A method for extracting and separating trace copper and cobalt from a super-lean vanadium-titanium magnetite flotation sulfur concentrate

Abstract

The invention discloses a method for extracting and separating trace copper and cobalt from super-lean vanadium-titanium magnetite flotation sulfur concentrate, belonging to the field of metallurgical technology. The super-lean vanadium-titanium magnetite flotation sulfur concentrate containing trace copper and cobalt is mixed with a sulfide salt and roasted in an air atmosphere to obtain a roasted material. The roasted material is acid-leached to obtain a copper-cobalt leaching solution. The copper-cobalt leaching solution is selectively extracted for copper to achieve copper-cobalt separation, obtaining a copper solution and a cobalt-containing raffinate. The copper solution is evaporated and crystallized to obtain a copper salt product. The cobalt-containing raffinate undergoes iron removal by hydrolysis, cobalt precipitation, acidolysis, and oxalate conversion to obtain a cobalt oxalate product. This method can not only achieve the efficient extraction of trace copper and cobalt from super-lean vanadium-titanium magnetite flotation sulfur concentrate, but also obtain a crystalline copper sulfate product with a purity greater than 99% and a cobalt oxalate product with a purity greater than 99.99% and a cobalt grade greater than 38%, realizing the resource utilization of super-lean vanadium-titanium magnetite flotation sulfur concentrate.

Description

Technical Field

[0001] The present invention relates to a method for extracting copper and cobalt from sulfur concentrate, and particularly to a method for extracting and separating trace copper and cobalt from flotation sulfur concentrate of ultra-lean vanadium-titanium magnetite, belonging to the field of metallurgical technology. Background Art

[0002] Copper and cobalt are widely used in fields such as aviation, electronics, electric power, machinery manufacturing, automobiles, chemical industry, new energy, and high-end equipment. The flotation sulfur concentrate of ultra-lean vanadium-titanium magnetite is usually directly used to produce sulfuric acid, which only contains trace amounts of valuable elements such as copper and cobalt, and it is difficult to extract directly. Existing flotation, pyrometallurgy, biological and other technologies are all difficult to efficiently recover copper and cobalt in flotation sulfur concentrate, especially the flotation sulfur concentrate of ultra-lean vanadium-titanium magnetite with only trace amounts of copper and cobalt. For example: flotation technology has high requirements for ore properties, the flotation process is relatively complex, and it is difficult to obtain ideal separation indexes; pyrometallurgy technology is relatively inefficient, energy-consuming, highly polluting, has high requirements for equipment, and the process is complex; biological technology has a slow leaching rate, strict environmental conditions requirements, and limited leaching rate. Summary of the Invention

[0003] Aiming at the deficiencies of the prior art, the purpose of the present invention is to provide a method for extracting and separating trace copper and cobalt from flotation sulfur concentrate of ultra-lean vanadium-titanium magnetite. This method adopts a process flow combining "roasting - leaching - extraction - hydrolysis precipitation - oxalate conversion", which can not only achieve the efficient extraction of trace copper and cobalt resources in flotation sulfur concentrate, but also obtain a crystalline copper sulfate product with a purity greater than 99% and an oxalate cobalt product with a purity greater than 99.99% and a cobalt grade greater than 38%, realizing the resource utilization of flotation sulfur concentrate containing ultra-lean vanadium-titanium magnetite.

[0004] To achieve the above technical purpose, the present invention provides a method for extracting and separating trace copper and cobalt from flotation sulfur concentrate of ultra-lean vanadium-titanium magnetite, which includes the following steps:

[0005] (1) The flotation sulfur concentrate of ultra-lean vanadium-titanium magnetite containing trace copper and cobalt is mixed with a sulfide salt and roasted in an air atmosphere to obtain a roasted material; the mass ratio of the flotation sulfur concentrate of ultra-lean vanadium-titanium magnetite containing trace copper and cobalt to the sulfide salt is 100:1 to 5; the sulfide salt is sodium sulfide;

[0006] (2) The roasted material is subjected to acid leaching to obtain a copper-cobalt leaching solution;

[0007] (3) The copper-cobalt leaching solution is selectively extracted for copper to achieve copper-cobalt separation, obtaining a copper solution and a cobalt-containing raffinate;

[0008] (4) The copper solution is obtained as a crystalline copper salt product through evaporation crystallization;

[0009] (5) The cobalt-containing raffinate is subjected to hydrolysis to remove iron, precipitate cobalt and acid hydrolyze to obtain a cobalt solution;

[0010] (6) The cobalt solution is converted into oxalate to obtain a cobalt oxalate product.

[0011] The invention aims at the distribution of copper and cobalt in the ultra-poor vanadium-titanium magnetite flotation sulfur concentrate containing trace copper and cobalt and the composition characteristics of the minerals. The key lies in using sulfide salt as a roasting aid to perform high-temperature decomposition of the ultra-poor vanadium-titanium magnetite flotation sulfur concentrate containing trace copper and cobalt under an air atmosphere. In the high-temperature process, the sulfide salt can dissociate sulfur anions, and the sulfur anions can be converted into sulfur dioxide gas, so that the materials are loose and not easy to agglomerate during the roasting process, and the strong reducing property of the sulfur anions can reduce the oxidation degree of metal ions such as cobalt, thereby facilitating the leaching effect of metal elements such as copper and cobalt. At the same time, roasting in an air atmosphere enables the sulfide ore to be fully oxidized and decomposed, which is conducive to the exposure of minerals such as copper and cobalt wrapped in the sulfide ore, and the decomposition of minerals such as copper sulfide and cobalt sulfide, and further improves the leaching effect of minerals containing copper and cobalt. In addition, since the copper and cobalt grades in the ultra-poor vanadium-titanium magnetite flotation sulfur concentrate are very low, the concentrations of copper and cobalt in the copper-cobalt leachate are very low. However, copper and cobalt can be enriched through extraction, stripping and hydrolysis to obtain high-concentration copper solution and cobalt solution, which is beneficial to the subsequent evaporation and crystallization to obtain crystalline copper sulfate products, as well as oxalate conversion to obtain cobalt oxalate products.

[0012] As a preferred solution, the copper and cobalt in the ultra-poor vanadium-titanium magnetite flotation sulfur concentrate containing trace copper and cobalt mainly exist in the form of copper sulfide and cobalt sulfide minerals, and a small amount exists in copper oxide and cobalt oxide minerals. The total copper grade in the ultra-poor vanadium-titanium magnetite flotation sulfur concentrate containing trace copper and cobalt is about 0.225%, and the total cobalt grade is about 0.096%, which is difficult to extract directly.

[0013] The mass ratio of the ultra-poor vanadium-titanium magnetite flotation sulfur concentrate containing trace copper and cobalt of the present invention to the sulfide salt is 100:1~5. If the mass ratio of the sulfide salt is too low, the material will be easily agglomerated, and the cobalt ions will be severely oxidized, which will eventually lead to incomplete leaching of copper and cobalt. As the mass ratio of the sulfide salt increases, it is beneficial to the leaching of copper and cobalt from the ultra-poor vanadium-titanium magnetite flotation sulfur concentrate containing trace copper and cobalt. However, when the mass ratio of the sulfide salt is too high, it will lead to excessive reduction, which is also unfavorable for the leaching effect of copper and cobalt. Therefore, the mass ratio of the ultra-poor vanadium-titanium magnetite flotation sulfur concentrate containing trace copper and cobalt to the sulfide salt is further preferably 100:2~4.

[0014] The sulfide salt of the present invention is sodium sulfide. Sodium sulfide can dissociate into sodium ions and sulfur anions at high temperatures. The sulfur anions are easily oxidized and converted into sulfur dioxide gas, which makes the material loose during emission and can control the oxidation degree of cobalt ions, reducing them to divalent cobalt ions that are more easily leached. The sodium ions can displace copper ions and cobalt ions. Of course, similar effects can also be achieved by using other sulfide salts, but other sulfide salts may introduce unfavorable impurity cations, which have an adverse impact on the subsequent purification of copper and cobalt. Therefore, the preferred sulfide salt of the present invention is sodium sulfide. If an auxiliary agent without sulfur anions is used, the ultrapoor vanadium-titanium magnetite flotation sulfur concentrate containing trace amounts of copper and cobalt is prone to caking and serious oxidation of cobalt ions during high-temperature roasting in an air atmosphere, resulting in low leaching rates of copper and cobalt elements in copper sulfide and cobalt sulfide ores.

[0015] As a preferred embodiment, the conditions for roasting are as follows: the roasting temperature is 500 - 700 °C, and the roasting time is 2 - 4 h. If the roasting temperature is too low, incomplete oxidation of sulfide minerals will occur, and the crystal lattices of copper sulfide minerals and cobalt sulfide minerals will not be fully destroyed, thus affecting the subsequent leaching of copper and cobalt. If the roasting temperature is too high, iron impurities will form dense ferrites with copper and cobalt, which is not conducive to leaching and is more costly. If the roasting time is too short, incomplete oxidation of sulfide minerals will occur; if the roasting time is too long, power consumption will increase, and the improvement of the leaching effect is not obvious. The further preferred roasting temperature is 600 - 650 °C. The further preferred roasting time is 2 - 3 h.

[0016] As a preferred embodiment, the conditions for acid leaching are as follows: sulfuric acid is used as the leaching agent, the solid-liquid ratio is 1 g: 4 mL to 1 g: 6 mL, the dosage of the leaching agent is 100 to 300 g / L, the stirring speed is 200 to 400 r / min, the leaching temperature is 30 to 80 °C, and the leaching time is 0.5 to 3.0 h. Based on the sulfide salt-assisted roasting of super-lean vanadium-titanium magnetite flotation sulfur concentrate containing trace amounts of copper and cobalt, by using the preferred acid leaching conditions, the leaching rate of copper can reach 90 to 100%, and the leaching rate of cobalt can reach 85 to 95%. During the acid leaching process, the dosage of the leaching agent and the leaching temperature are the main factors affecting the leaching effect. When the leaching temperature increases from room temperature to 50 °C, the leaching rates of copper and cobalt remain almost unchanged. However, when it continues to increase to 80 °C, the leaching rates of copper and cobalt increase by approximately 10%. The further preferred leaching temperature is 60 to 80 °C, the more preferred leaching temperature is 70 to 80 °C, and 80 °C is the optimal leaching temperature. When the dosage of the leaching agent increases from 100 g / L to 200 g / L, the leaching rates of copper and cobalt elements increase. When the dosage of the leaching agent is further increased to 300 g / L, the leaching rates of copper and cobalt elements slightly decrease. Therefore, the further preferred dosage of the leaching agent is 150 to 200 g / L. In comparison, the leaching effect is the best when the dosage of the leaching agent is 200 g / L. The solid-liquid ratio and the leaching time of the leaching solution also have a certain impact on the leaching effect. For example, when the solid-liquid ratio increases from 1 g: 4 mL to 1 g: 5 mL, the leaching rates of copper and cobalt increase. When the solid-liquid ratio is further increased to 1 g: 6 mL, the leaching rates of copper and cobalt do not increase significantly further. The leaching effect is the best when the solid-liquid ratio is 1 g: 5 mL. When the leaching time increases from 0.5 h to 1 h, the leaching rates of copper and cobalt increase slightly. After the leaching time is further increased to 1.5 h, the leaching rates of copper and cobalt do not increase significantly. The further preferred leaching time is 1 h to 1.5 h.

[0017] As a preferred embodiment, the process of selectively extracting copper includes extraction and stripping: The conditions for extraction are as follows: Lix984 is used as the extractant, the volume concentration of Lix984 in the organic extraction phase is 10 to 20%, the extraction phase ratio is O / A = 1: 0.5 to 2, and the extraction time is 3 to 7 min; The conditions for stripping are as follows: sulfuric acid solution is used as the stripping agent, the stripping phase ratio is A / O = 1: 0.5 to 2, and the stripping time is 2 to 11 min.

[0018] During the extraction process, the preferred Lix984 is used as a copper ion selective extractant for the extraction and separation of copper ions and cobalt ions in the copper-cobalt leaching solution. During the extraction process, the O / A has a relatively obvious influence on the extraction and separation of copper ions and cobalt ions. As the amount of the extraction aqueous phase increases, the extraction effect deteriorates. Therefore, the further preferred extraction ratio is O / A = 1:1 - 1.5. When O / A = 1:1, the extraction rate of copper ions is high and there are fewer impurities. In addition, as the extraction time increases, the extraction effect first increases and then decreases. Therefore, the extraction effect is the best at 4 - 6 minutes, and thus the optimal extraction time is 5 minutes. The extraction organic phase also contains a diluent, such as sulfonated kerosene.

[0019] During the stripping process, a sulfuric acid solution with a concentration of 150 - 200 g / L is used as the stripping agent, which can efficiently elute the copper ions in the copper-loaded organic phase to obtain a copper sulfate solution with a copper ion concentration of 8 - 10 g / L. The copper sulfate solution is further evaporated and crystallized at a temperature of 60 - 80 °C to obtain a crystalline copper sulfate product with a purity greater than 99%.

[0020] As a preferred solution, the conditions for iron removal are: adjusting the pH to 3.5 - 4.0, at a temperature of 80 - 85 °C, and the precipitation time is 0.5 - 2.5 h. Under the preferred conditions, the preferred hydrolysis precipitation of iron ion impurities is achieved, which is beneficial to obtaining a cobalt solution with a higher purity.

[0021] As a preferred solution, the conditions for cobalt precipitation are: adjusting the pH to 8.0 - 8.3, at a temperature of 60 - 70 °C, and the precipitation time is 0.25 - 1 h. Under the preferred conditions, the efficient hydrolysis precipitation of cobalt ions is achieved, which is beneficial to the enrichment of cobalt elements.

[0022] The concentration of cobalt element in the cobalt solution obtained by the present invention is 2.5 - 2.7 g / L.

[0023] The cobalt solution of the present invention is converted by adding ammonium oxalate in an amount of 2.5 - 2.7 g / L, and precipitating to obtain an ammonium oxalate cobalt product with a purity greater than 99.99% and a cobalt grade greater than 38%.

[0024] Compared with the prior art, the beneficial technical effects brought by the technical solution of the present invention:

[0025] 1. By performing a special roasting process on the ultra-lean vanadium-titanium magnetite flotation sulfur concentrate containing trace amounts of copper and cobalt, the present invention can improve the acid leaching extraction efficiency of copper and cobalt. The highest copper leaching rate can reach 95.15%, and the highest cobalt leaching rate can reach 86.55%. The concentrations of copper and cobalt in the leaching solution can reach 0.846 g / L and 0.261 g / L respectively. Compared with direct leaching, the leaching rate is greatly improved.

[0026] 2. By adopting the processes of extraction separation, evaporation crystallization, hydrolysis precipitation and oxalic acid conversion for the acid leaching solution, the present invention can obtain a copper sulfate solution with a copper concentration of 8 - 10 g / L and a cobalt-rich solution with a cobalt concentration of 2.5 - 2.7 g / L. Through evaporation crystallization and oxalic acid conversion, a crystalline copper sulfate product with a purity greater than 99% and an oxalic acid cobalt product with a purity greater than 99.99% and a cobalt grade greater than 38% can be obtained respectively, thus greatly improving the production efficiency and economic benefits of copper products and cobalt products. BRIEF DESCRIPTION OF THE DRAWINGS

[0027] Figure 1 FIG. is a process flow chart for extracting and separating copper and cobalt from a super-lean vanadium-titanium magnetite flotation sulfur concentrate containing trace amounts of copper and cobalt provided by the present invention.

[0028] Figure 2 FIG. is an XRD pattern of the super-lean vanadium-titanium magnetite flotation sulfur concentrate containing trace amounts of copper and cobalt of the present invention. DETAILED DESCRIPTION OF THE INVENTION

[0029] The following specific examples are intended to further illustrate the content of the present invention rather than limit the scope of protection of the claims.

[0030] The raw material of the super-lean vanadium-titanium magnetite flotation sulfur concentrate used in the following examples:

[0031] A certain domestic mining enterprise annually produces 18,000 tons of super-lean vanadium-titanium magnetite flotation sulfur concentrate, which contains about 0.225% copper and about 0.096% cobalt. The content of each component in the sample is shown in Table 1:

[0032]

[0033] The XRD pattern is as Figure 2 shown. It can be seen from Figure 2 that almost no characteristic peaks of copper minerals and cobalt minerals can be seen, indicating that its grade is extremely low.

[0034] Condition optimization experiment:

[0035] Experimental group 1:

[0036] Through a series of roasting tests, explore the influence of factors such as roasting temperature (500 - 700 °C), roasting time (2 - 4 h), types of roasting aids, and dosage of roasting aids on the roasting pretreatment of super-lean vanadium-titanium magnetite flotation sulfur concentrate.

[0037] The specific operation process is as follows: Weigh 50 g of ore sample (-43 μm) and 10% of the roasting aid into a porcelain boat, and put it into a muffle furnace. The roasting atmosphere is air. Set the corresponding temperature and time according to the muffle furnace rising 10 °C per 1 min. After roasting, cool it to room temperature and weigh it.

[0038] Effect of roasting temperature (500 - 700 °C) on leaching rate: Other test conditions were controlled as follows: 50 g of ore sample, roasting time of 3 h, roasting atmosphere of air, and no roasting aid. The direct leaching conditions were: temperature of 80 °C, stirring speed of 300 r / min, sulfuric acid concentration of 200 g / L, solid-liquid ratio of 1:5 g / mL, and leaching time of 1.5 h. The higher the temperature, the higher the leaching rate. When the temperature was increased to 650 °C, further increasing the temperature did not result in a significant increase in the leaching rate. Finally, 650 °C was determined as the optimal experimental condition.

[0039]

[0040] Effect of roasting time (2 - 4 h) on leaching rate: Other test conditions were controlled as follows: 50 g of ore sample, roasting temperature of 650 °C, roasting atmosphere of air, and no roasting aid. The direct leaching conditions were: temperature of 80 °C, stirring speed of 300 r / min, sulfuric acid concentration of 200 g / L, solid-liquid ratio of 1:5 g / mL, and leaching time of 1.5 h. The longer the time, the higher the leaching rate. Further increasing the time did not result in a significant increase in the leaching rate. Finally, 3 h was determined as the optimal experimental condition.

[0041]

[0042] Effect of type of roasting aid on leaching rate: Other test conditions were controlled as follows: 50 g of ore sample, roasting temperature of 650 °C, roasting atmosphere of air, roasting time of 3 h, and roasting aid dosage of 3%. The direct leaching conditions were: temperature of 80 °C, stirring speed of 300 r / min, sulfuric acid concentration of 200 g / L, solid-liquid ratio of 1:5 g / mL, and leaching time of 1.5 h. Finally, sodium sulfide was determined as the optimal roasting aid for the experiment.

[0043]

[0044] Effect of roasting aid dosage on leaching rate: Other test conditions were controlled as follows: 50 g of ore sample, roasting temperature of 650 °C, roasting atmosphere of air, roasting time of 3 h, and roasting aid of sodium sulfide. The direct leaching conditions were: temperature of 80 °C, stirring speed of 300 r / min, sulfuric acid concentration of 200 g / L, solid-liquid ratio of 1:5 g / mL, and leaching time of 1.5 h. Finally, a roasting aid dosage of 3% was determined as the optimal experimental condition.

[0045]

[0046] The roasting test study found that the optimal roasting conditions were: roasting temperature of 650 °C, roasting time of 3 h, roasting aid of sodium sulfide, roasting aid addition amount of 3%, and roasting in 1 stage.

[0047] Experimental group 2:

[0048] Through a series of direct leaching experiments on the roasted samples, the effects of conditions such as solid-liquid ratio (1g:4mL - 1g:6mL), leaching agent dosage (0 - 300g / L), stirring speed (200 - 400r / min), leaching temperature (60 - 80°C), and leaching time (0.5 - 3h) on the leaching of sulfur concentrate were investigated.

[0049] Effect of stirring speed (200 - 400r / min) on the leaching rate:

[0050] Other test conditions were controlled as follows: sulfuric acid concentration 200g / L, leaching temperature 80°C, solid-liquid ratio 1:5 g / mL, and leaching time 1.5h. The effect of stirring speed on the leaching rate was not obvious, and the leaching of copper and cobalt changed little with the increase of stirring speed. The optimal rotation speed condition was determined to be 300r / min.

[0051]

[0052] Effect of temperature (30 - 80°C) on the leaching rate:

[0053] Other test conditions were controlled as follows: stirring speed 300r / min, sulfuric acid concentration 200g / L, solid-liquid ratio 1:5 g / mL, and leaching time 1.5h. Temperature had a certain effect on the leaching of copper and cobalt elements. When the temperature increased from room temperature to 50°C, the leaching rate remained almost unchanged. However, when it continued to increase to 80°C, the leaching rates of copper and cobalt elements increased by approximately 20% and 30% respectively. Therefore, 80°C was determined as the optimal leaching temperature.

[0054]

[0055] Effect of leaching agent (sulfuric acid) concentration (0 - 300g / L) on the leaching rate:

[0056] Other test conditions were controlled as follows: stirring speed 300r / min, temperature 80°C, solid-liquid ratio 1:4 g / mL, and leaching time 1h. When the leaching agent dosage increased from 100g / L to 200g / L, the leaching rates of copper and cobalt elements increased. When the leaching agent dosage increased to 300g / L, the leaching rates of copper and cobalt elements decreased slightly. In comparison, the leaching effect was the best when the leaching agent dosage was 200g / L. Therefore, 200g / L was determined as the optimal leaching agent dosage.

[0057]

[0058] Effect of solid-liquid ratio (1:4 g / mL - 1:6 g / mL) on the leaching rate:

[0059] Other test conditions were controlled as follows: stirring speed 300 r / min, temperature 80 °C, sulfuric acid concentration 200 g / L, and leaching time 1.5 h. When the solid-liquid ratio was increased from 1:4 g / mL to 1:5 g / mL, the leaching rates of copper and cobalt increased. When the solid-liquid ratio was increased to 1:6 g / mL, the leaching rates of copper and cobalt did not increase significantly. In contrast, the leaching effect was the best at a solid-liquid ratio of 1:5 g / mL. Therefore, 1:5 g / mL was determined as the optimal solid-liquid ratio.

[0060]

[0061] Effect of leaching time (0.5 - 3.0 h) on the leaching rate:

[0062] Other test conditions were controlled as follows: stirring speed 300 r / min, temperature 80 °C, sulfuric acid concentration 200 g / L, and solid-liquid ratio 1:5 g / mL. When the leaching time was increased from 0.5 h to 1.5 h, the leaching rates of copper and cobalt increased slightly. After the leaching time was increased to 1.5 h, the leaching rates of copper and cobalt did not increase significantly. Therefore, the leaching time was determined to be 1.5 h.

[0063]

[0064] The optimal conditions for direct leaching were determined as follows: temperature 80 °C, stirring speed 300 r / min, particle size -43 μm, sulfuric acid concentration 200 g / L, solid-liquid ratio 1:5 g / mL, and leaching time 1.5 h.

[0065] The following examples were carried out under the optimal conditions obtained from the above optimization experiments:

[0066] Example 1

[0067] This example provides a method for extracting and separating trace copper and cobalt from super-lean vanadium-titanium magnetite flotation sulfur concentrate, including the following steps:

[0068] Oxidative roasting: Roasting was carried out under the conditions of roasting temperature 650 °C, roasting atmosphere air, roasting time 3 h, and sodium sulfide 3% (relative to the amount of super-lean vanadium-titanium magnetite flotation sulfur concentrate containing trace copper and cobalt). The super-lean vanadium-titanium magnetite flotation sulfur concentrate containing trace copper and cobalt contained about 0.225% copper and about 0.096% cobalt.

[0069] Leaching: Leaching was carried out for 1.5 h under the conditions of temperature 80 °C, stirring speed 300 r / min, sulfuric acid concentration 200 g / L, and solid-liquid ratio 1 g:5 mL. The copper leaching rate was 95.15%, and the cobalt leaching rate was 86.55%.

[0070] Copper extraction: 15% extractant Lix984 and 85% sulfonated kerosene as diluent were used, with an O / A ratio of 1:1 and a time of 5 min. The extraction rate was 99.59%.

[0071] Stripping of copper: using 180 g / L sulfuric acid solution, A / O = 1:1, reaction time 5 min, stripping rate 99.91%, copper concentration in the obtained copper sulfate solution was 9.3 g / L.

[0072] Evaporation crystallization: Place the copper sulfate solution on a magnetic stirrer and stir it thoroughly at a stirring speed of 200r / min. The temperature is controlled at 80°C and the reaction time is 5h to obtain a crystalline copper sulfate product with a purity greater than 99%.

[0073] Iron removal: adjust pH to 3.8, at 85℃, reaction time 2h. Iron removal rate is 99.86%, cobalt loss rate is 1.84%.

[0074] Cobalt precipitation: adjust pH to 8.1, at 65°C, reaction time 0.5h. Cobalt precipitation rate is 99.58%.

[0075] Acid dissolution: Use sulfuric acid solution with a pH of 0.5 and a temperature of 65°C to fully dissolve the cobalt enrichment. The concentration of the obtained cobalt solution is 2.57 g / L.

[0076] Precipitation: Mix the cobalt solution with an ammonium oxalate solution of pH=4.8, place the mixture on a magnetic stirrer, stir at 45°C, stir at a stirring speed of 200r / min, and react for 2h to obtain cobalt oxalate precipitation. During the precipitation process, 14~15g / L ammonium oxalate solution is used for every 100mL of cobalt solution, and a cobalt oxalate product with a purity greater than 99.99% and a cobalt grade greater than 38% is obtained by precipitation.

[0077] Although embodiments of the present invention have been shown and described, it will be appreciated by those skilled in the art that various changes, modifications, substitutions and variations may be made to the embodiments without departing from the principles and spirit of the present invention, and that the scope of the present invention is defined by the appended claims and their equivalents.

Claims

1. A method for extracting and separating trace copper and cobalt from ultra-poor vanadium-titanium magnetite flotation sulfur concentrate, characterized by: The following steps are involved: (1) The ultra-low vanadium-titanium magnetite flotation sulfur concentrate containing trace amounts of copper and cobalt is mixed with a sulfide salt and roasted in an air atmosphere to obtain a roasted material; the mass ratio of the ultra-low vanadium-titanium magnetite flotation sulfur concentrate containing trace amounts of copper and cobalt to the sulfide salt is 100:1-5; the sulfide salt is sodium sulfide; (2) acid leaching the calcined material to obtain a copper-cobalt leaching solution; (3) The copper-cobalt leaching solution is subjected to selective extraction of copper to achieve separation of copper and cobalt, thereby obtaining a copper solution and a cobalt-containing raffinate; (4) the copper solution is evaporated and crystallized to obtain a crystalline copper salt product; (5) The cobalt-containing raffinate is subjected to hydrolysis to remove iron, precipitate cobalt and acid hydrolyze to obtain a cobalt solution; (6) The cobalt solution is converted into oxalate to obtain a cobalt oxalate product.

2. The method for extracting and separating trace copper and cobalt from the flotation sulfur concentrate of ultra-poor vanadium-titanium magnetite according to claim 1, characterized in that: The copper and cobalt in the ultra-poor vanadium-titanium magnetite flotation sulfur concentrate containing trace amounts of copper and cobalt mainly exist in the form of copper sulfide minerals and cobalt sulfide minerals, and a small amount exists in copper oxide minerals and cobalt oxide minerals.

3. The method for extracting and separating trace copper and cobalt from the flotation sulfur concentrate of ultra-poor vanadium-titanium magnetite according to claim 1 or 2, characterized in that: The calcination conditions are as follows: the calcination temperature is 500-700° C. and the calcination time is 2-4 hours.

4. The method for extracting and separating trace copper and cobalt from the flotation sulfur concentrate of ultra-poor vanadium-titanium magnetite according to claim 1, characterized in that: The acid leaching conditions are: sulfuric acid is used as the leaching agent, the solid-liquid ratio is 1g:4mL~1g:6mL, the amount of the leaching agent is 100~300g / L, the stirring speed is 200~400r / min, the leaching temperature is 30~80℃, and the leaching time is 0.5~3.0h.

5. The method for extracting and separating trace copper and cobalt from the flotation sulfur concentrate of ultra-poor vanadium-titanium magnetite according to claim 1, characterized in that: The process of selectively extracting copper includes extraction and stripping: The extraction conditions are as follows: using Lix984 as the extractant, the volume concentration of Lix984 in the extracted organic phase is 10-20%, the extraction phase ratio is O / A=1:0.5-2, and the extraction time is 3-7 min; The stripping conditions are: using sulfuric acid solution as the stripping agent, the stripping phase ratio is A / O=1:0.5~2, and the stripping time is 2~11min.

6. The method for extracting and separating trace copper and cobalt from the flotation sulfur concentrate of ultra-poor vanadium-titanium magnetite according to claim 1, characterized in that: The conditions for iron removal are: adjusting the pH to 3.5-4.0, at a temperature of 80-85° C., and a precipitation time of 0.5-2.5 h.

7. The method for extracting and separating trace copper and cobalt from the flotation sulfur concentrate of ultra-poor vanadium-titanium magnetite according to claim 1, characterized in that: The cobalt precipitation conditions are as follows: adjusting the pH to 8.0-8.3, at a temperature of 60-70° C., and the precipitation time is 0.25-1 h.

Images