JPS5953335B2 - How to recover metals from ore - Google Patents

Description

DETAILED DESCRIPTION OF THE INVENTION The present invention relates to a method for recovering metals from sulfide ores containing sulfides of lead, silver and zinc.

To date, it has been proposed to convert metal sulfides into chlorides in metal recovery methods.

For example, metal sulfide concentrates can be chlorinated with ferric chloride and chlorine gas in an aqueous sodium chloride or calcium chloride solution (U.S. Pat. No. 1.736.659).
(see specification).

It has already been proposed to chlorinate dry concentrates of metal sulfides containing lead, zinc and silver sulfides with chlorine gas (Mining and Scientific Pres.
ss), Volume 112, May 27, 1916, “The Dry Chlorination of Complex Ors” by Ionides.
Dry Chlorination of Compl.
(see ex 0res).

The final chlorination takes place in the baking process in the presence of air, and the ferric chloride formed during the chlorination process is decomposed to produce chlorine to complete the chlorination of the metal sulfides.

This latter known method can be used for the production of zinc chloride, in which sulfur dioxide is produced during the baking process and
It is not a pollution-free method as it is released into the atmosphere.

Furthermore, when the chlorinated product is treated with sodium chloride to dissolve the metal chloride, undesired impurities, especially zinc chloride, are formed in the sodium chloride leachate.

This reduces the ability to dissolve silver and lead sulphides from the chlorinated ore after a period of time.

The present invention provides for (a) chlorinating the concentrate to convert metal sulfides to metal chlorides and converting sulfide sulfur to elemental sulfur; (b)
(a) leaching the non-gaseous products of the process with an aqueous sodium chloride solution to separate and dissolve the lead and silver chlorides therein from the insoluble solids; and (C) cooling the sodium chloride leachate to dissolve the lead and silver chlorides therein. Almost all of the lead chloride inside is precipitated,
Subsequently, precipitated lead chloride is separated from the leachate; (d) silver is recovered from the leachate from which lead chloride has been removed obtained in step (C); and (e) a portion of the solution resulting from step (d) is divided. , sending the remainder of the solution to the leaching solution in step (b), (f) removing substantially all of the zinc and other impurities from the portion of the solution split in step (e), and (g) dividing the portion of the solution into a portion. Electrolyze to produce chlorine gas while leaving dilute sodium chloride (h)
The remaining electrolyte is sent to step (b) and the chlorine gas produced in step (i) and (g) is sent to step (a) from sulfide ore concentrate containing lead, silver and zinc sulfides. A method for recovering metals is provided.

By means of the method of the invention, sulphide ore concentrates containing lead, silver and zinc sulphides can be treated to recover, in particular, silver and lead.

Metal recovery from the metal chlorides obtained from the chlorination process is carried out in such a way as to avoid the formation of impurities, including zinc chloride, in the sodium chloride leachate used to dissolve the metal chlorides formed in the chlorination process.

Instead of wet chlorination of sulfides, dry chlorination is carried out using dry chlorine gas, and heating converts sulfides to chlorides and vaporizes arsenic and antimony if present.

Dry chlorination has been found to be particularly effective when sulfides of the copperite-copperite system are present alone or in combination with other minerals such as galena.

After chlorinating the sulfides by either method, the metal chlorides are separated from the resulting solution, and the primary targets, lead and silver, are recovered from the separated chloride solution.

Lead chloride is crystallized by cooling, lead is recovered from lead chloride by molten salt electrolysis, and the obtained chlorine is sent to the chlorination process.

Silver can be separated from the lead chloride-free solution by cementation.

A portion of the lead and silver separated solution is divided and the remainder is sent to a sodium chloride leaching step.

Lead and silver removed by iron cementation precipitation,
A divided portion of the solution is washed with sodium carbonate, in particular to remove zinc and other metal impurities as carbonate.

Electrolysis of the resulting solution yields chlorine, which is sent to the ore chlorination process.

A portion of the resulting weak sodium chloride electrolyte is concentrated and sent to a sodium chloride leaching step to prevent the formation of zinc and other impurities if the process is carried out continuously.

The sodium hydroxide from the electrolysis is inter alia carbonated and the sodium carbonate obtained is used in the neutralization step.

By dividing a portion of the sodium chloride leachate after removal of lead chloride and subsequently adding it to the sodium chloride leaching process after removal of zinc and other metal impurities, chloride is added at approximately the same rate as that added from the ore. Zinc can be removed from the sodium chloride leachate.

This can prevent the formation of zinc chloride, which interferes with the dissolution of lead chloride in the sodium chloride leachate.

Additionally, most of the chlorine remains in the system since it does not leave the system as an impurity or other form of chloride.

The chlorine removed is removed as chlorine gas during electrolysis, and the chlorine is sent to the chlorination step with almost no loss.

The present invention provides a nearly pollution-free process in which no chlorine or lead vapors or compounds are released into the atmosphere.

Unlike pyrometallurgy, almost all sulfide sulfide is converted to elemental sulfur rather than sulfur dioxide.

Next, embodiments of the present invention will be described with reference to the drawings.

The concentrate was a galena-ammonite concentrate with the following composition: Silver 0.30-0.35
%Lead 68-70%
Antimony 0.80-1.4% total sulfur 14-17% zinc
4-6% iron
It is clear that other ores containing 2-4% lead, silver and zinc sulphides may be processed.

The reaction that occurs in the chlorination step is as follows: MS+C12→MC12 (M-Pb, Zn, Cu, Fe
, or Ag, etc.) S2+C12→S2Cl2 M5+82C1□→MCI□+3/2S2Sb2S2+
5C12→2SbC15+3/2S2 The concentrate is ground to U, S, Standard-65 mesh before chlorination, and the ground concentrate is dried.

Effective utilization of chlorine gas is achieved by contacting the crushed concentrate with chlorine in a countercurrent system.

As shown in Figure 2, the finely ground concentrate enters the upper end of the rotary kiln, and dry chlorine gas is introduced into the lower end of the kiln, which is the concentrate discharge end, so that the highest concentration of chlorine gas is produced. Contact with the concentrate, which has almost completed chlorination.

An inert purge gas such as nitrogen is supplied to the kiln along with chlorine to remove sulfur chloride as described below.

The entire length of the kiln is divided into two zones.

Zone 2 near the discharge end is operated at a temperature of approximately 115°C, and zone 1 at the upper end of the kiln is operated at a temperature of 80-115°C.

Chlorination occurred primarily in zone 1, and sulfur chloride gas was generated in zone 2.

The gas leaving the kiln containing vaporized antimony pentachloride (SbC15) is treated with a scrubber to remove recovered antimony compounds.

Arsenic can likewise be removed if the concentrate contains arsenic sulfide.

A quantity of chlorine gas per unit weight of concentrate is supplied to the kiln, and substantially all of the lead and silver in the concentrate are continuously converted to their respective chlorides.

It has been found that kiln temperatures of 50-150°C, especially in the range 80-115°C, are satisfactory.

The excess stickiness of the concentrate in the kiln will cause the temperature in the kiln to drop to about 1
This can be avoided by keeping the temperature below the melting point of sulfur at 19°C.

Since the reaction between lead sulfide (PbS) and chlorine is an exothermic reaction, cooling means are required in zone 1.

Alternatively, inert materials such as sand or recycle products can be added as concentrate diluents.

Heat can be applied within zone 2 to increase the vapor pressure of the sulfur chloride and purge it with an inert gas such as nitrogen.

It is desired that the amount of sulfur chloride in the chlorinated product be minimal to avoid hydrolysis during the subsequent sodium chloride leaching.

Such a hydrolysis reaction of 82C12 is expressed as follows: 52C12 + 2H20 = 2HC1 + H2S + 5O2 (or polythionic acid) The amount of chlorine gas to be treated varies depending on the composition of the concentrate to be treated.

In the case of galena-ammonite concentrate, most of the chlorine is used to chlorinate the galena (PBS).

Concentrates of this type typically contain 70% lead and the theoretical amount of chlorine required for the reaction PbS+Cl2=PbC12+S is 215 kg per tonne of concentrate.

The total addition of chlorine per ton of concentrate is 260-280 kg, with the remainder 45-65 kg of chlorine chlorinating the sulfides of copperite and other metals present such as zinc, iron, copper, etc.

The following example shows the dry chlorination of a lead sulfide concentrate and the subsequent dissolution of the resulting metal chloride with sodium chloride.

Example: Chlorination conditions: Equipment: 3-chamber rotary kiln Zone 1 Reaction: C12 addition 260-270 kg/ ) PbS ore leaching conditions: Pulp density Chlorination production $y50 per leachate 11% sb evaporated during chlorination -59 NaC1 % Ag = 99.3 Pb299.9 Sb = 96 Zn = 33 Fe = 47 Cu = 97 These results indicate that more than 99% of the lead and silver present in the concentrate is chloride. , indicating that it was extracted during the sodium chloride leaching.

Additionally, most of the antimony was recovered.

Almost all of the sulfide sulfur was converted to elemental sulfur during the dry chlorination process.

Controlled chlorine addition (260-280 kg per ton of concentrate)
) with low temperature (80-115°C) dry chlorination followed by sodium chloride leaching at 90-95°C for 1 hour to obtain 99% of silver, 99.9% of lead and 3% of zinc.
3%, 47% of iron, 97% of copper, and 96% of antimony were extracted.

During chlorination, antimony was probably vaporized as 5bC15 and recovered from the exhaust gas.

Arsenic, if present, can also be recovered in this manner.

Almost all the sulfide sulfide in metal sulfides is converted to elemental sulfur unlike in pyrometallurgical processes where sulfur is released as harmful sulfur dioxide.

According to the flow sheet shown in FIG. 1, the leaching shown in the above example is carried out as follows.

Regardless of whether dry or wet chlorination is used, the flow sheet of Figure 1 is the same after the chlorination step.

The chlorinated product is leached with sodium chloride solution to dissolve lead and silver chlorides and other metal chlorides that act as impurities.

After start-up, the sodium chloride leachate is continuously circulated as shown.) I) The ram is refilled.

The leachate of lead ore concentrate usually contains 26 to 280 sodium chloride, 15 to 30 g of lead, 15 to 30 g of ferrous iron, and small amounts of copper, antimony, calcium, magnesium, manganese, aluminum, etc. .

Regardless of the concentrate being treated, the sodium chloride leaching step is preferably carried out at a temperature of 80-100°C.

The leaching slurry is filtered while hot and the residue is discarded or.

Optionally treated for elemental sulfur recovery.

Next comes lead recovery.

Dissolved lead chloride is 80-100% sodium chloride leachate
℃ to about 15-20℃.

The resulting crystalline lead chloride is separated from the solution, for example by centrifugation, dried and electrolyzed in a molten salt cell to obtain lead, and the chlorine gas is circulated to the chlorination process.

Silver can then be recovered by cementation using metallic iron or lead from the sodium chloride leachate from which the lead chloride has been separated, yielding an impure silver sponge containing copper, lead, iron and other trace impurities. It will be done.

Pure silver can be obtained by refining this sponge.

A portion of the leachate from which lead and silver have been separated is removed and sent to the sodium chloride leach as shown.

Approximately 5-15% of the metal-separated leachate is then treated to remove impurities, particularly zinc chloride, and the decontaminated solution is sent to a sodium chloride leaching step.

This makes it possible to control the concentration of zinc chloride and other impurities in the leachate.

Zinc chloride significantly reduces the solubility of lead chloride in sodium chloride solution, so to dissolve large amounts of lead chloride, zinc chloride and other impurities were specifically used to separate the metal at approximately the same rate as those introduced in the chlorination process. It is removed from a portion of the exudate.

Treatment of a portion of the metal-separated solution makes it possible to remove impurities in a form other than chloride.

Removal as chloride would result in loss of chlorine from the system.

Chlorine is recovered as a gas and sent to the chlorination step, thereby avoiding loss of chlorine from the system.

As shown in FIG. 1, the lead remaining in a portion of the demetallized solution is removed by cementation with metallic iron, and the resulting sponge lead is cyclically sent to a silver cementation process.

All precipitated silver is also circulated and the concentration of lead in this portion of the solution is reduced from about 15 g to 0.2 parts.

This divided portion of the solution is then diluted with sodium carbonate to approx.
．． At temperatures of about 50 DEG to 80 DEG C., zinc, iron and other metal impurities precipitate as carbonates in easily filterable form.

Sodium carbonate is used because it reacts with zinc chloride to yield sodium chloride, which is subsequently electrolyzed so that chlorine is not lost from the system when zinc and other impurities are removed.

After removing solids, the split solution is electrolyzed to yield chlorine gas, sodium hydroxide, and dilute sodium chloride solution.

Electrolysis is greatly facilitated by removing zinc and other impurities from the solution beforehand.

This is because if zinc or other impurities are present in the electrolyte, electrolysis becomes almost impossible.

The sodium hydroxide is carbonated to obtain sodium carbonate which is recycled to the neutralization step.

The chlorine gas is sent to the chlorination step and the purified sodium chloride solution is concentrated and sent to the leaching step to prevent the formation of zinc in the leachate.

The process according to the invention can of course be carried out both continuously and batchwise.

Based on the results obtained using dry chlorination, the material balance for typical lead sulfide concentrates available on the market (galena-ammonite) is as follows:

All added chlorine gas was used internally.

This table shows that theoretically all of the lead and silver can be recovered from the system without loss of lead or silver.

After start-up, there is little need to add chlorine to the continuous working process due to losses due to mechanical operations such as filtration, concentration, etc.

Although the present invention has been described with respect to the treatment of copper ore concentrate containing lead, silver and zinc and the use of a dry chlorination process, the invention is not limited to the treatment or dry chlorination of this ore.

Dry or wet chlorination can generally be used for ores containing lead, zinc and silver.

The flow sheet of FIG. 1 allows the chlorination step and subsequent steps to be continued regardless of the method of chlorination.

Furthermore, metals can be recovered from the chlorides obtained by wet chlorination of sulfides and results comparable to those shown in the previous examples can be achieved.

[Brief explanation of drawings]

FIG. 1 is a flow sheet showing the method of the present invention, and FIG. 2 is a flow sheet showing the chlorination step.

Claims (1)

[Claims] 1. A method for recovering metal valuables from sulfide ore concentrate containing sulfides of lead, silver, and zinc, comprising: (a) chlorinating the concentrate to change metal sulfides into metal chlorides; and (b) leaching the non-gaseous product of step (a) with an aqueous sodium chloride solution to dissolve the lead and silver chlorides therein so that they can be separated from the insoluble solids. (C) cooling the sodium chloride leachate to precipitate almost all of the lead chloride therein, and then separating the precipitated lead chloride from the leachate; (d) separating the lead chloride obtained in step (C); (e) divide a portion of the solution resulting from step (d) and send the remainder of the solution to the leachate of step (b); (g) partially electrolyze the divided portions of the solution to obtain chlorine gas, leaving a dilute sodium chloride solution; (h)
A method for recovering metal from ore, characterized in that the remaining electrolyte is sent to step (b), and the chlorine gas obtained in step (i) and (g) is sent to step (a).