Journal Archive

Platinum Metals Rev., 2004, 48, (3), 125
doi: 10.1595/147106704X1667

Treatment of Platinum Flotation Products

  • A. V. Tatarnikov*
  • I. Sokolskaya
  • Ya. M. Shneerson*
  • A. Yu. Lapin
  • P. M. Goncharov
  • All-Russian R & D Institute of Chemical Technology,
  • 33 Kashirskoye Road, 115409 Moscow, Russia
  • Gipronickel Institute JS,
  • 11 Grazhdansky Prospekt, 195220 St. Petersburg, Russia

Article Synopsis

A flowsheet has been developed for the production of rich concentrates of precious and non-ferrous metals by a complex treatment of the flotation products from South African platinum-containing chrome ores. The procedure involves: autoclave leaching, roasting, hydrochlorination and precious metal recovery by sorption. Autoclave oxidative leaching of the initial material allows the non-ferrous metals to pass into solution from where they are recovered as a rich sulfide concentrate (> 30% nickel and copper). Recovering the precious metals into solution combines two operations: sinter roasting and hydrochlorination. Roasting destroys the precious metal acid-proof mineral forms. The precious metals are recovered from solution by ion exchange using anionites that are finally burned. The ash from the burning is a concentrate of precious metals (> 75% in total) which are recovered in three forms: ammonium chloroplatinate with purity > 98%, palladium dichlorodiamine with purity > 96%, and a mixture of rhodium, ruthenium and iridium hydrates. The flowsheet uses full water rotation and minimum consumption of reagents, and gives a good recovery of metals to commodity concentrates (nickel > 98%, copper > 80% and precious metals (in total) > 95%).

Most of the primary platinum group metals (pgms) comes from low-sulfide platinum ores (South Africa and the U.S.A) and from sulfide copper-nickel ores (Russian and Canadian deposits). Recent exploration in Russia and elsewhere has resulted in the discovery of some new deposits of platinum-bearing low-sulfide ores. Many existing technologies for treating pgm ore are based on mechanical beneficiation, high-temperature smelting and converting operations, and hydrometallurgical processing. Therefore, the development of complex hydrometallurgical technology to recover pgms and non-ferrous metals from low-sulfide pgm-bearing ores has considerably simplified the treatment, decreased the operating costs and improved the environmental conditions.

Our investigations involved flotation concentrates of low-sulfide platinum-bearing chrome ores from South Africa. The ore had the compositions shown in the Table. An X-ray microscopy study of the phase composition of the flotation concentrate showed that nickel is present as pentlandite and copper as chalcopyrite. Iron is present in these minerals and is also present as pyrrhotite and in rock-forming minerals (pyroxene, spinel and talc). Platinum group metals are found as their own sulfides and in mixed sulfide minerals, which are either individual, or associated with non-ferrous metals sulfide minerals and pyrrhotite, see Figure 1. Minor amounts of gold and ferroplatinum in metallic forms have also been discovered.

Fig. 1

Optical microscopy shows platinum group element minerals associated with pentlandite and chalcopyrite in reflected electrons (Cp is chalcopyrite)

Based on results of the investigation, the preferred treatment of the flotation concentrate is that shown in the flowsheet, see Figure 2.

Fig. 2

Flowsheet showing the treatment of flotation concentrates of low-sulfide platinum-containing ores

Autoclave oxidative leaching (AOL) at a temperature of 150°C and O2 partial pressure of 1 MPa enables non-ferrous metals to pass into solution selectively. This does not involve precious metal minerals as they are resistant to inorganic acids (1). The effect of sulfuric acid consumption and the process time on the AOL performance were studied. The AOL values were found to be dependent on the sulfur content in the initial material. For example, during dissolution of concentrate containing a relatively high sulfide content (about 7% S in a sulfide form) the AOL process is carried out in an automatic mode without sulfuric acid addition to the initial slurry. Nickel and copper recoveries are 96 and 80%, respectively, into solution.

Chemical Composition of the Initial Concentrates

Component Content

Concentrate 1, % Concentrate 2, %

Cu 0.62 2.07
Ni 1.04 3.11
Fe 7.9 12.4
S 1.72 7.09
Mg 9.3 9.45
Al 3.1 3.19
Cr 2.8 1.22
SiO2 42.24 38.5

Concentrate 1, gt-1 Concentrate 2, gt-1

Pt 156.0 450
Pd 74.7 310
Rh 26.3 106
Ru 52.7 151
Ir 0.59 37
Os 1.2 26
Au 1.91 4.2
Ag not analysed 17.4

Leaching of feed with relatively low sulfide content (< 2% S in a sulfide form) is possible under the same conditions, but with acidic additions to the initial slurry (30% of the solids' weight). In this case 94% Ni and 62.5% Cu are recovered into solution. Platinum group metals recoveries into solution are not high for these types of concentrate and are not more than 0.5% for platinum and palladium, and not more than 1.2% for rhodium and ruthenium. The AOL process is considered in (2).

Platinum group metals are present in the insoluble leach residues in natural chalcogenide mineral forms; they do not undergo chemical transformation at the AOL stage. Platinum metals may be recovered by a combination of oxidative roasting and hydrochlorination. At the roasting stage pgm minerals are broken down to produce metallic forms which are then dissolved during hydrochlorination to form pgm complex compounds.

The effect of the roasting temperature and time on the Pt and Pd recovered into the chlorination solution is shown in Figures 3(a) and 3(b). The optimum conditions for oxidative roasting are: temperature 1000°C, time 1 hour, and rate of heating the material 6.5–7.5°C min−1. In order to decrease sulfur dioxide production, a method to change sulfur to sulfate during roasting by adding calcium oxide to the initial feed was tested.

Fig. 3(a)

The effect of roasting temperature on platinum and palladium recovery into chlorination solution (roasting time = 2 hours)

Fig. 3(b)

Effect of roasting time on platinum and palladium recovery into chlorination solution (roasting temperature = 1000°C)

The effects of parameters, such as the type of oxidising agent, HCl concentration, slurry density, and time and sequence of the process on the final hydrochlorination values were studied so that maximum amounts of pgm could be recovered from the sinter. The effects on the pgm recoveries into solution of preliminary grinding before sintering and reduction by hydrazine were estimated. Chlorine gas was the preferred oxidising agent at the sinter hydrochlorination stage rather than hydrogen peroxide or potassium permanganate. The optimal acidity of the slurry for maximum pgm recovery into solution corresponds to an HCl concentration of 220 g l−1, see Figure 4, consistent with industrial practice (3). The platinum metals associates and rhodium, in particular, are sensitive to variations in the HCl concentration. The pgms recovered at the hydrochlorination stage decrease as the initial slurry density grows, see Figure 5. The maximum drop in recovery is seen with a L : S (liquid : solid) ratio change from 3 to 2. Variations in the slurry density in the range L : S = 3–5 have practically no effect on the platinum metals recoveries. When the chlorination time was increased from 2 to 4 hours, there was an increase in the pgms recoveries into solution. In general, platinum metals recoveries increase when the two-stage leaching process is used in a countercurrent mode. During a two-stage leaching process in solutions containing 170 g l−1 HCl, platinum extraction exceeded 99%, while extraction of palladium and platinum metals associates was over 90%.

Fig. 4

Platinum metals recoveries into the chlorination solution as a function of HCl concentration (L : S ratio = 10, chlorination time = 2 hours)

Fig. 5

Platinum metals recoveries into chlorination
solution as a function of initial slurry density (170 g l−1 HCl, chlorination time = 2 hours)

The platinum metals recoveries as a function of the time and sequence of the chlorination process are given in Figure 6. Sinter grinding and preliminary reduction by hydrazine allow the platinum metals recovered into solution to increase by 5–10% on average, see Figure 7.

Fig. 6

Effect of the chlorination time on platinum group metals recoveries into solution (120 g l−1 HCl, L:S ratio = 2)

Fig. 7

Effect of sinter grinding on platinum group metals recoveries into solution (220 g l−1 HCl, L : S ratio = 2, chlorination time = 2 hours)

Particle size analysis of the final chlorination cakes showed that the size distribution of the platinum metals is in proportion to the size yields. The maximum pgms content occurs at a size of ∼ 10 µm while the minimum content occurs in the range > 44 µm but < 74 µm. The minor concentration of rare platinum metals (Rh, Ru, Ir) in the ∼ 10 µm fraction is noted. Mineralogical analysis of the hydrochlorination cakes, see Figure 8, showed that all the platinum metals are present in acid-soluble forms (intermetallic compounds, dioxides). The incomplete recovery of platinum metals into the chlorination solution may be explained by their isolation by the rock minerals. An additional flotation recovery of platinum metals from the final hydrochlorination cakes enables the production of sufficiently rich concentrates (∼ 500 g t−1) with the mass of the initial material decreasing by up to three times. The concentrates may then be returned to the roasting stage of the flowsheet.

Fig. 8

Platinum group metals and their compounds in the final hydrochlorination cakes

Platinum metals recovery from hydrochlorination solutions was carried out by ion exchange (sorption). Anionite Rossion-11, a porous sorbent based on styrene and divinylbenzene with functional groups of primary, secondary and tertiary amines (−NH2, =NH, ≡N) groups, characterised by high selectivity towards the pgms, was used as the sorbent. It was found that for maximum amounts of pgms to pass to the sorbent, the HCl concentration in solution should not be higher than 125 g l−1, while the ferric iron concentration should not be more than 15 g l−1. The solution oxidation-reduction potential should be < 800 mV; this is also important to avoid sorbent destruction.

During the ion exchange process, with the rate of solution passing being: 2 volumes of solution to 1 volume of resin per hour, the following residual pgm content in the sorbate was (in mg l−1): platinum < 1; palladium 1–1.5; rhodium < 1; ruthenium 3–4; iridium 1–2. The residual contents correspond to a total pgms recovery of 95 to 98%. However, if the sorbate, containing residual amounts of pgms, is used to prepare hydrochlorination solutions – thus returning pgms to the chlorine leaching process – the pgm recoveries at the ion exchange stage will reach 99.5%.

During the investigations it was found that the optimum ion exchange modes developed using simulated solutions are reproduced in existing technological solutions for all the pgms, except rhodium. The reasons for the incomplete recovery of rhodium into the sorbent during ion exchange have been determined. Rhodium recovery into solution decreases as the sinter hydrochlorination time reduces due to the production of poorly sorbed forms of rhodium, such as [RhCl6]2−, [RhCl6−x(H2O)x]x−3, [RhCl6−x(H2O)x]x−2, etc. Based on the chlorination process time of 2 hours the sorption recovery figure is 70.3%. An additional two-hour treatment of the solution by chlorine at 90°C helps to increase rhodium recovery to 76%, while a six-hour treatment increases it to 93.8%. It is also noted that if the six-hour solution is held at 90°C (while mixing), then the free chlorine in solution facilites rhodium transformation to the most sorbed form, ([RhCl6]3−), and increases rhodium recovery into the sorbent up to 97%, see Figure 9.

Fig. 9

The degree of rhodium sorption recovery from solutions (ERh) as a function of the total chlorination time (sinter hydrochlorination and additional treatment of solution)

The resin produced from the sorption contains pgms and gold in the amount 90 kg of precious metal per tonne of air-dry resin. These studies showed there was the potential for pgm recovery from resin by both combined and selective pgm desorption and sorbent burning.

Burning the preliminary dried resin for 4 hours at 1000°C results in pgm-rich concentrates. The pgm and gold content in the concentrates is 80–85%. The ash yield (pgm concentrate) is 3–6% of the weight of the sorbent feed for burning. Gases, such as CO, CO2, CH4, are produced from the burning resin, while nitrogen oxides are formed when the resin is burned in an oxidising atmosphere. These gases require attention before being vented to the atmosphere.

It was found earlier, that anionites based on vinylpyridine sorb platinum and palladium well from nitrate and chloride solutions. Palladium is easily desorbed from anionite with dilute ammonia solution (5% on NH3). Technology based on this has been used to separate platinum and palladium from silver nitrate electrolyte at an affinage works since 1994. Electrolytes containing as much as 600 g l−1 silver and 80 g l−1 copper, have undergone purification. The platinum content of the electrolytes is in the range 30–500 mg l−1, and the palladium content is in the range 200–2500 mg l−1. The exchange resin capacity for platinum and palladium together is up to 45 g l−1. Palladium is desorbed from the resin by ammonia solution, precipitating palladium dichlordiamine salt of purity more than 99% (∼ 99.9%) from the palladium strippant. The resin is returned to the sorption. The platinum accumulated in the resin after ammonia flushing needs to be desorbed with thiourea, to precipitate an acid-soluble platinic-palladium concentrate from the strippant (5).

Similar technology has been used since 2000 to process solutions from leached pgm materials. A two-stage pgm sorption recovery from hydrochlorination solutions was developed (Figure 10).

Fig. 10

Two-stage pgm sorption recovery from hydrochlorination solutions

In the first stage Pt and Pd are recovered. Anionite containing pyridine groups has been synthesised for this. The distinctive feature of the anionite is that ammonia can desorb platinum and palladium from it. Quadrivalent platinum is reduced on anionite to bivalent platinum and as a result can be transferred to ammonia solution by desorption. The optimum composition of the resin functional groups and the structure have been determined. After oxidation, ammonia chloroplatinate and palladium dichlorodiamine are precipitated from the ammonia strippant. The anionite capacity is ∼ 35 g l−1 for platinum and palladium together.

In the second stage Rh, Ru and Ir are recovered by Rossion-11 anionite. Platinum group metals are desorbed from anionite with hot (60–80°C) thiourea solution. A mixture of Rh, Ru and Ir hydrates is precipitated from the thiourea solution. The iron and non-ferrous metals content of the concentrate produced does not exceed 4% of the total (Rh, Ru and Ir).

At the oxidative roasting stage of autoclave leaching the insoluble residue, and during burning the saturated resin, sublimation of volatile osmium and, partly, ruthenium oxides to the gas phase takes place. It is therefore necessary to collect, recover and then separate the products. Methods of collecting osmium tetraoxide from roaster gases are employed at existing plants in Russia (4).


A flowsheet for treating flotation concentrates produced from low-sulfide platinum-containing ores has been developed. It provides low consumption of cheap reagents, complete water circulation and gives a high recovery of platinum group metals and non-ferrous metals into rich concentrates. These are then suitable for refining.


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  2.  Ya. M. Shneerson, P. A. Goncharov, A. Yu. Lapin and V. M. Shpaer, ‘Selective recovery of non-ferrous metals from complex feed’, Proc. of Gipronickel Institute JS, St. Petersburg, 2000
  3.  M. A. Meretukov and A. M. Orlov, ‘Metallurgy of precious metals (foreign experience)’, Moscow, Metallurgy, 1990, 416 pp.
  4.  T. N. Greiver, E. L. Kassatsier, T. V. Vergizova and V. M. Khudyakov, et al., Russian Patent, 2,044,084, C22B 11/00, published in Bulletin 261995
  5.  A. Tatarnikov, I. Sokolskaya, Ya. Shneerson, A. Lapin and P. Goncharov, ‘Complex treatment of platinum flotation concentrates’, Metallurgy, refractories and environment; Proc. Vth Int. Conf. Metallurgy, Refractories and Environment, Stara Lesna, High Tatras, Slovakia, 2002
  6.  Ya. M. Shneerson, A. Yu. Lapin, P. A. Goncharov, V. M. Shpayer, T. N. Greiver, G. V. Petrov and A. V. Tatarnikov, ‘Hydrometallurgical processing technology for low-sulfide platinum-containing concentrates’, Tsvetnye Metally, 2001, (3)26

The Authors

A. V. Tatarnikov is a Senior Staff Scientist at the All-Russian R & D Institute of Chemical Technology, Moscow, Russia. He is involved in the recovery of noble and rare metals from ores and byproducts. His speciality is sorption concentrating of platinum group metals from solutions and slurries.

I. Sokolskaya is a Research Scientist at the All-Russian R & D Institute of Chemical Technology, Moscow, Russia. Her interests are in the synthesis of sorbents with given characteristics for application to the recovery of platinum group metals fromn solutions.

Ya. M. Shneerson is Head of the Hydrometallurgical Laboratory at the Gipronickel Institute JS, St. Petersburg, Russia. His interests are in the hydrometallurgical treatment of ores and concentrates containing non-ferrous and noble metals.

A. Yu. Lapin is a Leading Staff Scientist at the Gipronickel Institute JS, St. Petersburg, Russia. He is interested in non-ferrous and platinum group metals recovery from concentrates by means of hydrometallurgical autoclave treatment.

P. M. Goncharov is a Senior Staff Scientist at the Gipronickel Institute JS, St. Petersburg, Russia. He is involved in the treatment of ores and byproducts containing noble metals by means of hydrometallurgical autoclave treatment.

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