride media allows the adequate conditions for oxidising and solubilising the metals, as demonstrated by equilibrium calculations based on thermodynamic data. The experimental study of the leaching process revealed that Pd solubilisation is clearly easier than that of Rh. The factors temperature, time, and HCl and Cu(2+) concentrations were significant regarding Pd and Rh leaching, the latter requiring higher factor values to achieve the same results. Leaching yields of 95% Pd and 86% Rh were achieved under optimised conditions (T = 80 °C, t = 4h, [HCl] = 6M, [Cu(2+)] = 0.3M).

Liquid-Liquid Extraction of Platinum from Chloride Media by N,N′-Dimethyl-N,N′-Dicyclohexyltetradecylmalonamide

Liquid-liquid extraction of platinum(IV) from chloride media was carried out using N,N′-dimethyl-N,N′-dicyclohexyltetradecylmalonamide (DMDCHTDMA) in 1,2-dichloroethane. Platinum can be effectively extracted by DMDCHTDMA without addition of tin(II) chloride, since extraction percentages (%E) of 88% and 99% have been achieved from 6 M and 8 M HCl, respectively. Moreover, platinum can be successfully stripped through a simple contact with a 1 M HCl solution. The effect of some experimental parameters such as equilibration time, diluent, extractant and HCl concentrations was systematically investigated. The loading capacity of DMDCHTDMA was also evaluated. Data obtained from successive extraction-stripping cycles suggest a good stability pattern of DMDCHTDMA. Preliminary extraction data achieved with single metal ion solutions pointed out to a possible separation of platinum(IV) from palladium(II).

End-of-life Zn–MnO2 batteries: electrode materials characterization

Physical and chemical characterization of several sizes and shapes of alkaline and saline spent Zn–MnO2 batteries was carried out, aiming at contributing for a better definition of the applicable recycling processes. The characterization essays included the mass balance of the components, cathode and anode elemental analysis, the identification of zinc and manganese bearing phases and the morphology analysis of the electrode particles. The electrode materials correspond to 64–79% of the total weigh of the batteries, with the cathodes having clearly the highest contribution (usually more than 50%). The steel components, mainly from the cases, are also important (17–30%). Elemental analysis showed that the electrodes are highly concentrated in zinc (from 48–87% in anodes) and manganese (from 35–50% in cathodes). X-Ray powder diffraction allowed for identifying several phases in the electrodes, namely zinc oxide, in the anodes of all the types of saline and alkaline batteries tested, while zinc hydroxide chloride and ammine zinc chloride only appear in some types of saline batteries. The manganese found in the cathode materials is present as two main phases, MnO·Mn2O3 and ZnO·Mn2O3, the latter corroborating that zinc migration from anode to cathode occurs during the batteries lifespan. A unreacted MnO2 phase was also found presenting a low crystalline level. Leaching trials with diluted HCl solutions of alkaline and saline battery samples showed that all zinc species are reactive attaining easily over than 90% leaching yields, and about 30% of manganese, present as Mn(ii/iii) forms. The MnO2 phase is less reactive and requires higher temperatures to achieve a more efficient solubilization.

Separation of Cadmium, Cobalt, and Nickel by Solvent Extraction using the Nickel Salts of the Extractants

A new solvent‐extraction process for the separation of cadmium, cobalt, and nickel in sulphate solutions coming from the hydrometallurgical processing of spent Ni‐Cd batteries is proposed. The main innovation is to use nickel salts of the extractants, thus avoiding external pH control in the extraction operation. The extractants are first loaded with nickel in conditioning steps, using a neutralizer for pH control, and afterwards contacted with the aqueous processing solutions for extraction of interested metals with no further need of neutralization. This process is an alternative to the usual approach, which uses the sodium or ammonium salts of the extractants, avoiding introducing these cations in the process stream. Using this approach, the extraction of cadmium with nickel salt of 1 M DEHPA was performed at resulting pH values of 3.8–4.3 producing an organic phase loaded with 35 g/L Cd. Cobalt extraction with the nickel salt of Cyanex 272 was further achieved at resulting pH of 5.1–5.7 obtaining a organic loaded with 6.5 g/L Co.

Hydrometallurgical recycling of lithium-ion batteries by reductive leaching with sodium metabisulphite

The hydrometallurgical extraction of metals from spent lithium-ion batteries (LIBs) was investigated. LIBs were first dismantled and a fraction rich in the active material was obtained by physical separation, containing 95% of the initial electrode, 2% of the initial steel and 22% of plastic materials. Several reducers were tested to improve metals dissolution in the leaching step using sulphuric acid. Sodium metabisulphite led to the best results and was studied in more detail. The best concentration of Na2S2O5 was 0.1 M. The metals dissolution increased with acid concentration, however, concentrations higher than 1.25 M are unnecessary. Best results were reached using a stirring speed of 400 min-1. The metals leaching efficiency from the active material (Li, Mn, Ni, Co) increased with the temperature and was above 80% for temperatures higher than 60 °C. The dissolution of metals also rose with the increase in the liquid/solid ratio (L/S), however, extractions above 85% can be reached at L/S as lower as 4.5 L/kg, which is favourable for further purification and recovery operations. About 90% of metals extraction can be achieved after only 0.5 h of leaching. Sodium metabisulphite can be an alternative reducer to increase the leaching of Li, Mn, Co, and Ni from spent LIBs.

Scrubbing of Cadmium and Nickel from Cyanex 272 Loaded with Cobalt

Abstract A process for the scrubbing of cadmium and nickel from Cyanex 272 loaded with cobalt was studied and is presented in this paper, as a part of a global solvent extraction process proposed for the recycling of spent Ni‐Cd batteries. After cadmium recovery in a first extraction circuit, the resulting solution containing about 80 g/L Ni, 1 g/L Co, and a residual Cd concentration of 0.10 g/L is processed in the second circuit where cobalt is extracted by 0.5 M Cyanex 272. The resulting organic phase containing about 7 g/L Co is contaminated with about 2 g/L Ni and 0.5 g/L Cd. The removal of nickel was successfully attained by scrubbing with cobalt sulphate solutions containing more than 8.7 g/L Co, at organic/aqueous (O/A) = 4 (95–98% efficiency in a single stage), while cadmium removal by this process was inefficient. The scrubbing of cadmium was achieved with high efficiency using aqueous complexing agents, namely 1 M sodium thiosulphate or 3 M ammonium chloride (scrubbing efficiencies of 97% and 80%, respectively, in a single stage at O/A = 4). The selectivity Cd/Co observed when using these two inorganic ligands was very promising since only about 0.01% or 0.14% of cobalt was scrubbed. An overall countercurrent multistage system was proposed consisting on two scrubbing operations: (1) scrubbing with a cobalt sulphate solution (8.7 g/L Co) for Ni removal and (2) scrubbing with 1 M sodium thiosulphate or 3 M ammonium chloride for Cd removal. The overall scrubbing efficiencies were respectively, 98.6% for Ni and 99.1% or 97.9% for Cd. The final organic phase contained 9.1 g/L Co, 0.03 g/L Ni and 0.005 or 0.012 g/L Cd, anticipating the recovery of a cobalt product with good purity.

Kinetic approach to the study of froth flotation applied to a lepidolite ore

The number of published studies related to the optimization of lithium extraction from low-grade ores has increased as the demand for lithium has grown. However, no study related to the kinetics of the concentration stage of lithium-containing minerals by froth flotation has yet been reported. To establish a factorial design of batch flotation experiments, we conducted a set of kinetic tests to determine the most selective alternative collector, define a range of pulp pH values, and estimate a near-optimum flotation time. Both collectors (Aeromine 3000C and Armeen 12D) provided the required flotation selectivity, although this selectivity was lost in the case of pulp pH values outside the range between 2 and 4. Cumulative mineral recovery curves were used to adjust a classical kinetic model that was modified with a non-negative parameter representing a delay time. The computation of the near-optimum flotation time as the maximizer of a separation efficiency (SE) function must be performed with caution. We instead propose to define the near-optimum flotation time as the time interval required to achieve 95%–99% of the maximum value of the SE function.