CN102690942A - Method for reduction roasting and fortified acid dissolving of sulfur-containing gold concentrate calcine - Google Patents
Method for reduction roasting and fortified acid dissolving of sulfur-containing gold concentrate calcine Download PDFInfo
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- CN102690942A CN102690942A CN2012102002587A CN201210200258A CN102690942A CN 102690942 A CN102690942 A CN 102690942A CN 2012102002587 A CN2012102002587 A CN 2012102002587A CN 201210200258 A CN201210200258 A CN 201210200258A CN 102690942 A CN102690942 A CN 102690942A
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- roasting
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Abstract
The invention discloses a method for reduction roasting and fortified acid dissolving of sulfur-containing gold concentrate calcine, aiming to solve the problems that gold in conventional calcine is wrapped by fusible vitreous body in the high-temperature process and the gold leaching rate is difficult to increase during the follow-up cyaniding treatment. The method for the reduction roasting and the fortified acid dissolving of the sulfur-containing gold concentrate calcine, disclosed by the invention, comprises the following steps of: (1) carrying out the reduction roasting on the sulfur-containing gold concentrate calcine by using a reducing agent; (2) carrying out acid leaching on the calcine after being subjected to the reduction roasting by adopting a strong acid solution; and (3) filtering the reduced calcine after being subjected to the acid leaching and using the filter residue in cyaniding extraction of the gold. By using the method disclosed by the invention, the iron leaching rate can be up to more than 95%, which is one time higher than a traditional method and 8-10% higher than a method in which a cosolvent is used to enhance the leaching; the gold recovery rate is increased by 3-5%; the gold leaching rate in follow-up cyaniding extraction process is 95-98%, which is increased by 15-20% than a direct cyaniding method; and the acid use ratio is increased and the pickle liquor is more beneficial to the follow-up treatment.
Description
Technical field
The present invention relates to a kind of smelting process of gold ores containing sulfur, the method that intensified acid dissolves after the reducing roasting of particularly a kind of gold ores containing sulfur fluidized bed roasting slag.
Background technology
China exists sulfurous iron ore type and mispickel type gold mine in a large number, can be high sulphur gold concentrate with its enrichment through flotation operation, is difficult to directly adopt cyanide process therefrom to extract gold and silver.Fluidized bed roasting can generate SO with the sulphur oxidation in the concentrate
2Be used for relieving haperacidity, adopt the arsenic of two sections fluidized bed roastings in especially can the deep removal gold ore, in China's high sulphur gold concentrate smelting process, be used widely.But owing to passed through high-temperature calcination process, the gold in the calcining by flux body parcel, makes that gold leaching rate was difficult to improve when follow-up cyaniding was handled in pyroprocess; Contain gold after the cyaniding in the tailings still at 10～20g/t; Gold recovery less than 80%, serious waste of resources needs development new technologies to handle.
Mention among the Chinese patent CN1304609C with sulfuric acid or hydrochloric acid reinforcement acidleach and dissolve the iron in the calcining, make the fully exposure of parcel gold,, can significantly improve gold recovery acid leaching residue being carried out the cyaniding processing.Iron is in a large number with Fe in the calcining
2O
3There is Fe
2O
3Difficulty is gone out by acidleach, adopt solubility promoter YD 30 (EDTA), CuO etc. can to a certain degree improve dissolution rate and dissolution rate, but solubility promoter is expensive, needs to increase very big cost.
Summary of the invention
In order to remedy the deficiency of prior art, meet the needs of production, the present invention aims to provide the method that intensified acid dissolves after the reducing roasting of a kind of gold ores containing sulfur fluidized bed roasting slag, and this method is cheap and efficiently with the Fe in the roasting
2O
3Convert the lyotrope kind into, realize the efficient leaching of iron, and then the final recovery that improves gold.
To achieve these goals, the technical scheme of the present invention's employing is:
The method that a kind of sulfur containing aurin ore-roasting reducing roasting intensified acid dissolves is characterized in that, comprises the steps:
1) the sulfur containing aurin ore-roasting is carried out reducing roasting under the reductive agent effect;
2) adopt strong acid solution to carry out acidleach the calcining after the reducing roasting;
3) reduced calcine after the acidleach is filtered, filter residue is used for cyanidation gold-extracted.
According to embodiments of the invention, described reductive agent is the carbonaceous reducing agent of massfraction 5%～80%.Said carbonaceous reducing agent is preferably coal dust or coke powder, and when carbonaceous reducing agent was coal dust, pulverized coal consumption was 150～500kg/t calcining, and when carbonaceous reducing agent was coke powder, the coke powder consumption was 100～400kg/t calcining,
According to embodiments of the invention, described reductive agent is the gas base reductant of volume(tric)fraction 1%～50%.Said gas base reductant is preferably coal gas, hydrogen or Sweet natural gas; When selecting coal gas, gas concentration is 3%～20%; When selecting hydrogen, density of hydrogen is 1%～5%; When selecting Sweet natural gas, concentration of natural gas is 5%～30%.
According to embodiments of the invention, said strong acid solution is the hydrochloric acid or the sulfuric acid of massfraction 5～40%.
Further, in step 1) of the present invention, the temperature of preferred said reducing roasting is 300 ℃～1050 ℃, and the reducing roasting time is 0.5h～5h.More preferably said reducing roasting temperature is 500～800 ℃, and the reducing roasting time is 0.5h～3h.
In step 2 of the present invention) in, said acidleach adopts two sections adverse currents to leach, and wherein one section acidleach employing two-stage acid leaching liquid is leaching agent, and the acidleach temperature is 30～80 ℃, and calcining and leach liquor weight ratio are 1: 1～10, extraction time is 0.5～4 hour; One section acidleach finishes after-filtration; Filter residue after the filtration carries out two-stage acid leaching, and the sulfuric acid or the concentration of hydrochloric acid that are used for the two-stage acid leaching reduced calcine are 1～8mol/L, and the acidleach temperature is 30～80 ℃; Roasting and leach liquor weight ratio are 1: 1～10, and extraction time is 0.5h～4h; Two-stage acid leaching finishes after-filtration and obtains filtrating and filter residue, and filter residue is used for cyanidation gold-extracted, and filtrating is returned one section acidleach.
Preferred implementation condition of the present invention does, one section acidleach condition is 50～80 ℃ for the acidleach temperature, and calcining and leach liquor weight ratio do, 1: 2～6, and extraction time is 1～2 hour; The two-stage acid leaching condition is sulfuric acid or concentration of hydrochloric acid 2～6mol/L, and the acidleach temperature is 50～80 ℃, and calcining and leach liquor weight ratio are 1: 4～8, and extraction time is 2h～4h.
Below the present invention is done further description.
The method that a kind of sulfur containing aurin ore-roasting reducing roasting intensified acid dissolves comprises the steps:
With sulfur containing aurin ore-roasting and massfraction is that 5%～80% carbonaceous reducing agent coal dust or coke powder mix the back and get in the reduction roaster and reduce, or directly to adopt volume(tric)fraction be reduced calcine in reduction roaster such as 1%～50% gas base reductant coal gas, hydrogen, Sweet natural gas.
Wherein reducing roasting temperature is 300～1050 ℃, and the reducing roasting time is 0.5～5 hour, the hydrochloric acid or the sulfuric acid shrend of direct water of reduced calcine or massfraction 5～40% after roasting is accomplished.
Under the reductive action of reductive agent, the Fe in the calcining
2O
3Be converted into FeO, be easy to leached by sulfuric acid or hydrochloric acid.The acidleach process adopts two sections adverse currents to leach, and it is leaching agent that two-stage acid leaching liquid is adopted in one section acidleach, and the acidleach temperature is 30～80 ℃, and calcining and leach liquor weight ratio are 1: 1～10, and extraction time is 0.5～4 hour.One section acidleach finishes after-filtration, and filtrating is used for comprehensively reclaiming valuable constituents such as copper, iron, and filter residue carries out two-stage acid leaching.The sulfuric acid or the concentration of hydrochloric acid that are used for the two-stage acid leaching reduced calcine are 1～8mol/L, and the acidleach temperature is 30～80 ℃, and roasting and leach liquor weight ratio are 1: 1～10, and extraction time is 0.5～4 hour.Two-stage acid leaching finishes the back direct filtration and obtains filtrating and filter residue, and filter residue directly is used for cyanidation gold-extracted, and filtrating is returned one section acidleach.
The preferred solid reductant of reducing roasting in aforesaid method; It is that 150～500kg/t calcining or coke powder consumption are 100～400kg/t calcining that this solid reductant is preferably pulverized coal consumption; Certainly, reducing roasting also can be adopted the gas base reductant, the Sweet natural gas of the coal gas that said gas base reductant is a concentration 3%～20% or the hydrogen of concentration 1%～5% or concentration 5%～30%; The reducing roasting temperature is 500～800 ℃, and the reducing roasting time is 0.5～3 hour.
Preferred one section acidleach condition is 50～80 ℃ for the acidleach temperature in the aforesaid method, and calcining and leach liquor weight ratio are 1: 2～6, and extraction time is 1～2 hour; The two-stage acid leaching condition is sulfuric acid or concentration of hydrochloric acid 2～6mol/L, and the acidleach temperature is 50～80 ℃, and calcining and leach liquor weight ratio are 1: 4～8, and extraction time is 2～4 hours.
Key of the present invention is to use reductive agent at high temperature with Fe
2O
3Be reduced to FeO, the reduction reaction under the different reductive agent effects is following:
Fe
2O
3+C＝2FeO+CO
2Fe
2O
3+2C＝4FeO+CO
2
Fe
2O
3+CO＝2FeO+CO
2
Fe
2O
3+H
2＝2FeO+H
2O
3Fe
2O
3+CH
4＝6FeO+CO+2H
2O
Fe wherein
2O
3What obtain with FeO acidleach process is respectively ferric iron and ferrous iron solution; Ferric iron in the solution begins to generate deposition during greater than 2.5 left and right sides as pH; And ferrous iron also can stable existence in greater than 5 solution at pH; FeO obviously is more conducive to leach, and can obtain the low leach liquor of residual acidity, and the principle that leaches is strengthened in reducing roasting that Here it is.
Compared with prior art, the invention has the beneficial effects as follows:
(1) with Fe in the calcining
2O
3Leach after being reduced to FeO, under same acidleach temperature, leaching time and acid concentration condition, the iron leaching yield can reach more than 95%, exceeds one times than traditional method again, exceeds 8～10% than strengthening leaching method with solubility promoter.
(2) Fe in the calcining
2O
3Be reduced to FeO, gold can be by Fe in the hydrochloric acid leaching process
3+Oxidation is leached, and improves gold recovery 3～5%.
(3) leaching back slag rate is 10～15% of calcining, gold in acid leaching residue enrichment 7～10 times, and fully exposed, cyaniding leaching process gold leaching rate is compared direct cyaniding and is improved 15～20% 95～98%.
(4) two sections acidleach both can be guaranteed the iron leaching yield, can obtain one section very low leach liquor of residual acidity again, had improved sour utilization ratio, and pickling liquor also is more conducive to subsequent disposal.
Description of drawings
Fig. 1 is a process flow sheet of the present invention.
Embodiment
Embodiment 1
Gold ore calcining and the 100kg coal dust of getting 200kg mix the back under 800 ℃ in rotary kiln reductase 12 hour, calcining cooling back is subsequent use.Getting the 50kg reduced calcine, to adopt 3mol/L sulfuric acid be that 4: 1, extraction temperature are that 80 ℃, extraction time are to leach under the condition in 2 hours in liquid-solid ratio, and obtaining the iron leaching yield is 95%.
Above-mentioned leach liquor is continued to leach the 50kg reduced calcine; Liquid-solid ratio is that 4: 1, extraction temperature are that 80 ℃, extraction time are 2 hours; The iron leaching yield is 63%; Leach liquor pH is 4.4, and it is that 4: 1, extraction temperature are that 80 ℃, extraction time are to carry out two sections leachings under the condition in 2 hours in liquid-solid ratio that leached mud continue to use 3mol/L sulfuric acid, and the total leaching yield of iron is 98%.Leached mud after two sections leachings is carried out cyanidation gold-extracted, gold leaching rate is 96.4%.
Embodiment 2
The gold ore calcining of getting 200kg in the coal gas atmosphere of volume(tric)fraction 15% under 800 ℃ in rotary kiln reductase 12 hour, calcining cooling back is subsequent use.Getting the 50kg reduced calcine, to use 6mol/L hydrochloric acid be that 4: 1, extraction temperature are that 50 ℃, extraction time are to leach under the condition in 2.5 hours in liquid-solid ratio, and obtaining the iron leaching yield is 96%.
Above-mentioned leach liquor is continued to leach the 50kg reduced calcine; Liquid-solid ratio is that 4: 1, extraction temperature are that 50 ℃, extraction time are 2.5 hours; The iron leaching yield is 73%; Leach liquor pH is 5.1, and it is that 4: 1, extraction temperature are that 50 ℃, extraction time are to carry out two sections leachings under the condition in 2.5 hours in liquid-solid ratio that leached mud continue to use 6mol/L sulfuric acid, and the total leaching yield of iron is 99.2%.Leached mud after two sections leachings is carried out cyanidation gold-extracted, gold leaching rate is 97.2%.
The content that the foregoing description is illustrated is to be understood that to these embodiment and only is used to be illustrated more clearly in the present invention; And be not used in the restriction scope of the present invention; After having read the present invention, those skilled in the art all fall within the application's accompanying claims institute restricted portion to the modification of the various equivalent form of values of the present invention.
Claims (10)
1. the method that sulfur containing aurin ore-roasting reducing roasting intensified acid dissolves is characterized in that, comprises the steps:
1) the sulfur containing aurin ore-roasting is carried out reducing roasting with reductive agent;
2) adopt strong acid solution to carry out acidleach the calcining after the reducing roasting;
3) reduced calcine after the acidleach is filtered, filter residue is used for cyanidation gold-extracted.
2. the method that sulfur containing aurin ore-roasting reducing roasting intensified acid according to claim 1 dissolves is characterized in that described reductive agent is the carbonaceous reducing agent of massfraction 5%～80%.
3. the method that sulfur containing aurin ore-roasting reducing roasting intensified acid according to claim 2 dissolves; It is characterized in that; Said carbonaceous reducing agent is coal dust or coke powder, and when carbonaceous reducing agent was coal dust, pulverized coal consumption was 150～500kg/t calcining; When carbonaceous reducing agent was coke powder, the coke powder consumption was 100～400kg/t calcining.
4. the method that sulfur containing aurin ore-roasting reducing roasting intensified acid according to claim 1 dissolves is characterized in that described reductive agent is the gas base reductant of volume(tric)fraction 1%～50%.
5. the method that sulfur containing aurin ore-roasting reducing roasting intensified acid according to claim 3 dissolves is characterized in that said gas base reductant is coal gas, hydrogen or Sweet natural gas; When selecting coal gas, gas concentration is 3%～20%; When selecting hydrogen, density of hydrogen is 1%～5%; When selecting Sweet natural gas, concentration of natural gas is 5%～30%.
6. the method that sulfur containing aurin ore-roasting reducing roasting intensified acid according to claim 1 dissolves is characterized in that said strong acid solution is the hydrochloric acid or the sulfuric acid of massfraction 5～40%.
7. the method that sulfur containing aurin ore-roasting reducing roasting intensified acid according to claim 1 dissolves is characterized in that in step 1), the temperature of said reducing roasting is 300 ℃～1050 ℃, and the reducing roasting time is 0.5h～5h.
8. the method that sulfur containing aurin ore-roasting reducing roasting intensified acid according to claim 7 dissolves is characterized in that said reducing roasting temperature is 500～800 ℃, and the reducing roasting time is 0.5h～3h.
9. the method that sulfur containing aurin ore-roasting reducing roasting intensified acid according to claim 1 dissolves; It is characterized in that, in step 2) in, said acidleach adopts two sections adverse currents to leach; Wherein one section acidleach employing two-stage acid leaching liquid is leaching agent; The acidleach temperature is 30～80 ℃, and calcining and leach liquor weight ratio are 1: 1～10, and extraction time is 0.5～4 hour; One section acidleach finishes after-filtration; Filter residue after the filtration carries out two-stage acid leaching, and the sulfuric acid or the concentration of hydrochloric acid that are used for the two-stage acid leaching reduced calcine are 1～8mol/L, and the acidleach temperature is 30～80 ℃; Roasting and leach liquor weight ratio are 1: 1～10, and extraction time is 0.5h～4h; Two-stage acid leaching finishes after-filtration and obtains filtrating and filter residue, and filter residue is used for cyanidation gold-extracted, and filtrating is returned one section acidleach.
10. the method that sulfur containing aurin ore-roasting reducing roasting intensified acid according to claim 9 dissolves is characterized in that one section acidleach condition is 50～80 ℃ for the acidleach temperature, and calcining and leach liquor weight ratio are 1: 2～6, and extraction time is 1～2 hour; The two-stage acid leaching condition is sulfuric acid or concentration of hydrochloric acid 2～6mol/L, and the acidleach temperature is 50～80 ℃, and calcining and leach liquor weight ratio are 1: 4～8, and extraction time is 2h～4h.
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Cited By (10)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN103014319A (en) * | 2012-12-05 | 2013-04-03 | 中南大学 | Method for strengthening gold extraction of iron gold concentrate roasted product containing sulphur and arsenic |
| CN103194598A (en) * | 2013-04-19 | 2013-07-10 | 贵州东华工程股份有限公司 | Method for improving recovery rate of gold ore difficult to treat by adopting sulfuric acid leaching reduction process |
| CN103290232A (en) * | 2013-06-27 | 2013-09-11 | 成都理工大学 | Method for extracting gold from auriferous ores through potassium ferricyanide |
| CN103290231A (en) * | 2013-06-14 | 2013-09-11 | 潼关中金冶炼有限责任公司 | Method for improving gold leaching rate by deironing with acid leaching from arsenic material |
| CN104032141A (en) * | 2014-07-04 | 2014-09-10 | 西安建筑科技大学 | Method for recycling gold and silver from cyanidation tailings |
| CN104212974A (en) * | 2014-09-09 | 2014-12-17 | 中国科学院过程工程研究所 | Method for synchronously recovering gold, iron and cobalt or nickel from iron-containing cyanide tailings of gold ore |
| CN105671334A (en) * | 2016-03-09 | 2016-06-15 | 中南大学 | Method for extracting gold from roasted product |
| CN106086435A (en) * | 2016-07-31 | 2016-11-09 | 贵州宏达环保科技有限公司 | A kind of recovery copper, cobalt, nickel, silver method from cupric kiln slag |
| CN107760865A (en) * | 2016-08-17 | 2018-03-06 | 中国科学院过程工程研究所 | A kind of method of the dipped journey leachate gallium lithium ion enrichment of coal ash alkali |
| CN110668550A (en) * | 2019-10-18 | 2020-01-10 | 广西森合高新科技股份有限公司 | Gold concentrate non-cyanide beneficiation tailing liquid recycling treatment method |
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Cited By (12)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN103014319A (en) * | 2012-12-05 | 2013-04-03 | 中南大学 | Method for strengthening gold extraction of iron gold concentrate roasted product containing sulphur and arsenic |
| CN103194598A (en) * | 2013-04-19 | 2013-07-10 | 贵州东华工程股份有限公司 | Method for improving recovery rate of gold ore difficult to treat by adopting sulfuric acid leaching reduction process |
| CN103290231A (en) * | 2013-06-14 | 2013-09-11 | 潼关中金冶炼有限责任公司 | Method for improving gold leaching rate by deironing with acid leaching from arsenic material |
| CN103290232A (en) * | 2013-06-27 | 2013-09-11 | 成都理工大学 | Method for extracting gold from auriferous ores through potassium ferricyanide |
| CN104032141A (en) * | 2014-07-04 | 2014-09-10 | 西安建筑科技大学 | Method for recycling gold and silver from cyanidation tailings |
| CN104212974A (en) * | 2014-09-09 | 2014-12-17 | 中国科学院过程工程研究所 | Method for synchronously recovering gold, iron and cobalt or nickel from iron-containing cyanide tailings of gold ore |
| CN105671334A (en) * | 2016-03-09 | 2016-06-15 | 中南大学 | Method for extracting gold from roasted product |
| CN106086435A (en) * | 2016-07-31 | 2016-11-09 | 贵州宏达环保科技有限公司 | A kind of recovery copper, cobalt, nickel, silver method from cupric kiln slag |
| CN106086435B (en) * | 2016-07-31 | 2018-01-30 | 贵州宏达环保科技有限公司 | One kind reclaims copper, cobalt, nickel, silver-colored method from cupric kiln slag |
| CN107760865A (en) * | 2016-08-17 | 2018-03-06 | 中国科学院过程工程研究所 | A kind of method of the dipped journey leachate gallium lithium ion enrichment of coal ash alkali |
| CN110668550A (en) * | 2019-10-18 | 2020-01-10 | 广西森合高新科技股份有限公司 | Gold concentrate non-cyanide beneficiation tailing liquid recycling treatment method |
| CN110668550B (en) * | 2019-10-18 | 2022-06-21 | 广西森合高新科技股份有限公司 | Gold concentrate non-cyanide beneficiation tailing liquid recycling treatment method |
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Application publication date: 20120926 |