Method for selectively recovering nickel, cobalt and scandium from nickel laterite ore pickle liquor

Abstract

The invention provides a method for selectively recovering nickel, cobalt and scandium from a laterite-nickel ore pickle liquor, and relates to the technical field of resource recovery, the method comprises the following steps: adding magnesium bicarbonate into the laterite-nickel ore pickle liquor to obtain a multi-metal mixed precipitate; ammonia leaching is carried out, and nickel-cobalt-containing ammonia leaching liquid and scandium-containing leaching residues are obtained; ammonia is distilled, and nickel and cobalt are precipitated from the solution to obtain a nickel-cobalt enriched product; the scandium-containing leaching residues are taken to be subjected to alkaline leaching, scandium is enriched while aluminum is selectively recycled, scandium-rich residues are subjected to acid leaching, a scandium-loaded organic phase is obtained through extraction, and an Sc2O3 product is obtained through washing, reverse extraction, precipitation and roasting. According to the method, the problem of entrainment loss of nickel and cobalt during gradient separation of valuable metals in the conventional laterite-nickel ore pickle liquor is avoided, comprehensive utilization of nickel, cobalt and scandium in the laterite-nickel ore pickle liquor is achieved, the obtained nickel-cobalt enrichment is high in nickel and cobalt content and free of sulfur impurities, meanwhile, the leaching agent ammonia can be recycled, and reagent consumption is greatly reduced.

Description

Technical Field

[0001] The invention relates to the technical field of resource recovery, and in particular to a method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution. Background Art

[0002] Nickel, cobalt and scandium are all important strategic metals, widely used in national defense, aerospace, petrochemical, energy materials and other fields.

[0003] In order to achieve the comprehensive extraction of valuable components in laterite nickel ore, hydrometallurgical processes have been widely studied, among which the acid leaching process is a process that has been studied and applied more frequently, and the leaching media include sulfuric acid, nitric acid, hydrochloric acid, etc. After acid leaching, most of the valuable metals are efficiently leached, and the leachate is a complex acid leaching solution containing metals such as iron, nickel, cobalt, aluminum, scandium, and magnesium. After acid leaching, it is extremely important to separate the valuable metals from the complex leachate and prepare the products. At present, the widely used method is step-by-step neutralization precipitation, which gradually achieves the precipitation of iron, aluminum scandium, nickel cobalt, etc. by adjusting the pH of the leachate. However, during the neutralization precipitation process, the valuable metals will interact with each other and precipitate together, resulting in incomplete separation, especially in the process of neutralization precipitation of aluminum scandium, the generated aluminum hydroxide colloid will adsorb a large amount of nickel and cobalt, resulting in nickel and cobalt loss. In addition, the step-by-step neutralization precipitation process is long and the pH needs to be strictly controlled.

[0004] Chinese patent CN 117925998 A targets the nitric acid leachate of laterite nickel ore, and by adding a precipitant and a pH regulator, gradually controls the end point pH of the solution to 7.5~7.8 and 5~6 respectively to achieve the removal of iron, silicon and aluminum ions, and then adds surfactants, extractants, diluents, etc. to the solution to achieve the extraction and recovery of manganese, cobalt and nickel. Although the invention realizes the recovery of various valuable metals in the leachate, the valuable metal recovery process is long, the reagent consumption is large, and the aluminum slag obtained by removing aluminum ions is a hazardous waste containing various heavy metals. More importantly, the recovery of high-value scandium in laterite nickel ore is not considered. Laterite nickel ore acid leaching solution is a complex mixed solution containing multiple metals. The existing technology for separating and recovering valuable metals in laterite nickel ore acid leaching solution has problems such as long process, low resource utilization rate, and strict pH control. In summary, the short-process and efficient comprehensive recovery of valuable metals such as nickel, cobalt and scandium from laterite nickel ore acid leaching solution is of great significance to the comprehensive utilization of laterite nickel ore resources. Summary of the invention

[0005] The invention provides a method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution, so as to solve the technical problems in the prior art such as incomplete separation caused by interaction of valuable metals and co-precipitation, low metal extraction rate and long process.

[0006] In order to achieve the above-mentioned invention object, the technical solution provided by the present invention is as follows:

[0007] A method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution comprises the following steps:

[0008] S1, adding magnesium bicarbonate to the laterite nickel ore acid leaching solution to adjust the pH to 7.5-8.0 for neutralization precipitation, and separating the solid and liquid to obtain a multi-metal mixed precipitate containing iron, aluminum, scandium, nickel, cobalt and manganese;

[0009] S2, subjecting the multi-metal mixed precipitate to ammonia leaching to obtain an ammonia leaching solution containing nickel and cobalt and a leaching residue containing scandium;

[0010] S3, taking the nickel-cobalt-containing ammonia leaching solution to evaporate ammonia, nickel and cobalt are precipitated from the nickel-cobalt-containing ammonia leaching solution to obtain a nickel-cobalt enriched product, and the ammonia and the evaporated liquid are returned to the ammonia leaching process in S2 after absorption and liquid preparation;

[0011] S4, taking the scandium-containing leaching residue and subjecting it to alkaline leaching to selectively recover aluminum and enrich scandium at the same time, thereby obtaining a scandium-rich slag;

[0012] S5, acid leaching the scandium-rich slag to obtain an acid leaching solution, adding an extractant to the acid leaching solution to extract scandium, and obtaining a scandium-loaded organic phase;

[0013] S6. Washing, stripping, precipitating and calcining the scandium-loaded organic phase to obtain a Sc2O3 product.

[0014] Preferably, the laterite nickel ore in S1 is one or both of limonitic laterite nickel ore and magnesian laterite nickel ore.

[0015] Preferably, the nickel concentration in the laterite nickel ore acid leaching solution in S1 is ≥1.0 g / L, the cobalt concentration is ≥0.2 g / L, and the scandium concentration is ≥4 ppm.

[0016] Preferably, the reaction temperature of the neutralization precipitation process in S1 is 40° C. to 90° C., and the reaction time is 0.5 h to 2 h.

[0017] Preferably, the ammonia leaching in S2 adopts an ammonia-ammonium salt mixed solution for leaching, the leaching temperature is 40°C~80°C, the ammonia concentration is 4mol / L~6mol / L, the ammonium salt concentration is 1mol / L~3mol / L, the leaching liquid-solid ratio is 1:10-25g / mL, and the ammonia leaching time is 2h~6h.

[0018] Preferably, the ammonium salt is one or both of ammonium carbonate and ammonium bicarbonate.

[0019] Preferably, the alkaline medium in the alkali leaching process in S4 is one of sodium hydroxide or potassium hydroxide, the alkali leaching temperature is 80°C~100°C, the alkali concentration is 150g / L~250g / L, the liquid-to-solid ratio of the alkali leaching is 1:3~6g / mL, and the alkali leaching time is 1h~4h.

[0020] Preferably, in the acid leaching process described in S5, the acid medium is one or more of sulfuric acid, hydrochloric acid or nitric acid, the acid leaching temperature is 30°C~100°C, the concentration of the acid medium is 3mol / L~6mol / L, the liquid-to-solid ratio of the acid leaching is 1:3-6g / mL, and the acid leaching time is 2h~4h.

[0021] Preferably, the extractant in S5 is a sulfonated kerosene solution of a mixture of P204 and N1923, and the molar amount of N1923 in the extractant is 0.3-0.5 times the total molar amount of P204 and N1923.

[0022] Preferably, the solute concentration in the extractant in S5 is 0.15 mol / L~0.25 mol / L, the volume ratio of the extractant to the acid leaching solution is 1:8~15, and the extraction time is 10min~20min.

[0023] The method of the present invention adds magnesium bicarbonate to the acid leaching solution of laterite nickel ore for neutralization and precipitation, thereby promoting iron, aluminum, scandium, nickel, cobalt and manganese to form a multi-metal mixed precipitate; then, nickel and cobalt are selectively extracted from the mixed precipitate by ammonia leaching, and ammonia is evaporated to enrich the precipitate; finally, aluminum is extracted and scandium is enriched by combining alkaline leaching and acid leaching, and finally the preparation of high-purity Sc2O3 is achieved. Various metal ions (such as nickel² + 、Cobalt² + 、Aluminum + , Scandium + In acidic solutions, different metal ions interact with the solvent with different coordination groups. — ) are induced to form different precipitates. By adjusting the pH value, multiple metal ions can form hydroxide precipitates together under certain conditions, reducing the redissolution of single metal ions during the transfer process and improving the overall recovery rate. Multiple metal ions can form complexes or clusters during the precipitation process. When forming clusters, multiple ions of different metals can be bonded together by covalent or coordination bonds to form larger and more precipitable clusters, which enhances the capture effect of metal ions, and these particles can be better precipitated and separated. The existence of clusters not only increases the sedimentation efficiency of the precipitate, but also improves the recovery efficiency of the metal. When precipitating multiple metals, since the same precipitate contains multiple components, the interaction of the precipitates can improve its thermodynamic stability, that is, reduce the solubility of the precipitate. The stability of the precipitate is improved, and the redissolution of the metal caused by environmental changes (such as pH, temperature, etc.) is avoided, thereby significantly improving the recovery rate of the metal. In addition, when low-charged metal ions (such as nickel and cobalt) are precipitated together with high-charged metal ions (such as aluminum and scandium), their precipitation efficiency can be improved by forming mixed precipitates.

[0024] In the present invention, magnesium bicarbonate is added as a precipitant for multi-metal mixed precipitation, and the magnesium content in the generated multi-metal mixed precipitation is not high. During the alkali leaching of scandium leaching slag, a large amount of magnesium-aluminum hydrotalcite will not be formed with aluminum, thereby ensuring a high leaching rate of aluminum.

[0025] Compared with the prior art, the above technical solution has at least the following beneficial effects:

[0026] In the present invention, the multi-metals in the acid leaching solution are first precipitated together to obtain a multi-metal mixed precipitate, and then the selective recovery of nickel and cobalt is achieved through ammonia leaching and ammonia evaporation, and the evaporated ammonia and the residual ammonia evaporation liquid can be recycled. The scandium-containing slag after ammonia leaching is first alkali leached to selectively separate aluminum and simultaneously enrich scandium, and the obtained scandium-rich slag is acid leached, and the selective recovery of scandium is achieved through a solvent. The present invention avoids the entrainment loss during the separation of valuable metals in the acid leaching solution of laterite nickel ore, has the significant advantages of short process, high resource utilization, low reagent consumption, etc., and provides technical support for the comprehensive utilization of laterite nickel ore resources. BRIEF DESCRIPTION OF THE DRAWINGS

[0027] Figure 1 The present invention provides a flow chart of a method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution. DETAILED DESCRIPTION

[0028] In order to make the purpose, technical solution and advantages of the embodiment of the present invention clearer, the technical solution of the embodiment of the present invention will be clearly and completely described below in conjunction with the drawings of the embodiment of the present invention. Obviously, the described embodiment is a part of the embodiment of the present invention, not all of the embodiments. Based on the described embodiment of the present invention, all other embodiments obtained by ordinary technicians in this field without creative work are within the scope of protection of the present invention.

[0029] Example 1

[0030] like Figure 1 As shown, a method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution comprises the following steps:

[0031] (1) For the laterite nickel ore acid leaching solution obtained by pressurizing nitric acid and pre-neutralizing and removing iron from the limonitic laterite nickel ore, the nickel concentration in the laterite nickel ore acid leaching solution is ≥1.0g / L, the cobalt concentration is ≥0.2g / L, and the scandium concentration is ≥4ppm. Magnesium bicarbonate is added to adjust the pH of the leaching solution to 7.5, and the reaction is carried out at 40°C for 2h. The mixed hydroxide is obtained by filtration, and the precipitation rates of iron, aluminum, scandium, nickel, cobalt, and manganese are 100%, 99.93%, 99.96%, 99.82%, 98.56%, and 46.74%, respectively. The obtained multi-metal mixed precipitate has an iron content of 0.43%, an aluminum content of 23.76%, a scandium content of 370g / t, a nickel content of 5.43%, a cobalt content of 1.10%, a manganese content of 2.14%, and a magnesium content of 0.31%.

[0032] (2) The multi-metal mixture was leached with ammonia-ammonium carbonate, with the ammonia concentration being 6 mol / L, the ammonium carbonate concentration being 1 mol / L, the liquid-to-solid ratio being 1:15 g / mL, the leaching temperature being 60 °C, and the leaching time being 2 h. The leaching rates of nickel and cobalt were 96.78% and 97.54%, respectively, while the others were not leached.

[0033] (3) The ammonia leaching solution was evaporated at 150°C, and the evaporation volume was 30% of the initial solution volume. The nickel and cobalt precipitation rates were 95.16% and 93.86%, respectively. The nickel and cobalt content of the obtained nickel-cobalt enriched product was as high as 56.65%.

[0034] (4) The scandium-containing slag obtained after ammonia leaching was subjected to sodium hydroxide leaching. Under the conditions of leaching temperature of 80°C, sodium hydroxide concentration of 150 g / L, liquid-to-solid ratio of 1:3 g / mL and leaching time of 4 h, the aluminum leaching rate was 95.35%, and the scandium leaching rate was only 0.21%. Compared with the scandium-containing slag, the scandium in the scandium-rich slag was enriched more than 8 times.

[0035] (5) The scandium-rich slag was subsequently leached with sulfuric acid. Under the conditions of sulfuric acid concentration of 6 mol / L, liquid-solid ratio of 1:6 g / mL, leaching temperature of 30 °C, and leaching time of 2 h, the scandium leaching rate was 100%.

[0036] (6) Sulfonated kerosene solution of P204 and N1923 was used for the synergistic selective extraction of scandium from the sulfuric acid leachate. The molar amount of N1923 in the extractant was 0.3 times the total molar amount of P204 and N1923, the total solute (P204 and N1923) concentration in the extractant was 0.15 mol / L, the volume ratio of the organic phase to the aqueous phase was 1:8, and the extraction time was 20 min. The scandium extraction rate was 99.12%. The scandium-loaded organic phase was further washed, stripped, precipitated, and calcined to prepare the Sc2O3 product with a product purity of 98.47%. The raffinate was used for manganese recovery after removing iron and aluminum.

[0037] Example 2

[0038] A method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution comprises the following steps:

[0039] (1) For the leaching solution obtained by neutralizing and removing iron from magnesia-type laterite nickel ore with nitric acid at normal pressure, magnesium bicarbonate was added to adjust the pH of the leaching solution to 8.0, and the reaction was carried out at 90°C for 0.5h. The mixed hydroxide was filtered and the precipitation rates of iron, aluminum, scandium, nickel, cobalt and manganese were 100%, 99.96%, 99.87%, 99.92%, 99.03% and 49.87% respectively. The iron content in the mixed precipitate was 0.32%, the aluminum content was 21.86%, the scandium content was 190g / t, the nickel content was 4.31%, the cobalt content was 0.87%, the manganese content was 1.89% and the magnesium content was 0.27%.

[0040] (2) The multi-metal mixture was leached with ammonia-ammonium bicarbonate, with the ammonia concentration being 4 mol / L, the ammonium sulfate concentration being 2 mol / L, the liquid-to-solid ratio being 1:25 g / mL, the leaching temperature being 40 °C, and the leaching time being 6 h. The leaching rates of nickel and cobalt were 95.32% and 96.78%, respectively, while the others were not leached.

[0041] (3) The ammonia leaching solution was evaporated at 130°C, and the evaporation volume was 40% of the initial solution volume. The nickel and cobalt precipitation rates were 93.05% and 95.49%, respectively. The nickel and cobalt content of the obtained nickel-cobalt enriched product was as high as 52.37%.

[0042] (4) The scandium-containing slag obtained after ammonia leaching was subjected to potassium hydroxide leaching. Under the conditions of leaching temperature of 100°C, potassium hydroxide concentration of 250 g / L, liquid-solid ratio of 1:6 g / mL, and leaching time of 1 h, the aluminum leaching rate was 96.42%, and the scandium leaching rate was only 1.05%. Compared with the scandium-containing slag, the scandium in the scandium-rich slag was enriched more than 8 times.

[0043] (5) The scandium-rich slag was subsequently leached with nitric acid. Under the conditions of nitric acid concentration of 4 mol / L, liquid-to-solid ratio of 1:4 g / mL, leaching temperature of 80 °C, and leaching time of 2 h, the scandium leaching rate was 100%.

[0044] (6) P204 and N1923 were used to selectively extract scandium from the nitric acid leachate. The molar amount of N1923 in the extractant was 0.5 times the total molar amount of P204 and N1923, the solute concentration in the extractant was 0.25 mol / L, the volume ratio of the organic phase to the aqueous phase was 1:12, and the extraction time was 10 min. The scandium extraction rate was 99.03%. The scandium-loaded organic phase was further washed, stripped, precipitated, and calcined to prepare the Sc2O3 product with a product purity of 99.02%. The raffinate was used to recover manganese after removing iron and aluminum.

[0045] Example 3

[0046] A method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution comprises the following steps:

[0047] (1) The leachate obtained by pressurizing sulfuric acid and pre-neutralizing the limonitic laterite nickel ore to remove iron was adjusted to pH 7.5 by adding magnesium bicarbonate, reacted at 90°C for 1.5h, and filtered to obtain a mixed hydroxide. The precipitation rates of iron, aluminum, scandium, nickel, cobalt, and manganese were 99.96%, 99.04%, 99.86%, 99.32%, 98.78%, and 43.21%, respectively. The iron content in the mixed precipitate was 0.63%, the aluminum content was 18.79%, the scandium content was 870g / t, the nickel content was 6.31%, the cobalt content was 0.94%, the manganese content was 2.43%, and the magnesium content was 0.29%.

[0048] (2) The multi-metal mixture was leached with ammonia-ammonium carbonate, with the ammonia concentration being 5 mol / L, the ammonium sulfate concentration being 3 mol / L, the liquid-to-solid ratio being 1:10 g / mL, the leaching temperature being 80 °C, and the leaching time being 4 h. The leaching rates of nickel and cobalt were 95.43% and 96.05%, respectively, while the others were not leached.

[0049] (3) The ammonia leaching solution was evaporated at 130°C, and the evaporation volume was 50% of the initial solution volume. The nickel and cobalt precipitation rates were 92.04% and 96.65%, respectively. The nickel and cobalt content of the obtained nickel-cobalt enriched product was as high as 57.46%.

[0050] (4) The scandium-containing slag obtained after ammonia leaching was subjected to sodium hydroxide leaching. Under the conditions of leaching temperature of 90°C, sodium hydroxide concentration of 200 g / L, liquid-to-solid ratio of 1:5 g / mL and leaching time of 2 h, the aluminum leaching rate was 94.86%, and the scandium leaching rate was only 0.94%. Compared with the scandium-containing slag, the scandium in the scandium-rich slag was enriched more than 8 times.

[0051] (5) The scandium-rich slag was subsequently leached with hydrochloric acid. Under the conditions of hydrochloric acid concentration of 3 mol / L, liquid-to-solid ratio of 1:3 g / mL, leaching temperature of 100 °C, and leaching time of 4 h, the scandium leaching rate was 99.96%.

[0052] (6) P204 and N1923 were used to selectively extract scandium from the hydrochloric acid leachate. The molar amount of N1923 in the extractant was 0.4 times the total molar amount of P204 and N1923. The solute concentration in the extractant was 0.20 mol / L. The volume ratio of the organic phase to the aqueous phase was 1:15. The extraction time was 10 min. The scandium extraction rate was 99.64%. The scandium-loaded organic phase was further washed, stripped, precipitated, and calcined to prepare the Sc2O3 product with a product purity of 99.12%. The raffinate was used to recover manganese after removing iron and aluminum.

[0053] Example 4

[0054] A method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution comprises the following steps:

[0055] (1) For the leaching solution obtained after the hydrochloric acid atmospheric pressure leaching solution of the magnesium-type laterite nickel ore was extracted and iron was removed, magnesium bicarbonate was added to adjust the pH of the leaching solution to 8.0, and the reaction was carried out at 60°C for 2 hours. The mixed hydroxide was filtered and the precipitation rates of iron, aluminum, scandium, nickel, cobalt and manganese were 99.76%, 99.85%, 99.32%, 99.05%, 98.08% and 40.36% respectively. The iron content in the mixed precipitate was 0.31%, the aluminum content was 18.92%, the scandium content was 172g / t, the nickel content was 4.63%, the cobalt content was 0.94%, the manganese content was 46.86% and the magnesium content was 0.33%.

[0056] (2) The multi-metal mixture was leached with ammonia-ammonium carbonate, with the ammonia concentration being 6 mol / L, the ammonium carbonate concentration being 1.5 mol / L, the liquid-to-solid ratio being 1:20 g / mL, the leaching temperature being 60 °C, and the leaching time being 4 h. The leaching rates of nickel and cobalt were 95.62% and 94.67%, respectively, while the others were not leached.

[0057] (3) The ammonia leaching solution was evaporated at 140°C, and the evaporation volume was 40% of the initial solution volume. The nickel and cobalt precipitation rates were 93.56% and 95.71%, respectively. The nickel and cobalt content of the obtained nickel-cobalt enriched product was as high as 49.84%.

[0058] (4) The scandium-containing slag obtained after ammonia leaching was subjected to sodium hydroxide leaching. Under the conditions of leaching temperature of 95°C, sodium hydroxide concentration of 200 g / L, liquid-to-solid ratio of 1:4 g / mL and leaching time of 3 h, the aluminum leaching rate was 94.92% and the scandium leaching rate was only 0.86%. Compared with the scandium-containing slag, the scandium in the scandium-rich slag was enriched more than 8 times.

[0059] (5) The scandium-rich slag was subsequently leached with sulfuric acid. Under the conditions of sulfuric acid concentration of 4 mol / L, liquid-to-solid ratio of 1:3 g / mL, leaching temperature of 60 °C, and leaching time of 2 h, the scandium leaching rate was 100%.

[0060] (6) P204 and N1923 were used to selectively extract scandium from the sulfuric acid leachate. The molar amount of N1923 in the extractant was 0.3 times the total molar amount of P204 and N1923. The solute concentration in the extractant was 0.15 mol / L. The volume ratio of the organic phase to the aqueous phase was 1:10. The extraction time was 10 min. The scandium extraction rate was 98.83%. The scandium-loaded organic phase was further washed, stripped, precipitated, and calcined to prepare the Sc2O3 product with a product purity of 99.36%. The raffinate was used to recover manganese after removing iron and aluminum.

[0061] Example 5

[0062] A method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution comprises the following steps:

[0063] (1) The leachate in Example 1 was added with magnesium bicarbonate to adjust the pH of the leachate to 8.0, reacted at 60°C for 2h, and filtered to obtain a mixed hydroxide, wherein the precipitation rates of iron, aluminum, scandium, nickel, cobalt, and manganese were 100%, 99.86%, 99.91%, 99.74%, 99.02%, and 51.96%, respectively. The iron content in the mixed precipitate was 0.56%, the aluminum content was 21.63%, the scandium content was 365g / t, the nickel content was 5.31%, the cobalt content was 1.06%, the manganese content was 3.04%, and the magnesium content was 0.34%.

[0064] (2) The multi-metal mixture was leached with ammonia-ammonium carbonate, with the ammonia concentration being 6 mol / L, the ammonium carbonate concentration being 1 mol / L, the liquid-to-solid ratio being 1:15 g / mL, the leaching temperature being 60 °C, and the leaching time being 2 h. The leaching rates of nickel and cobalt were 96.52% and 97.03%, respectively, while the others were not leached.

[0065] (3) The ammonia leaching solution was evaporated at 140°C, and the evaporation volume was 30% of the initial solution volume. The nickel and cobalt precipitation rates were 94.15% and 92.93%, respectively. The nickel and cobalt content of the obtained nickel-cobalt enriched product was as high as 54.82%.

[0066] (4) The scandium-containing slag obtained after ammonia leaching was subjected to potassium hydroxide leaching. Under the conditions of leaching temperature of 90°C, sodium hydroxide concentration of 150 g / L, liquid-to-solid ratio of 1:5 g / mL and leaching time of 3 h, the aluminum leaching rate was 95.75%, and the scandium leaching rate was only 0.86%. Compared with the scandium-containing slag, the scandium in the scandium-rich slag was enriched more than 8 times.

[0067] (5) The scandium-rich slag was subsequently leached with hydrochloric acid. Under the conditions of hydrochloric acid concentration of 6 mol / L, liquid-solid ratio of 1:6 g / mL, leaching temperature of 30 °C, and leaching time of 2 h, the scandium leaching rate was 100%.

[0068] (6) P204 and N1923 were used to selectively extract scandium from the hydrochloric acid leachate. The molar amount of N1923 in the extractant was 0.3 times the total molar amount of P204 and N1923. The solute concentration in the extractant was 0.15 mol / L. The volume ratio of the organic phase to the aqueous phase was 1:10. The extraction time was 20 min. The scandium extraction rate was 99.43%. The scandium-loaded organic phase was further washed, stripped, precipitated, and calcined to prepare the Sc2O3 product with a product purity of 99.45%. The raffinate was used to recover manganese after removing iron and aluminum.

[0069] Comparative Example 1

[0070] This comparative example is the same as Example 1, except that magnesium bicarbonate is replaced with magnesium carbonate in this comparative example.

[0071] In this comparative example, during the multi-metal mixed precipitation process, the precipitation rates of iron, aluminum, scandium, nickel, cobalt, and manganese are 100%, 99.96%, 99.93%, 99.44%, 99.61%, and 71.37%, respectively. The iron content in the multi-metal mixed precipitation is 0.42%, the aluminum content is 18.47%, the scandium content is 264g / t, the nickel content is 4.28%, the cobalt content is 0.96%, the manganese content is 2.97%, and the magnesium content is 3.64%;

[0072] During the ammonia-ammonium carbonate leaching process, the leaching rates of nickel and cobalt were 95.86% and 96.54%, respectively;

[0073] During the alkaline leaching process, the aluminum leaching rate was only 76.89%, and the scandium leaching rate was 0.46%.

[0074] It can be seen that the use of magnesium carbonate precipitation will result in a higher magnesium content in the multi-metal mixed precipitate. The high content of magnesium will form a large amount of magnesium-aluminum hydrotalcite with aluminum during alkaline leaching of scandium-containing slag, reducing the aluminum leaching rate and making it impossible to achieve efficient and selective recovery of aluminum and enrichment of scandium.

[0075] Comparative Example 2

[0076] This comparative example is the same as Example 2, except that the magnesium bicarbonate in Example 2 is replaced by magnesium oxide in this comparative example.

[0077] In this comparative example, during the multi-metal mixed precipitation process, the precipitation rates of iron, aluminum, scandium, nickel, cobalt, and manganese are 100%, 100%, 99.97%, 99.64%, 99.71%, and 74.74%, respectively. The iron content in the mixed precipitation is 0.47%, the aluminum content is 19.81%, the scandium content is 194g / t, the nickel content is 4.02%, the cobalt content is 0.73%, the manganese content is 1.96%, and the magnesium content is 6.31%;

[0078] During the ammonia-ammonium carbonate leaching process, the leaching rates of nickel and cobalt were 94.93% and 94.61%, respectively;

[0079] During the alkaline leaching process, the aluminum leaching rate was 46.89%, and the scandium leaching rate was only 0.35%.

[0080] It can be seen that the use of magnesium oxide precipitation will result in a higher magnesium content in the multi-metal mixed precipitate. The high content of magnesium will form a large amount of magnesium-aluminum hydrotalcite with aluminum during alkaline leaching of scandium-containing slag, reducing the aluminum leaching rate and making it impossible to achieve efficient and selective recovery of aluminum and enrichment of scandium.

[0081] The above are only specific embodiments of the present invention, but the protection scope of the present invention is not limited thereto. The protection scope of the present invention shall be based on the protection scope of the claims.

Claims

1. A method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution, characterized in that: The following steps are involved: S1, adding magnesium bicarbonate to the laterite nickel ore acid leaching solution to adjust the pH to 7.5-8.0 for neutralization precipitation, and separating the solid and liquid to obtain a multi-metal mixed precipitate containing iron, aluminum, scandium, nickel, cobalt and manganese; S2, subjecting the multi-metal mixed precipitate to ammonia leaching to obtain an ammonia leaching solution containing nickel and cobalt and a leaching residue containing scandium; S3, taking the nickel-cobalt-containing ammonia leaching solution to evaporate ammonia, nickel and cobalt are precipitated from the nickel-cobalt-containing ammonia leaching solution to obtain a nickel-cobalt enriched product, and the ammonia and the evaporated liquid are returned to the ammonia leaching process in S2 after absorption and liquid preparation; S4, taking the scandium-containing leaching residue and performing alkaline leaching to selectively recover aluminum and enrich scandium to obtain scandium-rich slag; S5, acid leaching the scandium-rich slag to obtain an acid leaching solution, adding an extractant to the acid leaching solution to extract scandium, and obtaining a scandium-loaded organic phase; S6. Washing, stripping, precipitating and calcining the scandium-loaded organic phase to obtain a Sc2O3 product.

2. The method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution according to claim 1, characterized in that: The laterite nickel ore described in S1 is one or both of limonitic laterite nickel ore and magnesian laterite nickel ore.

3. The method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution according to claim 1, characterized in that: The nickel concentration in the laterite nickel ore acid leaching solution described in S1 is ≥1.0 g / L, the cobalt concentration is ≥0.2 g / L, and the scandium concentration is ≥4 ppm.

4. The method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution according to claim 1, characterized in that: The reaction temperature of the neutralization precipitation process in S1 is 40°C to 90°C, and the reaction time is 0.5h to 2h.

5. The method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution according to claim 1, characterized in that: The ammonia leaching described in S2 uses an ammonia-ammonium salt mixed solution for leaching, the leaching temperature is 40°C~80°C, the ammonia concentration is 4mol / L~6mol / L, the ammonium salt concentration is 1mol / L~3mol / L, the leaching liquid-solid ratio is 1:10-25g / mL, and the ammonia leaching time is 2h~6h.

6. The method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution according to claim 5, characterized in that: The ammonium salt is one or both of ammonium carbonate and ammonium bicarbonate.

7. The method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution according to claim 1, characterized in that: In the alkali leaching process described in S4, the alkaline medium is one of sodium hydroxide or potassium hydroxide, the alkali leaching temperature is 80°C~100°C, the alkali concentration is 150g / L~250g / L, the liquid-to-solid ratio of the alkali leaching is 1:3~6g / mL, and the alkali leaching time is 1h~4h.

8. The method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution according to claim 1, characterized in that: In the acid leaching process described in S5, the acid medium is one or more of sulfuric acid, hydrochloric acid or nitric acid, the acid leaching temperature is 30°C~100°C, the concentration of the acid medium is 3 mol / L~6mol / L, the liquid-to-solid ratio of the acid leaching is 1:3-6g / mL, and the acid leaching time is 2h~4h.

9. The method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution according to claim 1, characterized in that: The extractant in S5 is a sulfonated kerosene solution of a mixture of P204 and N1923, and the molar amount of N1923 in the extractant is 0.3-0.5 times the total molar amount of P204 and N1923.

10. The method for selectively recovering nickel, cobalt and scandium from laterite nickel ore acid leaching solution according to claim 1, characterized in that: The solute concentration in the extractant in S5 is 0.15 mol / L~0.25 mol / L, the volume ratio of the extractant to the acid leaching solution is 1:8~15, and the extraction time is 10min~20min.

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