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This paper describes a highly alkaline low-grade copper oxide ore. Copper can be selectively leached out while other metals are retained. A thermodynamic model of the CuO-(NH4)2SO4-NH3-H2O system was established for the leaching of tenorite (CuO) under conditions of mass and charge conservation. MATLAB's fitting functions, along with the diff and solve functions, were used to calculate the optimal ammonia concentration and total copper ion concentration of tenorite under different ammonium sulfate concentrations. The effects of various ammonia–ammonium salt solutions (ammonium sulfate, ammonium carbonate, ammonium chloride) on the copper leaching rate were investigated. Results show that under the conditions of an ammonia concentration of 1.2 mol/L, an ammonia–ammonium ratio of 2:1, a liquid–solid ratio of 3:1, a temperature of 25 °C, and a leaching time of 4 h, the copper leaching rate from the ammonium sulfate and ammonium chloride solutions reaches 70%, which is slightly higher than that of ammonium carbonate. Therefore, an ammonia–ammonium sulfate system is selected for leaching low-grade copper oxide due to its lower corrosion to equipment compared to the chlorination system. The impact of this study on industrial applications includes the potential to find more sustainable and cost-effective methods for resource recovery. The industry can reduce its dependence on resources and mitigate its environmental impact. Readers engaged in low-grade oxidized copper research will benefit from this study. [ABSTRACT FROM AUTHOR]

The recycling of spent lithium NMC-type batteries, widely used in electric vehicles, presents a challenge due to manganese content, which complicates metal separation and purification. This study explored a selective leaching process using ammonia to recover metals from high-manganese-content LMR-NMC cathodes. By adjusting the (NH4)2SO4 reagent concentration to 1–2 M and maintaining the temperature between 50–80 °C, the recovery rates of lithium, nickel and cobalt metals were enhanced, leaving manganese primarily as residue in the form of Mn(OH)₂ and (NH4)2Mn(SO4)2. A kinetics model, integrating an equilibrium-shrinking core model with a modified temperature-dependent Arrhenius approach, accurately simulates the metal recovery. The activation energies of the forward leaching reactions of Li, Ni, and Co were respectively (1.4331±0.0036)×105, (1.5494±0.0034)×105, and (1.5743±0.0040)×105 J/mol, while those for the backward reactions were (5.3307±0.0041)×105, (2.4753±0.0093)×105, and (1.6289±0.0092)×105 J/mol, respectively. The leaching mechanism was found to be exothermic, which allows maximum recovery at low temperatures. The findings highlight ammonia’s effectiveness as a selective leachant, significantly reducing manganese in the leaching solution, and streamlining nickel and cobalt separation, thus enhancing the recycling process’s efficiency and sustainability.

In this work, a new hydrometallurgical process was developed to treat zinc ash produced from the hot galvanizing industry. The theoretical analysis shows the feasibility of dissolving zinc ash in the NH3–NH4Cl–H2O system, and the dissolution products are predominantly composed of Zn (NH3)42+. The impacts of different experimental conditions were examined, and the leaching ratio of zinc was as high as 96.4% under the conditions of NH3/NH4+ ratio of 1:1, liquid/solid of 9:1, total ammonia concentration of 8 mol/L and the stirring speed of 250 rpm at 313 K for 120 min. The kinetics of the leaching process were investigated and the calculated apparent activation energy was approximately 4.69 kJ/mol, which indicated that the zinc ash leaching process was controlled by diffusion-controlled. As revealed by the determination of impurity ions, on one hand, there were fewer impurities in the leaching solution, and the concentrations of Fe2+ and Pb2+ in solution are less than 0.02 mg/L and 0.05 mg/L respectively; on the other hand, there was no need for further impurity removal in this process. The proposed process has a certain application value in treating zinc ash. [ABSTRACT FROM AUTHOR]

Zinc-containing dust can be found in ironmaking and steelmaking, and it is an important secondary resource of zinc. Zinc-containing dust from an electric furnace was used as a raw material to study the phase transformation behavior of the dust using a calcification roasting process and the zinc–iron separation behavior by using ammonia leaching. The zinc-bearing dust was mixed with CaO and roasted to transform the zinc ferrite into zinc oxide. The results showed that increasing the calcium oxide to dust ratio could promote the conversion of zinc ferrite to zinc oxide. When the calcium oxide ratio reached 60%, the peak of zinc ferrite in the calcined-roasted product in the zinc-containing dust basically disappeared. As the temperature increased, the zinc oxide grains increased but were still smaller than 10 µm. The calcined-roasted product was crushed and ground, and the zinc was leached by ammonia. A zinc–iron recovery rate of 86.12% was achieved by the ammonia leaching. The leachate could be used for zinc extraction by electrolysis. The leaching residue was mainly calcium ferrate, which could be used in sintering production. The proposed process may achieve on-site recovery of zinc-containing dust in steel-making plants. [ABSTRACT FROM AUTHOR]

Zinc ferrite can be found in zinc-bearing dust in ironmaking and steelmaking. It is difficult to be recovered due to its stable properties. The magnetization roasting and ammonia leaching method were used to separate iron and zinc from zinc ferrite in this study. Thermodynamic analysis showed that the key to the selective reduction of zinc ferrite to zinc oxide and ferric oxide is to control the appropriate temperature and atmosphere. The influences of the selective reduction roasting process of zinc ferrite on the conversion rate, phase change, and microevolution behavior were investigated. The microstructure analysis showed that the distribution area of iron was mainly gray, and the distribution area of zinc was mainly white grid lines. The zinc content in the white area was higher than that in the gray area. With the increase in temperature and PCO, the white area expanded and the migration of zinc and iron was accelerated, but the iron in the white area still existed. The ammonia leaching of the magnetization-roasted product showed that a zinc leaching rate of 78.12% was achieved under the following conditions: the roasting atmosphere of PCO/P(CO+CO2) = 25%, the roasting temperature of 750 °C, roasting duration for 45 min, n(NH3-H2O):n(NH4Cl) = 1:1, the solid–liquid ratio of 40 g/L, leachate concentration of 6 mol/L, leaching duration of 90 min, the leaching temperature of 50 °C, and the stirring rate of 200 rpm. [ABSTRACT FROM AUTHOR]

Pyrite cinders enriched in iron, gold-silver, copper, and other valuable metals which have an important comprehensive recycling value. Recovering copper from pyrite cinders by sulfuric acid leaching will cause the loss of iron and increase the consumption of alkaline in subsequent cyanide process of gold and silver, leading to high production cost. In order to solve these problems, copper from pyrite cinders was recovered by ammonia leach, the effects of leaching time, slurry pH and other process parameters on copper leaching rate were studied. The results show that the effects of leaching time, slurry pH, liquid-to-solid ratio, agitation speed and leaching temperature on copper leaching were significant. Under the optimal leaching conditions of pH 10. 81, liquid-to-solid ratio 7 : 1, leaching time 1 h, agitation speed 600 r/min, temperature 20±2 °C, the extraction rate of copper was 82.66%, and the iron content in ammonia leaching residues was high, which was beneficial to pyrite cinders pellet for iron making. [ABSTRACT FROM AUTHOR]

Copper is one of the most important valuable elements in waste printed circuit boards (WPCBs), which can be finally recovered by hydrometallurgy. Environmentally-friendly copper leaching technology has become critical. In this study, two environmentally friendly leaching systems, namely ammonia-ammonium chloride leaching system and glycine-hydrogen peroxide leaching system, on copper leaching behavior was concerned, the influence of leaching time and leaching agent concentration on leaching rate of copper concentrate with different size was revealed. In addition, a simple but effective leaching kinetics model were proposed. The results show that with the extension of time, the leaching rate increases rapidly at first (0–10 min) and then slowly (10–25 min) until stable. The optimum leaching concentration of ammonia is 3 mol/L. For ammonia leaching system, the leaching rate increased from 76.92% to 90.05% with the decrease of particle size. Meanwhile, the leaching rate increased from 77.32% to 87.93% in the glycine leaching system. Finer particles have a larger contact area, resulting in higher leaching rates. The simple leaching kinetics model proposed in this study can accurately describe the leaching process of copper. These two environmentally friendly leaching systems can be used as substitutes for traditional inorganic acids. [Display omitted] [ABSTRACT FROM AUTHOR]

Zinc ferrite can be found in zinc-bearing dust in ironmaking and steelmaking. It is difficult to be recovered due to its stable properties. The magnetization roasting and ammonia leaching method were used to separate iron and zinc from zinc ferrite in this study. Thermodynamic analysis showed that the key to the selective reduction of zinc ferrite to zinc oxide and ferric oxide is to control the appropriate temperature and atmosphere. The influences of the selective reduction roasting process of zinc ferrite on the conversion rate, phase change, and microevolution behavior were investigated. The microstructure analysis showed that the distribution area of iron was mainly gray, and the distribution area of zinc was mainly white grid lines. The zinc content in the white area was higher than that in the gray area. With the increase in temperature and PCO, the white area expanded and the migration of zinc and iron was accelerated, but the iron in the white area still existed. The ammonia leaching of the magnetization-roasted product showed that a zinc leaching rate of 78.12% was achieved under the following conditions: the roasting atmosphere of PCO/P(CO+CO2) = 25%, the roasting temperature of 750 °C, roasting duration for 45 min, n(NH3-H2O):n(NH4Cl) = 1:1, the solid–liquid ratio of 40 g/L, leachate concentration of 6 mol/L, leaching duration of 90 min, the leaching temperature of 50 °C, and the stirring rate of 200 rpm.

Roasting is an essential process before ammonia leaching of low-grade rhodochrosite ore. Thermal decomposition characteristics and phase transformation rules of low-grade rhodochrosite ore with a particle size of less than 74 µm in N2, CO2 and air were studied. MnCO3, MgCO3 and CaCO3 in rhodochrosite are decomposed into MnO, MgO and CaO at 339 ℃, 306 ℃ and 842 ℃, respectively. MnO is oxidized to MnO2, Mn2O3 and Mn3O4 at 339 ℃, 533 ℃ and 1005 ℃, respectively. The TG/DTA and phase transformation results indicated that by thermal decomposition of rhodochrosite ore in air, N2 and CO2, MnCO3 is converted into Mn2O3, MnO and MnO, respectively. MnCO3 is converted to MnO at a higher temperature in a carbon dioxide atmosphere (512 ℃) than in a nitrogen atmosphere (400 ℃). [ABSTRACT FROM AUTHOR]

Zinc ferrite (ZnFe2O4), one of the major forms of zinc in electric arc furnace dust and zinc leaching residue, is very difficult to chemically dissolve in either acidic or alkaline solutions. It is therefore necessary to study the extraction of zinc from zinc ferrite to achieve the full recovery of the zinc resources. In this paper, CaO was used to degrade the structure of zinc ferrite and then the ammonia leaching method was applied to recover zinc. The thermodynamic analysis indicated that the zinc ferrite could react with CaO to form zinc oxide and dicalcium ferrite. The CaO–Fe2O3–ZnO phase diagram shows that Ca2Fe2O5 can be in equilibrium with ZnO rather than ZnFe2O4 at a higher CaO area. The experimental results indicated that the phase transformation of zinc ferrite, reacted with CaO, to zinc oxide and dicalcium ferrite could be completely achieved with a CaO addition of 2.4:1 (mole ratio). The peaks of zinc ferrite in the X-ray diffraction (XRD) patterns cannot be clearly observed after calcium roasting at 1220 °C for 2 h. The ammonia leaching tests showed that the zinc leaching rate was about 90% under optimal roasting conditions. [ABSTRACT FROM AUTHOR]

The existence of metal impurities (Cu, Fe, Al) seriously affects the extraction process of valuable metals in spent Ni-Co-Mn ternary (NCM) materials. Prior to reclaiming valuable metals from spent NCM materials, removing metal impurities is the pivotal precondition for improving recovery efficiency. In this manuscript, the properties of metal impurities in spent NCM materials before and after the roasting are systematically studied. The results of thermogravimetric analysis, XPS analysis, and thermodynamic analysis elaborate that the decomposition products of the electrolyte react with metal impurities to form fluorides, which is unfavorable to the removal of metal impurities. On this basis, a novel route for selective extraction of metal impurities (Cu, Fe, Al) from the spent NCM materials before roasting is proposed, that is, a combined method of magnetic separation, alkaline leaching, and ammonia leaching are used to remove Fe, Al, and Cu in spent NCM materials. Ultimately, the 76.1% Fe, 100% Al, and 100% Cu are removed. The results of leaching thermodynamics further confirm the feasibility and effectiveness of the method. This work displays an efficient and economically feasible approach for removing metal impurities from spent NCM materials, which has great prospects for industrial-scale recycling. [ABSTRACT FROM AUTHOR]

An ammonia-based system was used to selectively leach cobalt (Co) from an African high-silicon low-grade Co ore, and the other elemental impurities were inhibited from leaching in this process. This process was simple and environmentally friendly. The results revealed that the leaching ratio of Co can reach up to 95.61% using (NH4)2SO4 as a leaching agent under the following materials and conditions: (NH4)2SO4 concentration 300 g/L, reductant dosage 0.7 g, leaching temperature 353 K, reaction time 4 h, and liquid—solid ratio 6 mL/g. The leaching kinetics of Co showed that the apparent activation energy of Co leaching was 76.07 kJ/mol (i.e., in the range of 40–300 kJ/mol). This indicated that the leaching of Co from the Co ore was controlled by an interfacial chemical reaction, and then the developed leaching kinetics model of the Co can be expressed as 1 − (1 − α) 1 / 3 = 28.01 × 10 3 × r 0 − 1 × C (NH 4) 2 SO 4 1.5 × exp (− 76073 / 8.314 T) × t , where α is the leaching ratio (%) of Co, r0 is the average radius (m) of the Co ore particles, T is the temperature (K), and C (NH 4) 2 SO 4 is the initial reactant concentration (kg/m3). [ABSTRACT FROM AUTHOR]

With the permeation of LIBs into electric vehicles, the recycling of spent LIBs appears inevitable from the perspective of health, economic, and environmental protection. In this work, a facile process was proposed to realize the recycling and reutilization of valuable metals from spent LIBs. By ammonia leaching, co-precipitation, and solid phase reaction, fresh LiNi0.6Co0.2Mn0.2O2 cathode material was regenerated. The results revealed that an increase in the concentration of sulfite and ammonia was beneficial to the dissolution of valuable metals except aluminum. And the dissolution of Mn was mainly dominated by pH value and the concentration of sulfite. The initial specific discharge capacity of the regenerated LiNi0.6Co0.2Mn0.2O2 reaches 178.4 mAh·g−1 at 0.1 C. After 100 cycles, the capacity retains 92.1% at 0.5 C, which shows excellent electrochemical performance. In general, the recycling process may be worth considering for the recycling of spent lithium-ion batteries. [ABSTRACT FROM AUTHOR]

In this study, the direct conversion behaviors of different alkaline earth metal solids (the hydroxides and the sulfates of alkaline earth metals Ca, Sr, Ba and Mg) to their corresponding carbonates in dissolved carbonate-containing pregnant solutions obtained by direct leaching of a smithsonite (ZnCO3) ore sample in aqueous ammonia solutions having different concentrations (4 M, 8 M and 13.3 M NH3) were investigated by using X-ray diffraction analyses at alkaline earth metal to dissolved carbonate mole ratios of 1:1 and 1:2, for revealing the conversion possibilities of dissolved carbonate in the pregnant solutions to solid carbonate by-products. The results of direct conversion experiments showed that Ca(OH)2, CaSO4·2H2O, Sr(OH)2·8H2O and Ba(OH)2·8H2O converted to their corresponding carbonates, SrSO4 partially converted to SrCO3 as observed by the presence of unreacted SrSO4 peaks in X-ray diffraction patterns of the converted solids, and BaSO4 did not convert to BaCO3 because of its lower solubility with respect to BaCO3. On the other hand, it was observed that Mg(OH)2 did not convert to MgCO3, but MgSO4·7H2O converted dominantly to an uncommon phase, which was tentatively identified as Mg5Zn3(CO3)2(OH)12·H2O. In the study, a complete discussion on the conversion behaviors of alkaline earth metal solids to their corresponding carbonates was given considering the differences between their solubility product constants and the changes in the free energies of the theoretical conversion reactions. In addition, infrared spectra and scanning electron microscope images of some of the converted solids were also presented for characterization purposes. [ABSTRACT FROM AUTHOR]

Smelting of copper concentrates in a flash furnace directly into a copper blister results in the production of slag with high copper and lead content. The recovery of these metals is carried out in an electric furnace, from which copper–lead–iron metallic alloy is then processed in a converter furnace. In the presented work, copper and lead extraction studies were carried out by hydrometallurgical route under leaching in ammonium acetate solutions. The obtained results indicate the possibility of receiving quite high copper leaching efficiency from slag with a relatively low yield for lead extraction. The use of ammonium acetate solutions with a concentration of 5 mol/dm3 leads to leaching from the slag over 80% of the copper contained in it after 90 min of the process carried out at a temperature of 70 ℃. Model calculations related to the study of the kinetics of the process indicate that after 120 min of leaching it can be expected that 90% of the copper contained in the slag could be carried out into the solution.

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In this study, firstly, the effects of ammonia concentration, leaching time and solid/liquid ratio on the leaching behaviour of zinc from a smithsonite (ZnCO3) ore sample in aqueous ammonia solutions were investigated at room temperature by chemical, X-ray diffraction (XRD) and Fouriertransform infrared (FT-IR) spectroscopy analyses. It was found that leaching ratio of zinc steeply increased from 30.1 to 76.2% with increasing ammonia concentration from 1.0 to 4.0 M and maximum zinc leaching ratio of 79.7% was reached after leaching in 13.3 M NH3 solution. The XRD pattern of the residue obtained after leaching in 4.0 M NH3 solution for 90 min at solid/liquid ratio of 0.15 g/mL, the optimum condition, showed that smithsonite phase in the ore sample almost completely dissolved whereas the gangue minerals goethite and calcite remained unaffected, confirming the selectivity of ammonia solution for zinc dissolution. Together with zinc, leaching ratios of cadmium were also determined. In second part of the study, precipitation tests (by complete drying at different temperatures) were conducted on dissolved zinc, carbonate and ammonia containing pregnant solutions obtained after selected leaching experiments. By complete drying of the pregnant solutions at low temperatures, i.e. 50°C, relatively pure solid zinc ammine carbonate (Zn(NH3)CO3) precipitates and at higher temperatures, i.e. 150°C, quite pure solid zinc carbonate hydroxide (Zn5(CO3)2(OH)6) precipitates could be prepared. High-temperature heating of Zn(NH3)CO3 and Zn5(CO3)2(OH)6 precipitates at 450°C yielded single-phase zinc oxide (ZnO). The chemical compositions, FT-IR spectra and scanning electron microscope (SEM) photographs of some of the precipitates were also presented. [ABSTRACT FROM AUTHOR]

• No impurities are introduced in GBIC low-temperature roasting process. • In-situ water in GBIC can react with willemite to form hemimorphite. • The formation of hemimorphite promotes the leaching efficiency of zinc. • The efficient selective separation of Zn and Ge is achieved in the AAC system. • AAC filtrate can be reused, and leaching residue is employed to extract Ge. Germanium-bearing iron cake (GBIC) produced through the germanium removal process in a zinc smelter is a valuable source of zinc and germanium. However, achieving selective separation of zinc and germanium using pyrometallurgy treatment and sulfuric acid leaching methods is challenging. To address these challenges, this study introduced a novel roasting–leaching process. During roasting, willemite transformed into ammonia-soluble hemimorphite, promoting zinc leaching. This transformation was mainly attributed to the reaction between in situ water in GBIC and willemite within 200–500 °C. Optimal roasting conditions included a particle size range of 49–74 µm, a roasting temperature of 490 °C, and a duration of 4 h. Under optimized leaching conditions—[NH 4 ++NH 3 ] (7 mol/L), NH 4 +/NH 3 molar ratio (2.5), temperature (50 °C), liquid-to-solid ratio (20 mL/g), agitation speed (600 rpm), and time (180 min)—zinc and germanium exhibited leaching yields of 84.92 % and 0.59 %, respectively, in an ammonia–ammonium carbonate system. After three cycles of leaching, the Zn concentration in leachate increased to 20 g/L, facilitating the subsequent recovery of zinc. Moreover, the Ge concentration in the germanium-bearing leaching residue increased by 1.5 times compared with the initial GBIC, making the residue a valuable source of germanium. [ABSTRACT FROM AUTHOR]

A low-grade copper oxide ore with highly alkaline was treated. According to the characteristics of the solution with the high content of calcium and magnesium, ammonia ammonium sulfate leaching system was adopted to conduct leaching experiment for recovering copper from a low-grade oxidized copper ore. The thermodynamic model of Cu2 (OH)2 CO3 - ( NH4) 2SO4 -NH3-H2O was established in the leaching system of malachite ( Cu2 (OH ) 2CO,) under the condition of double conservation of mass and charge. The fitting function of the Matlab software and the diff and solve functions were used to calculate the optimal ammonia concentration and total copper ion concentration of malachite under different ammonium sulfate concentrations, respectively. Parameters, such as ammonia concentration, ammonia-ammonium ratio and liquid solid ratio were optimized. The results indicate that the appropriate conditions include an ammonia concentration of 1. 2 mol/L, an ammonia-ammonium ratio of 2: 1 and a liquid-solid ratio of 3: 1. Under these conditions, the copper leaching rate reaches - 70% . The calculated results are basically consistent with the experimental measurements. [ABSTRACT FROM AUTHOR]

Zinc ferrite (ZnFe2O4), one of the major forms of zinc in electric arc furnace dust and zinc leaching residue, is very difficult to chemically dissolve in either acidic or alkaline solutions. It is therefore necessary to study the extraction of zinc from zinc ferrite to achieve the full recovery of the zinc resources. In this paper, CaO was used to degrade the structure of zinc ferrite and then the ammonia leaching method was applied to recover zinc. The thermodynamic analysis indicated that the zinc ferrite could react with CaO to form zinc oxide and dicalcium ferrite. The CaO–Fe2O3–ZnO phase diagram shows that Ca2Fe2O5 can be in equilibrium with ZnO rather than ZnFe2O4 at a higher CaO area. The experimental results indicated that the phase transformation of zinc ferrite, reacted with CaO, to zinc oxide and dicalcium ferrite could be completely achieved with a CaO addition of 2.4:1 (mole ratio). The peaks of zinc ferrite in the X-ray diffraction (XRD) patterns cannot be clearly observed after calcium roasting at 1220 °C for 2 h. The ammonia leaching tests showed that the zinc leaching rate was about 90% under optimal roasting conditions.

Smelting of copper concentrates in a flash furnace directly into a copper blister results in the production of slag with high copper and lead content. The recovery of these metals is carried out in an electric furnace, from which copper–lead–iron metallic alloy is then processed in a converter furnace. In the presented work, copper and lead extraction studies were carried out by hydrometallurgical route under leaching in ammonium acetate solutions. The obtained results indicate the possibility of receiving quite high copper leaching efficiency from slag with a relatively low yield for lead extraction. The use of ammonium acetate solutions with a concentration of 5 mol/dm3 leads to leaching from the slag over 80% of the copper contained in it after 90 min of the process carried out at a temperature of 70 ℃. Model calculations related to the study of the kinetics of the process indicate that after 120 min of leaching it can be expected that 90% of the copper contained in the slag could be carried out into the solution. [ABSTRACT FROM AUTHOR]

A novel method for recovering zinc from zinc ferrite by reduction roasting-ammonia leaching was studied in this paper. The reduction thermodynamic of zinc ferrite by CO was analyzed. The effects of roasting parameters on the phase transformation and conversion rate of zinc ferrite, and the leaching behavior of zinc from the reductive roasted samples by ammonia leaching, were experimentally investigated. The mineralogical phase compositions and chemical compositions of the samples were characterized by X-ray diffraction and chemical titration methods, respectively. The results showed that most of the zinc ferrite was transformed to zinc oxide and magnetite after weak reduction roasting. 86.43% of the zinc ferrite was transformed to zinc oxide under the optimum conditions: CO partial pressure of 25%, roasting temperature of 750°C, and roasting duration of 45 min. Finally, under the optimal leaching conditions, 78.12% of zinc was leached into the solution from the roasted zinc ferrite while all iron-bearing materials were kept in the leaching residue. The leaching conditions are listed as follows: leaching duration of 90 min, ammonia solution with 6 mol/L concentration, leaching temperature of 50°C, solid-to-liquid ratio of 40 g/L, and stirring speed of 200 r/min. [ABSTRACT FROM AUTHOR]

• This strategy improves the application of ammonia leaching in battery recycling. • The leaching time is shortened by its superior mechanism. • Valuable metals are well extracted within a short time. • Iron as impurity element is naturally precipitated along with the process. The normalization of the recycling process is considered a crucial guarantee for the sustainable development of lithium-ion batteries. Ammonia leaching is known for its capability of selective extraction, which is of vital importance to alleviate the burden of impurity separation. However, the sluggish reaction rate limits its practical application. Herein, an electrochemical reduction strategy is well-designed for ammonia leaching by introducing Fe/Fe2+ which functions as negative electrodes to form galvanic cells with cathode material (LiNi 0.33 Co 0.33 Mn 0.33 O 2 , NCM). Due to its superior mechanism, reaction time can be shortened with a satisfied result of valuable metals extraction. Around 100% Li and Ni, 91.74% Co can be extracted within 105 min by introducing Fe powder; 91.75% Li, 85.24% Ni and 75.25% Co can be extracted within an even shorter time of 15 min by introducing FeSO 4. This strategy provides inspiration and a new perspective on large-scale battery recycling, especially for the practical application of ammonia leaching. [ABSTRACT FROM AUTHOR]

Due to the high cost of the manganese solvent extraction process in the conventional recycling of spent NCM-ternary lithium-ion batteries (LIBs), we employed an anoxic complexation ammonia leaching-spontaneous precipitation process to selectively recover manganese from the black mass of spent ternary LIBs. This innovation serves as a substitution for the conventional acid leaching-manganese solvent extraction, aiming to reduce costs and enhance the efficiency of the recycling process. Specifically, the controlled reduction roasting was performed to achieve the dissociation of LiNi x Co y Mn z O 2 and the phase transition to MnO and metallic Ni and Co. Exploiting this difference, ammonia-ammonium carbonate were employed as leaching agents, which enabled the selective leaching of over 96 % of manganese, while those of nickel and cobalt were only 1.2 % and 2.6 %. After filtration, the MnCO 3 product with high purity was precipitated by allowing the leachate to stand for 12 h without adding any reagent. The ammonia solution could be circulated to leaching the subsequent batch of spent LIBs. Thermodynamic calculations were used to elucidate the leaching and crystallization mechanisms. Comparative process analysis and economic assessments confirmed that the proposed new method is extremely appealing due to its high selectivity, considerable economic advantages, and environmental benefits. A atom economic method was proposed to selectively extract manganese from spent LIBs via controlled reduction and ammonia leaching. [Display omitted] • The manganese is selectively recovered from spent ternary lithium-ion batteries. • 96 % of manganese was leached and those of nickel and cobalt were 1.2 % and 2.6 %. • The manganese was recovered as MnCO 3 by spontaneous precipitation. • The leaching and crystallization mechanism of manganese was revealed. • The process comparison confirmed the economic and environmental benefits. [ABSTRACT FROM AUTHOR]

The ammonia leaching method for treating low-grade rhodochrosite has the advantages of a good impurity removal effect and low environmental pollution. In this paper, aiming at the low leaching efficiency of low-grade rhodochrosite treated by the ammonia leaching method, studies on enhancing the leaching efficiency of manganese by using ammonium hydrogen fluoride as an additive are carried out. The effects of different ammonia concentrations, leaching temperatures, leaching times, liquid-solid ratios, stirring rates, and the addition of ammonium hydrogen fluoride on the leaching efficiency of manganese with and without ammonium hydrogen fluoride as an additive were comparatively studied, and the parameters of ammonia concentration, ammonia leaching temperature, and ammonium hydrogen fluoride dosage were optimized in the experimental study. The results indicated that ammonium hydrogen fluoride as an additive in the treatment of low-grade rhodochrosite by the ammonia leaching method could effectively increase the leaching efficiency of manganese, and the optimal process parameters were obtained. Meanwhile, the addition of ammonium hydrogen fluoride didn’t affect the quality of the steamed ammonia product.

Secondary zinc oxide (SZO), which comes from the zinc industry, is an important secondary resource of zinc and other valuable metals. In this study, the production feasibility and rationality of a cleaner zinc recovery process using SZO and a hydrometallurgical method were described. Zinc extraction is promoted by the addition of ammonium bicarbonate to a NH3–H2O system, and the maximum recovery of zinc could be close to 80% at the optimum leaching conditions of a stirring rate of 400 rpm, an ammonia/ammonium ratio of 7:3, a total ammonia concentration of 4 mol/L, and a liquid/solid ratio of 7 mL/g for 30 min at 35 °C. The kinetics of leaching were modeled using the shrinking core model of constant-size particles, and the rate-controlling step was determined to be the diffusion through the product layer. The apparent activation energy of the reaction was estimated to be 11.04 KJ·mol−1, while the order of reaction with respect to total ammonia concentration was 1.53 and the liquid/solid ratio was 2.26. The analysis results of the initial residue and the leached residue indicated that lead was transferred from PbCl2 to PbCO3 and that ZnFe2O4 was not leached in the NH3-NH4HCO3-H2O system.

The leaching of zinc from lead-zinc slag using ammonia and preparation of basic zinc carbonate by evaporating were researched. The effects of temperature, reaction time, total ammonia concentration and liquid-to-solid ratio on leaching of zinc were examined. The results show that factors influencing the zinc leaching rate are listed in the following order: liquid-to-solid ratio>reaction time>temperature>total ammonia concentration, and the optimal parameters are temperature of 30 °C, reaction time of 2 h, total ammonia concentration of 4 mol/L,liquid-to- solid ratio of 7 mL/g. Under the conditions,the leaching of zinc is 81. 99%. The amorphous basic zinc carbonate Zn5(CO3)2(OH)6 * 2H2O can be prepared from the leaching solution. [ABSTRACT FROM AUTHOR]

Copper smelter slag can be considered as an important source for using in the copper hydrometallurgy. Ammonia leaching seems to be attractive for processing copper slag. Under the influence of ammonia, copper, which is part of the minerals of copper, forms soluble ammonia complexes, while iron precipitates as insoluble compounds. In the present work, the possibility of leaching the copper smelter slag (1.26 wt.% Cu) of the Balkhash smelter using NH4OH solutions was considered. The effect of experimental factors (leaching duration, reagent concentration, temperature, stirring rate, as well as a solid-to-liquid ratio) on the extraction of copper into solution was studied. It was found that the extraction of copper into solution increases with increasing temperature (in the studied range 298 - 333 K), pulp density (up to 10 pct) and stirring rate (up to 800 rpm). The concentration of NH4OH almost does not affect on the level of copper extraction in the range of 1-4 M. The following experimental conditions provide the recovery of 65% of Cu into ammonia solution: 1M NHNHOH, T = 333 K, particle size 90% < 200 mesh, solid-to-liquid ratio 10 pct, stirring rate 800 rpm, and leaching duration 180 min, 65% of copper is extracted into the solution. Shrinking core model with mixture kinetics was used to describe the process of ammonia leaching of copper from smelter slag. The activation energy and pre-exponential factor for the reactions of copper dissolution were calculated to be 16.2 ± 0.7 kJ/mol and 0.138 ± 0.001 min-1, respectively. The relatively low value of the activation energy indicates that the rate of the overall leaching process is controlled mainly by the processes of mass transfer rather than by the rate of chemical reactions of copper dissolution. [ABSTRACT FROM AUTHOR]

• The cathode powder is obtained by wet impact crushing. • A ternary leaching system contains of NH 3 ·H 2 O, (NH 4) 2 SO 3 and NH 4 HCO 3. • (NH 4) 2 SO 3 is reductant while NH 4 HCO 3 acts as a pH buffer. • Ni and Cu fully leached out while Li and Co exhibit a moderate leaching efficiency. The cathode powder is obtained by wet crushing and screening, and the leaching behavior of Li, Ni, Co, Cu, and Al is then investigated using a ternary leaching system composed of ammonia, ammonium sulfite, and ammonium bicarbonate. Ammonium sulfite is necessary as a reductant to improve the Li, Ni, and Co leaching efficiencies, and ammonium bicarbonate acts as a buffer in ammoniacal solutions. A detailed understanding of the selective leaching process is obtained by investigating the effects of parameters such as the leaching reagent composition, leaching time (0–300 min), temperature (40–90 °C), solid-to-liquid ratio (10–50 g/L), and agitation speed (300–700 rpm). It is found that Ni and Cu could be almost fully leached out, while Al is hardly leached and Li(60.53%) and Co(80.99%) exhibit a moderate leaching efficiency. The results show that the optimum solid-liquid ratio of the leaching system is 20 g/L, and the increase of temperature and reaction time is beneficial to metal leaching. The leaching kinetics analysis shows that the chemical reaction control explains the leaching behavior of Li, Ni, and Co well. Therefore, this work may be beneficial for further recycling valuable metals from leaching solutions by introducing an extraction agent. [ABSTRACT FROM AUTHOR]

[Display omitted] • Complete method for the selection of metals of interest have been proposed. • Selective leaching of copper and zinc has been achieved. • Zinc was selectively precipitated. • Copper was electrodeposited. • Leaching solution was reused in a closed loop manner. In 2019, 224 Mt of Municipal Solid Wastes (MSW) were produced in Europe. A significant proportion of this MSW is treated by incineration producing each year 20 Mt of Municipal Solid Waste Incineration (MSWI) Bottom Ashes (BA). While the coarsest fraction (>2mm) can be reemployed in construction field, this is not the case for the finest fraction <2 mm (30% wt.) which is intended to be landfilled. The challenges of the treatment of this solid residue are multiple. The metals of interest are dispersed within this complex matrix and the gangue is composed by a large proportion of carbonate based minerals which reduces the choice of possible leaching agents. Moreover, the large amounts of iron and aluminum imply to choose a highly selective recovery method and the economic constrains steer the process towards low-cost technologies. The aim of this work is to study the feasibility of a selective copper and zinc recovery, whose contents in the bottom ash fine fraction are comparable to low-grade ores. A four-step process was designed consisting of a flushing, a leaching, a precipitation and an electrowinning. A selective leaching step using ammonia-ammonium chloride solution was carried out to avoid an iron co-leaching. After the leaching step, the precipitation of zinc was achieved by addition of aluminum nitrate followed by an electrodeposition of copper from ammonia medium. Up to 70% of Cu and Zn could be recovered with minimal modifications of the leachate composition allowing thus its reuse. The developed process enables the recovery of metallic copper and zinc in a spinel form from a waste originally intended for landfilling. [ABSTRACT FROM AUTHOR]

Roasting pretreatment and ammonia leaching were performed to extract zinc from complicated zinc ores. The residue of the unroasted ore showed that the zinc containing phase ZnCO3 cannot be effectively leached in the ammonia leaching system. Mineral phase transformation of ZnCO3 takes place at a roasting temperature of 673 K, and this is the reason for the improvement of zinc leaching recovery. Additionally, the parameters that can influence the leaching rate of zinc such as the calcined temperature, the total ammonia concentration, the ratio of liquid to solid, the stirring speed and the leaching time were also investigated. Zinc leaching recovery from the complicated zinc ores could reach 84.7% under the following optimized experiment conditions: roasting temperature of 673 K, leaching temperature of 298 K, stirring speed of 500 rpm, total ammonia concentration of 6 mol/l, liquid to solid ratio of 11:1 and leaching time of 60 min. Compared to the zinc leaching recovery from unroasted ore (49.7%, ammonia concentration 6 mol/l), roasting pretreatment and optimization of process parameters can increase the zinc leaching recovery by 70.4%.

Ammonia leaching is a promising method for processing low-grade copper ores, especially those containing large amounts of oxidized copper. In this paper, we study the effect of Si-containing minerals on the kinetics of Cu and Ag leaching from low-grade copper concentrates. The results of experiments on the pressure leaching of the initial copper concentrate in an ammonium/ammonium-carbonate solution with oxygen as an oxidizing agent are in good agreement with the shrinking core model in the intra-diffusion mode: in this case, the activation energies were 53.50 kJ/mol for Cu and 90.35 kJ/mol for Ag. Energy-dispersive X-ray spectroscopy analysis (EDX) analysis showed that reagent diffusion to Cu-bearing minerals can be limited by aluminosilicate minerals of the gangue. The recovery rate for copper and silver increases significantly after a preliminary alkaline desilication of the concentrate, and the new shrinking core model is the most adequate, showing that the process is limited by diffusion through the product layer and interfacial diffusion. The activation energy of the process increases to 86.76 kJ/mol for Cu and 92.15 kJ/mol for Ag. Using the time-to-a-given-fraction method, it has been shown that a high activation energy is required in the later stages of the process, when the most resistant sulfide minerals of copper and silver apparently remain.

In pyrometallurgical processes refining copper, the main source of loss in the conversion stage is from slag. This paper reports on research work treating converter slag containing high percentages of copper (36 wt%) using ammonium hydroxide at room temperature. Variables analyzed are solution pH, agitation, temperature, NH4OH concentration and particle size. Results showed that the hydronium ion resulting from ammonium hydroxide dissociation was the main oxidant of copper compounds in slag, such as CuO, Cu2O and Cu, with the exception of CuFeO2. The particles contain a large amount of microcracks (porosity) in their refractory structure (analyzed by compositional image capture (BSE)). Thus, the diffusion of the leaching solution through the microcracks making contact with the copper oxides would be allowed. Leaching mechanisms were corroborated by X-ray diffraction and scanning electron microscopy analysis. Increasing temperature and NH4OH concentration while decreasing particle size obtained higher copper recoveries, reaching values of 84.8%. Under the same conditions, the main impurity (iron) was minimal (

On the basis of an experiment on ultrasonic enhanced ammonia leaching of low grade zinc oxide ores, the effects of the total ammonia concentration, the ratio between the NH3 concentration (CNH3) and that of ammonium sulfate (NH4)2SO4 (C(NH4)2SO4), and the ultrasound power on zinc leaching rate were studied. The results showed that the leaching efficiency of zinc observably increased with the increase of the total ammonia concentration, and that the ratios between CNH3 and C(NH4)2SO4 can remarkably affect the zinc extraction from the low grade zinc oxide ores. When the ratio between CNH3 and C(NH4)2SO4 is 2 to 1, the leaching rates of zinc are observably improved. The ultrasound can improve the leaching efficiency of zinc, and its effect was especially pronounced when the NH3 concentrations were low. The higher ultrasound power cannot sensibly improve the leaching rate of zinc, but it can improve the leaching speed and shorten the leaching time. Under the optimized conditions - total ammonia concentration of 8 mol/l, ratio between CNH3 and C(NH4)2SO4 of 2 to 1, and ultrasound power of 600 W – 83.33% of zinc is recovered.

[Display omitted] • Solid waste from steel mills was used for Zn recycling. • The ammonia leaching mechanism of SZO was proposed. • Zinc sulfide promotes zinc leaching by resisting the interference of heavy metals. • Zinc leaching was inhibited by the presence of lead and iron. Recovering zinc from typical industrial solid waste secondary zinc oxide (SZO) dust, rich in zinc resources, presents a significant challenge when finding an economical and green approach. This study aimed to investigate the optimal conditions for ammonia leaching of SZO dust using hydrothermal ammonia leaching methods and to explore how impurities affect zinc resource recovery in an ammonia leaching system. The results of SZO characterization before and after leaching confirmed that zinc oxide was the primary component dissolved in the ammonia leaching system. Only a fraction of zinc sulfide, lead oxide, and ferrous oxide entered the liquid phase, while most remained in the leaching slag. The study concluded that ammonium salt acted as a "dipping aid" in the ammonia leaching process and resulted in a 93.40% Zn leaching rate under optimal leaching conditions. Additionally, the study found that while impurities inhibited the leaching of Zinc Oxide in SZO, they promoted the leaching of Zinc Sulfide. Additionally, the leachate was found to contain recoverable amounts of lead and iron. These findings offer a theoretical framework for the efficient and cost-effective recovery of resources from SZO, providing valuable guidance toward achieving this objective. [ABSTRACT FROM AUTHOR]

Extraction of nickel from laterites, especially low-grade limonitic ores has been facing challenges related to the lack of an effective beneficiation method. A chemo-physical laterite ore processing comprising oxalic acid leaching followed by size separation of the leach residue was recently introduced by the authors as a novel method to collect small amounts of nickel values dispersed in the ore matrix into distinct oxalate particles to form an intermediate product. In the present work, the ammonia leaching of such intermediate product containing about 12 wt% NiC 2 O 4 ·2H 2 O was accomplished. Factors such as initial ammonia concentration (NH 3 : 2.5–25 wt%), temperature (T: 20–50 °C), pulp density (S/L: 50–200 g/L), and sonication were examined in a series of agitated leaching tests. It was found that under the conditions of 25% NH 3 , T = 50 °C, S/L = 200 g/L, and 15 min of sonication, about 93% of nickel is dissolved resulting in a pregnant leach solution (PLS) with a nickel content of 7.4 g/L. Besides, the activation energy for dissolution of nickel oxalate in ammonia solution was found to be about 71 kJ mol–1 from a kinetic analysis of leaching data. The PLS was subsequently deammoniated by either boiling or vacuum-assisted gentle heating to reclaim the nickel as an oxalate salt. Electron microscopic analysis of nickel oxalate precipitates formed from ammonia solutions exhibited rod-like morphology with high aspect ratio. Deammoniation under boiling led to well-shaped and distinct particles (∼150 nm width), while the precipitates obtained under partial vacuum were splintered and finer (∼40 nm width). The pyrolysis at 400 °C in an air atmosphere of the nickel oxalate particles from the boiling and the vacuum-assisted deammoniation methods generated mesoporous NiO nanorods with average widths of about 100 and 30 nm, and specific surface areas of about 45 and 70 m2 g−1, respectively. [Display omitted] • Hydrometallurgical treatment of a chemo-physically beneficiated laterite ore • Selective extraction of Ni content in the feed by parametric ammoniacal leaching • Ammonia concentration the most influential factor on the leaching rate and efficiency • Chemically controlled dissolution with an activation energy of 73 kJ mol−1 • Deammoniation of the PLS to recover Ni-oxalate nanorods with high-aspect ratio • Pyrolysis of Ni-oxalate in air to mesoporous high surface area NiO nanoparticles [ABSTRACT FROM AUTHOR]

Recovery of zinc from low grade zinc oxide ore is attempted with ammonia and ammonium citrate solutions augmented by microwave roasting and ultrasound radiation. The influence of the ammonia-ammonium ratio, roasting temperature, ultrasound power, and leaching time were assessed on the recovery of zinc. A maximum zinc recovery of 88.57% could be achieved at a roasting temperature of 673 K, leaching temperature of 298 K, stirring speed of 300 rpm, total ammonia concentration of 5 mol/L with an ammonium citrate concentration of 1.2 mol/L, liquid to solid ratio of 5:1, the ultrasound power was 600 W and the leaching time was 120 min. The enhancement in recovery with increases in the roasting temperature up to 673 K was attributed to the conversion of ZnCO3 to ZnO. The phases of mineral samples and the reaction residues were characterized by X-ray diffraction (XRD). [ABSTRACT FROM AUTHOR]

A recycling process consisting of swelling, ammonia leaching and hydrochloric acid leaching was developed to recover Sn, Bi, Cu, and resin from Pb-free Sn-Bi-Cu solder paste. The resin was separated from metal powder by swelling the solder paste in methyl ethyl ketone at 30 °C, 200 rpm and 5% pulp density. The addition of distilled water precipitated the resin from methyl ethyl ketone. The recovered resin was confirmed to be epoxy resin by FT-IR analysis. Ammonia leaching followed by hydrochloric acid leaching was performed to separate Cu, Sn, Bi sequentially from metal powder. Copper was solubilized easily with leaching efficiency reaching 100% within 15 min under the following conditions: 5 M NH 3 solution, 1 M (NH 4 ) 2 CO 3 , 0.1 M CuCO 3 , 50 °C, and 1% pulp density. Tin and bismuth were recovered as leach residue in the ammonia leaching. Tin was selectively leached in 0.5 M HCl solution with 10,000 mg/L Sn 4 + and 1% pulp density at 50 °C and 400 rpm, while Bi was not detected in the leach solution. The recycling process has successfully separated Sn, Cu, Bi and resin. [ABSTRACT FROM AUTHOR]

[Display omitted] • Complexant is added during the ammonia leaching of low-grade rhodochrosite to improve the leaching efficiency of manganese. • Manganese leaching efficiency is improved because the complex formed by the complexant and manganese increases the solubility of manganese in the leaching process. • EDTA is considered to be the most effective complexant to improve the leaching efficiency of manganese. There is a problem of low manganese leaching efficiency in the treatment of high-phosphorus and low-grade rhodochrosite by ammonia leaching. In this work, a proven and effective method was proposed to improve the manganese leaching efficiency by adding complexant during ammonia leaching of high-phosphorus and low-grade rhodochrosite. The thermodynamics of leaching process of rhodochrosite calcine after adding citric acid (CA), ethylenediaminetetraacetic acid (EDTA), nitrilotriacetic acid (NTA), glycine (Gly) different complexants were analyzed, and the effects of complexants on the manganese leaching efficiency were studied experimentally. Thermodynamic analysis and experimental results indicated that the four complexants can improve the leaching efficiency of manganese to varying degrees, and addition the EDTA has the best effect on improving the leaching efficiency of manganese. Moreover, various leaching parameters when EDTA was added and the effect of adding EDTA on ammonia evaporation product were further studied. The leaching optimum technology of rhodochrosite calcine with EDTA was obtained. Under the optimum technology, the manganese leaching efficiency increased to 90.1%. The addition of EDTA in the leaching process did not affect the ammonia evaporation product, and MnCO 3 product with 43.26% manganese content were obtained after evaporating ammonia. [ABSTRACT FROM AUTHOR]