DE2060408C - Process for the recovery of lead and silver from residues of the sulfuric acid leaching of rusted zinc ore concentrates and / or of complex lead-zinc ore concentrates that arise in the course of the electrolytic extraction of zinc - Google Patents

Description

The invention relates to a method for the recovery of lead and silver from in the course of Production of electrolytic zinc residues from sulfuric acid leaching from consoled zinc ore concentrates or from complexes Lead-zinc ore concentrates.

The analyzes of the complex lead-zinc ore concentrations result in values which usually lie between the following figures: 10 to 45 ° / · Zn; up to 40% Pb; 0 to 25 ° / ₀ Cu; 0 to 100 g / metric ton of Ag; S to 30% Fe.

The sulfidic zinc ore concentrates are roasted in order to convert the zinc sulfide into zinc oxide, which is then dissolved in aqueous sulfuric acid.

The resulting zinc sulfate solution is purified and then electrolyzed to recover the zinc; at the same time regenerates during the Oxygen released by electrolysis, sulfuric acid, which is used to dissolve other zinc oxide. This acid, known as the "spent electrolyte," has a fine concentration between and 200 g per Lite. »», SO «is too vague • During the roasting of both the zinc ore concentrates and the complex lead-zinc ore concentrates, sieve connects the iron with the rest of the before existing metals and forms compounds, ferrites, with the general formula MeO-Fe ₅₁ O, in which Me represents any of the metals that are present with the iron in the concentrates.

The ferrites MeO - Fe ₁₁ O are insoluble in the aqueous solution of sulfuric acid, which is used for leaching or dissolving the zinc oxide, under normal operating conditions. The higher the concentration of iron in the mineral concentrate, the greater the number of metals that are stirred into the insoluble form during roasting. Following the leaching stages, a residue is obtained whose zinc content can vary within very wide limits, namely essentially between 15 and 35%. depending on whether the leached material is a concentrate of normal zinc ore or whether it is a complex concentrate

Over the past few years, both zinc and the rest have been in The valuable McUlIt contained in these residues has already been proposed in numerous or varied processes. Some did not go beyond mere laboratory tests; others have found suitable industrial use such. B. a method proposed by the inventor for the recovery of zinc from ferrites (German patent 1 295 840). In this process for the extraction of zinc from beferrite-containing residues by means of alkalis with hot excess sulfuric acid, removing the residue obtained in this way by decanting and / or filtering, neutralizing alkali solution with residual material containing zrakoxydeiu, the ferrites are leached using sulfuric acid of a certain concentration, in such a way that at the end the decomposition of ferrites a certain sulfuric acid concentration occurs, after which the heavy and Insoluble residues containing precious metals are separated off, the separated from the residue Lye diluted with cell acid of a certain acidity and then with sintered residual cover is neutralized, the temperature being kept between 90 and 95 ° C. throughout the treatment will.

In all of the proposed processes, the dissolution of the zinc and iron goes hand in hand with that simultaneous dissolution of other valuable elements, such as Cd, Cu, etc., taking place in high proportions

The disadvantage is that in all processes an insoluble final residue results, which is a very high one Percentage of Pb and Ag and other precious metals contained in the starting concentrates are. This insoluble residue is usually passed on to a lead smelter for processing. So far, no hydrometallurgical process is on industrial level known for the recovery of lead, silver and residues of precious metals, which is used for this purpose.

The present invention shows a way of working up the residues mentioned in an economical way hydrometallurgically.

When using the two above-mentioned types of concentrates as raw materials, it was in in both cases after the conventional leaching and after carrying out the process for the recovery of zinc from the ferrites, according to the Spanish patent 304 601, obtained an end product with the following analysis:,

■ Zntotal 1.49 · /. As = 0.006 · /. Mn = 0.34 ° /.

"i *. Pb 18" / · Sb = 0.01 · / · Ag = 650g / metric ton

Cu 0.12 · / » Ge = 0.0005V. Ni == 0.0005 · Λ

Cd, 0.011V. Fe = 3V «Co ^ = 0.0005 · /.

CaO = 15.85 x / x; MgO = 1.45 · / .; S - total = 15.80 · / .; S-SO ₄ "= 15.08 · /;; Insoluble residue = 4.68 · / ·.

Needless to say, this pro-fait is due to a separation of lead chloride as a result

Centers within the broadest limits fluctuate as its solubility decreases with decreasing

be able; this always depends on the original to- avoid temperature

^ composition of concentrates &. _{e) The} leaching time is determined by the

ΛίϋΤι ^ k ^^ tdasSa ^ f for the ^ ost part in the _{sense of the usBchkeit} the borides of Pb and Ag, under abSübewhlondiniddasBfaabBleKUuateni & alten. the best & ending conditions.

Even small amounts of silver from silver sulphide were sold to ^J

noted From this solution both metals, lead and

The method according to the invention is based on the fact that SÖber, as already mentioned, are insoluble that both Sflberchlorid and lead sulphate sw salts, ζ. B. as sulfides, or by one on top of the other an excess of precipitations present as anion, according to their position in the electro-chlorine Go into solution and corresponding compound chemical series, precipitated fertilize the chlorine soil. In the first case, after removal of the Cl'-

The object on which the invention is based is to produce the sulfides of Pb and Ag ions by washing solved in that the residues with a preference as recovery of both metals according to conventional chloride-saturated and acidic solution, preferred methods are fed to a lead smelter, wisely in the presence of substances that cause the oxy- In the second case, the silver can come out of the solution,

dicren of the sulphides of those present in the residues which contain lead and silver as chlorides, Favor metals - e.g. B. of copper chlorides - are precipitated with lead.

at a temperature between room temperature and boiling point. This precipitation takes place almost instantaneously the solution lying temperature and in a -lieh, and it is sufficient for this to be a slight stirring of the or more stages, the lead and the silver being solution. The result is a product whose can be obtained in the form of dissolved salts, the Ag content can be over 80%. For this product act. can the silver using conventional methods, such as

However, there are various factors that influence this refining, dissolving and subsequent felling with solution or solubility; among them are copper, etc., which are recovered: the Cl 'concentration, the acidity, the dissolved lead is precipitated with metallic zinc. Substances that _{convert the Ag 1} S into ClAg and the temperature In this way, a product is obtained whose temperature. Pb content is over 99% This precipitated lead shows

Tests showed the following: 40 some extremely good characteristics with regard to its late

further conversion to lead oxide.

a) The solubility of Pb and Ag increases with the. The subsequent recovery of the zinc, which was used in the precipitation of the concentration of Cl 'ions, the lead in the solution, occurs because it is present. Because of this, the insoluble salt is separated off. After previous use of a saturated water ₄₅ ClNa in washing out the chloride content, the zinc is preferred to he solution. used again in electrolytic production

b) It has been found that the acidification of the chloride D »e following working examples allow the solution containing the solution to increase the recovery of Ag tolerably due ^to two of the factors given above. The effect of H ₁ SO ₄ is known: the acid content and the existing ClCu; proven. The introduction of SO ₄ ions into the solution as they explain at the same time the practical implementation, however, has the disadvantage that the process according to the invention.

with the increasing concentration of SO ₄ ions

the solubility of the Pb decreases. This disadvantage in Example 1 can be remedied by adding CaCl ₁ , which removes it from the solution by the formation of CaSO _4. 300 g (dry weight) residue were also used. The ClII offers the advantage of adding more Cl 'ions to the solution with a content of 18% Pb and 550 g Ag per metric solution and thus helps to reduce the losses both mechanically and saturated with two liters of ClNa Solution for the duration of 2 hours, in the case of permanent type, which can occur during the implementation of room temperature, in a beaker with assistance procedure, to compensate. ^6o would take an agitator exhausted After the end of _x _ j. "J ... - _{,. N.} .., j this time the stirring was stopped, the fixed one

c) It was found> continue thatthe Vorhanden- _Sto ff abffltriert and washed out

its certain substances, such as the copper chloride, _{which became the fittest} p ^^ of _{this first leaching}

which convert the Ag ₁ S into AgQ, during the _exposure y _{t2 in the area saturated with clNa}

Solution of Ag «fie have a beneficial effect. ₆₅ ^, "^ _unier two hours of stirring at room temperature

d) The temperature has a favorable effect on the solution temperature. After this time, the speed of Pb and Ag a. Nonetheless, stop stirring, filter the solid product

i less this factor requires special care; χ separated, washed and dried.

The following results were achieved:

the solid material separated and washed out. The solid

2iSft

= 9.50 gfl /. Ag + = 0.004 g / l

After the second leach: Ph ++ -

Ae + =

ÄB5

ml of Ha (d = 1.15) acidified solution. Stirring was carried out for 2 hours at room temperature. After this time, stirring was stopped.

Weight of the final residue: 265 g. Investigation of this residue:

Pb =

These numbers represent a yield of Pb and Ag of 70.3% and 9.6%, respectively.

The experiment was repeated under the same conditions, with the only exception that 2 ml of HCl (Density 1.19) were added at each leach stage.

The following results were achieved:

Analysis of the obtained after the first leach Solution:

Pb ++ =

Pb ++ = 14.0 g / l, Ag + = 0.065 g / L

After the second leach: p _{b ++} _ ₁₀ 2 η -β-β

Z η mn η - o, U07g / L

Analysis of the obtained after the second leach

Solution:
Tf _

vL Z Wrn * ~ n

Weight of the final residue: 235 g. Analysis of this residue:

Ag = ^ metric ton. ·

These numbers represent 90 and 88 percent yields of Pb and Ag, respectively.

P? ^v e * rauct was repeated under the same conditions, with the only exception that 5 g of ClCu was added at each leaching stage. The following results were obtained:

Weight of the final residue: 253 g. '

Analysis of the examination of the residue obtained after the first leaching: 35 Solution:

^Pb ~ ² ^ ⁰ / *

Ag = 85 g / metric ton. . These numbers represent a yield of Pb and Ag from analysis of those obtained after the second leach

89.6 or 87.4%. 40 Solution:

Pb ++ = 9.72 g / L Ag + = 0.007 g / l.

Weight of the final residue = 235 g.

Analysis of this residue: Pb = 2.2%, Ag = 35 g / Tm.

These numbers represent 90 and 95 percent yields of Pb and Ag, respectively.

Example 2

There were 300 g (dry weight) residues with a content of 18% Pb and 550 g Ag per metric ton with 2 liters of a solution saturated with ClNa and acidified with 2 ml HQ (d = 1.19), at room temperature and for a period of time of 2 hours, leached in a beaker while stirring. After this time, stirring was stopped,

Claims (6)

Patent claims: 1. Process for the recovery of lead and silver from the residue of the zinc produced in the course of the eJektrolytischea extraction Sulfuric acid leaching of roasted zinc ore concentrates and / or complex lead-zinc ore con- "Centrates, characterized in that the residues of a leaching with a saw a solution of chlorides are subjected, preferably in the presence of substances that favor the oxidation of metal sulfides present in the residues, at a temperature between room temperature and the boiling point, and in one or more stages. 2. The method according to claim 1, characterized in that the acidic solution with chlorides saturated is ao 3. The method according to claim 1, characterized in that copper chloride are used as substances which favor the oxidation of the sulfides. 4. The method according to claim 1, characterized in indicates that the silver and lead present in the solution are separated by precipitation, lead and zinc are preferably used. 5. The method according to claim 4, characterized in that that for the precipitation of the lead zinc used is separated from the solution as an insoluble salt. 6. The method according to claim, characterized in that the obtained insoluble zinc Salt is washed out for the excretion of the chloride and introduced as a raw material for the extraction of electrolytic zinc.