A new method for cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings

Abstract

The present invention discloses a new method for the cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings. It comprises the following steps: step 1) using a hydrocyclone to preferentially separate the coarse-grained pyrite to obtain a coarse-grained sulfur concentrate; step 2) flotating the overflow of the hydrocyclone under alkaline conditions to further separate the pyrite with strong floatability in the alkaline environment to obtain a highly floatable sulfur concentrate; step 3) flotation by sequentially adding a collector, a regulator, acid mine wastewater and a frother to the slurry to obtain a weakly floatable sulfur concentrate. The present invention makes full use of the differentiated characteristics of the particle size and floatability of pyrite itself, adopts the methods of gravity separation and flotation to achieve its cascade beneficiation and enrichment, and the flotation process avoids the high production cost and operational safety risks brought about by the use of concentrated sulfuric acid as an activator; at the same time, it solves the problems of high cost alkali consumption in the conventional treatment of acid mine wastewater and unstable flotation when acid mine wastewater is used as a flotation activator for pyrite.

Description

Technical Field

[0001] The present invention belongs to the technical field of mineral processing, and relates to a flotation method for copper-sulfur separation tailings, and specifically to a method for activating pyrite suppressed by lime in the ore pulp without using concentrated sulfuric acid, thereby obtaining flotation sulfur concentrate. Background Art

[0002] Copper, a vital metal mineral resource, is widely used in numerous fields due to its excellent properties. my country's copper reserves rank seventh in the world, and with the advancement of industrialization, demand for copper has increased significantly. Studies have shown that by 2050, demand for copper will increase by 275%-350%. Most of my country's copper resources are copper sulfide ores, often coexisting with pyrite. Due to the similar chemical properties of these two minerals, and the fact that excessive pyrite content can reduce the grade of copper concentrate, suppressing pyrite during flotation is crucial for improving copper concentrate quality and achieving efficient sulfur utilization.

[0003] Currently, to achieve the selective separation of metal sulfide ores such as copper, lead, and zinc, most sulfide ore concentrators use a high-alkali process to suppress pyrite. In the subsequent flotation of the copper-sulfur separation tailings, large amounts of concentrated sulfuric acid are used as a pyrite activator. However, concentrated sulfuric acid is prone to corrosive equipment and poses a high safety risk. Therefore, research on inexpensive and safe high-alkali activators for pyrite suppression is extremely important. Furthermore, sulfide mining produces a large amount of acidic mine wastewater with a low pH. Using this activator instead of concentrated sulfuric acid as a pyrite activator could eliminate the safety issues associated with concentrated sulfuric acid.

[0004] Acid mine drainage is rich in Fe 3+ / Fe 2+ Mg 2+ 、Mn 2+ , Ca 2+ 、Al 3+ 、Zn 2+ Metal ions such as sulfuric acid and sulphuric acid have a significant effect on the flotation of pyrite. Therefore, when using acidic wastewater to float pyrite, it is important to find an effective pulp conditioner to eliminate the influence of metal ions so that acidic wastewater can replace concentrated sulfuric acid to float pyrite. Summary of the Invention

[0005] To address the shortcomings of the existing technology, the present invention aims to activate pyrite without using concentrated sulfuric acid, thereby obtaining pyrite concentrate while reducing the discharge of acidic wastewater from mineral processing. Currently, the flotation recovery of pyrite from copper-sulfur separation tailings is primarily achieved through an activated flotation process using concentrated sulfuric acid. This process carries safety risks and high production costs, and does not consider the use of a cascaded mineral processing and enrichment process that takes advantage of the differences in the physical and chemical properties of pyrite in copper-sulfur separation tailings.

[0006] Another technical problem to be solved by the present invention is to eliminate the influence of impurity ions in mine acid wastewater on flotation by optimizing the flotation reagent scheme.

[0007] In order to solve the above-mentioned drawbacks of the prior art, the present invention provides a new method for the cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings, which comprises the following steps:

[0008] Step 1) Hydrocyclone sulfur separation: taking the copper-sulfur separation tailings as the processing object, first use a hydrocyclone to preferentially separate the coarse-grained pyrite to obtain a coarse-grained sulfur concentrate;

[0009] Step 2) alkaline sulfur separation: The overflow of the hydrocyclone is subjected to preliminary flotation under alkaline conditions to separate the highly floatable pyrite to obtain a highly floatable sulfur concentrate; the pyrite collector used in the alkaline sulfur separation is butyl xanthate, and the dosage is 90-320 g / t; the pH value of the slurry is between 8.5 and 10.5;

[0010] Step 3) Sulfur separation with acid mine wastewater: Pyrite collector, regulator, acid mine wastewater and frother are added to the pulp in sequence for flotation to obtain weakly floatable sulfur concentrate. In step 3) activating flotation of pyrite with acid mine wastewater, the amount of acid mine wastewater used is 0.18m 3 / t to 1.3m 3 / t; flotation concentration is 30% to 40%; pulp adjuster is a combination of chitosan and sodium lignin sulfonate, the mass ratio of the two is 1:4 to 1:9, and the total dosage is 50 to 200 g / t; collector is at least one of butyl xanthate, amyl xanthate, and isopentyl xanthate, and the dosage is 120 to 350 g / t; flotation frother is 2 # One of oil or methyl isobutyl carbinol, the dosage is 15-55g / t.

[0011] Coarse-grained sulfur concentrate and strong floatability sulfur concentrate can be used alone as qualified sulfur concentrate; weak floatability sulfur concentrate needs to be mixed with the above two types of sulfur concentrate to be qualified sulfur concentrate; or the three can be mixed to be qualified sulfur concentrate.

[0012] In a preferred embodiment, the particle size of the copper-sulfur separation tailings is not less than 15% by weight of +0.074 mm.

[0013] In a preferred embodiment, the sulfur grade in the copper-sulfur separation tailings is not less than 15%.

[0014] In a preferred embodiment, the cyclone sulfur separation separates the coarser pyrite according to the different particle sizes of pyrite and sludge in the copper-sulfur separation tailings, while the finer sludge and pyrite are discharged through overflow as subsequent flotation slurry.

[0015] In a preferred embodiment, the flotation separation method preferentially floats out pyrite with strong floatability according to the different floatability of pyrite in the copper-sulfur separation tailings, while pyrite with poor floatability is activated by flotation using a combination of acid mine wastewater and a pulp conditioner.

[0016] In a preferred embodiment, in step 1), the diameter of the sand trap of the cyclone is 9.5 to 15 mm, preferably 10 to 14 mm, and more preferably 12 to 14 mm.

[0017] In a preferred embodiment, in step 1), the overflow port of the cyclone has a diameter of 20 to 26 mm, preferably 20 to 24 mm, and more preferably 22 to 24 mm.

[0018] In a preferred embodiment, in step 1), the feed pressure of the cyclone is 0.06-0.1 MPa, preferably 0.06-0.08 MPa.

[0019] In a preferred embodiment, in step 2), the pyrite collector is one of butyl xanthate, amyl xanthate, and isopentyl xanthate. In a preferred embodiment, the amount of the collector is 100-300 g / t, preferably 100-250 g / t, and more preferably 100-200 g / t.

[0020] In a preferred embodiment, in step 2), the flotation time is 3 to 5 minutes.

[0021] In a preferred embodiment, in step 2), the flotation pH value is between 8.5 and 10.5.

[0022] In a preferred embodiment, the slurry after flotation in step 2) is adjusted to a concentration of 30% to 60%, preferably 30% to 50%, and more preferably 35% to 45%. A pyrite collector, a conditioning agent, acid mine drainage water, and a frother are then added sequentially for flotation. In engineering applications, the thickening process is performed in a thickener; the result is a thickener underflow concentration of 40% to 70%, preferably 40% to 60%, and more preferably 40% to 50%.

[0023] In a preferred embodiment, in step 3), the flocculant is one of sodium polyacrylate and polyacrylamide, used in combination with polyaluminum chloride in a ratio of 5:1 to 10:1, with a total amount of 10 to 60 g / t, preferably 10 to 40 g / t, and more preferably 20 to 30 g / t.

[0024] In a preferred embodiment, in step 3), the pyrite collector is at least one of butyl xanthate, amyl xanthate, and isopentyl xanthate, and the amount used is 120 to 350 g / t, preferably 150 to 250 g / t.

[0025] In a preferred embodiment, in step 3), the flotation concentration is 30% to 50%, preferably 30% to 45%, and more preferably 30% to 40%. In engineering applications, the thickening process is carried out in a thickener; the thickening result is: the thickener underflow concentration is 40% to 70%, preferably 40% to 60%, and more preferably 40% to 50%.

[0026] In a preferred embodiment, in step 3), the pulp conditioner is a combination of chitosan and sodium lignin sulfonate, wherein the molecular weight of chitosan is 250,000, the weight ratio of the chitosan and sodium lignin sulfonate is 1:5 to 1:8, and the total amount is 50 to 200 g / t, preferably 80 to 150 g / t, and more preferably 100 to 120 g / t.

[0027] In the preferred embodiment, in step 3), the amount of acid mine wastewater used is 0.2m 3 / t to 1.2m 3 / t, preferably 0.2m 3 / t~0.8m 3 / t, more preferably 0.2m 3 / t~0.4m 3 / t.

[0028] In a preferred embodiment, in step 3), after adding the acid mine wastewater, the pH of the pulp is between 4.0 and 6.0, preferably between 4.0 and 5.0. That is, the flotation pH of the mine acid sulfur separation is between 4.0 and 6.0, preferably between 4.0 and 5.0. In a preferred embodiment, in step 3), the flotation frother is 2 # At least one of oil (i.e., No. 2 oil, composite higher alcohol) and methyl isobutyl carbinol is used in an amount of 5 to 50 g / t, preferably 10 to 40 g / t, and more preferably 10 to 20 g / t.

[0029] In a preferred embodiment, in step 3), the flotation time is 3 to 5 minutes.

[0030] In a preferred embodiment, in step 4), the final sulfur concentrate is the final product.

[0031] The invention discloses a new method for cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings. The sulfur grade in the obtained sulfur concentrate is 39.57%, and the sulfur recovery rate is 97.9%.

[0032] The present invention uses a step-by-step separation method to recover sulfur from copper-sulfur separation tailings. First, a cyclone is used to separate coarse pyrite. Then, pyrite with good floatability in the cyclone overflow is preferentially floated under alkaline conditions. Finally, pyrite with poor floatability is floated in an acidic mine wastewater environment. This ensures the sulfur grade and recovery rate of the final concentrate.

[0033] The main technical principles and innovations of the step-by-step separation method of the present invention are:

[0034] (1) According to the differentiated characteristics of pyrite's particle size and floatability, gravity separation and flotation methods are used to achieve its cascaded mineral enrichment. The sludge in the copper-sulfur separation tailings is relatively fine, and there are coarse and fine pyrite in it. According to this particle size characteristic, the use of a cyclone can directly separate the coarse pyrite concentrate with qualified grade. The fine pyrite in the overflow after the cyclone sulfur separation has good floatability and can be directly floated under alkaline conditions of 8.5-10.5. Therefore, according to the difference in the floatability of the pyrite itself, the step-by-step flotation method is used to preferentially float out the pyrite with good floatability under alkaline conditions, which is beneficial to the recovery of the pyrite. The pyrite in the alkaline sulfur dressing tailings has poor floatability. It is inhibited by lime and difficult to float directly by adding a collector. In the current process, concentrated sulfuric acid is used to activate the pyrite inhibited by lime. In the present invention, mine acidic wastewater is used instead of concentrated sulfuric acid to eliminate the inhibitory effect of hydroxyl calcium on the surface of pyrite, thereby floating out the pyrite with poor floatability in the alkaline sulfur dressing tailings.

[0035] (2) Acidic mine drainage is rich in Fe 3+ / Fe 2+ Mg 2+ 、Mn 2+ , Ca 2+ 、Al 3+ 、Zn 2+ Metal ions such as iron ions have a significant impact on the flotation of pyrite. Therefore, a pulp conditioner is used to eliminate the adverse effects of metal ions. The conditioner and acid mine drainage are used as a combination of reagents to enhance the sulfur separation effect of acid mine drainage.

[0036] (3) The collector is added before the acid mine wastewater, which can prevent the metal ions in the acid mine wastewater from consuming it to the greatest extent; combined with the use of slurry conditioner, it can avoid the influence of calcium sulfate and ferric hydroxide precipitation on flotation, so that pyrite with poor floatability can be fully collected.

[0037] The key to the present invention is to adopt a step-by-step enrichment method. First, coarse-grained pyrite is separated by a cyclone. Then, the pyrite with strong floatability that is not inhibited by lime in the cyclone overflow is preferentially floated out, avoiding the adverse effects of metal ions in the added acidic mine wastewater. Finally, a small amount of pyrite with poor floatability that is completely inhibited by lime is activated by flotation. The remaining pyrite is obtained by flotation separation by adding acidic mine wastewater and a slurry conditioner as a combined reagent. BRIEF DESCRIPTION OF THE DRAWINGS

[0038] Figure 1 Process flow chart DETAILED DESCRIPTION

[0039] The embodiments of the present invention are described in a clear and complete manner. Obviously, the embodiments described represent only a portion of the present invention, not all of it. All other embodiments that can be derived by a person of ordinary skill in the art based on the embodiments of the present invention without inventive effort are protected by the present invention.

[0040] Example 1

[0041] A copper sulfide mine in Jiangxi Province contains 9.6% sulfur and 0.48% copper. The main valuable metal is copper, and it is also accompanied by a considerable amount of gold and silver. The copper sulfide ore is mainly composed of chalcopyrite, chalcopyrite, and arsenic copper sulfide, and contains relatively high levels of pyrite and arsenopyrite; the gangue minerals are mainly quartz, feldspar, and sericite. Gold and silver are present in the sulfide ore in the form of associated minerals. The tailings after the separation of copper and sulfur are the feed for the sulfur selection stage, that is, the feed for flotation in this embodiment. Its sulfur grade is 20.74%, the copper grade is 0.05%, and the -200 mesh content is 82.7%. The acidic mine wastewater is taken from the sedimentation tank of the mining site for subsequent pyrite activation flotation. The acidic mine wastewater is rich in Fe 2+ 、Fe 3+ Mg 2+ 、Mn 2+ , Ca 2+ 、Al 3+ 、Zn 2+ The wastewater contained a high concentration of metal ions, including Fe and Mg, reaching 1989 mg / L and 2081 mg / L, respectively. Furthermore, the sulfur content reached a high of 5751 mg / L. The pH values ​​of five different wastewater batches were 2.87, 2.66, 3.21, 2.79, and 2.72, with an average of 2.85. The water sample had a high hardness, and after several days of storage, noticeable precipitation adhered to the inner walls of the container.

[0042] The present invention provides a new method for the cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings. Figure 1As shown, the process includes the following steps: directly separating the copper-sulfur separation tailings using a cyclone, with a cyclone grit port diameter of 10 mm, an overflow port diameter of 24 mm, and a feed pressure of 0.06 MPa, to obtain a coarse-grained pyrite concentrate K1. The cyclone overflow is then concentrated using a thickener to obtain an underflow slurry with a concentration of 40%. 100 g / t of butyl xanthate is then directly added to the thickener underflow, and pyrite with good floatability is selected through roughing I. The flotation pH is 10.3, and the flotation time is 5 minutes, to obtain a roughing I sulfur concentrate and roughing I tailings. The roughing I sulfur concentrate is then subjected to a single concentration operation (i.e., concentration I) to obtain a sulfur concentrate K2. No reagents are added during the concentration I operation, the flotation time is 5 minutes, and the flotation pH is 9.42. Then, the selected I tailings and the rougher I tailings were mixed and concentrated using a thickener to obtain a 40% slurry. Then, 120 g / t of butyl xanthate, 100 g / t of a combination of chitosan (molecular weight of chitosan is 250,000) and sodium lignin sulfonate in a mass ratio of 1:8, and 0.2 m3 of acid mine wastewater were added in sequence. 3 / t, 20g / t methyl isobutyl carbinol, followed by flotation for 5 minutes (i.e., roughing II). The pH of the slurry during flotation was 5.52, obtaining roughing II sulfur concentrate and roughing II tailings. The roughing II sulfur concentrate was subjected to a single concentration, i.e., concentration II. During concentration II, the slurry concentration was 35%. In this embodiment, water was added to the flotation concentration of concentration II to 35%. No reagents were added during the concentration II operation, the pH was 4.77, and the flotation time was 5 minutes. Concentrate K3 and concentration II tailings were obtained from concentration II. The roughing II tailings were subjected to a single scavenging operation to obtain the final tailings X. 20g / t butyl xanthate was added during the scavenging operation for 3 minutes. The concentration II tailings and scavenged concentrate were returned to the roughing II operation to form a closed circuit. Concentrates K1, K2, and K3 were used as the final concentrate product K. The recovery rate and grade of the sulfur concentration section were 97.9% and 39.57%, respectively.

[0043] Comparative Example 1

[0044] Other conditions were the same as those in Example 1, except that commercially available chemically pure concentrated sulfuric acid was used for activation and sulfur separation instead of acid mine wastewater. The final mineral processing indicators obtained through flotation testing were: a sulfur recovery rate of 97.65% and a sulfur grade of 39.62%. As can be seen from Example 1 and Comparative Example 1, the present invention can use acid mine wastewater instead of commercially available chemically pure concentrated sulfuric acid. Furthermore, the sulfur grade of the tailings after copper-sulfur separation in Example 1 is less than or equal to 0.79%, which further improves the sulfur recovery rate.

[0045] Comparative Example 2

[0046] The other conditions are the same as those in Example 1, except that: the highly floatable pyrite is not preferentially floated under alkaline conditions. That is, the copper-sulfur separation tailings are separated by a cyclone, the diameter of the cyclone sand settling port is 10 mm, the diameter of the overflow port is 24 mm, and the feed pressure is 0.06 MPa to obtain a coarse-grained pyrite concentrate K1. The overflow of the cyclone is then concentrated using a thickener to obtain a slurry with a concentration of 40%, and 120 g / t of butyl xanthate, 100 g / t of a combination of chitosan (molecular weight of chitosan is 250,000) and sodium lignin sulfonate in a ratio of 1:8, and 0.2 m3 of acid mine wastewater are added in sequence. 3 / t, 20g / t methyl isobutyl carbinol, followed by flotation for 5 minutes (i.e., roughing I). The slurry pH during flotation was 5.02, resulting in a rougher sulfur concentrate and rougher tailings. The rougher sulfur concentrate was then subjected to a single beneficiation operation to obtain sulfur concentrate K2 and beneficiated tailings. During beneficiation, the slurry concentration was 35%. In this embodiment, water was added to the beneficiation concentration to 35%. No reagents were added during the beneficiation operation. The flotation pH was 5.10, and the flotation time was 5 minutes. The rougher tailings were subjected to a single scavenging operation to obtain final tailings X. 20g / t butyl xanthate was added during the scavenging operation for 3 minutes. The beneficiated tailings and scavenged concentrate were returned to the roughing operation to form a closed circuit. Concentrates K1 and K2 were used as the final concentrate product K. The final beneficiation indicators obtained from flotation tests were: sulfur recovery rate of 96.76% and sulfur grade of 36.68%.

[0047] Comparative Example 3

[0048] The same process as in Example 1 was adopted, but no pulp conditioner was added. The rest of the reagent dosage was the same and flotation tests were conducted. The final mineral processing indicators obtained from the flotation tests were: sulfur recovery rate of 97.53% and sulfur grade of 36.92%.

[0049] Comparative Example 4

[0050] The same process as in Example 1 was used, except that acidic mine wastewater was not used as a flotation activator for the cyclone overflow. The copper-sulfur separation tailings were directly separated using a cyclone with a grit port diameter of 10 mm, an overflow port diameter of 24 mm, and a feed pressure of 0.06 MPa, yielding a coarse pyrite concentrate K1. The cyclone overflow was then concentrated using a thickener to obtain an underflow slurry with a concentration of 40%. 100 g / t of butyl xanthate was then added directly to the thickener underflow. The pyrite with good floatability was then selected through roughing I. The flotation pH was 10.3, and the flotation time was 5 minutes. This yielded a roughing I sulfur concentrate and roughing I tailings. The roughing I sulfur concentrate was then subjected to a single concentrating operation (i.e., concentrating I) to yield a roughing sulfur concentrate K2. (The concentrating conditions for the roughing I sulfur concentrate should be clearly stated, including the pH.) No reagents were added to the concentrating I operation, the flotation time was 5 minutes, and the flotation pH was 9.4. Then, the selected I tailings and the roughing I tailings were mixed and concentrated using a thickener to obtain a slurry with a concentration of 35%. 120 g / t of butyl xanthate, 100 g / t of a combination agent of chitosan (chitosan molecular weight is 250,000) and sodium lignin sulfonate in a ratio of 1:8, and 40 g / t of methyl isobutyl carbinol were added in sequence, followed by flotation for 5 minutes (i.e., roughing II). The pH of the slurry during flotation was 8.02, and roughing II sulfur rough concentrate and roughing II tailings were obtained. The roughing II sulfur rough concentrate was subjected to one concentration, i.e., selection II. During selection II, the slurry concentration was 35%. In this embodiment, water was added to the flotation concentration of selection II to 35%. No reagent was added to the selection II operation, the pH of selection II was 7.63, and the time was 5 minutes. Selection II obtained product concentrate K3 and selection II tailings. The rougher II tailings were subjected to a scavenging operation to obtain the final tailings X. 20 g / t butyl xanthate was added during the scavenging process for 3 minutes. The concentrated II tailings and scavenged concentrate were returned to the rougher II operation to form a closed loop. The concentrates K1, K2 and K3 were used as the final concentrate product K, with a recovery rate and grade of 93.49% and 36.72% respectively.

[0051] Comparative Example 5

[0052] The same process flow as in Example 1 was used, but the order of dosing was changed. Acidic mine wastewater was added as a regulator before the butyl xanthate and chitosan / sodium lignin sulfonate combination. The copper-sulfur separation tailings were directly separated using a cyclone with a grit inlet diameter of 10 mm, an overflow inlet diameter of 24 mm, and a feed pressure of 0.06 MPa to obtain a coarse-grained pyrite concentrate K1. The cyclone overflow was then concentrated using a thickener to obtain an underflow slurry with a concentration of 40%. 100 g / t of butyl xanthate was then added directly to the thickener underflow. The pyrite with good floatability was then selected through roughing I. The flotation pH was 10.3, and the flotation time was 5 min. Roughing I sulfur concentrate and roughing I tailings were obtained. The roughing I sulfur concentrate was then subjected to a single concentrating operation (i.e., concentrating I) to obtain a sulfur concentrate K2. No reagents were added during concentrating I, and the flotation time was 5 min. The flotation pH was 9.4. Then, the selected I tailings and the roughing I tailings were mixed and concentrated using a thickener to obtain a 35% slurry, and 0.2m of acid mine wastewater was added in sequence. 3 / t, butyl xanthate 120g / t, a combination agent of chitosan (chitosan molecular weight is 250,000) and sodium lignin sulfonate in a ratio of 1:8 100g / t, methyl isobutyl carbinol 20g / t, followed by flotation for 5min (i.e., roughing II), the pH of the slurry during flotation was 5.50, and roughing II sulfur rough concentrate and roughing II tailings were obtained. The roughing II sulfur rough concentrate was subjected to one concentration, i.e., concentration II. During concentration II, the pulp concentration was 35%. In this embodiment, water was added to the flotation concentration of concentration II to 35%. No reagent was added during the concentration II operation, the pH of concentration II was 4.81, the time was 5min, and concentration II obtained product concentrate K3 and concentration II tailings. The rougher II tailings were subjected to a scavenging operation to obtain the final tailings X. 20 g / t butyl xanthate was added during the scavenging process for 3 minutes. The concentrated II tailings and scavenged concentrate were returned to the rougher II operation to form a closed loop. The concentrates K1, K2 and K3 were used as the final concentrate product K, with a recovery rate and grade of 86.91% and 31.90% respectively.

[0053] Comparative Example 6

[0054] The same process as in Example 1 was adopted, with the addition of an excess of acid mine wastewater and the remaining reagent dosages being the same for flotation tests. The copper-sulfur separation tailings were directly separated using a cyclone, with a cyclone sand settling port diameter of 10 mm, an overflow port diameter of 24 mm, and a feed pressure of 0.06 MPa, to obtain a coarse-grained pyrite concentrate K1. The cyclone overflow was then concentrated using a thickener to obtain an underflow pulp with a concentration of 40%. 100 g / t of butyl xanthate was then directly added to the underflow of the thickener, the pulp pH was adjusted to 10.3, and flotation was performed for 5 minutes to obtain a roughing I concentrate and a roughing I tailing. The obtained roughing I concentrate was then subjected to fine concentration, i.e., fine concentration I, with a pH of 9.42. No reagents were added to obtain a sulfur rough concentrate K2 and a fine concentration I tailing. The selected I tailings and the rougher I tailings are mixed and concentrated using a thickener to obtain a 40% slurry. Then, a 1-roughing 1-fine 1-scavenging process (i.e., rougher II, selected II, scavenging I) is adopted to enrich the pyrite with poor floatability. During rougher II, 100g / t of a combination of chitosan (molecular weight of chitosan is 250,000) and sodium lignin sulfonate, 100g / t of butyl xanthate, and 1.6m 3 / t of acid mine wastewater, 40g / t of methyl isobutyl carbinol, and a pH of 2.96 were used. After flotation for 5 minutes, Rougher II concentrate and Rougher II tailings were obtained. The Rougher II concentrate was then subjected to beneficiation, namely Concentration II. The pulp concentration of Concentration II was 35%, the pH was 3.46, and flotation for 5 minutes produced sulfur concentrate K2 and Concentration II tailings. The Rougher II tailings were then subjected to scavenging, namely Scavenging I, with 20g / t of butyl xanthate used in Scavenging I and a pH of 3.07. After flotation for 3 minutes, Scavenging I concentrate and tailings X were obtained. The beneficiated tailings and scavenging concentrate were returned to the roughing operation, forming a closed-loop system. Concentrates K1, K2, and K3 were used as the final concentrate product K, with a recovery rate and grade of 73.36% and 29.63%, respectively.

[0055] Comparative Example 7

[0056] The same process as in Example 1 was adopted, except that the ratio of chitosan to sodium lignin sulfonate was changed to 5:1 and the remaining reagent dosages were the same. The flotation test was conducted, and the final mineral processing indicators obtained by the flotation test were: sulfur recovery rate of 89.93% and sulfur grade of 37.43%.

[0057] Example 2

[0058] A copper sulfide mine in Anhui Province contains 8.22% sulfur and 0.63% copper. Acidic mine wastewater was collected from the mine's mining area, with Fe, Mg, Ca, and S content of 1342 mg / L, 2366 mg / L, 853 mg / L, and 3693 mg / L, respectively, and a pH of 2.36. The tailings after copper-sulfur separation serve as feed for the sulfur selection stage. In this embodiment, the flotation feed has a sulfur grade of 15.81%, a copper grade of 0.07%, and a -200 mesh content of 76.60%. Using the process of the present invention, the copper-sulfur separation tailings are directly separated using a cyclone. The cyclone has a sand inlet diameter of 14 mm, an overflow inlet diameter of 20 mm, and a feed pressure of 0.1 MPa, yielding a coarse-grained pyrite concentrate, K1. The overflow of the cyclone is then concentrated using a thickener to obtain an underflow pulp with a concentration of 40%. 200g / t butyl xanthate is then added directly to the underflow of the thickener, the pulp pH is adjusted to 10.01, and flotation is performed for 5 minutes to obtain a roughing I concentrate and a roughing I tailing. The obtained roughing I concentrate is then fined, i.e., fine I, with a pH of 9.54. No reagents are added to obtain a sulfur rough concentrate K2 and a fine I tailing. The fine I tailings are then mixed with the roughing I tailings and concentrated using a thickener to obtain a pulp with a concentration of 40%. The pyrite with poor floatability is then enriched using a process flow of 1 roughing, 1 fine, and 1 scavenging (i.e., roughing II, fine II, scavenging I). During roughing II, a combination of chitosan (chitosan molecular weight is 300,000) and sodium lignin sulfonate at a mass ratio of 200g / t and 250g / t of butyl xanthate and 0.4m 3 / t acid mine wastewater, 50g / t methyl isobutyl carbinol, pH 4.23, and flotation for 5 minutes yielded Rougher II concentrate and Rougher II tailings. The Rougher II concentrate was then subjected to beneficiation, known as Concentration II. The pulp concentration of Concentration II was 35%, the pH was 4.25, and flotation for 5 minutes yielded sulfur concentrate K2 and Concentration II tailings. The Rougher II tailings were then subjected to scavenging, known as Scavenging I, with 40g / t butyl xanthate added and a pH of 5.22. Flotation for 3 minutes yielded Scavenging I concentrate and tailings X. The beneficiation tailings and scavenging concentrate were returned to the roughing operation, forming a closed-loop system. Concentrates K1, K2, and K3 were used as the final concentrate product K, with a recovery rate and grade of 97.57% and 38.25%, respectively.

[0059] Example 3

[0060] A copper sulfide mine in Yunnan Province contains 13.23% sulfur and 0.51% copper. Acidic mine wastewater was collected from the mine's mining area and contained 3620 mg / L, 1856 mg / L, 2604 mg / L, and 3870 mg / L of Fe, Mg, Ca, and S, respectively, with a pH of 3.01. The tailings after copper-sulfur separation serve as feed for the sulfur selection stage. In this embodiment, the flotation feed has a sulfur grade of 30.03%, a copper grade of 0.08%, and a -200 mesh content of 85.02%. Using the process of the present invention, the copper-sulfur separation tailings are directly separated using a cyclone. The cyclone has a sand inlet diameter of 13 mm, an overflow inlet diameter of 24 mm, and a feed pressure of 0.06 MPa, yielding a coarse-grained pyrite concentrate, K1. The overflow of the cyclone is then concentrated using a thickener to obtain an underflow pulp with a concentration of 40%; 300g / t butyl xanthate is then directly added to the underflow of the thickener, the pulp pH is adjusted to 10.50, and flotation is performed for 5 minutes to obtain a roughing I concentrate and a roughing I tailings. The obtained roughing I concentrate is then fined, i.e., fine I, with a pH of 10.11. No reagents are added to obtain a sulfur rough concentrate K2 and a fine I tailings. The fine I tailings are then mixed with the roughing I tailings and concentrated using a thickener to obtain a pulp with a concentration of 40%. The pyrite with poor floatability is then enriched using a process flow of 1 roughing, 1 fine, and 1 scavenging (i.e., roughing II, fine II, scavenging I); during roughing II, a combination of 100g / t chitosan (chitosan molecular weight is 500,000) and sodium lignin sulfonate with a mass ratio of 1:6 is added, 350g / t of butyl xanthate, and 1.2m 3 / t acid mine wastewater, 50g / t methyl isobutyl carbinol, pH 3.23, and flotation for 5 minutes yielded Rougher II concentrate and Rougher II tailings. The Rougher II concentrate was then subjected to beneficiation, i.e., Concentration II. The pulp concentration of Concentration II was 35%, the pH was 3.44, and flotation for 5 minutes yielded sulfur concentrate K2 and Concentration II tailings. The Rougher II tailings were then subjected to scavenging, i.e., Scavenging I. The butyl xanthate dosage for Scavenging I was 60g / t, the pH was 3.37, and flotation for 3 minutes yielded Scavenging I concentrate and tailings X. The beneficiation tailings and scavenging concentrate were returned to the roughing operation, forming a closed-loop system. Concentrates K1, K2, and K3 were used as the final concentrate product K, with a recovery rate and grade of 96.18% and 48.14%, respectively.

Claims

1. A new method for the cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings, characterized in that: The following steps are involved: Step 1) Separation of coarse-grained pyrite by hydrocyclone: ​​For the copper-sulfur high-alkali separation tailings, a hydrocyclone is first used to preferentially separate the coarse-grained pyrite to obtain a coarse-grained sulfur concentrate; the diameter of the hydrocyclone grit port used in the cyclone sulfur selection is 9.5-15 mm, the diameter of the overflow pipe is 20-26 mm, and the feed pressure is 0.06-0.1 MPa; Step 2) alkaline pulp flotation of easily floatable pyrite: Under alkaline conditions, the overflow of the hydrocyclone is subjected to preliminary flotation to preferentially separate the highly floatable pyrite to obtain a highly floatable sulfur concentrate; the pyrite collector used in the alkaline sulfur separation is butyl xanthate, and the dosage is 90-320 g / t; the pH value of the slurry is between 8.5 and 10.5; Step 3) Activating flotation of pyrite with acid mine wastewater: flotation is performed by sequentially adding a pyrite collector, a conditioning agent, acid mine wastewater, and a frother to the ore pulp to obtain a weakly floatable sulfur concentrate; The amount of acid mine wastewater used in the pyrite flotation stage of activated acid mine wastewater is 0.18m 3 / t to 1.3m 3 / t; flotation concentration is 30% to 40%; pulp adjuster is a combination of chitosan and sodium lignin sulfonate, the mass ratio of the two is 1:4 to 1:9, and the total dosage is 50 to 200 g / t; collector is at least one of butyl xanthate, amyl xanthate, and isopentyl xanthate, and the dosage is 120 to 350 g / t; flotation frother is 2 # One of oil or methyl isobutyl carbinol, the dosage is 15-55g / t.

2. A novel method for cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings according to claim 1, characterized in that: The diameter of the sand settling port of the cyclone used in cyclone sulfur selection is 10 to 14 mm.

3. A novel method for cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings according to claim 1, characterized in that: The pyrite collector for alkaline sulfur separation is butyl xanthate, the dosage is 100-300 g / t, the flotation time is 3-5 min; the pH value of the ore pulp is between 8.5 and 10.

5.

4. A novel method for cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings according to claim 1, characterized in that: The activator used in the sulfur separation of acid mine drainage comes from acid mine drainage naturally generated in open pits or underground mines of copper sulfide mines.

5. A novel method for cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings according to claim 3, characterized in that: In the stage of activated flotation of pyrite with acid mine wastewater, the order of adding reagents is collector, slurry conditioner, acid mine wastewater, and frother.

6. A novel method for cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings according to claim 3, characterized in that: The amount of acid mine wastewater used in the pyrite flotation stage of activated acid mine wastewater is 0.2m 3 / t to 1.2m 3 / t; flotation concentration is 30% to 40%; the pulp adjuster is a combination of chitosan and sodium lignin sulfonate, the molecular weight of chitosan is 250,000 to 500,000, the mass ratio of the two is 1:5 to 1:8, and the total dosage is 100 to 200 g / t; the collector is at least one of butyl xanthate, amyl xanthate, and isopentyl xanthate, and the dosage is 120 to 350 g / t; the flotation frother is 2 # One of oil or methyl isobutyl carbinol, the dosage is 20-50g / t; the flotation time is 3-5min.

7. A novel method for cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings according to claim 6, characterized in that: The flotation pH of acid mine drainage sulfur separation is between 4.0-6.

0.

8. A novel method for cascade beneficiation and enrichment of copper-sulfur high-alkali separation tailings according to any one of claims 1 to 7, characterized in that: The sulfur grade of the obtained total sulfur concentrate is above 38%, and the total sulfur recovery rate is above 95%.

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