AU2017100906A4 - Beneficiation method for high-clay mixed lead-zinc oxide-sulfide ore - Google Patents
Beneficiation method for high-clay mixed lead-zinc oxide-sulfide ore Download PDFInfo
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Abstract
A beneficiation method for high-clay mixed lead-zinc oxide-sulfide ore including the steps of: a. performing ore crushing and/or ball milling of a raw mixed lead-zinc oxide 5 sulfide ore to provide a ground ore where 75% or more of the ground ore has a particle size of less than 0.074mm; b. passing the ground ore to an agitator barrel, and carrying out froth flotation to produce a lead concentrate and lead deficient stream; c. passing the lead deficient stream to a zinc sulphide flotation step to obtain 10 a zinc sulfide concentrate and a zinc sulphide deficient stream; d. dewatering and optionally desliming, the zinc sulphide deficient stream and then performing a slurry mixing in a thick pulp procedure by the addition of water and chemical additives to enhance sulfatization on the surface of the zinc oxide ore in the zinc sulphide deficient stream and 15 e. performing froth flotation on the dewatered zinc sulphide deficient stream to obtain a zinc oxide concentrate. The beneficiation method for high-clay mixed lead-zinc oxide-sulfide ore not only enhances the grades of lead and zinc concentrates, but also during beneficiation, reagents and tailings water can be recycled, thereby reducing environmental pollution. 7 1 Lg sI Figure 1
Description
Patents Act 1990 2017100906 03 Μ 2017 Ρ/00/011 Regulation 3.2 Australia
COMPLETE SPECIFICATION INNOVATION PATENT
Invention Title: Beneficiation method for high-clay mixed lead-zinc oxide-sulfide ore
The following statement is a full description of this invention, including the best method of performing it known to us: 1001864525 1001863956 2017100906 03 Μ 2017
BENEFICIATION METHOD FOR HIGH-CLAY MIXED LEAD-ZINC
OXIDE-SULFIDE ORE
Field of the invention
The present invention relates to beneficiation method for high-clay mixed lead-5 zinc oxide-sulfide ore.
Background of the invention
High-clay mixed lead-zinc oxide-sulfide ore is one of the world's top ten refractory ores and usually has a relatively complex composition. The main usable minerals that high-clay mixed lead-zinc oxide-sulfide ores contain are galena, cerussite, sphalerite, 10 smithsonite, hemimorphite, and pyrite. There are a variety of gangue minerals, and the minerals are mixed in a complex manner. It has been found through screening, analysis, and experiments on the raw ore, that the content of ore clay for the of -20 μΜ fraction is over 20%, the content of zinc oxide in this fraction is over 60%. Additionally, the crushed ore contains a large quantity of clay minerals, and high levels of silicon and, 15 calcium. Therefore the molecule accumulation of zinc oxide is greatly affected.
In existing beneficiation process procedures, gravity separation and flotation are usually combined. After ball milling, common flotation reagents and flotation methods are used.
However, conventional gravity separation and flotation methods have the 20 following defects: 1. The total recycling rate of metal is relatively low, grades of concentrates are below standard, and a great quantity of mineral resources cannot be effectively recycled and utilized. 2. During beneficiation, reagents and tailings water are not recycled. As a 25 result, the environmental discharge requirements are not satisfied, and the cost of waste water treatment is relatively high.
Reference to any prior art in the specification is not an acknowledgment or suggestion that this prior art forms part of the common general knowledge in any 2 1001863956 2017100906 03 Μ 2017 jurisdiction or that this prior art could reasonably be expected to be understood, regarded as relevant, and/or combined with other pieces of prior art by a skilled person in the art.
Summary of the invention 5 The present invention provides a beneficiation method for high-clay mixed lead- zinc oxide-sulfide ore, so that not only can grades of lead and zinc concentrates be enhanced, but also during beneficiation, reagents and tailings water can be recycled, thereby reducing environmental pollution.
Accordingly the invention provides a beneficiation method for high-clay mixed 10 lead-zinc oxide-sulfide ore including the steps of: a. performing ore crushing and/or ball milling of a raw mixed lead-zinc oxide-sulfide ore to provide a ground ore where 75% or more of the ground ore has a particle size of less than 0.074mm; b. passing the ground ore to an agitator barrel, and carrying out froth flotation 15 for a period of 25 - 45 minute to produce a lead concentrate and lead deficient stream; the froth flotation step comprising adding to the agitator barrel per tonne of ground ore; • 500 g to 1500 g of Na2C03 to adjust the pH value of the ground ore to 8.3 to 8.4, 20 • 300 g to 1000 g of Na2Si03 per tonne of ground ore to disperse the ore clay, and then • 2000 g of ZnS04, 50 g to 100 g of ethyl thiocarbamate, and 20 g to 50 g of ammonium dibutyl dithiophosphate, c. passing the lead deficient stream to a zinc sulphide flotation step, the zinc sulphide flotation step of the lead deficient stream comprising a single roughing stage, 25 3 cleaning stages and 4 scavenging stages, for a total period of 25 - 45 minutes to obtain a zinc sulfide concentrate and a zinc sulphide deficient stream; d. dewatering and optionally desliming, the zinc sulphide deficient stream and then performing a slurry mixing in a thick pulp procedure by the addition of water 3 1001863956 2017100906 03 Μ 2017 and chemical additives to enhance sulfatization on the surface of the zinc oxide ore in the zinc sulphide deficient stream, the slurry mixing method including the steps of: • adding water and 350 g to 500 g of Na2S to every ton of dewatered zinc sulphide slurry (dewatered zinc sulphide deficient stream) to enhance 5 sulfatization on the surface of zinc oxide ore, and further conditioning the zinc sulphide deficient stream to form a conditioned zinc oxide deficient stream by • adding 250 g to 500 g of CuS04 to activate the zinc oxide, and finally • adding 200 g to 350 g of Na2C03, to adjust the pH value of the dewatered 10 zinc sulphide deficient stream to 11.4 to 11.6; and e. performing froth flotation on the dewatered zinc sulphide deficient stream for a total flotation time of 15 minutes to 40 minutes to obtain a zinc oxide concentrate by the steps of • subjecting the dewatered zinc sulphide deficient stream to a single roughing 15 stage, 3 cleaning stages and 4 scavenging stages, by adding 100 g to 1000 g of a depressant to every ton of dewatered zinc sulphide deficient stream, and adding 800 g to 1500 g of a combined collector that is formed of octadecylamine, dodecylamine, and sodium dodecyl sulfonate in a ratio of 1:1:1. 20 In Step d, the decision to perform desliming or not depends on a combination of the raw ore clay content and the recycling rate of the beneficiation device. When the raw ore clay content is less than 5wt%, desliming preferably is not performed, and when the raw ore clay content is 5wt% or greater, desliming is performed, so that the input to the dewatering device can be reduced. 25 In Step d, the slurry mixing water obtained after the thick pulp procedure is removed is recyclable, so as to effectively utilize tailings water and beneficiation reagents. 4 1001863956 2017100906 03 Μ 2017
In Step d, the additive in the thick pulp procedure is a sodium sulfide solution at a concentration of 1wt% to 5wt%.
The depressant in Step e is Na2Si03 or Na2C03.
The device to perform the thick pulp procedure in Step d uses a low-speed 5 agitator barrel.
The dewatering device in Step d is a solid-liquid separation device such as a disc filter or a belt filter.
The beneficial effects of the present invention are: 1. By adopting the process according to the invention, the recycling rate and the concentrate grade are greatly I0 increased.
To form a solid sulfide film on the surface of zinc oxide with high contents of silicon, calcium, and soluble salts from the refractory ores, in the present invention, requires proportioning of the thick pulp and proper slurry mixing. The properly proportioned ore slurry is then automatically conveyed to a low-speed agitator barrel 15 using a machine to perform the stirring, until flotation separation and accumulation of zinc oxide can be performed. Meanwhile, the flotation reagents used in the present invention are highly synthesized and processed combined collectors, so that separation of zinc oxide from gangue such as silicon and calcium is carried out synchronously, and the recycling rate of zinc oxide can be increased. 20 In the process of the invention, reagents in the flotation of lead-zinc sulfide ore are first removed, and the tailings water obtained after sulfide flotation is then used to perform sedimentation desliming and repeated beneficiation of the lead-zinc sulfide ore. After sedimentation and desliming are performed on the tailings water obtained after flotation of zinc oxide, repeated beneficiation of the zinc oxide ore may be performed, so 25 as to achieve recycling of tailings water, reduce waste water discharge, reduce environmental pollution, and at the same time further reduce the beneficiation cost.
As used herein, except where the context requires otherwise, the term "comprise" and variations of the term, such as "comprising", "comprises" and 5 1001863956 2017100906 03 Μ 2017 "comprised", are not intended to exclude further additives, components, integers or steps.
Further aspects of the present invention and further embodiments of the aspects described in the preceding paragraphs will become apparent from the following 5 description, given by way of example and with reference to the accompanying drawings.
Brief description of the drawings FIG. 1 is a schematic flowchart of a beneficiation method for high-clay mixed lead-zinc oxide-sulfide ore according to the present invention; and FIG. 2 is a flowchart of a process of treating ore by using a conventional method. 10 Detailed description of the embodiments
The present invention is further described below in detail with reference to specific implementation examples and the accompanying drawings.
Embodiments of the present invention describe in detail a beneficiation method for high-clay mixed lead-zinc oxide-sulfide ore (1). As shown in FIG. 1, the beneficiation 15 method includes the following steps: a. Perform ball milling (2) and ore crushing on raw ore (1) of mixed lead-zinc oxide-sulfide ore using a ball mill, where the particle size of more than 75% of the ground ore is less than 0.074 mm. b. Separate the ground ore whose particle size is less than 0.074 mm using a 20 separator where the ore preferably is graded (3) in a single grading pass, and then conveyed to an agitator barrel. The ore having a particle size greater than 0.074mm is then subjected to ball milling and regraded.
The graded ore then undergoes a lead sulphide flotation step (4). To the ore whose particle size is less than 0.074mm, 500 g to 1500 g of Na2C03 is added for every 25 ton of the ground ore in the agitator barrel, the pH value of the ground ore is adjusted to 8.3 to 8.4. 300 g to 1000 g of Na2Si03 per ton of ground ore is added to disperse the ore clay, and then per tonne of ground ore 2000 g of ZnS04, 50 g to 100 g of ethyl 6 1001863956 2017100906 03 Μ 2017 thiocarbamate, and 20 g to 50 g of ammonium dibutyl dithiophosphate, is added and froth flotation performed for 25 minutes to 45 minutes to obtain a lead concentrate (5). c. The tailings obtained after flotation of lead proceed to a zinc sulfide flotation step (6) comprising a single roughing stage, three cleaning stages, and four scavenging 5 stages, during the flotation of zinc sulfide. The flotation step includes adding 300 g to 600 g of CUSO4 per tonne of this tailings stream to activate the zinc sulfate, after which 100 g to 500 g of Na2C03 per tonne of this tailings stream is added. The pH of this tailings stream is adjusted to 9.4 to 9.6, and 50 g to 200 g of butyl xanthate and 5 g to 30 g of terpene oil per tonne of tailings is added to obtain a zinc sulfide concentrate. 10 The flotation time is 25 minutes to 45 minutes. d. Desliming and dewatering (8) is performed on the tailings obtained after flotation of zinc sulfide is completed, and then slurry mixing (9) is carried out using a thick pulp procedure (10).
The slurry mixing and thick pulp steps include adjusting the concentration of the 15 chemical agents. This is done by mixing water to the tailings and chemical agent to form the thick pulp. In this case, it is in the range of 170 to 250kg (preferably 180kg) Na2S per tonne of water is added. This adjusts the density of the pulp to enhance sulfatization on the surface of zinc oxide ore, enabling the slurry to reach the requirements for flotation after which thick pulp is removed ie the chemical reagents are removed from >0 the mineral residue. Water and flotation chemical agents are then added to perform slurry mixing. The slurry mixing method includes adding water, 350 g to 500 g of Na2S to every ton of ore slurry to perform sulfatization, then adding 250 g to 500 g of CUSO4 to perform activation, and finally adding 200 g to 350 g of Na2C03, to adjust the pH value of the ore slurry to 11.4 to 11.6. this results in a reduction in waste water 25 discharge to meet an environmental protection requirements, and at the same time to further reduce the beneficiation cost.
In Step c, Step d, and Step e, an excessively low pH value ie pH values less than the lower limit affects the concentrate grade, and an excessively high pH value ie above the specified pH range affects the recycling rate. 7 1001863956 2017100906 03 Μ 2017 e. Flotation process (11) for a flotation time of 25 to 40 minutes through a single roughing step, three cleaning steps, and four scavenging steps is performed for the ore slurry obtained after slurry mixing in Step d. 100 g to 1000 g of a depressant to every ton of ore slurry is added, and 800 g to 1500 g of a combined collector per tonne of ore 5 slurry is added. The combined collector is formed of octadecylamine, dodecylamine, and sodium dodecyl sulfonate in a ratio of 1:1:1, to obtain a zinc oxide concentrate (12) and tailing water stream (13).
In this embodiment, the flotation reagents used are highly synthesized and processed combined collectors, separation of zinc oxide from gangue such as silicon I0 and calcium is implemented synchronously, and the recycling rate of zinc oxide is increased.
In Step d and Step e, to form a solid sulfide film on the surface of zinc oxide with a high content of silicon, calcium, and soluble salts, as in the refractory ores used in the invention, proportioning of a thick pulp and proper slurry mixing are performed. The 15 properly proportioned ore slurry is then automatically conveyed to the agitator barrel provided with a stirrer to perform stirring, until effective flotation separation and accumulation of zinc oxide can occur, thereby increasing the grade of the zinc oxide concentrate.
In Step d, desliming is optional and the decision to perform desliming or not >0 depends on a combination of the raw ore clay content and the recycling rate of the beneficiation device. When the raw ore clay content is less than 5wt%, desliming preferably is not performed, and when the raw ore clay content is greater than 5wt%, desliming is performed, so that the input to the dewatering device can be reduced.
In Step d, the slurry mixing water obtained after the thick pulp is removed is 25 recyclable, and the water used for slurry mixing is partially tailings water, so as to effectively and properly utilize tailings water.
In Step d, the thick pulp is formed by the addition of an aqueous sodium sulfide solution containing of 1wt% to 5wt% sodium sulfide.
Preferably the depressant in Step d is Na2Si03 or Na2C03. 8 1001863956 2017100906 03 Μ 2017
The dewatering device in Step d uses a solid-liquid separation device such as a low-speed agitator barrel, a disc filter or a belt filter. In this embodiment, a low-speed agitator barrel is used.
This process is in contrast to the prior art process which do not include a 5 dewatering, slurry mixing and pulp thickening processing steps according to the present invention.
Embodiment 1:
The main usable minerals in the mixed oxide-sulfide ore in Fung Wong Shan, Lanping County, Yunnan are galena, cerussite, sphalerite, smithsonite, hemimorphite, I0 and pyrite. The recyclable valuable ore is lead sulfide, lead oxide, zinc sulfide, and zinc oxide. The content of lead is about 1wt% to wt3%, the content of zinc is 6wt% to 18wt%, the oxidization rate is up to 55 to 80% (signifying that 55%-80% of zinc in the raw ore is zinc oxide, and only 45%-20% is zinc sulphide), the grade of lead sulfide is above 0.6wt%, the grade of lead oxide is above 1wt%, the grade of zinc sulfide is above 15 4.5wt%, and the grade of zinc oxide is above 9.8wt%. The ore is brittle, the clay content is high, gangue minerals are complex, and the content of soluble salts is high and as a result it is difficult to perform flotation separation and accumulation.
Step 1. Utilize the method of the present invention, the raw ore is crushed and ball milled, to a fineness where 75% of the ore has a the particle size less than 0.074 >0 mm. After grading the finely ground or is conveyed to an agitator barrel. The ore which has a particle size greater than 0.074mm is ball milled again. To the agitator barrel 1000 g of Na2C03 to every ton of the ground ore in the agitator barrel is added, the pH value of the ground ore in the ground ore is adjusted to 8.3, and 800g of Na2Si03 per tonne of ore is added to disperse the ore clay, followed by 1500 g of ZnS04per tonne of ore, 60 g 25 of ethyl thiocarbamate, and 30 g of ammonium dibutyl dithiophosphate, and air is introduced and flotation performed for a period of 35 minutes to obtain a lead concentrate.
The tailings are then taken for flotation of lead sulfide. This is done by adding 600g of Na2C03 per tonne of tailings, adjusting the pH value of the tailings to 8.4, during 30 flotation of lead sulfide, adding to every ton of tailings: 1000 g of Na2Si03,100 g of ethyl 9 1001863956 2017100906 03 Μ 2017 thiocarbamate, 30 g of ammonium dibutyl dithiophosphate, and 600 g of ZnS04-7H20, and then performing a flotation process preferably comprising consisting of a single roughing stage, three cleaning stages, and three scavenging stages, to obtain a lead sulfide concentrate, where the flotation time was 35 minutes. 5 Step 2. After flotation of the lead sulphide, the tailings are passed to a further flotation process for flotation of zinc sulfide. The flotation process consists of a single roughing step, three cleaning steps, and four scavenging steps. During flotation of zinc sulfide, to every ton of tailings, 400 g of CuS04 is added to perform activation of zinc sulfate, then add 300 g of Na2C03, added to adjust the pH value of the tailings to 9.4, 10 and finally 100 g of butyl xanthate and 20 g of terpene oil is added, to obtain a zinc sulfide concentrate, where the flotation time is 38 minutes.
Step 3. After flotation of zinc sulfide is completed, the reagents for the flotation of lead-zinc sulfide ore is removed, then slurry mixing is performed using a thick pulp. The density of the pulp is adjusted to enhance sulfatization on the surface of zinc oxide and 15 makes the slurry reach a conditions required for flotation. The slurry mixing method includes the steps of adding for every ton of ore slurry, 350 g of Na2S to perform sulfatization, then flotation involves adding 300 g of CuS04 to perform activation, and finally adding 300 g of Na2C03, to adjust the pH value of the ore slurry to 11.4.
After flotation, the slurry mixing water obtained after the thick pulp is removed is >0 recycled, and the slurry mixing method is: adding 350 g of Na2S to every ton of ore slurry to perform sulfatization, then adding 300 g of CuS04 to perform activation, and finally adding 300 g of Na2C03, to adjust the pH value of the ore slurry to 11.4.
For the ore slurry obtained after slurry mixing in Step 3, a flotation step comprising a single roughing stage, three cleaning steps, and four scavenging steps 25 was performed. Then for every tonne of ore slurry, 600 g of a depressant Na2Si03 and finally 1000 g of a combined collector that is formed of octadecylamine, dodecylamine, and sodium dodecyl sulfonate in a ratio of 1:1:1 is added, to obtain a zinc oxide concentrate, where the flotation time is 35 minutes.
For the entire beneficiation process, refer to FIG. 1. For experimental results, 30 refer to Table 1. For comparison, the conventional gravity separation and full-flotation 10 1001863956 2017100906 03 Μ 2017 process in FIG. 2 was used to perform processing. For experimental results, also refer to Table 1.
Table 1
Process Product name Grade (%) Recyclinc rate (%) Pb Zn Pb Zn Process of the present invention Lead sulfide concentrate 57.71 1.56 60.48 1.55 Zinc sulfide concentrate 1.12 51.39 1.98 93.42 Zinc oxide concentrate 1.71 39.30 1.27 81.72 Conventional process Lead concentrate Lead sulfide concentrate 55.24 3.26 46.25 3.31 Zinc sulfide concentrate 1.12 51.39 1.98 93.42 Zinc oxide concentrate 0.35 20.18 0.65 32.13 5 As can be seen from Table 1, after carrying out the process according to the present invention, a comparison of the various data, it can be seen that the grades and recycling rates of the lead sulfide concentrate, the zinc sulfide concentrate, and the zinc oxide concentrate are all greatly increased. This is especially the case with the recycling rates of zinc sulfide and zinc oxide increased by 40% to 50% as compared with those in 10 a conventional process.
Embodiment 2: This embodiment changes the flotation times of various concentrates and amounts of various reagents and collectors used during flotation, so as to adjust pH values during flotation.
The mixed oxide-sulfide ore in Fung Wong Shan, Lanping County, Yunnan is still 15 used as raw ore. The ore milling method and requirements are the same as those in Embodiment 1. During flotation of various concentrates, the methods and steps are the same as those in Embodiment 1, and only the flotation times of various concentrates, use amounts of reagents, and pH values during flotation are changed. 11 1001863956 2017100906 03 Μ 2017
During flotation of the lead concentrate, per tonne of ore, 1200 g of Na2C03 is added to adjust the pH value of the ground ore to 8.4, then 1000 g of Na2Si03, is added and then1700 g of ZnS04, 80 g of ethyl thiocarbamate, and 50 g of ammonium dibutyl dithiophosphate is added, with a total flotation time of 35 minutes. 5 During flotation of the zinc sulfide concentrate, per tonne of ore 400 g of CUSO4 is added to perform activation of zinc sulfate, then 500 g of Na2C03 to adjust the pH value of the tailings to 9.5, and finally 120 g of butyl xanthate and 30 g of terpene oil is added, with a total flotation time of 36 minutes.
For the tailings obtained after flotation of zinc sulfide is completed, first the I0 reagents used in flotation of lead-zinc sulfide ore are removed, then slurry mixing by using a thick pulp is used to enhance sulfatization on the surface of zinc oxide. The thick pulp is then removed and the combined removed thick pulp and slurry mixing water is recyclable. The slurry mixing method includes the steps of is: adding 350g -500g of Na2S to every ton of ore slurry to perform sulfatization, then adding 500 g 15 CuS04 per tonne to perform activation, and finally adding 500 g of Na2C03, per tonne to adjust the pH value of the ore slurry to 11.5.
During flotation of the zinc oxide concentrate, a depressant ,800 g of Na2Si03 per tonne of ore slurry is added, and finally 1200 g per tonne of a combined collector that is formed of octadecylamine, dodecylamine, and sodium dodecyl sulfonate in a ratio of >0 1:1:1 was added with a total flotation time of 36 minutes.
The experimental results are shown in Table 2. For comparison, the conventional gravity separation and full-flotation process in FIG. 2 are used to perform processing. For experimental results, also refer to Table 2. 12 2017100906 03 Μ 2017 1001863956
Table 2 Process Product name Grac e (%) Recycling rate (%) Pb Zn Pb Zn Process of the present invention Lead sulfide concentrate 58.45 1.76 61.48 1.57 Zinc sulfide concentrate 1.22 53.39 2.18 93.63 Zinc oxide concentrate 1.82 41.16 1.32 82.66 Conventional process Lead sulfide concentrate 55.93 3.28 46.47 2.90 Zinc sulfide concentrate 1.12 51.39 1.98 93.42 Zinc oxide concentrate 0.37 20.37 0.67 32.43
As can be seen from Table 2, after performing the process according to the invention, a comparison of the tabulated data shows that both recycling rates and 5 concentrate grades are greatly increased as compared with those in a conventional process. Particularly, tailings water can be recycled, a use rate of resources is increased, at the same time discharge of toxic waste water is reduced, heavy metal pollution is also greatly reduced, and environmental protection is facilitated.
The foregoing provides only preferred embodiments of the present invention; 10 however, the present invention is not limited thereto. For a person skilled in the art, various changes and variations may be made to the present invention. Any changes, equivalent replacements, and improvements made within the spirit and principle of the present invention should fall within the protection scope of the present invention.
It will be understood that the invention disclosed and defined in this specification 15 extends to all alternative combinations of two or more of the individual features mentioned or evident from the text or drawings. All of these different combinations constitute various alternative aspects of the invention. 13
Claims (5)
1. A beneficiation method for high-clay mixed lead-zinc oxide-sulfide ore including the steps of: a. performing ore crushing and/or ball milling of a raw mixed lead-zinc oxide-sulfide ore to provide a ground ore where 75% or more of the ground ore has a particle size of less than 0.074mm; b. passing the ground ore to an agitator barrel, and carrying out froth flotation for a period of 25 - 45 minute to produce a lead concentrate and lead deficient stream; the froth flotation step comprising adding to the agitator barrel per tonne of ground ore; • 500 g to 1500 g of Na2C03 to adjust the pH value of the ground ore to 8.3 to 8.4, • 300 g to 1000 g of Na2Si03 per tonne of ground ore to disperse the ore clay, and then • 2000 g of ZnS04, 50 g to 100 g of ethyl thiocarbamate, and 20 g to 50 g of ammonium dibutyl dithiophosphate, c. passing the lead deficient stream to a zinc sulphide flotation step, the zinc sulphide flotation step of the lead deficient stream comprising a single roughing stage, 3 cleaning stages and 4 scavenging stages, for a totalperiod of 25 - 45 minutes to obtain a zinc sulfide concentrate and a zinc sulphide deficient stream; the d. dewatering and optionally desliming, the zinc sulphide deficient stream and then performing a slurry mixing in a thick pulp procedure by the addition of water and chemical additives to enhance sulfatization on the surface of the zinc oxide ore in the zinc sulphide deficient stream , the slurry mixing method including the steps of: • adding water and 350 g to 500 g of Na2S to every ton of dewatered zinc sulphide slurry to enhance sulfatization on the surface of zinc oxide, and further conditioning the zinc sulphide deficient stream to form a conditioned zinc sulphide deficient stream by • adding 250 g to 500 g of CUSO4 to activate the zinc oxide, and finally • adding 200 g to 350 g of Na2C03, to adjust the pH value of the zinc sulphide deficient stream to 11.4 to 11.6; and e. performing froth flotation on the conditioned zinc sulphide deficient stream for a total flotation time of 25 minutes to 40 minutes to obtain a zinc oxide concentrate by the steps of • subjecting the conditioned zinc sulphide deficient stream to a single roughing stage, 3 cleaning stages and 4 scavenging stages, by adding 100 g to 1000 g of a depressant to every ton of dewatered zinc sulphide deficient stream, and adding 800 g to 1500 g of a combined collector that is formed of octadecylamine, dodecylamine, and sodium dodecyl sulfonate in a ratio of 1:1:1. to produce a zinc oxide concentrate stream and tailings water stream.
2. The method of claim 1 wherein the optional desliming step in step d of claim 1 is not performed when the raw ore clay content is less than 5%wt.
3. The method of claim 1 wherein the optional desliming step in step d of claim 1 is performed when the raw ore clay content is 5wt% or greater.
4. The method of one of claims 1 -3 wherein the slurry mixing water obtained after the thick pulp procedure is removed in step d of claim 1 is is recycled so as to effectively reuse tailings water and beneficiation reagents.
5. The method of claim 1 -4 wherein the additive in the thick pulp procedure in step d of claim 1 is a sodium sulfide solution at a concentration of 1wt% to 5wt% and the depressant in step e is Na2SiC>3 or Na2CC>3.
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| CN110157924A (en) * | 2019-06-11 | 2019-08-23 | 长沙有色冶金设计研究院有限公司 | A kind of method of the low-grade secondary Zn accumulation of high oxidation rate oxysulphied zinc ore |
| CN111330741A (en) * | 2020-03-09 | 2020-06-26 | 中南大学 | Synchronous flotation method for oxidized-vulcanized mixed zinc ore |
| CN113042216A (en) * | 2021-03-09 | 2021-06-29 | 中南大学 | Flotation separation method for carbonaceous lead sulfide zinc minerals |
| CN113528855A (en) * | 2021-07-23 | 2021-10-22 | 昆明冶金研究院有限公司 | Method for extracting zinc from floating zinc oxide concentrate |
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2017
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| CN110157924B (en) * | 2019-06-11 | 2021-09-21 | 长沙有色冶金设计研究院有限公司 | Method for secondary zinc enrichment of low-grade high-oxidation-rate oxygen-sulfur mixed zinc ore |
| CN111330741A (en) * | 2020-03-09 | 2020-06-26 | 中南大学 | Synchronous flotation method for oxidized-vulcanized mixed zinc ore |
| CN113042216A (en) * | 2021-03-09 | 2021-06-29 | 中南大学 | Flotation separation method for carbonaceous lead sulfide zinc minerals |
| CN113042216B (en) * | 2021-03-09 | 2022-06-03 | 中南大学 | Flotation separation method for carbonaceous lead sulfide zinc minerals |
| CN113528855A (en) * | 2021-07-23 | 2021-10-22 | 昆明冶金研究院有限公司 | Method for extracting zinc from floating zinc oxide concentrate |
| CN113528855B (en) * | 2021-07-23 | 2023-02-14 | 昆明理工大学 | Method for extracting zinc from flotation zinc oxide concentrate |
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