Flotation method for recovering complex tin ore associated zinc-sulfur resources

Abstract

The invention discloses a flotation method for recovering complex tin ore associated zinc-sulfur resources, which realizes efficient recovery of tin polymetallic ore associated zinc-sulfur resources through the steps of raw ore grinding operation, floatable operation of zinc-sulfur and the like, analysis of zinc mineral monomer dissociation degree, zinc-sulfur separation operation, sulfur flotation operation and the like. And the influence on production caused by scaling of subsequent tin separation equipment is avoided.

Description

Technical Field

[0001] The present invention relates to the technical field of mineral processing, and in particular to a flotation method for recovering zinc and sulfur resources associated with complex tin ores. Background Art

[0002] Zinc and sulfur are common associated resources of tin polymetallic ores. The zinc and sulfur associated with tin ores are mostly sulfide minerals. Sulfide minerals account for a large proportion and are generally better in natural floatability than tin ores, which will interfere with the sorting of tin ores and affect the quality of tin concentrates. Therefore, before tin sorting, flotation technology is generally used for desulfurization, and high-grade zinc and sulfur ore resources are also comprehensively utilized.

[0003] The flotation process of traditional tin polymetallic ores often adopts "zinc-sulfur mixed flotation-mixed flotation concentrate zinc and sulfur re-separation" and "zinc and sulfur flotation in turn with priority". In order to ensure that tin separation is not affected by residual sulfide ores, a large amount of activators such as sulfuric acid and copper sulfate are added to the main process. Low-value and harmful sulfide minerals such as pyrite and arsenopyrite are separated from the ore pulp through flotation. In addition, in order to obtain high-quality zinc concentrate, a large amount of lime is added during the zinc concentrate flotation process to reduce the floatability of pyrite. For example, Chinese patent CN109926196B discloses a cyanide-free separation process for low-grade tin-lead-zinc polymetallic sulfide minerals, lead and zinc. In the zinc flotation process, whether in roughing or fine separation, lime is added as a pH adjuster to inhibit pyrite and adjust the pH to 10-12. This results in a high-alkaline, high-calcium environment in the ore pulp, which further causes scaling of equipment such as gravity separation and magnetic separation used in tin separation, thereby affecting the separation indicators of the tin concentrate. In addition, the scaled equipment requires frequent maintenance, which adds additional costs.

[0004] Therefore, the present invention aims to develop a lime-free, efficient and easy-to-implement flotation technology for the recovery of tin and sulfur resources associated with tin polymetallic ores, which, on the one hand, realizes the efficient comprehensive utilization of associated zinc and sulfur resources, and on the other hand, solves the impact of equipment scaling on tin concentrate separation. Summary of the Invention

[0005] The main purpose of the present invention is to provide a flotation method for recovering zinc and sulfur resources associated with complex tin ores, aiming to solve the technical problem of how to achieve efficient recovery of zinc and sulfur resources associated with tin and polymetallic ores in a lime-free flotation system, thereby avoiding scaling problems in subsequent tin sorting equipment that affect production.

[0006] In order to achieve the above object, the present invention provides a flotation method for recovering zinc and sulfur resources associated with complex tin ores, comprising the following steps:

[0007] S1: The raw ore is coarsely ground, slurried, pH regulator, activator, collector and frother are added, and then froth flotation is performed to obtain iso-floatable rough concentrate and iso-floatable rougher tailings;

[0008] S2: slurrying the floatable coarse concentrate and then beneficiating it to obtain zinc, sulfur and other floatable concentrates;

[0009] S3: Analyzing the grinding fineness and zinc mineral monomer dissociation degree of the zinc sulfur and other floatable concentrates, performing regrinding on the zinc sulfur and other floatable concentrates with a zinc mineral monomer dissociation degree of less than 85%; slurrying the zinc sulfur and other floatable concentrates with a zinc mineral monomer dissociation degree of 85% or more, adding a pH regulator, an adjusting agent A, an adjusting agent B, a collector and a frother, and then flotation to obtain a zinc rough concentrate and a zinc sulfur separation rougher tailings;

[0010] S4: scavenging the zinc-sulfur separation roughing tailings, adding a collector and a foaming agent before scavenging, and the scavenged tailings are sulfur concentrate I;

[0011] S5: slurrying the zinc coarse concentrate and then beneficiating the zinc concentrate to obtain zinc concentrate;

[0012] S6: scavenging the floatable rougher tailings, adding a collector and a frother before scavenging to obtain zinc-sulfur floatable tailings;

[0013] S7: adding an activator, a collector and a frother to the zinc-sulfur floatable tailings, and then performing froth flotation to obtain a sulfur rough concentrate and a sulfur rougher tailings;

[0014] S8: slurrying the crude sulfur concentrate and then beneficiating the concentrate to obtain sulfur concentrate II;

[0015] S9: The sulfur roughing tailings are subjected to froth flotation scavenging, and a collector and a frother are added before scavenging. The tailings obtained in the final scavenging step are the desulfurization products.

[0016] Preferably, in step S1, the feed fineness after coarse grinding is 60-65 wt% in the -200 mesh size fraction, the concentration after slurry adjustment is 32-36 wt%, and the pH value is adjusted to 5.5-6.0.

[0017] Preferably, the feed material is tin ore with an associated zinc grade greater than 1%, and the zinc element is present in sulfide minerals.

[0018] Preferably, in step S3, the slurry concentration of the floatable concentrate with a zinc mineral monomer dissociation degree of ≥85% is 22-25 wt %, and the pH value is adjusted to 5.0-5.5.

[0019] Preferably, in step S3: after adding the adjusting agent A, stir and react for 2 to 4 minutes; after adding the adjusting agent B, stir and react for 4 to 6 minutes; after adding the collecting agent and the frother, stir and react for 2 to 3 minutes respectively; the flotation time is 3 to 5 minutes.

[0020] Preferably, the scavenging step in S4 and / or S6 includes: scavenging for 1 to 2 stages, the flotation underflow product of each stage is the scavenging tailings, the flotation foam product of each stage is returned to the previous stage in turn, and the tailings of the last stage of scavenging are the obtained products.

[0021] Preferably, the pulp concentration of the coarsely ground ore in step S1 is controlled at 32-36 wt%; the pulp concentration of the equal floatable coarse concentrate in step S2 is controlled at 22-25 wt%; the pulp concentration of the zinc coarse concentrate in step S5 is controlled at 20-23 wt%; and the pulp concentration of the sulfur coarse concentrate in step S8 is controlled at 20-23 wt%.

[0022] Preferably, the beneficiation step in S2 and / or S5 and / or S8 includes: beneficiation is performed in 1 to 2 stages, the flotation foam product of each stage is the beneficiated concentrate, the flotation underflow products of each stage are returned to the previous stage in sequence, and the concentrate of the last stage is the obtained product.

[0023] Preferably, the adjusting agent A is sodium sulfite or sodium pyrosulfate, and the adjusting agent B is calcium polysulfide.

[0024] Preferably, the pH regulator is sulfuric acid, the activator is copper sulfate; the collector is at least one of butyl xanthate or isopentyl xanthate, and the foaming agent is No. 2 oil.

[0025] Beneficial effects:

[0026] The present invention provides a flotation method for recovering zinc and sulfur resources associated with complex tin ores. Targeting the zinc and sulfur resources associated with complex tin ores, the method employs a zinc-sulfur and other floatable processes in a lime-free, weakly acidic flotation environment, uses a small amount of activator and collector, and preferentially separates the vast majority of valuable zinc minerals and part of the pyrite from the main process to ensure the recovery rate of relatively high-value zinc concentrate. Sulfur-inhibiting zinc flotation is then used to separate high-value zinc concentrate and part of the sulfur concentrate from the zinc-sulfur and other floatable concentrates. Furthermore, an activator is employed to activate sulfur flotation to separate the remaining pyrite from the main process. The method of the present invention can effectively achieve efficient recovery of zinc and sulfur resources associated with tin polymetallic ores in a lime-free flotation system, addressing the impact of calcium buildup in traditional sulfide ore flotation on tin separation. The method is suitable for processing complex tin ores with high-value associated zinc and sulfur, thereby improving overall resource utilization.

[0027] The flotation method provided by the present invention separates zinc concentrate and sulfur concentrate in steps according to the mineral properties of the tin-zinc-sulfur ore. By adopting a lime-free, non-inhibiting, weakly acidic, and weakly activated flotation system, the method effectively avoids the influence of high-alkali and high-calcium slurries in traditional sulfide ore flotation on tin separation operations, and gives priority to floatable elements such as zinc and sulfur. Under the premise of ensuring the full recovery of valuable zinc minerals, the method minimizes the difficulty of zinc and sulfur separation and improves the product quality and utilization rate of zinc concentrate.

[0028] In the present invention, by conducting grinding fineness-zinc mineral monomer dissociation degree analysis, it can be effectively determined whether the ore needs to be re-grinded and the appropriate re-grinding fineness conditions, thereby achieving precise and effective control of zinc mineral dissociation; the inhibitors used in zinc-sulfur separation include pH regulators, activators, adjusters A / B and collectors, all of which are pure inorganic, low-dosage, easily degradable and highly selective reagent systems, which can effectively achieve effective inhibition of pyrite in zinc-sulfur and other floatable concentrates, and have no negative impact on the overall reuse of the ore pulp. BRIEF DESCRIPTION OF THE DRAWINGS

[0029] Figure 1 This is a process flow chart of Examples 1 to 5 of the present invention;

[0030] Figure 2 The present invention is a flow chart of the existing zinc-sulfur mixed flotation-mixed flotation concentrate zinc-sulfur re-separation process.

[0031] The purpose, features and advantages of the present invention will be further described with reference to the accompanying drawings and in conjunction with the embodiments. DETAILED DESCRIPTION

[0032] The present invention is further described below with reference to the examples. These examples are intended to illustrate the present invention only and are not intended to limit the scope of the present invention. Experimental methods in the following examples where specific conditions are not specified are generally performed in accordance with conventional conditions in the art or the conditions recommended by the manufacturer; the raw materials and reagents used, unless otherwise specified, are all commercially available from conventional markets. Any non-substantial changes and substitutions made by those skilled in the art based on the present invention fall within the scope of protection claimed in the present invention.

[0033] Example 1: Flotation method for recovering zinc and sulfur resources associated with complex tin ores

[0034] The original ore is a complex tin ore from Xilin Gol League, with the tin, zinc and sulfur grades of 0.54%, 1.14% and 3.40% respectively. The original zinc and sulfur separation process of the mine is "zinc and sulfur mixed flotation - mixed flotation concentrate zinc and sulfur re-separation" process (refer to Figure 2 ), the subsequent tin separation equipment operating rate is 68.47%. Figure 1 As shown, the implementation process of the process of the present invention is:

[0035] S1: By checking the classification and coarse-grain regrinding operations, the fineness of the floatable feed such as zinc and sulfur is controlled to -200 mesh size, accounting for 65wt%; the slurry is adjusted to a feed concentration of 36wt%, sulfuric acid is added to adjust the pH to 5.5, and then 120g / t copper sulfate is added to react for 3 minutes. 50g / t butyl xanthate and 20g / t No. 2 oil are added to the slurry after the activator reaction is completed, and stirred for reaction for 2 minutes respectively. Then, froth flotation is carried out to obtain floatable coarse concentrate such as flotation foam product and floatable rougher tailings such as flotation underflow product;

[0036] S2: slurrying the floatable coarse concentrate to control the pulp concentration at 25wt%, and then performing two stages of concentrating. The flotation foam product of each stage is the concentrating concentrate, and the flotation underflow product is the concentrating middlings. The concentrating middlings of each stage are sequentially returned to the previous stage. The concentrate of the second stage is the floatable concentrate of zinc, sulfur, etc.

[0037] S3: The zinc, sulfur and other floatable concentrates are analyzed for grinding fineness-zinc mineral monomer dissociation degree, and it is determined that the re-grinding fineness is -400 mesh particle size accounting for 75wt%, and the zinc mineral monomer dissociation degree is 86.39%; then slurrying is performed to control the slurry concentration at 25wt%, and sulfuric acid is added to control the pH value at 5.0; 200g / t sodium sulfite, 50g / t calcium polysulfide, 5g / t butyl xanthate and 2g / t No. 2 oil are added to the slurry with adjusted pH value in sequence, and the sodium sulfite is stirred and reacted for 2 minutes, the calcium polysulfide is stirred and reacted for 4 minutes, and the butyl xanthate and No. 2 oil are stirred and reacted for 2 minutes respectively, and then flotation is performed to obtain a flotation foam product zinc coarse concentrate and a flotation underflow product zinc-sulfur separation roughing tailings;

[0038] S4: The zinc-sulfur separation roughing tailings are subjected to two stages of scavenging, 2g / t of butyl xanthate and 1g / t of No. 2 oil are added in sequence before each stage of scavenging, the flotation foam product of each stage of scavenging is the zinc-sulfur separation scavenging ore, and the flotation underflow product is the zinc-sulfur separation scavenging tailings. The zinc-sulfur separation scavenging ore of each stage is returned to the previous stage in sequence, and the tailings of the second stage of scavenging is the sulfur concentrate I;

[0039] S5: The zinc coarse concentrate is slurried to control the slurry concentration at 23wt%, and then subjected to two stages of concentrating. The flotation froth product of each stage is the zinc-sulfur separation and concentrating concentrate, and the flotation underflow product is the zinc-sulfur separation and concentrating middlings. The zinc-sulfur separation and concentrating middlings of each stage are sequentially returned to the previous stage, and the concentrate of the second stage of concentrating is the zinc concentrate.

[0040] S6: scavenging the equal-floatable roughing tailings. Before scavenging, 20 g / t butyl xanthate and 10 g / t No. 2 oil are added and stirred for 2 minutes respectively. The flotation foam product of each stage is the equal-floatable scavenging tailings, and the flotation underflow product is the equal-floatable scavenging tailings. The scavenging tailings of each stage are returned to the previous stage in turn. The tailings of the second stage of equal-floatable scavenging are zinc-sulfur equal-floatable tailings.

[0041] S7: adding 240 g / t copper sulfate, 80 g / t butyl xanthate and 40 g / t No. 2 oil to the zinc-sulfur floatable tailings, stirring the copper sulfate for reaction for 3 minutes, stirring the butyl xanthate and No. 2 oil for reaction for 2 minutes respectively, and then performing two stages of froth flotation to obtain a sulfur rough concentrate and a sulfur rougher tailings;

[0042] S8: The sulfur crude concentrate is slurried to control the pulp concentration at 23wt%, and then subjected to two stages of concentration. The flotation foam product of each stage is the sulfur concentrated concentrate, and the flotation underflow product is the sulfur concentrated middlings. The concentrated middlings of each stage are sequentially returned to the previous stage. The concentrate of the second stage is sulfur concentrate II.

[0043] S9: The sulfur roughing tailings are subjected to two stages of foam flotation scavenging. Before scavenging, 20 g / t of butyl xanthate and 10 g / t of No. 2 oil are added and stirred for reaction for 2 minutes respectively. The foam product of each stage of scavenging flotation is the sulfur scavenging ore, and the flotation underflow product is the sulfur scavenging tailings. The scavenging ore of each stage is returned to the previous stage in turn, and the tailings of the second stage of sulfur scavenging is the desulfurization product.

[0044] The flotation indexes obtained in this example are shown in Table 1 below.

[0045] Table 1

[0046]

[0047] The zinc grade of the zinc concentrate is 45.87%, and the zinc recovery rate is 85.17%; the sulfur removal rate of the desulfurization product is 95.72%, and the tin recovery rate is 92.52%; through the implementation of this process, the operating rate of subsequent tin separation equipment is increased to 95.48%.

[0048] Example 2: Flotation method for recovering zinc and sulfur resources associated with complex tin ores

[0049] The original ore is a complex tin ore from Wenshan, Yunnan. The tin, zinc and sulfur grades of the original ore are 0.34%, 2.56% and 5.58% respectively. The original zinc and sulfur separation process of the mine is the "zinc and sulfur are preferentially floated in sequence" process. The subsequent tin separation equipment operation rate is 55.37%. Figure 1 As shown, the implementation method of the process of the present invention is similar to that of Example 1;

[0050] The difference from Example 1 is that in S1 of Example 2, the feed fineness is controlled at -200 mesh size accounting for 60wt%, the feed concentration is adjusted to 32wt%, sulfuric acid is added to adjust the pH to 6.0, 180g / t copper sulfate is added and reacted for 5min, and 180g / t isopentyl xanthate and 40g / t methyl isobutyl carbinol are added in sequence to the slurry after the activator reaction is completed, and stirred and reacted for 3min respectively;

[0051] The concentration of the equal floatable coarse concentrate slurry in S2 is controlled at 22wt%;

[0052] In S3, the zinc-sulfur and other floatable concentrates are subjected to a regrind fineness-zinc mineral monomer dissociation degree analysis, and it is determined that the regrind fineness is -400 mesh size accounting for 70wt%, and the zinc mineral monomer dissociation degree is 85.32%. The floatable concentrate having a regrind fineness of -400 mesh size accounting for 70wt% is slurried, the concentration is controlled at 22wt%, the pH value is adjusted to 5.5, and 150g / t sodium pyrosulfate, 80g / t calcium polysulfide, 10g / t isopentyl xanthate and 5g / t methyl isobutyl carbinol are added in sequence. The sodium pyrosulfate is stirred and reacted for 4 minutes, the calcium polysulfide is stirred and reacted for 5 minutes, and the isopentyl xanthate and methyl isobutyl carbinol are stirred and reacted for 3 minutes respectively.

[0053] Before scavenging in S4, add 5g / t of isopentyl xanthate and 2.5g / t of methyl isobutyl carbinol;

[0054] The slurry concentration of the zinc coarse concentrate in S5 is controlled at 20wt%;

[0055] Before scavenging in S6, add 30g / t butyl xanthate and 10g / t methyl isobutyl carbinol, and stir and react for 3 minutes respectively;

[0056] In S7, 150 g / t copper sulfate, 120 g / t butyl xanthate and 50 g / t methyl isobutyl carbinol are added to the zinc-sulfur and other floatable tailings slurry, the copper sulfate is stirred and reacted for 5 minutes, and the butyl xanthate and methyl isobutyl carbinol are stirred and reacted for 3 minutes respectively;

[0057] The slurry concentration of the sulfur concentrate in S8 is controlled at 20wt%;

[0058] Before scavenging in S9, add 30 g / t butyl xanthate and 15 g / t methyl isobutyl carbinol and stir to react for 3 minutes.

[0059] The flotation indexes obtained in this example are shown in Table 2 below.

[0060] Table 2

[0061]

[0062] The zinc grade of the zinc concentrate is 52.46%, and the zinc recovery rate is 92.08%; the sulfur removal rate of the desulfurization product is 93.90%, and the tin recovery rate is 94.98%; through the implementation of this process, the operating rate of subsequent tin separation equipment is increased to 96.74%.

[0063] Example 3: Flotation method for recovering zinc and sulfur resources associated with complex tin ores

[0064] The original ore is a complex tin ore from Gejiu, Yunnan. The tin, zinc and sulfur grades of the original ore are 0.23%, 0.72% and 2.58% respectively. The original zinc and sulfur separation process of the mine is "zinc and sulfur mixed flotation - mixed flotation concentrate zinc and sulfur re-separation" process. The subsequent tin separation equipment operation rate is 61.35%. Figure 1 As shown, the implementation method of the process of the present invention is similar to that of Example 1;

[0065] The difference from Example 1 is that in S1 of Example 3, the fineness of the feed is controlled at -200 mesh particle size, accounting for 62wt%, the slurry is adjusted to a feed concentration of 35wt%, sulfuric acid is added to adjust the pH to 5.8, 80g / t of copper sulfate is added and reacted for 4min, 60g / t of a floatable collector (butyl xanthate and isopentyl xanthate in a mass ratio of 3:1) and 20g / t of No. 2 oil are added to the slurry after the activator reaction is completed, and the mixture is stirred and reacted for 2min respectively;

[0066] The concentration of the iso-floatable coarse concentrate slurry in S2 is controlled at 24wt%;

[0067] In S3, the zinc-sulfur and other floatable concentrates are subjected to a regrind fineness-zinc mineral monomer dissociation degree analysis, and it is determined that the regrind fineness is 80 wt% of the -400 mesh size, and the zinc mineral monomer dissociation degree is 87.36%. The floatable concentrate having a regrind fineness of 80 wt% of the -400 mesh size is slurried, the concentration is controlled at 24 wt%, the pH value is adjusted to 5.2, and 250 g / t of sodium sulfite, 40 g / t of calcium polysulfide, 10 g / t of butyl xanthate and 4 g / t of No. 2 oil are added in sequence. The sodium sulfite is stirred and reacted for 3 minutes, the calcium polysulfide is stirred and reacted for 5 minutes, and the butyl xanthate and the foaming agent are stirred and reacted for 2.5 minutes respectively.

[0068] Before scavenging in S4, add 5g / t of isopentyl xanthate and 2.5g / t of No. 2 oil, and stir and react for 2.5 minutes respectively;

[0069] The slurry concentration of the zinc coarse concentrate in S5 is controlled at 22wt%;

[0070] Before scavenging in S6, add 20g / t of floatable collector (butyl xanthate and isopentyl xanthate in a mass ratio of 3:1) and 10g / t of No. 2 oil, and stir and react for 3 minutes respectively;

[0071] In S7, 120 g / t copper sulfate, 80 g / t butyl xanthate and 30 g / t No. 2 oil are added to the zinc-sulfur and other floatable tailings slurry, and the copper sulfate is stirred and reacted for 4 minutes, and the butyl xanthate and No. 2 oil are stirred and reacted for 2.5 minutes respectively;

[0072] The slurry concentration of the sulfur concentrate in S8 is controlled at 21wt%;

[0073] Before scavenging in S9, add 30g / t butyl xanthate and 15g / t No. 2 oil, and stir and react for 3 minutes respectively.

[0074] The flotation indexes obtained in this example are shown in Table 3 below.

[0075] Table 3

[0076]

[0077] The zinc grade of the zinc concentrate is 43.66%, and the zinc recovery rate is 81.78%; the sulfur removal rate of the desulfurization product is 95.64%, and the tin recovery rate is 93.09%; through the implementation of this process, the operating rate of subsequent tin separation equipment is increased to 98.37%.

[0078] Example 4: Flotation method for recovering zinc and sulfur resources associated with complex tin ores

[0079] The original ore is a complex tin ore from Nandan, with tin, zinc and sulfur grades of 0.27%, 2.08% and 3.14% respectively. The original zinc and sulfur separation process of the mine is the "zinc and sulfur flotation first" process, and the subsequent tin separation equipment operation rate is 59.64%. Figure 1 As shown, the implementation method of the process of the present invention is similar to that of Example 1;

[0080] The difference from Example 1 is that in S1 of Example 4, the fineness of the feed is controlled at -200 mesh particle size, accounting for 63wt%, the slurry is adjusted to a feed concentration of 33wt%, sulfuric acid is added to adjust the pH to 5.7, and then 120g / t copper sulfate is added to react for 5min. 100g / t of a floatable collector (butyl xanthate and isopentyl xanthate in a mass ratio of 4:1) and 40g / t of a foaming agent (No. 2 oil and methyl isobutyl carbinol in a mass ratio of 2:1) are sequentially added to the slurry after the activator reaction is completed, and the mixture is stirred and reacted for 3min respectively.

[0081] The concentration of the iso-floatable coarse concentrate slurry in S2 is controlled at 24wt%;

[0082] In S3, the zinc-sulfur and other floatable concentrates are subjected to a regrind fineness-zinc mineral monomer dissociation degree analysis, and it is determined that the regrind fineness is 82 wt% of the -400 mesh size, and the zinc mineral monomer dissociation degree is 85.37%. The floatable concentrate having a regrind fineness of 82 wt% of the -400 mesh size is slurried, the concentration is controlled at 23 wt%, the pH value is adjusted to 5.3, and 300 g / t of sodium sulfite, 80 g / t of calcium polysulfide, 30 g / t of isoamyl xanthate and 10 g / t of frother are added in sequence. The sodium sulfite is stirred and reacted for 4 minutes, the calcium polysulfide is stirred and reacted for 6 minutes, and the isoamyl xanthate and frother are stirred and reacted for 2 minutes respectively.

[0083] Before scavenging in S4, add 10g / t of isopentyl xanthate and 5g / t of foaming agent, and stir and react for 3 minutes respectively;

[0084] The slurry concentration of the zinc coarse concentrate in S5 is controlled at 21wt%;

[0085] Before scavenging in S6, add 20g / t of floatable collector and 10g / t of foaming agent, and stir and react for 3 minutes respectively;

[0086] In S7, 80 g / t copper sulfate, 100 g / t butyl xanthate and 30 g / t foaming agent are added to the zinc-sulfur and other floatable tailings slurry, and the copper sulfate is stirred and reacted for 4 minutes, and the butyl xanthate and foaming agent are stirred and reacted for 3 minutes respectively;

[0087] The slurry concentration of the sulfur concentrate in S8 is controlled at 21wt%;

[0088] Before scavenging in S9, add 30g / t butyl xanthate and 15g / t foaming agent, and stir and react for 3 minutes respectively.

[0089] The flotation indexes obtained in this example are shown in Table 4 below.

[0090] Table 4

[0091]

[0092]

[0093] The zinc grade of the zinc concentrate is 48.26%, and the zinc recovery rate is 89.56%; the sulfur removal rate of the desulfurization product is 95.67%, and the tin recovery rate is 84.60%; through the implementation of this process, the operating rate of subsequent tin separation equipment is increased to 94.71%.

[0094] Example 5: Flotation method for recovering zinc and sulfur resources associated with complex tin ores

[0095] The original ore is a complex tin ore from Nandan, with tin, zinc and sulfur grades of 1.74%, 1.23% and 2.45% respectively. The original zinc and sulfur separation process of the mine is "zinc and sulfur mixed flotation - mixed flotation concentrate zinc and sulfur re-separation" process, and the subsequent tin separation equipment operation rate is 72.94%. Figure 1 As shown, the implementation method of the process of the present invention is similar to that of Example 1;

[0096] The difference from Example 1 is that in S1 of Example 5, the feed fineness is controlled at -200 mesh size, accounting for 63wt%, the feed concentration is adjusted to 32wt%, sulfuric acid is added to adjust the pH to 5.8, 90g / t copper sulfate is added and reacted for 4 minutes, 100g / t butyl xanthate and 40g / t No. 2 oil are added to the slurry after the activator reaction is completed, and the mixture is stirred and reacted for 3 minutes respectively;

[0097] The concentration of the equal floatable coarse concentrate slurry in S2 is controlled at 23wt%;

[0098] In S3, the zinc-sulfur and other floatable concentrates are subjected to a regrind fineness-zinc mineral monomer dissociation degree analysis, and it is determined that the regrind fineness is -400 mesh size accounting for 82wt%, and the zinc mineral monomer dissociation degree is 85.37%. The floatable concentrate having a regrind fineness of -400 mesh size accounting for 82wt% is slurried, the concentration is controlled at 23wt%, the pH value is adjusted to 5.3, and 150g / t sodium sulfite, 50g / t calcium polysulfide, 30g / t butyl xanthate and 10g / t No. 2 oil are added in sequence. The sodium sulfite is stirred and reacted for 3 minutes, the calcium polysulfide is stirred and reacted for 5 minutes, and the isopentyl xanthate and No. 2 oil are stirred and reacted for 2 minutes respectively.

[0099] Before scavenging in S4, add 10g / t of isopentyl xanthate and 5g / t of No. 2 oil, and stir and react for 3 minutes respectively;

[0100] The slurry concentration of the zinc coarse concentrate in S5 is controlled at 21wt%;

[0101] Before scavenging in S6, add 20g / t butyl xanthate and 10g / t No. 2 oil, and stir and react for 3 minutes respectively;

[0102] In S7, 60g / t copper sulfate, 60g / t butyl xanthate and 30g / t No. 2 oil are added to the zinc-sulfur and other floatable tailings slurry, and the copper sulfate is stirred and reacted for 3 minutes, and the butyl xanthate and No. 2 oil are stirred and reacted for 2 minutes respectively;

[0103] The slurry concentration of the sulfur concentrate in S8 is controlled at 21wt%;

[0104] Before scavenging in S9, add 30g / t butyl xanthate and 15g / t No. 2 oil, and stir and react for 2 minutes respectively.

[0105] The flotation indices obtained in this example are shown in Table 5 below.

[0106] Table 5

[0107]

[0108] The zinc grade of the zinc concentrate is 46.32%, and the zinc recovery rate is 91.39%; the sulfur removal rate of the desulfurization product is 92.75%, and the tin recovery rate is 98.48%; through the implementation of this process, the operating rate of subsequent tin separation equipment is increased to 98.61%.

[0109] Comparative Experiment 1

[0110] The research object and implementation method of this comparative experiment are basically the same as those of Example 1. The difference lies in the difference in the fineness of the regrinding of zinc and sulfur floatable concentrates, which leads to differences in the zinc mineral dissociation degree and zinc concentrate separation index. The specific results are shown in Table 6.

[0111] Table 6 Changes in dissociation degree and separation index of zinc concentrate under different regrinding fineness

[0112]

[0113] Comparative Experiment 2

[0114] The research object and implementation method of this comparative experiment are basically the same as those of the embodiment, except that calcium polysulfide is not added in the zinc-sulfur separation operation, resulting in differences in the zinc concentrate separation indicators. The specific results are shown in Table 7.

[0115] Table 7 Effect of calcium polysulfide addition on zinc concentrate separation indexes

[0116]

[0117] From the above comparative experimental results, it can be seen that: the zinc mineral dissociation degree is divided into 85%, and the zinc-sulfur and other floatable concentrates with a fineness of -400 mesh size below 85% have a low proportion and still need to be re-grinded. The zinc-sulfur and other floatable concentrates with a mineral monomer dissociation degree ≥85% are subjected to subsequent slurry flotation operations, which are beneficial to the separation of zinc and sulfur. The grade and recovery rate of the obtained concentrates are relatively high; and the main function of calcium polysulfide is "selective sulfur suppression", which improves the grade and recovery rate of zinc ore and has almost no effect on tin ore.

[0118] Finally, it should be noted that the above description of the embodiments is intended to facilitate understanding and use of the invention by those skilled in the art. Those skilled in the art will readily be able to make various modifications to these embodiments and apply the general principles described herein to other embodiments without requiring inventive effort. Therefore, the present invention is not limited to the above-described embodiments. Improvements and modifications made by those skilled in the art based on the disclosure of the present invention, without departing from the scope of the present invention, should be within the scope of protection of the present invention.

Claims

1. A flotation method for recovering zinc and sulfur resources associated with complex tin ores, characterized in that: The following steps are involved: S1: The raw ore is coarsely ground, slurried, pH regulator, activator, collector and frother are added, and then froth flotation is performed to obtain iso-floatable rough concentrate and iso-floatable rougher tailings; S2: slurrying the floatable coarse concentrate and then beneficiating it to obtain zinc, sulfur and other floatable concentrates; S3: Analyzing the grinding fineness and zinc mineral monomer dissociation degree of the zinc sulfur and other floatable concentrates, performing regrinding on the zinc sulfur and other floatable concentrates with a zinc mineral monomer dissociation degree of less than 85%; slurrying the zinc sulfur and other floatable concentrates with a zinc mineral monomer dissociation degree of 85% or more, adding a pH regulator, an adjusting agent A, an adjusting agent B, a collector and a frother, and then flotation to obtain a zinc rough concentrate and a zinc sulfur separation rougher tailings; S4: scavenging the zinc-sulfur separation roughing tailings, adding a collector and a foaming agent before scavenging, and the scavenged tailings are sulfur concentrate I; S5: slurrying the zinc coarse concentrate and then beneficiating the zinc concentrate to obtain zinc concentrate; S6: scavenging the floatable rougher tailings, adding a collector and a frother before scavenging to obtain zinc-sulfur floatable tailings; S7: adding an activator, a collector and a frother to the zinc-sulfur floatable tailings, and then performing froth flotation to obtain a sulfur rough concentrate and a sulfur rougher tailings; S8: slurrying the crude sulfur concentrate and then beneficiating the concentrate to obtain sulfur concentrate II; S9: The sulfur roughing tailings are subjected to froth flotation scavenging, and a collector and a frother are added before scavenging. The tailings obtained in the final scavenging step are the desulfurization products.

2. The flotation method for recovering zinc and sulfur resources associated with complex tin ores according to claim 1, characterized in that: In step S1, the feed fineness after coarse grinding is 60-65 wt% in the -200 mesh size, the concentration after slurry adjustment is 32-36 wt%, and the pH value is adjusted to 5.5-6.

0.

3. The flotation method for recovering zinc and sulfur resources associated with complex tin ores according to claim 1, characterized in that: The feed material is tin ore with an associated zinc grade greater than 1%, and the zinc element is present in sulfide minerals.

4. The flotation method for recovering zinc and sulfur resources associated with complex tin ores according to claim 1, characterized in that: In step S3, the slurry concentration of the floatable concentrate with a zinc mineral monomer dissociation degree of ≥85% is 22-25wt%, and the pH value is adjusted to 5.0-5.

5.

5. The flotation method for recovering zinc and sulfur resources associated with complex tin ores according to claim 1, characterized in that: In step S3: after adding the adjusting agent A, stir and react for 2 to 4 minutes; after adding the adjusting agent B, stir and react for 4 to 6 minutes; after adding the collecting agent and the frother, stir and react for 2 to 3 minutes respectively; the flotation time is 3 to 5 minutes.

6. The flotation method for recovering zinc and sulfur resources associated with complex tin ores according to claim 1, characterized in that: The scavenging step in S4 and / or S6 includes: scavenging for 1 to 2 stages, each stage of flotation underflow product is its scavenging tailings, each stage of scavenging flotation foam product is returned to the previous stage in turn, and the tailings of the last stage of scavenging are the obtained products.

7. The flotation method for recovering zinc and sulfur resources associated with complex tin ores according to claim 1, characterized in that: The pulp concentration of the coarsely ground ore in step S1 is controlled at 32-36wt%; the pulp concentration of the floatable coarse concentrate in step S2 is controlled at 22-25wt%; the pulp concentration of the zinc coarse concentrate in step S5 is controlled at 20-23wt%; the pulp concentration of the sulfur coarse concentrate in step S8 is controlled at 20-23wt%.

8. The flotation method for recovering zinc and sulfur resources associated with complex tin ores according to claim 1, characterized in that: The beneficiation step in S2 and / or S5 and / or S8 includes: beneficiation for 1 to 2 stages, each stage of flotation foam product is the beneficiated concentrate, each stage of flotation underflow product is sequentially returned to the previous stage, and the last stage of beneficiation concentrate is the obtained product.

9. The flotation method for recovering zinc and sulfur resources associated with complex tin ores according to any one of claims 1 to 8, characterized in that: The adjusting agent A is sodium sulfite or sodium pyrosulfate, and the adjusting agent B is calcium polysulfide.

10. The flotation method for recovering zinc and sulfur resources associated with complex tin ores according to any one of claims 1 to 8, characterized in that: The pH regulator is sulfuric acid, and the activator is copper sulfate; The collector is at least one of butyl xanthate and isoamyl xanthate, and the foaming agent is at least one of No. 2 oil and methyl isobutyl carbinol.

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