Copper collecting agent GTB1 and preparation method and application thereof

Abstract

The invention provides a copper collecting agent GTB1 as well as a preparation method and application of the copper collecting agent GTB1. The copper collecting agent GTB1 is prepared from isobutyl methyl thionocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol and zinc dialkyl dithiophosphate. The preparation method comprises the step of uniformly mixing the isobutyl methyl thionocarbamate, the O-isopropyl-N-ethyl thionocarbamate, the methyl isobutyl carbinol and the zinc dialkyl dithiophosphate according to a formula ratio of 6: 2: 1: 1.

Description

Technical Field

[0001] The present invention belongs to the field of metallurgical chemistry, and particularly relates to a copper collector GTB1, a preparation method thereof and an application thereof. Background Art

[0002] Copper, lead, zinc and iron minerals usually coexist in nature. Due to the similar physical and chemical properties of copper, lead and zinc minerals and their dense coexistence with gangue minerals such as quartz and feldspar, it is necessary to develop an efficient separation process during ore dressing to improve the quality and reduce the impurities of copper, lead and zinc concentrates. In recent years, with the development of the economy, the industrial demand for copper, lead and zinc has increased significantly, resulting in a gradual reduction of easily beneficiated ores in China and an increasing number of refractory polymetallic complex associated ores. Industrially, ores with a grade lower than 3% are usually regarded as low-grade ores. Such ores usually contain valuable elements such as sulfur and iron that can be comprehensively recovered. The content and distribution of these elements have a great impact on the subsequent metal smelting process. Therefore, different separation and recovery methods are needed for associated metals.

[0003] In order to comprehensively utilize copper, lead and zinc ores, processes such as preferential flotation process, bulk flotation process, iso-flotation process and asynchronous flotation process can be adopted. Which process to adopt specifically needs to comprehensively consider factors such as the floatability difference of minerals, mineral properties, process test results and economic benefits, and analyze the influence of different flotation sequences on ore dressing indexes through tests to determine the optimal flotation recovery process.

[0004] Since chalcopyrite in the ore is easily oxidized by weathering to precipitate copper ions, which activates pyrrhotite and pyrite, the use of xanthate collectors requires "high pressure and strong pulling", which greatly causes waste of reagents, directly resulting in high cost of ore dressing wastewater treatment and increased environmental protection pressure. For complex polymetallic ores, especially copper-iron polymetallic ores containing pyrrhotite, there is an urgent need to develop a copper collector that can solve the above problems. Summary of the Invention

[0005] The first object of the present invention is to provide a copper collector GTB1; the second object is to provide a preparation method of the copper collector GTB1; the third object is to provide an application of the copper collector GTB1.

[0006] The first object of the present invention is achieved as follows. The copper collector GTB1 is composed of isobutyl methyl thiocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol and zinc dialkyldithiophosphate.

[0007] The second object of the present invention is achieved as follows. It is obtained by mixing isobutyl methyl thiocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol and zinc dialkyldithiophosphate in the formulated ratio.

[0008] The third object of the present invention is achieved as follows: the application of the copper collector GTB1 in the comprehensive utilization and recovery process of polymetals.

[0009] The copper collector described in the present invention has the advantages of low dosage, good selective collection property for copper minerals, poor selectivity for lead, zinc and sulfur minerals without collection and adsorption, good degradation property of the reagent, low dosage of the supporting inhibitor, simple and convenient process, etc. Detailed implementation mode

[0010] The following further illustrates the present invention in conjunction with embodiments, but does not limit the present invention in any way. Any transformation or replacement based on the teachings of the present invention falls within the protection scope of the present invention.

[0011] The copper collector GTB1 described in the present invention is composed of isobutyl methyl thionocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol and zinc dialkyldithiophosphate.

[0012] The mass ratio of the isobutyl methyl thionocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol and zinc dialkyldithiophosphate is 6:2:1:1.

[0013] The preparation method of the copper collector GTB1 described in the present invention is to mix the isobutyl methyl thionocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol and zinc dialkyldithiophosphate in the formulated ratio.

[0014] The application of the copper collector GTB1 described in the present invention is the application of the copper collector GTB1 in the comprehensive utilization and recovery process of polymetals.

[0015] The following further illustrates the present invention with specific implementation cases: Example 1 Weigh isobutyl methyl thionocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol and zinc dialkyldithiophosphate = 6:2:1:1, and obtain the target copper collector GTB1 through stirring and mixing.

[0016] Example 2 Weigh isobutyl methyl thionocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol and zinc dialkyldithiophosphate = 5:3:1:1, and obtain the target copper collector GTB1 through stirring and mixing.

[0017] Example 3 Weigh isobutyl methyl thionocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol and zinc dialkyldithiophosphate = 6:1:2:1, and obtain the target copper collector GTB1 through stirring and mixing.

[0018] Example 4 The GTB1 copper collector prepared in Example 3 was used for the test, and the specific details are as follows: 1 Sample properties 1.1 Chemical components and mineral composition Observed with the naked eye, the ore is mostly dark brown and gray-green, and the mineral aggregates are distributed without orientation, forming a sparse disseminated structure of the ore. The results of the chemical multi-element analysis of the ore sample are shown in Table 1, and the results of the chemical phase analysis of lead, zinc, copper, and iron are shown in Tables 2, 3, 4, and 5.

[0019] Table 1 Chemical analysis results of main elements in the ore sample / %

[0020] Table 2 Lead phase analysis results of the raw ore / %

[0021] Table 3 Zinc phase analysis results of the raw ore / %

[0022] Table 4 Copper phase analysis results of the raw ore / %

[0023] Table 5 Iron phase analysis results of the raw ore / %

[0024] The ore sample is a lead-zinc polymetallic sulfide ore, and the main valuable elements are lead, zinc, and iron, with their contents being 2.64%, 1.45%, and 26.81 respectively. Other associated valuable elements are Au 0.10 g / t, Ag 10.8 g / t, Cu 0.045%, In 55.3 g / t, and sulfur 3.94%. Through observation under a polarized light microscope, X-ray diffraction analysis, and MLA analysis, it was found that there are 23 minerals in five categories, namely sulfides, oxides, silicates, sulfates, and phosphates, in the ore. The main minerals in the ore are galena, marmatite, magnetite, pyrite / marcasite, pyrrhotite, epidote,hedenbergite, talc, quartz, calcite, etc.

[0025] 1.2 Analysis of the dissemination state of main minerals Under the condition that the grinding fineness is 64.39% - 0.075mm, MLA is used to measure and count the dissemination sizes of the ore and the main target minerals in the ore, namely galena, marmatite, pyrrhotite, pyrite / marcasite, magnetite, hematite / limonite, epidote, johannsenite, andradite, epidote, and talc. The dissemination sizes of marmatite, epidote, and johannsenite are the coarsest, with the +75μm size fraction accounting for 33.29%, 38.54%, and 31.23% respectively, and their dissemination sizes are similar to the ore size; followed by galena, pyrrhotite, pyrite / marcasite, magnetite, andradite, and talc, with slightly finer dissemination sizes, and the +75μm size fraction accounting for 28.17%, 26.98%, 25.16%, 29.55%, 27.37%, and 26.39% respectively.

[0026] 1.3 Symbiotic relationships of main minerals Under the current grinding conditions, the free surface areas of marmatite, pyrite, and pyrrhotite are relatively good, all greater than 91%, while the free surface areas of galena, magnetite, and hematite / limonite are slightly lower than the former, between 83% and 86%.

[0027] Galena is mainly intergrown with marmatite, magnetite, epidote, and johannsenite, and the proportion of their common surface area is 3.60%, 1.18%, 2.10%, and 1.99 respectively. The free surface area proportion of galena is 86.50%.

[0028] Marmatite is mainly intergrown with epidote and johannsenite, and the proportion of their common surface area is 1.24% and 1.30% respectively. The free surface area proportion of marmatite is 93.89%.

[0029] Pyrite / marcasite is mainly intergrown with marmatite, pyrrhotite, and epidote, and the proportion of their common surface area is 1.24%, 1.64%, and 1.69% respectively. The free surface area proportion of pyrite / marcasite is 91.18%.

[0030] Pyrrhotite is mainly intergrown with pyrite / marcasite, magnetite, and epidote, and the proportion of their common surface area is 1.05%, 1.45%, and 1.24% respectively. The free surface area proportion of pyrrhotite is 92.54%.

[0031] Magnetite is mainly intergrown with epidote, johannsenite, calcite, and talc, and the proportion of their common surface area is 4.20%, 2.12%, 1.30%, and 1.96% respectively. The free surface area proportion of magnetite is 85.34%.

[0032] 1.4 Dissemination characteristics of main minerals The dissemination size of galena ranges from 0.001 to 1.37 mm, and the aggregates are granular and dense massive. It has a close symbiotic relationship with marmatite, magnetite, and pyrite, mostly occurring as intergrown or mutually enveloped. Among them, the symbiotic relationship between some galena and magnetite is particularly complex. This part of galena is in the form of fine grains enveloped in magnetite and is not easily monomer-dissociated.

[0033] The dissemination size of marmatite ranges from 0.005 to 4.75 mm, presenting as xenomorphic granular. It is mostly distributed among transparent minerals such as epidote and johannsenite, and some is enveloped in epidote and johannsenite, having partial intergrowth or envelopment with galena, magnetite, pyrite, and pyrrhotite. Among them, some galena, magnetite, and chalcopyrite are in the form of fine grains or opacified and enveloped in marmatite, and may partially enter the zinc concentrate, affecting the grade of the zinc concentrate.

[0034] The dissemination size of chalcopyrite ranges from 0.001 to 1.00 mm, presenting as xenomorphic granular. It is mostly intergrown with marmatite, galena, pyrite, and pyrrhotite, and some is in the form of fine grains or opacified and enveloped in marmatite; a small number is distributed in a star-like pattern among transparent minerals.

[0035] The dissemination size of pyrrhotite ranges from 0.005 to 0.55 mm, and pyrrhotite presents as xenomorphic granular. It is mostly intergrown mosaically with pyrite, marcasite, magnetite, marmatite, and galena; some is replaced by marcasite and magnetite, showing an interbedded or mixed distribution with them.

[0036] The dissemination size of pyrite / marcasite ranges from 0.005 to 6.00 mm, mostly presenting as euhedral - subhedral - xenomorphic granular. It is mostly scattered and disseminated among transparent minerals such as epidote and johannsenite; it is mostly intergrown with marmatite, pyrrhotite, galena, etc.; some pyrite is replaced by marcasite and magnetite, showing a mixed distribution with them.

[0037] The dissemination size of magnetite ranges from 0.005 to 1.25 mm, and magnetite presents as xenomorphic granular. It has a close symbiotic relationship with hematite, galena, marmatite, and pyrite. Some fills between acicular and tabular hematite grains, some is intergrown mosaically or mutually enveloped with galena, and some is simply intergrown with marmatite and pyrite; a small amount of magnetite is replaced by hematite, and a small amount of magnetite replaces pyrite.

[0038] 1.5 Analysis of Mineralogical Factors Affecting Mineral Processing Indexes The carrier minerals of lead in the ore are only galena, and the carrier minerals of zinc are only marmatite, both of which are independent sulfide minerals, facilitating independent recovery by flotation. The Fe content in marmatite is 12.69%, and this part of iron cannot be removed only by physical beneficiation methods. Therefore, the upper limit of the zinc concentrate grade belongs to Grade VI. Most of the marmatite in the ore encloses fine-grained and milky chalcopyrite, which is difficult to dissociate during grinding and may enter the zinc concentrate, affecting the zinc concentrate grade. Some galena has a close symbiotic relationship with marmatite, which may cause marmatite to enter the lead concentrate with galena, affecting the lead concentrate grade. There are many types of iron carrier minerals in the ore, and only magnetite is the useful iron mineral, with the distribution rate of iron in it being only 12.91%.

[0039] 2 Mineral processing test research From the results of the ore property research, it can be seen that the valuable minerals in the ore are mainly galena, marmatite, chalcopyrite, pyrrhotite, magnetite, and pyrite. Considering the surface characteristics and magnetism of the above minerals, the main technological processes for mineral recovery are the combined process of magnetic separation first and then flotation and the combined process of flotation first and then magnetic separation. Since the amount removed by magnetic separation is relatively small and magnetic separation is placed before lead flotation, thickening and dehydration are also required before flotation, increasing the complexity of the process. Therefore, the combined process of flotation first and then magnetic separation is adopted. The particles of galena with an embedded particle size <5μm account for 3.72%. Most of these fine-grained galena are enclosed in magnetite. Selective collectors with strong selectivity can be used to pre-concentrate them into the rough concentrate and then re-ground, which can not only prevent the high cost of fine grinding of the raw ore but also reduce the over-grinding degree of the target minerals. Since there are various recoverable minerals in the ore, the difficulty of the test lies in how to fully and efficiently recover various minerals and obtain high-quality concentrates. Since the copper content in the raw ore is extremely low, direct preferential flotation is not economically viable. Therefore, the principle technological process of copper-lead bulk flotation - zinc flotation - sulfur flotation - iron magnetic separation is preliminarily determined.

[0040] 2.1 Grinding fineness test The full dissociation of valuable minerals and gangue minerals is a prerequisite for efficient separation and recovery of minerals. Different grinding finenesses not only affect the dissociation of minerals but also ensure that the appropriate mineral particle size has a high fresh surface activity, strengthening the full interaction between the mineral surface and the collector, thereby enhancing the recovery and separation of minerals. To ensure high indexes in flotation, it is of great significance to determine the grinding fineness in combination with the ore properties. Since the ore contains easily slime-forming gangues such as calcite and iron talc, when selecting the appropriate grinding fineness, both the monomer dissociation degree of valuable minerals and the reduction of over-crushing of gangue to deteriorate the flotation pulp environment should be considered. Through tests, a grinding fineness of -74μm accounting for 60% is more appropriate. At this grinding fineness, when the flotation pulp concentration is 30% and the stirring speed of the flotation machine is 2000r / min, the air inflow of the flotation machine is 0.4m 3 / h. At natural pH values, using different collectors, a one-stage roughing and one-stage scavenging process flow was adopted to conduct tests on the types of copper-lead collectors. The test results are shown in Table 6. It can be seen from Table 6 that the five collectors are all copper-lead collectors with good selectivity screened in the early stage, and basically do not collect pyrite at natural pH values, effectively reducing the tests of pyrite inhibitors. When Pa is used alone as the copper-lead collector, the lead grade in the rough concentrate is the highest, and the recovery rates of lead and copper are the lowest; when using the combination of Pb-1 and F5B, the loss rate of lead in the tailings is the lowest, but the loss rate of copper is as high as 52.54%; when using the combination of No. 25 black medicine and GTB1, the loss rate of copper in the tailings is the lowest, and the recovery rate of lead in the rough concentrate obtained by roughing is the highest. In order to comprehensively recover copper and lead minerals, it is more appropriate to choose the combination of No. 25 black medicine and GTB1 for dosing.

[0041] Table 6 Test results of grinding fineness

[0042] 2.2 Test on the location of regrinding for copper-lead bulk concentrate The rough concentrate obtained from copper-lead roughing was subjected to three-stage cleaning to improve the copper-lead quality of the final concentrate. In order to investigate the influence of the regrinding location on copper-lead recovery, comparative tests were carried out on non-regrinding of the rough concentrate, regrinding of the rough concentrate to 90% passing -0.049mm, regrinding of the first-stage cleaned concentrate to 90% passing -0.049mm, and regrinding of the second-stage cleaned concentrate to 90% passing -0.049mm to investigate the feasibility of reducing the copper-lead grade in the middlings and improving the quality of the copper-lead bulk concentrate. Compared with non-grinding of the rough concentrate, the lead grade in the first-stage cleaned concentrate can be greatly improved after regrinding the rough concentrate. Generally speaking, the lead grade of the copper-lead bulk concentrate obtained by carrying out the third-stage cleaning operation after regrinding the second-stage cleaned concentrate is the highest. The lead grades of the copper-lead bulk concentrate are as follows: regrinding the second-stage cleaned concentrate > regrinding the first-stage cleaned concentrate > regrinding the rough concentrate > non-grinding. Since the recovery rate loss of lead in the copper-lead bulk concentrate is the largest after carrying out the third-stage cleaning operation after regrinding the second-stage cleaned concentrate, it is more appropriate to choose regrinding the first-stage cleaned concentrate after comprehensive consideration. The lead grade of the copper-lead bulk concentrate obtained under the conditions of regrinding the first-stage cleaned concentrate is 64.85%, and the copper grade is 0.91%. Since the copper grade is relatively low, subsequent copper-lead separation is not carried out.

[0043] 2.3 Test on the types of zinc-selective pyrite inhibitors By X-ray energy spectrum analysis, sphalerite contains 53.54% Zn and 12.69% Fe. As the iron content in sphalerite increases, its iron ions change its surface properties, affecting the adsorption of collectors on its surface and resulting in a gradual decline in floatability. The ore contains 1.44% marcasite / pyrite and 2.54% pyrrhotite. Combining X-ray energy spectrum analysis and MLA analysis, it can be known that the pyrrhotite in the ore is monoclinic, and the pyrrhotite contains 60.56% Fe. The floatability of the three minerals, marcasite / pyrite, pyrrhotite, and sphalerite, gradually decreases. In order to obtain high-quality zinc concentrate, it is necessary to use pyrite inhibitors to inhibit marcasite / pyrite and pyrrhotite. Lime (3000 g / t), inorganic inhibitor LY1 (2000 g / t), and organic inhibitor LY2 (600 g / t) were used respectively to carry out tests on different types of pyrite inhibitors, and to examine the effects of different inhibitors on zinc-sulfur separation. Using lime as the inhibitor for marcasite / pyrite and pyrrhotite, the effect is the best. Through one-stage rough zinc flotation and one-stage scavenging zinc flotation, the foam product (the combination of rough zinc concentrate and zinc middlings) has the highest zinc grade, reaching 29.64%, and the highest zinc recovery rate, which is 85.39%.

[0044] 2.4 Tests on Types of Zinc Collectors Valuable minerals such as pyrite and gangue minerals such as iron talc in the ore are prone to react with collectors, interfering with the flotation process. A suitable zinc collector can not only significantly improve the flotation efficiency of zinc minerals but also reduce reagent consumption. Therefore, a suitable zinc collector is an important prerequisite for realizing an efficient and low-consumption flotation process. Comparative tests were carried out using ethyl xanthate and ethyl thionocarbamate with wide adaptability and strong collecting ability, and collector A2 and GZ-1 with good selectivity to screen the most suitable collector for this type of high-iron sphalerite. Using GZ-1 and ethyl xanthate as collectors, the loss rate of zinc in the tailings is about 2%. However, the sulfur recovery rate of the tailings obtained by GZ-1 is higher than that of ethyl xanthate, indicating that ethyl xanthate has strong collecting ability and collects more pyrite into the rough zinc concentrate and zinc middlings; the collecting performances of A2 and ethyl thionocarbamate are basically similar, and the zinc loss rate and sulfur recovery rate in the tailings are relatively similar. In order to obtain the highest zinc recovery rate and minimize the entry of pyrite into the rough zinc concentrate and zinc middlings, it is more suitable to choose GZ-1.

[0045] 2.4 Tests on Improving Quality and Reducing Impurities of Zinc Second Concentrate It was found through the beneficiation test of zinc bulk concentrate that by adding lime three times respectively through three-stage beneficiation, the zinc grade of the concentrate 1 increased from 29.76% to 36.26%, an increase of 6.50 percentage points; through beneficiation 2, the zinc grade of the concentrate 2 increased from 36.26% to 39.12%, an increase of 2.86 percentage points, with a relatively small increase; through beneficiation 3, the zinc grade of the concentrate 3 (i.e., zinc secondary concentrate) increased from 39.12% to 41.36%, an increase of 2.24 percentage points, with a relatively small increase. When the fourth-stage beneficiation was continued, the increase in the zinc grade of zinc concentrate was only 1.5 percentage points, but the loss rate of zinc increased significantly, and it was difficult to increase the zinc grade of zinc concentrate to over 45%. Microscopic examination of the concentrate 3 found that marmatite was xenomorphic granular, mostly monomer-dissociated particles, some sphalerite contained fine-grained chalcopyrite, and some sphalerite was intergrown with transparent minerals, pyrrhotite, pyrite, etc.; pyrrhotite was xenomorphic granular, mostly monomer-dissociated particles, some of which were intergrown with sphalerite, etc.; pyrite was xenomorphic granular, mostly intergrown with sphalerite; chalcopyrite was xenomorphic granular, mostly encapsulated in sphalerite. The results showed that the main factors affecting the zinc grade of zinc concentrate were pyrrhotite and pyrite. The flotability of this part of pyrrhotite and pyrite was good, and it was difficult to inhibit by adding lime, and excessive addition of lime would cause obvious inhibition to marmatite. On the one hand, a flotation test of adding a dispersant to assist lime was carried out on zinc secondary concentrate to improve quality and reduce impurities, and on the other hand, a magnetic separation test for impurity removal was carried out on zinc secondary concentrate. Considering that pyrrhotite was prone to magnetic agglomeration and encapsulate marmatite, sodium hexametaphosphate of polyphosphate type and sodium silicate of widely used silicate type were used as dispersants. The results showed that when magnetic separation was used to improve the quality and reduce impurities of zinc concentrate, the operating recovery rate of zinc was relatively high, basically above 99%. Through the comparison of the results of four groups of tests, it was found that when magnetic separation of zinc secondary concentrate was carried out with a magnetic field of 0.25T, zinc concentrate with a zinc grade of 45.77% and a zinc operating recovery rate of 99.04% could be obtained. Adding a dispersant for flotation to increase the zinc grade of zinc concentrate resulted in a relatively large zinc loss rate. Considering comprehensively, it was more appropriate to carry out magnetic separation impurity removal on zinc secondary concentrate with 0.25T.

[0046] 2.5 Closed-circuit test of the whole process Under the optimal reagent regime, optimal regrinding fineness, and optimal magnetic separation field strength of the conditional test, a full-circuit test was carried out to examine the separation and recovery of each target mineral in copper-lead concentrate, zinc concentrate, sulfur concentrate, iron concentrate, and tailings, as well as the metal distribution of the middlings in each operation. In the copper-lead bulk flotation operation, copper-lead bulk concentrate was obtained through one roughing, two scavengings, and three cleanings; after two roughings, three cleanings, and one scavenging in the zinc flotation operation, zinc secondary concentrate was obtained, and zinc concentrate was obtained by subjecting the zinc secondary concentrate to magnetic separation with 0.25T; sulfur concentrate 1 was obtained through one roughing, one cleaning, and one scavenging of the tailings after zinc flotation; iron rough concentrate and tailings 1 were obtained by subjecting the tailings after sulfur flotation to magnetic separation. The iron rough concentrate was reground and then underwent one desulfurization roughing, one desulfurization cleaning, and one desulfurization scavenging to obtain sulfur concentrate 2 and desulfurized iron concentrate. The desulfurized iron concentrate was subjected to magnetic separation with a 0.15T magnetic field to obtain the final iron concentrate. The test results are shown in Table 7.

[0047] Table 7 Results of the full-circuit test

[0048] (1) The obtained copper-lead bulk concentrate has a lead grade of 60.12%, a copper grade of 0.73%, a copper recovery rate of 33.36%, and a lead recovery rate of 94.72%. The main factors affecting the quality of the copper-lead bulk concentrate are pyrite and marmatite. Since the associated copper grade is extremely low (copper content 0.045%), and chalcopyrite is mostly intergrown with marmatite, galena, pyrite, and pyrrhotite, the copper recovery rate in the copper-lead bulk concentrate is low, and the copper content in the copper-lead bulk concentrate is only 0.73%. The subsequent copper-lead separation is not economically viable, so copper-lead separation is not carried out; some galena is mainly closely intergrown with marmatite, resulting in a zinc grade as high as 6.89% in the copper-lead bulk concentrate, which can be further studied in depth to improve the separation efficiency of lead and zinc.

[0049] (2) The zinc grade of the zinc concentrate is 46.99%, and the zinc recovery rate is 88.26%. Since some chalcopyrite is in the form of fine grains and emulsion and is wrapped in marmatite, 30.34% of the copper metal is lost in the zinc concentrate.

[0050] (3) Sulfur concentrate 1 and sulfur concentrate 2 are combined into sulfur concentrate, with a sulfur grade of 37.41%, an iron grade of 47.69%, a sulfur recovery rate of 34.57%, and an iron recovery rate of 6.23%. In the ore, pyrite / marcasite:pyrrhotite = 1:1.76. The theoretical limit sulfur grade of the sulfur concentrate that can be obtained from this ore is 48.38%, and the iron grade is 51.62%. The results show that some pyrite / marcasite is lost in the copper-lead bulk concentrate and zinc concentrate, resulting in a lower sulfur grade in the sulfur concentrate.

[0051] (4) After magnetic separation of the tailings from sulfur scavenging 1, the obtained iron rough concentrate has an iron grade of 61.00%, a sulfur grade of 0.97%, and an iron recovery rate of 10.79%. In order to obtain high-quality iron concentrate with an iron grade greater than 65% and a sulfur grade less than 0.5%, the iron rough concentrate needs to be reground and then subjected to flotation desulfurization and magnetic separation for impurity removal.

[0052] (5) The iron recovery rate of the iron concentrate obtained by magnetic separation is 9.33%. Combining the phase analysis and MLA analysis, it can be known that the iron in magnetite accounts for 13.15% of the total iron, indicating that the phase recovery rate of magnetite in the iron concentrate is 70.95%. Analyzing the monomer dissociation degree of iron in the tailings, it can be seen that most of the magnetite lost in the tailings is undissociated, and this part of magnetite is difficult to recover economically.

[0053] 3 Conclusions (1) The main valuable elements in a polymetallic ore in Qinghai are lead, zinc, and iron, with their contents being 2.64%, 1.45%, and 26.81% respectively. Other associated valuable elements are Au 0.10 g / t, Ag 10.8 g / t, Cu 0.045%, In 55.3 g / t, and sulfur 3.94%.

[0054] (2) The ore is composed of five categories and 23 kinds of minerals in total. The independent minerals of copper, lead, and zinc are chalcopyrite, galena, and marmatite respectively. There are 12 independent minerals of iron in total. Iron mainly occurs in the form of independent minerals in epidote, johannsenite, and magnetite, and the distribution rates of iron in them are 46.83%, 16.71%, and 13.15% respectively. Only the iron in magnetite has the value of recovery.

[0055] (3) The ore is treated by the combined flotation and magnetic separation process of copper-lead bulk flotation - zinc flotation - sulfur flotation - iron magnetic separation - iron flotation, obtaining a copper-lead bulk concentrate with a lead grade of 60.12%, a copper grade of 0.73%, a copper recovery rate of 33.36%, a lead recovery rate of 94.72%, a zinc concentrate with a zinc grade of 46.99% and a zinc recovery rate of 88.26%, and an iron concentrate with an iron grade of 67.22% and an iron recovery rate of 9.33%.

[0056] Example 5 Experiments were carried out respectively with the copper collector GTB1 prepared in Example 1, Example 2, Example 3, and Example 4. The method was the same as that in Example 4. The results all showed that the copper collector GTB1 of the present invention was more suitable in the process of comprehensively recovering copper and lead minerals.

Claims

1. A copper collector GTB1, characterized in that, The copper collector GTB1 described above is composed of isobutyl methyl thionocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol, and zinc dialkyldithiophosphate.

2. The copper collector GTB1 according to claim 1, characterized in that, The mass ratio of isobutyl methyl thionocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol, and zinc dialkyldithiophosphate is 6:2:1:

1.

3. A method for preparing the copper collector GTB1 according to claim 1 or 2, characterized in that, It is obtained by mixing isobutyl methyl thionocarbamate, O-isopropyl-N-ethyl thionocarbamate, methyl isobutyl carbinol, and zinc dialkyldithiophosphate in the formula ratio.

4. Use of the copper collector GTB1 described in claim 1 or 2, characterized in that, The application of the copper collector GTB1 in the multi-metal comprehensive utilization and recovery process.

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