CN102732735A - Method for recovering valued metals from melting slag of copper anode slime Kaldo furnace - Google Patents
Method for recovering valued metals from melting slag of copper anode slime Kaldo furnace Download PDFInfo
- Publication number
- CN102732735A CN102732735A CN2012102545668A CN201210254566A CN102732735A CN 102732735 A CN102732735 A CN 102732735A CN 2012102545668 A CN2012102545668 A CN 2012102545668A CN 201210254566 A CN201210254566 A CN 201210254566A CN 102732735 A CN102732735 A CN 102732735A
- Authority
- CN
- China
- Prior art keywords
- reaction
- filter residue
- gold
- silver
- filtrating
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Granted
Links
Classifications
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Abstract
The invention discloses a method for recovering valued metals from melting slag of a copper anode slime Kaldo furnace. The method comprises the following steps of: leaching and preliminarily removing silicon from melting slag of the copper anode slime Kaldo furnace with hydrochloric acid, and then removing acid soluble base metals such as copper and arsenic by utilizing a mixed solution of hydrochloric acid and sodium chloride to obtain a second filter residue; removing lead from the second filter residue by utilizing an acetic acid system to obtain a third filter residue; adding the third filter residue to a carbonate solution to convert barium in the third filter residue into barium carbonate, and reacting the barium carbonate with hydrochloric acid to obtain gold and silver enriched slag; and respectively extracting gold and silver from the gold and silver enriched slag. Compared with the prior art, a mode of removing the base metals such as the silicium, the lead and the barium step by step is adopted by the method, the gold and the silver are separated from the base metals and effectively enriched, and then the gold and the silver are further recovered, so the method has the advantages of high removal rates of the base materials, high enrichment degrees of the gold and the silver and high recovery rates of the gold and the silver. Proved by experiment results, the recovery rates of the gold and the silver are both higher than 98%.
Description
Technical field
The present invention relates to the metal recovery technical field, more particularly, relate to a kind of method that from many stoves of copper anode mud Ka Er smelting slag, reclaims valuable metal.
Background technology
Precious metal has unique physics and chemical property, is widely used in every field such as military project, oil, chemical industry, organic synthesis, microelectronics and tail gas clean-up, plays irreplaceable effect in modern times in the high-tech industry.Copper anode mud is the important source of extracting precious metal.Copper anode mud pyrogenic attack technology has that treatment capacity is big, the simple advantage of operation of equipment, how to be adopted by the big copper smelting enterprise of industrial scale.The many stoves of Ka Er are handled technology and the environment-friendly advantage of copper anode mud thermal process with himself, progressively obtain promotion and application.Many stove fires of Ka Er method is handled the smelting slag (being called for short many stoves of copper anode mud Ka Er smelting slag) that copper anode mud produces; Mainly consist of sodium, lead, silicon, barium; Contain valuable metals such as a certain amount of gold and silver bismuth; Its concrete composition is: Si 3 ~ 15wt%, Na 5 ~ 10wt%, Cu 0.1 ~ 2wt%, As 0.5 ~ 5wt%, Ba 8 ~ 20wt%, Pb 20 ~ 40wt%, Bi 1 ~ 5wt%, Au 50 ~ 300g/t, Ag 0.3 ~ 1%.Therefore, many stoves of copper anode mud Ka Er smelting slag is the good raw material of valuable metals such as comprehensive recovery of gold, silver, bismuth, thereby the regenerated using of many stoves of copper anode mud Ka Er smelting slag has great importance for the famine that solves China's precious metal resource.
The mode of both at home and abroad many stoves of Ka Er smelting slag being handled is at present smelted the host system recovery for returning copper; Copper smelting enterprise for there not being plumbous smelting system returns the many slags of Ka Er to the copper smelting system, and is more harm than good; Because can only reclaiming gold and silver, the copper smelting system can't reclaim lead, bismuth; Return the copper smelting system simultaneously and can cause lead in big system, to form vicious cycle, influence copper smelting system and cupric electrolysis system, and have a strong impact on the quality of anode copper and cathode copper.
Treatment process to noble metal smelting slag mainly comprises fire reduction melting technology and the direct process for extracting gold and silver of wet method at present.Wherein,, thereby adopt the fire reduction melting technology will improve melt temperature, thereby the high energy consumption of bringing, refractory materials work-ing life is short, slag is clamminess disadvantageous effects such as the mechanical folder of gold and silver because the many slags of Ka Er contain dystectic silicon and permanent white.And utilize wet method directly to extract many stoves of art breading Ka Er smelting slag of gold and silver; Mainly have following problem: because gold and silver content is relatively low in many stoves of copper anode mud Ka Er smelting slag, and the slag mesosilicic acid is plumbous, barium sulfate content is high, and is big to gold and silver parcel degree; Leach medicament and be difficult to the comprehensive engagement gold and silver; Thereby cause the gold and silver leaching yield low, recovery of gold and silver is lower, less than 80%.The inventor considers, a kind of method that from many stoves of copper anode mud Ka Er smelting slag, reclaims valuable metal is provided, valuable metals such as comprehensive recovery of gold and silver, and the recovery of gold and silver is high.
Summary of the invention
In view of this, the technical problem that the present invention will solve is to provide a kind of method that from many stoves of copper anode mud Ka Er smelting slag, reclaims valuable metal, comprehensive recovery of gold and silver, and the recovery of gold and silver is high.
In order to solve above technical problem, the present invention provides a kind of method that from many stoves of copper anode mud Ka Er smelting slag, reclaims valuable metal, may further comprise the steps:
Step a) is carried out hydrochloric acid with many stoves of copper anode mud Ka Er smelting slag and is leached processing, obtains first filtrating and first filter residue after the reaction, and said many stoves of copper anode mud Ka Er smelting slag comprises Si, Na, Ba, Pb, Bi, Au, Cu, As and Ag;
Step b) utilizes the mixing solutions of hydrochloric acid and sodium-chlor that said first filter residue is leached processing, obtains second filtrating and second filter residue after the reaction;
Step c) is mixed said second filter residue with acetate system, obtain the 3rd filtrating and the 3rd filter residue after the reaction;
Step d) adds said the 3rd filter residue in carbonate solution, obtain the 4th filtrating and the 4th filter residue after the reaction, with said the 4th filter residue and mixed in hydrochloric acid, obtains gold and silver enrichment slag and the 5th filtrating after the reaction;
Step e) is extracted Jin Heyin respectively from said gold and silver enrichment slag.
Preferably, said many stoves of copper anode mud Ka Er smelting slag part component content is:
Si?3~15wt%；
Na?5~10wt%；
Cu?0.1~2wt%；
As?0.5~5wt%；
Ba?8~20wt%；
Pb?20~40wt%；
Bi?1~5wt%；
Au?0.05~0.3wt%；
Ag?0.3~1wt%。
Preferably, said step a) is specially:
Utilize hydrochloric acid that many stoves of copper anode mud Ka Er smelting slag is leached processing, the reaction after-filtration obtains first filtrating and first filter residue, and said many stoves of copper anode mud Ka Er smelting slag comprises Si, Na, Ba, Pb, Bi, Au and Ag; The concentration of said hydrochloric acid is 2 ~ 5mol/L, and liquid-solid ratio is 4 ~ 10:1, and the temperature of reaction is 20 ℃ ~ 70 ℃, and the time of reaction is 2 ~ 5h.
Preferably, also comprise:
Regulating said first pH value of filtrate is 2, obtains the bismuth concentrate.
Preferably, said step b) is specially:
Utilize the mixing solutions of hydrochloric acid and sodium-chlor that said first filter residue is leached processing; Obtain second filtrating and second filter residue after the reaction; The concentration of said hydrochloric acid in mixing solutions is 2 ~ 4mol/L, and the concentration of sodium-chlor in mixing solutions is 0.5 ~ 2mol/L, and liquid-solid ratio is 4 ~ 10:1; The temperature of reaction is 40 ℃ ~ 70 ℃, and the time of reaction is 1 ~ 4h.
Preferably, said step c) is specially:
Said second filter residue is mixed with acetate system; Said acetate system is the mixing solutions of acetate and acetate concentration; Obtain the 3rd filtrating and the 3rd filter residue after the reaction, said acetate and the acetic acid concentration in mixing solutions is respectively 3 ~ 5mol/L, and liquid-solid ratio is 6 ~ 10:1; The temperature of reaction is 60 ℃ ~ 90 ℃, and the time of reaction is 1 ~ 5h.
Preferably, also comprise:
In said the 3rd filtrating, add sulfuric acid, obtain lead sulfate after the reaction.
Preferably, said step d) is specially:
In sodium carbonate solution, add said the 3rd filter residue, obtain the 4th filtrating and the 4th filter residue after the reaction, the concentration of said sodium carbonate solution is 3 ~ 4mol/L, and liquid-solid ratio is 10 ~ 14:1, and the temperature of reaction is 80 ℃ ~ 100 ℃, and the time of reaction is 2 ~ 5h;
With said the 4th filter residue and mixed in hydrochloric acid, obtain gold and silver enrichment slag and the 5th filtrating after the reaction.
Preferably, also comprise:
In said the 5th filtrating, add sulfuric acid, obtain permanent white and hydrochloric acid after the reaction.
Preferably, said step e) is specially:
Said gold and silver enrichment slag is carried out the VAL-DROP chlorination leach processing, obtain the 6th filtrating and the 6th filter residue after the reaction;
Said the 6th filtrating is mixed with oxalic acid, obtain bronze after the reduction reaction;
Said the 6th filter residue is carried out S-WAT leach processing, obtain the 7th filtrating, obtain silver powder after the reduction.
The invention provides a kind of method that from many stoves of copper anode mud Ka Er smelting slag, reclaims valuable metal; Many stoves of copper anode mud Ka Er melting hydrochloric acid for dreg is leached preparatory desiliconization; Utilize the mixing solutions of hydrochloric acid and sodium-chlor to remove easy molten sour base metal such as copper, arsenic again, obtain second filter residue; Utilize acetate system to take off lead second filter residue, obtain the 3rd filter residue; In carbonate solution, add said the 3rd filter residue, the barium in the 3rd filter residue is converted into barium carbonate, and obtain gold and silver enrichment slag behind the hydrochloric acid reaction; From said gold and silver enrichment slag, extract Jin Heyin respectively.Compared with prior art, the present invention adopts substep to remove base metal modes such as silicon, lead, barium, and gold and silver is separated with base metal; Gold and silver has obtained effective enrichment, further reclaims gold and silver again, therefore; Base metal clearance of the present invention is high, and the gold and silver enrichment degree is high, and the recovery of gold and silver is high.Experimental result shows that the gold and silver recovery of method provided by the invention is more than 98%.
Description of drawings
Fig. 1 is the embodiment of the invention 1 disclosed schema that from many stoves of copper anode mud Ka Er smelting slag, reclaims valuable metal.
Embodiment
Carry out clear, intactly description in the face of the technical scheme in the embodiment of the invention down, obviously, described embodiment only is the present invention's part embodiment, rather than whole embodiment.Based on the embodiment among the present invention, those of ordinary skills are not making the every other embodiment that is obtained under the creative work prerequisite, all belong to the scope of the present invention's protection.
The present invention provides a kind of method that from many stoves of copper anode mud Ka Er smelting slag, reclaims valuable metal, may further comprise the steps:
Step a) is carried out hydrochloric acid with many stoves of copper anode mud Ka Er smelting slag and is leached processing, obtains first filtrating and first filter residue after the reaction, and said many stoves of copper anode mud Ka Er smelting slag comprises Si, Na, Ba, Pb, Bi, Au, Cu, As and Ag;
Step b) utilizes the mixing solutions of hydrochloric acid and sodium-chlor that said first filter residue is leached processing, obtains second filtrating and second filter residue after the reaction;
Step c) is mixed said second filter residue with acetate system, obtain the 3rd filtrating and the 3rd filter residue after the reaction;
Step d) adds said the 3rd filter residue in carbonate solution, obtain the 4th filtrating and the 4th filter residue after the reaction, with said the 4th filter residue and mixed in hydrochloric acid, obtains gold and silver enrichment slag and the 5th filtrating after the reaction;
Step e) is extracted Jin Heyin respectively from said gold and silver enrichment slag.
In above-mentioned preparation process, the present invention adopts substep to remove base metal modes such as silicon, lead, barium, and gold and silver is separated with base metal, and gold and silver has obtained effective enrichment, further reclaims gold and silver again.Said many slags of Ka Er principal constituent thing phase: sodium mainly exists with water glass, sodium arseniate or sodium antimonate form, and silicon exists with the silicate form, and barium exists with the permanent white form, and lead mainly exists with lead silicate and plumbous oxide form.
Many stoves of copper anode mud Ka Er smelting slag is leached before the processing, preferably it is carried out break process, more preferably be crushed to more than 100 orders.Said many stoves of copper anode mud Ka Er smelting slag part component content that the present invention adopts is: Si 3 ~ 15wt%; Na 5 ~ 10wt%; Cu 0.1 ~ 2wt%; As 0.5 ~ 5wt%; Ba 8 ~ 20wt%; Pb 20 ~ 40wt%; Bi 1 ~ 5wt%; Au0.05 ~ 0.3wt%; Ag 0.3 ~ 1wt%.Preferably, Si 3 ~ 10wt%; Na 6 ~ 8wt%; Cu 0.1 ~ 1wt%; As 0.5 ~ 3wt%; Ba 8 ~ 15wt%; Pb 20 ~ 40wt%; Bi 1 ~ 4wt%; Au 0.05 ~ 0.2wt%; Ag 0.3 ~ 0.8wt%.
In above-mentioned preparation process; Many stoves of copper anode mud Ka Er melting hydrochloric acid for dreg is leached preparatory desiliconization; Said step a) is specially: utilize hydrochloric acid that many stoves of copper anode mud Ka Er smelting slag is leached processing; The reaction after-filtration obtains first filtrating and first filter residue, and said many stoves of copper anode mud Ka Er smelting slag comprises Si, Na, Ba, Pb, Bi, Au and Ag; The concentration of said hydrochloric acid is 2 ~ 5mol/L, and liquid-solid ratio is 4 ~ 10:1, and the temperature of reaction is 20 ℃ ~ 70 ℃, and the time of reaction is 2 ~ 5h.Wherein, above-mentioned liquid-solid ratio is liquid-solid mass ratio, i.e. mass ratio between leach liquor and the solid materials.As preferred version, also comprise: regulating said first pH value of filtrate is 2, obtains the bismuth concentrate.In the above-mentioned desiliconization process, the reaction of most water glass and minor amount of silicon lead plumbate generates lyotrope silicic acid and chlorate entering liquid phase in hydrochloric acid and many stoves of the copper anode mud Ka Er smelting slag, and most of MOX is got into liquid phase by dissolving with hydrochloric acid simultaneously.
The desilication reaction formula:
2HCl+Na
2SiO
3=2NaCl+H
2SiO
3
HCl+PbSiO
3=PbCl
2+H
2SiO
3
After step a) is accomplished, utilize the mixing solutions of hydrochloric acid and sodium-chlor to remove easy molten sour base metal such as copper, arsenic.Said step b) is specially: utilize the mixing solutions of hydrochloric acid and sodium-chlor that said filter residue is leached processing; Obtain second filtrating and second filter residue after the reaction; The concentration of said hydrochloric acid in mixing solutions is 2 ~ 4mol/L, and the concentration of sodium-chlor in mixing solutions is 0.5 ~ 2mol/L, and liquid-solid ratio is 4 ~ 10:1; The temperature of reaction is 40 ℃ ~ 70 ℃, and the time of reaction is 1 ~ 4h.Wherein, above-mentioned liquid-solid ratio is liquid-solid mass ratio, i.e. mass ratio between leach liquor and the solid materials.In this step, the mixing solutions of hydrochloric acid and sodium-chlor is leached to MOX remaining in said first filter residue such as bismuth oxide, cupric oxide, arsenate etc. in second filtrating.
Then second filter residue is taken off lead with acetate system, obtain the 3rd filter residue and promptly take off lead skim.Said step c) is specially: said second filter residue is mixed with acetate system; Said acetate system is the mixing solutions of acetate and acetate concentration; Obtain the 3rd filtrating and the 3rd filter residue after the reaction, said acetate and the acetic acid concentration in mixing solutions is respectively 3 ~ 5mol/L, and liquid-solid ratio is 6 ~ 10:1; The temperature of reaction is 60 ℃ ~ 90 ℃, and the time of reaction is 1 ~ 5h.Wherein, above-mentioned liquid-solid ratio is liquid-solid mass ratio, i.e. mass ratio between leach liquor and the solid materials.Said acetate is preferably one or more in sodium-acetate, Potassium ethanoate and the ammonium acetate, more preferably sodium-acetate.
Said second filter residue comprises a certain amount of lead silicate, and this step utilizes the lead silicate of low solubility and acetate system reaction to generate the lyotrope plumbic acetate, is liquid phase thereby get into the 3rd filtrating, and reaction formula is following:
2HAc+2NaAc+2PbSiO
3=2Pb(Ac)
2+Na
2SiO
3+H
2SiO
3
2HAc+PbO=Pb(Ac)
2+H
2O
As preferred version, also comprise: in said the 3rd filtrating, add sulfuric acid, obtain lead sulfate after the reaction.
After step c) is accomplished, in carbonate solution, add the 3rd filter residue that said step c) obtains, the barium in the 3rd filter residue is converted into barium carbonate, and obtain gold and silver enrichment slag behind the hydrochloric acid reaction.Said carbonate is preferably one or more in yellow soda ash, salt of wormwood and the volatile salt, more preferably yellow soda ash.
Said step d) is specially: in sodium carbonate solution, add said the 3rd filter residue; Obtain the 4th filtrating and the 4th filter residue after the reaction; Said sodium carbonate solution is a saturated sodium carbonate solution, and the concentration of sodium carbonate solution is 3 ~ 4mol/L, and liquid-solid ratio is 10 ~ 14:1; The temperature of reaction is 80 ℃ ~ 100 ℃, and the time of reaction is 2 ~ 5h; With said the 4th filter residue and mixed in hydrochloric acid, obtain gold and silver enrichment slag and the 5th filtrating after the reaction.Wherein, above-mentioned liquid-solid ratio is liquid-solid mass ratio, i.e. mass ratio between leach liquor and the solid materials.In step d), permanent white principle transition utilizes the high concentration carbon silicate system under hot conditions, permanent white to be converted into barium carbonate, uses the dissolving with hydrochloric acid barium carbonate again, reaches the purpose except that barium.The main reaction formula:
Na
2CO
3+BaSO
4=Na
2SO
4+BaCO
3
2HCl+BaCO
3=BaCl
2+H
2CO
3
As preferred version, also comprise: in said the 5th filtrating, add sulfuric acid, obtain permanent white and hydrochloric acid after the reaction, the hydrochloric acid reusable edible.
At last, from said gold and silver enrichment slag, extract Jin Heyin respectively.Said step e) is specially: said gold and silver enrichment slag is carried out the VAL-DROP chlorination leach processing, obtain the 6th filtrating and the 6th filter residue after the reaction; Said the 6th filtrating is mixed with oxalic acid, obtain bronze after the reduction reaction; Said the 6th filter residue is carried out S-WAT leach processing, obtain the 7th filtrating, obtain silver powder after the reduction.The said reduction that obtains silver powder is handled, preferably with Hydrazine Hydrate 80 as going back original reagent.
Can find out that from above scheme the method that from many stoves of copper anode mud Ka Er smelting slag, reclaims valuable metal provided by the invention has following characteristics:
(1), the base metal clearance is high, the gold and silver enrichment is high, recovery of gold and silver is high.
The present invention adopts indirect removal of impurities mode, after processing is leached in conversion through substep with many stoves of copper anode mud Ka Er smelting slag principal constituent genus lead, barium, silicon, sodium etc. earlier, but slag loss of weight 90%-95%, but the gold and silver enrichment is extracted gold and silver to gold and silver enrichment slag more than 10 times again.Separation and concentration process gold and silver rate of loss is extremely low, and the integrated artistic recovery of gold and silver is more than 98%.
(2), utilize full wet processing to remove the HMP material.
According to HMP material silicon-dioxide in many stoves of copper anode mud Ka Er smelting slag and permanent white thing characteristics mutually, specific aim is carried out chemical conversion and is isolated silicon and barium, has solved the difficult problem that thermal process is difficult to handle HMP material permanent white.
(3), lead, bismuth, barium reclaim comprehensive utilization of resources rate height.
The present invention is carrying out the isolating while to base metal and gold and silver, and valuable element lead, bismuth, barium are reclaimed, and has accomplished comprehensive utilization of resources conscientiously.
(4), environment protecting is good, good operational environment.
The present invention is pure hydrometallurgical processes, does not have waste gas to produce, and the waste residue generation is few, and most of process waste liquor can utilize in other operations, and water use efficiency is high, and environment protecting is good.And method provided by the invention does not have lead steam to pollute good operational environment.
In order to further specify technical scheme of the present invention; Below in conjunction with embodiment the preferred embodiment of the invention is described; Describe just to further specifying feature and advantage of the present invention but should be appreciated that these, rather than to the restriction of claim of the present invention.
Embodiment 1
As shown in Figure 1, be the schema that from many stoves of copper anode mud Ka Er smelting slag, reclaims valuable metal provided by the invention.
(1), many raw slag of Ka Er composition: Si 3.7%, Na 6.2%, Cu 0.7%, As2.1%, Ba 13.9%, Pb 35.9%, Bi 1.2%, precious metal Au 0.0208%, Ag 0.48%.Many stoves of Ka Er smelting slag is broken and be finely ground to granularity 100 orders.
(2), desiliconization
The many slags of 2000g Ka Er are leached desiliconization with hydrochloric acid, the desilication process condition: concentration of hydrochloric acid 3mol/L, liquid-solid mass ratio 6:1,40 ℃ of temperature of reaction, reaction times 3h.After reaction finishes; Filter, the heavy 1718.4g of filter residue, desiliconization process slag rate of weight loss is 14.08%; Filter residue principal constituent: Si 1.61%, Na0.89%, Cu0.34%, As0.737%, Ba 16.28%, Pb 40.95%, Bi 0.2%, precious metal Au0.0242%, Ag 0.558%.Filtrating adds alkali and transfers to pH=2, obtains the bismuth concentrate, and bi content is 9% in the bismuth concentrate.This process silicon removes major part, and major parts such as sodium arsenic bismuth copper are removed, and the precious metal gold and silver is loss not.
(3), reinforcement removes base metal
Leach the desiliconization slag reinforcement with hydrochloric acid and sodium chloride solution and remove base metal, strengthen removing the base metal processing condition: concentration of hydrochloric acid 2.5mol/L, sodium-chlor 1mol/L, liquid-solid mass ratio 7:1,50 ℃ of temperature of reaction, reaction times 2h.After reaction finishes; Filter, the heavy 1574.2g of filter residue removes base metal process slag rate of weight loss 8.38%; Filter residue principal constituent: Si 1.12%, Na0.22%, Cu0.033%, As0.002%, Ba 17.45%, Pb 44.34%, Bi 0.01%, precious metal Au 0.0264%, Ag 0.603%.Leach through strengthening, basic removal such as base metal sodium, arsenic, copper is cleaner, and gold is loss not, and silver loses about 1%.
(4), take off lead
Remove the base metal slag and get in the acetate system and take off lead, take off splicer's condition of planting: sodium-acetate is identical with acetate concentration all to be 5mol/L, liquid-solid mass ratio 8:1,80 ℃ of temperature of reaction, reaction times 3h.After reaction finishes, filter, take off the heavy 460.3g of lead skim, take off plumbous process slag rate of weight loss 70.73%, take off plumbous filter residue principal constituent: Si 2.17%, Pb 19.8%, Ba 59.74%, precious metal Au 0.0904%, Ag 2.04%.Take off plumbous filtrating and add sulfuric acid and obtain lead sulfate, acetic acid is able to reprocessing cycle and uses.
(5), take off barium
Taking off lead skim gets in the saturated sodium carbonate solution permanent white in the slag is barium carbonate, transformation process condition: concentration of sodium carbonate 3.5mol/L, liquid-solid mass ratio 12:1,90 ℃ of temperature, reaction times 3h transition.After-filtration makes the transition; Transition, the back mother liquor can be used as the required alkali lye use of bismuth concentrate, and transition, filter residue was used dissolving with hydrochloric acid, and filter residue is a gold and silver enrichment slag; The heavy 127.9g of slag; Take off barium process slag rate of weight loss 72.2%, take off gold and silver enrichment slag principal constituent behind the barium: Si 0.19%, Pb 2.3%, Ba 12.7%, precious metal Au 0.325%, Ag 7.35%.Filtrating adds sulfuric acid, obtains permanent white and hydrochloric acid, the hydrochloric acid reusable edible.
(6), gold and silver reclaims
Adopt the chlorination of strong oxidizer VAL-DROP to leach gold and silver enrichment slag, gold obtains bronze with the oxalic acid reduction in leach liquor;
Silver leaches through S-WAT earlier and obtains silver powder with hydrazine hydrate reduction again in slag.Finally obtaining bronze is 0.416g, silver powder 9.4g, gold recovery 98.5%, silver raising recovery rate 98%.Bronze, silver powder purity are very high, free from foreign meter basically.
Embodiment 2
(1), many raw slag of Ka Er composition: Si 4.1%, Na 7.6%, Cu 1.5%, As4.6%, Ba 14.8%, Pb 38.9%, Bi 2.6%, precious metal Au 0.015%, Ag 0.68%.Many stoves of Ka Er smelting slag is broken and be finely ground to granularity 150 orders.
(2), desiliconization
The many slags of 2000g Ka Er are leached desiliconization with hydrochloric acid, the desilication process condition: concentration of hydrochloric acid 4mol/L, liquid-solid mass ratio 8:1,60 ℃ of temperature of reaction, reaction times 2h.After reaction finishes; Filter, the heavy 1688.4g of filter residue, desiliconization process slag rate of weight loss is 15.58%; Filter residue principal constituent: Si 1.92%, Na0.68%, Cu0.43%, As 1.84%, Ba 17.43%, Pb 45.1%, Bi 0.5%, precious metal Au0.0177%, Ag 0.805%.Filtrating adds alkali and transfers to pH=2, obtains the bismuth concentrate, and bi content is 18.4% in the bismuth concentrate.This process silicon removes major part, and major parts such as sodium arsenic bismuth copper are removed, and the precious metal gold and silver is loss not.
(3), reinforcement removes base metal
Leach the desiliconization slag reinforcement with hydrochloric acid and sodium chloride solution and remove base metal, strengthen removing the base metal processing condition: concentration of hydrochloric acid 2.5mol/L, sodium-chlor 2mol/L, liquid-solid mass ratio 8:1,60 ℃ of temperature of reaction, reaction times 3h.After reaction finishes; Filter, the heavy 1541.4g of filter residue takes off base metal process slag rate of weight loss 8.72%; Filter residue principal constituent: Si 1.09%, Na0.17%, Cu0.001%, As0.005%, Ba 19.12%, Pb 48.7%, Bi 0.02%, precious metal Au0.0195%, Ag 0.878%.Leach through strengthening, basic removal such as base metal sodium, arsenic, copper is cleaner, and gold is loss not, and silver loses about 0.5%.
(4), take off lead
Remove the base metal slag and get in the acetate system and take off lead, take off splicer's condition of planting: acetate is all 4mol/L mutually with acetate concentration, liquid-solid mass ratio 10:1,70 ℃ of temperature of reaction, reaction times 4h.After reaction finishes, filter, take off the heavy 467.5g of lead skim, take off plumbous process slag rate of weight loss 69.7%, take off plumbous filter residue principal constituent: Si 1.39%, Pb 21.2%, Ba 33.45%, precious metal Au 0.0642%, Ag 2.897%.Take off plumbous filtrating and add sulfuric acid and obtain lead sulfate, acetic acid is able to reprocessing cycle and uses.
(5), take off barium
Taking off lead skim gets in the saturated sodium carbonate solution permanent white in the slag is barium carbonate, transformation process condition: concentration of sodium carbonate 3.8mol/L, liquid-solid mass ratio 10:1,95 ℃ of temperature, reaction times 4h transition.After-filtration makes the transition; Transition, the back mother liquor can be used as the required alkali lye use of bismuth concentrate, and transition, filter residue was used dissolving with hydrochloric acid, and filter residue is a gold and silver enrichment slag; The heavy 110g of slag; Take off barium process slag rate of weight loss 76.4%, take off gold and silver enrichment slag principal constituent behind the barium: Si 0.17%, Pb 3.2%, Ba 13.9%, precious metal Au 0.273%, Ag 12.29%.Filtrating adds sulfuric acid, obtains permanent white and hydrochloric acid, the hydrochloric acid reusable edible.
(6), gold and silver reclaims
Adopt the chlorination of strong oxidizer VAL-DROP to leach gold and silver enrichment slag, gold obtains bronze with the oxalic acid reduction in leach liquor;
Silver leaches through S-WAT earlier and obtains silver powder with hydrazine hydrate reduction again in slag.Bronze, silver powder purity are 99.99%.Finally obtaining bronze is 0.295g, silver powder 13.34g, whole process gold recovery 98%, silver raising recovery rate 98%.
Embodiment 3
(1), many raw slag of Ka Er composition: Si 6.8%, Na 8.2%, Cu 1.2%, As 1.7%, Ba 12.5%, Pb 35.4%, Bi 1.6%, precious metal Au 0.025%, Ag 0.89%.Many stoves of Ka Er smelting slag is broken and be finely ground to granularity 180 orders.
(2), desiliconization
The many slags of 2000g Ka Er are leached desiliconization with hydrochloric acid, the desilication process condition: concentration of hydrochloric acid 3mol/L, liquid-solid mass ratio 7:1,70 ℃ of temperature of reaction, reaction times 2h.After reaction finishes; Filter, the heavy 1675.3g of filter residue, desiliconization process slag rate of weight loss is 16.25%; Filter residue principal constituent: Si 2.13%, Na 1.25%, Cu 0.45%, As 0.86%, Ba 14.93%, Pb 41.3%, Bi 0.6%, precious metal Au 0.0298%, 1.06%.Filtrating adds alkali and transfers to pH=2, obtains the bismuth concentrate, and bi content is 18.4% in the bismuth concentrate.This process silicon removes major part, and major parts such as sodium arsenic bismuth copper are removed, and the precious metal gold and silver is loss not.
(3), reinforcement removes base metal
Leach the desiliconization slag reinforcement with hydrochloric acid and sodium chloride solution and remove base metal, strengthen removing the base metal processing condition: concentration of hydrochloric acid 2.5mol/L, sodium-chlor 2mol/L, liquid-solid mass ratio 8:1,60 ℃ of temperature of reaction, reaction times 2h.After reaction finishes; Filter, the heavy 1560.3g of filter residue takes off base metal process slag rate of weight loss 6.85%; Filter residue principal constituent: Si 1.09%, Na0.13%, Cu0.001%, As0.005%, Ba 16%, Pb 43.2%, Bi 0.02%, precious metal Au 0.032%, Ag 1.135%.Leach through strengthening, basic removal such as base metal sodium, arsenic, copper is cleaner, and gold is loss not, and silver loses about 0.5%.
(4), take off lead
Remove the base metal slag and get in the acetate system and take off lead, take off splicer's condition of planting: acetate is all 4mol/L mutually with acetate concentration, liquid-solid mass ratio 12:1,80 ℃ of temperature of reaction, reaction times 4h.After reaction finishes, filter, take off the heavy 535g of lead skim, take off plumbous process slag rate of weight loss 65.7%, take off plumbous filter residue principal constituent: Si1.2%, Pb 12.2%, Ba 46%, precious metal Au 0.0934%, Ag 3.32%.Take off plumbous filtrating and add sulfuric acid and obtain lead sulfate, acetic acid is able to reprocessing cycle and uses.
(5), take off barium
Taking off lead skim gets in the saturated sodium carbonate solution permanent white in the slag is barium carbonate, transformation process condition: concentration of sodium carbonate 4mol/L, liquid-solid mass ratio 11:1,90 ℃ of temperature, reaction times 4h transition.After-filtration makes the transition; Transition, the back mother liquor can be used as the required alkali lye use of bismuth concentrate, and transition, filter residue was used dissolving with hydrochloric acid, and filter residue is a gold and silver enrichment slag; The heavy 104.3g of slag; Take off barium process slag rate of weight loss 80.5%, take off gold and silver enrichment slag principal constituent behind the barium: Si 0.5%, Pb 1.2%, Ba 10.7%, precious metal Au 0.47%, Ag 17.01%.Filtrating adds sulfuric acid, obtains permanent white and hydrochloric acid, the hydrochloric acid reusable edible.
(6), gold and silver reclaims
Adopt the chlorination of strong oxidizer VAL-DROP to leach gold and silver enrichment slag, gold available oxalic acid reduction in leach liquor obtains bronze;
Silver leaches through S-WAT earlier and obtains silver powder with hydrazine hydrate reduction again in slag.Bronze, silver powder purity are 99.99%.Finally obtaining bronze is 0.295g, silver powder 13.34g, whole process gold recovery 98%, silver raising recovery rate 98%.
To the above-mentioned explanation of the disclosed embodiments, make this area professional and technical personnel can realize or use the present invention.Multiple modification to these embodiment will be conspicuous concerning those skilled in the art, and defined General Principle can realize under the situation that does not break away from the spirit or scope of the present invention in other embodiments among this paper.Therefore, the present invention will can not be restricted to these embodiment shown in this paper, but will meet and principle disclosed herein and features of novelty the wideest corresponding to scope.
Claims (10)
1. method that from many stoves of copper anode mud Ka Er smelting slag, reclaims valuable metal may further comprise the steps:
Step a) is carried out hydrochloric acid with many stoves of copper anode mud Ka Er smelting slag and is leached processing, obtains first filtrating and first filter residue after the reaction, and said many stoves of copper anode mud Ka Er smelting slag comprises Si, Na, Ba, Pb, Bi, Au, Cu, As and Ag;
Step b) utilizes the mixing solutions of hydrochloric acid and sodium-chlor that said first filter residue is leached processing, obtains second filtrating and second filter residue after the reaction;
Step c) is mixed said second filter residue with acetate system, obtain the 3rd filtrating and the 3rd filter residue after the reaction;
Step d) adds said the 3rd filter residue in carbonate solution, obtain the 4th filtrating and the 4th filter residue after the reaction, with said the 4th filter residue and mixed in hydrochloric acid, obtains gold and silver enrichment slag and the 5th filtrating after the reaction;
Step e) is extracted Jin Heyin respectively from said gold and silver enrichment slag.
2. method according to claim 1 is characterized in that, said many stoves of copper anode mud Ka Er smelting slag part component content is:
Si?3~15wt%；
Na?5~10wt%；
Cu?0.1~2wt%；
As?0.5~5wt%；
Ba?8~20wt%；
Pb?20~40wt%；
Bi?1~5wt%；
Au?0.05~0.3wt%；
Ag?0.3~1wt%。
3. method according to claim 1 is characterized in that, said step a) is specially:
Utilize hydrochloric acid that many stoves of copper anode mud Ka Er smelting slag is leached processing, the reaction after-filtration obtains first filtrating and first filter residue, and said many stoves of copper anode mud Ka Er smelting slag comprises Si, Na, Ba, Pb, Bi, Au and Ag; The concentration of said hydrochloric acid is 2 ~ 5mol/L, and liquid-solid ratio is 4 ~ 10:1, and the temperature of reaction is 20 ℃ ~ 70 ℃, and the time of reaction is 2 ~ 5h.
4. method according to claim 1 is characterized in that, also comprises:
Regulating said first pH value of filtrate is 2, obtains the bismuth concentrate.
5. method according to claim 1 is characterized in that, said step b) is specially:
Utilize the mixing solutions of hydrochloric acid and sodium-chlor that said first filter residue is leached processing; Obtain second filtrating and second filter residue after the reaction; The concentration of said hydrochloric acid in mixing solutions is 2 ~ 4mol/L, and the concentration of sodium-chlor in mixing solutions is 0.5 ~ 2mol/L, and liquid-solid ratio is 4 ~ 10:1; The temperature of reaction is 40 ℃ ~ 70 ℃, and the time of reaction is 1 ~ 4h.
6. method according to claim 1 is characterized in that, said step c) is specially:
Said second filter residue is mixed with acetate system; Said acetate system is the mixing solutions of acetate and acetate concentration; Obtain the 3rd filtrating and the 3rd filter residue after the reaction, said acetate and the acetic acid concentration in mixing solutions is respectively 3 ~ 5mol/L, and liquid-solid ratio is 6 ~ 10:1; The temperature of reaction is 60 ℃ ~ 90 ℃, and the time of reaction is 1 ~ 5h.
7. method according to claim 1 is characterized in that, also comprises:
In said the 3rd filtrating, add sulfuric acid, obtain lead sulfate after the reaction.
8. method according to claim 1 is characterized in that, said step d) is specially:
In sodium carbonate solution, add said the 3rd filter residue, obtain the 4th filtrating and the 4th filter residue after the reaction, the concentration of said sodium carbonate solution is 3 ~ 4mol/L, and liquid-solid ratio is 10 ~ 14:1, and the temperature of reaction is 80 ℃ ~ 100 ℃, and the time of reaction is 2 ~ 5h;
With said the 4th filter residue and mixed in hydrochloric acid, obtain gold and silver enrichment slag and the 5th filtrating after the reaction.
9. method according to claim 1 is characterized in that, also comprises:
In said the 5th filtrating, add sulfuric acid, obtain permanent white and hydrochloric acid after the reaction.
10. method according to claim 1 is characterized in that, said step e) is specially:
Said gold and silver enrichment slag is carried out the VAL-DROP chlorination leach processing, obtain the 6th filtrating and the 6th filter residue after the reaction;
Said the 6th filtrating is mixed with oxalic acid, obtain bronze after the reduction reaction;
Said the 6th filter residue is carried out S-WAT leach processing, obtain the 7th filtrating, obtain silver powder after the reduction.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN201210254566.8A CN102732735B (en) | 2012-07-20 | 2012-07-20 | Method for recovering valued metals from melting slag of copper anode slime Kaldo furnace |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN201210254566.8A CN102732735B (en) | 2012-07-20 | 2012-07-20 | Method for recovering valued metals from melting slag of copper anode slime Kaldo furnace |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| CN102732735A true CN102732735A (en) | 2012-10-17 |
| CN102732735B CN102732735B (en) | 2014-02-19 |
Family
ID=46989029
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CN201210254566.8A Expired - Fee Related CN102732735B (en) | 2012-07-20 | 2012-07-20 | Method for recovering valued metals from melting slag of copper anode slime Kaldo furnace |
Country Status (1)
| Country | Link |
|---|---|
| CN (1) | CN102732735B (en) |
Cited By (9)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN103752415A (en) * | 2014-01-09 | 2014-04-30 | 铜陵有色金属集团股份有限公司 | Method for recycling precious metal such as gold and silver in Kaldo furnace by using flotation technology |
| CN104878209A (en) * | 2015-06-15 | 2015-09-02 | 中南大学 | Method of leaching gold and silver targets in copper anode mud |
| CN105063361A (en) * | 2015-08-20 | 2015-11-18 | 江西铜业股份有限公司 | Method for comprehensively recovering valuable metal from copper anode slime |
| CN105886782A (en) * | 2016-06-29 | 2016-08-24 | 江西铜业集团公司 | Method for recycling valuable metal in silver residue by using full-wet method |
| CN106086440A (en) * | 2016-08-04 | 2016-11-09 | 西北矿冶研究院 | Method for separating and recovering valuable metals in Kaldo furnace smelting slag by wet method |
| CN106636660A (en) * | 2016-11-05 | 2017-05-10 | 北京工业大学 | Copper anode furnace slag comprehensive utilization method |
| CN106967884A (en) * | 2017-03-20 | 2017-07-21 | 北矿力澜科技咨询（北京）有限公司 | A kind of method of silver separating residue of copper anode slime step by step arithmetic |
| CN108034821A (en) * | 2017-11-14 | 2018-05-15 | 河南中原黄金冶炼厂有限责任公司 | A kind of wet processing process of earth of positive pole smelting slag |
| CN116281884A (en) * | 2023-03-20 | 2023-06-23 | 金川集团铜业有限公司 | Method for removing selenium and bismuth from copper anode slime pressurized leaching residues |
Citations (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101338368A (en) * | 2008-08-08 | 2009-01-07 | 中南大学 | Method for preprocessing anode sludge and recovering dissipated metal |
| CN101434385A (en) * | 2008-12-08 | 2009-05-20 | 阳谷祥光铜业有限公司 | Process for extracting tellurium from copper anode mud |
-
2012
- 2012-07-20 CN CN201210254566.8A patent/CN102732735B/en not_active Expired - Fee Related
Patent Citations (2)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101338368A (en) * | 2008-08-08 | 2009-01-07 | 中南大学 | Method for preprocessing anode sludge and recovering dissipated metal |
| CN101434385A (en) * | 2008-12-08 | 2009-05-20 | 阳谷祥光铜业有限公司 | Process for extracting tellurium from copper anode mud |
Non-Patent Citations (2)
| Title |
|---|
| 夏彬: "铜阳极泥稀贵金属回收工艺及优化", 《铜业工程》, no. 3, 31 March 2011 (2011-03-31), pages 34 - 37 * |
| 涂百乐等: "卡尔多炉处理铜阳极泥技术及应用实践", 《黄金》, vol. 32, no. 3, 31 March 2011 (2011-03-31), pages 45 - 48 * |
Cited By (15)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN103752415A (en) * | 2014-01-09 | 2014-04-30 | 铜陵有色金属集团股份有限公司 | Method for recycling precious metal such as gold and silver in Kaldo furnace by using flotation technology |
| CN104878209A (en) * | 2015-06-15 | 2015-09-02 | 中南大学 | Method of leaching gold and silver targets in copper anode mud |
| CN105063361B (en) * | 2015-08-20 | 2017-03-22 | 江西铜业股份有限公司 | Method for comprehensively recovering valuable metal from copper anode slime |
| CN105063361A (en) * | 2015-08-20 | 2015-11-18 | 江西铜业股份有限公司 | Method for comprehensively recovering valuable metal from copper anode slime |
| CN105886782B (en) * | 2016-06-29 | 2018-01-23 | 江西铜业集团公司 | A kind of method of valuable metal in Whote-wet method recovery silver separating residues |
| CN105886782A (en) * | 2016-06-29 | 2016-08-24 | 江西铜业集团公司 | Method for recycling valuable metal in silver residue by using full-wet method |
| CN106086440A (en) * | 2016-08-04 | 2016-11-09 | 西北矿冶研究院 | Method for separating and recovering valuable metals in Kaldo furnace smelting slag by wet method |
| CN106086440B (en) * | 2016-08-04 | 2017-12-22 | 西北矿冶研究院 | Method for separating and recovering valuable metals in Kaldo furnace smelting slag by wet method |
| CN106636660A (en) * | 2016-11-05 | 2017-05-10 | 北京工业大学 | Copper anode furnace slag comprehensive utilization method |
| CN106636660B (en) * | 2016-11-05 | 2018-03-30 | 北京工业大学 | A kind of method of copper anode furnace comprehensive utilization of slag |
| CN106967884A (en) * | 2017-03-20 | 2017-07-21 | 北矿力澜科技咨询（北京）有限公司 | A kind of method of silver separating residue of copper anode slime step by step arithmetic |
| CN106967884B (en) * | 2017-03-20 | 2019-05-21 | 北矿力澜科技咨询（北京）有限公司 | A kind of method of silver separating residue of copper anode slime step by step arithmetic |
| CN108034821A (en) * | 2017-11-14 | 2018-05-15 | 河南中原黄金冶炼厂有限责任公司 | A kind of wet processing process of earth of positive pole smelting slag |
| CN108034821B (en) * | 2017-11-14 | 2019-08-02 | 河南中原黄金冶炼厂有限责任公司 | A kind of wet processing process of earth of positive pole smelting slag |
| CN116281884A (en) * | 2023-03-20 | 2023-06-23 | 金川集团铜业有限公司 | Method for removing selenium and bismuth from copper anode slime pressurized leaching residues |
Also Published As
| Publication number | Publication date |
|---|---|
| CN102732735B (en) | 2014-02-19 |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| CN102732735B (en) | Method for recovering valued metals from melting slag of copper anode slime Kaldo furnace | |
| CN108118157B (en) | Wiring board burns the recovery method of cigarette ash pretreatment and bromine | |
| CN105734299B (en) | A kind of method of oxygen pressure treatment tin anode mud comprehensively recovering valuable metal | |
| CN106011488B (en) | A kind of method of high arsenic-and copper-bearing cigarette ash comprehensively recovering valuable metal | |
| CN103397213B (en) | Method for decomposing and extracting Baotou rare earth ore through mixed alkali roasting process | |
| CN102286665B (en) | Comprehensive recovery method for complicated materials containing arsenic and valuable metal slag dust | |
| CN102534255B (en) | Wet-fire combined smelting process for antimony or bismuth | |
| CN107338454B (en) | A method of recycling copper and arsenic from white metal | |
| CN105347315A (en) | Method for extracting crude tellurium from tellurium-containing smelting waste residue | |
| CN107502744B (en) | A kind of processing method of high lead barium silver separating residues | |
| CN102502531A (en) | Novel and environment-friendly method for extracting tellurium from tellurium slag | |
| CN102363522A (en) | Technology for extracting selenium from low-grade selenium-containing material | |
| WO2012171481A1 (en) | Hydrometallurgical process for complete and comprehensive recovery with substantially no wastes and zero emissions | |
| CN1948522A (en) | Metho of recovering valuable metal in gold silver smelting furance waste lining brick | |
| CN103589873A (en) | Method for recovering valuable metals from silver-zinc slag | |
| CN104911366A (en) | Method for recovering valuable metals from silver-bismuth slag through using aqua regia | |
| CN108950199A (en) | A method of it being used for the nickel and cobalt solution of synthesis of ternary presoma using the preparation of nickel sulfide cobalt ore | |
| CN102586584A (en) | Method for selectively separating valuable metals from complex lead-containing precious metal material | |
| CN112063850B (en) | Method for recovering valuable metals after alkaline leaching and dehalogenation of circuit board smelting smoke dust | |
| CN110592385B (en) | Harmless recovery method for waste circuit board smelting soot | |
| CN104445105B (en) | Method for enriching and recovering tellurium dioxide from tellurium-containing soda residue | |
| CN109055764B (en) | Comprehensive recovery method of high-chlorine low-zinc material | |
| CN102321824B (en) | Method for desiliconizing cobalt white alloy | |
| CN109252057B (en) | Molten salt chlorination extraction method of low-grade zirconium concentrate | |
| CN112111647B (en) | Method for pre-treating gold leaching by using gold ore calcine or roasting cyanidation tailings |
Legal Events
| Date | Code | Title | Description |
|---|---|---|---|
| C06 | Publication | ||
| PB01 | Publication | ||
| C10 | Entry into substantive examination | ||
| SE01 | Entry into force of request for substantive examination | ||
| C14 | Grant of patent or utility model | ||
| GR01 | Patent grant | ||
| CF01 | Termination of patent right due to non-payment of annual fee |
Granted publication date: 20140219 Termination date: 20210720 |
|
| CF01 | Termination of patent right due to non-payment of annual fee |