阅读说明：本技术 一种金矿捕收剂及其应用 (Gold ore collecting agent and application thereof ) 是由 马立成 叶树峰 刘翔 李正辰 闫敬民 于 2021-07-08 设计创作，主要内容包括：本发明涉及一种金矿捕收剂及其应用,所述捕收剂包括：异戊基黄药和N-烷基二硫代氨基甲酸次甲基膦酸二烷基脂的混合物。所述应用包括如下步骤：(1)将金原矿进行磨矿分级后,加入包含所述捕收剂的浮选药剂依次进行粗选、精选和扫选,得到金精矿；(2)将步骤(1)所述金精矿依次进行第一焙烧、第二焙烧、浸出和固液分离,得到含金液相和浸出渣,所述浸出渣经磁化焙烧后进行磁选,得到铁精矿。通过对捕收剂配方的调整,实现了对高硫卡林型型金矿中金的高效富集,同时特定的焙烧工艺,实现了该矿石中的硫和铁的综合利用。获得的高硫金精矿金品位＞8g/t,硫品位>30％,金回收率可达到90％以上。(The invention relates to a gold ore collecting agent and application thereof, wherein the collecting agent comprises: a mixture of isoamyl xanthate and N-alkyl dithiocarbamate dialkylester of phosphinic acid. The application comprises the following steps: (1) grinding and grading the gold raw ore, and adding a flotation agent containing the collecting agent to perform rough concentration, fine concentration and scavenging in sequence to obtain gold concentrate; (2) and (2) sequentially carrying out first roasting, second roasting, leaching and solid-liquid separation on the gold concentrate obtained in the step (1) to obtain a gold-containing liquid phase and leaching residues, and carrying out magnetic separation on the leaching residues after magnetization roasting to obtain iron concentrate. By adjusting the formula of the collecting agent, the high-efficiency enrichment of gold in the high-sulfur Carlin type gold ore is realized, and the comprehensive utilization of sulfur and iron in the ore is realized by the specific roasting process. The gold grade of the obtained high-sulfur gold concentrate is more than 8g/t, the sulfur grade is more than 30 percent, and the gold recovery rate can reach more than 90 percent.)

1. A gold ore collector, comprising: a mixture of isoamyl xanthate and N-alkyl dithiocarbamate dialkylester of phosphinic acid.

2. The gold ore collector of claim 1 wherein the mass ratio of isoamyl xanthate to dialkyl N-alkyl dithiocarbamate methylphosphonate is (1-4): 1.

3. Use of a gold ore collector according to claim 1 or claim 2, wherein the use comprises the steps of:

(1) grinding and grading gold raw ore, and adding a flotation reagent containing the collecting agent according to claim 1 or 2 to perform rough concentration, fine concentration and scavenging in sequence to obtain gold concentrate;

(2) and (2) sequentially carrying out first roasting, second roasting, leaching and solid-liquid separation on the gold concentrate obtained in the step (1) to obtain a gold-containing liquid phase and leaching residues, and carrying out magnetic separation on the leaching residues after magnetization roasting to obtain iron concentrate.

4. The use according to claim 3, wherein the gold grade in the gold raw ore of step (1) is 1-3g/t, the sulphur content > 5%;

preferably, the ore grinding classification of step (1) comprises: the gold raw ore is subjected to primary ball milling and then primary screening to obtain a primary undersize product, and a secondary ball milling product obtained by secondary ball milling of the primary oversize product and the primary undersize product are mixed to obtain a classification product and slurry preparation is carried out;

preferably, the-200-mesh particles in the primary undersize product account for 50-70% of the total mass of the primary undersize product in mass percentage;

preferably, the-200-mesh particles in the classified product account for 80-95% of the total mass of the classified product in terms of mass percentage content.

5. The use as claimed in claim 3 or 4 wherein the collector of step (1) is added in an amount of 100-500 g/t;

preferably, the flotation reagent in the step (1) further comprises a regulator, a depressant, an activator and a foaming agent;

preferably, the conditioning agent comprises 1 or a combination of at least 2 of sodium carbonate, oxalic acid, sulfuric acid, or ammonium sulfate;

preferably, the addition amount of the regulator is 500-2000 g/t;

preferably, the inhibitor comprises water glass and/or sodium hexametaphosphate;

preferably, the addition amount of the inhibitor is 800-1500 g/t;

preferably, the activator comprises copper sulfate and/or lead sulfate;

preferably, the addition amount of the activating agent is 200-600 g/t;

preferably, the foaming agent comprises pinitol oil and/or diethyl phthalate;

preferably, the blowing agent is added in an amount of 10 to 100 g/t.

6. Use according to any one of claims 3 to 5, wherein said concentration of step (1) is carried out at least 2 times;

preferably, the scavenging of step (1) is performed at least 3 times.

7. The use as claimed in any one of claims 3 to 6, wherein the temperature of the first calcination in step (2) is 350-550 ℃;

preferably, the time for the first roasting in the step (2) is 0.5-1.5 h;

preferably, the temperature of the second roasting in the step (2) is 550-800 ℃;

preferably, the time of the second roasting in the step (2) is 0.5-2 h.

8. Use according to any one of claims 3 to 7, wherein the leached solid phase of step (2) contains particles of-400 mesh in a mass percentage content of 75 to 95% of the total mass of the leached solid phase;

preferably, the solid-to-liquid ratio g/mL in the leaching in the step (2) is 1 (1-5);

preferably, the leaching of step (2) comprises a pretreatment with an acid followed by leaching with a cyanate;

preferably, the acid comprises sulfuric acid and/or hydrochloric acid;

preferably, the time of the pretreatment is 2-6 h;

preferably, the cyanate is added in an amount of 0.5-4000 g/t;

preferably, the density of the bottom carbon in the leaching is 8-16 g/L;

preferably, the leaching time is 24-60 h.

9. The use of any one of claims 3 to 8, wherein in step (2) the magnetizing roasting is carried out by adding a reducing agent;

preferably, the reductant comprises pulverized coal;

preferably, the addition amount of the reducing agent is 2-12% of the mass of the leaching residue;

preferably, the temperature of the magnetizing roasting in the step (2) is 600-850 ℃;

preferably, the magnetizing roasting time in the step (2) is 0.5-1.5 h.

10. Use according to any of claims 3-9, characterized in that the use comprises the steps of:

(1) grinding and grading the gold raw ore, and adding a flotation agent containing the collecting agent to perform rough concentration, fine concentration and scavenging in sequence to obtain gold concentrate;

(2) sequentially carrying out first roasting, second roasting, leaching and solid-liquid separation on the gold concentrate obtained in the step (1) to obtain a gold-containing liquid phase and leaching residues, and carrying out magnetic separation on the leaching residues after magnetization roasting to obtain iron concentrate;

the gold grade in the gold raw ore in the step (1) is 1-3g/t, and the sulfur content is more than 5%; the ore grinding classification comprises the following steps: the gold raw ore is subjected to primary ball milling and then primary screening to obtain a primary undersize product, and a secondary ball milling product obtained by secondary ball milling of the primary oversize product and the primary undersize product are mixed to obtain a classification product and slurry preparation is carried out; the mass percentage of particles with a particle size of-200 meshes in the primary undersize product accounts for 50-70% of the total mass of the primary undersize product; the content of particles with the particle size of-200 meshes in the classified product accounts for 80-95% of the total mass of the classified product in percentage by mass; the addition amount of the collecting agent is 100-500 g/t;

the temperature of the first roasting in the step (2) is 350-550 ℃; the first roasting time is 0.5-1.5 h; the temperature of the second roasting is 550-800 ℃; the second roasting time is 0.5-2 h; the leached solid phase contains particles with the size of-400 meshes in a mass percentage accounting for 75-95% of the total mass of the leached solid phase.

Technical Field

The invention relates to the field of flotation, and particularly relates to a gold ore collecting agent and application thereof.

Background

The Carlin type gold deposit refers to a fine-dip type medium-low temperature hydrothermal gold deposit produced in clastic rock, carbonate rock and silicalite which are not locally deteriorated, and is also called a fine-dip type gold deposit, a permeable hot water type gold deposit, a sedimentary rock type gold deposit and a "dip-dye type deposit in a sedimentary rock stratum which is chemically favorable for mineral formation". The gold particles in the ore are extremely fine and are wrapped in carrier minerals such as arsenopyrite, sulfide, carbonate or silicalite to exist in an invisible gold or sub-microscopic gold state.

At present, along with the exploitation and utilization of gold mineral resources, easily-exploited and high-content rich mineral resources are continuously reduced, and refractory gold smelting ores containing arsenic, carbon and micro-fine particles are concerned more and gradually become main sources of gold products. In recent years, China discovers a large number of Carlin type gold ores in areas such as Qinling mountains of West and Guiqian, and the like, has abundant resource quantity and has mining and utilization values.

Gold particles in the Carlin type gold ore are uniformly distributed in sulfide ore and gangue ore and are embedded with carrier ore very finely, so that even if the sulfide ore is completely enriched into gold concentrate, a better enrichment ratio and a better gold recovery rate cannot be obtained; in addition, the prior ore grinding technology and assembly are difficult to effectively dissociate the gangue mineral from gangue mineral monomers, so that the flotation recovery rate is low. However, at present, the flotation difficulty of the microfine dip-dyeing gold ore is higher.

For example, CN107213992A discloses a copper-gold-silver ore flotation collector and a flotation method, which belong to the technical field of mineral flotation and solve the outstanding problems of low beneficiation efficiency, high cost and the like of copper-gold-silver ore in the prior art. The collector ester-205 was made from the following raw materials: 30-40 parts of N, N-diethyl dithiocarbamic acid propionitrile ester, 20-25 parts of diisopropyl ammonium dithiophosphate, 10-15 parts of nigroic acid, 20-30 parts of pine oil and 2-5 parts of ethanol; the flotation method comprises the steps of preparing flotation ore pulp, preparing flotation and carrying out copper-gold-silver ore flotation. Multiple circulation and metal loss of target metal minerals in the flotation process are avoided, the main flotation process is shortened, and investment and operation cost are reduced; the collecting agent ester-205 has strong collecting capability, improves the flotation speed, improves the recovery rate of copper, gold and silver, reduces the scavenging times and reduces the types and the dosage of flotation reagents.

CN104259010A discloses a novel gold flotation collector, which is composed of butyl xanthate, isoamyl xanthate and butylamine black powder. The weight ratio of the butyl xanthate to the isoamyl xanthate to the butylamine black is 4-8:1-5: 0.5-2. Preferably, the weight ratio of the butyl xanthate to the isoamyl xanthate to the butylamine black is 6: 3: 1. the traditional xanthate with a single carbon chain is broken through, different collectors and selectivities of different carbon chains are adopted, the collector of the butyl xanthate is complementary to the selectivity of the isoamyl xanthate, the selectivity of the butylamine black is increased, the three reagents are fully mixed, the spear of the collecting capacity and the selectivity is resolved, and the collecting performance and the selectivity are exerted to the maximum. Compared with the traditional medicament, the recovery rate of the novel medicament is improved by 0.5 percent, the dosage of the medicament is reduced by 12g/t, and the concentrate grade is improved by 0.74 g/t.

High-sulfur Carlin type gold ore has poor flotation effect due to high sulfur (> 5%) content, and high gold concentrate grade is difficult to obtain, so the ore is difficult to be effectively utilized.

Disclosure of Invention

In view of the problems in the prior art, the invention aims to provide a gold ore collecting agent and application thereof, which realize the high-efficiency enrichment of gold in high-sulfur Carlin type gold ores by adjusting the formula of the collecting agent, and realize the comprehensive utilization of sulfur and iron in the ores by a specific roasting process.

In order to achieve the purpose, the invention adopts the following technical scheme:

in a first aspect, the present invention provides a gold ore collector comprising: a mixture of isoamyl xanthate and N-alkyl dithiocarbamate dialkylester of phosphinic acid.

The collector provided by the invention has the advantages that through the specific matching between the isoamyl xanthate and the N-alkyl dithiocarbamic acid methylene phosphonic acid dialkyl ester, the ion complex generated after the dissociation of the isoamyl xanthate and the N-alkyl dithiocarbamic acid methylene phosphonic acid dialkyl ester and the metal ion form a chelate complex, the collector has higher selectivity on minerals such as gold-loaded pyrite, arsenopyrite and the like, the high-efficiency enrichment of gold in high-sulfur gold ores is realized, meanwhile, the high-efficiency utilization of the high-sulfur Carlin gold ores is realized through a specific roasting process, and the comprehensive utilization of sulfur and iron in the ores is realized.

In the invention, other solvents for dissolving the collector or components convenient for exerting the performance of the medicament may also exist in the collector, but the collector does not contain other specific effective components contained in the gold collector disclosed in the prior art; the preparation process only needs to mix according to the proportion or be supplemented with a certain solvent.

In a preferred embodiment of the present invention, the mass ratio of the isopentyl xanthate to the dialkyl N-alkyldithiocarbamic acid phosphinate in the collector is (1-4):1, and the collector may be, for example, 1:1, 1:1.2, 1:1.4, 1:1.6, 1:1.8, 1:2, 1:2.2, 1:2.4, 1:2.6, 1:2.8, 1:3, 1:3.2, 1:3.4, 1:3.6, 1:3.8, or 1:4, but is not limited to the above-mentioned values, and other combinations not listed within this range are also applicable.

In a second aspect, the present invention provides the use of a gold ore collector as in the first aspect, the use comprising the steps of:

(1) grinding and grading gold raw ore, and adding a flotation reagent containing the collecting agent according to claim 1 or 2 to perform rough concentration, fine concentration and scavenging in sequence to obtain gold concentrate;

(2) and (2) sequentially carrying out first roasting, second roasting, leaching and solid-liquid separation on the gold concentrate obtained in the step (1) to obtain a gold-containing liquid phase and leaching residues, and carrying out magnetic separation on the leaching residues after magnetization roasting to obtain iron concentrate.

As a preferred embodiment of the present invention, the gold grade in the gold raw ore in step (1) is 1 to 3g/t, the sulfur content is > 5%, for example, 1g/t, 1.5g/t, 2g/t, 2.5g/t or 3g/t, for example, 5.5%, 6%, 7%, 8%, 9% or 10%, etc., but not limited to the values listed, and other combinations not listed within this range are also applicable.

Preferably, the ore grinding classification of step (1) comprises: and after primary ball milling, primary screening the gold raw ore to obtain a primary undersize product, and mixing a secondary ball milling product obtained by secondary ball milling of the primary oversize product with the primary undersize product to obtain a classification product and mixing the classification product with slurry.

Preferably, the content of the particles of the primary undersize product of-200 mesh in percentage by mass is 50 to 70% of the total mass of the primary undersize product, for example, 50%, 51%, 52%, 53%, 54%, 55%, 56%, 57%, 58%, 59%, 60%, 61%, 62%, 63%, 64%, 65%, 66%, 67%, 68%, 69%, or 70% and the like, but not limited to the values listed, and other combinations not listed in this range are also applicable.

Preferably, the content of the particles of-200 mesh in the fractionated product is 80 to 95% by mass based on the total mass of the fractionated product, and may be, for example, 80%, 81%, 82%, 83%, 84%, 85%, 86%, 87%, 88%, 89%, 90%, 91%, 92%, 93%, 94%, or 95%, and the like, but is not limited to the values listed, and other combinations not listed in this range are also applicable.

As a preferred embodiment of the present invention, the amount of the collector added in step (1) is 100-500g/t, and may be, for example, 100g/t, 150g/t, 200g/t, 250g/t, 300g/t, 350g/t, 400g/t, 450g/t or 500g/t, but is not limited to the values listed, and other combinations not listed within this range are also applicable.

Preferably, the flotation reagent in step (1) further comprises a regulator, a depressant, an activator and a foaming agent.

Preferably, the conditioning agent comprises 1 or a combination of at least 2 of sodium carbonate, oxalic acid, sulfuric acid or ammonium sulfate.

When the regulating agent is oxalic acid, the regulating agent and the collecting agent can further generate a composite effect to further activate the pyrite object and enhance the adsorption site of the collecting agent, so that the collecting effect of the collecting agent on gold ores is improved.

Preferably, the amount of the modifier is 500-2000g/t, such as 500g/t, 600g/t, 700g/t, 800g/t, 900g/t, 1000g/t, 1100g/t, 1200g/t, 1300g/t, 1400g/t, 1500g/t, 1600g/t, 1700g/t, 1800g/t, 1900g/t or 2000g/t, but not limited to the values listed, and other combinations not listed within this range are equally applicable.

Preferably, the inhibitor comprises water glass and/or sodium hexametaphosphate.

Preferably, the inhibitor is added in an amount of 800-1500g/t, such as 800g/t, 850g/t, 900g/t, 950g/t, 1000g/t, 1050g/t, 1100g/t, 1150g/t, 1200g/t, 1250g/t, 1300g/t, 1350g/t, 1400g/t, 1450g/t, or 1500g/t, but not limited to the recited values, and other combinations not recited within this range are equally applicable.

Preferably, the activator comprises copper sulfate and/or lead sulfate.

Preferably, the activator is added in an amount of 200-600g/t, such as 200g/t, 250g/t, 300g/t, 350g/t, 400g/t, 450g/t, 500g/t, 550g/t or 600g/t, but not limited to the recited values, and other combinations not recited within this range are equally applicable.

Preferably, the foaming agent comprises pinitol oil and/or diethyl phthalate.

Preferably, the blowing agent is added in an amount of 10 to 100g/t, for example, 10g/t, 20g/t, 30g/t, 40g/t, 50g/t, 60g/t, 70g/t, 80g/t, 90g/t or 100g/t, etc., but not limited to the recited values, and other combinations not recited within this range are also applicable.

As a preferred embodiment of the present invention, the concentration in the step (1) is carried out at least 2 times, for example, 2 times, 3 times, 4 times or 5 times, but the present invention is not limited to the above-mentioned values, and other combinations not shown in the above-mentioned range are also applicable.

Preferably, the sweeping in step (1) is performed at least 3 times, such as 3, 4, 5 or 6 times, but not limited to the recited values, and other combinations not recited within this range are equally applicable.

As a preferred embodiment of the present invention, the temperature of the first baking in the step (2) is 350-550 ℃, for example 350 ℃, 360 ℃, 370 ℃, 380 ℃, 390 ℃, 400 ℃, 410 ℃, 420 ℃, 430 ℃, 440 ℃, 450 ℃, 460 ℃, 470 ℃, 480 ℃, 490 ℃, 500 ℃, 510 ℃, 520 ℃, 530 ℃, 540 ℃ or 550 ℃, but is not limited to the values listed, and other combinations not listed within the range are also applicable.

Preferably, the first calcination time in step (2) is 0.5-1.5h, such as 0.5h, 0.6h, 0.7h, 0.8h, 0.9h, 1h, 1.1h, 1.2h, 1.3h, 1.4h, or 1.5h, but not limited to the recited values, and other combinations not recited in this range are also applicable.

Preferably, the temperature of the second baking in step (2) is 550-.

Preferably, the second calcination in step (2) is carried out for 0.5-2h, such as 0.5h, 0.6h, 0.7h, 0.8h, 0.9h, 1h, 1.1h, 1.2h, 1.3h, 1.4h, 1.5h, 1.6h, 1.7h, 1.8h, 1.9h or 2h, etc., but not limited to the recited values, and other combinations not recited in this range are also applicable.

In a preferred embodiment of the present invention, the content of the particles of-400 mesh in the leached solid phase in step (2) is 75 to 95% by mass of the total mass of the leached solid phase, and may be, for example, 75%, 78%, 80%, 82%, 84%, 86%, 88%, 90%, 92%, 94%, or 95%, but not limited to the above-mentioned values, and other combinations not shown in the above-mentioned range are also applicable.

Preferably, the solid-to-liquid ratio g/mL in the leaching in the step (2) is 1 (1-5), and may be, for example, 1:1, 1:1.5, 1:2, 1:2.5, 1:3, 1:3.5, 1:4, 1:4.5 or 1:5, but is not limited to the values listed, and other combinations not listed within the range are also applicable.

Preferably, the leaching of step (2) comprises a pretreatment with an acid followed by leaching with a cyanate.

Preferably, the acid comprises sulfuric acid and/or hydrochloric acid.

Preferably, the pretreatment time is 2 to 6 hours, and may be, for example, 2 hours, 2.5 hours, 3 hours, 3.5 hours, 4 hours, 4.5 hours, 5 hours, 5.5 hours, or 6 hours, etc., but is not limited to the recited values, and other combinations not recited within the range are also applicable.

Preferably, the cyanate is added in an amount of 0.5 to 4000g/t, and may be, for example, 0.5g/t, 1g/t, 2g/t, 4g/t, 6g/t, 8g/t, 10g/t, 50g/t, 100g/t, 200g/t, 400g/t, 600g/t, 800g/t, 1000g/t, 1500g/t, 2000g/t, 2500g/t, 3000g/t, 3500g/t or 4000g/t, etc., but is not limited to the values recited, and other combinations not recited within this range are equally suitable.

Preferably, the density of the bed carbon in the leach is 8-16g/L, and may be, for example, 8g/L, 8.5g/L, 9g/L, 9.5g/L, 10g/L, 10.5g/L, 11g/L, 11.5g/L, 12g/L, 12.5g/L, 13g/L, 13.5g/L, 14g/L, 14.5g/L, 15g/L, 15.5g/L, or 16g/L, and the like, but is not limited to the recited values, and other non-recited combinations within this range are equally applicable.

Preferably, the leaching time is 24-60h, such as 24h, 25h, 30h, 35h, 40h, 45h, 50h, 55h or 60h, etc., but not limited to the recited values, and other combinations not recited within this range are equally applicable.

As a preferable technical scheme of the invention, a reducing agent is added in the magnetizing roasting in the step (2) for roasting.

Preferably, the reductant comprises coal fines.

Preferably, the amount of the reducing agent added is 2 to 12% by mass of the leached residue, and may be, for example, 2%, 3%, 4%, 5%, 6%, 7%, 8%, 9%, 10%, 11%, 12% or the like, but is not limited to the values listed, and other combinations not listed in this range are also applicable.

Preferably, the magnetizing roasting temperature in step (2) is 600-.

Preferably, the magnetizing roasting time in step (2) is 0.5-1.5h, such as 0.5h, 0.6h, 0.7h, 0.8h, 0.9h, 1h, 1.1h, 1.2h, 1.3h, 1.4h or 1.5h, but not limited to the recited values, and other combinations not recited in the range are also applicable.

As a preferred technical scheme of the invention, the application comprises the following steps:

(1) grinding and grading the gold raw ore, and adding a flotation agent containing the collecting agent to perform rough concentration, fine concentration and scavenging in sequence to obtain gold concentrate;

(2) sequentially carrying out first roasting, second roasting, leaching and solid-liquid separation on the gold concentrate obtained in the step (1) to obtain a gold-containing liquid phase and leaching residues, and carrying out magnetic separation on the leaching residues after magnetization roasting to obtain iron concentrate;

the gold grade in the gold raw ore in the step (1) is 1-3g/t, and the sulfur content is more than 5%; the ore grinding classification comprises the following steps: the gold raw ore is subjected to primary ball milling and then primary screening to obtain a primary undersize product, and a secondary ball milling product obtained by secondary ball milling of the primary oversize product and the primary undersize product are mixed to obtain a classification product and slurry preparation is carried out; the mass percentage of particles with a particle size of-200 meshes in the primary undersize product accounts for 50-70% of the total mass of the primary undersize product; the content of particles with the particle size of-200 meshes in the classified product accounts for 80-95% of the total mass of the classified product in percentage by mass; the addition amount of the collecting agent is 100-500 g/t;

the temperature of the first roasting in the step (2) is 350-550 ℃; the first roasting time is 0.5-1.5 h; the temperature of the second roasting is 550-800 ℃; the second roasting time is 0.5-2 h; the leached solid phase contains particles with the size of-400 meshes in a mass percentage accounting for 75-95% of the total mass of the leached solid phase.

In the flotation process, the addition amount unit g/t of the medicament in the flotation process is to add a certain mass of the medicament according to the mass of the solid in the ore pulp.

In the invention, the scavenged concentrate returns to the previous stage of flotation, scavenged tailings are fed into scavenged or returned to roughing, or the scavenged 1 tailings and the scavenged 1 tailings are combined and returned to roughing operation, the tailings in the fine concentration 2 return to the fine concentration 1,

compared with the prior art, the invention at least has the following beneficial effects:

(1) aiming at high-sulfur Carlin type gold ores, sectional grinding and classification are adopted, so that the phenomenon of argillization caused in the fine grinding process is effectively avoided, and the recovery rate of flotation gold is improved. The gold grade of the obtained high-sulfur gold concentrate is more than 8g/t, the sulfur grade is more than 30%, the gold recovery rate can reach more than 90%, and the flue gas generated by roasting is used for preparing sulfuric acid after the obtained high-sulfur gold concentrate is roasted.

(2) The purpose of comprehensively utilizing the sulfur and the iron in the ore can be achieved by roasting, leaching and roasting gold concentrate, leaching and cyaniding tailings magnetizing roasting and magnetic separation of iron. The magnetic separation tailings are used for making bricks, cementing materials or cement admixtures and the like.

Detailed Description

To better illustrate the invention and to facilitate the understanding of the technical solutions thereof, typical but non-limiting examples of the invention are as follows:

example 1

This embodiment provides a gold ore collector, the collector includes: a mixture of isoamyl xanthate and N-alkyl dithiocarbamate methylene phosphonic acid dialkyl ester; the mass ratio of isoamyl xanthate to N-alkyl dithiocarbamic acid methylene phosphonic acid dialkyl ester in the collecting agent is 2: 1;

the application is carried out by the following process:

step one, grinding the raw ore by 55% with the fineness of-200 meshes at the first stage, grinding the classified coarse fraction product by the second stage, and finally grinding the coarse fraction product by 85% with the fineness of-200 meshes.

Adding 1000g/t of sodium carbonate serving as a regulator into the ore pulp respectively; inhibitor water glass and sodium hexametaphosphate are 800 and 400 g/t; activator copper sulfate 400 g/t; 400g/t of collecting agent; 40g/t of a foaming agent, namely, terpineol oil; the flotation process comprises 2 times of concentration and 3 times of scavenging.

Step two, roasting the high-sulfur gold concentrate in a fluidized bed for two sections, wherein flue gas generated by roasting is used for preparing sulfuric acid, and the roasting conditions are as follows: the first-stage roasting temperature is 500 ℃, the roasting time is 1.0h, and the second-stage roasting temperature is 700 ℃, and the roasting time is 1 h.

And step three, leaching gold from the calcine, wherein the grinding fineness of the calcine is-400 meshes and accounts for 90%, the ratio of the calcine to water is 1:4, the pretreatment time of sulfuric acid is 4 hours, the pH value is adjusted to 11-13 by calcium oxide, the using amount of sodium cyanide is 2000g/t, and after gold extraction is carried out for 48 hours under the condition that the density of bottom carbon is 12g/L, cyanide tailings are obtained.

And step four, recycling iron resources in the cyanidation slag, adding 8% of coal powder into the cyanidation tailings, carrying out magnetization roasting for 1h at the temperature of 750 ℃, and then obtaining iron ore concentrate by adopting a coarse scanning closed-loop magnetic separation process.

The gold grade of the high-sulfur gold concentrate obtained by flotation is 8.17g/t, the sulfur grade is 35.63 percent, and the gold recovery rate is 91.23 percent; the leaching rate of gold in the roasting leaching operation is 82.26 percent, and the grade of magnetized roasting iron is 58.38 percent.

Example 2

This embodiment provides a gold ore collector, the collector includes: a mixture of isoamyl xanthate and N-alkyl dithiocarbamate methylene phosphonic acid dialkyl ester; the mass ratio of isoamyl xanthate to N-alkyl dithiocarbamic acid methylene phosphonic acid dialkyl ester in the collecting agent is 2.5: 1;

the application is carried out by the following process:

step one, grinding the raw ore by a primary grinding fineness of-200 meshes accounting for 60%, and grinding the classified coarse fraction product by a secondary grinding to obtain the ground ore with a final grinding fineness of-200 meshes accounting for 88%.

Respectively adding 1500g/t of oxalic acid serving as a regulator into the ore pulp; inhibitor water glass and sodium hexametaphosphate are 1000+500 g/t; activator copper sulfate 500 g/t; the dosage of the collecting agent is 350 g/t; the dosage of a foaming agent diethyl phthalate is 100 g/t; the flotation process comprises 2 times of concentration and 3 times of scavenging.

Step two, roasting the high-sulfur gold concentrate in a fluidized bed for two sections, wherein flue gas generated by roasting is used for preparing sulfuric acid, and the roasting conditions are as follows: the first-stage roasting temperature is 550 ℃, the roasting time is 1.0h, and the second-stage roasting temperature is 700 ℃, and the roasting time is 1 h.

And step three, leaching gold from the calcine, wherein the grinding fineness of the calcine is 92 percent with-400 meshes, the ratio of the calcine to water is 1:4, the pretreatment time of sulfuric acid is 4 hours, the pH value is adjusted to 11-13 by calcium oxide, the using amount of sodium cyanide is 2000g/t, and after the gold is leached for 48 hours under the condition that the density of bottom carbon is 12g/L, cyanide tailings are obtained.

And step four, recycling iron resources in the cyanidation slag, adding 10% of coal powder into the cyanidation tailings, carrying out magnetization roasting for 1.5h at the temperature of 750 ℃, and then obtaining iron ore concentrate by adopting a coarse scanning closed-loop magnetic separation process.

The gold grade of the high-sulfur gold concentrate obtained by flotation is 8.43g/t, the sulfur grade is 38.86 percent, and the gold recovery rate is 92.36 percent; the leaching rate of gold in the roasting leaching operation is 83.74%, and the grade of magnetized roasting iron is 60.02%.

Example 3

This embodiment provides a gold ore collector, the collector includes: a mixture of isoamyl xanthate and N-alkyl dithiocarbamate methylene phosphonic acid dialkyl ester; the mass ratio of isoamyl xanthate to N-alkyl dithiocarbamic acid methylene phosphonic acid dialkyl ester in the collecting agent is 3: 1;

the application is carried out by the following process:

step one, grinding the raw ore by a first stage to obtain a coarse fraction product with a fineness of-200 meshes accounting for 65%, and grinding the coarse fraction product by a second stage to obtain a final ground ore with a fineness of-200 meshes accounting for 88%.

Adding 1200g/t of oxalic acid serving as a regulator into the ore pulp respectively; inhibitor water glass and sodium hexametaphosphate are 800 and 600 g/t; activator copper sulfate 400 g/t; the dosage of the collecting agent is 350 g/t; 40+20g/t of pinitol oil and diethyl phthalate serving as foaming agents; the flotation process comprises 2 times of concentration and 3 times of scavenging.

Step two, roasting the high-sulfur gold concentrate in a fluidized bed for two sections, wherein flue gas generated by roasting is used for preparing sulfuric acid, and the roasting conditions are as follows: the first-stage roasting temperature is 500 ℃, the roasting time is 1.0h, and the second-stage roasting temperature is 700 ℃, and the roasting time is 1 h.

And step three, leaching gold from the calcine, wherein the grinding fineness of the calcine is 92 percent with-400 meshes, the ratio of the calcine to water is 1:3, the pretreatment time of sulfuric acid is 6 hours, the pH value is adjusted to 11-13 by calcium oxide, the using amount of sodium cyanide is 3000g/t, and after gold extraction is carried out for 48 hours under the condition that the density of bottom carbon is 14g/L, cyanide tailings are obtained.

And step four, recycling iron resources in the cyanidation slag, adding 8% of coal powder into the cyanidation tailings, carrying out magnetization roasting for 1h at the temperature of 750 ℃, and then obtaining iron ore concentrate by adopting a coarse scanning closed-loop magnetic separation process.

The gold grade of the high-sulfur gold concentrate obtained by flotation is 8.05g/t, the sulfur grade is 36.74 percent, and the gold recovery rate is 91.83 percent; the leaching rate of gold in the roasting leaching operation is 83.71 percent, and the grade of magnetized roasting iron is 59.64 percent.

Example 4

This embodiment provides a gold ore collector, the collector includes: a mixture of isoamyl xanthate and N-alkyl dithiocarbamate methylene phosphonic acid dialkyl ester; the mass ratio of isoamyl xanthate to N-alkyl dithiocarbamic acid methylene phosphonic acid dialkyl ester in the collecting agent is 4: 1;

the application is carried out by the following process:

step one, grinding the raw ore by a primary grinding fineness of-200 meshes accounting for 60%, and grinding the classified coarse fraction product by a secondary grinding to obtain the ground ore with a final grinding fineness of-200 meshes accounting for 85%.

Respectively adding 800g/t of 10% sulfuric acid serving as a regulator into the ore pulp; inhibitor water glass 1000 g/t; the activator is copper sulfate and lead nitrate 300+100 g/t; 400g/t of collecting agent and 50g/t of foaming agent terpineol oil; the flotation process comprises 2 times of concentration and 3 times of scavenging.

Step two, roasting the high-sulfur gold concentrate in a fluidized bed for two sections, wherein flue gas generated by roasting is used for preparing sulfuric acid, and the roasting conditions are as follows: the first-stage roasting temperature is 550 ℃, the roasting time is 1.0h, and the second-stage roasting temperature is 700 ℃, and the roasting time is 1.5 h.

And step three, leaching gold from the calcine, wherein the grinding fineness of the calcine is 92 percent with-400 meshes, the ratio of the calcine to water is 1:4, the pretreatment time of sulfuric acid is 4 hours, the pH value is adjusted to 11-13 by calcium oxide, the using amount of sodium cyanide is 2500g/t, and after gold extraction is carried out for 48 hours under the condition that the density of bottom carbon is 13g/L, cyanide tailings are obtained.

And step four, recycling iron resources in the cyanidation slag, adding 10% of coal powder into the cyanidation tailings, carrying out magnetization roasting for 1.5h at the temperature of 750 ℃, and then obtaining iron ore concentrate by adopting a coarse scanning closed-loop magnetic separation process.

The gold grade of the high-sulfur gold concentrate obtained by flotation is 8.16g/t, the sulfur grade is 35.24 percent, and the gold recovery rate is 90.25 percent; the leaching rate of gold in the roasting and leaching operation is 82.68 percent, and the grade of magnetized and roasted iron is 58.46 percent.

Example 5

The difference from the embodiment 2 is only that oxalic acid is replaced by sulfuric acid with the same concentration, and the gold grade of the high-sulfur gold concentrate obtained by flotation is 7.86g/t, the sulfur grade is 34.14 percent, and the gold recovery rate is 88.25 percent; the leaching rate of gold in the roasting leaching operation is 81.38%, and the grade of magnetized roasting iron is 57.16%.

Comparative example 1

The difference from the embodiment 1 is only that the isoamyl xanthate in the collector is replaced by the same amount of butyl black, and the gold grade of the high-sulfur gold concentrate obtained by flotation is 7.66g/t, the sulfur grade is 33.42 percent, and the gold recovery rate is 89.25 percent; the leaching rate of gold in the roasting leaching operation is 80.68%, and the grade of magnetized roasting iron is 57.15%.

Comparative example 2

The difference from the embodiment 1 is only that the isoamyl xanthate in the collector is replaced by the same amount of isobutyl black, and the gold grade of the high-sulfur gold concentrate obtained by flotation is 7.98g/t, the sulfur grade is 33.78 percent, and the gold recovery rate is 87.63 percent; the leaching rate of gold in the roasting leaching operation is 81.37%, and the grade of magnetized roasting iron is 57.32%.

Comparative example 3

The difference from the example 1 is only that the dialkyl N-alkyl dithiocarbamic acid methylene phosphonic acid ester in the collector is replaced by equal amount of alcohol black (sodium diisoamyl dithiophosphate), and the gold grade of the high-sulfur gold concentrate obtained by flotation is 7.89g/t, the sulfur grade is 32.63 percent, and the gold recovery rate is 88.69 percent; the leaching rate of gold in the roasting leaching operation is 81.08 percent, and the grade of magnetized roasting iron is 58.06 percent.

Comparative example 4

The method is different from the example 1 only in that the dialkyl N-alkyl dithiocarbamic acid methylene phosphonic acid ester in the collector is replaced by the same amount of the siloxane alcohol black (sodium polyoxybutenol dithiophosphate), the gold grade of the high-sulfur gold concentrate obtained by flotation is 7.82g/t, the sulfur grade is 32.39 percent, and the gold recovery rate is 85.58 percent; the leaching rate of gold in the roasting leaching operation is 81.57%, and the grade of magnetized roasting iron is 57.67%.

Comparative example 5

The method is different from the embodiment 1 only in that the mass ratio of isoamyl xanthate to N-alkyl dithiocarbamic acid methylene phosphonic acid dialkyl ester in the collecting agent is 0.5:1, the gold grade of the high-sulfur gold concentrate obtained by flotation is 8.04g/t, the sulfur grade is 33.98 percent, and the gold recovery rate is 87.68 percent; the leaching rate of gold in the roasting leaching operation is 80.39%, and the grade of magnetized roasting iron is 57.85%.

Comparative example 6

The difference from the example 1 is only that the mass ratio of the isoamyl xanthate to the N-alkyl dithiocarbamic acid methylene phosphonic acid dialkyl ester in the collecting agent is 6:1, the gold grade of the high-sulfur gold concentrate obtained by flotation is 7.51g/t, the sulfur grade is 33.67 percent, and the gold recovery rate is 89.47 percent; the leaching rate of gold in the roasting leaching operation is 81.47%, and the grade of magnetized roasting iron is 57.86%.

According to the results of the above examples and comparative examples, the invention realizes the high-efficiency enrichment of gold in the high-sulfur gold ores by adjusting the formula of the collecting agent, and simultaneously realizes the high-efficiency utilization of the high-sulfur Carlin gold ores and the comprehensive utilization of sulfur and iron in the ores by using a specific roasting process.

The applicant declares that the present invention illustrates the detailed structural features of the present invention through the above embodiments, but the present invention is not limited to the above detailed structural features, that is, it does not mean that the present invention must be implemented depending on the above detailed structural features. It should be understood by those skilled in the art that any modifications of the present invention, equivalent substitutions of selected components of the present invention, additions of auxiliary components, selection of specific modes, etc., are within the scope and disclosure of the present invention.

The preferred embodiments of the present invention have been described in detail, however, the present invention is not limited to the specific details of the above embodiments, and various simple modifications may be made to the technical solution of the present invention within the technical idea of the present invention, and these simple modifications are within the protective scope of the present invention.

It should be noted that the various technical features described in the above embodiments can be combined in any suitable manner without contradiction, and the invention is not described in any way for the possible combinations in order to avoid unnecessary repetition.

In addition, any combination of the various embodiments of the present invention is also possible, and the same should be considered as the disclosure of the present invention as long as it does not depart from the spirit of the present invention.