Investigations into complex processing of zinc cakes

Investigations are carried out into the processing of zinc cakes with the purpose of transferring zinc, copper, and iron from ferrites into a solution and of concentrating noble metals in the lead-containing silicate product. Zinc cakes are subjected to sulfatization by oleum followed by leaching with a concentrated H₂SO₄ solution. With the use of a planned multifactor experiment according to the Box plan, the effect of the H₂SO₄ content in the leaching solution on the cake leaching process, its duration, and temperature has been studied. Due to this, the models of the dependence of the listed factors on the extraction of Zn, Fe, and Cu into the solution have been obtained. The possibility of an almost complete extraction of these metals into the solution is demonstrated. The following leaching conditions are suggested as optimal: H₂SO₄ concentration of 8.33 g/l in the leaching and washing solutions, leaching time of 2 h, and leaching temperature of 75°C.     

Leaching of vanadium from carbonaceous shale

The leaching of vanadium from carbonaceous shale using dilute H₂SO₄ was investigated, and the mechanism of leaching determined. The results showed that higher leaching efficiency of vanadium was obtained by increasing initial concentration of H₂SO₄, raising leaching temperature and prolonging leaching time. Addition of ammonium fluoride also enhanced recovery. A recovery of 92% was obtained using a liquid to solid ratio 4:1, initial H₂SO₄ concentration 18%, NH₄F addition 4.8 wt.% of carbonaceous shale, leaching temperature 95 °C and contact time 8 h; the recovery was only 56% without NH₄F. The presence of NH₄F enhanced the leaching of vanadiferous mica.     

Dissolution kinetics of malachite in ammonia/ammonium carbonate leaching

The leaching of oxide copper ore containing malachite, which is the unique copper mineral in the ore, by aqueous ammonia solution has been studied. The effect of leaching time, ammonium hydroxide, and ammonium carbonate concentration, pH, [NH₃]/[NH₄⁺] ratio, stirring speed, solid/liquid ratio, particle size, and temperature were investigated. The main important parameters in ammonia leaching of malachite ore are determined as leaching time, ammonia/ammonium concentration ratio, pH, solid/liquid ratio, leaching temperature, and particle size. Optimum leaching conditions from malachite ore by ammonia/ammonium carbonate solution are found as ammonia/ammonium carbonate concentrations: 5 M NH₄OH+0.3 M (NH₄)₂CO₃; solid/liquid ratio: 1:10 g/mL; leaching times: 120 min; stirring speed: 300 rpm; leaching temperature: 25 °C; particle size finer than 450 μm. More than 98% of copper was effectively recovered. During the leaching, copper dissolves as in the form of Cu(NH₃)₄⁺² complex ion, whereas gangue minerals do not react with ammonia. It was determined that interface transfer and diffusion across the product layer control the leaching process. The activation energy for dissolution was found to be 15 kJ mol⁻¹.     

Coordination-reduction leaching process of ion-adsorption type rare earth ore with ascorbic acid

The magnesium sulfate(MgSO₄)-ascorbic acid(Vc) compound leaching technique can extract rare earth elements(REEs) existing in ion-exchangeable phase and colloidal phase from ion-adso rption type rare earth ore through the synergy effect of coordination and reduction,but its reaction process and mechanism remain unclear.In this paper,the coordination-reduction leaching mechanism was analyzed from the perspectives of leaching thermodynamics and kinetics,which provide theoretical guidance...     

Leaching of oxidized zinc ores in various media

The leaching of various oxidized zinc ores in different solvents has been studied in laboratory tests by agitation leaching and in some cases by percolation leaching. Some tests have also been carried out on synthetic zinc compounds.The comparison of the leaching of four different Belgian ores in sulfuric acid, sulfurous acid, ammonium hydroxide and sodium hydroxide, shows that the better results are obtained with sulfuric acid and with caustic soda when the concentration of the latter is high enough. All the ores respond very well to leaching except one which is very rich in iron and can be processed only with concentrated caustic soda. Mineralogical studies have shown that smithsonite is completely leached but that hemimorphite is refractory to leaching in any solvent studied.The leaching in alkaline media (ammonia, caustic soda, diethylenetriamine) has been studied on five different ores. It is confirmed that hemimorphite is difficult to dissolve. Good results are obtained only with caustic soda at high temperature. In ammoniacal solutions, the presence of ammonium carbonate enhances the leaching of hemimorphite but decreases the leaching of smithsonite. The optimum concentration of diethylenetriamine is 250 g/l.Caustic soda leaching tests were carried out on pure compounds and showed that ZnO, Zn(OH)₂, PbO, PbCO₃ and 2PbCO₃βPb(OH)₂ dissolve very rapidly and completely, that the leaching of Zn₂SiO₄ is controlled by kinetics and that the leaching of ZnS, PbS and PbSiO₄ is very poor, their solubility product being very quickly attained.The percolation leaching by sulfuric acid has been tested on Belgian ores and the good results obtained support the idea that this kind of ore could be successfully treated in this way.     

High-efficiency simultaneous extraction of rare earth elements and iron from NdFeB waste by oxalic acid leaching

Iron can not be recovered at high value because only rare earth elements are effectively recovered from NdFeB waste via oxidation roasting-hydrochloric acid leaching process.In this study,a new method for leaching NdFeB waste with oxalic acid was developed.The high-efficiency,simultaneous and high-value recovery of rare earth elements and iron was realized to simplify the process and improve the economic benefit.Results of the oxalic acid leaching experiments show that under the optimum leaching conditions at 90℃ for 6 h in the aqueous solution of oxalic acid(2 mol/L) with a liquid-solid ratio of60 mL/g,the iron leaching efficiency and precipitation rate of rare earth oxalate reach 93.89% and 93.17%,respectively.Rare earth oxalate and Fe(C₂O₄)3^3- were left in the residue and the leaching solution,respectively.The leaching mechanism was further analyzed by characterising the leach residues obtained through X-ray powder diffraction(XRD) and scanning electron microscopy-energy dispersive X-ray spectroscopy(SEM-EDS).Results of the leaching kinetics study indicate that the process of oxalic acid leaching follows the shrinking nucleus model,and the leaching kinetics model is controlled by the mixed factors of diffusion and chemical reaction.The leaching residue was calcined at 850℃ for 3 h and then decomposed into rare earth oxide,which can be directly used to prepare rare earth alloy via molten salt electrolysis.For the leaching solution,ferric oxalate solution was reduced using Fe powder to prepare the ferrous oxalate(FeC₂O₄-2H₂O).     

Effect of roasting activation of rare earth molten salt slag on extraction of rare earth,lithium and fluorine

A new process was proposed to extract rare earth elements(REEs),Li and F from electrolytic slag of rare earth molten salt by synergistic roasting and acid leaching.Firstly,the thermodynamic analysis of roasting reaction was carried out,then the effects of roasting factors on leaching REEs,Li and F in slag were investigated.In additions,the mineral phase and morphology of molten salt slag,roasting slag and acid leaching slag were characterized,and the migration mechanism of REES,Li and F minerals...     

Leaching of copper(I) sulphide by sulphuric acid solution with addition of sodium nitrate

Finely grained samples of copper(I) sulphide were leached by H₂SO₄ solution with added NaNO₃. The occurrence probability of chemical reactions was analysed based on literature data and products which were formed during the process and the overall leaching reaction was defined. The effect of temperature, concentration of NaNO₃ and H₂SO₄, stirring speed, phase ratio and time, on the leaching degree of copper was studied. The quantity of copper dissolved increases with growth of the values of all the parameters. Kinetic analysis shows that the leaching mechanism is very complex. By using appropriate mathematical kinetic models, it is found that the leaching rate is chemically controlled. It was concluded that the leaching reaction is first order with respect to the concentration of NaNO₃ and second order with respect to the concentration of H₂SO₄.     

Mineralogical changes occurring during the ferric lon leaching of bornite

The reaction products formed during the leaching of bornite in either ferric chloride or ferric sulfate media depend on the leaching conditions as well as the particle size of the bornite. The extent of dissolution is always more vigorous in the ferric chloride system and increases with increasing temperature in either system. The reaction initially involves the rapid outward diffusion of copper to form slightly nonstoichiometric bornite (Cu_5-xFeS₄), chalcopyrite, and covellite. The non-stoichiometric bornite is progressively converted to a Cu₃FeS₄ phase, which varies considerably in its composition, and to covellite. Although the reaction at low temperature terminates at the Cu₃FeS₄ phase, leaching at higher temperatures results in further dissolution to elemental sulfur and soluble Cu²⁺ and Fe²⁺. The leaching ofmassive bornite illustrates the complexities of the leaching reaction more clearly than is observed for the finelypaniculate bornite. In leached massive bornite, a distinct covellite zone appears in the Cu₃FeS₄ phase; as well, chalcopyrite exsolution lamellae rimmed by a copper sulfide (possibly digenite) appear in the covellite zone, in the Cu₃FeS₄ phase, and in the nonstoichiometric bornite. The experimental leaching results, especially those involving massive bornite, are generally consistent with the mineralogical trends produced by supergene alteration of bornite ores, but a significant difference is that the Cu₃FeS₄ phase does not correspond closely to the mineral idaite.     

Selective Leaching Process for Neodymium Recovery from Scrap Nd-Fe-B Magnet

Neodymium-iron-boron (Nd-Fe-B) magnets were most widely applied to permanent magnetic products in the world due to their high magnetic force. The increasing growth of scrap Nd-Fe-B magnets resulted in disposal problems and the reduction of neodymium (Nd) valuable resources. In this study, we developed a simple hydrometallurgical precipitation process with pH adjustment to separate and recover Nd 100 pct recovery from scrap Nd-Fe-B magnets. Several physical and chemical methods such as demagnetization, grinding, screening, and leaching processes were also adopted to investigate the recovery of Nd and other metals from scrap Nd-Fe-B magnets. The leaching process was carried out with four leaching reagents such as NaOH, HCl, HNO₃, and H₂SO₄. Batch studies were also conducted to optimize the leaching operating conditions with respect to leaching time, concentration of leaching reagent, temperature, and solid/liquid ratio for both HCl and H₂SO₄ leaching reagents. Nd was successfully separated and recovered with 75.41 wt pct from optimized H₂SO₄ leaching solution through precipitation. Further, the purity and weight percentage of the obtained Nd product was analyzed using scanning electron microscopy–energy-dispersive spectroscopy (SEM-EDS) analysis. An X-ray diffraction (XRD) study confirmed the obtained product of Nd was in the form of NdOOH and Nd(OH)₃.     

Extraction of scandium from red mud by acid leaching with CaF_2 and solvent extraction with P507

The extraction of Sc by acid leaching with CaF_2 and solvent extraction with P507 from red mud was proposed.The influence of acid leaching and solvent extraction on recovery of Sc was investigated.The CaF_2 can obviously improve the leaching efficiency of Sc and reduce the acid consumption.The leaching efficiency of Sc increases from 74% to 92% and the dosage of acid reduces under suitable conditions by adding 5% CaF_2.The minerals in red mud can easily be decomposed and leached into the acid solution with CaF_2 through analysis of XRD pattern.The particles of red mud become smaller and multihole.The Sc can be selectively extracted with 10% P507 at the pH value of 0.1 from the acid leaching solution.More than 98% of Sc and less than 10% of Al and Fe are extracted.The SC_2O_3 with purity of 99% is obtained after the process of reverse extraction with NaOH,H_2SO_4 dissolution,precipitation by oxalic acid and roasting at 750℃.     

Kinetics of silver leaching from manganese-silver associated ores in sulfuric acid solution in the presence of hydrogen peroxide

Kinetics of silver leaching from a manganese-silver associated ore in sulfuric acid solution in the presence of H₂O₂ has been investigated in this article. It is found that sulfuric acid and hydrogen peroxide have significant effects on the leaching rate of silver. The reaction orders of H₂SO₄ and H₂O₂ were determined as 0.80 and 0.68, respectively. It is found that the effects of temperature on the leaching rate are not marked, the apparent activation energy is attained to be 8.05 kJ/mol within the temperature range of 30 °C to 60 °C in the presence of H₂O₂. Silver leaching is found to be diffusion-controlled and follows the kinetic model: 1−2x/3−(1−x)^2/3=Kt. It is also found that particle size presents a clear effect on silver leaching rate, and the rate constant (k) is proportional to d ₋₂ ⁰ .     

Simulation ofin situ uraninite leaching-part III: The effects of solution concentration

The effects of variations in the concentrations of leaching reagents have been simulated forin situ leaching of UO₂ by H₂O-(NH₄)₂CO₃-NH₄HCO₃. The model used in the simulations incorporates rate laws for the mineral reactions, equilibrium reactions among the solution species, and a mixing cell representation of solution flow. Of the component concentrations, the major factor affecting the rate of uraninite dissolution is the oxidant concentration. High peroxide concentrations lead to more rapid reaction with an early maximum in the U(VI) concentration. If lower oxidant concentrations are used, the reaction is under mixed kinetic and mass transfer control and the U(VI) concentration is lower but approximately constant for an extended period. Because they increase the concentration of the HCO _3/- anion, high ammonium carbonate and ammonium bicarbonate concentrations also result in some enhancement in the rate of U leaching; the reaction is known to be half-order in both HCO₃ ^- and H₂O₂. A 10:1 ratio of (NH₄)₂CO₃ to NH₄HCO₃ concentrations was found to result in a nearly constant pH during most of the leaching process. Calcite-containing gangue causes an immediate pH increase from about 8.9 to 9.4. The rate of the calcite reaction, calcite saturation index, and porosity are all independent of the lixiviant concentrations. Detailed calculations of solution speciation are necessary to predict the concentrations of individual species from those of components.     

Effect of solution composition on the optimum redox potential for chalcopyrite leaching in sulfuric acid solutions

The leaching rate of chalcopyrite (CuFeS₂) by Fe³⁺ in H₂SO₄ solutions depends on the redox potential determined by the Fe³⁺/Fe²⁺ concentration ratio, and there is a maximum leaching rate at an optimum redox potential. The present study investigated the effects of solution composition on the optimum redox potential by electrochemical measurements using a CuFeS₂ electrode and electrolyte solutions containing 0.01–1 kmol m^− 3 of H₂SO₄, Fe²⁺, and Cu²⁺ at 298 K in nitrogen.Anodic-polarization curves of the CuFeS₂ electrode showed that there was a current peak on the curves in the presence of Cu²⁺ and Fe²⁺, corresponding to the maximum leaching rate. The redox potential of the peak increased markedly with increasing Cu²⁺ concentration, while it was little affected by the H₂SO₄ and Fe²⁺ concentrations. These results agree with the results of leaching experiments reported previously, and indicate that the optimum redox potential for chalcopyrite leaching is a function of the Cu²⁺ concentration. An empirical equation for the optimum redox potential for CuFeS₂ leaching is proposed.     

Leaching of a sphalerite concentrate with H2SO4–HNO3 solutions in the presence of C2Cl4

The enhanced leaching of sphalerite concentrates in H₂SO₄–HNO₃ solutions and the extraction of sulfur with tetrachloroethylene were studied. Variables of the process were investigated including leaching temperature, reaction time, liquid / solid ratio, and tetrachloroethylene concentration. The number of cycles that tetrachloroethylene could be recycled did not have a significant effect on zinc extraction. The results indicated that 99.6% zinc extraction was obtained after three hours of leaching at 85 °C and 0.1 MPa O₂, when 20 g of sphalerite concentrate were leached in a 200 ml solution containing 2.0 mol/L H₂SO₄ and 0.2 mol/L HNO₃, in the presence of 10 ml C₂Cl₄. Leaching rates were significantly improved under these conditions.     

Leaching of a sphalerite concentrate with H2SO4–HNO3 solutions in the presence of C2Cl4

The enhanced leaching of sphalerite concentrates in H₂SO₄–HNO₃ solutions and the extraction of sulfur with tetrachloroethylene were studied. Variables of the process were investigated including leaching temperature, reaction time, liquid / solid ratio, and tetrachloroethylene concentration. The number of cycles that tetrachloroethylene could be recycled did not have a significant effect on zinc extraction. The results indicated that 99.6% zinc extraction was obtained after three hours of leaching at 85 °C and 0.1 MPa O₂, when 20 g of sphalerite concentrate were leached in a 200 ml solution containing 2.0 mol/L H₂SO₄ and 0.2 mol/L HNO₃, in the presence of 10 ml C₂Cl₄. Leaching rates were significantly improved under these conditions.     

Recovery of valuable metals from jarosite by sulphuric acid roasting using microwave and water leaching

In this paper, jarosite residue (JR) blended with concentrated H₂SO₄ was subjected to a process comprising microwave roasting and water leaching. The effects of H₂SO₄/JR weight ratio, microwave roasting temperature and time, water leaching conditions on the recovery of Fe, Zn, In, Cu, Cd, Ag and Pb were investigated utilising a series of experiments.

Based on energy conservation and environmental protection, optimum conditions for metals recovery from JR were determined as: H₂SO₄/JR weight ratio?=?0.36, microwave roasting temperature, 250°C; roasting time, 30?min; leaching temperature, 50°C; leaching time, 1?h; and liquid–solid ratio, 4:1 (mL/g), thus, the extraction of Fe, Zn, In, Cu, Ag and Cd were 89.4, 80.7, 85.1, 90.7, 61.3 and 48.8% respectively, while the Pb was concentrated in the final residue. Scanning electron microscope-energy dispersive spectrometer (SEM-EDS) patterns were used to characterise and analyse the transformation of valuable metals in the residue after roasting and leaching.     

The dissolution of sphalerite in ferric sulfate media

The dissolution of sphalerite, (Zn,Fe)S, in ferric sulfate media was investigated using closely sized fractions of crushed sphalerite crystals. Linear kinetics were observed, and the rate increased in proportion to the surface area, as the average particle size of the sphalerite decreased. The predominant reaction products are ZnSO₄, FeSO₄, and elemental sulfur. The leaching rate increases with increasing temperature, and the apparent activation energy is 44 kJ/mol. The relatively high apparent activation energy suggests that the rate is chemically controlled, a conclusion supported by the insensitivity of the rate of the rotation speed that was observed in complementary rotating disk experiments. The rate increases as the 0.3 to 0.4 power of the Fe(SO₄)_1.5 concentration, and is nearly independent of the pulp density, in the presence of a stoichiometric excess of ferric sulfate. In 0.3 M Fe(SO₄)_1.5 media, the rate increases with increasing acid concentrations >0.1 M H₂SO₄, but is insensitive to more dilute acid concentrations. In the absence of ferric ions, the rate increases rapidly with increasing H₂SO₄ concentrations, and relatively rapid rates are observed in solutions containing >0.5 M H₂SO₄. The rate decreases with increasing initial concentrations of ZnSO₄, MgSO₄, or FeSO₄ in the ferric sulfate leaching solution, and this emphasizes the importance of maintaining the dissolved iron in a fully oxidized state in a commercial leaching operation.     

Leaching kinetics of ionic rare-earth in ammonia-nitrogen wastewater system added with impurity inhibitors

Ammonia-nitrogen wastewater is produced during the dressing and smelting process of rare-earth ores.Such wastewater includes a very high concentration of NH4＋, as well as other ions（e.g., NH4＋, RE3＋, Al3＋, Fe3＋, Ca2＋, Cl–, and Si O32–） with a p H of 5.4–5.6.Its direct discharge will pollute, yet it can be recycled and used as a leaching reagent for ionic rare-earth ores.In this study, leaching kinetics studies of both rare earth ions and impurity ion Al3＋ were conducted in the ammonia-nitrogen wastewater system with the aid of impurity inhibitors.Results showed that the leaching process of rare-earth followed the internal diffusion kinetic model.When the temperature was 298 K and the concentration of NH4＋ was 0.3 mol/L, the leaching reaction rate constant of ionic rare-earth was 1.72 and the apparent activation energy was 9.619 k J/mol.The leaching rate was higher than that of conventional leaching system with ammonium sulfate, which indicated that ammonia-nitrogen wastewater system and the addition of impurity inhibitors could promote ionic rare-earth leaching.The leaching kinetic process of impurity Al3＋ did not follow either internal diffusion kinetic model or chemical reaction control, but the hybrid control model which was affected by a number of process factors.Thus, during the industrial production the leaching of impurity ions could be reduced by increasing the concentration of impurity inhibitors, reducing the leaching temperature to a proper range, accelerating the seepage velocity of leaching solution, or increasing the leaching rate of rare earths.     

Simultaneous recovery of rare earth elements and phosphorus from phosphate rock by phosphoric acid leaching and selective precipitation:Towards green process

Phosphate rock has been considered as one of the potential promising resources for rare earth elements(REEs). But the cost issues and the technical challenges caused by the low content of REEs in ores did hinder the further development of REEs recovery technologies. In order to explore a green process for the recovery of REEs from phosphate rock, this study investigates the effects of phosphoric acid concentration, liquid-to-solid ratio(L/S ratio), leaching time and temperature on the leaching efficiencies of the major components from phosphate rock. A REEs recovery of 94.3% and a phosphorus recovery of 95.3%are achieved under the optimal conditions of attacking phosphate rock using 30%P_2 O_5 acid with an L/S ratio of 10:1 and a stirring speed of 250 r/min at 25 ℃ for 4 h. Then,the selective precipitation of REEs with 81.3% REEs recovery is realized by heating up the leaching solution from 25 to 90 ℃ and keeping for4 h. Thereafter, more than 95% phosphoric acid is recovered by H_2 SO_4 and high purity gypsum, more than95% CaSO_4(tested by XRF), is also produced at the same time. Ultimately, a green process that leaches phosphate rock with H_3 PO_4, selectively precipitates REEs from leaching solution by heating up, recovers H_3 PO_4 with H2 SO4 is proposed. Compared with REE recovery in traditional processes, this process owns the merits of simple operation, energy saving and minimum wastes.