Oxygen pressure leaching method for copper-cobalt slag

Abstract

The invention relates to an oxygen pressure leaching method of copper and cobalt slag, which comprises the following steps: after mixing the copper and cobalt slag to be treated into pulp, grinding the pulp to obtain ore pulp; after the ore pulp and concentrated sulfuric acid are mixed and react, mixed slurry is obtained; and after oxygen pressure leaching is conducted on the mixed slurry, solid-liquid separation is conducted, and oxygen pressure leaching liquid and oxygen pressure leaching residues are obtained. According to the method, the ore pulp is treated with the concentrated sulfuric acid firstly, then oxygen pressure leaching is conducted, the cobalt leaching rate can be effectively increased, and the iron leaching rate can be reduced.

Description

Technical Field

[0001] This invention relates to an oxygen pressure leaching method for copper-cobalt slag, belonging to the field of hydrometallurgy. Background Technology

[0002] Currently, most copper smelting is done using pyrometallurgical methods, which are divided into two main categories: flash smelting and pool smelting. Pyrometallurgical copper smelting has the advantages of high copper recovery rate and comprehensive recovery of associated valuable metals (gold, silver, cobalt, nickel, etc.). In particular, the use of slow-cooling flotation process for copper smelting slag can reduce the copper content of the waste slag to <0.3%, essentially achieving "complete utilization." However, before the adoption of slow-cooling flotation process, copper smelting slag was water-quenched, resulting in a higher copper content of about 1%. For copper concentrate rich in cobalt, pyrometallurgical smelting will result in cobalt loss, with the waste slag (i.e., copper-cobalt slag) having a high cobalt content of about 0.8%, low sulfur content, and impurities mainly consisting of iron and silica-calcium gangue components. This type of waste slag was generated before the adoption of slow-cooling flotation process and is found in large quantities in copper-rich areas such as Africa. Because the water-quenched slag produced by pyrometallurgical processes is characterized by the tight encapsulation of copper, cobalt, iron, and gangue, it is impossible to recover valuable metals such as copper and cobalt from the waste slag using flotation. Compared with high-temperature high-acid leaching (HPAL) of laterite nickel ore (Ni: ~1.5%, Co: ~0.1%, Fe: 20-40%), laterite nickel ore containing about 2% Ni+Co has economic value for recovery. Similarly, copper-cobalt slag containing about 2% Cu+Co also has economic value for recovery.

[0003] However, the leaching rates of copper and cobalt in copper-cobalt slag are low when using the traditional oxygen pressure leaching process. Furthermore, the high iron content in copper-cobalt slag leads to scale buildup inside the autoclave during oxygen pressure leaching, affecting normal production. Summary of the Invention

[0004] In view of the shortcomings of the prior art, the purpose of this invention is to provide an oxygen pressure leaching method for copper-cobalt slag, so as to improve the leaching rate of copper and cobalt in copper-cobalt slag.

[0005] To solve the above-mentioned technical problems, the technical solution of the present invention is as follows:

[0006] An oxygen pressure leaching method for copper-cobalt slag includes the following steps:

[0007] S1. After preparing the copper-cobalt slag to be treated into a slurry, grind it to obtain a slurry.

[0008] S2. The slurry is mixed with concentrated sulfuric acid and reacted to obtain a mixed slurry;

[0009] S3. After oxygen pressure leaching of the mixed slurry, solid-liquid separation is performed to obtain oxygen pressure leaching solution and oxygen pressure leaching residue.

[0010] Thus, in this invention, after grinding to obtain the slurry, the slurry is first mixed with concentrated sulfuric acid. Concentrated sulfuric acid has strong oxidizing properties and releases a large amount of heat, allowing oxides such as iron oxide, copper oxide, and cobalt oxide in the minerals, as well as silicate minerals, to fully react with the sulfuric acid. This also allows sulfide minerals such as cobalt sulfide to be better exposed, preparing for the subsequent oxygen pressure leaching process. Simultaneously, some of the added concentrated sulfuric acid is fully consumed by the oxides, so the actual sulfuric acid concentration in the resulting mixed slurry is not high. This prepares for the simultaneous iron removal via hematite removal during the subsequent oxygen pressure leaching process. Therefore, this invention further utilizes oxygen pressure leaching in a high-pressure reactor. As the reaction proceeds, the copper-cobalt slag consumes both the sulfuric acid in the mixed slurry and the sulfuric acid produced from hematite iron removal. This allows valuable metals such as copper and cobalt to be fully leached, while also allowing iron to enter the slag phase, reducing the iron leaching rate. This facilitates subsequent treatment of the oxygen pressure leaching solution and helps reduce the loss of valuable metals such as copper and cobalt in subsequent treatment processes.

[0011] The main chemical reaction formulas in this invention are as follows:

[0012] CuFe2O4+4H2SO4=CuSO4+Fe2(SO4)3+4H2O (1)

[0013] CuO + H₂SO₄ = CuSO₄ + H₂O (2)

[0014] CoO + H₂SO₄ = CoSO₄ + H₂O (3)

[0015] CoS+H2SO4+1 / 2O2=CoSO4+H2O+S (4)

[0016] Fe2O3+3H2SO4=Fe2(SO4)3+3H2O (5)

[0017] Fe2(SO4)3+(x+3)H2O=Fe2O3·xH2O+3H2SO4 (6)

[0018] Furthermore, in S1, the proportion of minerals with a particle size of less than 40 μm in the solid phase of the slurry is ≥90%.

[0019] Further, in S1, the copper-cobalt slag to be treated is mixed with water at a mass ratio of 60-70:30-40 to form a slurry, and then ground.

[0020] Further, in S2, the amount of concentrated sulfuric acid added is 50-70 wt% of the dry weight of the copper-cobalt slag corresponding to the slurry, preferably 55-65 wt%, more preferably 58-62 wt%; the concentration of the concentrated sulfuric acid is ≥80 wt%, preferably 85-99 wt%, more preferably 90-99 wt%.

[0021] Furthermore, in S2, the reaction time is 0.5-1h.

[0022] Furthermore, between S2 and S3, there is also a step of adjusting the liquid-solid ratio of the mixed slurry to 1.5-2:1 (mL:g); preferably, one or two of the following are used to adjust the liquid-solid ratio: water or magnesium-removed liquid.

[0023] Optionally, the concentration of sulfuric acid in the mixed slurry after adjusting the liquid-solid ratio is 40-50 g / L.

[0024] Furthermore, in S3, during oxygen pressure leaching, the temperature is controlled at 180-190℃, preferably 182-188℃, the total pressure is 1-1.3 MPa, preferably 1.1-1.2 MPa, and the oxygen partial pressure is 0.2-0.3 MPa, preferably 0.22-0.28 MPa.

[0025] Furthermore, in S3, the oxygen pressure leaching time is 1.5-2 hours, preferably 1.6-1.8 hours.

[0026] Furthermore, in S3, during oxygen pressure leaching, oxygen-enriched gas with an oxygen concentration of 90 vol% or more is used, preferably oxygen-enriched gas with an oxygen concentration of 95-99 vol%.

[0027] Further, the oxygen pressure leachate is mixed with calcium hydroxide and reacted. When the pH of the reaction system is 2.5-3.5, the solid and liquid are separated to obtain iron slag and iron-removed liquid. Optionally, the reaction temperature is controlled at 80-90℃ during the reaction. Generally, the reaction time is 3-4 hours.

[0028] After extracting copper from the iron-removed liquid by extraction, a raffinate is obtained; optionally, the extractant is an alkane or an olefin; preferably, a two-stage extraction is performed; optionally, the organic phase obtained by extraction is back-extracted with a copper electrodeposition lean solution to obtain a copper electrodeposition rich solution, which is then sent to the copper electrodeposition production process.

[0029] The raffinate is mixed with magnesium oxide and reacted. When the pH of the reaction system is 7-8, the solid and liquid are separated to obtain crude cobalt hydroxide product and cobalt precipitation liquid. Optionally, the reaction temperature is controlled at 60-70℃, and the reaction time is generally 3-4 hours.

[0030] Further, the cobalt-precipitated liquid is mixed with calcium hydroxide and reacted. When the pH of the reaction system is 10-11, the solid and liquid are separated to obtain calcium-magnesium slag and magnesium-removed liquid. Preferably, the magnesium-removed liquid is returned to S1 and / or S3. Optionally, the reaction time is 1.5-2 hours. Optionally, the calcium-magnesium slag is washed with water and then stored.

[0031] Further, in S3, after oxygen pressure leaching of the mixed slurry, the temperature and pressure are reduced, and solid-liquid separation is performed to obtain oxygen pressure leaching solution and oxygen pressure leaching residue.

[0032] Optionally, the oxygen pressure leaching residue can be stored or sold after being washed with water.

[0033] Further, in S3, an oxygen pressure leaching system is used for oxygen pressure leaching. This system includes a first pipeline, N high-pressure reactors, a second pipeline, a third pipeline, a fourth pipeline, N ninth valves, tenth valves, twelfth valves, a flash tank, and a solid-liquid separation unit. The flash tank is connected to the solid-liquid separation unit. Each high-pressure reactor includes a tank body with an inlet, a outlet, an air inlet, and a manhole. A stirring mechanism is installed inside the tank body. Each inlet is connected in parallel to the first pipeline, each outlet is connected in parallel to the third pipeline, and each air inlet is connected in parallel to the second pipeline. A fourth valve and a seventh valve are located between the first pipeline and the inlet. The inlet, fourth valve, and seventh valve are sequentially connected to the first pipeline. The pipeline between the fourth and seventh valves is connected to the third pipeline. A third valve is located between the second pipeline and the air inlet. A fifth valve is located between the third pipeline and the outlet. One port of the ninth valve is connected to the pipeline between the fifth valve and the third pipeline, and the other port of the ninth valve is connected to the fourth pipeline.

[0034] An eleventh valve is installed on the third pipeline between the inlet and outlet of the same tank; an eighth valve is installed on the third pipeline between two adjacent tanks.

[0035] One port of the tenth valve is connected to one end of the third pipeline, one port of the twelfth valve is connected to the other end of the third pipeline, and the other port of the tenth valve and the other port of the twelfth valve are connected in parallel to the inlet end of the flash tank.

[0036] Where N is an integer ≥ 2;

[0037] Preferably, the solid-liquid separation unit is a thickener;

[0038] Preferably, N is 4-6;

[0039] Preferably, the autoclave is a vertical autoclave.

[0040] Optionally, the copper-cobalt slag contains 0.4-1.5% Cu, 0.5-1.2% Co, 15-25% Fe, and 0.1-0.8% S.

[0041] Optionally, the copper-cobalt slag contains 0.6-1.2% Cu, 0.7-0.9% Co, 18-22% Fe, and 0.3-0.5% S.

[0042] Furthermore, the main phase composition of the copper-cobalt slag is: magnetite 31-35%, fir olivine 33-38%, pyroxene 10-16%, glassy phase 8-12%, and matte 1-5%.

[0043] Generally, the high cobalt content in copper-cobalt slag is due to the high cobalt content in the copper concentrate, making it highly valuable for recovery. Since the copper grade in the copper concentrate is much higher than the cobalt grade, during pyrometallurgical copper smelting, most of the sulfur bound to copper is removed, while only a small portion of the sulfur bound to cobalt is removed. Therefore, this invention employs oxygen pressure leaching to convert cobalt sulfide into cobalt sulfate, aiming to recover cobalt.

[0044] Compared with the prior art, the beneficial effects of the present invention are as follows:

[0045] (1) The present invention first treats the slurry with concentrated sulfuric acid and then performs oxygen pressure leaching, which can effectively improve the cobalt leaching rate and ensure a low iron leaching rate.

[0046] (2) In this invention, under high temperature conditions of 180–190°C and the presence of oxygen, Fe in the solution... 3+ A hydrolysis reaction will occur, reducing the iron content in the solution (containing 5-9 g / L of Fe), which is beneficial for subsequent iron removal processes. This will also reduce the loss of copper and cobalt.

[0047] (4) The present invention uses a smaller liquid-to-solid ratio (1.5-2 mL: 1 g) during oxygen pressure leaching, which can increase the copper and cobalt content in the solution, making it easier to enrich and recover copper and cobalt in the future. At the same time, it helps to reduce the volume of the system and reduce equipment investment.

[0048] (5) In the oxygen pressure leaching process of the present invention, due to the small solid-liquid ratio of the leaching liquid, there will be mineral slurry deposition and scaling in the high pressure vessel (such as the formation of hematite slag during the process). Therefore, by using high pressure vessels in series, the deposited slag can be cleaned periodically through the manhole of a single high pressure vessel. Each high pressure vessel can serve as a backup for the others, ensuring the continuous operation of other high pressure vessels and helping to improve processing efficiency.

[0049] (6) Under high temperature, high acid and oxygen pressure conditions, the copper leaching rate can reach more than 88%, and the cobalt leaching rate can reach more than 79%. Attached Figure Description

[0050] Figure 1 This is a flow chart of an oxygen pressure leaching process for copper-cobalt slag according to the present invention.

[0051] Figure 2 This is a simplified structural diagram of an oxygen pressure leaching system according to the present invention.

[0052] Figure 3 This is a digital photograph of the oxygen pressure leaching residue obtained in Example 1 of the present invention. Detailed Implementation

[0053] The present invention will be described in detail below with reference to embodiments. It should be noted that, unless otherwise specified, the embodiments and features described in the embodiments of the present invention can be combined with each other. For ease of description, the terms "upper," "lower," "left," and "right" appearing below only indicate that they correspond to the upper, lower, left, and right directions of the accompanying drawings themselves, and do not limit the structure. Unless otherwise specified, the relevant percentages refer to mass percentages.

[0054] Example 1

[0055] See Figure 1 The oxygen pressure leaching method for copper-cobalt slag in this embodiment includes the following steps:

[0056] S1. After preparing 400g of copper-cobalt slag (containing 0.92% Cu, 0.72% Co, 18% Fe, and 0.3% S) into a slurry, grind it to obtain a slurry.

[0057] S2. The slurry is mixed with 240g of concentrated sulfuric acid and reacted to obtain a mixed slurry;

[0058] S3. After oxygen pressure leaching of the mixed slurry, the temperature and pressure are reduced, followed by solid-liquid separation to obtain 800 mL of oxygen pressure leaching solution (containing 4.08 g / L Cu, 2.84 g / L Co, and 8.63 g / L Fe) and 465 g of oxygen pressure leaching residue (containing 0.09% Cu, 0.13% Co, and 14% Fe; see [reference]). Figure 3 The red color indicates that the iron has settled in the form of hematite.

[0059] In S1, the proportion of minerals with a particle size of less than 40 μm in the solid phase of the slurry is 90%.

[0060] In S1, the copper-cobalt slag to be treated is mixed with water at a mass ratio of 65:35 to form a slurry, and then ground.

[0061] In S2, the amount of concentrated sulfuric acid added is 60 wt% of the dry weight of the copper-cobalt slag corresponding to the slurry; the concentration of the concentrated sulfuric acid is 98 wt%.

[0062] Between S2 and S3, there is also a step of adding magnesium-removing liquid to adjust the liquid-to-solid ratio (volume-to-mass ratio, mL:g) of the mixed slurry to 2:1.

[0063] In S3, during oxygen pressure leaching, the temperature is controlled at 190°C, the total pressure is 1.3 MPa, the oxygen partial pressure is 0.3 MPa, and the oxygen pressure leaching time is 2 hours; during this period, oxygen-enriched gas with an oxygen concentration of 99 vol% is introduced.

[0064] The oxygen pressure leachate was mixed and reacted with calcium hydroxide emulsion at 88°C. When the pH of the reaction system reached 2.8, the solid and liquid were separated to obtain iron slag and iron-removed liquid.

[0065] After extracting copper from the iron-removed liquid, a raffinate is obtained.

[0066] The raffinate was mixed with magnesium oxide emulsion at 67°C and reacted. When the pH of the reaction system reached 7.6, the solid and liquid were separated to obtain crude cobalt hydroxide product and cobalt precipitation liquid.

[0067] The cobalt-precipitated liquid is mixed with calcium hydroxide powder and reacted. When the pH of the reaction system reaches 10.6, the solid and liquid are separated to obtain calcium-magnesium slag and magnesium-removed liquid. The magnesium-removed liquid is returned to S1 and S3.

[0068] See Figure 2 In S3, an oxygen pressure leaching system is used for oxygen pressure leaching. This system includes a pump 1, a first valve 8 and a second valve 9, a first pipeline 5, five high-pressure reactors 2, a second pipeline 6, a third pipeline 7, a fourth pipeline 19, five ninth valves 16, tenth valves 17, twelfth valves 20, a sixth valve 13, a flash tank 3, and a solid-liquid separation unit 4. The flash tank 3 is connected to the solid-liquid separation unit 4, and the outlet of the pump 1 is connected to the first pipeline 5. The high-pressure reactor 2 includes a tank body with an inlet, a outlet, an air inlet, an exhaust outlet (not shown), and a manhole (not shown). The inlet and outlet are located at the top of the tank body, and the air inlet is located at the bottom. The tank body contains... The mixing mechanism has inlets connected in parallel to a first pipeline 5, outlets connected in parallel to a third pipeline 7, and air inlets connected in parallel to a second pipeline 6. A fourth valve 11 and a seventh valve 14 are provided between the first pipeline 5 and the inlets. The inlets, fourth valve 11, seventh valve 14, and first pipeline 5 are sequentially connected. The pipeline between fourth valve 11 and seventh valve 14 is connected to the third pipeline 7. A third valve 10 is provided between the second pipeline 6 and the air inlets. A fifth valve 12 is provided between the third pipeline 7 and the outlets. One port of a ninth valve 16 is connected to the pipeline between the fifth valve 12 and the third pipeline 7, and the other port of the ninth valve 16 is connected to the fourth pipeline 19.

[0069] An eleventh valve 18 is provided on the third pipeline 7 between the inlet and outlet of the same tank. The eleventh valve 18 is located between the connection position of the fourth valve 11 and the third pipeline 7 of the corresponding tank and the connection position of the fifth valve 12 and the third pipeline 7. In two adjacent tanks, an eighth valve 15 is provided on the third pipeline 7 between the outlet of the preceding tank and the inlet of the following tank.

[0070] One port of the tenth valve 17 is connected to one end of the third pipeline 7, and one port of the twelfth valve 20 is connected to the other end of the third pipeline 7. The other port of the tenth valve 17 and the other port of the twelfth valve 20 are connected in parallel to the inlet of the sixth valve 13. The outlet of the sixth valve 13 is connected to the inlet of the flash tank 3. The solid-liquid separation unit 4 is a thickener. The autoclave is a vertical autoclave. The outlet of the first valve 8 and the outlet of the second valve 9 are connected in parallel to one end of the second pipeline 6. A steam source can be connected to the inlet of the first valve 8, and an oxygen source can be connected to the inlet of the second valve 9 to supply oxygen and steam to meet the needs of reaction and heating. The sixth valve 13, the flash tank 3, and the solid-liquid separation unit are connected in sequence.

[0071] The specific usage method of the above-mentioned oxygen pressure leaching treatment system includes the following steps: (1) Open the first valve 8 (regulating valve) and the second valve 9 (regulating valve), open the third valve 10, adjust the steam flow rate and the oxygen-enriched gas flow rate, and control the temperature and pressure of the high-pressure reactor to the target value. (2) Open the pump 1 (pressurizing pump) and the seventh valve 14 and the fourth valve 11 of the first high-pressure reactor to pump the slurry into the first high-pressure reactor 2. (3) Open the fifth valve 12, the eighth valve 15 and the sixth valve 13 of the first high-pressure reactor 2, and after the leaching slurry that has been processed by the second, third, fourth and fifth high-pressure reactors 2 in sequence is cooled and depressurized by the flash tank 3, it is thickened by the thickener to obtain oxygen pressure leaching solution and oxygen pressure leaching residue.

[0072] When it is necessary to clean the slurry deposits and scale in the No. 1 high-pressure reactor, close the fourth valve 11, the fifth valve 12, and the third valve 10 of the No. 1 high-pressure reactor, short-circuit the No. 1 high-pressure reactor, and connect the feed pipes and discharge pipes of the No. 2, No. 3, No. 4, and No. 5 high-pressure reactors to allow the other four high-pressure reactors to operate continuously. After the No. 1 high-pressure reactor is cleaned, open the ninth valve 16 of the No. 5 high-pressure reactor, and open the ninth valve 16, the fifth valve 12, the fourth valve 11, and the tenth valve 17 of the No. 1 high-pressure reactor. The mixed slurry flowing out after being processed by the No. 2, No. 3, No. 4, No. 5, and No. 1 high-pressure reactors 2 in sequence is then cooled and depressurized in the flash evaporator 3 and thickened to obtain oxygen pressure leaching solution and oxygen pressure leaching residue.

[0073] Similarly, when other autoclaves need to be cleaned or repaired, the relevant valves of the autoclave can be closed, allowing the other autoclaves to be connected in series and run continuously, thereby improving processing efficiency.

[0074] Calculations show that the copper leaching rate is 88.63% and the cobalt leaching rate is 79.01%.

[0075] Example 2

[0076] The oxygen pressure leaching method for copper-cobalt slag in this embodiment includes the following steps:

[0077] S1. After preparing 400g of copper-cobalt slag (containing 0.67% Cu, 0.81% Co, 21% Fe, and 0.5% S) into a slurry, grind it to obtain a slurry.

[0078] S2. The slurry is mixed with 200g of concentrated sulfuric acid and reacted to obtain a mixed slurry;

[0079] S3. After oxygen pressure leaching of the mixed slurry, the temperature and pressure are reduced, and then solid-liquid separation is performed to obtain 600 mL of oxygen pressure leaching solution (containing 3.85 g / L Cu, 4.17 g / L Co, and 8.25 g / L Fe) and 460 g of oxygen pressure leaching residue (containing 0.08% Cu, 0.16% Co, and 17% Fe).

[0080] In S1, the proportion of minerals with a particle size of less than 40 μm in the solid phase of the slurry is 90%.

[0081] In S1, the copper-cobalt slag to be treated is mixed with water at a mass ratio of 68:32 to form a slurry, and then ground.

[0082] In S2, the amount of concentrated sulfuric acid added is 50 wt% of the dry weight of the copper-cobalt slag corresponding to the slurry; the concentration of the concentrated sulfuric acid is 98 wt%.

[0083] Between S2 and S3, there is also a step of adding magnesium-removing liquid to adjust the liquid-solid ratio of the mixed slurry to 1.5:1.

[0084] In S3, during oxygen pressure leaching, the temperature is controlled at 180°C, the total pressure is 1.0 MPa, the oxygen partial pressure is 0.2 MPa, and the oxygen pressure leaching time is 1.5 h; during this period, oxygen-enriched gas with an oxygen concentration of 99 vol% is introduced.

[0085] The oxygen pressure leachate is mixed with calcium hydroxide and reacted. When the pH of the reaction system reaches 3.0, the solid and liquid are separated to obtain iron slag and iron-removed liquid.

[0086] After extracting copper from the iron-removed liquid, a raffinate is obtained.

[0087] The raffinate was mixed with magnesium oxide and reacted. When the pH of the reaction system reached 7.5, the solid and liquid were separated to obtain crude cobalt hydroxide product and cobalt precipitation liquid.

[0088] The cobalt-precipitated liquid is mixed with calcium hydroxide and reacted. When the pH of the reaction system reaches 10.8, the solid and liquid are separated to obtain calcium-magnesium slag and magnesium-removed liquid. The magnesium-removed liquid is returned to S1 and S3.

[0089] In S3, oxygen pressure leaching is performed using the oxygen pressure leaching system described in Example 1.

[0090] Calculations show that the copper leaching rate is 86.27% and the cobalt leaching rate is 77.28%.

[0091] Comparative Example 1

[0092] Example 1 was repeated, except that in S3, no oxygen-enriched gas was introduced, the temperature was controlled at 190°C, and the pressure was 1.0 MPa, resulting in 800 mL of oxygen pressure leaching solution (containing 3.78 g / L Cu, 1.51 g / L Co, and 14.44 g / L Fe) and 440 g of oxygen pressure leaching residue (containing 0.15% Cu, 0.38% Co, and 13.0% Fe). Calculations showed that the copper leaching rate was 82.07% and the cobalt leaching rate was 41.94%.

[0093] Comparative Example 2

[0094] Example 1 was repeated, except that in S3, no oxygen-enriched gas was introduced, the temperature was controlled at 200°C, and the pressure at 1.3 MPa, resulting in 800 mL of oxygen pressure leaching solution (containing 3.87 g / L Cu, 1.58 g / L Co, and 11.53 g / L Fe) and 450 g of oxygen pressure leaching residue (containing 0.13% Cu, 0.36% Co, and 13.5% Fe). Calculations showed that the copper leaching rate was 84.11% and the cobalt leaching rate was 43.75%.

[0095] Comparative Examples 1 and 2 show that the copper leaching rate decreased due to the absence of oxygen, while the cobalt leaching rate decreased more significantly. The reason for the larger decrease in cobalt leaching rate is that although cobalt sulfide leaching occurs under high temperature and high acid conditions, cobalt is difficult to leach without oxygen, resulting in a higher concentration of Fe in the solution. 3+ The hydrolysis reaction is also weakened due to the lack of oxygen, which increases the iron content in the solution, increases the burden on subsequent iron removal, and may lead to a greater loss of copper and cobalt.

[0096] Comparative Example 3

[0097] Repeat Example 1, except that the mixing step in S2 is omitted and the slurry and concentrated sulfuric acid are directly fed into the oxygen pressure leaching system.

[0098] As a result, the copper leaching rate was 84.20%, and the cobalt leaching rate was 69.50%.

[0099] Comparative Example 4

[0100] Example 1 is repeated, except that in S2, the slurry is mixed with 336g of 70% sulfuric acid solution to obtain a mixed slurry.

[0101] As a result, the copper leaching rate was 85.62%, and the cobalt leaching rate was 72.70%.

[0102] The comparison shows that omitting the mixing step in S2 or using a low-concentration sulfuric acid solution in S2 will lead to a decrease in the leaching rate of copper and a significant decrease in the leaching rate of cobalt, which is not conducive to the recovery of high-value elements such as cobalt.

[0103] The above embodiments should be understood as being used only to illustrate the present invention more clearly, and not to limit the scope of the present invention. After reading the present invention, any modifications of the present invention in various equivalent forms by those skilled in the art fall within the scope defined by the appended claims.

Claims

1. A method for oxygen pressure leaching of copper-cobalt slag, characterized in that, Includes the following steps: S1. After preparing the copper-cobalt slag to be treated into a slurry, grind it to obtain a slurry. S2. The slurry is mixed with concentrated sulfuric acid and reacted to obtain a mixed slurry; S3. After oxygen pressure leaching of the mixed slurry, solid-liquid separation is performed to obtain oxygen pressure leaching solution and oxygen pressure leaching residue.

2. The oxygen pressure leaching method according to claim 1, characterized in that, In S1, the proportion of minerals with a particle size of less than 40 μm in the solid phase of the slurry is ≥90%.

3. The oxygen pressure leaching method according to claim 1, characterized in that, In S1, the copper-cobalt slag to be treated is mixed with water at a mass ratio of 60-70:30-40 and then ground.

4. The oxygen pressure leaching method according to claim 1, characterized in that, In S2, the amount of concentrated sulfuric acid added is 50-70 wt% of the dry weight of the copper-cobalt slag corresponding to the slurry, preferably 55-65 wt%, more preferably 58-62 wt%; the concentration of the concentrated sulfuric acid is ≥80 wt%, preferably 85-99 wt%, more preferably 90-99 wt%.

5. The oxygen pressure leaching method according to claim 1, characterized in that, In S2, the reaction time is 0.5-1h.

6. The oxygen pressure leaching method according to claim 1, characterized in that, Between S2 and S3, there is also a step of adjusting the liquid-solid ratio of the mixed slurry to 1.5-2 mL:1 g; preferably, one or two of the following are used to adjust the liquid-solid ratio: water or magnesium-removed liquid.

7. The oxygen pressure leaching method according to claim 1, characterized in that, In S3, during oxygen pressure leaching, the temperature is controlled at 180-190℃, the total pressure is 1-1.3 MPa, and the oxygen partial pressure is 0.2-0.3 MPa; preferably, the oxygen pressure leaching time is 1.5-2 hours.

8. The oxygen pressure leaching method according to claim 1, characterized in that, The oxygen pressure leachate is mixed with calcium hydroxide and reacted. When the pH of the reaction system is 2.5-3.5, the solid and liquid are separated to obtain iron slag and iron-removed liquid. After extracting copper from the iron-removed liquid, a raffinate is obtained. The raffinate is mixed with magnesium oxide and reacted. When the pH of the reaction system is 7-8, the solid and liquid are separated to obtain crude cobalt hydroxide product and cobalt precipitation liquid.

9. The oxygen pressure leaching method according to claim 8, characterized in that, The cobalt-precipitated liquid is mixed with calcium hydroxide and reacted. When the pH of the reaction system reaches 10-11, the solid and liquid are separated to obtain calcium-magnesium slag and magnesium-removed liquid. Preferably, the magnesium-removed liquid is returned to S1 and / or S3.

10. The oxygen pressure leaching method according to any one of claims 1-9, characterized in that, In S3, an oxygen pressure leaching system is used for oxygen pressure leaching. The oxygen pressure leaching system includes a first pipeline (5), N high-pressure reactors (2), a second pipeline (6), a third pipeline (7), a fourth pipeline (19), N ninth valves (16), tenth valves (17), twelfth valves (20), a flash tank (3), and a solid-liquid separation unit (4). The flash tank (3) is connected to the solid-liquid separation unit (4). The high-pressure reactor (2) includes a tank body with a feed inlet, a discharge outlet, an air inlet, and a manhole. A stirring mechanism is provided inside the tank body. Each feed inlet is connected to the first pipeline (5), each discharge outlet is connected to the third pipeline (7), and each air inlet is connected to the second pipeline (5). 6) Above; A fourth valve (11) and a seventh valve (14) are provided between the first pipeline (5) and the feed inlet. The feed inlet, the fourth valve (11), the seventh valve (14) and the first pipeline (5) are connected in sequence. The pipeline between the fourth valve (11) and the seventh valve (14) is connected to the third pipeline (7). A third valve (10) is provided between the second pipeline (6) and the air inlet. A fifth valve (12) is provided between the third pipeline (7) and the discharge port. One port of the ninth valve (16) is connected to the pipeline between the fifth valve (12) and the third pipeline (7). The other port of the ninth valve (16) is connected to the fourth pipeline (19). An eleventh valve (18) is provided on the third pipeline (7) between the inlet and outlet of the same tank; an eighth valve (15) is provided on the third pipeline (7) between two adjacent tanks; One port of the tenth valve (17) is connected to one end of the third pipeline (7), one port of the twelfth valve (20) is connected to the other end of the third pipeline (7), and the other port of the tenth valve (17) and the other port of the twelfth valve (20) are connected in parallel to the inlet end of the flash tank (3); Where N is an integer ≥ 2; Preferably, the solid-liquid separation unit (4) is a thickener; Preferably, N is 4-6; Preferably, the autoclave is a vertical autoclave.

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