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Recovery of precious metals from electronic scrap, in particular from waste products of the thick-layer technique
14ydrometallurgy, 25 (1990) 99-110
99
Elsevier Science Publishers B.V., Amsterdam - - Printed in The Netherlands
Recovery of precious metals from electronic scrap, in particular from waste products of the thick-layer technique K. Gloe ~, P. Mfihl ~ and M. Knothe 2 Central Institute of Solid-State Physics and Materials Research, Dresden (German Democratic Republic) 2VEB Mansfeld Kombinat "Wilhelm Pieck", Research Institute of Non-Ferrous Metals, Freiberg (German Democratic Republic) (Received July 3, 1989; revised and accepted November 13, 1989 )
ABSTRACT Gloe, K., Mfihl, P. and Knothe, M., 1990. Recovery of precious metals from electronic scrap, in particular from waste products of the thick-layer technique. Hydrometallurgy, 25: 99-110. From electronic scrap, in particular from the waste products of the thickqayer technique (which is applied to produce passive electronic components) the Ag, Au, Ru and Pd contents are recovered by hydrometallurgical procedures. Ag, Pd and Au are dissolved at 20°C in 5 M to 9 M HNO3 or HCI/ C12 mixtures, being then precipitated from the leach liquors as concentrates. Coatings on the base of glass inhibiting the leaching of precious metals are removed by leaching with fluoride solutions, which results also in the stripping of Ru-bearing layers from the ceramic substrate. From the solid Ru-concentrates obtained, ruthenium is recovered and purified by alkaline fusion, precipitation and dissolution of RuOz-xH20 followed by cation exchange. After the precious-metal and glass interlayers have been stripped, the ceramic substrates can be re-used to produce passive components.
INTRODUCTION
Precious metals such as Au, Ag, Pt, Pd and Ru, due to their specific physical
and chemical properties, are essential constituents of electronic components. In the future, development will give rise to an increase in consumption of these metals. This inevitably results in an increased amount of waste products, residues and scrap containing precious metals. Thus, the recovery of the valuable constituents from these waste products is an important and beneficial task. In this paper possibilities of reprocessing waste products of the so-called thick-layer technique will be discussed. This technique developed to produce passive electronic components is now being applied to manufacturing hybrid circuits. To accomplish this, a paste consisting either of conducting metal 0304-386X/90/$03.50
© 1990 Elsevier Science Publishers B.V.
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powder (e.g., Ag, Au, Pd, Pt) or non-conducting oxide powder (RuO2, [Bi,Pb]Ru2OT), inorganic binder (fritted glass for adhesion) and organic binding agents for providing the physical properties necessary to the silk-screen process is printed onto a substrate (in general, A1203 ceramics). After removal of the organic constituents subsequent burning results in sintering of the metallic or oxidic constituents, thus providing the desired electrical properties. Due to the complicated manufacturing process and the necessity high precision of the components, even during manufacturing, waste products and residues are obtained. The waste products can be divided into two categories: (1) wastes of paste (incorrect charges and residues) which amongst other components contain the organic binding agents; (2) printed substrates in various processing states (unburned and burned, in part, with connection contacts and more or less enclosed). By reprocessing these residues, one primarily strives for the recovery of the precious metals, but secondarily also for the recovery and re-use of the A1203 substrates. So far pyrometallurgical processes have been proposed for recovering the precious metals (after crushing the starting material) [ 1 ]. However, these processes that involve extraction and concentration of the precious metals by melting with lead generally require a complicated lead blast furnace operation when reprocessing materials high in A1203. Without exception they result in high consumption of energy and time, causing marked losses due to the great number of process steps to be performed before the precious metals can be obtained. In addition, it is in general impossible to re-use the ceramic substrates. The low precious-metal contents of the starting material represents a further problem. For these reasons we were induced to recover the valuable constituents by hydrometallurgical procedures. REALIZATION AND RESULTS OF THE EXPERIMENTS
Experim en tal
The investigations were carried out on typical waste products from all process stages of the thick-layer technique as applied in the electronic industry. The ceramic A1203 substrates (dimension 30 m m × 15 m m and 50 m m X 30 m m ) had been printed over with Ag/Pd-, Au- and Ru-bearing conductor or resistor strips as well as with multilayers and covering layers on the base of glass, the precious-metal content on the substrates having been in the order of some milligrams. Several typical printed substrates are shown in Fig. 1. Prior to the treatment metallic connections were removed. Capsules were broken in order that the leaching solution might have access. (However, a satisfactory solution to the problem has not yet been found. )
RECOVERY OF P R E C I O U S METALS F R O M ELECTRONIC SCRAP
I 01
Fig. 1. Printed ceramic substrate.
Paste wastes containing 10-20 wt.% of binding agents were first subjected to a thermal treatment at about 400 ° C in flowing air in order to remove these constituents. The leaching experiments were carried out on the laboratory scale in a rotary apparatus (about 250 substrates in 200-ml solution) and on the pilotplant scale in a commercial 10-1 plating barrel (about 4 kg substrates in 1.5 1 solution). The process solutions were reprocessed on both laboratory and pilot-plant scale; for pilot-plant reprocessing 50-100 1 solutions were used. The studies on the purification of ruthenium were performed with various leaching residues as well as by model experiments with ruthenium metal (labelled Ru- 106 ). Qualitative analysis of the solids was performed by emission spectral analysis and quantitative analysis by mass spectrometry (solid-state mass spectrograph type MS 7, AEJ, UK). In addition, the O, Si and A1 contents were determined by neutron activation with the neutron generator type AN 3 (KFKI Budapest). The contents in the solutions were usually determined by AAS (AAS photometer type 300 S, Perkin Elmer) in accordance with instructions for analysis established at the Research Institute of Non-Ferrous Metals. The Ru-content of the solution was determined by a photometric method with dithio-oxamide (after it had been transferred into the chloride solution). The composition of the Ru-complexes was found from electron spectra
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(Specord UV/VIS, VEB Carl Zeig Jena). The oxidation degree of ruthenium was determined by a coulometric method as described in [ 2 ]. When radio-actively labelled solutions were used, the gamma radiation was measured by means of the well-type scintillation probe type VA-S-968 coupled with the measuring unit VA 20026 (VEB Megelektronik Dresden).
Results of the leaching experiments Leaching of the ignited paste residues Ag- and Pd-bearing products could be leached without difficulties with 5 M to 9 M nitric acid, whereas Au-bearing products were dissolved by aqua regia or hydrochloric acid/chlorine mixtures. In both cases any residue occasionally obtained was free from precious metals. Ru-rich residues, however, were not capable of being digested by hydrometallurgical processes, but only by fusion (see below).
Leaching of the ceramic substrates The experiments have shown that the success of leaching was strongly dependent upon how the leaching agent had access to the layers containing the precious metals. Uncoated layers of Ag/Pd or Au were quickly dissolved by 7 M to 9 M nitric acid or aqua regia even at room temperature. (For the dissolution of gold layers, mixtures of hydrochloric acid and chlorine obtained, e.g., by adding K]VInO4to 3-6 M hydrochloric acid, have proved also useful.) At a solid/solution ratio of about 2 : 1 with agitation the leach was completed within 1 to 2 h, which was suggested by the observation that the strips containing the precious metal had disappeared. (Leaching without agitation required considerably longer duration due to the fact that the parts are covering each other. ) Glass-containing layers on and between the conductor and resistor strips largely inhibited the dissolution of the precious metals. Therefore, they had to be removed before the leaching of the precious metals. This was accomplished by the use of diluted > 0.5 M hydrofluoric acid or 10 M t o 12 M N H 4 F or NH4HF2 solution. When the layer structure was complex, the leach had to be performed in several stages alternately by the use of acid and fluoride solutions in order that all the layers containing precious metals might be stripped. The Ru-containing layers were not attacked by the solutions mentioned above, rather they were stripped as a whole by the fluoride-containing solutions which attacked and dissolved the glass-containing covering and adhesive layers. This process was considerably accelerated by continuous agitation of the parts. After the leach had been finished, the ruthenium was found exclusively in a solid residue together with abraded A1203 and glass components. The practice of multi-stage leach can be seen from the following pilot-plant scale experiment.
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About 4 kg of ceramic substrates were leached in a 100-1 plating barrel with 1.5 1 of 9 M HNO3 for 1.5-2 h at room temperature with a rotation velocity of 20 rev/min. The nitric acid containing Ag and Pd was then decanted and used to leach fresh material. After an intermediate wash with water the ceramic substrates were treated with 1.5 1 of 10 A/NH4F solution for 18-20 h at room temperature. The Ru-containing black deposit was rinsed and the ceramic substrates, after having been thoroughly washed with water, were subjected to visual inspection. When new Ag/Pd layers were observed, the leach with HNO3 was repeated. When Au-containing layers appeared, the leach was performed with 1.51 of 6 M hydrochloric acid with the addition of a small amount of K]VInO4 ( for 1.5-2 h ). The stripping by steps of the layers containing the precious metals is shown in Fig. 2. Table 1 shows the composition of the fractions obtained after leaching several charges of ceramic substrates. In any case, the fluoride solutions were virtually free from precious metals.
Fig. 2. Ceramic substrates after various leaching stages. TABLE 1 Composition of leach liquor and residue Fraction
Composition
HNO3 solution (g/l) HC1/CI2 solution (g/l) Ru-bearing residue (wt.%)
5-15 Pd, 20-30 Ag, -<0.005 Ru, 1-5 Cu, Zn, Pb, Bi 1-5 Au, <0.02 Pd, <0.01 Ag, <0.001 Ru 5-7 Ru, 0.2-1 Ag, -<0.002 Pd, 7-10 Bi, Pb, remainder A1203
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Experiments on further processing of the leached products (a) Ceramic substrates. After having been thoroughly washed in water and dried at about 200°C, the leached ceramic substrates were re-used to be coated by the thick-layer technique. It became evident that the success of re-use was primarily dependent upon the conditions of the preceding leach with fluoride solution. It was vital to regulate the activity of fluoride in such a way that the glass-containing covering and intermediate layers were stripped completely without attack on the substrate. This can be ensured by leaching with NHaF or NH4HF2 solutions [3]. However, the conditions of oxidizing leach have no effect, provided the Ag, Au and Pd layers are quantitatively stripped. Such recovered ceramic substrates have been used in industry for several years without disturbance in production. (b) Recovery of precious metals. As can be seen from Table 1, the leach resulted in a preliminary separation of fractions rich in Ag/Pd, Au and Ru. From these the precious metals had to be transferred into re-usable compounds. From the gold-containing solutions the gold was precipitated by reduction with SO2 (other appropriate reducing agents are Fe (II) SO4 and formic acid ). The precipitation was quick and complete; thus, the solutions could be discarded without further treatment. Ag and Pd were precipitated together from the nitric-acid solution by a mixture of KSCN and K4Fe (CN) 6. As described earlier [4,5 ], a high degree of separation of the precious metals was achieved by a simple process (final contents in mg/l: Ag 1-2; Pd 4-6). Excessive precipitating agent is precipitated by the present non-ferrous metals, since these also form slightly soluble compounds. Therefore, the nitric-acid solution can be re-used for the leach. From the concentrates obtained the refined metals can be recovered by conventional procedures. For that it is expedient to transfer the concentrates into a separation plant. No applicable procedures were available, however, for reprocessing the Rucontaining solid products. The following indications could be found in the literature. The digestion can be accomplished only by fusion [6] as follows: (a) fusion wit CI2/NaC1 resulting in Ru ( I I I ) / R u (IV) chlorocomplexes [ 7 ]; (b) fusion with alkalies in the presence of oxidizing agents (KNO3, Na202 ) [ 810 ]; (c) fusion with alkalies after the Ru has been converted into RuO2 [ 11 ]. In both cases the water-soluble ruthenates RuO~-2 are obtained. R u t h e n i u m is refined almost exclusively by selective formation and separation of the volatile RuO4. To convert most of the existent Ru-compounds into RuO4, and, in general, to handle this unstable toxic c o m p o u n d is problematical. The existing conditions (type and a m o u n t of other compounds, complex form of the Ru) are of great importance so that several methods have been worked out [8,12,13]. Oxidation with C12 in alkaline solution
RECOVERYOFPRECIOUSMETALSFROMELECTRONICSCRAP
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(which by a special procedure also in the presence of chloride allows RuO4 to be formed in large amounts [ 14 ] and with NaBrO3, KIO4, K2S2Os and others in sulfuric-acid solution is preferentially used. The RuO4 separated by distillation, volatilization in the gas stream or by liquid-liquid extraction with CC14 [ 15 ] is in general trapped as Ru (IV) chlorocomplex in mixtures of 6 M HC1 and reducing agents (SO2, oxalate, H202). From these processes elemental Ru is recovered mostly by the precipitation of hydrated ruthenium oxide RuO2" H20 followed by the reduction in a hydrogen stream. Moreover, similar to the purification of Ir and Pt, the recovery of Ru by precipitation and ignition of (NH4)2RuC16 has been described [9,16]. However, the data on generation and properties of this c o m p o u n d are inconsistent [8,12,16].
Digestion of the Ru-containing material Fusion with C12/NaC1 yielded unsatisfactory results, presumably, because of the high Al203 contents of the starting material. However, fusion with a mixture consisting of 1 part of Na202 and 2-3 parts of KOH (2 h, 500 ° C ) as well as with KOH (2 h, 700°C) was carried out with good success with all starting products, amongst them also metallic Ru. Because fusion with KOH had a considerably less corrosive effect and, thus, crucibles from Ni and Fe could be used, in the following investigations fusion was performed always with KOH. (We were surprised to find that, when using NaOH, worse results were obtained.) The melt was easily dissolved in water at 40-50°C. After filtration through a glass filter (no paper filter! ) a reddish-brown clear solution was obtained. The residue after fusion amounted to 5 + 3% of the charge weight ( n = 10 experiments ), after it had been treated with 1 M to 2 M H2SO4 to eliminate contaminations by crucible material. This Ru-containing fraction was fed to the next fusion. The leach liquor contained ruthenium as K2RuO4 and varying amounts of several accompanying elements (A1, Bi, Pd, Ag). In the absence of reducing agents this solution was stable over a long period.
Purification of ruthenium Two procedures have been tested. ( 1 ) Separation as RuO4. Being observant of the usual safety instructions, we have generated the RuO4 in a glass apparatus by introducing chlorine into the alkaline ruthenate solution and trapped it in three receiving flasks with 6 M HCI. The temperature rose to about 70 ° C during the introduction of C12, dropped to 35 °C after the main reaction had quieted down and was then again raised to 60-70 °C by heating. The reaction was finished when the solution in the distillation flask was almost colourless (final Ru content l - 1 0 p p m relative to the evaporated residue). (2) Precipitation of RuO2"xH20, dissolution in HC1 and purification with
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ion exchangers. During the reaction of the alkaline ruthenate solution with weak reducing agents (ethanol, formic acid, oxalic acid) as well as with strong ones (hypophosphite, hydrazine hydrate, metallic A1) in any case hydrated Ru (OV) oxide was precipitated, but never metallic Ru. The precipitation was complete (final contents in the solution < 1 mg Ru per litre); filtration of the finely divided precipitate was relatively easy (conditions of precipitation: temperature 60-70 ° C, excess reducing agent). The precipitate could be washed without difficulty with hot water or 0.5 M to 1 M NaOH. To our surprise, this was also possible with strong acids such as 0.5 M t o 1 M H C 1 0 4 o r H 2 S O 4 which do not form complexes with Ru. The fresh precipitate could be easily dissolved in hot 6 M hydrochloric acid. In the solution obtained Ru was found to be present as a mixture of various chloro- and hydroxocomplexes of Ru (IV). By this precipitation and washing of the precipitate with acids and bases the concentration of a n u m b e r of accompanying elements such as K +, Na +, A13+, Fe 3+, Ni 2÷ was largely decreased. (Thus, in the solutions obtained the total concentration o f N a +, K +, Fe 3+ and Ni 2+ was in the range of 0.03-0.01 eq./1. ) When the precipitation was performed with strong reducing agents such as hydrazine hydrate, in addition, the present precious metals were separated because these were reduced into the metallic state and thus remained undissolved after the treatment of the precipitate with hydrochloric acid. The data are given in Table 2. Further purification was accomplished with the cation exchangers Wofatite KPS or KS 11. Due to the low selectivity c o m m o n to all ion exchangers of this type it was necessary to eliminate the excessive hydrochloric acid by evaporation. Because of the great tendency of the Ru to hydrolyze, which at an acidity of the solution < 0.1 M resulted in deposition of hydrolysis products on the ion exchanger, the acidity of the feed solution was set to 0.3-0.5 M. Under these conditions interfering cations were separated to a high degree. Thus, the following concentrations of typical impurities were measured in the purified solutions (in % relative to the Ru content): Mg 0.0027; Ca TABLE 2 Depletion of Ag and Pd concentration by precipitation and dissolution of RuOz.xH20 Fraction
Contents (rag/l) Ru
Feed solution Alkaline RuO 2- solution Final solution Ru02 dissolved in HC1
780 2070
Ag
9.0 < 1
Pd
7.8 < 1
RECOVERYOFPRECIOUSMETALSFROMELECTRONICSCRAP
107
0.018; Pb < 0.04; Ag 0.0045. Ru-adsorption by the cation exchanger was low, being in the range of <0.3 g/1 of resin.
Recovery of Ru (a) Precipitation of (NH4)2RuC16. Black ( N H 4 ) 2 R u C 1 6 w a s precipitated from Ru-containing solutions with NH4C1 in the presence of an oxidizing agent (H202 or NaCIO3) after the concentration of HC1 had been set to 6-7 M. At elevated temperatures the precipitation process was accelerated. The precipitation rate of Ru was strongly dependent upon the working conditions and varied within the range of 70-90%. This unsatisfying precipitation rate resulted primarily from the fact that in the feed solution besides RuC126- Ru complexes are also present which cannot be precipitated with NH~- ions. These complexes were identified in most cases as [ Ru2OCli0 ] 4-. (b) Precipitation of RuO2-xH20 and reduction in flowing H2. In accordance with the description by Brandstetr [ 17] brown-black hydrated ruthen i u m oxide was precipitated from the Ru solution with (NH4)2CO3 solution at boiling point. At pH ~ 6 the precipitation was complete (CRu < 1 m g / l ) . The precipitate was easy to filter. After washing with hot water the precipitate was dried, ignited at 600°C and reduced at 600-700°C in flowing H2. Table 3 shows the purity of ruthenium obtained by various procedures as determined by mass spectrometric analysis. The following purification procedures were used. Sample 1: distillation of RuO4. Samples 2-4: precipitation of RuOz-xH20 and purification by ion exchangers. Final precipitation: samples 1-3 hydrated oxide; sample 4 (NH4)2RuC16. As can be seen, pure Ru can be recovered by both purification procedures. The results of sample 3 may indicate some lack of reproducibility, but the analysis of the Ru solution by other procedures during purification always revealed sufficient separation. It has yet to be clarified by additional work which of the analytical results prove correct. TABLE 3 Composition of recovered Ru Sample No.
l 2 3 4
Impurity contents (wt.%) Fe
Pb
Bi
A1
Na
Si
O
0.016-0.05 0.033 0.047 0.008
0.04-0.06 0.014 0.3(?) 0.09
0.03 0.007 0.2 0.03
n.d. 0.014 0.16 0.04
0.01 0.027 0.076 0.009
n.d. 0.006 0.023 0.03
n.d. 0.6 1.4 0.9
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K. GLOE EY AL.
DISCUSSION
The results of these investigations have shown that the valuable constituents of the waste products of the thick-layer technique can be recovered by hydrometallurgical procedures. The flow sheets of the processes of reprocessing the ceramic substrates and purification of ruthenium are shown in Figs. 3 and 4. The paste residues are the starting material of HNO3 or HC1/C12 leach; in this case the leach with fluoride is omitted. The Ru-containing residues are directly reprocessed according to Fig. 4. The procedure proposed is distinctive in that the precious metals can be
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RECOVERY OF PRECIOUS METALS FROM ELECTRONIC SCRAP
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recovered and the ceramic substrates can be prepared for re-use by the use of cheap chemicals and simple equipment. Because all process solutions that are discarded containing only minimal amounts of precious metals (which, in addition, can be further decreased by known procedures such as precipitation with base metals) the losses during reprocessing are very low. Thus, the hydrometallurgical recovery is distinctly superior to thermal reprocessing because the latter results in high losses of precious metals and the ceramic substrafes are entirely lost. By this procedure Ag, Au and Pd are precipitated from the process solutions as concentrates. Beyond that there is still the possibility of separating
1 10
K. GLOE ET AL.
a n d s i m u l t a n e o u s l y r e f i n i n g the p r e c i o u s m e t a l s b y l i q u i d - l i q u i d e x t r a c t i o n . Similarly to the p r o c e d u r e s o f p r e c i o u s - m e t a l s r e c o v e r y f r o m p r i m a r y raw m a t e r i a l s [ 18 ] Au c a n be e x t r a c t e d with solvating solvents (e.g., d i b u t y l carb i t o l ) a n d Ag a n d P d with alkyl s u l p h i d e s or alkyl sulphoxides. P r e l i m i n a r y results o f i n v e s t i g a t i o n s c a r r i e d o u t o n leach liquors o b t a i n e d as a result o f this w o r k h a v e p r o v e d these e x t r a c t i o n p r o c e d u r e s to be applicable [ 19 ]. T h i s allows s u b s e q u e n t p u r i f i c a t i o n e i t h e r to be e n t i r e l y o m i t t e d or to be m a r k e d l y facilitated. T h e new p r o c e d u r e o f R u p u r i f i c a t i o n d e v e l o p e d in this w o r k is distinguished by the fact t h a t f r o m v e r y c o m p l e x starting m a t e r i a l the R u is rec o v e r e d with high p u r i t y a n d yield with the c o m p l i c a t e d RuO4 distillation b e i n g o m i t t e d . T h i s p r o c e d u r e is n o t c o n f i n e d to the starting m a t e r i a l u s e d in this work. It can also be a p p l i e d to r e c o v e r r u t h e n i u m effectively f r o m o t h e r solid m a t e r i a l s [ 20 ].
REFERENCES 1 Day, J.G. and Green, H., 1980. Method for the Recovery of Precious Metals. DE 3047, 194. 2 Weltrick, G., Phillips, G. and Milnev, G.W.C., 1969. Analyst, 96: 840-843. 3 Oppermann, H. Reichelt, W., Mfihl, P. and Grimm, G., 1982. Method for the Recovery of Feedstock from Electronics Scrap. DD208 480. 4 Knothe, M., Pfrepper, G. and Pfrepper, R., 1989. Hydrometallurgy, 21: 293-304. 5 Pfrepper, G., Pfrepper, R. and Knothe, M., 1985. Method for Separating Ag and Pd from Waste Solutions. DD 229 393. 6 Rard, A.J., 1985. Chem. Rev., 85: 1-39. 7 Mit'kin, V.N., Nikonorov, Ju.J. and Zemskov, S.V., 1977. Izv. Sib. Otd. Akad. Nauk SSSR, Ser. Khim. Nauk, 5: 92-100. 8 Gmelin, 1938. Handbuch der anorganischen Chemic. System Nr. 63, Ru Berlin. 9 Clements, F.S., 1962. Ind. Chem., 38: 345-354. 10 Japan Carbit Co., 1978. Method ofRu-Recovery. GB 1 527 758. 11 Pittie, W.H. and Overbeek, G., 1975. Separation of Ru and Os from Ir and Rh. US Patent 3 997 337. 12 Brauer, G., 1981. Handbuch Pr~ip. anorg. Chemic. Stuttgart, Bd. 3, p. 1749. 13 Gorski, B. and F~rsterling, H.U., 1984. Isotopenpraxis, 20: 201-204. 14 Knothe, M., 1979. Method for Ru-Extraction. DD 135 214. 15 Khan, A. and Morris, D.F.C., 1967. Sep. Sci., 2:201-204 (l.c.: 635-644). 16 Gilchrist, R., 1943. Res. Nat. Bur. Stand., 30: 279-372. 17 Brandstetr, J. and Vrestal, J., 1962. Collect. Czech. Commun., 27:1798-1810. 18 Barnes, J.E. and Edwards, J.D., 1982. Chem. Ind., 6: 151-155. 19 Gloe, K. and Mfihl, P., 1986. Proc. XXV Syrup. Bergm~nnisches Pribram in Wissenschaft und Technik, Pribam, Sekt. N, p. 390. 20 Knothe, M. and K~hne, M., 1987. Method for Ru-Extraction. DD 261 811.
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Elsevier Science Publishers B.V., Amsterdam - - Printed in The Netherlands
Recovery of precious metals from electronic scrap, in particular from waste products of the thick-layer technique K. Gloe ~, P. Mfihl ~ and M. Knothe 2 Central Institute of Solid-State Physics and Materials Research, Dresden (German Democratic Republic) 2VEB Mansfeld Kombinat "Wilhelm Pieck", Research Institute of Non-Ferrous Metals, Freiberg (German Democratic Republic) (Received July 3, 1989; revised and accepted November 13, 1989 )
ABSTRACT Gloe, K., Mfihl, P. and Knothe, M., 1990. Recovery of precious metals from electronic scrap, in particular from waste products of the thick-layer technique. Hydrometallurgy, 25: 99-110. From electronic scrap, in particular from the waste products of the thickqayer technique (which is applied to produce passive electronic components) the Ag, Au, Ru and Pd contents are recovered by hydrometallurgical procedures. Ag, Pd and Au are dissolved at 20°C in 5 M to 9 M HNO3 or HCI/ C12 mixtures, being then precipitated from the leach liquors as concentrates. Coatings on the base of glass inhibiting the leaching of precious metals are removed by leaching with fluoride solutions, which results also in the stripping of Ru-bearing layers from the ceramic substrate. From the solid Ru-concentrates obtained, ruthenium is recovered and purified by alkaline fusion, precipitation and dissolution of RuOz-xH20 followed by cation exchange. After the precious-metal and glass interlayers have been stripped, the ceramic substrates can be re-used to produce passive components.
INTRODUCTION
Precious metals such as Au, Ag, Pt, Pd and Ru, due to their specific physical
and chemical properties, are essential constituents of electronic components. In the future, development will give rise to an increase in consumption of these metals. This inevitably results in an increased amount of waste products, residues and scrap containing precious metals. Thus, the recovery of the valuable constituents from these waste products is an important and beneficial task. In this paper possibilities of reprocessing waste products of the so-called thick-layer technique will be discussed. This technique developed to produce passive electronic components is now being applied to manufacturing hybrid circuits. To accomplish this, a paste consisting either of conducting metal 0304-386X/90/$03.50
© 1990 Elsevier Science Publishers B.V.
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K. G L O E ET AL.
powder (e.g., Ag, Au, Pd, Pt) or non-conducting oxide powder (RuO2, [Bi,Pb]Ru2OT), inorganic binder (fritted glass for adhesion) and organic binding agents for providing the physical properties necessary to the silk-screen process is printed onto a substrate (in general, A1203 ceramics). After removal of the organic constituents subsequent burning results in sintering of the metallic or oxidic constituents, thus providing the desired electrical properties. Due to the complicated manufacturing process and the necessity high precision of the components, even during manufacturing, waste products and residues are obtained. The waste products can be divided into two categories: (1) wastes of paste (incorrect charges and residues) which amongst other components contain the organic binding agents; (2) printed substrates in various processing states (unburned and burned, in part, with connection contacts and more or less enclosed). By reprocessing these residues, one primarily strives for the recovery of the precious metals, but secondarily also for the recovery and re-use of the A1203 substrates. So far pyrometallurgical processes have been proposed for recovering the precious metals (after crushing the starting material) [ 1 ]. However, these processes that involve extraction and concentration of the precious metals by melting with lead generally require a complicated lead blast furnace operation when reprocessing materials high in A1203. Without exception they result in high consumption of energy and time, causing marked losses due to the great number of process steps to be performed before the precious metals can be obtained. In addition, it is in general impossible to re-use the ceramic substrates. The low precious-metal contents of the starting material represents a further problem. For these reasons we were induced to recover the valuable constituents by hydrometallurgical procedures. REALIZATION AND RESULTS OF THE EXPERIMENTS
Experim en tal
The investigations were carried out on typical waste products from all process stages of the thick-layer technique as applied in the electronic industry. The ceramic A1203 substrates (dimension 30 m m × 15 m m and 50 m m X 30 m m ) had been printed over with Ag/Pd-, Au- and Ru-bearing conductor or resistor strips as well as with multilayers and covering layers on the base of glass, the precious-metal content on the substrates having been in the order of some milligrams. Several typical printed substrates are shown in Fig. 1. Prior to the treatment metallic connections were removed. Capsules were broken in order that the leaching solution might have access. (However, a satisfactory solution to the problem has not yet been found. )
RECOVERY OF P R E C I O U S METALS F R O M ELECTRONIC SCRAP
I 01
Fig. 1. Printed ceramic substrate.
Paste wastes containing 10-20 wt.% of binding agents were first subjected to a thermal treatment at about 400 ° C in flowing air in order to remove these constituents. The leaching experiments were carried out on the laboratory scale in a rotary apparatus (about 250 substrates in 200-ml solution) and on the pilotplant scale in a commercial 10-1 plating barrel (about 4 kg substrates in 1.5 1 solution). The process solutions were reprocessed on both laboratory and pilot-plant scale; for pilot-plant reprocessing 50-100 1 solutions were used. The studies on the purification of ruthenium were performed with various leaching residues as well as by model experiments with ruthenium metal (labelled Ru- 106 ). Qualitative analysis of the solids was performed by emission spectral analysis and quantitative analysis by mass spectrometry (solid-state mass spectrograph type MS 7, AEJ, UK). In addition, the O, Si and A1 contents were determined by neutron activation with the neutron generator type AN 3 (KFKI Budapest). The contents in the solutions were usually determined by AAS (AAS photometer type 300 S, Perkin Elmer) in accordance with instructions for analysis established at the Research Institute of Non-Ferrous Metals. The Ru-content of the solution was determined by a photometric method with dithio-oxamide (after it had been transferred into the chloride solution). The composition of the Ru-complexes was found from electron spectra
102
K. GLOE ET AL.
(Specord UV/VIS, VEB Carl Zeig Jena). The oxidation degree of ruthenium was determined by a coulometric method as described in [ 2 ]. When radio-actively labelled solutions were used, the gamma radiation was measured by means of the well-type scintillation probe type VA-S-968 coupled with the measuring unit VA 20026 (VEB Megelektronik Dresden).
Results of the leaching experiments Leaching of the ignited paste residues Ag- and Pd-bearing products could be leached without difficulties with 5 M to 9 M nitric acid, whereas Au-bearing products were dissolved by aqua regia or hydrochloric acid/chlorine mixtures. In both cases any residue occasionally obtained was free from precious metals. Ru-rich residues, however, were not capable of being digested by hydrometallurgical processes, but only by fusion (see below).
Leaching of the ceramic substrates The experiments have shown that the success of leaching was strongly dependent upon how the leaching agent had access to the layers containing the precious metals. Uncoated layers of Ag/Pd or Au were quickly dissolved by 7 M to 9 M nitric acid or aqua regia even at room temperature. (For the dissolution of gold layers, mixtures of hydrochloric acid and chlorine obtained, e.g., by adding K]VInO4to 3-6 M hydrochloric acid, have proved also useful.) At a solid/solution ratio of about 2 : 1 with agitation the leach was completed within 1 to 2 h, which was suggested by the observation that the strips containing the precious metal had disappeared. (Leaching without agitation required considerably longer duration due to the fact that the parts are covering each other. ) Glass-containing layers on and between the conductor and resistor strips largely inhibited the dissolution of the precious metals. Therefore, they had to be removed before the leaching of the precious metals. This was accomplished by the use of diluted > 0.5 M hydrofluoric acid or 10 M t o 12 M N H 4 F or NH4HF2 solution. When the layer structure was complex, the leach had to be performed in several stages alternately by the use of acid and fluoride solutions in order that all the layers containing precious metals might be stripped. The Ru-containing layers were not attacked by the solutions mentioned above, rather they were stripped as a whole by the fluoride-containing solutions which attacked and dissolved the glass-containing covering and adhesive layers. This process was considerably accelerated by continuous agitation of the parts. After the leach had been finished, the ruthenium was found exclusively in a solid residue together with abraded A1203 and glass components. The practice of multi-stage leach can be seen from the following pilot-plant scale experiment.
RECOVERYOFPRECIOUSMETALSFROMELECTRONICSCRAP
103
About 4 kg of ceramic substrates were leached in a 100-1 plating barrel with 1.5 1 of 9 M HNO3 for 1.5-2 h at room temperature with a rotation velocity of 20 rev/min. The nitric acid containing Ag and Pd was then decanted and used to leach fresh material. After an intermediate wash with water the ceramic substrates were treated with 1.5 1 of 10 A/NH4F solution for 18-20 h at room temperature. The Ru-containing black deposit was rinsed and the ceramic substrates, after having been thoroughly washed with water, were subjected to visual inspection. When new Ag/Pd layers were observed, the leach with HNO3 was repeated. When Au-containing layers appeared, the leach was performed with 1.51 of 6 M hydrochloric acid with the addition of a small amount of K]VInO4 ( for 1.5-2 h ). The stripping by steps of the layers containing the precious metals is shown in Fig. 2. Table 1 shows the composition of the fractions obtained after leaching several charges of ceramic substrates. In any case, the fluoride solutions were virtually free from precious metals.
Fig. 2. Ceramic substrates after various leaching stages. TABLE 1 Composition of leach liquor and residue Fraction
Composition
HNO3 solution (g/l) HC1/CI2 solution (g/l) Ru-bearing residue (wt.%)
5-15 Pd, 20-30 Ag, -<0.005 Ru, 1-5 Cu, Zn, Pb, Bi 1-5 Au, <0.02 Pd, <0.01 Ag, <0.001 Ru 5-7 Ru, 0.2-1 Ag, -<0.002 Pd, 7-10 Bi, Pb, remainder A1203
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K. GLOE ET AL.
Experiments on further processing of the leached products (a) Ceramic substrates. After having been thoroughly washed in water and dried at about 200°C, the leached ceramic substrates were re-used to be coated by the thick-layer technique. It became evident that the success of re-use was primarily dependent upon the conditions of the preceding leach with fluoride solution. It was vital to regulate the activity of fluoride in such a way that the glass-containing covering and intermediate layers were stripped completely without attack on the substrate. This can be ensured by leaching with NHaF or NH4HF2 solutions [3]. However, the conditions of oxidizing leach have no effect, provided the Ag, Au and Pd layers are quantitatively stripped. Such recovered ceramic substrates have been used in industry for several years without disturbance in production. (b) Recovery of precious metals. As can be seen from Table 1, the leach resulted in a preliminary separation of fractions rich in Ag/Pd, Au and Ru. From these the precious metals had to be transferred into re-usable compounds. From the gold-containing solutions the gold was precipitated by reduction with SO2 (other appropriate reducing agents are Fe (II) SO4 and formic acid ). The precipitation was quick and complete; thus, the solutions could be discarded without further treatment. Ag and Pd were precipitated together from the nitric-acid solution by a mixture of KSCN and K4Fe (CN) 6. As described earlier [4,5 ], a high degree of separation of the precious metals was achieved by a simple process (final contents in mg/l: Ag 1-2; Pd 4-6). Excessive precipitating agent is precipitated by the present non-ferrous metals, since these also form slightly soluble compounds. Therefore, the nitric-acid solution can be re-used for the leach. From the concentrates obtained the refined metals can be recovered by conventional procedures. For that it is expedient to transfer the concentrates into a separation plant. No applicable procedures were available, however, for reprocessing the Rucontaining solid products. The following indications could be found in the literature. The digestion can be accomplished only by fusion [6] as follows: (a) fusion wit CI2/NaC1 resulting in Ru ( I I I ) / R u (IV) chlorocomplexes [ 7 ]; (b) fusion with alkalies in the presence of oxidizing agents (KNO3, Na202 ) [ 810 ]; (c) fusion with alkalies after the Ru has been converted into RuO2 [ 11 ]. In both cases the water-soluble ruthenates RuO~-2 are obtained. R u t h e n i u m is refined almost exclusively by selective formation and separation of the volatile RuO4. To convert most of the existent Ru-compounds into RuO4, and, in general, to handle this unstable toxic c o m p o u n d is problematical. The existing conditions (type and a m o u n t of other compounds, complex form of the Ru) are of great importance so that several methods have been worked out [8,12,13]. Oxidation with C12 in alkaline solution
RECOVERYOFPRECIOUSMETALSFROMELECTRONICSCRAP
105
(which by a special procedure also in the presence of chloride allows RuO4 to be formed in large amounts [ 14 ] and with NaBrO3, KIO4, K2S2Os and others in sulfuric-acid solution is preferentially used. The RuO4 separated by distillation, volatilization in the gas stream or by liquid-liquid extraction with CC14 [ 15 ] is in general trapped as Ru (IV) chlorocomplex in mixtures of 6 M HC1 and reducing agents (SO2, oxalate, H202). From these processes elemental Ru is recovered mostly by the precipitation of hydrated ruthenium oxide RuO2" H20 followed by the reduction in a hydrogen stream. Moreover, similar to the purification of Ir and Pt, the recovery of Ru by precipitation and ignition of (NH4)2RuC16 has been described [9,16]. However, the data on generation and properties of this c o m p o u n d are inconsistent [8,12,16].
Digestion of the Ru-containing material Fusion with C12/NaC1 yielded unsatisfactory results, presumably, because of the high Al203 contents of the starting material. However, fusion with a mixture consisting of 1 part of Na202 and 2-3 parts of KOH (2 h, 500 ° C ) as well as with KOH (2 h, 700°C) was carried out with good success with all starting products, amongst them also metallic Ru. Because fusion with KOH had a considerably less corrosive effect and, thus, crucibles from Ni and Fe could be used, in the following investigations fusion was performed always with KOH. (We were surprised to find that, when using NaOH, worse results were obtained.) The melt was easily dissolved in water at 40-50°C. After filtration through a glass filter (no paper filter! ) a reddish-brown clear solution was obtained. The residue after fusion amounted to 5 + 3% of the charge weight ( n = 10 experiments ), after it had been treated with 1 M to 2 M H2SO4 to eliminate contaminations by crucible material. This Ru-containing fraction was fed to the next fusion. The leach liquor contained ruthenium as K2RuO4 and varying amounts of several accompanying elements (A1, Bi, Pd, Ag). In the absence of reducing agents this solution was stable over a long period.
Purification of ruthenium Two procedures have been tested. ( 1 ) Separation as RuO4. Being observant of the usual safety instructions, we have generated the RuO4 in a glass apparatus by introducing chlorine into the alkaline ruthenate solution and trapped it in three receiving flasks with 6 M HCI. The temperature rose to about 70 ° C during the introduction of C12, dropped to 35 °C after the main reaction had quieted down and was then again raised to 60-70 °C by heating. The reaction was finished when the solution in the distillation flask was almost colourless (final Ru content l - 1 0 p p m relative to the evaporated residue). (2) Precipitation of RuO2"xH20, dissolution in HC1 and purification with
106
K. GLOEET AL.
ion exchangers. During the reaction of the alkaline ruthenate solution with weak reducing agents (ethanol, formic acid, oxalic acid) as well as with strong ones (hypophosphite, hydrazine hydrate, metallic A1) in any case hydrated Ru (OV) oxide was precipitated, but never metallic Ru. The precipitation was complete (final contents in the solution < 1 mg Ru per litre); filtration of the finely divided precipitate was relatively easy (conditions of precipitation: temperature 60-70 ° C, excess reducing agent). The precipitate could be washed without difficulty with hot water or 0.5 M to 1 M NaOH. To our surprise, this was also possible with strong acids such as 0.5 M t o 1 M H C 1 0 4 o r H 2 S O 4 which do not form complexes with Ru. The fresh precipitate could be easily dissolved in hot 6 M hydrochloric acid. In the solution obtained Ru was found to be present as a mixture of various chloro- and hydroxocomplexes of Ru (IV). By this precipitation and washing of the precipitate with acids and bases the concentration of a n u m b e r of accompanying elements such as K +, Na +, A13+, Fe 3+, Ni 2÷ was largely decreased. (Thus, in the solutions obtained the total concentration o f N a +, K +, Fe 3+ and Ni 2+ was in the range of 0.03-0.01 eq./1. ) When the precipitation was performed with strong reducing agents such as hydrazine hydrate, in addition, the present precious metals were separated because these were reduced into the metallic state and thus remained undissolved after the treatment of the precipitate with hydrochloric acid. The data are given in Table 2. Further purification was accomplished with the cation exchangers Wofatite KPS or KS 11. Due to the low selectivity c o m m o n to all ion exchangers of this type it was necessary to eliminate the excessive hydrochloric acid by evaporation. Because of the great tendency of the Ru to hydrolyze, which at an acidity of the solution < 0.1 M resulted in deposition of hydrolysis products on the ion exchanger, the acidity of the feed solution was set to 0.3-0.5 M. Under these conditions interfering cations were separated to a high degree. Thus, the following concentrations of typical impurities were measured in the purified solutions (in % relative to the Ru content): Mg 0.0027; Ca TABLE 2 Depletion of Ag and Pd concentration by precipitation and dissolution of RuOz.xH20 Fraction
Contents (rag/l) Ru
Feed solution Alkaline RuO 2- solution Final solution Ru02 dissolved in HC1
780 2070
Ag
9.0 < 1
Pd
7.8 < 1
RECOVERYOFPRECIOUSMETALSFROMELECTRONICSCRAP
107
0.018; Pb < 0.04; Ag 0.0045. Ru-adsorption by the cation exchanger was low, being in the range of <0.3 g/1 of resin.
Recovery of Ru (a) Precipitation of (NH4)2RuC16. Black ( N H 4 ) 2 R u C 1 6 w a s precipitated from Ru-containing solutions with NH4C1 in the presence of an oxidizing agent (H202 or NaCIO3) after the concentration of HC1 had been set to 6-7 M. At elevated temperatures the precipitation process was accelerated. The precipitation rate of Ru was strongly dependent upon the working conditions and varied within the range of 70-90%. This unsatisfying precipitation rate resulted primarily from the fact that in the feed solution besides RuC126- Ru complexes are also present which cannot be precipitated with NH~- ions. These complexes were identified in most cases as [ Ru2OCli0 ] 4-. (b) Precipitation of RuO2-xH20 and reduction in flowing H2. In accordance with the description by Brandstetr [ 17] brown-black hydrated ruthen i u m oxide was precipitated from the Ru solution with (NH4)2CO3 solution at boiling point. At pH ~ 6 the precipitation was complete (CRu < 1 m g / l ) . The precipitate was easy to filter. After washing with hot water the precipitate was dried, ignited at 600°C and reduced at 600-700°C in flowing H2. Table 3 shows the purity of ruthenium obtained by various procedures as determined by mass spectrometric analysis. The following purification procedures were used. Sample 1: distillation of RuO4. Samples 2-4: precipitation of RuOz-xH20 and purification by ion exchangers. Final precipitation: samples 1-3 hydrated oxide; sample 4 (NH4)2RuC16. As can be seen, pure Ru can be recovered by both purification procedures. The results of sample 3 may indicate some lack of reproducibility, but the analysis of the Ru solution by other procedures during purification always revealed sufficient separation. It has yet to be clarified by additional work which of the analytical results prove correct. TABLE 3 Composition of recovered Ru Sample No.
l 2 3 4
Impurity contents (wt.%) Fe
Pb
Bi
A1
Na
Si
O
0.016-0.05 0.033 0.047 0.008
0.04-0.06 0.014 0.3(?) 0.09
0.03 0.007 0.2 0.03
n.d. 0.014 0.16 0.04
0.01 0.027 0.076 0.009
n.d. 0.006 0.023 0.03
n.d. 0.6 1.4 0.9
108
K. GLOE EY AL.
DISCUSSION
The results of these investigations have shown that the valuable constituents of the waste products of the thick-layer technique can be recovered by hydrometallurgical procedures. The flow sheets of the processes of reprocessing the ceramic substrates and purification of ruthenium are shown in Figs. 3 and 4. The paste residues are the starting material of HNO3 or HC1/C12 leach; in this case the leach with fluoride is omitted. The Ru-containing residues are directly reprocessed according to Fig. 4. The procedure proposed is distinctive in that the precious metals can be
-
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RECOVERY OF PRECIOUS METALS FROM ELECTRONIC SCRAP
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109
(Su)
I /esr,~in~iTi: y ttz: k /ira / i O
:~e: i ~ u ~ (8uJ
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recovered and the ceramic substrates can be prepared for re-use by the use of cheap chemicals and simple equipment. Because all process solutions that are discarded containing only minimal amounts of precious metals (which, in addition, can be further decreased by known procedures such as precipitation with base metals) the losses during reprocessing are very low. Thus, the hydrometallurgical recovery is distinctly superior to thermal reprocessing because the latter results in high losses of precious metals and the ceramic substrafes are entirely lost. By this procedure Ag, Au and Pd are precipitated from the process solutions as concentrates. Beyond that there is still the possibility of separating
1 10
K. GLOE ET AL.
a n d s i m u l t a n e o u s l y r e f i n i n g the p r e c i o u s m e t a l s b y l i q u i d - l i q u i d e x t r a c t i o n . Similarly to the p r o c e d u r e s o f p r e c i o u s - m e t a l s r e c o v e r y f r o m p r i m a r y raw m a t e r i a l s [ 18 ] Au c a n be e x t r a c t e d with solvating solvents (e.g., d i b u t y l carb i t o l ) a n d Ag a n d P d with alkyl s u l p h i d e s or alkyl sulphoxides. P r e l i m i n a r y results o f i n v e s t i g a t i o n s c a r r i e d o u t o n leach liquors o b t a i n e d as a result o f this w o r k h a v e p r o v e d these e x t r a c t i o n p r o c e d u r e s to be applicable [ 19 ]. T h i s allows s u b s e q u e n t p u r i f i c a t i o n e i t h e r to be e n t i r e l y o m i t t e d or to be m a r k e d l y facilitated. T h e new p r o c e d u r e o f R u p u r i f i c a t i o n d e v e l o p e d in this w o r k is distinguished by the fact t h a t f r o m v e r y c o m p l e x starting m a t e r i a l the R u is rec o v e r e d with high p u r i t y a n d yield with the c o m p l i c a t e d RuO4 distillation b e i n g o m i t t e d . T h i s p r o c e d u r e is n o t c o n f i n e d to the starting m a t e r i a l u s e d in this work. It can also be a p p l i e d to r e c o v e r r u t h e n i u m effectively f r o m o t h e r solid m a t e r i a l s [ 20 ].
REFERENCES 1 Day, J.G. and Green, H., 1980. Method for the Recovery of Precious Metals. DE 3047, 194. 2 Weltrick, G., Phillips, G. and Milnev, G.W.C., 1969. Analyst, 96: 840-843. 3 Oppermann, H. Reichelt, W., Mfihl, P. and Grimm, G., 1982. Method for the Recovery of Feedstock from Electronics Scrap. DD208 480. 4 Knothe, M., Pfrepper, G. and Pfrepper, R., 1989. Hydrometallurgy, 21: 293-304. 5 Pfrepper, G., Pfrepper, R. and Knothe, M., 1985. Method for Separating Ag and Pd from Waste Solutions. DD 229 393. 6 Rard, A.J., 1985. Chem. Rev., 85: 1-39. 7 Mit'kin, V.N., Nikonorov, Ju.J. and Zemskov, S.V., 1977. Izv. Sib. Otd. Akad. Nauk SSSR, Ser. Khim. Nauk, 5: 92-100. 8 Gmelin, 1938. Handbuch der anorganischen Chemic. System Nr. 63, Ru Berlin. 9 Clements, F.S., 1962. Ind. Chem., 38: 345-354. 10 Japan Carbit Co., 1978. Method ofRu-Recovery. GB 1 527 758. 11 Pittie, W.H. and Overbeek, G., 1975. Separation of Ru and Os from Ir and Rh. US Patent 3 997 337. 12 Brauer, G., 1981. Handbuch Pr~ip. anorg. Chemic. Stuttgart, Bd. 3, p. 1749. 13 Gorski, B. and F~rsterling, H.U., 1984. Isotopenpraxis, 20: 201-204. 14 Knothe, M., 1979. Method for Ru-Extraction. DD 135 214. 15 Khan, A. and Morris, D.F.C., 1967. Sep. Sci., 2:201-204 (l.c.: 635-644). 16 Gilchrist, R., 1943. Res. Nat. Bur. Stand., 30: 279-372. 17 Brandstetr, J. and Vrestal, J., 1962. Collect. Czech. Commun., 27:1798-1810. 18 Barnes, J.E. and Edwards, J.D., 1982. Chem. Ind., 6: 151-155. 19 Gloe, K. and Mfihl, P., 1986. Proc. XXV Syrup. Bergm~nnisches Pribram in Wissenschaft und Technik, Pribam, Sekt. N, p. 390. 20 Knothe, M. and K~hne, M., 1987. Method for Ru-Extraction. DD 261 811.