CN117718138B - A method for separating and recovering complex tin-lead-zinc polymetallic ores and its application - Google Patents

A method for separating and recovering complex tin-lead-zinc polymetallic ores and its application Download PDF

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CN117718138B
CN117718138B CN202410175681.9A CN202410175681A CN117718138B CN 117718138 B CN117718138 B CN 117718138B CN 202410175681 A CN202410175681 A CN 202410175681A CN 117718138 B CN117718138 B CN 117718138B
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concentrate
flotation
tailings
tin
lead
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CN117718138A (en
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吴伯增
胡明振
邱鸿鑫
孙晓豪
阙山东
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Guangxi Huaxi Nonferrous Metals Co ltd
Guangxi Senhe High Technology Co ltd
China University of Mining and Technology Beijing CUMTB
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Guangxi Huaxi Nonferrous Metals Co ltd
Guangxi Senhe High Technology Co ltd
China University of Mining and Technology Beijing CUMTB
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Abstract

本发明涉及选矿技术领域，具体而言，涉及一种复杂锡铅锌多金属矿分离回收方法及其应用。复杂锡铅锌多金属矿分离回收方法包括分级和抛废、等可浮浮选、铅与锌硫的分离浮选、锌‑硫分离浮选、硫‑砷混浮、硫‑砷分离、脱硫浮选、浮锡和收锡等步骤。该方法能够实现锡铅锌多金属矿中有价元素的有效富集和综合回收。

The present invention relates to the field of mineral processing technology, and in particular to a complex tin-lead-zinc polymetallic ore separation and recovery method and its application. The complex tin-lead-zinc polymetallic ore separation and recovery method comprises the steps of classification and discarding, iso-floatation, lead and zinc-sulfur separation flotation, zinc-sulfur separation flotation, sulfur-arsenic mixed flotation, sulfur-arsenic separation, desulfurization flotation, tin flotation and tin collection. The method can achieve the effective enrichment and comprehensive recovery of valuable elements in tin-lead-zinc polymetallic ore.

Description

Separation and recovery method for complex tin-lead-zinc multi-metal ore and application thereof
Technical Field
The invention relates to the technical field of mineral separation, in particular to a separation and recovery method for complex tin-lead-zinc polymetallic ores and application thereof.
Background
Tin, lead, antimony and zinc are used as important nonferrous metal mineral resources required by industrial development, and development and utilization of the important nonferrous metal mineral resources are always paid attention. However, with continuous exploitation of high-quality rich mineral resources, the comprehensive grade of tin-lead-zinc multi-metal sulfide ores is gradually reduced, and the ore bodies mainly have the following characteristics: (1) The ore is of various types and has great separation difficulty, and the ore body usually contains cassiterite, marmatite, jamesonite, pyrite, pyrrhotite, arsenopyrite, small amount of galena, sphalerite, chalcopyrite, tetrahedrite and the like; (2) The ores are unevenly embedded, the sulfide ores are unevenly embedded mainly of fine grains, the mutual embedding is compact, and the cassiterite crystallization granularity is unevenly embedded mainly of coarse grains or mainly of fine grains; (3) high gangue mineral content.
At present, as tin, lead and zinc sulfide ores have different crystal structures and large property differences, the comprehensive recovery mainly has the following difficulties: (1) lean ore grade is low: the quartz stone content in the gangue is high and can reach 65% -70%, the cassiterite is embedded unevenly in thickness, the sulfide minerals and the gangue are mutually wrapped in fine particles, and the gangue minerals have great influence on the subsequent flow; (2) The cassiterite is seriously over-pulverized in the ore grinding process, and the sulfide ore is hard to float: the cassiterite is brittle and is difficult to grind in the sulfide ore, so that the cassiterite is extremely easy to excessively grind in the grinding process, and a large amount of generated secondary slime influences the subsequent sorting effect, and a large amount of tin resources are lost if the secondary slime is not recycled; (3) The floatation of the sulphide ores has the advantages of high medicine consumption, high tin inclusion and high separation difficulty of the sulphide ores: on the premise of coarser granularity of the mixed floating ore feeding, the method of increasing the dosage and 'strong pulling' is adopted to recover the sulphide ore. The circulation quantity of the mixed floating middlings of the sulphide ores is large, the floatation action time is long, and the floatability difference of various sulphide ores is very large, so that excessive medicament is adsorbed on the surface of part of the minerals with high easy floating adsorption capacity, and middlings are repeatedly circulated. In order to float the sulphide ore thoroughly, only a lengthening process is adopted, and a large amount of medicament is added, so that the medicament consumption is increased. The foam layer of the concentrate is thin, the concentration water leakage phenomenon is serious, fine-grain cassiterite is easy to be mixed in sulfide ore foam, so that the sulfide ore is high in tin inclusion, and the loss of tin metal is caused. (4) The amount of the floating tin mud is large, the property is complex, and the beneficiation and recovery difficulty is large: when treating lean tin ores with uneven cassiterite crystallization thickness, the ore needs to be further finely ground, and the grinding difficulty of the lean ores is high, so that the amount of ore slurry is increased, the cassiterite loss during desliming is serious, and the loss rate is up to 45% -50%.
Based on the method, the flotation system and the process flow are critical for comprehensive recovery of complex tin-lead-zinc polymetallic ores. However, no full-component recovery process for multi-metallic ores containing tin, lead and zinc has been reported at present.
In view of this, the present invention has been made.
Disclosure of Invention
The first aim of the invention is to provide a separation and recovery method for complex tin-lead-zinc multi-metal ores, which can realize effective enrichment and comprehensive recovery of valuable elements in the tin-lead-zinc multi-metal ores. Solves the problems that the tin, lead and zinc sulfide ores have different crystal structures and large property difference, are difficult to comprehensively recover and the full-component recovery process of the multi-metal ores containing tin, lead and zinc does not exist at present.
The second purpose of the invention is to provide the application of the separation and recovery method of the complex tin-lead-zinc multi-metal ore in ore dressing.
In order to achieve the above object of the present invention, the following technical solutions are specifically adopted:
the invention firstly provides a method for separating and recovering complex tin-lead-zinc multi-metal ores, which comprises the following steps:
(a) Screening complex tin-lead-zinc multi-metal ores into low-grade ores, medium-grade ores and high-grade ores, wherein the medium-grade ores are subjected to jigging and beneficiation to obtain jigged rough concentrates and jigged middlings, and the jigged rough concentrates are subjected to secondary grinding and then are mixed with the low-grade ores and subjected to reselection to obtain front heavy concentrates and overflows; carrying out deck floating table flotation on the front heavy concentrate to obtain first tin concentrate and deck floating tailings; mixing the table floating tailings with the jigged middlings and carrying out three-stage ore grinding to obtain combined ores;
(b) Performing primary roughing, secondary concentration and tertiary scavenging on the combined ore obtained in the step (a) to obtain iso-floatable concentrate and iso-floatable tailings;
(c) Separation flotation of lead and zinc sulphur: concentrating the equal floatable concentrate obtained in the step (b) and performing four-stage grinding, then adjusting the pH value and performing primary roughing flotation to obtain roughing concentrate and roughing tailings;
lead concentrate flotation: carrying out twice concentration flotation on the roughing concentrate to obtain lead-containing concentrate, wherein the lead-containing concentrate comprises at least one of lead-antimony concentrate, lead-silver concentrate and lead-copper concentrate;
(d) Zinc-sulfur separation: carrying out three scavenging flotation on the rougher tailings obtained in the step (c) to obtain scavenging concentrate and scavenging tailings; the concentration and the pH value of ore pulp of the scavenging tailings are regulated, and then primary roughing flotation is carried out to obtain rough zinc concentrate and rough sulfur tailings; carrying out three-time concentration flotation on the coarse zinc concentrate to obtain zinc concentrate and concentrated middlings; performing twice scavenging flotation on the coarse sulfur tailings to obtain first sulfur concentrate and zinc sulfur tailings;
(e) Sulfur-arsenic mixed floating: magnetically separating the zinc-sulfur tailings obtained in the step (d) to obtain magnetic tailings and magnetic concentrate, and performing primary roughing, secondary concentration and tertiary scavenging on the underflow obtained after the classification treatment of the magnetic tailings to obtain sulfur-arsenic mixed flotation concentrate and mixed flotation tailings;
Performing first shaking table reselection on the mixed tailings to obtain low-grade tin reselected concentrate;
(f) Sulfur-arsenic separation: carrying out primary roughing, secondary concentration and tertiary scavenging on the sulfur-arsenic mixed concentrate obtained in the step (e) to obtain second sulfur concentrate;
mixing the second sulfur concentrate with the magnetic concentrate obtained in the step (e) to obtain a sulfur concentrate product;
(g) Desulfurizing and floating: concentrating the overflow obtained in the step (a) to obtain settled sand, regulating the concentration and pH of the settled sand, and then performing desulfurization flotation of one roughing, two concentration and three scavenging to obtain third sulfur concentrate and first desulfurization flotation tailings; the third sulfur concentrate returns to the step (c) and is mixed with the floatable concentrate;
(h) Float tin and harvest tin: adjusting the concentration and the pH value of the first desulfurization flotation tailings obtained in the step (g), and then performing primary roughing, secondary concentration and tertiary scavenging to obtain a tin concentrate; carrying out magnetic separation on the float tin concentrate to obtain magnetic separation tailings, and carrying out gravity separation on the magnetic separation tailings by a second shaking table to obtain fine tin concentrate;
(i) Tin collection: performing third shaking table reselection on the floatable tailings obtained in the step (b) to obtain second tin concentrate and shaking table tailings; mixing the second tin concentrate with the first tin concentrate obtained in the step (a) to obtain a tin concentrate product; the concentration and the pH value of the shaking table tailings are regulated after five-stage grinding, and then desulfurization flotation of primary roughing, secondary concentration and tertiary scavenging is carried out to obtain fourth sulfur concentrate and second desulfurization flotation tailings; the fourth sulfur concentrate returns to the step (c) and is mixed with the floatable concentrate; and carrying out fourth shaking table gravity separation desulfurization on the second desulfurization flotation tailings to obtain moderate tin concentrate.
The invention also provides application of the complex tin-lead-zinc multi-metal ore separation and recovery method in ore dressing.
Compared with the prior art, the invention has the beneficial effects that:
(1) The separation and recovery method for the complex tin-lead-zinc multi-metal ore can realize effective enrichment and comprehensive recovery of valuable elements in the complex tin-lead-zinc multi-metal ore.
(2) The method for separating and recovering the complex tin-lead-zinc polymetallic ore provided by the invention adopts the flow ideas of 'stage grinding', 'stage dressing', 'coarse grain early collecting', 'edge grinding and losing', 'coarse fine grinding', and 'fine mud return and concentrated treatment', realizes the comprehensive recovery of tin, lead, zinc and sulfur, has simple separation process, is easy to popularize and has high recovery rate.
(3) According to the method for separating and recovering the complex tin-lead-zinc polymetallic ore, provided by the invention, aiming at the difficult problems that the gangue mineral content is high and the subsequent separation process is easy to influence, the multistage combined waste discarding process consisting of front-stage reselection, jigging and the like is adopted, 30% -40% of waste stones are discarded in advance, and the sorting grade and the processing capacity of the main process can be improved.
(4) According to the separation and recovery method for the complex tin-lead-zinc multi-metal ore, provided by the invention, aiming at the difficult problems that the difference of grindability of the cassiterite and the sulfide ore is large and the excessive grinding is easy, ore branches with different grades and different grindabilities are respectively ground, the excessive grinding of the cassiterite in the ore grinding process is reduced, and the mixed flotation and the selection granularity of the sulfide ore is reduced.
Drawings
In order to more clearly illustrate the embodiments of the present invention or the technical solutions in the prior art, the drawings that are needed in the description of the embodiments or the prior art will be briefly described, and it is obvious that the drawings in the description below are some embodiments of the present invention, and other drawings can be obtained according to the drawings without inventive effort for a person skilled in the art.
Fig. 1 is a schematic flow chart of a separation and recovery method of complex tin-lead-zinc multi-metal ores.
Detailed Description
The technical solution of the present invention will be clearly and completely described below with reference to the accompanying drawings and detailed description, but it will be understood by those skilled in the art that the examples described below are some, but not all, examples of the present invention, and are intended to be illustrative of the present invention only and should not be construed as limiting the scope of the present invention. All other embodiments, which can be made by those skilled in the art based on the embodiments of the invention without making any inventive effort, are intended to be within the scope of the invention. The specific conditions are not noted in the examples and are carried out according to conventional conditions or conditions recommended by the manufacturer. The reagents or apparatus used were conventional products commercially available without the manufacturer's attention.
In a first aspect, the present invention provides a method for separating and recovering complex tin-lead-zinc multi-metal ores, as shown in fig. 1, which is a schematic flow chart of the method, specifically includes the following steps:
(a) Classification and waste disposal: screening complex tin-lead-zinc multi-metal ores (raw ores) into low-grade ores, medium-grade ores and high-grade ores, wherein the granularity of the high-grade ores is larger than that of the medium-grade ores; the medium-sized ore is subjected to jigging and beneficiation to obtain jigged rough concentrate, jigged middling and jigged tailings, and the jigged tailings are directly thrown to waste; mixing the jigged rough concentrate with the low-grade ore after secondary grinding and reselecting to obtain front heavy concentrate and overflow; carrying out deck floating bed flotation on the front heavy concentrate to obtain first tin concentrate and deck floating tailings; and mixing the table floating tailings with the jigged middlings, and carrying out three-stage ore grinding to obtain combined ores.
(b) And (c) performing primary roughing, secondary concentration (namely, primary concentration and secondary concentration) and tertiary scavenging (namely, primary scavenging, secondary scavenging and tertiary scavenging) on the combined ore obtained in the step (a) to obtain the primary floatable concentrate and the primary floatable tailings. Among them, the purpose of the floatable flotation is to comprehensively recover lead-zinc (lead-zinc ore and zinc-sulfur ore).
(c) Separation flotation of lead (lead antimony, lead silver or lead copper) and zinc sulphur: concentrating and four-stage grinding the iso-floatable concentrate obtained in the step (b), then adjusting the pH value, and performing primary roughing flotation to obtain roughing concentrate and roughing tailings.
Lead concentrate flotation: and carrying out twice concentration flotation on the roughing concentrate to obtain final lead-containing concentrate, wherein the lead-containing concentrate comprises at least one of lead-antimony concentrate, lead-silver concentrate and lead-copper concentrate.
(d) Zinc-sulfur separation: carrying out three scavenging flotation on the rougher tailings obtained in the step (c) to obtain scavenging concentrate and scavenging tailings; then, the concentration and the pH value of ore pulp of the scavenging tailings are regulated, and after mixing for 3-5 minutes in a stirring barrel, primary roughing flotation is carried out, so that rough zinc concentrate and rough sulfur tailings are obtained; carrying out three-time concentration flotation on the coarse zinc concentrate to obtain final zinc concentrate and concentrated middlings; and carrying out twice scavenging flotation on the coarse sulfur tailings to obtain first sulfur concentrate and zinc sulfur tailings.
(e) Sulfur-arsenic mixed floating: and (d) carrying out magnetic separation on the zinc-sulfur tailings obtained in the step (d) to remove magnetic minerals to obtain magnetic separation tailings and magnetic concentrate, and carrying out primary roughing, secondary concentration and tertiary scavenging on the underflow obtained after the classification treatment of the magnetic separation tailings to obtain sulfur-arsenic mixed flotation concentrate and mixed flotation tailings.
And (3) carrying out first shaking table reselection on the mixed tailings to obtain low-grade tin reselected concentrate. And the tailings obtained by the gravity separation of the mixed floating tailings through the first shaking table are discarded.
(f) Sulfur-arsenic separation: and (3) carrying out one-time roughing, two-time concentration and three-time scavenging on the sulfur-arsenic mixed concentrate obtained in the step (e) to obtain second sulfur concentrate and tailings (waste).
Mixing the second sulfur concentrate with the magnetic concentrate obtained in the step (e) to obtain a final sulfur concentrate product. The sulfur concentrate product can be used as a chemical raw material, and real resources are recycled in a green way.
(g) Desulfurizing and floating: concentrating the overflow obtained in the step (a) to obtain settled sand, regulating the concentration and pH of the settled sand, and then performing desulfurization flotation of one roughing, two concentration and three scavenging to obtain third sulfur concentrate and first desulfurization flotation tailings; and (c) returning the third sulfur concentrate to the step (c) and mixing with the floatable concentrate.
(h) Float tin and harvest tin: adjusting the concentration and the pH value of the first desulfurization flotation tailings obtained in the step (g), and then performing primary roughing, secondary concentration and tertiary scavenging to obtain a tin concentrate; and carrying out magnetic separation on the float tin concentrate to remove magnetic minerals, so as to obtain magnetic separation tailings, and carrying out gravity separation on the magnetic separation tailings through a second shaking table to obtain fine tin concentrate.
(i) Tin collection: performing third shaking table reselection on the floatable tailings obtained in the step (b) to obtain second tin concentrate and shaking table tailings; mixing the second tin concentrate with the first tin concentrate obtained in the step (a) to obtain a tin concentrate product; the concentration and the pH value of the shaking table tailings are regulated after five-stage grinding, and then desulfurization flotation of primary roughing, secondary concentration and tertiary scavenging is carried out to obtain fourth sulfur concentrate and second desulfurization flotation tailings; the fourth sulfur concentrate returns to the step (c) and is mixed with the floatable concentrate; and (3) carrying out fourth shaking table gravity separation desulfurization on the second desulfurization flotation tailings according to the difference of mineral density to obtain moderate tin concentrate and tailings (waste).
The separation and recovery method for the complex tin-lead-zinc multi-metal ore can realize effective enrichment and comprehensive recovery of valuable elements in the complex tin-lead-zinc multi-metal ore.
Specifically, aiming at the difficult problems of high gangue mineral content and easiness in influencing the subsequent separation process, the invention adopts a multi-stage combined waste discarding process consisting of front-stage reselection, jigging and the like, and 30% -40% of waste stones are discarded in advance, so that the selection grade and the processing capacity of the main process can be improved.
In addition, the invention respectively grinds the ore branches with different grades and different grindabilities aiming at the difficult problems that the difference of grindability of the cassiterite and the sulfide ore is large and the excessive grinding is easy to occur, thereby reducing the excessive grinding of the cassiterite in the grinding process and reducing the mixed flotation and the selection granularity of the sulfide ore.
The invention adopts the flow thinking of ' stage grinding ', ' stage beneficiation ', ' coarse grain early collection ', ' edge grinding and losing ', ' coarse fine grinding ', and ' fine mud return to queue ', and centralized treatment ', realizes the comprehensive recovery of all components of the tin-lead-zinc multi-metal sulfide ore, and has high recovery rate.
In some specific embodiments, in the step (a), the granularity of the low-grade ore is less than 2mm, the granularity of the medium-grade ore is 2-4 mm, and the granularity of the high-grade ore is greater than 4mm.
In some specific embodiments, in step (a), the high-grade ore is subjected to one-stage grinding and then returned to the upper-stage screening process.
In some embodiments, in step (a), the reselection is performed in a conical spiral chute system. The high-efficiency gravity separation equipment of the conical spiral chute system is adopted for waste disposal, which is beneficial to improving the grade and recovery rate of valuable elements.
In some specific embodiments, in the step (a), the jigged rough concentrate is mixed with the low-grade ore after secondary ore grinding and enters a conical spiral chute system for gravity separation, and the bottom flow after enrichment is front heavy concentrate and enters a table floating table for table floating table flotation; and overflows into a thickener for concentration.
In some specific embodiments, in the step (a), the bench flotation table has a flotation stroke of 5-20 mm, including but not limited to a point value of any one of 5mm, 8mm, 10mm, 13mm, 15mm, 18mm, 20mm or a range value between any two; the table floating bed is subjected to floating at a frequency of 240-360 times/min, including but not limited to any one point value or any range value between any two of 240 times/min, 250 times/min, 280 times/min, 300 times/min, 330 times/min, 350 times/min and 360 times/min.
In some embodiments, in step (a), the collector used in the bench flotation bed comprises cinnamic hydroxamic acid having the formula C 6 H 5 -ch=ch- (c=o) NHOH, the amount of cinnamyl hydroxamic acid is 200-400 g/t, including but not limited to a point value of any one of 200g/t, 250g/t, 300g/t, 350g/t, 400g/t, or a range value between any two.
Cinnamic hydroxamic acid exhibits a stronger capture capacity than other collectors, while retaining Ca on gangue mineral surfaces 2+ Such as calcite, fluorite, etc., exhibit a weak adsorption capacity, which is advantageous for the improvement of the quality of cassiterite.
The preparation method of the cinnamon hydroxamic acid comprises the following steps: and (3) mixing the styryl, the N-hydroxyformamide and the nitric acid in an inert atmosphere, heating to react, and sequentially distilling and drying after the reaction is completed to obtain the cinnamon hydroxamic acid.
In some specific embodiments, the molar ratio of styryl, N-hydroxyformamide and nitric acid is 1:2-4:1-3.
In some specific embodiments, the heating temperature is 200-300 ℃, including but not limited to any one of 200 ℃, 220 ℃, 250 ℃, 280 ℃, 300 ℃ or a range between any two; the reaction time is 2-5 h, including but not limited to any one of the point values of 2h, 3h, 4h and 5h or a range value between any two of the point values; the distillation time is 4-8 h, including but not limited to any one of 4h, 5h, 6h, 7h and 8h or any range value between the two; the drying time is 8-24 h, including but not limited to any one of 8h, 9h, 10h, 12h, 15h, 18h, 20h and 24h or a range between any two.
In some embodiments, the inert atmosphere comprises a nitrogen atmosphere and/or an argon atmosphere.
In some specific embodiments, in the step (b), the collector used for the primary roughing of the iso-floatable flotation comprises butyl xanthate, and the dosage of the collector used for the primary roughing of the iso-floatable flotation in the step (b) is 30-80 g/t; including but not limited to a point value of any one of 30g/t, 40g/t, 50g/t, 60g/t, 80g/t, or a range value between any two.
In some specific embodiments, in the step (b), the collecting agent used for the first and second sweeps in the three sweeps of the floatable flotation comprises butyl xanthate (no collecting agent is added in the third sweep), and in the step (b), the collecting agent used for the first and second sweeps in the three sweeps of the floatable flotation is respectively and independently 30-80 g/t; including but not limited to a point value of any one of 30g/t, 40g/t, 50g/t, 60g/t, 80g/t, or a range value between any two.
In some specific embodiments, in the step (b), the activating agent used for the primary roughing of the iso-floatable flotation comprises copper sulfate, and in the step (b), the using amount of the activating agent used for the primary roughing of the iso-floatable flotation is 40-60 g/t; including but not limited to a point value of any one of 40g/t, 50g/t, 60g/t, or a range value between any two.
In some specific embodiments, in step (b), the floatable flotation is performed after the slurry concentration of the combined ore obtained in step (a) is adjusted to 30wt.% to 50wt.% (including but not limited to a point value of any one of 30wt.%, 35wt.%, 40wt.%, 45wt.%, 50wt.%, or a range value between any two).
In some embodiments, in step (c), the floatable concentrate obtained in step (b) is concentrated by a thickener.
In some specific embodiments, in the step (c), the pH is adjusted to 10 to 12; including but not limited to a point value of any one of 10, 10.5, 11, 11.5, 12 or a range value between any two.
In some embodiments, in step (c), quicklime is used to adjust the pH.
In some embodiments, in step (c), the inhibitors used for the one-time rougher flotation include a mass ratio of 1-3:3-5:1-4:2-6 (e.g., 1:3:1:2, 1:3:1:3, 1:3:1:4, 1:3:1:5, 1:3:1:6, 1:3:2:2, 1:3:2:3, 1:3:2:5, 1:3:2:6, 1:3:3:2, 1:3:3, 1:3:4, 1:3:3:5, 1:3:3:6, 1:3:4:2, 1:3:4:3, 1:3:4:4, 1:3:5, 1:3:4:5, 1:3:4:6, 1:4:4:2: 1:4:3, 1:4:4:4, 1:4:4:5, 1:4:4:61:5:4:2, 1:5:4:3, 1:5:4:4, 1:5:4:5, 1:5:4:6, 2:5:4:2, 2:5:4:3, 2:5:4:4, 2:5:4:5, 2:5:4:6, 3:5:4:2, 3:5:4:3, 3:5:4:4, 3:5:4:5 or 3:5:4:6) sodium carbocyante, thiourea-containing lye, polymeric ferric sulfate-containing lye and carbonate.
According to the invention, the sodium carbo-cyanite, the alkali liquor containing thiourea, the alkali liquor containing polymeric ferric sulfate and the carbonate are used as inhibitors, so that the harm of the sodium cyanide-based marmatite inhibitor to the environment can be reduced, the process adaptability is strong, the cleaning and the environmental protection are realized, meanwhile, the inhibition effect on the pyrite is good, and the lime consumption is reduced; in addition, the mixture of the sodium carbo-isocyanurate, the alkali liquor containing thiourea, the alkali liquor containing polymeric ferric sulfate and the carbonate can increase the stability of the sodium carbo-isocyanurate, and is favorable for the adsorption of cyanide on the surface of sphalerite, so that the zinc content in lead concentrate is reduced, and the cost of the medicament is reduced.
The amount of the inhibitor used in the primary roughing flotation in the step (c) (the mass sum of the sodium carbo-cyanite, the lye containing thiourea, the lye containing polymeric ferric sulfate and the carbonate) is 1000-1500 g/t, including but not limited to any one of the point values of 1000g/t, 1100g/t, 1200g/t, 1300g/t, 1400g/t, 1500g/t or a range between any two.
Wherein the molecular formula of the sodium carbo-cyanuric acid is Na 3 (CN) 3 C 3 H 3 N 6 O 3 The structural formula is。
The alkali liquor containing thiourea refers to a mixed liquor containing thiourea and alkali.
In some embodiments, the pH of the thiourea-containing lye is greater than 7.5, preferably greater than 10.
As an example, the preparation method of the thiourea-containing alkali solution includes: alkaline solution with pH=11-12 and thiourea solution with mass fraction of 1% -5% are mixed according to volume ratio=1-2: 3-4, wherein the alkaline solution comprises sodium hydroxide solution or potassium hydroxide solution, etc., but the preparation method of the thiourea-containing alkali solution is not limited thereto.
The alkali liquor containing the polymeric ferric sulfate refers to a mixed liquor containing the polymeric ferric sulfate and alkali. Wherein the base includes, but is not limited to, sodium hydroxide, potassium hydroxide, and the like.
In some embodiments, the pH of the polymeric ferric sulfate containing lye is greater than 7.5, preferably greater than 10.
In some embodiments, the carbonate salt includes, but is not limited to, sodium carbonate, potassium carbonate, and the like, for example.
In some embodiments, in step (c), the collector used in the primary rougher flotation comprises butyl xanthate; the amount of the collector used in the primary rougher flotation in the step (c) is 100-120 g/t, including but not limited to a point value of any one of 100g/t, 110g/t and 120g/t or a range value between any two.
In some embodiments, in step (c), the frother used in the primary rougher flotation comprises a # 2 oil; the amount of the foaming agent used in the primary roughing flotation in the step (c) is 30-50 g/t, including but not limited to a point value of any one of 30g/t, 35g/t, 40g/t, 45g/t and 50g/t or a range value between any two.
In some embodiments, in step (c), inhibitors used for the first of the two beneficiation floats include a mass ratio of 1-3:3-5:1-4:2-6 (e.g., 1:3:1:2, 1:3:1:3, 1:3:1:4, 1:3:1:5, 1:3:1:6, 1:3:2:2, 1:3:2:3, 1:3:2:4, 1:3:2:5, 1:3:2:6, 1:3:3:2, 1:3:3:3, 1:3:4, 1:3:3:5, 1:3:3:6, 1:3:4:2, 1:4:3:3, 1:3:4:4, 1:3:4:5, 1:3:4:6, 1:4:4:2, 1:3:4:4:4:4:5: 1:4:3, 1:4:4:4, 1:4:4:5, 1:4:4:61:5:4:2, 1:5:4:3, 1:5:4:4, 1:5:4:5, 1:5:4:6, 2:5:4:2, 2:5:4:3, 2:5:4:4, 2:5:4:5, 2:5:4:6, 3:5:4:2, 3:5:4:3, 3:5:4:4, 3:5:4:5, or 3:5:4:6) sodium carbocyante, thiourea-containing lye, polymeric ferric sulfate-containing lye, and carbonate (no inhibitor added for the second beneficiation).
The amount of the inhibitor used in the first selection in the two selection flotation in the step (c) is 150-400 g/t, including but not limited to any one of 150g/t, 200g/t, 250g/t, 300g/t, 350g/t, 400g/t or a range between any two of the point values.
In some embodiments, in step (d), the scavenger concentrate is returned to the third scavenger of the three scavenger stages above, forming a closed cycle.
In some specific embodiments, in step (d), the collectors used in the first and second of the three-pass screen include butyl xanthate (no collector added in the third pass); the amount of the collector used for the first and second scanning in the three-scanning flotation in the step (d) is 20-80 g/t, including but not limited to a point value of any one of 20g/t, 30g/t, 40g/t, 50g/t, 60g/t, 80g/t or a range value between any two.
In some specific embodiments, in step (d), the concentration of the pulp of the scavenger tail is adjusted to 30wt.% to 40wt.%; including but not limited to a point value of any one of 30wt.%, 33wt.%, 35wt.%, 38wt.%, 40wt.%, or a range value therebetween.
In some specific embodiments, in the step (d), the pH of the ore pulp of the scavenger tail is adjusted to 11-12; including but not limited to a point value of any one of 11, 11.5, 12 or a range value between any two.
In some embodiments, in step (d), the pH of the pulp of the scavenger tail is adjusted with quicklime.
In some specific embodiments, in the step (d), the inhibitor used for the primary rougher flotation includes the following components in mass ratio of 1-2: 1 polyglutamic acid and polyaspartic acid; the inhibitor used in the primary rougher flotation in the step (d) is 1600-630 g/t, including but not limited to any one point value or any range value between 1600g/t, 1650g/t, 1700g/t, 1750g/t and 1800 g/t.
According to the invention, in the primary roughing flotation in the step (d), polyglutamic acid and polyaspartic acid inhibitors are adopted according to the property characteristics of gangue minerals, so that the method is environment-friendly and has good inhibition effect.
In some embodiments, in step (d), the collector used in the primary rougher flotation comprises butyl xanthate; the amount of the collector used in the primary roughing flotation in the step (d) is 50-100 g/t; including but not limited to a point value of any one of 50g/t, 60g/t, 80g/t, 100g/t, or a range value between any two.
In some embodiments, in step (d), the beneficiated middlings are returned to a third beneficiation of the three beneficiation flotation stages above to form a closed cycle.
In some specific embodiments, in the step (d), the inhibitor used for the twice-scavenging flotation comprises the following components in percentage by mass: 1 polyglutamic acid and polyaspartic acid; the dosage of the inhibitor used in the twice scavenging flotation in the step (d) is 300-700 g/t; including but not limited to a point value of any one of 300g/t, 400g/t, 500g/t, 600g/t, 700g/t, or a range value between any two.
In some embodiments, in step (d), the collector used in the twice-swiped flotation comprises butyl xanthate; the amount of the collector used in the twice-sweeping flotation in the step (d) is 30-100 g/t, including but not limited to a point value of any one of 30g/t, 40g/t, 50g/t, 60g/t, 80g/t and 100g/t or a range value between any two of the two.
In some embodiments, in step (d), the first sulfur concentrate is returned to the second sweep of the two-sweep flotation stages above to form a closed cycle.
In some specific embodiments, in the step (e), the magnetic field strength of the magnetic separation is 5000-15000 gs; including but not limited to a point value of any one of 5000Gs, 6000Gs, 8000Gs, 10000Gs, 12000Gs, 13000Gs, 15000Gs, or a range value therebetween.
In some specific embodiments, in step (e), the magnetic separation tailings are classified by a spiral classifier to obtain an underflow and an overflow, the underflow is subjected to one roughing, two concentrating and three scavenging, and the overflow is discarded as tailings.
In some embodiments, in step (e), the collector used in the primary roughing comprises butyl xanthate; the consumption of the collecting agent used for the primary roughing in the step (e) is 30-80 g/t; including but not limited to a point value of any one of 30g/t, 40g/t, 50g/t, 60g/t, 80g/t, or a range value between any two.
In some specific embodiments, in step (e), the collectors used in the first and second of the three sweeps include butyl xanthate (no collector added in the third sweep); the amount of the collecting agent used for the first scanning and the second scanning in the step (e) is 30-80 g/t, including but not limited to any one point value or any range value between any two of 30g/t, 40g/t, 50g/t, 60g/t and 80 g/t.
In some specific embodiments, in the step (e), the activating agent used for the primary roughing comprises copper sulfate, and the using amount of the activating agent used for the primary roughing in the step (e) is 40-60 g/t; including but not limited to a point value of any one of 40g/t, 50g/t, 60g/t, or a range value between any two.
In some specific embodiments, in the step (e), the stroke of the first shaking table reselection is 16-22 mm, including but not limited to a point value of any one of 16mm, 18mm, 20mm, 22mm or a range value between any two; in the step (e), the first shaking table reselects the stroke frequency 200-260 times/min, including but not limited to any one of 200 times/min, 220 times/min, 240 times/min, 250 times/min, 260 times/min or any range between the two.
In some embodiments, in step (e), the first shaking table reselection is performed in a fine mud shaking table.
In some embodiments, in step (f), the collector used in the primary roughing comprises butyl xanthate; the amount of the collecting agent used in the primary roughing in the step (f) is 30-50 g/t, including but not limited to any one point value or any range value between the two points of 30g/t, 40g/t and 50 g/t.
In some specific embodiments, in step (f), the collectors used in the first and second of the three sweeps include butyl xanthate (no collector added in the third sweep); the amount of the collecting agent used for the first scanning and the second scanning in the step (f) is 30-50 g/t, including but not limited to any one point value or any range value between any two of 30g/t, 40g/t and 50 g/t.
In some embodiments, in step (f), the inhibitor used in the one-time roughing comprises sodium thioglycolate; the inhibitor used in the first roughing in the step (f) is 2000-3000 g/t, including but not limited to any one point value or any range value between 2000g/t, 2300g/t, 2500g/t, 2800g/t and 3000 g/t.
Wherein, the molecular formula of the sodium thioglycolate is:。
the invention adopts sodium thioglycolate as inhibitor, which can completely replace the virulent inhibitor sodium cyanide, thus ensuring the production index of zinc concentrate grade and recovery rate and meeting the environmental protection requirement.
In some embodiments, in step (f), the inhibitor used in the first and second of the two refinements comprises sodium thioglycolate; the inhibitor used in the first and second beneficiations of step (f) is used in an amount of 2000-3000 g/t, including but not limited to a point value of any one of 2000g/t, 2300g/t, 2500g/t, 2800g/t, 3000g/t, or a range value between any two.
In some embodiments, in step (g), the overflow is concentrated by a thickener.
In some specific embodiments, in step (g), the concentration of the grit is adjusted to 30wt.% to 40wt.%; including but not limited to a point value of any one of 30wt.%, 33wt.%, 35wt.%, 38wt.%, 40wt.%, or a range value therebetween.
In some specific embodiments, in the step (g), the pH of the settled sand is adjusted to 6-7; including but not limited to a point value of any one of 6, 6.3, 6.5, 6.7, 7 or a range value between any two.
In some embodiments, in step (g), the collector used in the primary roughing comprises butyl xanthate; the consumption of the collecting agent used for the primary roughing in the step (g) is 30-120 g/t; including but not limited to a point value of any one of 30g/t, 50g/t, 80g/t, 100g/t, 120g/t, or a range value between any two.
In some specific embodiments, in step (g), the collectors used in the first and second of the three sweeps include butyl xanthate (no collector added in the third sweep); the amount of the collector used in the first and second scans in the step (g) is 30-120 g/t, including but not limited to any one of 30g/t, 50g/t, 80g/t, 100g/t, 120g/t or a range between any two.
In some embodiments, in step (g), the activator used in the primary roughing comprises copper sulfate; the dosage of the activating agent used in the primary roughing in the step (g) is 65-200 g/t, including but not limited to a point value of any one of 65g/t, 80g/t, 100g/t, 120g/t, 150g/t, 180g/t and 200g/t or a range value between any two of the two.
In some specific embodiments, in step (h), the concentration of the first desulphurized flotation tailings is adjusted to 25wt.% to 35wt.%; including but not limited to a point value of any one of 25wt.%, 28wt.%, 30wt.%, 33wt.%, 35wt.%, or a range value therebetween.
In some specific embodiments, in the step (h), adjusting the pH of the first desulphurized flotation tailings to 6-7; including but not limited to a point value of any one of 6, 6.3, 6.5, 6.8, 7 or a range value between any two.
In some embodiments, in step (h), the collector used in the first roughing comprises cinnamic hydroxamic acid and a P86 collector, wherein the cinnamic hydroxamic acid has the formula C 6 H 5 -ch=ch- (c=o) NHOH, the amount of cinnamyl hydroxamic acid being 40-220 g/t, including but not limited to a point value of any one of 40g/t, 50g/t, 80g/t, 100g/t, 130g/t, 150g/t, 180g/t, 200g/t, 220g/t, or a range value therebetween; the dosage of the P86 collector is 60-200 g/t, including but not limited to any one point value or range value between any two of 60g/t, 80g/t, 100g/t, 150g/t and 200 g/t.
In some embodiments, in step (h), the collectors used in the first and second of the three sweeps include cinnamic hydroxamic acid having formula C and a P86 collector (no collector added in the third sweep) 6 H 5 -ch=ch- (c=o) NHOH, the amount of cinnamyl hydroxamic acid being 40-220 g/t, including but not limited to a point value of any one of 40g/t, 50g/t, 80g/t, 100g/t, 130g/t, 150g/t, 180g/t, 200g/t, 220g/t, or a range value therebetween; the dosage of the P86 collector is 60-200 g/t, including but not limited to any one point value or range value between any two of 60g/t, 80g/t, 100g/t, 150g/t and 200 g/t.
In some specific embodiments, in step (h), the method of preparing cinnamon hydroxamic acid comprises: and (3) mixing the styryl, the N-hydroxyformamide and the nitric acid in an inert atmosphere, heating to react, and sequentially distilling and drying after the reaction is completed to obtain the cinnamon hydroxamic acid.
Wherein the heating temperature is 200-300 ℃, including but not limited to any one of 200 ℃, 220 ℃, 250 ℃, 280 ℃ and 300 ℃ or a range value between any two; the reaction time is 2-5 h, including but not limited to any one of the point values of 2h, 3h, 4h and 5h or a range value between any two of the point values; the distillation time is 4-8 h, including but not limited to any one of 4h, 5h, 6h, 7h and 8h or any range value between the two; the drying time is 8-24 h, including but not limited to any one of 8h, 9h, 10h, 12h, 15h, 18h, 20h and 24h or a range between any two. The inert atmosphere comprises a nitrogen atmosphere and/or an argon atmosphere. The molar ratio of the styryl, the N-hydroxyformamide and the nitric acid is 1:2-4:1-3.
In some embodiments, in step (h), the inhibitor used in the one-time roughing comprises a sodium silicate-containing acid solution; the dosage of the inhibitor for the primary roughing in the step (h) is 600-1800 g/t; including but not limited to a point value of any one of 600g/t, 800g/t, 1000g/t, 1500g/t, 1800g/t, or a range value between any two.
In some embodiments, in step (h), the inhibitor used in the first of the two beneficiations comprises an acid solution comprising sodium silicate (the second beneficiation does not add an inhibitor). Wherein the acid solution containing sodium silicate is a mixed solution containing sodium silicate and acid, and the mixed solution is acidic.
According to the invention, the acid solution containing sodium silicate is used as an inhibitor, compared with water glass, the acidified sodium silicate has stronger inhibition effect on gangue minerals at a lower dosage, and the quality of the minerals is improved mainly by further inhibiting silicate and calcium, so that the influence of the gangue minerals on the quality of concentrate is reduced; at the same time, the aggregation of bubbles is reduced, entrainment is reduced, grade is improved, and the effect is irreversible.
As an example, the method for preparing the sodium silicate-containing acid solution includes: mixing an acid solution with a sodium silicate solution for acidification; wherein the acid solution comprises at least one of oxalic acid solution, sulfuric acid solution and hydrochloric acid solution; the mass fraction of the sodium silicate solution is 30% -70%; the molar concentration of the acid solution is 0.5-5 mol/l; the volume ratio of the acid solution to the sodium silicate solution is 1-5:1-5; the mixing time is 30-60 min. The method for preparing the sodium silicate-containing acid solution is not limited thereto.
The dosage of the inhibitor used for the first selection in the two selections in the step (h) is 600-1800 g/t; including but not limited to a point value of any one of 600g/t, 800g/t, 1000g/t, 1500g/t, 1800g/t, or a range value between any two.
In some embodiments, in step (h), the activator used in the primary roughing comprises lead nitrate; the dosage of the activating agent used for the primary roughing in the step (h) is 50-80 g/t; including but not limited to a point value of any one of 50g/t, 60g/t, 70g/t, 80g/t, or a range value therebetween.
In some specific embodiments, in the step (h), the magnetic field strength of the magnetic separation is 4500-8000 gs; including but not limited to a point value of any one of 4500Gs, 5000Gs, 6000Gs, 7000Gs, 8000Gs, or a range value therebetween.
In some specific embodiments, in the step (h), the stroke of the second shaking table reselection is 5-20 mm, including but not limited to a point value of any one of 5mm, 8mm, 10mm, 13mm, 15mm, 18mm, 20mm or a range value between any two; the impulse is 260-340 times/min, including but not limited to any one point value or range value between any two of 260 times/min, 280 times/min, 300 times/min, 320 times/min and 340 times/min.
In some embodiments, in step (h), the second shaking table reselection is performed in a fine mud shaking table.
In some specific embodiments, in the step (i), the stroke of the third shaking table reselection is 11-16 mm, including but not limited to any one of 11mm, 12mm, 13mm, 15mm, 16mm or a range between any two; the impulse is 200-300 r/min, including but not limited to any one point value or any range value between two points of 200 times/min, 230 times/min, 250 times/min, 260 times/min, 280 times/min and 300 times/min.
In some embodiments, in step (i), the third shaker reselection is performed in a fine sand shaker.
In some specific embodiments, in step (i), the concentration of the mineral passing through the five stages of grinding is adjusted to 30wt.% to 40wt.%; including but not limited to a point value of any one of 30wt.%, 33wt.%, 35wt.%, 38wt.%, 40wt.%, or a range value therebetween.
In some specific embodiments, in the step (i), the pH of the mineral subjected to the five-stage grinding is adjusted to 6-7; including but not limited to a point value of any one of 6, 6.3, 6.5, 6.8, 7 or a range value between any two.
In some specific embodiments, in the step (i), the collector used in the primary roughing comprises butyl xanthate, and the dosage of the collector is 30-120 g/t; including but not limited to a point value of any one of 30g/t, 40g/t, 50g/t, 80g/t, 100g/t, 120g/t, or a range value between any two.
In some specific embodiments, in step (i), the collectors used in the first and second of the three sweeps include butyl xanthate (no collector added in the third sweep); the amount of the collector used in the first and second scans in the step (i) is 30-120 g/t, including but not limited to any one of 30g/t, 40g/t, 50g/t, 80g/t, 100g/t, 120g/t, or a range between any two.
In some embodiments, in step (i), the activator used in the primary roughing comprises copper sulfate; the dosage of the activating agent used in the primary roughing in the step (i) is 100-200 g/t, including but not limited to any one point value or any range value between the two points of 100g/t, 120g/t, 150g/t, 180g/t and 200 g/t.
In some specific embodiments, in the step (i), the stroke of the fourth shaking table reselection is 10-25 mm, including but not limited to a point value of any one of 10mm, 13mm, 15mm, 18mm, 20mm, 22mm, 25mm or a range value between any two; the fourth shaking table reselects the stroke frequency to be 200-400 times/min, including but not limited to any one point value or any range value between two points of 200 times/min, 250 times/min, 300 times/min, 350 times/min and 400 times/min.
In some embodiments, in step (i), the fourth shaking table reselection is performed in a middling shaking table.
In some specific embodiments, in step (i), the ratio of the material particle size obtained after five stages of grinding to less than 200 mesh is 70wt.% to 80wt.%.
By adopting the beneficiation reagent and the beneficiation parameters, higher yield can be obtained, and the grade of valuable metal elements can be improved.
The unit of the dosage of each reagent "g/t" refers to the mass of the reagent (medicament) added into each ton of mineral or ore pulp is n grams. For example, 100g/t refers to a mass of 100 grams of reagent added per ton of pulp.
The first-stage grinding, the second-stage grinding, the third-stage grinding, the fourth-stage grinding and the fifth-stage grinding refer to the first grinding, the second grinding, the third grinding, the fourth grinding and the fifth grinding respectively.
It will be appreciated that the tailings produced in each of the beneficiation steps described above are not addressed to disposal.
In a second aspect, the invention provides an application of the complex tin-lead-zinc multi-metal ore separation and recovery method in ore dressing.
Embodiments of the present invention will be described in detail below with reference to examples, but it will be understood by those skilled in the art that the following examples are only for illustrating the present invention and should not be construed as limiting the scope of the present invention. The specific conditions are not noted in the examples and are carried out according to conventional conditions or conditions recommended by the manufacturer. The reagents or apparatus used were conventional products commercially available without the manufacturer's attention.
Example 1
The complex tin-lead-zinc polymetallic ore (abbreviated as raw ore) provided in this embodiment is selected from tin polymetallic sulfide ore in certain places in guangxi, wherein main metal minerals are cassiterite, jamesonite, marmatite, pyrite and arsenopyrite, gangue minerals are mainly quartz and calcite, wherein the tin element content is 0.23wt.% (wt.%), the lead element content is 0.18wt.%, the antimony element content is 0.13wt.%, the zinc element content is 1.46wt.%, the sulfur element content is 5.74wt.%, and the arsenic element content is 1.25wt.%, and the separation and recovery method of the complex tin-lead-zinc polymetallic ore specifically comprises the following steps:
(1) Grading and jigging waste disposal: screening raw ore into three particle size grades, carrying out primary grinding on the ore with the particle size larger than 4mm (namely high-sized ore), returning to an upper screening process, and carrying out jigging and beneficiation on the ore with the particle size of 2-4 mm (namely medium-sized ore) to obtain jigged rough concentrate and jigged middling (jigged tailings are directly thrown waste); the jigged rough concentrate is subjected to secondary grinding (the particle size of the ore after secondary grinding is less than 200 meshes is 45 wt.%) and then is mixed with the ore with the particle size of less than 2mm (namely, the low-size ore) and is subjected to reselection through a conical spiral chute system, so that front heavy concentrate (underflow) and overflow are obtained; the front heavy concentrate enters a deck floating table to carry out deck floating table flotation, the stroke is 12mm, the stroke frequency is 300 times/min, and the collector used for deck floating table flotation is cinnamon hydroxamic acid (the molecular formula is C 6 H 5 -ch=ch- (c=o) NHOH in an amount of 300g/t to obtain a first tin concentrate and a deck tailings; mixing the table floating tailings with the jigged middlings, and carrying out three-stage grinding to obtain combined ores.
The preparation method of the cinnamon hydroxamic acid comprises the following steps: under inert atmosphere, mixing styryl, N-hydroxyformamide and nitric acid in a molar ratio of 1:2:1, heating to react, and sequentially distilling and drying after the reaction is completed to obtain cinnamyl hydroxamic acid; wherein the heating temperature is 250 ℃, the reaction time is 4 hours, the distillation time is 6 hours, and the drying time is 12 hours.
(2) And (3) adjusting the concentration of the ore pulp of the combined ore obtained in the step (1) to 40wt.%, and then performing primary roughing, secondary concentration and tertiary scavenging and the like to obtain the equifloatable concentrate and the equifloatable tailings. Wherein, the collecting agent used for one roughing is butyl xanthate, and the dosage of the collecting agent is 50g/t (namely, the mass of the butyl xanthate added into each ton of ore pulp is 50 g); the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 50g/t; the collecting agents used in the first and second times of the three times of scavenging are butyl xanthate, and the dosage of each time of scavenging is 50g/t.
(3) Separating and floating lead-antimony and zinc-sulfur: concentrating the equi-floatable concentrate obtained in the step (2) by a thickener, performing four-stage ore grinding (the particle ratio of the ore with the granularity smaller than 200 meshes after the four-stage ore grinding is 70 wt.%), adjusting the pH value to be 11 by quicklime, performing primary roughing flotation, and obtaining roughing concentrate and roughing tailings. The inhibitor for primary roughing flotation comprises sodium carbo-cyanuric acid, alkali liquor containing thiourea, alkali liquor containing polymeric ferric sulfate and carbonate in a mass ratio of (2:4:3:4), wherein the dosage of the inhibitor is 1000g/t, the dosage of the collector butyl xanthate for primary roughing flotation is 100g/t, and the dosage of the foamer 2# oil for primary roughing flotation is 35g/t.
Wherein the alkali liquor containing thiourea is a mixed liquor containing thiourea and sodium hydroxide, and the pH value of the alkali liquor is 11.5. The alkali liquor containing the polymeric ferric sulfate is a mixed liquor containing the polymeric ferric sulfate and sodium hydroxide, and the pH value of the alkali liquor is 11.5. The carbonate is sodium carbonate.
Lead concentrate flotation: and (3) carrying out twice concentration flotation on the roughing concentrate to obtain the final lead-antimony concentrate. Wherein the inhibitor used for the first concentration in the two concentration flotation is the same as the inhibitor used for the first roughing flotation in the step (3), and the dosage of the inhibitor is 350g/t.
(4) Zinc-sulfur separation: and (3) carrying out three scavenging flotation on the rougher tailings obtained in the step (3) to obtain scavenging concentrate and scavenging tailings. Wherein, the scavenging concentrate returns to the third scavenging in the three scavenging of the upper stage to form a closed cycle; the collecting agents used in the first and second times of the three times of the scavenging flotation are butyl xanthate, and the amount of each time of scavenging is 50g/t.
And then adjusting the concentration of the ore pulp of the scavenging tailings to 30wt.%, and adjusting the pH value of the ore pulp to be 12 by using quicklime, and then performing roughing flotation for one time to obtain rough zinc concentrate and rough sulfur tailings. Wherein, the inhibitor used for primary roughing flotation is mass ratio=2: 1, wherein the dosage of the inhibitor is 1800g/t; the collector used for primary roughing flotation is butyl xanthate, and the dosage of the collector is 60g/t.
And (3) carrying out three-time concentration flotation on the coarse zinc concentrate to obtain final zinc concentrate and concentrated middlings. Wherein, the middling is selected for the third time in the upper three-time selecting floatation to form a closed cycle.
And (3) carrying out twice scavenging flotation on the coarse sulfur tailings to obtain first sulfur concentrate and zinc sulfur tailings. Wherein, the inhibitor used for the twice scavenging flotation is that the mass ratio=2: 1, the dosage of the inhibitor is 500g/t, the collector used for the twice scavenging flotation is butyl xanthate, and the dosage of the butyl xanthate is 40g/t. And returning the first sulfur concentrate to the second scavenging in the upper-level twice scavenging flotation to form closed cycle.
(5) Sulfur-arsenic mixing and floating: and (3) carrying out magnetic separation on the zinc-sulfur tailings obtained in the step (4), wherein the magnetic field strength is 10000Gs, and removing magnetic minerals to obtain magnetic tailings and magnetic concentrate. And grading the magnetic separation tailings by a spiral grader to obtain underflow and overflow, discarding the overflow as tailings, and carrying out primary roughing, secondary concentration and tertiary scavenging on the underflow to obtain sulfur-arsenic mixed flotation concentrate and mixed flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 50g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 30g/t; the collecting agents used in the first and second times of the three times of scavenging are butyl xanthate, and the dosage of each time of scavenging is 50g/t.
And (3) carrying out first shaking table gravity separation on the mixed floating tailings in a fine mud shaking table, wherein the stroke is 20mm, the stroke frequency is 230 times/min, the first shaking table gravity separation concentrate is low-grade tin gravity separation concentrate, and the first shaking table gravity separation tailings are classified as final tailings.
(6) Sulfur-arsenic separation: and (3) carrying out primary roughing, secondary concentration and tertiary scavenging on the sulfur-arsenic mixed concentrate obtained in the step (5) to obtain second sulfur concentrate. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 40g/t; the inhibitor used for one-time roughing is sodium thioglycolate, and the dosage of the inhibitor is 2500g/t; the inhibitor used in the first concentration and the second concentration in the two concentrations is sodium thioglycolate, and the dosage of each concentration is 2500g/t; the collecting agents used in the first and second times of the three times of scavenging are butyl xanthate, and the dosage of each time of scavenging is 40g/t.
Mixing the second sulfur concentrate with the magnetic concentrate obtained in the step (5) to obtain a final sulfur concentrate product.
(7) Desulfurizing and floatation: concentrating the overflow obtained in the step (1) by a thickener to obtain settled sand, adjusting the concentration of the settled sand to 30wt.% and the pH=6, and then performing desulfurization flotation of one roughing, two concentration and three scavenging to obtain third sulfur concentrate and first desulfurization flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 50g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 100g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage of the collecting agents is 50g/t. And (3) returning the third sulfur concentrate to the step (3) for mixing with the floatable concentrate (continuing the concentration and four-stage grinding process).
(8) Floating tin and collecting tin: and (3) adjusting the concentration of the first desulfurization flotation tailings obtained in the step (7) to 25wt.% and the pH=6, and then performing one roughing, two concentration and three scavenging to obtain the tin concentrate. Wherein the collectors used for one roughing are cinnamon hydroxamic acid (the preparation method is the same as the step (1) of the embodiment) and P86 collector, the dosage of the cinnamon hydroxamic acid is 180g/t, and the dosage of the P86 collector is 130g/t; the inhibitor used for one roughing is oxalic acid solution containing sodium silicate (prepared by mixing oxalic acid solution with the molar concentration of 1mol/l and sodium silicate solution with the mass fraction of 50% in a volume ratio of = 1:1 for 50 min), and the dosage of the inhibitor is 1400g/t; the activator used in one roughing is lead nitrate, and the dosage of the activator is 50g/t. The inhibitor used in the first of the two beneficiations was an oxalic acid solution containing sodium silicate (preparation method as above) in an amount of 1400g/t. The collectors used in the first and second of the three scans were cinnamon hydroxamic acid (the preparation method is the same as step (1) of this example) and P86 collector, and the amount of cinnamon hydroxamic acid used in each scan was 180g/t, and the amount of P86 collector used in each scan was 130g/t.
And (3) carrying out magnetic separation on the float tin concentrate to remove magnetic minerals, wherein the magnetic field strength is 8000Gs, and obtaining the magnetic separation tailings. And (3) carrying out second shaking table reselection on the magnetic separation tailings in a fine mud shaking table, wherein the stroke is 12mm, and the stroke frequency is 300 times/min, so as to obtain fine mud tin concentrate.
(9) Tin collection: and (3) carrying out third shaking table reselection on the iso-floatable tailings obtained in the step (2) in a fine sand shaking table, wherein the stroke is 14mm, and the flushing frequency is 250r/min, so as to obtain second tin concentrate and shaking table tailings. Mixing the second tin concentrate with the first tin concentrate obtained in the step (1) to obtain a final tin concentrate product.
And (3) grinding the table tailings in five stages (the particle ratio of less than 200 meshes is 70 wt.%), adjusting the concentration to 30wt.% and the pH=6, and then performing desulfurization flotation of one roughing, two concentration and three scavenging to obtain fourth sulfur concentrate and second desulfurization flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 50g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 100g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage of the collecting agents is 50g/t. And (3) returning the fourth sulfur concentrate to the step (3) for mixing with the floatable concentrate (continuing the concentration and four-stage grinding process).
And (3) carrying out fourth shaking table gravity separation desulfurization on the second desulfurization flotation tailings in a middling shaking table, wherein the stroke is 18mm, and the stroke frequency is 300 times/min, so as to obtain moderate tin concentrate.
Through the steps, in the lead-antimony concentrate obtained in the embodiment, the lead grade (namely the mass fraction of lead element) is 24.74%, the lead element recovery rate is 83.56%, the antimony grade (namely the mass fraction of antimony element) is 19.03%, and the antimony element recovery rate is 91.37%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 48.28 percent, and the recovery rate of the zinc element is 89.20 percent; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 44.40%, and the recovery rate of tin element is 78.81%.
Example 2
The complex tin-lead-zinc polymetallic ore (abbreviated as raw ore) provided in this embodiment is selected from tin polymetallic sulfide ore in certain places of guangxi university, wherein main metal minerals are cassiterite, jamesonite, marmatite, pyrite and arsenopyrite, gangue minerals are mainly quartz and calcite, wherein the tin element content is 0.33 wt%, the lead element content is 0.20 wt%, the antimony element content is 0.14 wt%, the zinc element content is 1.63 wt%, the sulfur element content is 6.85 wt%, and the arsenic element content is 1.56 wt%, and the separation and recovery method of the complex tin-lead-zinc polymetallic ore specifically comprises the following steps:
(1) Grading and jigging waste disposal: screening raw ore into three particle size grades, carrying out primary grinding on the ore with the particle size larger than 4mm (namely high-sized ore), returning to an upper screening process, and carrying out jigging and beneficiation on the ore with the particle size of 2-4 mm (namely medium-sized ore) to obtain jigged rough concentrate and jigged middling (jigged tailings are directly thrown waste); the jigged rough concentrate is subjected to secondary grinding (the particle size of the ore after secondary grinding is less than 200 meshes is 50 wt.%) and then is mixed with the ore with the particle size of less than 2mm (namely, the low-size ore) and is subjected to reselection through a conical spiral chute system, so that front heavy concentrate (underflow) and overflow are obtained; the front heavy concentrate enters a deck floating table to carry out deck floating table flotation, the stroke is 8mm, the stroke frequency is 250 times/min, the collector used for deck floating table flotation is cinnamon hydroxamic acid, the dosage of the collector is 200g/t, and first tin concentrate and deck floating tailings are obtained; mixing the table floating tailings with the jigged middlings, and carrying out three-stage grinding to obtain combined ores.
The preparation method of the cinnamon hydroxamic acid comprises the following steps: under inert atmosphere, mixing styryl, N-hydroxyformamide and nitric acid in a molar ratio of 1:2:1, heating to react, and sequentially distilling and drying after the reaction is completed to obtain cinnamyl hydroxamic acid; wherein the heating temperature is 200 ℃, the reaction time is 5 hours, the distillation time is 4 hours, and the drying time is 8 hours.
(2) And (3) adjusting the concentration of the ore pulp of the combined ore obtained in the step (1) to 35wt.%, and then performing primary roughing, secondary concentration and tertiary scavenging and the like to obtain the equifloatable concentrate and the equifloatable tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 40g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 40g/t; the collecting agents used in the first and second scanning in the three scanning are butyl xanthate, and the dosage is 40g/t.
(3) Separating and floating lead-antimony and zinc-sulfur: concentrating the equi-floatable concentrate obtained in the step (2) by a thickener, performing four-stage grinding (the particle ratio of the mineral with the granularity smaller than 200 meshes is 75 wt.%) and then adjusting the pH value to be 10.5 by quicklime, and performing primary roughing flotation to obtain roughing concentrate and roughing tailings. The inhibitor for primary roughing flotation is sodium carbo-cyanuric acid, alkali liquor containing thiourea, alkali liquor containing polymeric ferric sulfate and carbonate in a mass ratio of (2:3:3:4), the dosage of the inhibitor is 1200g/t, the dosage of the collector butyl xanthate for primary roughing flotation is 110g/t, and the dosage of the foamer 2# oil for primary roughing flotation is 30g/t.
Wherein the alkali solution containing thiourea is a mixed solution containing thiourea and potassium hydroxide, and the pH value of the alkali solution is 12. The alkali liquor containing the polymeric ferric sulfate is a mixed liquor containing the polymeric ferric sulfate and potassium hydroxide, and the pH value of the alkali liquor is 12. The carbonate is potassium carbonate.
Lead concentrate flotation: and (3) carrying out twice concentration flotation on the roughing concentrate to obtain the final lead-antimony concentrate. Wherein the inhibitor used for the first concentration in the two concentration flotation is the same as the inhibitor used for the first roughing flotation in the step (3), and the dosage of the inhibitor is 300g/t.
(4) Zinc-sulfur separation: and (3) carrying out three scavenging flotation on the rougher tailings obtained in the step (3) to obtain scavenging concentrate and scavenging tailings. Wherein, the scavenging concentrate returns to the third scavenging in the three scavenging of the upper stage to form a closed cycle; the collecting agents used in the first and second scavenging in the third scavenging floatation are butyl xanthate, and the dosage of the collecting agents is 30g/t.
And then adjusting the concentration of the ore pulp of the scavenging tailings to 35wt.%, adjusting the pH=11.5 of the ore pulp by using quicklime, and performing primary roughing flotation to obtain rough zinc concentrate and rough sulfur tailings. Wherein, the inhibitor used for primary roughing flotation is mass ratio=1: 1, wherein the dosage of the inhibitor is 1600g/t; the collector used for primary roughing flotation is butyl xanthate, and the dosage of the collector is 50g/t.
And (3) carrying out three-time concentration flotation on the coarse zinc concentrate to obtain final zinc concentrate and concentrated middlings. Wherein, the middling is selected for the third time in the upper three-time selecting floatation to form a closed cycle.
And (3) carrying out twice scavenging flotation on the coarse sulfur tailings to obtain first sulfur concentrate and zinc sulfur tailings. Wherein, the inhibitor used for the twice scavenging flotation is that the mass ratio=1: 1, the dosage of the inhibitor is 400g/t, the collector used for twice scavenging flotation is butyl xanthate, and the dosage of the butyl xanthate is 50g/t. And returning the first sulfur concentrate to the second scavenging in the upper-level twice scavenging flotation to form closed cycle.
(5) Sulfur-arsenic mixing and floating: and (3) carrying out magnetic separation on the zinc-sulfur tailings obtained in the step (4), wherein the magnetic field strength is 8000Gs, and removing magnetic minerals to obtain magnetic tailings and magnetic concentrate. And grading the magnetic separation tailings by a spiral grader to obtain underflow and overflow, discarding the overflow as tailings, and carrying out primary roughing, secondary concentration and tertiary scavenging on the underflow to obtain sulfur-arsenic mixed flotation concentrate and mixed flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 40g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 40g/t; the collecting agents used in the first and second scanning in the three scanning are butyl xanthate, and the dosage is 40g/t.
And (3) carrying out first shaking table gravity separation on the mixed floating tailings in a fine mud shaking table, wherein the stroke is 18mm, the stroke frequency is 200 times/min, the first shaking table gravity separation concentrate is the low-grade tin gravity separation concentrate, and the first shaking table gravity separation tailings are classified as final tailings.
(6) Sulfur-arsenic separation: and (3) carrying out primary roughing, secondary concentration and tertiary scavenging on the sulfur-arsenic mixed concentrate obtained in the step (5) to obtain second sulfur concentrate. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 30g/t; the inhibitor used for one roughing is sodium thioglycolate, and the dosage of the inhibitor is 2200g/t; the inhibitors used in the first selection and the second selection in the two selections are sodium thioglycolate, and the dosage of the inhibitors is 2100g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage of the collecting agents is 30g/t.
Mixing the second sulfur concentrate with the magnetic concentrate obtained in the step (5) to obtain a final sulfur concentrate product.
(7) Desulfurizing and floatation: concentrating the overflow obtained in the step (1) by a thickener to obtain settled sand, adjusting the concentration of the settled sand to 35wt.% and the pH=7, and then performing desulfurization flotation of one roughing, two concentration and three scavenging to obtain third sulfur concentrate and first desulfurization flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 30g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 65g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage of the collecting agents is 30g/t. And (3) returning the third sulfur concentrate to the step (3) for mixing with the floatable concentrate (continuing the concentration and four-stage grinding process).
(8) Floating tin and collecting tin: and (3) adjusting the concentration of the first desulfurization flotation tailings obtained in the step (7) to 30wt.% and the pH=7, and then performing one roughing, two concentration and three scavenging to obtain the tin concentrate. Wherein the collectors used for one roughing are cinnamon hydroxamic acid (the preparation method is the same as the step (1) of the embodiment) and a P86 collector, the dosage of the cinnamon hydroxamic acid is 100g/t, and the dosage of the P86 collector is 60g/t; the inhibitor used for one roughing is a sulfuric acid solution containing sodium silicate (prepared by mixing a sulfuric acid solution with the molar concentration of 2mol/l and a sodium silicate solution with the mass fraction of 60% in a volume ratio of = 1:2 for 30 min), and the dosage of the inhibitor is 1500g/t; the activator used in one roughing is lead nitrate, and the dosage of the activator is 70g/t. The inhibitor used in the first of the two beneficiations was a sodium silicate-containing sulfuric acid solution (preparation method as above) in an amount of 1500g/t. The collectors used in the first and second of the three scans were cinnamic hydroxamic acid (the preparation method is the same as step (1) of this example) and P86 collector, the amounts of cinnamic hydroxamic acid and P86 collector were 200g/t and 180g/t, respectively.
And (3) carrying out magnetic separation on the tin concentrate, removing magnetic minerals, and obtaining the magnetic separation tailings, wherein the magnetic field strength is 6000 Gs. And (3) carrying out second shaking table reselection on the magnetic separation tailings in a fine mud shaking table, wherein the stroke is 18mm, and the stroke frequency is 330 times/min, so as to obtain fine mud tin concentrate.
(9) Tin collection: and (3) carrying out third shaking table reselection on the iso-floatable tailings obtained in the step (2) in a fine sand shaking table, wherein the stroke is 16mm, and the stroke frequency is 300r/min, so as to obtain second tin concentrate and shaking table tailings. Mixing the second tin concentrate with the first tin concentrate obtained in the step (1) to obtain a final tin concentrate product.
And (3) grinding the table tailings in five stages (the particle ratio of less than 200 meshes is 75 wt.%), adjusting the concentration to 35wt.% and the pH=7, and then performing desulfurization flotation of one roughing, two concentration and three scavenging to obtain fourth sulfur concentrate and second desulfurization flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 40g/t; the activating agent used for one roughing is copper sulfate, and the dosage of the activating agent is 180g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage of the collecting agents is 40g/t. And (3) returning the fourth sulfur concentrate to the step (3) for mixing with the floatable concentrate (continuing the concentration and four-stage grinding process).
And (3) carrying out fourth shaking table gravity separation desulfurization on the second desulfurization flotation tailings in a middling shaking table, wherein the stroke is 25mm, and the stroke frequency is 350 times/min, so as to obtain moderate tin concentrate.
Through the steps, the lead grade (namely the mass fraction of lead element) in the lead-antimony concentrate obtained by the embodiment is 18.12%, the lead element recovery rate is 22.98%, the antimony grade (namely the mass fraction of antimony element) is 14.99%, and the antimony element recovery rate is 95.89%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 36.97 percent, and the recovery rate of the zinc element is 88.95 percent; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 41.56%, and the recovery rate of tin element is 75.61%.
Example 3
The complex tin-lead-zinc polymetallic ore (abbreviated as raw ore) provided in this embodiment is selected from tin polymetallic sulfide ore in a certain place in the inner mongolic tin huge market, wherein the metal minerals are cassiterite, galena, marmatite, pyrite and silver, the gangue minerals are mainly quartz and calcite, wherein the tin element content is 0.24wt.% (wt.%) and the lead element content is 0.17wt.%, the silver element content is 20g/t, the zinc element content is 1.41wt.%, the sulfur element content is 5.53wt.%, and the arsenic element content is 1.02%, and the separation and recovery method of the complex tin-lead-zinc polymetallic ore specifically comprises the following steps:
(1) Grading and jigging waste disposal: screening raw ore into three particle size grades, carrying out primary grinding on the ore with the particle size larger than 4mm (namely high-sized ore), returning to an upper screening process, and carrying out jigging and beneficiation on the ore with the particle size of 2-4 mm (namely medium-sized ore) to obtain jigged rough concentrate and jigged middling (jigged tailings are directly thrown waste); the jigged rough concentrate is subjected to secondary grinding (the particle size of the ore after secondary grinding is less than 200 meshes is 60 wt.%) and then is mixed with the ore with the particle size of less than 2mm (namely, the low-size ore) and is subjected to reselection through a conical spiral chute system, so that front heavy concentrate (underflow) and overflow are obtained; the front heavy concentrate enters a deck floating table to carry out deck floating table flotation, the stroke is 15mm, the stroke frequency is 330 times/min, the collector used for deck floating table flotation is cinnamon hydroxamic acid, the dosage of the collector is 350g/t, and first tin concentrate and deck floating tailings are obtained; mixing the table floating tailings with the jigged middlings, and carrying out three-stage grinding to obtain combined ores.
The preparation method of the cinnamon hydroxamic acid comprises the following steps: under inert atmosphere, mixing styryl, N-hydroxyformamide and nitric acid in a molar ratio of 1:2:1, heating to react, and sequentially distilling and drying after the reaction is completed to obtain cinnamyl hydroxamic acid; wherein the heating temperature is 300 ℃, the reaction time is 3 hours, the distillation time is 8 hours, and the drying time is 20 hours.
(2) And (3) adjusting the concentration of the ore pulp of the combined ore obtained in the step (1) to 45wt.%, and then performing primary roughing, secondary concentration and tertiary scavenging and the like to obtain the equifloatable concentrate and the equifloatable tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 60g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 60g/t; the collecting agents used in the first and second times of three times of scavenging are butyl xanthate, and the dosage is 60g/t.
(3) Separating and floating lead, silver and zinc and sulfur: concentrating the equi-floatable concentrate obtained in the step (2) by a thickener, performing four-stage grinding (the particle ratio of the mineral with the granularity smaller than 200 meshes is 80 wt.%) and then adjusting the pH value to be 10 by quicklime, and performing primary roughing flotation to obtain roughing concentrate and roughing tailings. The inhibitor for primary roughing flotation is sodium carbo-cyanuric acid, alkali liquor containing thiourea, alkali liquor containing polymeric ferric sulfate and carbonate in a mass ratio of = 2:4:1:6, the dosage of the inhibitor is 1400g/t, the dosage of the collector butyl xanthate for primary roughing flotation is 110g/t, and the dosage of the foamer 2# oil for primary roughing flotation is 50g/t.
Wherein the alkali liquor containing thiourea is mixed liquor containing thiourea and sodium hydroxide, and the pH value of the alkali liquor is 11. The alkali liquor containing the polymeric ferric sulfate is a mixed liquor containing the polymeric ferric sulfate and sodium hydroxide, and the pH value of the alkali liquor is 11. The carbonate is ammonium carbonate.
Lead concentrate flotation: and (3) carrying out twice concentration flotation on the roughing concentrate to obtain the final lead-silver concentrate. Wherein the inhibitor used for the first concentration in the two concentration flotation is the same as the inhibitor used for the first roughing flotation in the step (3), and the dosage of the inhibitor is 150g/t.
(4) Zinc-sulfur separation: and (3) carrying out three scavenging flotation on the rougher tailings obtained in the step (3) to obtain scavenging concentrate and scavenging tailings. Wherein, the scavenging concentrate returns to the third scavenging in the three scavenging of the upper stage to form a closed cycle; the collecting agents used in the first and second scavenging in the third scavenging floatation are butyl xanthate, and the dosage of the collecting agents is 70g/t.
And then adjusting the concentration of the ore pulp of the scavenging tailings to 35wt.%, adjusting the pH value of the ore pulp to be 12 by using quicklime, and performing primary roughing flotation to obtain rough zinc concentrate and rough sulfur tailings. Wherein, the inhibitor used for primary roughing flotation is mass ratio=2: 1, the dosage of the inhibitor is 1700g/t; the collector used for primary roughing flotation is butyl xanthate, and the dosage of the collector is 80g/t.
And (3) carrying out three-time concentration flotation on the coarse zinc concentrate to obtain final zinc concentrate and concentrated middlings. Wherein, the middling is selected for the third time in the upper three-time selecting floatation to form a closed cycle.
And (3) carrying out twice scavenging flotation on the coarse sulfur tailings to obtain first sulfur concentrate and zinc sulfur tailings. Wherein, the inhibitor used for the twice scavenging flotation is that the mass ratio=2: 1, the dosage of the inhibitor is 600g/t, the collector used for the twice scavenging flotation is butyl xanthate, and the dosage of the butyl xanthate is 80g/t. And returning the first sulfur concentrate to the second scavenging in the upper-level twice scavenging flotation to form closed cycle.
(5) Sulfur-arsenic mixing and floating: and (3) carrying out magnetic separation on the zinc-sulfur tailings obtained in the step (4), wherein the magnetic field strength is 5000Gs, and removing magnetic minerals to obtain magnetic tailings and magnetic concentrate. And grading the magnetic separation tailings by a spiral grader to obtain underflow and overflow, discarding the overflow as tailings, and carrying out primary roughing, secondary concentration and tertiary scavenging on the underflow to obtain sulfur-arsenic mixed flotation concentrate and mixed flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 70g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 60g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage is 70g/t.
And (3) carrying out first shaking table gravity separation on the mixed floating tailings in a fine mud shaking table, wherein the stroke is 18mm, the stroke frequency is 240 times/min, the first shaking table gravity separation concentrate is low-grade tin gravity separation concentrate, and the first shaking table gravity separation tailings are classified as final tailings.
(6) Sulfur-arsenic separation: and (3) carrying out primary roughing, secondary concentration and tertiary scavenging on the sulfur-arsenic mixed concentrate obtained in the step (5) to obtain second sulfur concentrate. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 50g/t; the inhibitor used for one-time roughing is sodium thioglycolate, and the dosage of the inhibitor is 2300g/t; the inhibitor used in the first concentration and the second concentration in the two concentrations is sodium thioglycolate, and the dosage of the inhibitor is 2300g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage of the collecting agents is 50g/t.
Mixing the second sulfur concentrate with the magnetic concentrate obtained in the step (5) to obtain a final sulfur concentrate product.
(7) Desulfurizing and floatation: concentrating the overflow obtained in the step (1) by a thickener to obtain settled sand, adjusting the concentration of the settled sand to 40wt.% and the pH=7, and then performing desulfurization flotation of one roughing, two concentration and three scavenging to obtain third sulfur concentrate and first desulfurization flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 100g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 70g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage of the collecting agents is 100g/t. And (3) returning the third sulfur concentrate to the step (3) for mixing with the floatable concentrate (continuing the concentration and four-stage grinding process).
(8) Floating tin and collecting tin: and (3) adjusting the concentration of the first desulfurization flotation tailings obtained in the step (7) to 35wt.% and the pH=7, and then performing one roughing, two concentration and three scavenging to obtain the tin concentrate. Wherein the collectors used for one roughing are cinnamon hydroxamic acid (the preparation method is the same as the step (1) of the embodiment) and a P86 collector, the dosage of the cinnamon hydroxamic acid is 50g/t, and the dosage of the P86 collector is 100g/t; the inhibitor used for one roughing is a hydrochloric acid solution containing sodium silicate (prepared by mixing a hydrochloric acid solution with the molar concentration of 1mol/l and a sodium silicate solution with the mass fraction of 40% in a volume ratio of = 1:1.3 for 40 min), and the dosage of the inhibitor is 600g/t; the activator used in one roughing is lead nitrate, and the dosage of the activator is 80g/t. The inhibitor used in the first of the two beneficiations was a sodium silicate-containing hydrochloric acid solution (preparation method as above) in an amount of 600g/t. The collectors used in the first and second of the three scans were cinnamic hydroxamic acid (the preparation method is the same as step (1) of this example) and P86 collector, the amounts of cinnamic hydroxamic acid and P86 collector were 50g/t, and 100g/t, respectively.
And (3) carrying out magnetic separation on the float tin concentrate to remove magnetic minerals, wherein the magnetic field strength is 4500Gs, and obtaining the magnetic separation tailings. And (3) carrying out second shaking table reselection on the magnetic separation tailings in a fine mud shaking table, wherein the stroke is 7mm, and the stroke frequency is 280 times/min, so as to obtain fine mud tin concentrate.
(9) Tin collection: and (3) carrying out third shaking table reselection on the iso-floatable tailings obtained in the step (2) in a fine sand shaking table, wherein the stroke is 12mm, and the stroke frequency is 200r/min, so as to obtain second tin concentrate and shaking table tailings. Mixing the second tin concentrate with the first tin concentrate obtained in the step (1) to obtain a final tin concentrate product.
And (3) grinding the table tailings in five stages (the particle ratio of less than 200 meshes is 80 wt.%), adjusting the concentration to 40wt.% and the pH value to be 7, and then performing desulfurization flotation of one roughing, two concentration and three scavenging to obtain fourth sulfur concentrate and second desulfurization flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 100g/t; the activating agent used in the primary roughing is copper sulfate, and the dosage of the activating agent is 120g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage of the collecting agents is 100g/t. And (3) returning the fourth sulfur concentrate to the step (3) for mixing with the floatable concentrate (continuing the concentration and four-stage grinding process).
And (3) carrying out fourth shaking table gravity separation desulfurization on the second desulfurization flotation tailings in a middling shaking table, wherein the stroke is 10mm, and the stroke frequency is 220 times/min, so as to obtain the moderate tin concentrate.
Through the steps, the lead grade (namely the mass fraction of lead element) in the lead-silver concentrate obtained by the embodiment is 24.11%, the recovery rate of lead element is 88.54%, the silver grade (namely the mass fraction of silver element) is 155g/t, and the recovery rate of silver element is 90.71%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 45.48 percent, and the recovery rate of zinc element is 88.27 percent; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 40.99%, and the recovery rate of tin element is 74.04%.
Example 4
The complex tin-lead-zinc polymetallic ore (abbreviated as raw ore) provided in this embodiment is selected from tin polymetallic sulfide ore in some places in Yunnan, wherein main metal minerals are cassiterite, galena, chalcopyrite, marmatite, pyrite and arsenopyrite, gangue minerals are mainly quartz and calcite, wherein the tin element content is 0.21 wt%, the lead element content is 0.17 wt%, the copper element content is 0.30 wt%, the zinc element content is 1.32 wt%, the sulfur element content is 7.15 wt%, and the arsenic element content is 1.66 wt%, and the separation and recovery method of the complex tin-lead-zinc polymetallic ore specifically comprises the following steps:
(1) Grading and jigging waste disposal: screening raw ore into three particle size grades, carrying out primary grinding on the ore with the particle size larger than 4mm (namely high-sized ore), returning to an upper screening process, and carrying out jigging and beneficiation on the ore with the particle size of 2-4 mm (namely medium-sized ore) to obtain jigged rough concentrate and jigged middling (jigged tailings are directly thrown waste); the jigged rough concentrate is subjected to secondary grinding (the particle size of the ore after secondary grinding is less than 200 meshes is 60 wt.%) and then is mixed with the ore with the particle size of less than 2mm (namely, the low-size ore) and is subjected to reselection through a conical spiral chute system, so that front heavy concentrate (underflow) and overflow are obtained; the front heavy concentrate enters a deck floating table to carry out deck floating table flotation, the stroke is 20mm, the stroke frequency is 350 times/min, the collector used for deck floating table flotation is cinnamon hydroxamic acid, the dosage of the collector is 400g/t, and first tin concentrate and deck floating tailings are obtained; mixing the table floating tailings with the jigged middlings, and carrying out three-stage grinding to obtain combined ores.
The preparation method of the cinnamon hydroxamic acid comprises the following steps: under inert atmosphere, mixing styryl, N-hydroxyformamide and nitric acid in a molar ratio of 1:2:1, heating to react, and sequentially distilling and drying after the reaction is completed to obtain cinnamyl hydroxamic acid; wherein the heating temperature is 280 ℃, the reaction time is 3 hours, the distillation time is 6 hours, and the drying time is 8 hours.
(2) And (3) adjusting the concentration of the ore pulp of the combined ore obtained in the step (1) to 50wt.%, and then performing primary roughing, secondary concentration and tertiary scavenging and the like to obtain the equifloatable concentrate and the equifloatable tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 80g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 60g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage is 80g/t.
(3) Separating and floating lead copper and zinc sulfur: concentrating the equi-floatable concentrate obtained in the step (2) by a thickener, performing four-stage grinding (the particle ratio of the mineral with the granularity smaller than 200 meshes is 80 wt.%) and then adjusting the pH value to be 12 by quicklime, and performing primary roughing flotation to obtain roughing concentrate and roughing tailings. The inhibitor for primary roughing flotation comprises sodium carbo-cyanuric acid, alkali liquor containing thiourea, alkali liquor containing polymeric ferric sulfate and carbonate in a mass ratio of = 2:5:4:2, wherein the dosage of the inhibitor is 1500g/t, the dosage of the collector butyl xanthate for primary roughing flotation is 120g/t, and the dosage of the foamer 2# oil for primary roughing flotation is 50g/t.
Wherein the alkali solution containing thiourea is a mixed solution containing thiourea and potassium hydroxide, and the pH value of the alkali solution is 10.6. The alkali liquor containing the polymeric ferric sulfate is a mixed liquor containing the polymeric ferric sulfate and potassium hydroxide, and the pH value of the alkali liquor is 10.6. The carbonate is sodium carbonate.
Lead concentrate flotation: and (3) carrying out twice concentration flotation on the roughing concentrate to obtain the final lead copper concentrate. Wherein the inhibitor used for the first concentration in the two concentration flotation is the same as the inhibitor used for the first roughing flotation in the step (3), and the dosage of the inhibitor is 400g/t.
(4) Zinc-sulfur separation: and (3) carrying out three scavenging flotation on the rougher tailings obtained in the step (3) to obtain scavenging concentrate and scavenging tailings. Wherein, the scavenging concentrate returns to the third scavenging in the three scavenging of the upper stage to form a closed cycle; the collecting agents used in the first and second scavenging in the third scavenging floatation are butyl xanthate, and the dosage of the collecting agents is 80g/t.
And then adjusting the concentration of the ore pulp of the scavenging tailings to 38wt.%, adjusting the pH value of the ore pulp to be 11 by using quicklime, and performing primary roughing flotation to obtain rough zinc concentrate and rough sulfur tailings. Wherein, the inhibitor used for primary roughing flotation is mass ratio=2: 1, wherein the dosage of the inhibitor is 1800g/t; the collector used for primary roughing flotation is butyl xanthate, and the dosage of the collector is 100g/t.
And (3) carrying out three-time concentration flotation on the coarse zinc concentrate to obtain final zinc concentrate and concentrated middlings. Wherein, the middling is selected for the third time in the upper three-time selecting floatation to form a closed cycle.
And (3) carrying out twice scavenging flotation on the coarse sulfur tailings to obtain first sulfur concentrate and zinc sulfur tailings. Wherein, the inhibitor used for the twice scavenging flotation is that the mass ratio=2: 1, the dosage of the inhibitor is 700g/t, the collector used for twice scavenging flotation is butyl xanthate, and the dosage of the butyl xanthate is 100g/t. And returning the first sulfur concentrate to the second scavenging in the upper-level twice scavenging flotation to form closed cycle.
(5) Sulfur-arsenic mixing and floating: and (3) carrying out magnetic separation on the zinc-sulfur tailings obtained in the step (4), wherein the magnetic field strength is 8000Gs, and removing magnetic minerals to obtain magnetic tailings and magnetic concentrate. And grading the magnetic separation tailings by a spiral grader to obtain underflow and overflow, discarding the overflow as tailings, and carrying out primary roughing, secondary concentration and tertiary scavenging on the underflow to obtain sulfur-arsenic mixed flotation concentrate and mixed flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 80g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 80g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage is 80g/t.
And (3) carrying out first shaking table gravity separation on the mixed floating tailings in a fine mud shaking table, wherein the stroke is 22mm, the stroke frequency is 250 times/min, the first shaking table gravity separation concentrate is the low-grade tin gravity separation concentrate, and the first shaking table gravity separation tailings are classified as final tailings.
(6) Sulfur-arsenic separation: and (3) carrying out primary roughing, secondary concentration and tertiary scavenging on the sulfur-arsenic mixed concentrate obtained in the step (5) to obtain second sulfur concentrate. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 50g/t; the inhibitor used for one roughing is sodium thioglycolate, and the dosage of the inhibitor is 2200g/t; the inhibitor used in the first concentration and the second concentration in the two concentrations is sodium thioglycolate, and the dosage of the inhibitor is 2200g/t; the collecting agents used in the first and second scans are butyl xanthate, and the dosage of the collecting agents is 50g/t.
Mixing the second sulfur concentrate with the magnetic concentrate obtained in the step (5) to obtain a final sulfur concentrate product.
(7) Desulfurizing and floatation: concentrating the overflow obtained in the step (1) by a thickener to obtain settled sand, adjusting the concentration of the settled sand to 40wt.% and the pH=6.5, and then performing desulfurization flotation of one roughing, two concentration and three scavenging to obtain third sulfur concentrate and first desulfurization flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 120g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 80g/t; the collecting agents used in the first and second scanning in the three scanning are butyl xanthate, and the dosage of the collecting agents is 120g/t. And (3) returning the third sulfur concentrate to the step (3) for mixing with the floatable concentrate (continuing the concentration and four-stage grinding process).
(8) Floating tin and collecting tin: and (3) adjusting the concentration of the first desulfurization flotation tailings obtained in the step (7) to 38wt.% and the pH=7, and then performing one roughing, two concentration and three scavenging to obtain the tin concentrate. Wherein the collectors used for one roughing are cinnamon hydroxamic acid (the preparation method is the same as the step (1) of the embodiment) and P86 collector, the dosage of the cinnamon hydroxamic acid is 180g/t, and the dosage of the P86 collector is 130g/t; the inhibitor used for one roughing is oxalic acid solution containing sodium silicate (prepared by mixing oxalic acid solution with the molar concentration of 1mol/l and sodium silicate solution with the mass fraction of 50% in a volume ratio of = 1:1 for 30 min), and the dosage of the inhibitor is 1400g/t; the activator used in one roughing is lead nitrate, and the dosage of the activator is 50g/t. The inhibitor used in the first of the two beneficiations was an oxalic acid solution containing sodium silicate (preparation method as above) in an amount of 1400g/t. The collectors used in the first and second of the three scans were cinnamic hydroxamic acid (the preparation method is the same as step (1) of this example) and P86 collector, the amounts of cinnamic hydroxamic acid and P86 collector were 180g/t, and 130g/t, respectively.
And (3) carrying out magnetic separation on the float tin concentrate to remove magnetic minerals, wherein the magnetic field strength is 8000Gs, and obtaining the magnetic separation tailings. And (3) carrying out second shaking table reselection on the magnetic separation tailings in a fine mud shaking table, wherein the stroke is 15mm, and the stroke frequency is 320 times/min, so as to obtain fine mud tin concentrate.
(9) Tin collection: and (3) carrying out third shaking table reselection on the iso-floatable tailings obtained in the step (2) in a fine sand shaking table, wherein the stroke is 15mm, and the stroke frequency is 280r/min, so as to obtain second tin concentrate and shaking table tailings. Mixing the second tin concentrate with the first tin concentrate obtained in the step (1) to obtain a final tin concentrate product.
And (3) grinding the table tailings in five stages (the particle ratio of less than 200 meshes is 80 wt.%), adjusting the concentration to 40wt.% and the pH value to be 7, and then performing desulfurization flotation of one roughing, two concentration and three scavenging to obtain fourth sulfur concentrate and second desulfurization flotation tailings. Wherein, the collecting agent used for one-time roughing is butyl xanthate, and the dosage is 120g/t; the activating agent used for primary roughing is copper sulfate, and the dosage of the activating agent is 100g/t; the collecting agents used in the first and second scanning in the three scanning are butyl xanthate, and the dosage of the collecting agents is 120g/t. And (3) returning the fourth sulfur concentrate to the step (3) for mixing with the floatable concentrate (continuing the concentration and four-stage grinding process).
And (3) carrying out fourth shaking table gravity separation desulfurization on the second desulfurization flotation tailings in a middling shaking table, wherein the stroke is 20mm, and the stroke frequency is 350 times/min, so as to obtain moderate tin concentrate.
Through the steps, the lead grade (namely the mass fraction of lead element) in the lead-copper concentrate obtained by the embodiment is 24.96%, the recovery rate of lead element is 89.48%, the copper grade (namely the mass fraction of antimony element) is 17.25%, and the recovery rate of copper element is 94.88%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 45.09%, and the recovery rate of zinc element is 88.85%; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 40.96%, and the recovery rate of tin element is 72.89%.
Comparative example 1
The separation and recovery method of the complex tin-lead-zinc polymetallic ore provided by the comparative example is basically the same as that of the embodiment 1, and the difference is that in the step (1), the jigged rough concentrate is mixed with the ore with the grain size smaller than 2mm after two-stage ore grinding, and then the mixture is not subjected to reselection through a conical spiral chute system, but directly enters a table floating table to be subjected to table floating.
In the lead-antimony concentrate obtained in the comparative example, the lead grade (namely, the mass fraction of lead element) is 20.34%, the lead element recovery rate is 80.69%, the antimony grade (namely, the mass fraction of antimony element) is 10.23%, and the antimony element recovery rate is 79.22%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 40.39 percent, and the recovery rate of the zinc element is 85.36 percent; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 39.68%, and the recovery rate of tin element is 68.25%.
Comparative example 2
The separation and recovery method of the complex tin-lead-zinc polymetallic ore provided by the comparative example is basically the same as that of the embodiment 1, and is different in that the 2-4 mm size fraction ore in the step (1) is not subjected to jigging beneficiation, but the 2-4 mm size fraction ore is directly mixed with the table floating tailings and subjected to three-stage grinding.
In the lead-antimony concentrate obtained in the comparative example, the lead grade (namely, the mass fraction of lead element) is 20.02%, the lead element recovery rate is 80.39%, the antimony grade (namely, the mass fraction of antimony element) is 10.97%, and the antimony element recovery rate is 79.36%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 40.32%, and the recovery rate of zinc element is 86.31%; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 38.65%, and the recovery rate of tin element is 67.21%.
Comparative example 3
The separation and recovery method of the complex tin-lead-zinc multi-metal ore provided in this comparative example is basically the same as that of example 1, except that four-stage ore grinding is not performed in step (3), and five-stage ore grinding is not performed in step (9).
In the lead-antimony concentrate obtained in the comparative example, the lead grade (i.e. mass fraction of lead element) is 19.56%, the lead element recovery rate is 79.98%, the antimony grade (i.e. mass fraction of antimony element) is 12.03%, and the antimony element recovery rate is 78.99%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 39.70 percent, and the recovery rate of the zinc element is 85.69 percent; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 37.25%, and the recovery rate of tin element is 66.98%.
Comparative example 4
The separation and recovery method of the complex tin-lead-zinc polymetallic ore provided by the comparative example is basically the same as that of comparative example 1, and the difference is that in the step (1), the collecting agent cinnamon hydroxamic acid used for table floating bed flotation is replaced by butyl xanthate with equal mass.
In the lead-antimony concentrate obtained in the comparative example, the lead grade (namely, the mass fraction of lead element) is 18.56%, the lead element recovery rate is 75.64%, the antimony grade (namely, the mass fraction of antimony element) is 10.25%, and the antimony element recovery rate is 75.39%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 38.12 percent, and the recovery rate of the zinc element is 84.10 percent; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 35.12%, and the recovery rate of tin element is 63.21%.
Comparative example 5
The method for separating and recovering the complex tin-lead-zinc polymetallic ore provided in the present comparative example is basically the same as comparative example 1, except that in the step (3), the inhibitor used for the primary roughing flotation is replaced with sodium sulfite of equal mass, and the inhibitor used for the primary refining in the secondary refining flotation is replaced with sodium sulfite of equal mass.
In the lead-antimony concentrate obtained in the comparative example, the lead grade (i.e., mass fraction of lead element) is 17.69%, the lead element recovery rate is 76.21%, the antimony grade (i.e., mass fraction of antimony element) is 11.03%, and the antimony element recovery rate is 74.12%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 38.14 percent, and the recovery rate of the zinc element is 84.13 percent; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 36.18%, and the recovery rate of tin element is 64.13%.
Comparative example 6
The method for separating and recovering the complex tin-lead-zinc polymetallic ore provided in the present comparative example is basically the same as comparative example 1, except that in the step (6), sodium thioglycolate as an inhibitor for primary roughing is replaced with sodium sulfite of equal mass, and the inhibitors for primary refining and secondary refining in the two refinements are replaced with sodium sulfite of equal mass.
In the lead-antimony concentrate obtained in the comparative example, the lead grade (namely, the mass fraction of lead element) is 17.15%, the lead element recovery rate is 71.26%, the antimony grade (namely, the mass fraction of antimony element) is 11.20%, and the antimony element recovery rate is 70.17%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 35.14 percent, and the recovery rate of the zinc element is 81.01 percent; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 36.18%, and the recovery rate of tin element is 61.29%.
Comparative example 7
The separation and recovery method of the complex tin-lead-zinc polymetallic ore provided in the comparative example is basically the same as that of comparative example 1, except that in the step (8), the cinnamic hydroxamic acid in the collector used in the primary roughing is replaced by the P86 collector with the same mass, and the collectors used in the primary scavenging and the secondary scavenging in the three scavenging are replaced by the P86 collector with the same mass, namely, the cinnamic hydroxamic acid is not added in the step (8).
In the lead-antimony concentrate obtained in the comparative example, the lead grade (i.e., mass fraction of lead element) is 18.66%, the lead element recovery rate is 77.45%, the antimony grade (i.e., mass fraction of antimony element) is 10.11%, and the antimony element recovery rate is 76.12%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 37.14 percent, and the recovery rate of the zinc element is 81.26 percent; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 35.12%, and the recovery rate of tin element is 65.13%.
Comparative example 8
The separation and recovery method of the complex tin-lead-zinc polymetallic ore provided in this comparative example is basically the same as that of comparative example 1, except that in step (8), the oxalic acid solution containing sodium silicate as the inhibitor used in the primary roughing is replaced with water glass (i.e., an aqueous solution of sodium silicate, in which the mass fraction of sodium silicate is 25%), and the oxalic acid solution containing sodium silicate as the inhibitor used in the primary refining in the secondary refining is replaced with water glass (in which the mass fraction of sodium silicate is 25%).
In the lead-antimony concentrate obtained in the comparative example, the lead grade (namely, the mass fraction of lead element) is 17.23%, the lead element recovery rate is 75.17%, the antimony grade (namely, the mass fraction of antimony element) is 11.02%, and the antimony element recovery rate is 74.16%; the zinc grade (namely the mass fraction of zinc element) in the zinc concentrate is 34.13 percent, and the recovery rate of zinc element is 86.10 percent; the grade of tin (namely the mass fraction of tin element) in the tin concentrate is 35.25%, and the recovery rate of tin element is 61.21%.
As can be seen from the comparison of the embodiment 1 and the comparison of the embodiment 1 to the comparison 3, the method for separating and recovering the complex tin-lead-zinc polymetallic ore provided by the invention can be suitable for complex tin-lead-zinc polymetallic ore containing high gangue minerals and having great difference in ore grindability, valuable metal resources are efficiently recovered to have higher grade and high recovery rate, meanwhile, the medicament consumption is reduced, the production cost is reduced, the process flow is simple, the adaptability is strong, and the popularization is easy.
As can be seen from comparison of comparative examples 1 and 4 to 8, the recovery rate or grade of valuable metal elements can be further improved by adopting a suitable reagent in each step.
While the invention has been illustrated and described with reference to specific embodiments, it is to be understood that the above embodiments are merely illustrative of the technical aspects of the invention and not restrictive thereof; those of ordinary skill in the art will appreciate that: modifications may be made to the technical solutions described in the foregoing embodiments, or equivalents may be substituted for some or all of the technical features thereof, without departing from the spirit and scope of the present invention; such modifications and substitutions do not depart from the spirit of the corresponding technical solutions; it is therefore intended to cover in the appended claims all such alternatives and modifications as fall within the scope of the invention.

Claims (10)

1.一种复杂锡铅锌多金属矿分离回收方法，其特征在于，包括如下步骤：1. A method for separating and recovering complex tin-lead-zinc polymetallic ores, characterized in that it comprises the following steps: （a）、将复杂锡铅锌多金属矿筛分为低粒级矿、中粒级矿和高粒级矿，所述中粒级矿经跳汰选矿后得到跳汰粗精矿和跳汰中矿，所述跳汰粗精矿经二段磨矿后与所述低粒级矿混合并进行重选得到前重精矿和溢流；将所述前重精矿进行枱浮摇床浮选得到第一锡精矿和枱浮尾矿；所述枱浮尾矿与所述跳汰中矿混合并进行三段磨矿，得到合并矿；(a) screening a complex tin-lead-zinc polymetallic ore into low-grade ore, medium-grade ore and high-grade ore; the medium-grade ore is subjected to jigging to obtain a jigging coarse concentrate and a jigging middling; the jigging coarse concentrate is subjected to two-stage grinding to be mixed with the low-grade ore and subjected to gravity separation to obtain a pre-heavy concentrate and an overflow; the pre-heavy concentrate is subjected to table flotation shaking table flotation to obtain a first tin concentrate and a table flotation tailings; the table flotation tailings is mixed with the jigging middlings and subjected to three-stage grinding to obtain a combined ore; （b）、将步骤（a）得到的所述合并矿进行一次粗选、两次精选和三次扫选的等可浮浮选，得到等可浮精矿和等可浮尾矿；(b) subjecting the combined ore obtained in step (a) to isoflurane flotation with one roughing selection, two cleaning selections and three scavenging selections to obtain isoflurane concentrate and isoflurane tailings; （c）、铅与锌硫的分离浮选：将步骤（b）得到的所述等可浮精矿进行浓缩和四段磨矿，然后调节pH并进行一次粗选浮选，得到粗选精矿和粗选尾矿；(c) Separation and flotation of lead and zinc and sulfur: the floatable concentrate obtained in step (b) is concentrated and subjected to four-stage grinding, and then the pH is adjusted and a roughing flotation is performed once to obtain a roughing concentrate and a roughing tailing; 铅精矿浮选：将所述粗选精矿进行两次精选浮选，得到含铅精矿，所述含铅精矿包括铅锑精矿、铅银精矿和铅铜精矿中的至少一种；Lead concentrate flotation: subjecting the rougher concentrate to two concentration flotations to obtain a lead-containing concentrate, wherein the lead-containing concentrate includes at least one of a lead-antimony concentrate, a lead-silver concentrate and a lead-copper concentrate; （d）、锌-硫分离：将步骤（c）得到的所述粗选尾矿进行三次扫选浮选，得到扫选精矿和扫选尾矿；调节所述扫选尾矿的矿浆的浓度和pH，然后进行一次粗选浮选，得到粗锌精矿和粗硫尾矿；将所述粗锌精矿进行三次精选浮选，得到锌精矿和精选中矿；将所述粗硫尾矿进行两次扫选浮选，得到第一硫精矿和锌硫尾矿；(d) Zinc-sulfur separation: subjecting the rougher tailings obtained in step (c) to scavenging flotation three times to obtain scavenging concentrate and scavenging tailings; adjusting the concentration and pH of the slurry of the scavenging tailings, and then subjecting the rougher flotation to one time to obtain a coarse zinc concentrate and a coarse sulfur tailings; subjecting the coarse zinc concentrate to three times of concentrating flotation to obtain a zinc concentrate and a concentrating middlings; subjecting the coarse sulfur tailings to two times of scavenging flotation to obtain a first sulfur concentrate and a zinc-sulfur tailings; （e）、硫-砷混浮：将步骤（d）得到的所述锌硫尾矿进行磁选，得到磁选尾矿和磁性精矿，将所述磁选尾矿经分级处理后所得的底流进行一次粗选、两次精选和三次扫选，得到硫砷混浮精矿和混浮尾矿；(e) Sulfur-arsenic mixed flotation: subjecting the zinc-sulfur tailings obtained in step (d) to magnetic separation to obtain magnetic tailings and magnetic concentrates, subjecting the underflow obtained after the magnetic tailings are graded to one roughing separation, two cleaning separations and three scavenging separations to obtain sulfur-arsenic mixed flotation concentrates and mixed flotation tailings; 将所述混浮尾矿进行第一摇床重选得到低度锡重选精矿；The mixed flotation tailings are subjected to first shaking table gravity separation to obtain a low-concentration tin gravity separation concentrate; （f）、硫-砷分离：将步骤（e）得到的所述硫砷混浮精矿进行一次粗选、两次精选和三次扫选，得到第二硫精矿；(f) Sulfur-arsenic separation: the sulfur-arsenic mixed flotation concentrate obtained in step (e) is subjected to one roughing selection, two cleaning selections and three scavenging selections to obtain a second sulfur concentrate; 将所述第二硫精矿与步骤（e）得到的所述磁性精矿混合得到硫精矿产品；mixing the second sulfur concentrate with the magnetic concentrate obtained in step (e) to obtain a sulfur concentrate product; （g）、脱硫浮选：将步骤（a）得到的所述溢流进行浓缩处理得到沉砂，调节所述沉砂的浓度和pH，然后进行一次粗选、两次精选和三次扫选的脱硫浮选，得到第三硫精矿和第一脱硫浮选尾矿；所述第三硫精矿返回步骤（c）与所述等可浮精矿混合；(g) Desulfurization flotation: the overflow obtained in step (a) is concentrated to obtain sediment, the concentration and pH of the sediment are adjusted, and then a roughing selection, a cleaning selection and a scavenging selection are performed to obtain a third sulfur concentrate and a first desulfurization flotation tailing; the third sulfur concentrate is returned to step (c) to be mixed with the isofloatable concentrate; （h）、浮锡和收锡：调节步骤（g）得到的所述第一脱硫浮选尾矿的浓度和pH，然后进行一次粗选、两次精选和三次扫选，得到浮锡精矿；将所述浮锡精矿进行磁选得到磁选尾矿，所述磁选尾矿经第二摇床重选后得到细泥锡精矿；(h), floating tin and collecting tin: adjusting the concentration and pH of the first desulfurization flotation tailings obtained in step (g), and then performing one roughing selection, two cleaning selections and three scavenging selections to obtain floating tin concentrate; performing magnetic separation on the floating tin concentrate to obtain magnetic separation tailings, and performing re-selection on the second shaking table on the magnetic separation tailings to obtain fine mud tin concentrate; （i）、收锡：将步骤（b）得到的所述等可浮尾矿进行第三摇床重选，得到第二锡精矿和摇床尾矿；所述第二锡精矿和步骤（a）得到的所述第一锡精矿混合得到锡精矿产品；所述摇床尾矿经五段磨矿后调节浓度和pH，然后进行一次粗选、两次精选和三次扫选的脱硫浮选，得到第四硫精矿和第二脱硫浮选尾矿；所述第四硫精矿返回步骤（c）与所述等可浮精矿混合；所述第二脱硫浮选尾矿进行第四摇床重选脱硫，得到中度锡精矿。(i) Tin collection: subjecting the iso-floatable tailings obtained in step (b) to a third shaking table gravity separation to obtain a second tin concentrate and shaking table tailings; the second tin concentrate is mixed with the first tin concentrate obtained in step (a) to obtain a tin concentrate product; the shaking table tailings are subjected to five-stage grinding to adjust the concentration and pH, and then subjected to one roughing separation, two cleaning separations and three scavenging separations for desulfurization flotation to obtain a fourth sulfur concentrate and a second desulfurization flotation tailings; the fourth sulfur concentrate is returned to step (c) to be mixed with the iso-floatable concentrate; the second desulfurization flotation tailings are subjected to a fourth shaking table gravity separation for desulfurization to obtain a medium tin concentrate. 2.根据权利要求1所述复杂锡铅锌多金属矿分离回收方法，其特征在于，步骤（a）中，满足以下条件（1）至（5）中的至少一个：2. The method for separating and recovering complex tin-lead-zinc polymetallic ores according to claim 1, characterized in that, in step (a), at least one of the following conditions (1) to (5) is satisfied: （1）所述低粒级矿的粒度小于2mm，所述中粒级矿的粒度为2~4mm，所述高粒级矿的粒度大于4mm；(1) The particle size of the low-grade ore is less than 2 mm, the particle size of the medium-grade ore is 2-4 mm, and the particle size of the high-grade ore is greater than 4 mm; （2）将所述高粒级矿进行一段磨矿后返回上级的筛分工序；(2) grinding the high-grade ore and returning it to the upper screening process; （3）所述重选在圆锥螺旋溜槽系统中进行；(3) The gravity separation is carried out in a conical spiral chute system; （4）所述枱浮摇床浮选的冲程为5~20mm，冲次为240~360次/min；(4) The flotation stroke of the table flotation shaker is 5-20 mm, and the frequency is 240-360 times/min; （5）所述枱浮摇床浮选所用的捕收剂包括肉桂异羟肟酸，所述肉桂异羟肟酸的分子式为C6H5-CH=CH–(C=O)NHOH，所述肉桂异羟肟酸的用量为200~400g/t；所述肉桂异羟肟酸的制备方法包括：在惰性气氛下，将苯乙烯基、N-羟基甲酰胺和硝酸混合后加热进行反应，待所述反应完成后依次进行蒸馏和干燥，得到所述肉桂异羟肟酸；所述加热的温度为200~300℃；所述反应的时间为2~5h；所述蒸馏的时间为4~8h；所述干燥的时间为8~24h；(5) The collector used for the table flotation includes cinnamic hydroxamic acid, the molecular formula of the cinnamic hydroxamic acid is C 6 H 5 -CH=CH–(C=O)NHOH, and the amount of the cinnamic hydroxamic acid used is 200-400 g/t; the preparation method of the cinnamic hydroxamic acid includes: in an inert atmosphere, mixing styryl, N-hydroxyformamide and nitric acid, heating and reacting, and after the reaction is completed, distilling and drying in sequence to obtain the cinnamic hydroxamic acid; the heating temperature is 200-300° C.; the reaction time is 2-5 h; the distillation time is 4-8 h; and the drying time is 8-24 h; 或者，步骤（b）中，满足以下条件（6）至（9）中的至少一个：Alternatively, in step (b), at least one of the following conditions (6) to (9) is satisfied: （6）所述等可浮浮选的一次粗选所用的捕收剂包括丁基黄药，所述捕收剂的用量为30~80g/t；(6) The collector used in the primary roughing of the iso-floatable flotation comprises butyl xanthate, and the amount of the collector used is 30-80 g/t; （7）所述等可浮浮选的三次扫选中的第一次扫选和第二次扫选所用的捕收剂包括丁基黄药，所述捕收剂的用量为30~80g/t；(7) The collector used in the first and second scavenging of the three scavenging of the equal flotation flotation comprises butyl xanthate, and the amount of the collector used is 30-80 g/t; （8）所述等可浮浮选的一次粗选所用的活化剂包括硫酸铜，所述活化剂的用量为30~60g/t；(8) The activator used in the primary roughing of the iso-floatable flotation comprises copper sulfate, and the amount of the activator used is 30-60 g/t; （9）将所述合并矿的浆料浓度调节至30wt.%~50wt.%后再进行所述等可浮浮选。(9) The slurry concentration of the combined ore is adjusted to 30wt.%~50wt.% before performing the iso-flotation flotation. 3.根据权利要求1所述复杂锡铅锌多金属矿分离回收方法，其特征在于，步骤（c）中，满足以下条件至少其一：3. The method for separating and recovering complex tin-lead-zinc polymetallic ores according to claim 1, characterized in that, in step (c), at least one of the following conditions is met: （1）所述调节pH至10~12；(1) adjusting the pH to 10-12; （2）所述一次粗选浮选所用的抑制剂包括质量比为1~3：3~5：1~4：2~6的碳化三聚氰酸钠、含硫脲的碱液、含聚合硫酸铁的碱液和碳酸盐，所述抑制剂的用量为1000~1500g/t；(2) The depressant used in the primary roughing flotation comprises carbonized sodium cyanurate, alkali solution containing thiourea, alkali solution containing polyferric sulfate and carbonate in a mass ratio of 1-3:3-5:1-4:2-6, and the amount of the depressant is 1000-1500 g/t; （3）所述一次粗选浮选所用的捕收剂包括丁基黄药，所述捕收剂的用量为100~120g/t；(3) The collector used in the primary roughing flotation comprises butyl xanthate, and the amount of the collector used is 100-120 g/t; （4）所述一次粗选浮选所用的起泡剂包括2#油，所述起泡剂的用量为30~50g/t；(4) The frother used in the primary roughing flotation includes 2# oil, and the amount of the frother is 30-50 g/t; （5）所述两次精选浮选中的第一次精选所用的抑制剂包括质量比为1~3：3~5：1~4：2~6的碳化三聚氰酸钠、含硫脲的碱液、含聚合硫酸铁的碱液和碳酸盐，所述抑制剂的用量为150~400g/t。(5) The depressant used in the first of the two flotation separations comprises carbonized sodium cyanurate, alkali solution containing thiourea, alkali solution containing polyferric sulfate and carbonate in a mass ratio of 1-3:3-5:1-4:2-6, and the amount of the depressant used is 150-400 g/t. 4.根据权利要求1所述复杂锡铅锌多金属矿分离回收方法，其特征在于，步骤（d）中，满足以下条件至少其一：4. The method for separating and recovering complex tin-lead-zinc polymetallic ores according to claim 1, characterized in that in step (d), at least one of the following conditions is met: （1）将所述扫选精矿返回上级的所述三次扫选中的第三次扫选；(1) returning the scavenged concentrate to the third scavenging of the three scavengings at the upper level; （2）所述三次扫选浮选中的第一次扫选和第二次扫选所用的捕收剂包括丁基黄药，所述捕收剂的用量为20~80g/t；(2) The collector used in the first scavenging and the second scavenging of the three scavenging flotation comprises butyl xanthate, and the amount of the collector used is 20-80 g/t; （3）将所述扫选尾矿的矿浆的浓度调节至30wt.%~40wt.%；(3) adjusting the concentration of the slurry of the scavenged tailings to 30wt.%~40wt.%; （4）将所述扫选尾矿的矿浆的pH调节至11~12；(4) adjusting the pH of the slurry of the scavenged tailings to 11-12; （5）所述一次粗选浮选所用的抑制剂包括质量比为1~2：1的聚谷氨酸和聚天冬氨酸，所述抑制剂的用量为1600~1800g/t；(5) The depressant used in the primary roughing flotation comprises polyglutamic acid and polyaspartic acid in a mass ratio of 1 to 2:1, and the amount of the depressant is 1600 to 1800 g/t; （6）所述一次粗选浮选所用的捕收剂包括丁基黄药，所述捕收剂的用量为50~100g/t；(6) The collector used in the primary roughing flotation comprises butyl xanthate, and the amount of the collector used is 50-100 g/t; （7）将所述精选中矿返回上级的所述三次精选浮选中的第三次精选；(7) returning the concentrated middlings to the upper level for the third concentration of the three concentration flotation; （8）所述两次扫选浮选所用的抑制剂包括质量比为1~2：1的聚谷氨酸和聚天冬氨酸，所述抑制剂的用量为300~700g/t；(8) The depressant used in the two scavenging flotations comprises polyglutamic acid and polyaspartic acid in a mass ratio of 1 to 2:1, and the amount of the depressant is 300 to 700 g/t; （9）所述两次扫选浮选所用的捕收剂包括丁基黄药，所述捕收剂的用量为30~100g/t；(9) The collector used in the two scavenging flotation steps includes butyl xanthate, and the amount of the collector used is 30-100 g/t; （10）将所述第一硫精矿返回上级的所述两次扫选浮选中的第二次扫选。(10) Returning the first sulfur concentrate to the upper level for the second scavenging of the two scavenging flotation operations. 5.根据权利要求1所述复杂锡铅锌多金属矿分离回收方法，其特征在于，步骤（e）中，满足以下条件至少其一：5. The method for separating and recovering complex tin-lead-zinc polymetallic ores according to claim 1, characterized in that, in step (e), at least one of the following conditions is met: （1）所述磁选的磁场强度为5000~15000Gs；(1) The magnetic field strength of the magnetic separation is 5000~15000Gs; （2）所述一次粗选所用的捕收剂包括丁基黄药，所述捕收剂的用量为30~80g/t；(2) The collector used in the primary roughing includes butyl xanthate, and the amount of the collector used is 30-80 g/t; （3）所述三次扫选中的第一次扫选和第二次扫选所用的捕收剂包括丁基黄药，所述捕收剂的用量为30~80g/t；(3) The collector used in the first and second sweeps of the three sweeps includes butyl xanthate, and the amount of the collector used is 30-80 g/t; （4）所述一次粗选所用的活化剂包括硫酸铜，所述活化剂的用量为30~60g/t；(4) The activator used in the primary roughing includes copper sulfate, and the amount of the activator is 30-60 g/t; （5）所述第一摇床重选的冲程为16~22mm，冲次为200~260次/min；(5) The stroke of the first shaking table is 16-22 mm, and the stroke frequency is 200-260 times/min; （6）所述第一摇床重选在细泥摇床中进行。(6) The first shaking table gravity selection is carried out in a fine mud shaking table. 6.根据权利要求1所述复杂锡铅锌多金属矿分离回收方法，其特征在于，步骤（f）中，满足以下条件至少其一：6. The method for separating and recovering complex tin-lead-zinc polymetallic ores according to claim 1, characterized in that in step (f), at least one of the following conditions is met: （1）所述一次粗选所用的捕收剂包括丁基黄药，所述捕收剂的用量为30~50g/t；(1) The collector used in the primary roughing includes butyl xanthate, and the amount of the collector used is 30-50 g/t; （2）所述三次扫选中的第一次扫选和第二次扫选所用的捕收剂包括丁基黄药，所述捕收剂的用量为30~50g/t；(2) The collector used in the first and second sweeps of the three sweeps includes butyl xanthate, and the amount of the collector used is 30-50 g/t; （3）所述一次粗选所用的抑制剂包括巯基乙酸钠，所述抑制剂的用量为2000~3000g/t；(3) The inhibitor used in the primary roughing includes sodium thioglycolate, and the amount of the inhibitor used is 2000-3000 g/t; （4）所述两次精选中的第一次精选和第二次精选所用的抑制剂包括巯基乙酸钠，所述抑制剂的用量为2000~3000g/t。(4) The inhibitor used in the first and second concentrations of the two concentrations includes sodium thioglycolate, and the amount of the inhibitor used is 2000-3000 g/t. 7.根据权利要求1所述复杂锡铅锌多金属矿分离回收方法，其特征在于，步骤（g）中，满足以下条件至少其一：7. The method for separating and recovering complex tin-lead-zinc polymetallic ores according to claim 1, characterized in that in step (g), at least one of the following conditions is met: （1）调节所述沉砂的浓度至30wt.%~40wt.%；(1) adjusting the concentration of the sediment to 30wt.%~40wt.%; （2）调节所述沉砂的pH至6~7；(2) adjusting the pH of the sediment to 6-7; （3）所述一次粗选所用的捕收剂包括丁基黄药，所述捕收剂的用量为30~120g/t；(3) The collector used in the primary roughing includes butyl xanthate, and the amount of the collector used is 30-120 g/t; （4）所述三次扫选中的第一次扫选和第二次扫选所用的捕收剂包括丁基黄药，所述捕收剂的用量为30~120g/t；(4) The collector used in the first and second sweeps of the three sweeps includes butyl xanthate, and the amount of the collector used is 30-120 g/t; （5）所述一次粗选所用的活化剂包括硫酸铜，所述活化剂的用量为65~200g/t。(5) The activator used in the primary roughing includes copper sulfate, and the amount of the activator used is 65-200 g/t. 8.根据权利要求1所述复杂锡铅锌多金属矿分离回收方法，其特征在于，步骤（h）中，满足以下条件至少其一：8. The method for separating and recovering complex tin-lead-zinc polymetallic ores according to claim 1, characterized in that in step (h), at least one of the following conditions is met: （1）调节所述第一脱硫浮选尾矿的浓度至25wt.%~35wt.%；(1) adjusting the concentration of the first desulfurization flotation tailings to 25wt.%~35wt.%; （2）调节所述第一脱硫浮选尾矿的pH至6~7；(2) adjusting the pH of the first desulfurization flotation tailings to 6-7; （3）所述一次粗选所用的捕收剂包括肉桂异羟肟酸和P86捕收剂，所述肉桂异羟肟酸的分子式为C6H5-CH=CH–(C=O)NHOH，所述肉桂异羟肟酸的用量为40~220g/t，所述P86捕收剂的用量为60~200g/t；(3) The collector used in the primary roughing includes cinnamic hydroxamic acid and P86 collector, the molecular formula of the cinnamic hydroxamic acid is C 6 H 5 -CH=CH–(C=O)NHOH, the amount of the cinnamic hydroxamic acid is 40-220 g/t, and the amount of the P86 collector is 60-200 g/t; （4）所述三次扫选中的第一次扫选和第二次扫选所用的捕收剂包括肉桂异羟肟酸和P86捕收剂，所述肉桂异羟肟酸的分子式为C6H5-CH=CH–(C=O)NHOH，所述肉桂异羟肟酸的用量为40~220g/t，所述P86捕收剂的用量为60~200g/t；(4) The collectors used in the first and second sweeps of the three sweeps include cinnamic hydroxamic acid and P86 collector, the molecular formula of the cinnamic hydroxamic acid is C 6 H 5 -CH=CH–(C=O)NHOH, the amount of the cinnamic hydroxamic acid is 40-220 g/t, and the amount of the P86 collector is 60-200 g/t; （5）所述一次粗选所用的抑制剂包括含硅酸钠的酸溶液，所述抑制剂的用量为600~1800g/t；(5) The inhibitor used in the primary roughing includes an acid solution containing sodium silicate, and the amount of the inhibitor used is 600-1800 g/t; （6）所述两次精选中的第一次精选所用的抑制剂包括含硅酸钠的酸溶液，所述抑制剂的用量为600~1800g/t；(6) The inhibitor used in the first of the two concentrations comprises an acid solution containing sodium silicate, and the amount of the inhibitor used is 600-1800 g/t; （7）所述一次粗选所用的活化剂包括硝酸铅，所述活化剂的用量为50~80g/t；(7) The activator used in the primary roughing includes lead nitrate, and the amount of the activator used is 50-80 g/t; （8）所述磁选的磁场强度为4500~8000Gs；(8) The magnetic field strength of the magnetic separation is 4500~8000Gs; （9）所述第二摇床重选的冲程为5~20mm，冲次为260~340次/min；(9) The stroke of the second shaking table is 5-20 mm, and the stroke frequency is 260-340 times/min; （10）所述第二摇床重选在细泥摇床中进行。(10) The second table gravity selection is carried out in a fine mud table. 9.根据权利要求1所述复杂锡铅锌多金属矿分离回收方法，其特征在于，步骤（i）中，满足以下条件至少其一：9. The method for separating and recovering complex tin-lead-zinc polymetallic ores according to claim 1, characterized in that, in step (i), at least one of the following conditions is met: （1）所述第三摇床重选的冲程为11~16mm，冲次为200~300r/min；(1) The stroke of the third shaking table for gravity selection is 11-16 mm, and the stroke frequency is 200-300 r/min; （2）所述第三摇床重选在细砂摇床中进行；(2) The third shaking table gravity selection is carried out in a fine sand shaking table; （3）将经过所述五段磨矿的矿物的浓度调节至30wt.%~40wt.%；(3) adjusting the concentration of the mineral after the five-stage grinding to 30wt.%~40wt.%; （4）将经过所述五段磨矿的矿物的pH调节至6~7；(4) adjusting the pH of the mineral after the five-stage grinding to 6-7; （5）所述一次粗选所用的捕收剂包括丁基黄药，所述捕收剂的用量为30~120g/t；(5) The collector used in the primary roughing includes butyl xanthate, and the amount of the collector used is 30-120 g/t; （6）所述三次扫选中的第一次扫选和第二次扫选所用的捕收剂包括丁基黄药，所述捕收剂的用量为30~120g/t；(6) The collector used in the first and second sweeps of the three sweeps includes butyl xanthate, and the amount of the collector used is 30-120 g/t; （7）所述一次粗选所用的活化剂包括硫酸铜，所述活化剂的用量为100~200g/t；(7) The activator used in the primary roughing comprises copper sulfate, and the amount of the activator used is 100-200 g/t; （8）所述第四摇床重选的冲程为10~25mm，冲次为200~400次/min；(8) The stroke of the fourth shaking table for gravity selection is 10-25 mm, and the stroke frequency is 200-400 times/min; （9）所述第四摇床重选在中矿摇床中进行。(9) The fourth shaking table gravity separation is carried out in the middling shaking table. 10.如权利要求1~9任一项所述复杂锡铅锌多金属矿分离回收方法在选矿中的应用。10. Application of the complex tin-lead-zinc polymetallic ore separation and recovery method as described in any one of claims 1 to 9 in mineral processing.
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