DE2340157A1 - PROCESS FOR SEPARATING AND PURIFYING GOLD IN A MIXTURE WITH THE METALS OF THE PLATING GROUP - Google Patents

Description

at "nt" nvint

Dr. KBoettner ' P 6107

• DipUng. H.-J. Müller

Dr.Th.Berendt

D8 Munich 80

Vuoile Graha Slr. 38, TeL 47 51 55

SVARSAB HXNXNG, EXPLORATION & DEVELOPMENT COMPANY (PROPRIETARY) LIMITED, 56, Main Street, Johannesburg, Transvaal Province

(South African Republic)

Process for the separation and purification of gold, which is mixed with the metals of the platinum,

group is present

The present invention relates to the separation and purification of gold in one. Mixture with the metals the platinum group is present.

In general, the five metals of the platinum group come namely platinum, palladium, iridium, ruthenium and rhodium, in nature together with gold, and these six elements must be separated and purified, around in trade and. Industry to be able to find a useful use. As a rule, the first step is a breakup of these elements carried out in two groups, namely by treating .the basic mixture, which in general from a RohsteinCmatteJ leach residue or sludge

09/0909 '

consists of using aqua regia, and in this case go the platinum, palladium and gold in solution, while the iridium, ruthenium and rhodium are essentially in the Remaining residue that is produced in this process. The two groups of metals made in this way are then subjected to separation procedures which are generally lengthy and complicated, to separate and clean the individual metals.

The present invention relates to a process for the separation of gold from the metals of the platinum group, which is easier to carry out, at least in the case where a high-purity gold is required as the end product is, and / or which facilitates the separation of the metals of the platinum group and gold;

The inventive method for separating the gold from the metals of the platinum group consists in that one solution containing gold and one or more platinum group metals at least once with nitric acid evaporate to dryness, dissolve the residue in nitric acid, add hydrogen peroxide and the resulting Solution boils to precipitate and coagulate the gold as a metal.

Other features of the invention include that the solution, the gold and one or more metals of the platinum group is evaporated to dryness twice with nitric acid before treatment with hydrogen peroxide occurs that the process to be carried out with the precipitated gold after separation thereof is repeated and that the filtrate from the process with formic acid

Λ 09809/0909

^f - ο ~ υ ο

is cooked »about the precipitation of everything contained in it To ensure gold.

Evaporation to dryness with nitric acid is carried out, to the nitrosochloride complexes of the existing To form metals. '

Still further features of the invention are apparent from the following description of a preferred embodiment thereof which includes a specific example in which the amounts of the used Reagents are shown in parentheses.

Generally concentrates of the platinum group metals are used and obtained from gold as raw stone leach residue or sludge, which - processed starting from this stage to obtain the desired metals, and for this reason the following description continues with the processing of such a raw stone leach residue or sludge.

In the course of the description, reference is made to the attached flow charts Referenced by those

FIG. 1 shows a flow diagram which shows the general processing of the raw stone leach residue illustrates, and

FIG. 2 is a flow diagram illustrating the further Separation and purification of the gold explained.

The sludge used as the starting material in this aspect of the invention had the following composition on:

409809/0909

element

Platinum (Pt) 6.92

Palladium (Pd) '3.04

Gold - (Au) 0.59

Rhodium (Rh) 0.32

Ruthenium • (Ru) O.6o

Iridium (Ir) 0.08

Silver (Ag) 0.03

Aluminum oxide (Al ₂ O-,) 3.60

Antimony. (Sb) 0.05

Bismuth (Bi) 0.11

Calcium oxide (CaO) '0.49

Copper (Cu) 6.91

Cobalt (Co) 0.19

Iron (Fe) '4.57

Magnesium oxide (MgO) 0.35

Nickel (Ni) 3.50

organic substances ^ ~ ^ n ^ 2 ^ m-7 23.1

Potassium '(K)' 0.03

Silica '(SiO ₂ ) 12.92

Sodium (Na) 9.25

Sulfur (total) (S) 19.63 Tellurium ■ (Te). 0.29

Zinc (Zn) <0.001

other components

+ O ₂ + H ₂ O (as difference) 3.43

'■ 100.00

First, 2.5 kg of the sludge were roasted in a stream of air for two hours at 600 ° C. in order to convert all sulfides and the free sulfur into oxides in this stage 1 .

A09809 / 0909

In stage 2, the roasted material was refluxed with stirring for two hours with 20 # H ₂ SO ^ (3 liters), then cooled to about 55 ° C. and filtered to dissolve out most of the base metals present.

Calcium oxide (CaO) was added to the filtrate in step 3 in order to precipitate all of the nickel, copper and the metals of the platinum group and the gold contained in the piltrate. This precipitate was collected and it contained 30 mg platinum, 100 mg palladium, 10 mg gold, 50 mg rhodium and 40 mg iridium. In practical operation, this precipitation would be fed back into the raw stone melting plant (matte smelter).

The residue remaining after the sulfuric acid treatment was then leached with aqua regia (1.5 liters) for three hours in stage 4. This aqua regia leaching and all other aqua regia leachings were carried out as follows:

The material was refluxed with the required amount of hydrochloric acid (HCl) for 30 minutes. The required Amount of nitric acid (ENT.,) Was then gradually within added for 60 minutes. The mixture was allowed to continue boiling for the remainder of the time (i.e. 1 1/2 hours).

Then sodium bromate (NaBrO-,) (30 g) was added to the solution, which was simmered for another 30 minutes after the solution had been diluted with water (1.5 to 3 liters) to add the platinum group metals and gold to oxidize their highest degrees of oxidation. The p "was then with sodium carbonate (about 60 ° C) (1 kg) to 6.5

409809/0909

posed to precipitate all metals - except platinum - as hydrated oxides. The solution was left to stand for 20 minutes and then filtered or suction filtered under vacuum. In this way, the platinum was removed so as to reduce by about 50% the majority of the platinum group metals and gold which had to be alloyed with lead.

To the filtrate (+ 90 % platinum (Pt) in the feed) was added formic acid (0.2 liter) in step 5 and the solution was refluxed with stirring for 5 hours. In stage 6 , sodium carbonate (Na ₂ CO 2) (0.5 kg) was added in portions over the course of 1 hour until a value of

5.0 was reached. The solution was then boiled for a further 60 minutes, cooled to 60 ° C. and under vacuum filtered. These stages 5 and 6 were carried out to to precipitate the platinum in the solution.

The filtrate was passed through an anion exchange column in stage 7 and the effluent was separated out. This process contained 100 mg Pt, 10 mg Pd, 10 mg Au, 1 mg Rh and 25 mg Ir.

The precipitate was dissolved in aqua regia (1.0 liter) in step 8, the solution cooled and filtered. The resulting residue consisted of silver chloride (AgCl).

This aqua regia filtrate contained most of the platinum. This platinum was made with the bulk of the palladium and gold combined in the manner described below, and the separation of these platinum group metals was carried out using a separate procedure.

409809/0909

The combination of residue and precipitation - accumulated during the initial aqua regia leaching and the sodium carbonate precipitation in stage 4 (+ 14OO g) - was mixed with a flux A and a flux B, which are defined below, and in stage 9 75 Melted together at 1140 ° C for minutes. The consequence of this is that most of the rhodium is converted from an aqua regia insoluble form into an aqua regia soluble form. At the same time, the solubility of iridium and ruthenium in aqua regia is further reduced. 2.5 kg of flux A per kg of residue and 335 g of flux B per 100 g of platinum group metals plus gold to be collected were used. These fluxes had the following compositions:

Flux A (active flux) (3.5 kg)

(1) Borax Na ₃ B ₄ O ₇ : 20 % (0.75 kg)

(2) Sodium carbonate Na ₃ CO ₅ : 80 % (2.75 kg)

Flux B (collector) (0.55 kg)

(1) Black lead PbO: 90 % (0.5 kg)

(2) Charcoal C:. 10 % (0.05 kg)

The molten material was poured into iron molds and allowed to cool in them. The lead regulus obtained in this way was separated from the slag, and the slag was crushed and collected, but would in practice Operation to be returned to the raw stone smelting plant. The slag contained 520 mg Pt, 230 mg Pd, 45 mg Au, 24 mg Rh, 45 mg Ru (ruthenium) and 5 mg Ir.

Λ09809 / 0909

The lead regulus was crushed and then refluxed with 20 # HNO, (5.5 liters) in step 10 for 5 hours to remove the lead. The solution was cooled to 55 ° C and filtered under vacuum. The residue was sent to a later aqua regia leach stage referred to as stage 11.

In stage 12, formic acid (0.05 liters) was added to the Flltrat from the nitric acid leaching of the stage 10 and adjusted the p _H with sodium carbonate Na ₂ CO, (0.3 kg) to 2.0 to all the metals of the platinum group and to precipitate the gold that had been dissolved by the nitric acid. The solution was then stirred at room temperature for 5 hours and filtered under vacuum.

The resulting precipitate was combined with the residue obtained from the nitric acid leach step in step 10 and this combination of materials was fed to the aqua regia leach step in step 11. In stage 13 , NHj.OH was added to the filtrate to precipitate the lead and any platinum group metals present plus gold, and this precipitate was dried and calcined and, in practical operation, would be returned to the melt bleaching stage> ie, the stage

The combination of residue and precipitate was refluxed in stage 11 with aqua regia (0.75 liters) for 3 hours cooked. The solution was cooled to 55 ° C and filtered under vacuum. The residue formed part of the Feed for the process for the separation of rhodium, iridium and ruthenium from one another. This residue contained 850 mg Pt, 700 mg Pd, 100 mg Au, 500 mg Rh,

409809/0909

14720 mg Ru and 1420 mg Ir. The precious metals usually make up approximately 50 % of this as a metal concentrate, while the remaining 50 % consists of lead chloride (PbCl ₂ ).

To the filtrate from the aqua regia leaching stage of the stage, the stoichiometric amount of HgSO ₂ ^ plus 10 % excess (100 ml of a 50% H ₂ SO ^) was added in stage 14 to precipitate the lead contained in this solution. The solution was
filtered.

Solution was boiled for 30 minutes, cooled to 55 ° C and

The precipitated lead sulphate was converted into hot lead (PbO) in stage 15 by annealing and was ready for use again for return to the stage of enamel lead.

The filtrate from the lead precipitation stage was evaporated to dryness in stage 16, and the salts obtained were calcined at 600 ° C. for 2 hours. As has been found, this temperature is important in order to make the rhodium in the residue insoluble in aqua regia.

The glow product was refluxed with aqua regia (0.75 liters) for three hours in step 17. The solution was cooled and filtered under vacuum.

The residue was mixed with the residue from the aqua regia leach stage of Stage 11 carried out after alloying, were combined to give a combined Feed material for the separation of the secondary group of platinum group metals, namely rhodium, Ruthenium and iridium, the following contents of platinum metal

4Q9809 / 09Q9

len and gold: 1130 mg Pt, 85O mg Pd, 120 mg Au,

me

7480 mg of Rh, l4920 ^ Ru and 1720 mg of Ir.

That in the last aqua regia leach stage of the stage The resulting filtrate was combined with the filtrate that was used in the aqua regia leaching stage of the stage 8, which had been carried out with the metals that were dissolved in the leaching stage 4 carried out at the beginning had been accrued in order to produce a feed material for the processes for the separation of platinum, palladium, and win gold. This charge had the following levels of metals, platinum group and gold on: 1712IO mg Pt, 7 ^ 790 mg Pd, 14560 mg Au, 200 mg Rh, 25 mg Ru and 210 mg Ir.

This precipitate obtained through the practice of the present invention contained the following ingredients: 50 mg Pt, 10 mg Pd, 13750 mg Au, 40 mg Rh and a trace of ruthenium.

This impure gold was dissolved in aqua regia (0.06 liters) in stage 20. The solution was evaporated to dryness and the salts were boiled in stage 21 with NH ^ OH (0.6 liters) to dissolve all of the silver chloride. The solution was cooled to 55 ° C and filtered. The filtrate, which contained only a trace of silver, was discarded.

The residue was dissolved in boiling 5 # HNO (0.6 liter) in step 22. The insoluble residue was filtered off, and in practical operation it would be returned to a suitable point in the work-up process.

409809/0909

In stage 25 , the filtrate would be passed over the cation exchange resin "Dowex 50 χ 8 ^M ; the resin retained the base metals and the gold was washed out with water.

HpO ₂ (0.03 liter) was added to the drain in step 24 and the whole was boiled for 30 minutes in order to precipitate the gold. After cooling to 55 ° C., the solution was filtered.

The dried precipitate consisted of the finished gold, which had a purity of over 99.97 % gold.

The total gold yield was 9000 mg, which corresponds to a value of approximately 60 % .

The gold in the filtrate was in stage 25 solution precipitated by boiling the solution with formic acid (12 ml) and adjustment of the p _H of the Lö with Na ₂ CO, (10 g) at 5 »0th The solution was cooled and filtered. The filtrate was discarded; it did not contain any precious metals. The precipitated metal would in practical operation be recycled to an appropriate stage in the separation process.

By using the method according to the invention, gold of high purity can thus be obtained in one go easy and effective way to perform.

409809/0909

Claims (5)

Claims 1. Process for the separation of gold from the metals of the Platinum group, characterized in that there is a solution containing gold and one or more metals of the platinum group contains, evaporated to dryness at least once with nitric acid, the residue again in nitric acid dissolves, adds hydrogen peroxide and the resulting solution is heated to boiling in order to precipitate the gold as metal and to coagulate. 2. The method according to claim 1, characterized in that the solution containing the gold and the metals of the platinum group contains, is evaporated to dryness twice with nitric acid before precipitation with hydrogen peroxide he follows. 3. The method according to any one of the preceding claims, characterized in characterized in that the precipitated gold is redissolved in aqua regia, then using ammonium hydroxide is precipitated, dissolved in nitric acid and precipitated again with hydrogen peroxide to make a finished product Win gold product. 4. The method according to any one of the preceding claims, characterized in that the hydrogen peroxide in an amount of about 1 ml per gram of gold to be precipitated is used. 5. Gold, characterized in that it is made by any method as claimed in the preceding claims is, has been won. 09809/0909 Blank page