Pollution Control and Resource Reuse for Alkaline Hydrometallurgy of Amphoteric Metal Hazardous Wastes

Various wastes containing zinc and lead are generated in industries such as galvanizing, casting, scrap recycling, smelting, and pyrometallurgical and hydrometallurgical process, as forms of dust, tailings, residues, sludge, and lower-grade lean ores. The disposals of these wastes are now becoming expensive due to the need for the treatment to render the wastes nonhazardous. The options currently available can be comprehensively classified as security landfilling and pyrometallurgical and hydrometallurgical processes. As the wastes containing zinc and lead have been classified as hazardous, conventional landfilling processes should be modified to security landfilling in order to meet the environmental constraints required for the hazardous. The pyrometallurgical processes to treat hazardous wastes containing zinc and lead include Inred, Wala Kiln, Plasma, etc., in which zinc and lead can be easily extracted due to the high temperature used, though a large amount of energy will be consumed and serious secondary pollutions arise. The hydrometallurgical treatment method for wastes containing zinc and lead is by dissolution in mineral acids and alkaline solutions. For acidic leaching process, though zinc and lead will be dissolved to an acceptable high level, the bulk materials, iron, calcium, etc., will also be dissolved completely, and the dissolved iron and other elements have to be precipitated from the leach solutions, possibly leading to the generation of new hazardous wastes and wastewaters. Moreover, a big fraction of zinc exists as zinc ferrites in the dust, which cannot be attacked effectively by acidic leaching processes. Therefore, the acidic leaching process seems not to be economically viable for the treatment of dust. The alkaline process is considered to be a cleaner technology for extracting zinc and lead from the hazardous waste bearing zinc and lead and will be fully introduced in this book.

The thermodynamics of alkaline leaching of solid wastes bearing zinc or lead is carried out for the hydrometallurgy of zinc or lead hazardous wastes. The E-pH equilibrium diagrams of leaching systems of zinc or lead in different forms are drawn. In strong alkaline solution, most of zinc ions exist in the form of \( \mathrm{Zn}{\left(\mathrm{OH}\right)}_4^{2-} \). ZnO, ZnCO₃, and Zn₂SiO₄ are the predominant mineralogical phases in solid wastes bearing zinc. ZnO and ZnCO₃ can be dissolved directly by strong alkaline solution, while higher concentration of OH⁻ is required to dissolve Zn₂SiO₄. PbO, PbS, PbCO₃, and PbSO₄ are the predominant mineralogical phases in solid wastes bearing lead. PbO, PbSO₄, and PbCO₃ can be dissolved in concentrated NaOH solutions, while the dissolution of PbS may be negligible. The equilibrium concentrations of Cd in leach solutions at 180 ~ 240 g/L NaOH is about 0.45 ~ 0.79 g/L. The dissolution of other impurities such as Fe, Cu, Co, Ni, Mg, Ca, etc. can be negligible.

The leaching reaction kinetics of waste bearing zinc and lead are studied. For waste bearing zinc, plot of 1 − (1 − η)^1/3 versus leaching time giving linear relationship indicates that the velocity of leaching is controlled by the chemical reaction step; the apparent activation energy is obtained to be 49.22 kJ/mol from an Arrhenius plot. For smithsonite and willemite, during the early stage of leaching reaction, 1 − (2/3)η − (1 − η)^2/3 is linear with leaching time, with the apparent activation energy of 19.95 kJ/mol and 19.32 kJ/mol, respectively. In the later stage, 1 − (1 − η)^1/3 is linear with leaching time, with the apparent activation energy of 46.54 kJ/mol and 32.64 kJ/mol, respectively. Thus, it indicates that the leaching mechanism of smithsonite changes from inner diffusion control to chemical reaction control and that of willemite changes from inner diffusion control to mixed control, during the leaching process in alkaline solution. The leaching of lead oxide may be controlled by diffusion of NaOH solution passing through the solid-phase layer at lower temperature and less NaOH concentration and by surface chemical reaction at high temperature and high NaOH concentration.

ZnO, ZnCO₃, and Zn₂SiO₄ are the predominant mineralogical phases in solid wastes bearing zinc. ZnO, ZnCO₃, PbO, PbCO₃, and PbSO₄ can be easily dissolved directly by strong alkaline solutions. However, quite a big portion of zinc in the dust exists as zinc ferrites or zinc sulfide and lead as lead sulfide which cannot be dissolved completely in alkaline solution. Integrated hydrometallurgical processes for leaching the zinc and lead from zinc ferrites, zinc sulfide, and lead as lead sulfide in alkaline medium are developed. For zinc ferrites in EAF dust, the zinc is extracted by direct leaching-melting-leaching of melt or hydrolysis-melting-leaching processes. For the zinc sulfide in solid wastes, the leaching rate of zinc in alkaline solution is improved via chemical conversion with PbCO₃ so that the zinc in zinc sulfide can be extracted in NaOH solution with lead carbonate as additive, and the lead sulfide in the leach residues can be converted to lead carbonate by reacting with sodium carbonate solution using air as the oxidizing agent, and then the lead can be recycled in the whole process by dissolving the lead carbonate in NaOH solution. Over 90% of zinc can be extracted from the zinc sulfide, and over 95% of lead sulfide in leach residues can be converted. The Pb in leaded glass can be extracted in NaOH solution after mechanochemical reduction with metallic iron as additive. Over 90% of Pb can be extracted from the mechanochemical-reduced leaded glass, compared with less than 5% of Pb extraction for alkaline leaching of nonactivated samples, and 67% Zn extraction for alkaline leaching of activated samples for 2 h by planetary ball mill. Alkaline leaching process of smoke dust and lead oxide ore is conducted, which found that the leaching rate was over 90% after 30 min or longer. The resultant leaching residue contains lower than 1–2% of zinc, 0.5% of lead, 0.3% of copper, and 0.1% of cadmium, and over 35% of Fe, and may be classified as nonhazardous waste according to the results of leaching tests on the residues.

Zinc and lead are usually concomitantly present in zinc and lead hazardous wastes such as electric arc furnace (EAF) dust. The selective and quantitative separation of lead from zinc is an important step in these alkaline hydrometallurgy processes. The separation of lead from zinc in alkaline zinc solution, in leaching solutions of zinc and lead hazardous wastes, can be achieved by sodium sulfide precipitation. When the weight ratio of sodium sulfide (average molecular weight 222) added to the lead present in alkaline solution is over 1.8–2.0 (molar ratio approximately 1.5–1.7), lead could be separated selectively and quantitatively, while the zinc remained in the solution without concomitant loss. The residues from the precipitation step are identified as mixtures of PbS and Na₂Pb(OH)₂S, with PbS being the predominant compound. Sn can be cemented quantitatively from alkaline zinc solution by the addition of zinc powder at 95 °C, 500 rpm with Zn/Sn molar ratio 6. Al in the leaching solution can be removed by Na₂SiO₃, As can be removed by Fe₂(SO₄)₃ at low NaOH concentration, Cl⁻ can be removed by increasing NaOH content, and Cu can be removed from alkaline lead leaching solution by Pb power replacement.

Electrowinning of zinc and lead from alkaline solutions depends on zinc or lead concentrations, temperature, current density, NaOH concentration, electrode distance and electrode material on current efficiency, and energy consumption of zinc or lead electrowinning. The optimum conditions for zinc electrowinning are zinc concentration 30–40 g/L, current density 800–1000 A/m², NaOH concentration 180–200 g/L, and temperature 30–50 °C, and those for lead electrowinning are lead concentration 20 g/L, 5 g/L lead concentration left in the spent electrolyte, below 50 °C, 400 A/m² current density, 5 mol/L NaOH concentration, and stainless steel of electrode material. Impurities ions have great effect on the quality of the metallic zinc or lead powder in alkaline leaching electrowinning process. The electrolysis deteriorates when the concentration of arsenide is over 500 mg/L, but \( {\mathrm{CO}}_3^{2-} \), \( {\mathrm{SO}}_4^{2-} \), \( {\mathrm{SiO}}_3^{2-} \), and F^- have no effects on zinc electrowinning. Copper will be electrodeposited prior to lead in the cathode. Tin, \( {\mathrm{C}}_4{\mathrm{H}}_4{\mathrm{O}}_6^{2-} \), \( {\mathrm{CO}}_4^{2-} \), and gluten in the solution decrease the current efficiency of the lead electrowinning process, while As, Sb, and W, as well as \( {\mathrm{SO}}_4^{2-} \), \( {\mathrm{PO}}_4^{3-} \), \( {\mathrm{SO}}_3^{2-} \), \( {\mathrm{SiO}}_3^{2-} \), SCN⁻, \( {\mathrm{CO}}_3^{2-} \), Cl⁻, and AC⁻, have no effects.

Low-grade smithsonite ores are leached using alkaline leaching solution. The zinc and lead concentrates can be obtained from lean oxidized zinc ores by alkaline leaching followed by two-step sulfide precipitation of lead and zinc sulfides. It is found that over 85% of both Zn and Pb and less than 10% of Al can be leached from the ore when the leaching operation is conducted at over 95 °C using 5 M NaOH solution as leaching agent. The dissolution of impurities such as Fe, Ca, etc. is negligible. Leaching of Pb can be improved remarkably with addition of NaCl to the leaching systems. Optimum conditions of lead precipitation are found to be 1.8 Na₂S/Pb weight ratio, 90 °C, 60 min, and those for zinc precipitation are 2.4 Na₂S/Zn weight ratio, 90 °C, 180 min. According to scale-up experiment, total extractions of both lead and zinc reach above 80% from the ores, and the quality of lead and zinc concentrates can meet industrial requirements for roasting process. The Pb-free solution is then used for the electrolysis of metallic Zn using stainless steel electrodes. Zn metal powder with purity higher than 99.5% is obtained. The specific energy for the Zn electrolysis in alkaline leach solution is around 2.4–2.6 kWh/kg Zn, which is lower than the energy consumption of 3.3 kWh/kg Zn required in the conventional process of Zn electrolysis in acidic sulfate electrolytes.

The impurities in the electrowinning solution can be accumulated during the cycle of alkaline solutions. The integrated process of regeneration and purification of spent electrolyte by causticization is developed. The addition of solid NaOH can be used to separate Na₂CO₃ in the spent electrolyte. The Na₂CO₃ is causticized using CaO to regenerate NaOH. During this separation and causticization process, the impurity ions in the spent electrolyte are removed simultaneously. The causticized alkaline solution is recycled to the next leaching step. Moreover, Mo, W, Nb, Ta, etc., will be also dissolved in alkaline solution under the leaching conditions of zinc and lead hazardous wastes. The Mo and W can be recovered with ion flotation and adsorbing colloid flotation process using primary amine dodecylamine (RNH₂) as collector and ferric ions or magnesium chloride as co-precipitants or depressant (modifier). Adsorbing colloid flotation (ACF) and ion flotation (IF) using ferric ions as the co-precipitant or sorbent and sodium dodecyl sulfate as the collector can be used for the effective removal of arsenic and molybdenum from aqueous solution. The synergistic extraction of phosphorus, arsenic, and silica from tungstate and molybdate solutions with primary amines and TBP can be realized. Mixed solutions of 12-molybdophosphoric acid and its reduced molybdenum blue solutions are found to be an effective scrubbing process for the combined removal of H₂S, SO₂, and NO_x from gas streams. When the molar ratio of NO₂/(SO₂ + H₂S) in the waste gases is just equal to 1/2, H₂S, SO₂, and NO_x will be removed quantitatively, while the regeneration of scrubbing solution is not needed. H₂S and SO₂ react with molybdophosphoric acid solution to form elemental sulfur and H₂SO₄, respectively, while molybdophosphoric acid is reduced into blue species, molybdenum blue solution. The reactions of H₂S are rather quickly, while those of SO₂ are much more slowly. The resulting molybdenum blue solution is used further for the removal of NO_x, in which molybdenum blue is rapidly oxidized back into molybdophosphoric acid and NO_x is reduced into N₂. Porous white tungstic acid (white tungstic acid), whose composition is firstly revealed in this work, is used to produce new tungstoniobates and tungstotantalates and a series of heteropoly compounds. Moreover, a quick determination method of tungsten concentration is proposed according to the quantitative relationship between the precipitate of white tungstic acid and pH value of a certain range.

An integrated process consisted of raw materials milling and grinding, leaching, purification, electrowinning, regeneration of the spent electrolytes, vacuum drying, analysis, etc. is developed, and the industrial application flow sheet is designed, with annual yield 1500–2000 tons metallic zinc powder. The process developed has been successfully applied at several plants with a scale of 1000, 1500, and 2000 tons of metallic zinc powder annually. The long-term industrial operations have been shown that the process can use any raw materials containing Cl, F, Si, Fe, As, Cu, etc. The total Zn content of zinc powder produced is over 98% and the metallic zinc content over 94%, which reaches the Class 2 of the national standards of zinc powder quality. Hence, the process is quite different to the acidic process and can apply the raw materials unacceptable for the acidic process. The life cycle assessment method is used to investigate the environmental impacts of the new alkaline processes of production of Zn powder. The results indicate that the environmental impacts of alkaline leaching-electrowinning technology are less than the conventional acid leaching hydrometallurgical processes and pyrometallurgical processes. The energy consumption, greenhouse effect, and acidification in the new processes are 87.76%, 47.80%, and 0.08% of those in the conventional acid leaching hydrometallurgical processes and are 93.94%, 9.25%, and 0.14% of those in the pyrometallurgical processes.