Treatment method of alkaline leaching antimony removal slag

Abstract

The invention relates to a treatment method of alkaline leaching antimony-removed slag, which comprises the following steps: carrying out size mixing on the alkaline leaching antimony-removed slag to be treated to obtain slurry with the pH value of 0-1; carrying out pressure leaching treatment on the slurry under the conditions that the temperature is 210-230 DEG C and the air partial pressure is 0.5-1.5 MPa, cooling and depressurizing, and carrying out solid-liquid separation to obtain pressure leaching residues; and after cyanidation leaching is conducted on the pressure leaching residues, solid-liquid separation is conducted, and cyanidation residues and leaching liquid rich in gold are obtained. According to the treatment method, the gold leaching rate is high and reaches 97.5% or above, and the content of Au in the cyaniding slag is only about 2 g / t.

Description

Technical Field

[0001] The invention relates to a method for treating alkali leaching antimony-removing slag, and belongs to the field of hydrometallurgy. Background Art

[0002] High-antimony ore, which contains 10-20% antimony, is typically flotated to produce antimony concentrate (40-50% antimony), with a yield of about one-quarter that of the high-antimony ore. The flotation tailings are antimony-gold concentrate (7-8% antimony), with a yield of about three-quarters that of the high-antimony ore. The gold content of antimony-gold concentrate can reach over 85g / t. The gold valuation factor depends on the gold content of the material. For example, cyanide slag containing ~8g / t Au has a valuation factor of 40-50%; roasting slag containing ~30g / t Au has an Au valuation factor of approximately 82%; and raw materials containing over 80g / t Au have a valuation factor of up to 91%. Therefore, antimony-gold concentrate has an extremely high overall recovery value.

[0003] Common methods for processing antimony-gold concentrates include pyrometallurgical smelting, slurry electrolysis, or alkaline leaching to remove antimony and then extract gold. Pyrometallurgical smelting utilizes the readily oxidizable and volatilized properties of stibnite to roast it at temperatures between 800°C and 1000°C to produce relatively pure antimony trioxide powder. This is then further reduced, smelted, and refined to produce antimony ingots. This method is not suitable for processing mixed sulfur-oxygen antimony ores or materials containing low-melting-point ores.

[0004] Chinese invention patent application CN02103704.3 discloses a method for producing antimony by electrolyzing slurry from antimony-containing sulfide minerals. This method involves electrolyzing the slurry in a hydrochloric acid-ammonium chloride medium to produce metallic antimony plates. However, the gold concentrate after electrowinning contains a large amount of chloride ions, which adversely affects subsequent gold extraction processes and equipment. Alkaline leaching for antimony removal and subsequent gold extraction also presents the challenge of low gold leaching rates.

[0005] Chinese invention patent application CN105063354A discloses a cascade recovery method for arsenic- and antimony-containing difficult-to-treat gold ores. This method involves subjecting the leached residue after alkaline leaching to antimony removal to a two-stage roasting process of "weak oxidation roasting to remove arsenic and then oxidation roasting to desulfurize" before extracting gold and silver. This method has a long process flow and high energy consumption. Summary of the Invention

[0006] In view of the shortcomings of the prior art, the present invention aims to provide a method for treating alkaline leaching antimony removal slag to improve the gold leaching rate.

[0007] In order to solve the above technical problems, the technical solutions of the present invention are as follows:

[0008] A method for treating alkaline leaching antimony removal slag comprises the following steps:

[0009] S1. Slurrying the alkali-leached antimony-removed slag to be treated to obtain a slurry with a pH value of 0-1;

[0010] S2. After pressure leaching the slurry at 210-230° C. and an air partial pressure of 0.5-1.5 MPa, the temperature and pressure are reduced, and the solid-liquid is separated to obtain pressure leaching residue;

[0011] S3. After cyaniding the pressure leaching residue, solid-liquid separation is performed to obtain cyanided residue and gold-rich leachate.

[0012] Furthermore, in S1, the alkaline leaching antimony removal slag is obtained by alkaline leaching of antimony-gold concentrate.

[0013] Furthermore, when alkaline leaching is performed to remove antimony, the antimony-gold concentrate is mixed with water in a mass-to-volume ratio of 1 g:3-5 mL to prepare a slurry, and a water-soluble metal sulfide and a strong base are added so that the initial concentration of the water-soluble metal sulfide in the reaction system is 20-50 g / L, and the amount of the strong base added is 1.2-1.5 times the molar amount of Sb in the antimony-gold concentrate; preferably, the water-soluble metal sulfide is Na2S, and preferably, the strong base is NaOH.

[0014] Furthermore, the content of Sb in the alkaline leaching antimony slag is less than 1.5 wt %, thereby avoiding the problem of antimony consuming alkali and cyanide after entering the cyanidation process and forming a coating that affects the cyanidation leaching rate.

[0015] Furthermore, in the antimony-gold concentrate, the Sb content is 5-9wt%, and the Au content is 50-100g / t.

[0016] Furthermore, in S1, the alkaline leached antimony-removed slag is slurried at a liquid-solid ratio of 2-4 mL:1 g, and concentrated sulfuric acid is added to adjust the pH to obtain a slurry with a pH of 0-1. Preferably, the concentration of the concentrated sulfuric acid is 90-98 wt%.

[0017] Furthermore, in S2, the pressure leaching temperature is 215-225°C, preferably 218-223°C; and the air partial pressure is 0.8-1.2 MPa.

[0018] Furthermore, in S2, the pressure leaching time is 1-3 hours, preferably 1.5-2.5 hours.

[0019] Furthermore, in S3, when cyanide leaching is performed, the pressurized leaching residue is mixed with water at a mass volume ratio of 1 g: 3-6 mL to prepare a slurry, the pH value is adjusted to 9-11, and a water-soluble metal cyanide is added. After leaching, the solid and liquid are separated to obtain cyanide residue and a gold-rich leachate;

[0020] During the leaching process, the pH of the reaction system is controlled to be 9-11; by adding water-soluble metal cyanide, the CN in the solution of the reaction system is controlled. - The mass concentration is 0.25-0.3%.

[0021] Furthermore, in S3, the leaching temperature is 10-50°C, and the leaching time is 24-48h.

[0022] The present invention can effectively improve the leaching rate of gold by sequentially performing acid adjustment, pressure leaching and cyanidation on the alkaline leaching antimony-removing slag. In the first step of the present invention, the alkaline leaching process is used to treat the antimony-removing slag produced by the antimony-gold concentrate to avoid a large amount of Sb2S3 from entering the cyanidation process, which leads to increased NaOH consumption. The alkaline leaching antimony-removing slag is acidified by adding concentrated sulfuric acid, which has strong oxidizing properties and a fast acid adjustment speed. The acid adjustment process is ready for pressure leaching; the pressure leaching process effectively destroys the structure of the antimony-removing slag, which can effectively improve the leaching rate of gold in the subsequent cyanidation process. Pressure leaching can also convert a small amount of Sb2S3 in the material into Sb2O3, reducing the alkali consumption required for adjusting the pH in the subsequent cyanidation process. More importantly, it destroys the structure of the antimony-removing slag, facilitating the subsequent gold and CN - Binding reaction.

[0023] In some embodiments, the main chemical reactions occurring during the alkaline leaching process are as follows:

[0024] Sb2O3+3Na2S+3H2O=Sb2S3+6NaOH;

[0025] Sb2S3+3Na2S=2Na3SbS3;

[0026] Sb2S3+4NaOH=NaSbO2+Na3SbS3+2H2O.

[0027] Antimony trioxide in the antimony gold concentrate is first converted into antimony sulfide, which is then dissolved in an alkaline solution of sodium sulfide. When there is insufficient sodium sulfide in the solution, Sb2S3 will react directly with NaOH. The generated Na3SbS3 dissolves into the solution and is separated from the alkaline leaching residue. Some unleached Sb2S3 and a very small amount of Sb2O3 remain in the antimony removal residue and enter the subsequent pressure leaching process, where the following reaction occurs:

[0028] Sb2S3+6O2+3H2O=Sb2O3+3H2SO4

[0029] Sb2O3 is amphoteric and alkaline. During the pH adjustment process of cyanide leaching, Sb2O3 in the pressure leaching residue does not react with NaOH and NaCN, so it will not lead to an increase in the consumption of alkali and cyanide in the cyanide process.

[0030] If the antimony-gold concentrate is directly subjected to cyanide leaching without pressure leaching, Sb2S3 will have a very adverse effect on cyanide leaching. During the pH adjustment process, Sb2S3 will undergo the following reactions:

[0031] Sb2S3+6OH — =SbO33— +SbS3 3— +3H2O.

[0032] Generated SbS3 3— Part of it reacts with alkali to form SbO3 3— and S 2— :

[0033] 2SbS3 3— +12OH — =2SbO3 3— +6S 2— +6H2O.

[0034] Another part SbS3 3— With CN — , O2 reacts to form CNS — :

[0035] 2SbS3 3— +6CN — +3O2=6CNS — +2SbO3 3— .

[0036] It may also directly react to:

[0037] Sb2S3+3S 2— =2SbS3 3— .

[0038] According to the above reaction, the cyanide leaching of Sb2S3 in the raw material will lead to an increase in the consumption of NaOH and NaCN, and an increase in cost. At the same time, the decomposition product of antimony sulfide SbS3 accumulated in the cyanide solution 3— 、S 2— 、SbO3 3— Producing a thin and dense film on the gold surface to hinder CN — The flow of oxygen and oxygen to the gold particles dramatically slows down the dissolution rate of gold, resulting in a significant reduction in the gold leaching rate. This is the main reason why antimony-containing gold concentrates are extremely difficult to cyanide. Generally, direct cyanide treatment is only possible when the antimony content is very low (Sb < 0.5%).

[0039] Compared with the prior art, the present invention has the following beneficial effects:

[0040] (1) The gold leaching rate of the treatment method of the present invention is high, reaching more than 97.5%, and the cyanide slag contains only about 2g / t of Au.

[0041] (2) The present invention first adjusts the acidity of the alkaline leaching antimony removal slag to neutralize the alkali remaining in the alkaline leaching process, significantly improving the gold leaching rate. If the acid adjustment process is not performed, the pressurized leaching solution will be neutral or even alkaline, resulting in a low gold leaching rate in the subsequent cyanide leaching process. The pH of the acid adjustment is controlled to 0-1, and after pressurized leaching, the pressurized leaching slag obtained contains less acidic substances. In some embodiments, the NaOH consumption required for pH adjustment during the cyanide process is only 5-30 kg / t of ore, which is low in cost.

[0042] (3) The present invention performs pressure leaching on the antimony-removed slag after acid adjustment, destroying the structure of the antimony-removed slag under high temperature and high pressure conditions, thereby providing favorable conditions for subsequent cyanide leaching.

[0043] (4) The pressurized leachate obtained after the pressure leaching of the present invention contains 0-25g / L H2SO4. Controlling the acid concentration in the leachate can enhance the degree of destruction of the antimony removal slag and provide favorable conditions for subsequent cyanide leaching. When low-concentration sulfuric acid is sent to the sewage treatment station for neutralization, the amount of lime consumed is small. The neutralization process can remove impurity ions such as arsenic and iron from the solution, and the neutralized liquid is returned to the antimony removal slag for slurry preparation.

[0044] (5) The treatment process of the present invention is relatively short. After acid adjustment, pressure leaching is directly carried out, which is simple and efficient. BRIEF DESCRIPTION OF THE DRAWINGS

[0045] Figure 1 This is a processing flow chart of antimony-gold concentrate of the present invention. DETAILED DESCRIPTION

[0046] The present invention will be described in detail below with reference to the embodiments. It should be noted that, unless otherwise specified, the embodiments and features of the embodiments of the present invention may be combined with each other. Unless otherwise specified, the relevant percentages are percentages by mass.

[0047] Example 1

[0048] 400g of antimony-gold concentrate alkali leaching antimony removal slag (containing Au 88.98g / t, Ag 0.86g / t, Fe 6.60%, S5.54%, As 2.5%) was taken, 1.2L of neutralizing solution was added to slurry, the liquid-solid ratio was controlled to 3mL:1g, 108g of concentrated sulfuric acid with a mass concentration of 98wt% was added under normal pressure, and the pH value of the pulp was adjusted to 0.5; the acid-adjusted pulp was added to an autoclave for pressure leaching, the temperature was controlled to 215°C, the time was 2.0h, the air partial pressure was 1.0Mpa, the temperature and pressure were lowered after the reaction, and the liquid-solid separation was performed to obtain 382.3g of pressure leaching slag (containing Au 93.10g / t, Ag 0.90g / t, Fe 6.47%, S10.52%, As 2.47%), 1.2L of pressure leaching solution (containing acid 18g / L, Fe 1.40g / L, As 0.46g / L).

[0049] The pressure leaching residue was sent to cyanide leaching. 300g of the pressure leaching residue was added with 1.5L of water to prepare the slurry. After adding NaOH to adjust the pH to 10, NaCN was added to leach the gold. After stirring and leaching at room temperature for 36h, liquid-solid separation was performed to obtain 276.2g of cyanide residue (containing 2.2g / t Au) and a gold-rich leachate with a gold leaching rate of 97.82%.

[0050] During the leaching process, the pH of the solution was controlled to 10 by adding NaOH; the CN content in the solution was controlled by adding NaCN. - The mass concentration of NaOH is 25 parts per ten thousand; the cumulative amount of NaOH added is 18.5 kg / t pressure leaching residue, and the cumulative amount of NaCN added is 13.32 kg / t pressure leaching residue.

[0051] The pressurized leaching liquid is sent to the sewage treatment station for neutralization to remove impurities such as iron, and a neutralized liquid with a pH of 6.5 is obtained, which is returned to the alkaline leaching antimony removal slag for slurry preparation.

[0052] Example 2

[0053] 400g of antimony-gold concentrate alkali leaching antimony removal slag (containing Au 88.98g / t, Ag 0.86g / t, Fe 6.60%, S5.54%, As 2.5%) was taken, 1.2L of neutralizing solution was added to slurry, the liquid-solid ratio was controlled to 3mL:1g, 110g of concentrated sulfuric acid with a mass concentration of 93wt% was added under normal pressure, the pH of the pulp was adjusted to 0.8, the acid-adjusted pulp was added to an autoclave for pressure leaching, the temperature was controlled to 225°C, the time was 2.5h, the air partial pressure was 1.2Mpa, and after the reaction, the temperature was lowered and the pressure was lowered, and liquid-solid separation was performed to obtain 378g of pressure leaching slag (containing Au 94.16g / t, Ag 0.91g / t, Fe 6.70%, S10.80%, As 2.48%), 1.2L of pressure leaching solution (containing acid 20g / L, Fe0.90g / L, As 0.54g / L).

[0054] The pressure leaching residue was sent to cyanide leaching. 300g of the pressure leaching residue was added with 1.5L of water to prepare the slurry. After adding NaOH to adjust the pH to 10.5, NaCN was added to leach the gold. After stirring and leaching at room temperature for 36h, the liquid-solid separation was carried out to obtain 275g of cyanide residue (containing 1.8g / t Au) and a gold-rich leachate with a gold leaching rate of 98.25%.

[0055] During the leaching process, the pH of the solution was controlled at 10.5 by adding NaOH; the CN content in the solution was controlled by adding NaCN. - The mass concentration of NaOH is 30 parts per ten thousand; the cumulative amount of NaOH added is 19.4 kg / t pressure leaching residue, and the cumulative amount of NaCN added is 14.5 kg / t pressure leaching residue.

[0056] The pressurized leaching liquid is sent to the sewage treatment station for neutralization to remove impurities such as iron, and a neutralized liquid with a pH of 7 is obtained, which is returned to the alkaline leaching antimony removal slag for slurry preparation.

[0057] Comparative Example 1

[0058] Example 1 was repeated, except that after the alkaline leaching of the antimony-removed slag was slurried, concentrated sulfuric acid was not added to adjust the pH, and pressure leaching and cyanide leaching were performed directly. The results showed that the pH of the pressure leaching solution was 8, the pressure leaching residue mass was 370.9 g (containing 95.96 g / t Au), and cyanide leaching of 300 g of the pressure leaching residue yielded 286.17 g of cyanide residue (containing 49.42 g / t Au) and 1.5 L of gold-rich leachate, with a gold leaching yield of only 50.87%. The added amounts of NaOH and NaCN were 6.6 kg / t and 29.1 kg / t of pressure leaching residue, respectively.

[0059] It can be seen from this that the gold leaching rate is low when the alkaline leaching antimony removal slag is directly subjected to pressure leaching-cyanidation without acid adjustment.

[0060] Comparative Example 2

[0061] Example 1 was repeated, except that a roasting process was used instead of the acid adjustment-pressure leaching process. After the alkaline-leached antimony-removed slag was roasted at 615°C for 3.5 hours, 300 g of the roasted ore was subjected to cyanide leaching. 262.69 g of cyanide slag (containing 24.30 g / t Au) and 1.5 L of gold-rich leachate were obtained, with a gold leaching yield of 76.09%. During the cyanide leaching process, the cumulative addition of NaOH was 10.0 kg / t of pressure leaching slag, and the cumulative addition of NaCN was 20.5 kg / t of pressure leaching slag.

[0062] It can be seen from this that the gold leaching rate is low when the "roasting" process is used instead of "acid adjustment-pressure leaching".

[0063] Comparative Example 3

[0064] Example 1 was repeated except that a "roasting-pressure leaching" process was used instead of the "acid adjustment-pressure leaching" process. Specifically, the alkaline leached antimony-removed slag was roasted at 615° C. for 3.5 h, and then 300 g of the roasted ore was pressure leached to obtain 282.6 g of pressure leaching slag (containing 100.32 g / t of Au). 200 g of the pressure leaching slag was subjected to cyanide leaching to obtain 188.90 g of cyanide slag (containing 8.40 g / t of Au), with a gold leaching rate of 92.09%.

[0065] It can be seen that using the "roasting-pressure leaching" process instead of the "acid adjustment-pressure leaching" process is more complex and the gold leaching rate is not high. The above examples should be understood as merely illustrating the present invention and are not intended to limit the scope of the present invention. After reading this disclosure, various equivalent modifications of the present invention by those skilled in the art fall within the scope of the appended claims.

Claims

1. A method for treating alkali leaching antimony removal slag, characterized in that: The steps include: S1. Slurrying the alkali-leached antimony-removed slag to be treated to obtain a slurry with a pH value of 0-1; S2. After pressure leaching the slurry at 210-230° C. and an air partial pressure of 0.5-1.5 MPa, the temperature and pressure are reduced, and the solid-liquid is separated to obtain pressure leaching residue; S3. After cyaniding the pressure leaching residue, solid-liquid separation is performed to obtain cyanided residue and gold-rich leachate.

2. The processing method according to claim 1, characterized in that In S1, the alkaline leaching antimony removal slag is obtained by alkaline leaching of antimony-gold concentrate.

3. The processing method according to claim 2, characterized in that When alkaline leaching is performed to remove antimony, the antimony-gold concentrate is mixed with water in a mass-to-volume ratio of 1 g:3-5 mL to prepare a slurry, and a water-soluble metal sulfide and a strong base are added so that the initial concentration of the water-soluble metal sulfide in the reaction system is 20-50 g / L, and the amount of the strong base added is 1.2-1.5 times the molar amount of Sb in the antimony-gold concentrate; preferably, the water-soluble metal sulfide is Na2S, and preferably, the strong base is NaOH.

4. The processing method according to claim 1, characterized in that The content of Sb in the alkaline leaching antimony removal slag is less than 1.5 wt%.

5. The processing method according to claim 1 or 2, characterized in that: In the antimony-gold concentrate, the content of Sb is 5-9wt%, and the content of Au is 50-100g / t.

6. The processing method according to claim 1, characterized in that In S1, the alkaline leaching antimony removal slag is slurried at a liquid-solid ratio of 2-4 mL:1 g, and concentrated sulfuric acid is added to adjust the pH to obtain a slurry with a pH value of 0-1; preferably, the concentration of concentrated sulfuric acid is 90-98 wt%.

7. The processing method according to claim 1, characterized in that In S2, the pressure leaching temperature is 215-225°C, preferably 218-223°C; and the air partial pressure is 0.8-1.2 MPa.

8. The processing method according to claim 1, characterized in that In S2, the pressure leaching time is 1-3 hours, preferably 1.5-2.5 hours.

9. The processing method according to claim 1, characterized in that: In S3, when cyanide leaching is performed, the pressurized leaching residue is mixed with water at a mass volume ratio of 1 g: 3-6 mL to prepare a slurry, the pH value is adjusted to 9-11, and a water-soluble metal cyanide is added. After leaching, the solid and liquid are separated to obtain cyanide residue and a gold-rich leachate; During the leaching process, the pH of the reaction system is controlled to be 9-11; by adding water-soluble metal cyanide, the CN in the solution of the reaction system is controlled. - The mass concentration is 0.25-0.3%.

10. The processing method according to claim 9, characterized in that: In S3, the leaching temperature is 10-50°C and the leaching time is 24-48h.

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