An industrial method for recovering pyrite from lead-zinc tailings by quality-based separation

Abstract

The application discloses an industrial method for grading and sorting lead-zinc tailings to recover pyrite, and adopts a black medicine type capturing agent to carry out a lead-zinc flotation process, and a lead-zinc tailings slurry obtained by not adopting a lime to inhibit pyrite process in the flotation process is subjected to magnetic separation according to the properties of main components of the pyrite and the pyrrhotite, weak magnetic tailings concentrated treatment is carried out on the tailings after grading and sorting, underflow liquid is obtained, and a combined method of roughing, scavenging and flotation is adopted to obtain final sulfur concentrate. The application realizes grading and sorting of the lead-zinc tailings, and efficiently and comprehensively recovers the pyrite resources in the lead-zinc tailings.

Description

Technical Field

[0001] This invention belongs to the field of lead-zinc tailings sorting technology, and particularly relates to an industrial method for sorting and recovering pyrite from lead-zinc tailings. Background Technology

[0002] The associated pyrite obtained after lead-zinc ore beneficiation, also known as lead-zinc tailings, is usually composed of pyrite and pyrrhotite, and is generally referred to as pyrite. The associated pyrite resources and output of non-ferrous mines occupy an important position in China's sulfur production. Pyrite has long been a major raw material for my country's sulfuric acid industry, and its slag can be used for iron and steel smelting. The sulfur grade in pyrite-based sulfur concentrate reaches over 46.5%, and the iron ore grade in the sulfuric acid slag after acid production can reach 62%. In contrast, the sulfur grade in pyrrhotite-based sulfur concentrate reaches over 36.0%, and the iron ore grade in the sulfuric acid slag after acid production can reach over 66%, achieving efficient recovery and utilization of pyrite and iron resources. With the continuous deepening of research on beneficiation theory, beneficiation technology has made significant progress; however, the beneficiation process still depends on the ore properties. Since the lead-zinc ore beneficiation process focuses on the recovery and indicators of valuable main elements such as lead, zinc, and copper, it usually adopts a priority flotation process, separating them one by one according to the order of easy to difficult (usually lead, zinc, and sulfur). Therefore, associated pyrite is often suppressed and enters the tailings, and its value is relatively low. Therefore, the efficient recovery of pyrite from the tailings has become the focus of research.

[0003] Currently, the technological processes for producing sulfur concentrate from lead-zinc tailings include conventional flotation, single gravity separation, and combined gravity-flotation processes. Research focuses primarily on the development of beneficiation techniques. Chinese patent application 201310380715.X discloses a combined gravity-flotation separation method for producing high-grade sulfur concentrate, mainly using gravity separation to recover pyrite. However, while gravity separation is effective for coarse particles, it performs poorly for fine particles smaller than 80 μm, and its throughput is relatively low, making operation and management complex.

[0004] For lead-zinc tailings containing both pyrite and pyrrhotite, two types of iron sulfide minerals, and whose gangue is mainly carbonate, how to process high-calcium, high-iron, and severely muddy lead-zinc tailings on a large scale, how to classify and sort them and recover the pyrite resources in a high-value comprehensive manner, and how to select a reasonable process are urgent problems to be solved. Summary of the Invention

[0005] In order to overcome the technical problems in the prior art, the present invention provides an industrial method for sorting and recovering pyrite from lead-zinc tailings.

[0006] To solve the above-mentioned technical problems, the present invention proposes the following technical solution:

[0007] This invention provides an industrial method for separating and recovering pyrite from lead-zinc tailings, comprising the following steps:

[0008] S1. Magnetic separation is performed on the lead-zinc tailings slurry to obtain weakly magnetic magnetic tailings after separation.

[0009] S2. The magnetic separation tailings obtained in step S1 are concentrated to obtain underflow and overflow.

[0010] S3. After adding an activating agent to the underflow obtained in step S2 and stirring, the sulfur flotation is carried out to obtain rough concentrate and rough tailings.

[0011] S4. After adding activating agents to the roughing tailings obtained in step S3 and stirring, perform flotation sulfur scavenging to obtain scavenging concentrate. Then, return the scavenging concentrate to the flotation sulfur roughing process in step S3.

[0012] S5. Add activating agent to the rough concentrate obtained in step S3 and stir before carrying out sulfur flotation. The sulfur flotation is carried out at least four times to obtain the final sulfur concentrate.

[0013] The lead-zinc tailings slurry is obtained by selectively pulling and pressing lead and zinc ore using black reagents and a "starvation" dosing method. It is obtained through a flotation process of lead first and then zinc, without using lime to suppress pyrite during the flotation process. The main components of the lead-zinc tailings slurry are pyrite and pyrrhotite.

[0014] This invention employs a xanthate-based flocculant in the extraction of lead and zinc from lead-zinc tailings. Its advantages include: firstly, its foaming properties eliminate the need for additional foaming agents; secondly, it allows lead to float at lower pH levels, eliminating the need for lime; thirdly, it offers good selectivity, reducing the use of zinc sulfate and sulfite, and replacing sodium sulfate and other inhibitors, thus minimizing environmental pollution during lead-zinc-sulfur separation; and fourthly, it facilitates the recovery of associated silver. Because this xanthate-based flocculant has a weaker collecting capacity but stronger selectivity than xanthates, particularly its weak collecting capacity for pyrite, significantly reduces the difficulty of sulfur recovery from lead-zinc tailings due to the lack of strong pyrite inhibition.

[0015] Lime acts as both an alkali and a depressant for pyrite. Its depressant function arises from the formation of a hydrophilic film of ferrous hydroxide and ferric hydroxide on the mineral surface. In the lead-zinc flotation process, lime is not used as a depressant, which facilitates the activation and recovery of pyrite. In the subsequent re-concentration of lead-zinc tailings, only a small amount of activator is needed, and a collector with weaker collecting power than xanthate can be used for flotation collection.

[0016] The lead-zinc tailings slurry in this invention includes pyrite and pyrrhotite. Pyrrhotite has poor floatability and poor flotation effect, but it has a certain degree of magnetism. The pyrite in the lead-zinc tailings slurry is weakly magnetic, and the pyrrhotite includes strongly magnetic monoclinic pyrrhotite and weakly magnetic hexagonal pyrrhotite. Based on the difference in magnetism, magnetic separation is first performed to obtain weakly magnetic tailings. Then, the obtained magnetic tailings are concentrated to reduce the amount of reagents used in subsequent processes. A portion of magnetic pyrite is recovered through magnetic separation. Although the sulfur grade is low, its iron content is high, and it can be sold as a high-iron-sulfur concentrate.

[0017] As an optional implementation, in the industrial method provided by the present invention, the pH value of the lead-zinc tailings slurry is 7.0 to 8.0, the proportion of solid matter with a mineral particle size of less than 74 μm in the slurry is 65 to 75%, the content of solid matter in the slurry is 15 to 25%, and the sulfur grade in the slurry is 13 to 23%.

[0018] Furthermore, given that the proportion of solids with a particle size smaller than 74 μm in the slurry of this invention is 65-75%, gravity separation is ineffective for fine-grained raw materials with high mud content. Therefore, flotation is the only option in this invention. The concentration process is employed due to the low fineness and concentration of the feed material. The slurry pH of 7.0-8.0 indicates a lime-free process, and the sulfur content range demonstrates its recovery value.

[0019] As an optional implementation, in the industrial method provided by the present invention, the proportion of solid matter with mineral particle size less than 20 μm in the lead-zinc tailings slurry is greater than 38%.

[0020] The finer the mineral particles, the more difficult the separation. In this invention, the proportion of solid matter with mineral particles smaller than 20 μm is greater than 38%, which increases the separation difficulty. Therefore, this invention adopts a method of magnetic separation followed by coarse separation, scavenging and fine separation, and the fine separation is performed at least four times.

[0021] As an optional implementation, in the industrial method provided by the present invention, the content of carbonate minerals in the slurry is 45% or more, and the content of calcite and dolomite in the carbonate minerals is ≥30%.

[0022] Since the lead-zinc tailings in this invention are carbonate minerals, with a carbonate mineral content of 45-55% in the slurry, and due to the good floatability and high content of carbonates, they are difficult to separate from the useful minerals. Adding acid during flotation would require a huge amount of acid, which is not conducive to industrial processing. Furthermore, due to the high carbonate content, sulfate-based activators are not used; only acidic substances such as copper sulfate can be used. Carbonates are generally relatively soft and easily become muddy, which can easily cover the surface of pyrite minerals, directly affecting the separation effect. Therefore, multiple cleaning processes and special middlings treatment methods are necessary. Thus, this invention employs a combination of roughing, scavenging, and cleaning processes, with the cleaning process involving at least four stages to gradually separate carbonates from the useful minerals and prepare a sulfur concentrate.

[0023] Calcite and dolomite, as gangue minerals, have good floatability and account for more than 30%, which increases the difficulty of sorting lead-zinc tailings in this invention.

[0024] As an optional implementation, in the industrial method provided by the present invention, in step S5, during the sulfur flotation process, the tailings obtained after the first sulfur flotation process and the tailings obtained after the second sulfur flotation process are combined, an activating agent is added, and after stirring, flotation is performed again.

[0025] As an optional implementation, in the industrial method provided by the present invention, the tailings of the first and second refinements are subjected to flotation again to obtain a re-refinement concentrate and a re-refinement tailings, wherein the re-refinement concentrate is returned to the sulfur roughing process in step S3, and the re-refinement tailings are returned to the sulfur scavenging process in step S4.

[0026] In this invention, the tailings obtained after two rounds of fine selection still have a high mud content, so it is necessary to add activating agents again for flotation. The resulting re-selected concentrate is returned to the roughing operation, and the re-selected tailings are returned to the scavenging operation.

[0027] As an optional implementation, in the industrial method provided by the present invention, in step S5, during the sulfur flotation process, the refined tailings obtained after the third sulfur flotation process are returned to the first sulfur flotation process in step S5.

[0028] As an optional implementation, in the industrial method provided by the present invention, in step S5, the tailings obtained after the fourth and subsequent sulfur flotation and cleaning processes are sequentially returned to the previous cleaning process.

[0029] In this invention, based on the elemental grade, particle size, and other physicochemical properties of the product after flotation, the tailings obtained after the third sulfur flotation process are returned to the first sulfur flotation process, and the tailings obtained after the fourth and subsequent sulfur flotation processes are returned to the previous stage of the refining operation. This facilitates the separation of lead and zinc tailings to obtain the final sulfur concentrate.

[0030] As an optional implementation, in the industrial method provided by the present invention, the activating agent includes an activator, a composite collector, and a foaming agent. The activator is copper sulfate or lead nitrate, the composite collector is a mixture of acrylonitrile dimethyl dithiocarbamate and sodium butadiene black powder, and the foaming agent is No. 2 oil or MIBC.

[0031] In this invention, acrylonitrile dimethyl dithiocarbamate and sodium butadiene black powder are used as a composite collector. When used in combination with copper sulfate or lead nitrate, the collecting effect is better.

[0032] As an optional implementation, in the industrial method provided by the present invention, the ratio of the activator to the composite collector is 1:2 to 2:1.

[0033] As an optional implementation method, in the industrial method provided by the present invention, the amount of activator is 50-300 g / t, the amount of compound collector is 50-300 g / t, and the amount of foaming agent is 10-50 g / t.

[0034] As an optional implementation, in the industrial method provided by the present invention, in step S2, the content of solids in the underflow liquid is 30-45%.

[0035] As an optional implementation, in the industrial method provided by the present invention, in steps S3-S5, the stirring time after adding the activating agent is 10-30 min.

[0036] Compared with the prior art, the beneficial effects of the present invention are as follows:

[0037] Lead-zinc ore mainly contains galena, sphalerite, and pyrite. Based on the difference in floatability and flotation velocity between galena and sphalerite, a selective collector (black powder-type collector) and depressant are used in a "starvation" dosing method. This selective, gentle pulling and pressing process achieves lime-free lead-zinc tailings prepared by sequential preferential flotation. The lead-zinc tailings mainly consist of two iron sulfide minerals: pyrite and pyrrhotite. Pyrite is a weakly magnetic mineral, and pyrrhotite is composed of strongly magnetic monoclinic pyrrhotite and weakly magnetic hexagonal pyrrhotite. This invention utilizes the difference in specific magnetization coefficients between pyrite and monoclinic pyrrhotite, employing the magnetic force generated by a permanent magnet weak magnetic separator to selectively enrich the strongly magnetic monoclinic pyrrhotite. Separation and treatment will produce a high-iron-grade sulfur concentrate, addressing the problem of poor floatability and difficulty in activation and recovery of pyrrhotite, thus reducing the recovery rate. It also prepares for subsequent separation, reducing volume, and improving quality. Magnetic separation separates high-iron-grade, highly magnetic sulfur concentrate from pyrrhotite. The remaining tailings from the magnetic separation are then subjected to flotation treatment, employing a combination of roughing, scavenging, and cleaning processes, with the cleaning process involving at least four stages, to gradually separate carbonates from valuable minerals and prepare sulfur concentrate. This method achieves the graded separation of lead-zinc tailings, efficiently and comprehensively recovering pyrite resources. This invention fully utilizes the wastewater from the magnetic separation and concentration operations for tiered treatment and reuse. The relatively high concentration of sulfur concentrate in the flotation process significantly reduces the amount of sulfur-removing reagents required, exhibiting characteristics of high targeting, high selectivity, high recovery rate, and low cost. Attached Figure Description

[0038] To more clearly illustrate the technical solutions in the embodiments of the present invention or the prior art, the drawings used in the description of the embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0039] Figure 1 This is an industrial process diagram for the separation and recovery of pyrite from lead-zinc tailings in this invention. Detailed Implementation

[0040] To facilitate understanding of the present invention, the present invention will be described more fully and in detail below with reference to the accompanying drawings and preferred embodiments, but the scope of protection of the present invention is not limited to the following specific embodiments.

[0041] Unless otherwise defined, all technical terms used herein have the same meaning as commonly understood by those skilled in the art. The technical terms used herein are for the purpose of describing particular embodiments only and are not intended to limit the scope of the invention.

[0042] Unless otherwise specified, all raw materials, reagents, instruments and equipment used in this invention can be purchased from the market or prepared by existing methods.

[0043] In this invention, the preparation method of lead-zinc tailings used in the following embodiments and comparative examples is as follows: The lead-zinc ore mainly contains galena, sphalerite, and pyrite. Based on the difference between the natural floatability and flotation velocity of galena and the floatability of sphalerite, the slurry after grinding and classification is treated with a selective collector (black powder-type collector) and an inhibitor (a combination of zinc sulfate and sodium sulfite). A "starvation" dosing method is used to selectively pull and press lightly, preferentially floating lead. Then, the lead tailings are treated with sulfur suppression and zinc activation. The zinc tailings are prepared by first adding an inhibitor (sodium thioglycolate inhibitor) and then adding an activator (copper sulfate) in sequence to obtain the lead-zinc tailings slurry of this invention.

[0044] Example 1

[0045] Lead-zinc tailings from a certain area in Hunan Province were selected as the raw ore. The main minerals include pyrite, pyrrhotite, siderite, garnet, calcite, quartz, and feldspar. Analysis showed that the main components of the ore sample were: S 15.72%, Fe 21.44%, Pb 0.26%, Zn 0.31%, CaO 21.55%, and SiO2 14.19%.

[0046] (1) Feed 26t / h of lead-zinc tailings slurry into a CTB-1218 wet permanent magnet drum separator, control the magnetic field strength to 0.25 Tesla, remove pyrrhotite, a magnetic mineral with a yield (the weight ratio of product to raw material) of 7.65%, and obtain high-iron sulfur concentrate and magnetic tailings. The sulfur grade in the high-iron sulfur concentrate reaches 36.07% (after roasting, iron concentrate with an iron grade of more than 66% can be obtained).

[0047] (2) The magnetic separation tailings are fed into the inclined plate thickening box for concentration, and the underflow concentration is controlled at about 30%. The overflow is returned to the previous process as recycled water.

[0048] (3) The bottom flow of the thickening tank is fed into the floating sulfur mixing tank, and 100g / t of copper sulfate activator, 200g / t of compound collector and 10g / t of frother No. 2 oil are added. After stirring for 20 minutes, it enters the roughing operation and the roughing concentrate and roughing tailings are obtained through separation.

[0049] (4) Add 50g / t of copper sulfate activator, 50g / t of compound collector, and 10g / t of No. 2 oil frother to the roughing tailings and then proceed to the scavenging operation to obtain scavenging concentrate and final tailings. The scavenging concentrate is returned to the roughing operation. Part of the final tailings is used as backfill for the goaf, and the remainder is discharged into the tailings dam.

[0050] (5) The rough concentrate is added with 20g / t of activator copper sulfate, 50g / t of compound collector and 10g / t of foaming agent No. 2 oil and then enters the first refining operation. After separation, the first refining concentrate and the first refining tailings are obtained.

[0051] (6) The concentrate from the first stage enters the second stage of the process, and after aeration separation, the second stage concentrate and the second stage tailings are obtained.

[0052] (7) After the tailings of Refinement 1 and Refinement 2 are combined, 20g / t of copper sulfate activator, 50g / t of compound collector and 10g / t of frother No. 2 oil are added to enter the re-selection operation. After separation, the re-selection concentrate and re-selection tailings are obtained. The re-selection concentrate is returned to the roughing operation and the re-selection tailings are entered into the scavenging operation.

[0053] (8) The No. 2 concentrate enters the No. 3 process. After separation, the No. 3 concentrate and No. 3 tailings are obtained. The No. 3 tailings are returned to the No. 1 process. The No. 3 concentrate enters the No. 4 process. After separation, the No. 4 concentrate and No. 4 tailings are obtained. The No. 4 tailings are returned to the No. 3 process. The No. 4 concentrate is the final sulfur concentrate. The sulfur grade of this product reaches 47.34% (after roasting, an iron concentrate with an iron grade of 63% can be obtained).

[0054] The mineral processing indicators are shown in Table 1 below.

[0055] Table 1: Results of the pyrite recovery process test in Example 1

[0056] Product Name Yield / % Grade S / % S recovery rate / % Pyrrhotite 7.65 36.07 17.68 Sulfur concentrate 20.97 47.34 63.61 Final tailings 71.38 4.09 18.71 raw ore 100.00 15.61 100.00

[0057] Example 2

[0058] Lead-zinc tailings from a certain area in Hunan Province were selected as the raw ore. The main minerals include pyrite, pyrrhotite, siderite, garnet, calcite, quartz, and feldspar. Analysis showed that the main components of the ore sample were: S 19.35%, Fe 23.62%, Pb 0.33%, Zn 0.25%, CaO 19.75%, and SiO2 16.33%.

[0059] (1) Feed 26t / h of lead-zinc tailings slurry into a CTB-1218 wet permanent magnet drum separator, control the magnetic field strength to 0.30 Tesla, remove pyrrhotite, a magnetic mineral with a yield of 8.05%, and obtain high-iron sulfur concentrate and magnetic tailings. The sulfur grade in the high-iron sulfur concentrate reaches 37.56% (after roasting, iron concentrate with an iron grade of more than 66% can be obtained).

[0060] (2) The magnetic separation tailings are fed into the inclined plate thickening box for concentration, and the underflow concentration is controlled at about 45%. The overflow is returned to the previous process as recycled water.

[0061] (3) The bottom flow of the thickening tank is fed into the floating sulfur mixing tank, and 120g / t of copper sulfate activator, 220g / t of compound collector and 15g / t of frother No. 2 oil are added. After stirring for 30 minutes, it enters the roughing operation and the roughing concentrate and roughing tailings are obtained through separation.

[0062] (4) Add 50g / t of copper sulfate activator, 40g / t of compound collector, and 10g / t of No. 2 oil frother to the roughing tailings and then proceed to the scavenging operation to obtain scavenging concentrate and final tailings. The scavenging concentrate is returned to the roughing operation. Part of the final tailings is used as backfill for the goaf, and the remainder is discharged into the tailings dam.

[0063] (5) The rough concentrate is added with 25g / t of activator copper sulfate, 50g / t of compound collector and 10g / t of frother No. 2 oil and then enters the first refining operation. After separation, the first refining concentrate and the first refining tailings are obtained.

[0064] (6) The concentrate from the first stage enters the second stage of the process, and after aeration separation, the second stage concentrate and the second stage tailings are obtained.

[0065] (7) After the tailings of Refinement 1 and Refinement 2 are combined, 25g / t of copper sulfate activator, 40g / t of compound collector and 10g / t of frother No. 2 oil are added to enter the re-selection operation. After separation, the re-selection concentrate and re-selection tailings are obtained. The re-selection concentrate is returned to the roughing operation and the re-selection tailings are entered into the scavenging operation.

[0066] (8) The No. 2 concentrate enters the No. 3 process. After separation, the No. 3 concentrate and No. 3 tailings are obtained. The No. 3 tailings are returned to the No. 1 process. The No. 3 concentrate enters the No. 4 process. After separation, the No. 4 concentrate and No. 4 tailings are obtained. The No. 4 tailings are returned to the No. 3 process. The No. 4 concentrate is the final sulfur concentrate. The sulfur grade of this product reaches 47.88% (after roasting, an iron concentrate with an iron grade of 61% can be obtained).

[0067] The mineral processing indicators are shown in Table 2 below.

[0068] Table 2: Results of the pyrite recovery process test in Example 2

[0069] Product Name Yield / % Grade S / % S recovery rate / % Pyrrhotite 8.05 37.56 15.55 Sulfur concentrate 25.06 47.88 61.69 Final tailings 66.89 6.62 22.77 raw ore 100.00 19.45 100.00

[0070] Comparative Example 1

[0071] Lead-zinc tailings from a certain area in Hunan Province were selected. The main minerals include pyrite, pyrrhotite, siderite, garnet, calcite, quartz, and feldspar. Analysis showed that the main components of the ore sample were: S 15.72%, Fe 21.44%, Pb 0.26%, Zn 0.31%, CaO 21.55%, and SiO2 14.19%.

[0072] (1) Feed 26t / h of lead-zinc tailings slurry into a CTB-1218 wet permanent magnet drum separator, control the magnetic field strength to 0.25T, and remove pyrrhotite, a magnetic mineral with a yield of 7.80%.

[0073] (2) The magnetic separation tailings are fed into the inclined plate thickening box, and the underflow concentration is controlled at about 30%. The overflow is returned to the previous process as recycled water.

[0074] (3) The bottom flow of the thickening tank is fed into the floating sulfur mixing tank, and 100g / t of copper sulfate activator, 200g / t of compound collector and 10g / t of frother No. 2 oil are added. After stirring for 20 minutes, it enters the roughing operation and the roughing concentrate and roughing tailings are obtained through separation.

[0075] (4) Add 50g / t of copper sulfate activator, 50g / t of compound collector, and 10g / t of No. 2 oil frother to the roughing tailings and then proceed to the scavenging operation to obtain scavenging concentrate and final tailings. The scavenging concentrate is returned to the roughing operation. Part of the final tailings is used as backfill for the goaf, and the remainder is discharged into the tailings dam.

[0076] (5) The rough concentrate is added with 20g / t of activator copper sulfate, 50g / t of compound collector and 10g / t of foaming agent No. 2 oil and then enters the first refining operation. After separation, the first refining concentrate and the first refining tailings are obtained.

[0077] (6) The tailings of the first fineness are returned to the roughing operation; the concentrate of the first fineness enters the second fineness operation, and after separation, the second fineness concentrate and the second fineness tailings are obtained. The tailings of the second fineness are returned to the first fineness operation; the concentrate of the second fineness enters the third fineness operation, and after separation, the third fineness concentrate and the third fineness tailings are obtained. The tailings of the third fineness are returned to the second fineness operation; the concentrate of the third fineness enters the fourth fineness operation, and after separation, the fourth fineness concentrate and the fourth fineness tailings are obtained. The tailings of the fourth fineness are returned to the third fineness operation. The fourth fineness concentrate is the final sulfur concentrate.

[0078] The difference between Comparative Example 1 and Example 1 is that: tailings from Refinement 1 are returned to the roughing operation, tailings from Refinement 2 are returned to Refinement 1, tailings from Refinement 3 are returned to Refinement 2, and tailings from Refinement 4 are returned to Refinement 3.

[0079] The mineral processing indicators are shown in Table 3 below.

[0080] Table 3: Results of the pyrite recovery process test in Comparative Example 1

[0081] Product Name Yield / % Grade S / % S recovery rate / % Pyrrhotite 7.80 36.22 18.01 Sulfur concentrate 22.21 43.36 61.38 Final tailings 69.99 4.62 20.61 raw ore 100.00 15.69 100.00

[0082] Comparative Example 2

[0083] Lead-zinc tailings from a certain area in Hunan Province were selected as the raw ore. The main minerals include pyrite, pyrrhotite, siderite, garnet, calcite, quartz, and feldspar. Analysis showed that the main components of the ore sample were: S 15.72%, Fe 21.44%, Pb 0.26%, Zn 0.31%, CaO 21.55%, and SiO2 14.19%.

[0084] (1) Feed 26t / h of lead-zinc tailings slurry into a CTB-1218 wet permanent magnet drum separator, control the magnetic field strength to 0.25 Tesla, remove pyrrhotite, a magnetic mineral with a yield of 7.68%, and obtain high-iron sulfur concentrate and magnetic tailings. The sulfur grade in the high-iron sulfur concentrate reaches 36.15% (after roasting, iron concentrate with an iron grade of more than 66% can be obtained).

[0085] (2) The magnetic separation tailings are fed into the inclined plate thickening box for concentration, and the underflow concentration is controlled at about 30%. The overflow is returned to the previous process as recycled water.

[0086] (3) The bottom flow of the thickening tank is fed into the floating sulfur mixing tank, and 100g / t of copper sulfate activator, 200g / t of butyl xanthate collector and 10g / t of No. 2 oil frother are added. After stirring for 20 minutes, the roughing operation is carried out, and the roughing concentrate and roughing tailings are obtained by separation.

[0087] (4) Add 50g / t of copper sulfate activator, 50g / t of butyl xanthate collector, and 10g / t of No. 2 oil frother to the roughing tailings and then proceed to the scavenging operation to obtain scavenging concentrate and final tailings. The scavenging concentrate is returned to the roughing operation. Part of the final tailings is used for backfilling of the goaf, and the remainder is discharged into the tailings dam.

[0088] (5) The rough concentrate is added with 20g / t of copper sulfate activator, 50g / t of butyl xanthate collector and 10g / t of No. 2 oil frother and then enters the first refining operation. After separation, the first refining concentrate and the first refining tailings are obtained.

[0089] (6) The concentrate from the first stage enters the second stage of the process, and after aeration separation, the second stage concentrate and the second stage tailings are obtained.

[0090] (7) After the tailings of Refinement 1 and Refinement 2 are combined, 20g / t of copper sulfate activator, 50g / t of butyl xanthate collector and 10g / t of No. 2 oil frother are added to enter the re-selection operation. After separation, the re-selection concentrate and re-selection tailings are obtained. The re-selection concentrate is returned to the roughing operation and the re-selection tailings are entered into the scavenging operation.

[0091] (8) The No. 2 concentrate enters the No. 3 process. After separation, the No. 3 concentrate and No. 3 tailings are obtained. The No. 3 tailings are returned to the No. 1 process. The No. 3 concentrate enters the No. 4 process. After separation, the No. 4 concentrate and No. 4 tailings are obtained. The No. 4 tailings are returned to the No. 3 process. The No. 4 concentrate is the final sulfur concentrate. The sulfur grade of this product reaches 47.34% (after roasting, an iron concentrate with an iron grade of 61.19% can be obtained).

[0092] The difference between Comparative Example 2 and Example 1 is that the trapping agent is butyl xanthate.

[0093] The mineral processing indicators are shown in Table 4 below.

[0094] Table 4: Results of the pyrite recovery process test in Comparative Example 2

[0095] Product Name Yield / % Grade S / % S recovery rate / % Pyrrhotite 7.68 36.15 17.58 Sulfur concentrate 21.55 46.55 63.51 Final tailings 70.77 4.22 18.91 raw ore 100.00 15.79 100.00

[0096] The above description, in conjunction with specific preferred embodiments, provides a further detailed explanation of the present invention. However, it should not be construed that the specific implementation of the present invention is limited to these descriptions. For those skilled in the art, various simple deductions or substitutions can be made without departing from the concept of the present invention, and all such modifications and substitutions should be considered within the scope of protection of the present invention.

Claims

1. An industrial method for separating and recovering pyrite from lead-zinc tailings, characterized in that, Includes the following steps: S1. Magnetic separation is performed on the lead-zinc tailings slurry, and weak magnetic tailings are obtained after separation. S2. The magnetic separation tailings obtained in step S1 are concentrated to obtain underflow and overflow. S3. After adding an activating agent to the underflow obtained in step S2 and stirring, the sulfur flotation roughing is carried out to obtain roughing concentrate and roughing tailings. S4. Add activating agent to the roughing tailings obtained in step S3, stir and then carry out flotation scavenging to obtain scavenging concentrate. Then return the scavenging concentrate to the flotation roughing process in step S3. S5. Add activating agent to the rough concentrate obtained in step S3 and stir before carrying out sulfur flotation and cleaning. The sulfur flotation and cleaning is carried out at least four times to obtain the final sulfur concentrate. In the sulfur flotation process, the tailings obtained from the first sulfur flotation and the tailings obtained from the second sulfur flotation are combined, activating agents are added, and after stirring, they are floated again. The tailings obtained from the first and second sulfur flotation processes are floated again to obtain a re-selection concentrate and a re-selection tailings. The re-selection concentrate is returned to the sulfur flotation roughing process in step S3 for sulfur flotation roughing, and the re-selection tailings are returned to the sulfur flotation scavenging process in step S4 for sulfur flotation scavenging. In the sulfur flotation process, the tailings obtained after the third sulfur flotation process are returned to the sulfur flotation process in step S5 as raw material for further sulfur flotation. The tailings obtained after the fourth and subsequent sulfur flotation and beneficiation processes are returned to the previous beneficiation process as raw materials for further sulfur flotation and beneficiation. The activating agent includes an activator, a compound collector, and a foaming agent. The activator is copper sulfate or lead nitrate, and the compound collector is a mixture of acrylonitrile dimethyl dithiocarbamate and sodium butadiene black powder. The lead-zinc tailings slurry is obtained by selectively pulling and pressing lead and zinc ore using black reagents and a "starvation" dosing method. It is obtained through a flotation process of lead first and then zinc, without the use of lime to suppress pyrite during the flotation process. The main components of the lead-zinc tailings slurry are pyrite and pyrrhotite, and the carbonate mineral content of the slurry is above 45%.

2. The industrial method for separating and recovering pyrite from lead-zinc tailings according to claim 1, characterized in that, The pH value of the lead-zinc tailings slurry is 7.0-8.0, the proportion of solid matter with a mineral particle size of less than 74μm in the slurry is 65-75%, the content of solid matter in the slurry is 15-25%, and the sulfur grade in the slurry is 13-23%.

3. The industrial method for separating and recovering pyrite from lead-zinc tailings according to claim 1, characterized in that, The proportion of solid matter with mineral particles smaller than 20 μm in the lead-zinc tailings slurry is greater than 38%.

4. The industrial method for separating and recovering pyrite from lead-zinc tailings according to claim 1, characterized in that, The content of calcite and dolomite in the carbonate minerals is ≥30%.

5. The industrial method for separating and recovering pyrite from lead-zinc tailings according to claim 1, characterized in that, The ratio of the activator to the composite collector is 1:2 to 2:1.

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