Method for comprehensively recovering valuable metals from crude nickel cobalt hydroxide

Abstract

The invention belongs to the field of hydrometallurgy technology and high-purity valuable metal recovery, and particularly relates to a method for comprehensively recovering valuable metal from crude nickel cobalt hydroxide, which comprises the following steps: (1) acid leaching and manganese recovery; (2) recycling scandium; and (3) nickel and cobalt are recycled. The invention provides an integrated short-process green process for recovering four high-purity products from crude nickel cobalt hydroxide, and the process integrates key technologies such as three-section acid leaching manganese extraction, C272 extraction-hydrogen peroxide electrolysis synergistic purification, carbon dioxide selective reverse extraction and the like, so that efficient recovery of metals and cost reduction and efficiency improvement in the process are realized.

Description

TECHNICAL FIELD

[0001] The present application belongs to the field of hydrometallurgy technology and high-purity valuable metal recovery, and particularly relates to a method for comprehensively recovering valuable metals from crude nickel-cobalt hydroxide. BACKGROUND

[0002] Crude nickel-cobalt hydroxide (YS / T 1460-2021 "Crude nickel-cobalt hydroxide") is generally prepared by high-pressure acid leaching technology using laterite nickel ore as raw material, and can be used to produce nickel sulfate and electrolytic nickel. According to the component and actual demand, other elements such as cobalt and manganese are also recovered therefrom, for example, to produce products such as cobalt sulfate, electrolytic cobalt, and manganese sulfate. The existing hydrometallurgical technology for crude nickel-cobalt hydroxide generally adopts the processes of leaching, impurity removal, P204 impurity extraction, P507 cobalt-magnesium extraction, nickel sulfate evaporation, and manganese recovery. Part of the manganese is back-extracted with calcium, copper, and other impurities during P204 impurity extraction, and the P204 back-extraction solution needs to go through multiple impurity removal processes before it can meet the production requirements of battery-grade manganese sulfate. The composition of crude nickel-cobalt hydroxide is complex, and the control of impurities is extremely strict when recovering battery-grade nickel sulfate, high-purity nickel, high-purity cobalt, and high-purity scandium oxide from it. Therefore, it is often difficult to obtain high-purity products using conventional processes.

[0003] In the production of electrolytic nickel, the existing technology often adopts the P204-P507 extraction process, but needs to add various additives (such as boric acid, sodium salt, magnesium salt, surfactant, etc.) to the cathode solution to improve the appearance quality of electrolytic nickel, enhance the conductivity of the solution, and reduce the internal stress of the nickel plate. On the one hand, the extraction depth of P507 for magnesium and cobalt in the solution is low, and the main component of the back-extraction solution is cobalt sulfate, but it also usually contains several tens of milligrams per liter of nickel and several tens of milliliters per liter to several grams per liter of magnesium. Although the solution obtained by P507 back-extraction can meet the requirements of the electrolytic nickel process, the impurities in the raffinate often do not meet the requirements for directly producing battery-grade nickel sulfate (the cleanliness of the pre-crystallization solution is more critical when producing battery-grade nickel sulfate, and the presence of a small amount of nickel and magnesium will cause co-crystallization with nickel sulfate during evaporation crystallization, affecting the purity of the crystallized product), and the magnesium and other impurities therein need to be removed in depth. On the other hand, the impurities in the anode solution accumulate after electrolysis, and various impurities in the anode solution will co-crystallize during the evaporation crystallization process of nickel sulfate, affecting the purity of the product. Therefore, the anode solution cannot be directly used to produce high-purity nickel sulfate products, but needs to be added to an alkaline precipitating agent to precipitate nickel, and then go through washing-acid dissolution-impurity removal processes before it can be used to produce nickel sulfate. This process is long, complex, and costly, and it is difficult to switch to producing battery-grade nickel sulfate.

[0004] In the existing technology of hydrometallurgy or solvent extraction, Mg 2+ , Co 2+The conventional method is to use inorganic strong acid, such as hydrochloric acid, sulfuric acid. Strong acid can provide high concentration of H + , H + Combined with the loaded organic phase, the metal ions are replaced and returned to the aqueous phase. However, there are problems as follows: (1) large consumption of acid and base, high cost of wastewater treatment: a large amount of acid is consumed for stripping magnesium, and a large amount of base is consumed for subsequent neutralization treatment, which is high in cost; the conventional method can be regarded as an acid-base consumption cycle process: acid addition stripping → neutralization precipitation of the stripping solution after stripping by adding alkali → generation of wastewater, the stripping solution after stripping contains high concentration of salt (such as MgCl2, MgSO4, CoCl2, CoSO4, NiCl2, NiSO4, etc.), and the introduction of a large amount of SO4 2- , Cl - anions increases the difficulty and cost of wastewater treatment. (2) poor selectivity: if not properly controlled, strong acid will strip all the metal ions that can be stripped at one time, it is difficult to achieve complete separation of magnesium and cobalt, and the product purity cannot reach high purity / ultra-high purity level; a small amount of nickel, cobalt and other ions are also stripped together during magnesium stripping, and complex steps such as impurity removal and evaporation crystallization of the magnesium-containing solution are generally required for magnesium recovery, and the process flow is long. (3) strong acid has strong corrosiveness to equipment and pipelines, and has high requirements for the material of equipment and pipelines. SUMMARY

[0005] The present application aims to solve the problems of efficient separation and high-value recovery of multiple metals, and provides an integrated short process green process for recovering four high-purity products from crude nickel-cobalt hydroxide. The process integrates three-stage acid leaching of manganese, C272 extraction-hydrogen peroxide electrolysis synergistic purification, and carbon dioxide selective stripping, and realizes efficient recovery of metals and cost reduction and efficiency improvement of the process.

[0006] The technical scheme of the present application is as follows:

[0007] A method for comprehensively recovering valuable metals from crude nickel-cobalt hydroxide, comprising the following steps:

[0008] (1) acid leaching and manganese recovery: slurry the crude nickel-cobalt hydroxide with water according to a liquid-solid mass ratio (1-2): 1, heat to 50-95℃, add sulfuric acid for first-stage acid leaching, control the terminal pH value to be 2.5-3.0, separate the solid and liquid after reaction, obtain the first-stage acid leaching liquid and the first-stage acid leaching residue, recover manganese from the first-stage acid leaching residue, and obtain battery-grade manganese sulfate;

[0009] The main reaction occurring during the first stage of acid leaching is: M(OH)₂ + H₂SO₄ = MSO₄ + 2H₂O (M = Ni, Co, Mg, etc.). Since some manganese (approximately 30-80%) in the crude nickel-cobalt hydroxide exists as manganese dioxide, which has strong oxidizing properties and does not react when sulfuric acid is added alone, the main component of the first stage acid leaching residue is manganese dioxide, along with unreacted and entrained impurities such as nickel, cobalt, iron, and silicon. To fully recover manganese from the raw material and obtain a high-purity product, it is necessary to dissolve the impurities in the first stage acid leaching residue while avoiding or minimizing the dissolution of manganese, thereby recovering battery-grade manganese sulfate from the relatively pure manganese dioxide residue.

[0010] Maintaining a relatively high pH (2.5-3.0) at the end of the first stage of acid leaching can precipitate some impurities (such as iron ions) into the first-stage acid leaching residue, while simultaneously inhibiting the leaching rate of other impurities. This helps prevent impurities from accumulating in zirconium hydroxide during subsequent scandium recovery, thereby improving the purity of the final scandium extract and reducing the amount of precipitant used during scandium precipitation, thus lowering costs.

[0011] (2) Scandium recovery: Zirconium hydroxide is added to the first acid leaching solution to precipitate scandium, and the pH is controlled at 4.0-4.5. After the reaction is completed, solid and liquid are separated to obtain scandium precipitation liquid and scandium precipitation residue. Water is added to the scandium precipitation residue at a liquid-solid mass ratio of (1-2):1 to form a slurry and sulfuric acid is added for desorption. The final pH is controlled at 1.0-1.5. After solid-liquid separation, a scandium-containing solution is obtained. Scandium is recovered from the scandium-containing solution to obtain high-purity scandium oxide.

[0012] During the scandium precipitation process, scandium ions in a portion of the acid leaching solution are transferred to the solid phase through the combined effects of adsorption and neutralization by zirconium hydroxide, forming scandium precipitate residue. Subsequently, the adsorbed or precipitated scandium is desorbed using sulfuric acid to obtain a scandium-containing solution for further recovery. Compared with the conventional technique in this field of directly extracting scandium from low-concentration scandium-containing solutions using an extractant and then back-extracting, this method has the following advantages: ① The pH value is controlled at a low level (4.0-4.5) during scandium precipitation. Under this condition, fewer impurity ions are adsorbed, the scandium enrichment is high, and the scandium concentration in the subsequent sulfuric acid desorption solution is high (reaching 1-10 g / L) and the solution purity is high (thanks to the high solubility of scandium sulfate in water, which is conducive to the full desorption and dissolution of scandium); ② The precipitation and desorption processes of scandium are simple. During scandium precipitation, zirconium hydroxide only needs to be added directly to the acid leaching solution and the pH and other conditions need to be controlled. During desorption, sulfuric acid only needs to be added to the scandium precipitate residue and the pH and other conditions need to be controlled.

[0013] (3) Nickel and cobalt recovery: The scandium precipitation solution is subjected to P204 extraction to remove impurities, and the P204 raffinate is obtained after phase separation; the P204 raffinate is extracted with C272 extractant with a volume concentration of 15-25%, and the C272 raffinate and C272 organic phase loaded with cobalt and magnesium are obtained after phase separation; cobalt is recovered from the C272 organic phase to obtain high-purity cobalt; nickel is recovered from the C272 raffinate to obtain high-purity nickel or battery-grade nickel sulfate.

[0014] The scandium precipitation solution mainly contains ions such as nickel, cobalt, manganese, and magnesium, as well as small amounts of other impurities. P204 extraction can deeply remove these impurity ions and manganese. The resulting P204 raffinate is then subjected to C272 extraction for cobalt and magnesium. Subsequently, magnesium and cobalt in the C272 organic phase are back-extracted sequentially, and the cobalt back-extraction solution is further purified to produce high-purity cobalt products. High-purity nickel-based products can be obtained by degreasing, evaporating and crystallizing, or electrolyzing the C272 raffinate.

[0015] Compared to using P507 to extract cobalt and magnesium, C272 achieves higher extraction rates for both under the same conditions, with a higher cobalt-nickel separation coefficient (approximately 25 times that of P507), resulting in better selective extraction and separation of cobalt from magnesium and nickel. Furthermore, C272 has lower water solubility than P507, introducing less oil into the raffinate. Therefore, the resulting C272 raffinate has higher purity and can be directly used to produce high-purity nickel-based products. It also allows for flexible switching between different nickel-based product production systems, avoiding the need for deep magnesium and cobalt removal from the raffinate before producing battery-grade nickel sulfate due to incomplete magnesium or cobalt extraction, which would increase extractant consumption and prolong the process.

[0016] Further, in step (1), the method for recovering manganese from the first acid leaching residue includes:

[0017] a. Add water to the first stage of acid leaching residue at a wet base liquid-solid mass ratio of (1-3):1 to slurry, heat to 85-95℃, add concentrated sulfuric acid to carry out the second stage of normal pressure acid leaching, control the residual acid concentration at the reaction endpoint to be 70-150 g / L, and separate the solid and liquid after 2-4 hours of reaction to obtain the second stage of acid leaching solution and the second stage of acid leaching residue.

[0018] b. The second stage acid leaching residue is slurried with water at a wet base liquid-solid mass ratio of (1-3):1. The third stage of pressurized acid leaching is carried out under the conditions of pressure 0.1-0.2MPa and temperature 100-150℃, with an excess sulfuric acid coefficient of 1.1-1.3. After reacting for 1-2 hours, the solid and liquid are separated to obtain the third stage acid leaching solution.

[0019] c. Evaporate and crystallize the third acid leaching solution to obtain battery-grade manganese sulfate;

[0020] The second stage of acid leaching solution is returned to step (1) for the pulping of the crude nickel-cobalt hydroxide and the first stage of acid leaching.

[0021] The second stage of acid leaching mainly involves controlling the acidity and reaction temperature to fully dissolve impurities such as nickel, cobalt, iron, and silicon that were not completely leached from the first stage leaching residue. This improves the purity of the second stage leaching residue (mainly composed of manganese dioxide), which in turn benefits the purity of the final product, manganese sulfate. The main reaction occurring during the third stage of acid leaching is: 2MnO2 + 2H2SO4 → 2MnSO4 + 2H2O + O2↑.

[0022] Regarding the amount of sulfuric acid used: Sulfuric acid provides H₂ for redox reactions. + Insufficient dosage or concentration can lead to a decrease in reaction rate or even cessation of reaction. Therefore, an excess sulfuric acid coefficient of 1.1-1.3 is preferred.

[0023] Further, in step (2), the specific method for recovering high-purity scandium oxide from the scandium-containing solution is as follows: the scandium-containing solution is extracted using C272 extractant, and the phases are separated to obtain an organic phase loaded with scandium and a raffinate. Then, oxalic acid solution is added to the organic phase loaded with scandium for precipitation, and scandium oxalate solid is separated. High-purity scandium oxide is obtained by high-temperature calcination.

[0024] Further, in step (3), the method for recovering high-purity nickel from the C272 raffinate is as follows: the C272 raffinate is de-oiled to obtain a de-oiled liquid with a nickel concentration of 90-100 g / L, and 5-20 L / m³ of nickel is added to the de-oiled liquid. 3 High-purity nickel is obtained by electrolysis of hydrogen peroxide; the electrolysis conditions are: electrolytic cell temperature 55-65℃, current density 180-220 A / m 2 The distance between the same poles is 10-15 cm.

[0025] The main difference between the electrolytic nickel system of this invention and existing technologies is that existing technologies typically require the addition of various additives (such as boric acid, sodium sulfate, magnesium sulfate, surfactants, etc.) to improve the appearance quality of the electrolytic nickel plate, enhance the conductivity of the solution, and reduce internal stress. However, this invention does not introduce these additives during electrolytic production; instead, it only adds hydrogen peroxide while controlling the nickel ion concentration in the cathode solution at 90-100 g / L to achieve deep purification of the solution, improve the appearance quality of the electrolytic nickel, and reduce internal stress. This design eliminates the need for complex impurity removal treatment when the anolyte needs to be switched to produce crystalline nickel sulfate. Specific advantages are as follows:

[0026] ① By introducing hydrogen peroxide, controlling the nickel concentration in the cathode solution, and using C272 to extract cobalt and magnesium in a synergistic effect, the impurity content in the cathode solution can be ensured to be extremely low, while avoiding the introduction of new impurities. This makes the system switching between electrolytic nickel and nickel sulfate crystal production highly flexible, and high-purity nickel sulfate crystals can be produced without complex treatment of the anolyte, thereby shortening the anolyte reuse process (in contrast, in the existing technology, if nickel sulfate crystals are produced using anolyte, impurities such as sodium, boron, and magnesium must be removed before evaporation to obtain high-purity products, which is a complex and lengthy process).

[0027] ② Deep purification solution: Hydrogen peroxide can effectively decompose organic contaminants in the electrolyte (such as incompletely removed oil, COD, etc.), preventing their adsorption on the cathode surface, thereby improving the purity and density of the electrolytic nickel plate. In acidic electrolytes, hydrogen peroxide can also synergistically enhance the purification effect with sulfuric acid, stabilize the electrolysis process, and adjust the redox potential of the solution to inhibit harmful side reactions (such as Fe). 2+ The process involves the co-deposition of low-valence metal ions, thereby increasing the purity of the final electrolytic nickel product. Due to the high purity and low impurity content of the solution, this process allows for seamless switching between high-purity electrolytic nickel and battery-grade nickel sulfate production systems without complex processing.

[0028] ③ Improve the appearance quality of electrolytic nickel: Adding hydrogen peroxide to the electrolyte can improve solution convection and gas emission. The oxygen microbubbles generated by the decomposition of hydrogen peroxide can enhance electrolyte convection, reduce concentration polarization, and promote the escape of hydrogen and other bubbles, thus preventing the formation of pinholes or loose structures in the deposit layer due to bubble retention.

[0029] ④ Reduce internal stress: By purifying the solution and optimizing the deposition environment, hydrogen peroxide helps form a dense and smooth metal deposition layer, reducing lattice defects and internal stress, and improving the mechanical properties of the product.

[0030] In practical applications, the amount of hydrogen peroxide used needs to be controlled to avoid excessive oxidation due to excessive amount or poor purification effect due to insufficient amount.

[0031] Further, in step (3), the method for recovering cobalt from the C272-loaded organic phase includes: introducing carbon dioxide gas into the C272-loaded organic phase for slow stirring reaction, and obtaining a cobalt-loaded organic phase and a magnesium carbonate-containing slurry after phase separation; separating the magnesium carbonate-containing slurry into solid and liquid phases to obtain magnesium carbonate by-product; adding dilute sulfuric acid to the cobalt-loaded organic phase for back-extraction, controlling the pH of the aqueous phase to 2.0-2.5, and obtaining a cobalt sulfate solution and a regenerated organic phase after phase separation, wherein the regenerated organic phase is returned for cobalt extraction, and the cobalt sulfate solution is used to produce high-purity cobalt.

[0032] This method involves introducing CO2 into a supported organic phase, utilizing the weakly acidic environment (pH 4.5-5.5) formed in water and the heterogeneous interfacial reaction to achieve selective back-extraction of magnesium and generate magnesium carbonate, while cobalt remains in the organic phase. This results in efficient separation of magnesium and cobalt, with the advantages of a short process and environmental friendliness.

[0033] Further, the method for producing high-purity cobalt from the cobalt sulfate solution is as follows: the cobalt sulfate solution is subjected to deep impurity removal using D402 resin to obtain a purified solution; the purified solution is then subjected to diaphragm-free electrowinning to obtain high-purity cobalt; the electrowinning conditions are: electrolytic cell temperature 55-65℃, current density 180-220 A / m 2 The electrode spacing is 10-15 cm, the cathode is a cobalt starting plate, and the anode is a titanium plated with iridium and tantalum.

[0034] Further, in step (3), the method for recovering battery-grade nickel sulfate from the C272 raffinate is as follows: the C272 raffinate is passed through activated carbon for dynamic oil removal to obtain an oil-removed liquid; the oil-removed liquid is then subjected to evaporation crystallization, centrifugation separation and drying in sequence to obtain battery-grade nickel sulfate.

[0035] Furthermore, in step (3), the volume concentration of the P204 extractant during use is also 15-25%.

[0036] The process flow of this invention is as follows: Figure 1 As shown.

[0037] Due to the adoption of the above technical solution, the beneficial effects of the present invention are as follows:

[0038] (1) Nickel, cobalt, manganese and scandium can be recovered simultaneously from crude nickel-cobalt hydroxide to obtain their high-purity products;

[0039] (2) The crude nickel-cobalt hydroxide is subjected to three-stage acid leaching. Impurity ions are dissolved in advance, and then the acid leaching residue is leached under pressure. The manganese leaching rate is high, and finally, battery-grade manganese sulfate is obtained by evaporation and crystallization (high product purity).

[0040] (3) When electrolyzing nickel, boric acid, sodium salt, magnesium salt, surfactant and other additives are not introduced. Instead, hydrogen peroxide is added and the nickel concentration in the cathode liquid is controlled to achieve multiple purposes such as purifying the solution, improving the appearance quality of electrolyzed nickel and reducing internal stress, and finally producing high-purity nickel.

[0041] (4) The present invention uses C272 instead of the traditional P507 for deep extraction and separation of nickel, cobalt and magnesium. Through the synergistic effect of C272 extraction and the addition of hydrogen peroxide during nickel electrolysis, the impurity content in the raffinate can be controlled to be extremely low, while avoiding the introduction of impurities in the cathode liquid. This makes the system switching flexibility in the production of electrolytic nickel and nickel sulfate crystals, and high-purity nickel sulfate crystals can be produced without complex treatment of the anolyte, thus shortening the anolyte recycling process.

[0042] (5) By using carbon dioxide as a mild and controllable acidifying agent, efficient and selective back-extraction of magnesium and cobalt is achieved, which solves the problems of poor selectivity, high cost, long magnesium recovery process and high wastewater treatment cost in the traditional strong acid method for back-extraction of magnesium and cobalt. It provides a green and economical new method for the deep removal of impurities in hydrometallurgy, especially cobalt smelting, and the production of high-purity / ultra-high-purity electrolytic cobalt. Attached Figure Description

[0043] Figure 1 This is a process flow diagram of the present invention. Detailed Implementation

[0044] The technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0045] Example 1

[0046] This embodiment provides a method for the comprehensive recovery of valuable metals from crude nickel-cobalt hydroxide, the specific steps of which are as follows:

[0047] (1) Acid leaching and manganese recovery: 1000 kg of crude nickel-cobalt hydroxide (dry basis) was slurried with water at a liquid-to-solid mass ratio of 1.5:1, heated to 70°C, and concentrated sulfuric acid was added for the first stage of acid leaching. The final pH value was controlled at 2.8. After reacting for 2 hours, the mixture was filtered to obtain the first stage acid leaching solution and the first stage acid leaching residue (wet weight approximately 1500 kg). The acid leaching residue was slurried with water at a wet basis liquid-to-solid mass ratio of 2:1, heated to 85°C, and concentrated sulfuric acid was added for the second stage of atmospheric pressure acid leaching. The final residual acid concentration was controlled at 150 g / L. After reacting for 3 hours, the mixture was filtered to obtain the second stage acid leaching solution and the second stage acid leaching residue. The second stage acid leaching residue was slurried with the same slurry ratio, pumped into a pressure vessel, and subjected to a third stage of pressurized acid leaching at a pressure of 0.1 MPa and a temperature of 100°C. The excess sulfuric acid coefficient was 1.2. After reacting for 1.5 hours, the pressure was released and the mixture was filtered to obtain the third stage acid leaching solution. The third stage acid leaching solution is evaporated, concentrated, cooled, crystallized, and centrifuged to obtain battery-grade manganese sulfate. The second stage acid leaching solution is returned to the first stage acid leaching feed.

[0048] (2) Scandium recovery: Zirconium hydroxide slurry was slowly added to the first-stage acid leaching solution, and the pH of the system was controlled at 4.2. The mixture was stirred and reacted for 1 hour. After pressure filtration, the scandium-precipitated liquid and scandium-precipitated residue were obtained. Water was added to the scandium-precipitated residue at a liquid-to-solid mass ratio of 1.5:1 to form a slurry. Sulfuric acid was slowly added, and the final pH was controlled at 1.2. After desorption for 1 hour, the mixture was pressure filtered to obtain a scandium-containing solution (Sc). 3+ (Concentration approximately 5 g / L); the scandium-containing solution was subjected to three-stage countercurrent extraction using a 20% (v / v) C272 extractant. The scandium-enriched loaded organic phase was back-extracted with a 0.5 mol / L oxalic acid solution to obtain scandium oxalate precipitate, which was then calcined at 800℃ for 2 hours to obtain high-purity scandium oxide.

[0049] (3) Nickel and cobalt recovery: The scandium precipitation solution was subjected to three-stage extraction with 20% volume concentration of P204 extractant to remove impurities, resulting in P204 raffinate. The P204 raffinate was then subjected to five-stage extraction with 20% volume concentration of C272 extractant. After phase separation, C272 raffinate (main component is nickel sulfate) and C272 organic phase loaded with cobalt and magnesium were obtained.

[0050] Cobalt recovery: Food-grade carbon dioxide gas is introduced into the supported organic phase, and the gas flow rate is controlled to allow the system to bubble slowly while stirring for 40 minutes. After stopping the gas flow, the system is allowed to stand and separate, yielding a lower layer of magnesium carbonate-containing slurry and an upper layer of cobalt-supported organic phase. The slurry is then filtered, washed, and dried to obtain magnesium carbonate as a byproduct. 1.5 mol / L dilute sulfuric acid is added to the cobalt-supported organic phase for three-stage back-extraction, controlling the final aqueous phase pH to 2.3 to obtain a pure cobalt sulfate solution. This solution is then subjected to deep adsorption purification using a D402 resin column, followed by diaphragm-free electrodeposition (conditions: temperature 60℃, current density 200 A / m). 2 (with a pole distance of 12 cm), a high-purity cobalt plate was obtained.

[0051] Nickel recovery: The C272 raffinate is pumped into an activated carbon adsorption tower for dynamic oil removal, yielding an oil-removed liquid with a nickel concentration of 90 g / L. A portion of the oil-removed liquid is used as catholyte and added to a 20 L / m³ filter. 3 After being treated with hydrogen peroxide, it is fed into an electrolytic cell for electrolysis (conditions: temperature 60℃, current density 200 A / m). 2 (with a pole spacing of 12 cm), a high-purity nickel plate was obtained.

[0052] Example 2

[0053] This embodiment provides a method for the comprehensive recovery of valuable metals from crude nickel-cobalt hydroxide, the specific steps of which are as follows:

[0054] (1) Acid leaching and manganese recovery: 1000 kg of crude nickel-cobalt hydroxide (dry basis) was slurried with water at a liquid-to-solid mass ratio of 1.5:1, heated to 50°C, and concentrated sulfuric acid was added for the first stage of acid leaching. The final pH value was controlled at 2.5. After reacting for 2 hours, the mixture was filtered to obtain the first stage acid leaching solution and the first stage acid leaching residue (wet weight approximately 1500 kg). The acid leaching residue was slurried with water at a wet basis liquid-to-solid mass ratio of 2:1, heated to 90°C, and concentrated sulfuric acid was added for the second stage of atmospheric pressure acid leaching. The final residual acid concentration was controlled at 100 g / L. After reacting for 3 hours, the mixture was filtered to obtain the second stage acid leaching solution and the second stage acid leaching residue. The second stage acid leaching residue was slurried with the same slurry ratio, pumped into a pressure vessel, and subjected to a third stage of pressurized acid leaching at a pressure of 0.15 MPa and a temperature of 120°C. The excess sulfuric acid coefficient was 1.3. After reacting for 1.5 hours, the pressure was released and the mixture was filtered to obtain the third stage acid leaching solution. The third stage acid leaching solution is evaporated, concentrated, cooled, crystallized, and centrifuged to obtain battery-grade manganese sulfate. The second stage acid leaching solution is returned to the first stage acid leaching feed.

[0055] (2) Scandium recovery: Zirconium hydroxide slurry was slowly added to the first-stage acid leaching solution, and the pH of the system was controlled at 4.0. The mixture was stirred and reacted for 1 hour. After filtration, the scandium-precipitated liquid and scandium-precipitated residue were obtained. Water was added to the scandium-precipitated residue at a liquid-to-solid mass ratio of 2:1 to form a slurry. Sulfuric acid was slowly added, and the final pH was controlled at 1.0. After desorption for 1 hour, the mixture was filtered to obtain a scandium-containing solution (Sc). 3+ (Concentration approximately 5 g / L); the scandium-containing solution was subjected to three-stage countercurrent extraction using a 15% (v / v) C272 extractant. The scandium-enriched loaded organic phase was back-extracted with a 0.5 mol / L oxalic acid solution to obtain scandium oxalate precipitate, which was then calcined at 800℃ for 2 hours to obtain high-purity scandium oxide.

[0056] (3) Nickel and cobalt recovery: The scandium precipitation solution was subjected to three-stage extraction with 15% volume concentration of P204 extractant to remove impurities, resulting in P204 raffinate. The P204 raffinate was then subjected to five-stage extraction with 15% volume concentration of C272 extractant. After phase separation, C272 raffinate (main component is nickel sulfate) and C272 organic phase loaded with cobalt and magnesium were obtained.

[0057] Cobalt recovery: Food-grade carbon dioxide gas is introduced into the supported organic phase, and the gas flow rate is controlled to allow the system to bubble slowly while stirring for 40 minutes. After stopping the gas flow, the system is allowed to stand and separate, yielding a lower layer of magnesium carbonate-containing slurry and an upper layer of cobalt-supported organic phase. The slurry is then filtered, washed, and dried to obtain magnesium carbonate as a byproduct. 1.5 mol / L dilute sulfuric acid is added to the cobalt-supported organic phase for three-stage back-extraction, controlling the final aqueous phase pH to 2.0 to obtain a pure cobalt sulfate solution. This solution is then subjected to deep adsorption purification using a D402 resin column, followed by diaphragm-free electrodeposition (conditions: temperature 55℃, current density 220 A / m). 2 (with a pole spacing of 10 cm), a high-purity cobalt plate was obtained.

[0058] Nickel recovery: The C272 raffinate is pumped into an activated carbon adsorption tower for dynamic oil removal, yielding a de-oiled liquid with a nickel concentration of 100 g / L. A portion of the de-oiled liquid is used as catholyte and added at 5 L / m³. 3 After being treated with hydrogen peroxide, it is fed into an electrolytic cell for electrolysis (conditions: temperature 55℃, current density 220 A / m). 2 (with a pole spacing of 10 cm), a high-purity nickel plate was obtained.

[0059] Example 3

[0060] This embodiment provides a method for the comprehensive recovery of valuable metals from crude nickel-cobalt hydroxide, the specific steps of which are as follows:

[0061] (1) Acid leaching and manganese recovery: 1000 kg of crude nickel-cobalt hydroxide (dry basis) was slurried with water at a liquid-to-solid mass ratio of 1.5:1, heated to 95℃, and concentrated sulfuric acid was added for the first stage of acid leaching. The final pH value was controlled at 3.0. After reacting for 2 hours, the mixture was filtered to obtain the first stage acid leaching solution and the first stage acid leaching residue (wet weight approximately 1500 kg). The acid leaching residue was slurried with water at a wet basis liquid-to-solid mass ratio of 2:1, heated to 95℃, and concentrated sulfuric acid was added for the second stage of atmospheric pressure acid leaching. The final residual acid concentration was controlled at 70 g / L. After reacting for 3 hours, the mixture was filtered to obtain the second stage acid leaching solution and the second stage acid leaching residue. The second stage acid leaching residue was slurried with the same slurry ratio, pumped into a pressure vessel, and subjected to a third stage of pressurized acid leaching at a pressure of 0.2 MPa and a temperature of 150℃. The excess sulfuric acid coefficient was 1.1. After reacting for 1.5 hours, the pressure was released and the mixture was filtered to obtain the third stage acid leaching solution. The third stage acid leaching solution is evaporated, concentrated, cooled, crystallized, and centrifuged to obtain battery-grade manganese sulfate. The second stage acid leaching solution is returned to the first stage acid leaching feed.

[0062] (2) Scandium recovery: Zirconium hydroxide slurry was slowly added to the first-stage acid leaching solution, and the pH of the system was controlled at 4.5. The mixture was stirred and reacted for 1 hour. After pressure filtration, the scandium-precipitated liquid and scandium-precipitated residue were obtained. Water was added to the scandium-precipitated residue at a liquid-to-solid mass ratio of 1:1 to form a slurry. Sulfuric acid was slowly added, and the final pH was controlled at 1.5. After desorption for 1 hour, the mixture was pressure filtered to obtain a scandium-containing solution (Sc). 3+ (Concentration approximately 5 g / L); the scandium-containing solution was subjected to three-stage countercurrent extraction using a 25% (v / v) C272 extractant. The scandium-enriched loaded organic phase was back-extracted with a 0.5 mol / L oxalic acid solution to obtain scandium oxalate precipitate, which was then calcined at 800℃ for 2 hours to obtain high-purity scandium oxide.

[0063] (3) Nickel and cobalt recovery: The scandium precipitation solution was subjected to three-stage extraction with 25% volume concentration of P204 extractant to remove impurities, resulting in P204 raffinate. The P204 raffinate was then subjected to five-stage extraction with 25% volume concentration of C272 extractant. After phase separation, C272 raffinate (main component is nickel sulfate) and C272 organic phase loaded with cobalt and magnesium were obtained.

[0064] Cobalt recovery: Food-grade carbon dioxide gas is introduced into the supported organic phase, and the gas flow rate is controlled to allow the system to bubble slowly while stirring for 40 minutes. After stopping the gas flow, the system is allowed to stand and separate, yielding a lower layer of magnesium carbonate-containing slurry and an upper layer of cobalt-supported organic phase. The slurry is then filtered, washed, and dried to obtain magnesium carbonate as a byproduct. 1.5 mol / L dilute sulfuric acid is added to the cobalt-supported organic phase for three-stage back-extraction, controlling the final aqueous phase pH to 2.5 to obtain a pure cobalt sulfate solution. This solution is then subjected to deep adsorption purification using a D402 resin column, followed by diaphragm-free electrodeposition (conditions: temperature 65℃, current density 180 A / m). 2 (with a pole spacing of 15 cm), a high-purity cobalt plate was obtained.

[0065] Nickel sulfate production: All C272 raffinate is pumped into an activated carbon adsorption tower for dynamic oil removal, yielding a de-oiled liquid with a nickel concentration of 100 g / L. The de-oiled liquid is then transferred to an evaporator crystallizer, where it is evaporated and concentrated to supersaturation under vacuum, followed by cooling and crystallization. The crystal slurry is centrifuged, and the crystals are rinsed with a small amount of pure water and dried at 80°C to obtain battery-grade nickel sulfate hexahydrate.

[0066] To further illustrate the technical effects of the present invention, the applicant has set up the following comparative examples and conducted experimental comparisons as follows:

[0067] Comparative Example 1

[0068] Compared to Example 1, Comparative Example 1 changed the recovery sequence: instead of first precipitating scandium, it directly used P204 to extract a first stage of the acid leaching solution to remove impurities. The specific steps were as follows: the first stage of the acid leaching solution was directly extracted with 20% P204 extractant to remove impurities (intended to remove impurities such as iron, aluminum, and manganese). However, P204 has a very strong extraction ability for scandium, causing scandium to be co-extracted into the organic phase. When attempting to recover scandium from the loaded organic phase, it was found that scandium was difficult to separate from the various co-extracted impurities, requiring complex acid washing, alkali back-extraction, and other steps, and the final purity of the obtained scandium oxide could only reach 95%. At the same time, the extraction capacity of the P204 organic phase decreased due to the accumulation of scandium, requiring frequent regeneration.

[0069] Comparative Example 2

[0070] Compared to Example 1, Comparative Example 2 changed the cobalt-magnesium separation method: it adopted traditional sulfuric acid staged back-extraction (first using a weak acid to back-extract magnesium, then using a strong acid to back-extract cobalt). The specific steps are as follows: First, magnesium is back-extracted with 1 mol / L dilute sulfuric acid, controlling the pH at the back-extraction endpoint at 4.25. Under these conditions, approximately 15-20% of the cobalt is co-extracted. The separated cobalt- and magnesium-containing solution requires further separation of cobalt and magnesium, increasing the process length. Then, cobalt is back-extracted with 2.0 mol / L sulfuric acid, controlling the pH at the back-extraction endpoint at 2.0. This method consumes approximately 30% more acid and alkali than the CO2 method in Example 1, and the treatment of the resulting cobalt-containing magnesium sulfate wastewater is complex.

[0071] Comparative Example 3

[0072] Compared to Example 1, this comparative example used a 20% volume concentration of P507 extractant instead of C272 for cobalt-magnesium extraction. Under the same extraction stages and conditions, P507 had a lower magnesium extraction rate, resulting in a higher residual magnesium ion content (>50 mg / L) in the C272 raffinate. The purity of the nickel plates obtained by electrolysis using this raffinate was slightly reduced, and magnesium accumulated in the anolyte, making it unsuitable for direct use in the production of battery-grade nickel sulfate; a dedicated magnesium removal step was required.

[0073] The experimental data for the examples and comparative examples are shown in Table 1.

[0074] Table 1: Comparison of key data between the examples and comparative examples

[0075]

[0076] Data Analysis: As shown in Table 1, the method of the present invention (Examples 1-3) achieves efficient and high-purity recovery of four valuable metals through key technologies such as the "scandium precipitation followed by extraction" sequence, the use of C272 extractant combined with CO2 selective back-extraction of magnesium, and hydrogen peroxide-assisted electrolysis. Furthermore, the process is simple and the system switching is flexible. In contrast, changing the recovery sequence (Comparative Example 1) leads to a significant decrease in scandium recovery rate and purity; traditional acid back-extraction (Comparative Example 2) has poor separation effect and high cost; changing the extractant (Comparative Example 3) results in insufficient removal of key impurities, affecting product purity and process flexibility. This fully demonstrates that the technical solution of the present invention has non-obviousness and outstanding comprehensive effects.

[0077] The above description is a detailed description of the preferred embodiments of the present invention. However, the embodiments are not intended to limit the scope of the patent application of the present invention. All equivalent changes or modifications made under the technical spirit of the present invention should fall within the patent scope covered by the present invention.

Claims

1. A method for the comprehensive recovery of valuable metals from crude nickel-cobalt hydroxide, characterized in that, Includes the following steps: (1) Acid leaching and manganese recovery: Crude nickel-cobalt hydroxide is slurried with water, heated to 50-95℃, and sulfuric acid is added for the first stage of acid leaching. The final pH value is controlled to be 2.5-3.

0. After the reaction, solid and liquid are separated to obtain the first stage acid leaching solution and the first stage acid leaching residue. Manganese is recovered from the first stage acid leaching residue to obtain battery-grade manganese sulfate. (2) Scandium recovery: Zirconium hydroxide is added to the first acid leaching solution to precipitate scandium, and the pH is controlled at 4.0-4.

5. After the reaction is completed, solid and liquid are separated to obtain scandium precipitation liquid and scandium precipitation residue; water is added to the scandium precipitation residue to slurry and sulfuric acid is added to desorb it. After solid-liquid separation, a scandium-containing solution is obtained. Scandium is recovered from the scandium-containing solution to obtain high-purity scandium oxide. (3) Nickel and cobalt recovery: The scandium precipitation solution is subjected to P204 extraction to remove impurities, and the P204 raffinate is obtained after phase separation; the P204 raffinate is extracted with C272 extractant, and the C272 raffinate and C272 organic phase loaded with cobalt and magnesium are obtained after phase separation; cobalt is recovered from the C272 organic phase to obtain high-purity cobalt; nickel is recovered from the C272 raffinate to obtain high-purity nickel or battery-grade nickel sulfate.

2. The method according to claim 1, characterized in that, In step (1), the method for recovering manganese from the first acid leaching residue includes: a. Add water to the first stage acid leaching residue and slurry it. Heat the temperature to 85-95℃ and add concentrated sulfuric acid to carry out the second stage of normal pressure acid leaching. Control the residual acid concentration at the end of the reaction to be 70-150 g / L. After reacting for 2-4 hours, separate the solid and liquid to obtain the second stage acid leaching solution and the second stage acid leaching residue. b. The second stage acid leaching residue is slurried with water, and the third stage of pressurized acid leaching is carried out under the conditions of pressure 0.1-0.2 MPa and temperature 100-150℃. After reacting for 1-2 hours, solid-liquid separation is performed to obtain the third stage acid leaching solution. c. Evaporate and crystallize the third acid leaching solution to obtain battery-grade manganese sulfate; The second stage of acid leaching solution is returned to step (1) for the pulping of the crude nickel-cobalt hydroxide and the first stage of acid leaching.

3. The method according to claim 2, characterized in that, In the third stage of pressurized acid leaching described in step b, the excess coefficient of sulfuric acid is 1.1-1.

3.

4. The method according to claim 1, characterized in that, In step (2), when adding sulfuric acid to the scandium precipitate for desorption, water is added to make a slurry according to the liquid-solid mass ratio (1-2):1, and sulfuric acid is added to control the final pH to 1.0-1.

5. The method for recovering high-purity scandium oxide from the scandium-containing solution is as follows: the scandium-containing solution is extracted using C272 extractant, and the phases are separated to obtain the scandium-loaded organic phase and the raffinate. Then, oxalic acid solution is added to the scandium-loaded organic phase for precipitation, and scandium oxalate solid is separated. High-purity scandium oxide is obtained by high-temperature calcination.

5. The method according to claim 1, characterized in that, In step (3), the method for recovering high-purity nickel from the C272 raffinate is as follows: the C272 raffinate is de-oiled to obtain a de-oiled liquid with a nickel concentration of 90-100 g / L; hydrogen peroxide is added to the de-oiled liquid and electrolysis is performed to obtain high-purity nickel; the amount of hydrogen peroxide added is 5-20 L / m 3 The electrolysis conditions are: electrolytic cell temperature 55-65℃, current density 180-220 A / m², and electrode spacing 10-15 cm.

6. The method according to claim 1, characterized in that, In step (3), the method for recovering cobalt from the C272-loaded organic phase includes: introducing carbon dioxide gas into the C272-loaded organic phase for reaction, and obtaining a cobalt-loaded organic phase and a magnesium carbonate-containing slurry after phase separation; separating the magnesium carbonate-containing slurry into solid and liquid phases to obtain magnesium carbonate; adding dilute sulfuric acid to the cobalt-loaded organic phase for back-extraction, controlling the pH of the aqueous phase to be 2.0-2.5, and obtaining a cobalt sulfate solution and a regenerated organic phase after phase separation, wherein the regenerated organic phase is returned for cobalt extraction, and the cobalt sulfate solution is used to produce high-purity cobalt.

7. The method according to claim 6, characterized in that, The method for producing high-purity cobalt from the cobalt sulfate solution is as follows: The cobalt sulfate solution is subjected to deep impurity removal using D402 resin to obtain a purified solution; the purified solution is then electrolyzed to obtain high-purity cobalt; the electrolysis conditions are: electrolytic cell temperature 55-65℃, current density 180-220 A / m³. 2 The electrode spacing is 10-15 cm, the cathode is a cobalt starting plate, the anode is a titanium-plated iridium-tantalum plate, and a diaphragm-less electrode is used.

8. The method according to claim 1, characterized in that, In step (3), the method for recovering battery-grade nickel sulfate from the C272 raffinate is as follows: the C272 raffinate is passed through activated carbon for dynamic oil removal to obtain an oil-removed liquid; the oil-removed liquid is then subjected to evaporation crystallization, centrifugation separation and drying in sequence to obtain battery-grade nickel sulfate.

9. The method according to claim 1, characterized in that, In step (3), the volume concentration of both the P204 extractant and the C272 extractant is 15-25% when used.

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