A polymetallic tin placer beneficiation process

Abstract

The application discloses a multi-metallic sand tin ore beneficiation process, adopts the equal floatation-mixed floatation-gravity separation backwater process, lead concentrate yield is 0.39%, lead grade is 16.62%, recovery rate is 15.93%; zinc concentrate yield is 1.08%, zinc grade is 39.32%, zinc recovery rate is 46.92%, the gravity separation indexes are as follows: tin concentrate yield is 0.39%, tin content is 40.57%, tin recovery rate is 51.66%; low-grade tin concentrate yield is 0.51%, tin content is 3.17%, tin recovery rate is 5.28%. The beneficiation process recovers lead, zinc and tin in the sand tin ore, has high beneficiation indexes and good beneficiation effect.

Description

TECHNICAL FIELD

[0001] The present application belongs to the field of mineral processing, and particularly relates to a multi-metallic tin placer ore processing technology. BACKGROUND

[0002] The tin-lead-zinc ore is a multi-metallic sulfide ore, and in addition to lead, zinc, sulfur and tin, it also contains silver, cadmium, germanium and other rare and precious metals. The main metallic minerals are galena, marmatite, pyrite, pyrrhotite and cassiterite. For this type of ore, the current flotation process mainly has two types:

[0003] 1. Lead-zinc floatable-zinc-sulfur mixed flotation separation process

[0004] The method comprises the following steps: grinding the raw ore to a fineness of 60%-75% of 74 microns, regrinding the middlings, under the condition of not adjusting the pH of the ore slurry, floating the lead sulfide minerals and part of the zinc sulfide minerals similar to the floatability of the lead sulfide minerals with xanthate or black drug to obtain a lead-zinc mixed concentrate, then, using lime and zinc sulfate method to float and separate the lead concentrate and the zinc concentrate, the tailings are activated with copper sulfate, and then mixed with xanthate to obtain a zinc-sulfur mixed concentrate, and then the zinc concentrate and the sulfur concentrate are obtained by using lime and cyanide for floatation separation, and the tailings after the mixed flotation are subjected to gravity separation to recover tin metal. The process has the advantages of considering the recovery of lead, zinc, silver and tin, and the disadvantages are that the separation efficiency between lead, zinc and sulfur is low, the zinc content in the lead concentrate is high, the grade of the zinc concentrate is low, the zinc recovery rate is low, and the process uses cyanide to pollute the environment.

[0005] 2. Fine grinding high alkalinity preferential flotation process

[0006] In order to improve the lead concentrate recovery rate and the zinc concentrate index, the ore is finely ground to more than 80% of 74 microns, the lead mineral is floated by using a strong collector high alkalinity process, a sufficient amount of lime and xanthate is added, the newly dissociated pyrite is strongly inhibited under the high alkalinity condition of the ore slurry pH>12, the surface of the newly dissociated galena is protected by xanthate, and the lead is preferentially floated and recovered; the tailings after the flotation are activated with copper sulfate, and then the zinc sulfide rough concentrate is recovered by using butyl xanthate, and then the rough concentrate is cleaned under the high alkalinity condition to obtain the zinc concentrate; the tailings after the zinc flotation are thickened and activated with sulfuric acid, and then the sulfur concentrate is recovered by using xanthate; and finally, the tailings are subjected to gravity separation to recover tin metal. The process has the advantages of high grade of the lead concentrate and the zinc concentrate, and good zinc recovery rate; and the disadvantages are low lead and silver recovery rates and low tin recovery rate.

[0007] In order to comprehensively improve the multi-metallic beneficiation recovery efficiency of the tin placer ore, the present application is proposed. SUMMARY

[0008] The technical problem to be solved by the present application is to overcome the shortcomings of the prior art, and to provide a multi-metallic tin placer ore processing technology. In order to solve the above technical problems, the basic idea of the technical solution of the present application is:

[0009] A polymetallic tin placer beneficiation process, comprising the following steps:

[0010] Step 1 grinding classification: the raw ore of-2mm is subjected to closed-circuit grinding classification to obtain-0.2mm ore pulp; wherein the metallic minerals in the raw ore are sphalerite, stibnite, galena, cassiterite, arsenopyrite, pyrite and limonite, and the gangue minerals mainly are calcite and quartz; in the raw ore, the grade of tin is 0.30%, the grade of lead is 0.40%, the grade of zinc is 0.90%, and the grade of silver is 31.11g / t; the closed-circuit grinding classification in step 1 adopts a combination of a ball mill and a hydrocyclone to form a closed-circuit process for grinding classification.

[0011] Step 2 equal floatation: 2 # Step 2 equal floatation: 2

[0012] Step 3 lead mixed flotation: the lead-zinc-sulfur mixed concentrate is subjected to flotation separation through one roughing and two cleanings and two scavengings to obtain lead concentrate and zinc-sulfur tailings; when the dosages of the reagents are 208g / t of sodium humate, 890g / t of calcium oxide, 332g / t of zinc sulfate, 332g / t of sodium sulfite and 312g / t of JFR-1, the lead recovery rate is the highest; in step 3, the yield of lead concentrate is 0.34%, the grade of lead contained in the lead concentrate is 18.53%, the recovery rate of lead is 15.68%, and the lead concentrate contains 1421g / t of silver.

[0013] Step 4 re-concentration of primary tailings: the primary tailings in step 2 are subjected to one roughing and two cleanings and two scavengings to obtain flotation tailings and secondary zinc-sulfur concentrate; the dosages of the reagents are 5110g / t of sulfuric acid, 348g / t of copper sulfate, 348g / t of butyl xanthate and 41g / t of 2# oil;

[0014] Step 5 zinc-sulfur separation: the secondary zinc-sulfur concentrate and the zinc-sulfur tailings are combined and subjected to one roughing and two cleanings and two scavengings to obtain zinc concentrate and sulfur concentrate; the dosages of the reagents are 120g / t of sodium humate, 7890g / t of calcium oxide, 143g / t of copper sulfate, 61g / t of butyl xanthate and 27g / t of 2# oil; in step 5, the yield of zinc concentrate is 1.11%, the grade of zinc contained in the zinc concentrate is 41.31%, the recovery rate of zinc is 51.10%, and the zinc concentrate contains 217g / t of silver.

[0015] Step 6 reselection: the tailings are screened to obtain +0.074mm material, -0.074mm+0.038mm material and -0.038mm material, the +0.074mm material is subjected to two times of table reselection, the -0.074mm+0.038mm material is subjected to one time of table reselection, the middlings are regrinded and reselected, the -0.038mm material is subjected to reselection by one time of suspension disc concentrator, one time of table reselection, to obtain tin concentrate, secondary tin concentrate and tailings, wherein the yield of the medium tin concentrate is 0.39%, the tin content is 40.57%, and the tin recovery rate is 51.66%; the yield of the low-grade tin concentrate is 0.51%, the tin content is 3.17%, and the tin recovery rate is 5.28%.

[0016] The -2mm raw ore is subjected to closed-circuit grinding and grading to obtain -0.2mm ore slurry, wherein the metal minerals in the raw ore are sphalerite, stibnite, galena, cassiterite, arsenopyrite, pyrite and limonite, and the gangue minerals mainly include calcite and quartz; the elements of tin, lead and zinc in the ore have industrial recovery value, and the elements of silver and sulfur have comprehensive recovery value; the grade of lead in the raw ore is, the grade of zinc is, and the grade of tin is, 2# oil, sulfuric acid, lead nitrate and butyl xanthate are added into the -0.2mm raw ore slurry to perform flotation, one roughing, one cleaning and two scavenging are performed to obtain lead-zinc mixed concentrate, the lead-zinc mixed concentrate is separated by mixed flotation, the sulfur minerals and tailings are separated by re-flotation, and the equal-flotation closed-circuit test obtains lead concentrate with a yield of 0.34%, a lead grade of 18.53% and a recovery rate of 15.68%; zinc concentrate with a yield of 1.11%, a zinc grade of 41.31% and a recovery rate of 51.10%.

[0017] The tailings are added into the suspension disc concentrator to perform reselection to obtain tin concentrate with a yield of 0.39%, a tin content of 40.57% and a tin recovery rate of 51.66%; low-grade tin concentrate with a yield of 0.51%, a tin content of 3.17% and a tin recovery rate of 5.28%.

[0018] Compared with the prior art, the present application has the following beneficial effects.

[0019] The present application adopts the equal-flotation-mixed flotation process: the yield of lead concentrate is 0.34%, the lead grade is 18.53%, the recovery rate is 15.68%, and the lead concentrate contains silver of 1421g / t; the yield of zinc concentrate is 1.11%, the zinc grade is 41.31%, the zinc recovery rate is 51.10%, and the zinc concentrate contains silver of 217g / t; the equal-flotation-mixed flotation water process: the yield of lead concentrate is 0.39%, the lead grade is 16.62%, the recovery rate is 15.93%; the yield of zinc concentrate is 1.08%, the zinc grade is 39.32%, and the zinc recovery rate is 46.92%. The reselection indexes are as follows: the yield of tin concentrate is 0.39%, the tin content is 40.57%, and the tin recovery rate is 51.66%; the yield of low-grade tin concentrate is 0.51%, the tin content is 3.17%, and the tin recovery rate is 5.28%.

[0020] The specific embodiments of the present application will be further described in detail below with reference to the accompanying drawings. BRIEF DESCRIPTION OF DRAWINGS

[0021] The accompanying drawings, which are incorporated in and constitute a part of this specification, illustrate exemplary embodiments of the application and together with the description, serve to explain the application. Obviously, the drawings described below are only some embodiments of the present application, and other drawings can be obtained by those skilled in the art without creative labor. In the drawings:

[0022] Figure 1 is a process flow diagram of the present application's sink-float process;

[0023] Figure 2 is a process flow diagram of the present application's gravity separation process.

[0024] It should be noted that these drawings and written descriptions are not intended to limit the scope of the concept of the present application in any way, but to illustrate the concept of the present application to those skilled in the art by referring to specific embodiments. DETAILED DESCRIPTION

[0025] In order to make the purpose, technical scheme and advantages of the embodiments of the present application more clear, the technical scheme in the embodiments will be described clearly and completely below with reference to the drawings in the embodiments of the present application. The following embodiments are used to illustrate the present application, but not to limit the scope of the present application.

[0026] Example 1

[0027] The Dachang mining area is an old mining area with a production scale of more than sixty years, and the mining of underground rich ore and high-grade cassiterite resources has basically ended. In order to extend the life of the mine and ensure the stable development of the enterprise, resource reserves exploration and verification are carried out on five cassiterite mining areas, namely Suishuwan, Lengshuichong, Laochangpo-Team, Shaping and Hongtang, which contain a large amount of low-grade cassiterite resources, and beneficiation process exploration test research is carried out to comprehensively recover valuable metals such as tin, lead, zinc and silver.

[0028] The core samples are taken from the core samples of Suishuwan and Lengshuichong mining areas, and the core samples are classified into high-grade (containing tin 0.35% or more), medium-grade (containing tin 0.35% to 0.15%) and low-grade (containing tin 0.15% or less) three ore samples, which are respectively dried and crushed to 2mm particle size, respectively stacked, weighed, sampled and tested. According to the test results of the three ore samples, a total sample with tin grade of about 0.3-0.35% is prepared, and the total weight is 850kg. The total sample is mixed and divided to separate the raw ore multi-element analysis sample, rock identification, particle size analysis sample and total backup sample, and the test raw ore sample is 800g per part.

[0029] The results of the chemical multi-element analysis of the raw ore are shown in Table 1, the results of the mineral composition analysis of the raw ore are shown in Table 2, the results of the main element phase analysis of the raw ore are shown in Table 3, and the results of the raw ore analysis of the particle size of 2 mm are shown in Table 4.

[0030] Table 1: Results of the chemical multi-element analysis of the raw ore (%)

[0031] Element Sn Pb Zn Sb S As C Content 0.30 0.40 0.90 0.33 3.04 0.29 2.22 Element Fe SiO2 Ca0 In Ag Content 7.03 50.79 5.10 0 31.11

[0032] Ag is in g / t.

[0033] Table 2: Results of the mineral composition analysis of the raw ore (%)

[0034] Mineral name Cassiterite Zinc blende Jamesite Galena Mimetite Arsenopyrite Content 0.35 1.34 0.08 0.08 0.48 0.63 Mineral name Pyrrhotite Pyrite Calcite Quartz Others Total Content 3.86 13.69 9.11 50.79 19.59 100.00

[0035] Table 3: Results of the main element phase analysis of the raw ore (%)

[0036]

[0037] From the phase analysis results, it can be seen that: 1. The proportion of lead sulfide in the lead minerals is only 22.22%, and the proportion of lead oxide and lead-iron-vanadium is 77.78%.

[0038] 2. The proportion of zinc sulfide in the zinc minerals is 64.44%, and the proportion of zinc oxide is 30.00%.

[0039] Table 4: Results of the raw ore analysis of the particle size of 2 mm (%)

[0040]

[0041] From the raw ore analysis results of the particle size of 2 mm, it can be seen that the content of -0.074 mm in the raw ore is 51.24%, and the content of -0.019 mm is 31.70%. The ore sample has a high mud content, which increases the difficulty of separation.

[0042] Summary:

[0043] 1) The mineral composition of the sand tin ore in the Dachang mining area is complex, the grade is low, the oxidation degree is high, and the separation is difficult. The main metal minerals in the ore are sphalerite, stibnite, galena, cassiterite, arsenopyrite, pyrite, and limonite, and the gangue minerals are mainly calcite and quartz. The elements of tin, lead, and zinc in the ore have industrial recycling value, and the elements of silver and sulfur have comprehensive recycling value.

[0044] 2) Zinc exists in the form of sphalerite, which is black, irregular granular particles, and has a brittle conchoidal fracture and is not transparent. The proportion of zinc sulfide in the zinc minerals is 64.44%.

[0045] 3) Antimony exists in the form of stibnite, which is black, dark gray, and exists as a single body, and is mainly distributed in fine particle size.

[0046] 4) Lead mainly occurs in the form of galena, lead fahlore, etc. Galena is distributed in all particle sizes, with yellow-brown color, round corners, non-magnetic particles, and shell-like fracture. The proportion of lead sulfide in lead minerals is only 22.22%, and the proportion of lead oxide and lead iron vanadium is 77.78%. Therefore, the recovery of lead minerals is very difficult.

[0047] 5) The content of -0.074 mm in -2 mm raw ore is as high as 51.24%, and the content of -0.019 mm is as high as 31.70%. The ore sample contains a lot of mud, which increases the difficulty of useful metal recovery and separation.

[0048] 6) The test ore sample of this sand tin ore is not the original sand tin ore. The high content of -0.074 mm in the -2 mm screen analysis result and the data that the proportion of lead oxide and lead iron vanadium in the ore sample is as high as 77.78% confirm that part of the minerals in the core sample are tailings after separation.

[0049] According to the research results of the properties of the raw ore and the beneficiation technology of tin, lead and zinc ore in Dachang mine area, the test scheme is as follows: taking the recovery of tin, lead and zinc as the main purpose, and comprehensively recovering silver and sulfur. The test is planned to be carried out in the principle of "gravity-flotation-gravity" and "flotation-gravity", and the recovery of tin, lead, zinc, silver and sulfur is planned. The test is planned to be carried out in the principle of "gravity-flotation-gravity" and "flotation-gravity", and the recovery of tin, lead, zinc, silver and sulfur is planned.

[0050] (1) Mixed flotation process: explore and determine the flotation size: -2 mm raw ore is pre-screened and ground to -0.3 mm, -0.2 mm, and mixed flotation test of sulfide ore is carried out to determine the size of the selected particles; carry out desliming condition test; carry out condition test of flotation reagent; under the best conditions, carry out closed circuit test to obtain lead concentrate, zinc concentrate, sulfur concentrate and flotation tailings; the flotation tailings are gravity separated to obtain tin concentrate.

[0051] (2) Equal floatation-mixed floatation process: the best conditions of each reagent in the mixed floatation process are used as the reference for the equal floatation-mixed floatation process, and the best reagent conditions are obtained after the closed circuit test to obtain lead concentrate, zinc concentrate, sulfur concentrate and flotation tailings; the flotation tailings are gravity separated to obtain tin concentrate.

[0052] Because the ore sample used in this test is a core sample of sand tin ore resource exploration, the weight of the prepared sample is too small to meet the requirements of jigging and discarding test, so the jigging and discarding test cannot be carried out. Therefore, the pre-gravity discarding test only carried out spiral discarding test.

[0053] Test conditions: -2 mm raw ore is pre-deslimed using spiral classifier, the desliming results are shown in Table 5; the returned sand is separated by spiral chute to obtain spiral chute rough concentrate, middlings and tailings; the spiral chute tailings are deslimed with 0.074 mm screen. The spiral chute condition test results are shown in Table 6.

[0054] Table 5 Pre-desliming results of spiral classifier (%)

[0055]

[0056] Table 6 Spiral chute test results (%)

[0057]

[0058]

[0059] The spiral chute test results show that: ① When the pulp concentration is 25% and the processing capacity is 286 Kg / h, the operating recovery rates of tin, lead and zinc are 23.56%, 23.46% and 33.89% respectively, and the loss rate is the lowest. Therefore, this condition is used as the parameter for feeding. ② The grade ranges of tin, lead and zinc in the tailings of each group of spiral chute test are between 0.13-0.16%, 0.17-0.22% and 0.56-0.66% respectively, which are relatively high compared with the original ore grade.

[0060] Feeding test conditions: -2 mm raw ore is pre-deslimed by spiral classifier, and the returned sand is separated by spiral chute. The feeding concentration is 25%, and the processing capacity is 286 Kg / h.

[0061] The spiral chute rough concentrate, middlings and tailings are obtained. The tailings are deslimed by a 0.074 mm sieve, and the sieve is the final tailings, and the sieve is the mud. The feeding test results are shown in Table 7.

[0062] Table 7 Spiral chute feeding test results (%)

[0063]

[0064] 1) -2 mm raw ore is pre-deslimed by spiral classifier, and the returned sand is separated by spiral chute. The spiral chute separation best feeding condition concentration is 25%, and the processing capacity is 286 Kg / h. After one separation, the discarded tailings yield of the raw ore is 19.96%, the tin, lead and zinc grades are 0.14%, 0.22% and 0.51% respectively, and the tin, lead and zinc loss rates are 6.14%, 14.67% and 9.18%.

[0065] 2) From the spiral chute discarded feeding test results, the spiral chute has a certain effect on the discarded raw ore, but the tailings grade of the discarded spiral chute is relatively high compared with the original ore tin, lead and zinc grades of 0.30%, 0.40% and 0.90% respectively, especially the lead loss is large. Therefore, the front heavy uses the spiral chute equipment as the discarded equipment, which has a lower feasibility in production.

[0066] 3) According to the spiral chute discarded test results, the "flotation-heavy" principle process is used in this test flotation process.

[0067] The analysis results of the ore grinding to -0.2 mm particle size show that the content of -0.074 mm in the ore sample is 88.29%, and the content of -0.010 mm is 31.10%, and the ore sample has high mud content. The mud has a great influence on the flotation. The specific gravity of the ore sample is measured, and the settling time of -0.010 mm is calculated as the final desliming particle size; the settling time and the number of extraction are fixed as the test condition variables. The desliming test results of the number of extraction are shown in Table 8.

[0068] Table 8 Desliming test results (%)

[0069]

[0070] The test results show that: (1) under the same reagent conditions, the lead and zinc recovery rate indicators are not much different; the lead grade in the slime is the same as the original ore grade, and the zinc grade is slightly lower than the original ore grade;

[0071] (2) with the increase of the number of extraction, the slime yield and the lead and zinc loss rate also increase; for every 1% increase in yield, the lead loss is 1% recovery rate, and the zinc loss is 0.76% recovery rate;

[0072] (3) considering the lead and zinc loss rate and the influence of slime on the selection, the desliming yield is controlled at about 5% for the best.

[0073] The test results of the sulfuric acid reagent conditions from -0.3 mm and -0.2 mm selection particle size show that the best dosage of sulfuric acid reagent is 4290 g / t. The verification test of the sulfuric acid dosage for -0.2 mm selection particle size is carried out, the sulfuric acid dosage verification test results are shown in Table 9, and the open-circuit test process is shown in Figure 1 .

[0074] The open-circuit test results of the sulfuric acid dosage verification show that when the sulfuric acid dosage is 4220 g / t, the lead and zinc recovery rate is the highest, which is consistent with the best dosage of sulfuric acid reagent of 4290 g / t from the test of -0.3 mm and -0.2 mm selection particle size.

[0075] Table 9 Sulfuric acid dosage test results (%)

[0076]

[0077] The test results of the sulfuric acid copper dosage are shown in Table 10.

[0078] The test results of the sulfuric acid copper dosage show that when the sulfuric acid copper dosage is 735 g / t, the lead and zinc recovery rate reaches the peak.

[0079] Table 10 Sulfuric acid copper dosage test results (%)

[0080]

[0081] In order to improve the lead recovery rate, lead nitrate open-circuit test was carried out, and the results showed that the lead recovery rate was not obviously improved when the dosage of lead nitrate was between 484 g / t and 630 g / t, and the lead nitrate did not play a role in activating the lead oxide. The results of the lead nitrate open-circuit test are shown in Table 11.

[0082] Table 11 Test results of lead nitrate dosage (%)

[0083]

[0084]

[0085] Based on the dosage test of sulfuric acid and copper sulfate, dosage tests of butyl xanthate and combination of butyl xanthate and butyl ammonium black medicine were carried out.

[0086] 1. The test results of butyl xanthate dosage are shown in Table 12. The test results of butyl xanthate dosage show that the better dosage of butyl xanthate is 573 g / t.

[0087] Table 12 Test results of butyl xanthate dosage (%)

[0088]

[0089] 2. Combination dosage test of butyl xanthate and butyl ammonium black medicine. The test results are shown in Table 4-16. The test results of combination dosage of butyl xanthate and butyl ammonium black medicine show that the better dosage of butyl xanthate is 692 g / t and the better dosage of butyl ammonium black medicine is 104 g / t.

[0090]

[0091]

[0092] The test results of sodium sulfide dosage are shown in Table 13. The test results of sodium sulfide dosage show that the better dosage of sodium sulfide is 593 g / t.

[0093] Table 13 Test results of sodium sulfide dosage (%)

[0094]

[0095] The conditions of open-circuit verification test of the best combination of various medicines are: the dosage of sulfuric acid is 4550 g / t, the dosage of copper sulfate is 818 g / t, the dosage of sodium sulfide is 585 g / t, the dosage of butyl xanthate is 818 g / t, the dosage of butyl ammonium black medicine is 65 g / t, and the dosage of 2# oil is 85 g / t. The test results are shown in Table 14.

[0096] Table 14 Test results of open-circuit test of the best combination of medicines (%)

[0097]

[0098]

[0099] The reagent conditions for the fixed mixed-flotation process are: sulfuric acid dosage of 4550 g / t, copper sulfate dosage of 818 g / t, sodium sulfide dosage of 585 g / t, butyl xanthate dosage of 818 g / t, butylamine black medicine dosage of 65 g / t, and 2# oil 85 g / t, to obtain a mixed concentrate. The mixed concentrate is subjected to lead-zinc separation condition test. The test results of lead-zinc separation reagent YS-1 are shown in Table 15.

[0100] From the test results of zinc inhibitor YS-1 reagent, it can be seen that with the continuous increase of the dosage, the zinc grade and recovery rate in the lead rough concentrate do not decrease significantly, and the zinc inhibition effect is not good.

[0101] Table 15 Test results of YS-1 reagent dosage (%)

[0102]

[0103] The mixed concentrate lead-zinc separation inhibitor mainly conducts sodium humate, lime, zinc sulfate, sodium sulfite and JFR-1 reagent dosage test, and the test results are shown in Table 16.

[0104] From the test results of the inhibitor dosage, it can be seen that when the reagent dosage is sodium humate 348 g / t, lime 5360 g / t, zinc sulfate 456 g / t, sodium sulfite 456 g / t and JFR-1 dosage 362 g / t, the zinc grade and recovery rate in the lead rough concentrate decrease significantly, and the zinc inhibition effect is better.

[0105] Table 16 Test results of inhibitor dosage (%)

[0106]

[0107]

[0108] In view of the low lead recovery rate in the mixed concentrate lead-zinc separation, the mixed concentrate regrinding condition test is conducted under the same reagent conditions. The test results are shown in Table 17.

[0109] From the test results, it can be seen that compared with the mixed concentrate without grinding, the lead grade and recovery rate of the lead rough concentrate do not improve significantly.

[0110] Table 17 Test results of mixed concentrate regrinding conditions (%)

[0111]

[0112]

[0113] Based on the conditional tests, the closed-circuit test of bulk flotation of sulphide ore and separation of lead and zinc was carried out. The indexes of the closed-circuit test of bulk flotation process are shown in Table 18.

[0114] The yield of lead concentrate and zinc concentrate obtained by the closed-circuit test of bulk flotation is 0.37% and 0.93% respectively, the grade of lead and zinc is 10.89% and 42.13% respectively, and the recovery rate is 10.26% and 43.60% respectively.

[0115] Table 18 Total indexes of closed-circuit test of bulk flotation process (%)

[0116]

[0117] The conditional tests of rough scavenging with addition of activator sodium sulfide, zinc sulfide flotation scavenging with addition of activator sodium sulfide and zinc sulfide flotation with addition of copper sulfate were carried out in the equal floatation process, mainly to investigate the effect of activator sodium sulfide. The test results are shown in Table 19.

[0118] Table 19 Test results of butyl xanthate, sodium sulfide and copper sulfate dosage (%)

[0119]

[0120] The test results show that the addition of activator sodium sulfide in rough scavenging and zinc sulfide flotation scavenging in the equal floatation process has no obvious effect on the recovery of lead and zinc.

[0121] The test mainly investigates the effect of lead nitrate. The test results are shown in Table 20.

[0122] Table 20 Test results of equal floatation lead nitrate dosage (%)

[0123]

[0124]

[0125] The test results show that: (1) when the dosage of lead nitrate reaches 300 g / t, the lead recovery rate is the highest; (2) under the condition of the same total dosage of collector, the addition of a small amount of ammonium black drug in rough selection of equal floatation ensures the lead recovery rate, but part of the zinc with good floatability is also mixed in the lead concentrate with lead, which increases the zinc grade of lead concentrate, making it difficult to separate lead and zinc and affecting the zinc recovery rate.

[0126] The test mainly investigates the effect of zinc sulfide inhibitor combination in rough selection of lead operation. The test results are shown in Table 21.

[0127] Table 21 Test results of zinc sulfide inhibitor combination dosage in lead operation (%)

[0128]

[0129]

[0130]

[0131] The test results show that the lead recovery rate is the highest when the sodium humate dosage is 208 g / t, the zinc sulfate dosage is 332 g / t, the sodium sulfite dosage is 332 g / t, and the JFR-1 dosage is 312 g / t.

[0132] Based on the conditional test, the equal floatation closed-circuit test is carried out. The test indexes are shown in Table 22.

[0133] The equal floatation closed-circuit test obtains the lead concentrate yield of 0.34%, the lead grade of 18.53%, and the recovery rate of 15.68%; the zinc concentrate yield of 1.11%, the zinc grade of 41.31%, and the recovery rate of 51.10%.

[0134] Table 22 Equal floatation closed-circuit test indexes (%)

[0135]

[0136] The reagent dosage and cost of the mixed floatation process and the equal floatation process closed-circuit test are shown in Table 23.

[0137] Table 23 Reagent dosage and cost table

[0138]

[0139] Note: The reagent price is the reagent price of Chehe Concentrator.

[0140] The flotation is carried out in two test process flows of the mixed floatation process and the equal floatation-mixed floatation process. The flotation test index is shown in Table 24.

[0141] Table 24 Flotation process closed-circuit test indexes (%)

[0142]

[0143] The test results of the two process flows of the mixed floatation process and the equal floatation-mixed floatation process show that the equal floatation process is better than the mixed floatation process in terms of test indexes, and the reagent cost changes little.

[0144] According to the flotation process test results, the flotation tailings of the equal floatation process with the-0.2 mm beneficiation particle size are used as the ore sample of the gravity separation process. The ore sample contains 0.33% of tin, the original ore yield is 77.20%, and the tin recovery rate is 83.68%.

[0145] The flotation tailings ore sample of the equal floatation process is screened into three particle sizes of +0.074 mm, +0.038 mm, and-0.038 mm. The +0.074 mm and +0.038 mm ore samples are selected by the shaking table, and the-0.038 mm ore sample is respectively tested by the shaking table and the suspension disc concentrator.

[0146] The results of the micro-fine slime table test are shown in Table 25, the results of the suspension disc concentrator condition test are shown in Table 26, and the results of the best condition feeding of the suspension disc concentrator are shown in Table 27.

[0147] Table 25 - Results of table test of -0.038 mm ore sample (%)

[0148]

[0149] Table 26 - Results of suspension disc concentrator test of -0.038 mm ore sample (%)

[0150]

[0151]

[0152] Table 27 - Results of best condition feeding of suspension disc concentrator of -0.038 mm ore sample (%)

[0153]

[0154] The results of the test of the -0.038 mm ore sample using the table and the suspension disc concentrator show that the recovery rate of the suspension disc concentrator is higher than that of the table.

[0155] The total index of the gravity separation of the tailings of the equal floatation process is shown in Table 28, and the separation process is shown in Figure 2 .

[0156] Table 28 - Total index of gravity separation (%)

[0157]

[0158] The comprehensive product of the equal floatation process was subjected to chemical multi-element analysis, and the results of the analysis of the lead and zinc concentrates are shown in Tables 29 and 30.

[0159] Table 29 - Results of chemical multi-element analysis of lead concentrate (%)

[0160]

[0161]

[0162] Ag is in g / t.

[0163] Table 30 - Results of chemical multi-element analysis of zinc concentrate (%)

[0164] Element Sn Pb Zn Sb S Content 0.35 1.54 41.26 0.77 29.18 Element As Cd Fe Ag Content 0.20 0.31 8.35 217.25

[0165] Ag is in g / t.

[0166] The product granularity analysis was performed on the products of the equal floatation process, and the granularity analysis results of the -0.2 mm flotation tailings are shown in Table 31.

[0167] Table 31 Granularity analysis results of -0.2 mm flotation tailings (%)

[0168] Product name Yield Tin grade Tin distribution rate +0.074 12.12 0.43 15.97 Water separation + 0.074 10.15 0.87 27.05 0.037 15.33 0.41 19.25 0.019 16.84 0.29 14.96 0.010 39.24 0.17 20.44 -0.010 6.32 0.12 2.33 Total 100.00 0.33 100.00

[0169] 1) The sand tin ore in the Dachang mining area has complex mineral composition, low grade, high oxidation degree, and is difficult to separate. The main metal minerals in the ore are sphalerite, stibnite, galena, cassiterite, arsenopyrite, pyrite, and limonite, and the gangue minerals are mainly calcite and quartz. The elements of tin, lead, and zinc in the ore have industrial recovery value, and the elements of silver and sulfur have comprehensive recovery value. The comprehensive grade of the test sample is: tin 0.30%, lead 0.40%, zinc 0.90%, and silver 31.11 g / t.

[0170] 2) Problems and matters needing attention:

[0171] (1) The proportion of lead sulfide in the lead minerals of the sand tin ore is only 22.22%, and the proportions of lead oxide and lead-iron-vanadium are 77.78%; since the recovery of cassiterite is considered, the oxidized ore collector cannot be added in the test. Therefore, the recovery of lead minerals is very difficult.

[0172] (2) The content of -0.074 mm in the -2 mm raw ore is 51.24%, and the content of -0.019 mm is even 31.70%, which means the ore sample has high mud content, increasing the difficulty of recovery and separation of various useful metals.

[0173] 3) The flotation test indexes are as follows:

[0174] Equal floatation-mixed floatation process: lead concentrate yield 0.34%, lead grade 18.53%, lead recovery rate 15.68%, lead concentrate containing silver 1421 g / t; zinc concentrate yield 1.11%, zinc grade 41.31%, zinc recovery rate 51.10%, and zinc concentrate containing silver 217 g / t.

[0175] 4) The gravity separation indexes are as follows: tin concentrate yield 0.39%, tin content 40.57%, and tin recovery rate 51.66%; low-grade tin concentrate yield 0.51%, tin content 3.17%, and tin recovery rate 5.28%.

[0176] 5) It is suggested that the production process adopts the "equal floatation-mixed floatation separation-gravity separation" process.

[0177] The above merely describes the preferred embodiments of the present application, and is not intended to limit the present application in any form. Although the present application has been disclosed with the preferred embodiments as above, it is not intended to limit the present application, and any person skilled in the art can make some changes or modifications to the above-mentioned technical content with the above-mentioned prompt as equivalent embodiments with equivalent changes without departing from the technical solution of the present application. Any simple modification, equivalent change and modification made to the above embodiments according to the technical essence of the present application without departing from the technical solution of the present application still belongs to the scope of the present application.

Claims

1. A multi-metallic tin placer beneficiation process, comprising the following steps: Step 1: grinding classification: the raw ore of-2 mm is subjected to closed-circuit grinding classification to obtain ore pulp of-0.2 mm; Three ore samples of high grade containing tin of 0.35% or more, medium grade containing tin of 0.35% to 0.15%, and low grade containing tin of 0.15% or less are taken, and a total sample containing tin of about 0.3-0.35% is prepared as raw ore, the tin grade in the raw ore is 0.30%, the lead grade is 0.40%, the zinc grade is 0.90%, and the silver grade is 31.11 g / t; Among them, the metallic minerals in the raw ore are sphalerite, stibnite, galena, cassiterite, arsenopyrite, pyrite and limonite, and the gangue minerals are mainly calcite and quartz; the proportion of lead sulfide in the lead minerals is only 22.22%, and the proportion of lead oxide and lead-iron-vanadium is 77.78%; the proportion of zinc sulfide in the zinc minerals is 64.44%, and the proportion of zinc oxide is 30.00%; Step 2: Flotation of floatable: 2 # Lead, butyl xanthate, and sulfuric acid were used for flotation, and through one roughing, one cleaning, and two scavenging, lead-zinc-sulfur mixed concentrate and primary tailings were obtained. Step 3: lead mixed flotation: the lead-zinc-sulfur mixed concentrate is subjected to flotation separation by one roughing, two cleanings and two scavengings to obtain lead concentrate and zinc-sulfur tailings, the yield of lead concentrate is 0.34%, the lead grade is 18.53%, and the recovery rate is 15.68%, and the lead concentrate contains silver of 1421 g / t; Step 4: re-concentration of primary tailings: the primary tailings in step 2 are subjected to one roughing, two cleanings and two scavengings to obtain flotation tailings and secondary zinc-sulfur concentrate; Step 5: zinc-sulfur separation: the secondary zinc-sulfur concentrate and zinc-sulfur tailings are combined, and subjected to one roughing, two cleanings and two scavengings to obtain zinc concentrate and sulfur concentrate, the yield of zinc concentrate is 1.11%, the zinc grade is 41.31%, the zinc recovery rate is 51.10%, and the zinc concentrate contains silver of 217 g / t; Step 6: gravity separation: the flotation tailings are subjected to screening to obtain +0.074 mm material, -0.074 mm+0.038 mm material, and -0.038 mm material, the +0.074 mm material is subjected to two table gravity separations, the -0.074 mm+0.038 mm material is subjected to one table gravity separation, the middlings are re-ground and re-separated, and the -0.038 mm material is subjected to one suspension disc separator gravity separation, one table gravity separation, to obtain tin concentrate, secondary tin concentrate, and tailings; the yield of tin concentrate is 0.39%, the tin content is 40.57%, and the tin recovery rate is 51.66%; the yield of low-grade tin concentrate is 0.51%, the tin content is 3.17%, and the tin recovery rate is 5.28%.

2. A polymetallic placer tin ore beneficiation process according to claim 1, characterized in that, The closed-circuit grinding classification in step 1 adopts a combination of a ball mill and a hydrocyclone to form a closed-circuit process for grinding classification.

3. A polymetallic placer tin ore beneficiation process according to claim 1, characterized in that, In step 2, the dosage of reagents is: the dosage of sulfuric acid is 4550 g / t, the dosage of copper sulfate is 818 g / t, the dosage of sodium sulfide is 585 g / t, the dosage of butyl xanthate is 818 g / t, the dosage of ammonium black drug is 65 g / t, and the dosage of 2# oil is 85 g / t.

4. A polymetallic placer tin ore beneficiation process according to claim 1, characterized in that, In step 3, the dosage of reagents is: the dosage of sodium humate is 208 g / t, the dosage of calcium oxide is 890 g / t, the dosage of zinc sulfate is 332 g / t, the dosage of sodium sulfite is 332 g / t, and the dosage of JFR-1 is 312 g / t, the lead recovery rate is the highest.

5. A polymetallic placer tin ore beneficiation process according to claim 1, characterized in that, In step 4, the dosage of the reagents is as follows: the dosage of sulfuric acid is 5110 g / t, the dosage of copper sulfate is 348 g / t, the dosage of butyl xanthate is 348 g / t, and the dosage of 2# oil is 41 g / t.

6. A polymetallic placer tin ore beneficiation process according to claim 1, characterized in that, In step 5, the dosage of the reagents is as follows: the dosage of sodium humate is 120 g / t, the dosage of calcium oxide is 7890 g / t, the dosage of copper sulfate is 143 g / t, the dosage of butyl xanthate is 61 g / t, and the dosage of 2# oil is 27 g / t.

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