Note : Les descriptions sont présentées dans la langue officielle dans laquelle elles ont été soumises.
~2~98~
Backqround of the Invention
The present invention is concerned with
a process for recovering precious metals from
materials containing same. More particularly, the
invention is directed towards the recovery of
platinum and palladium from automobile catalytic
converters
Platinum and palladium are widely used for
catalytic conversion of toxic exhaust gases from
automobiles to non polluting products. Since 1975,
about 140 tons of platinum and 47 tons of palladium
have been consumed annually in the United States and
Canada for this purpose, which represents a major
consumption of these metals. Also in Japan, new
environmental protection rules now oblige most cars
to be equipped with catalytic converters. It is
expected that in some European countries the same
restriction will be introduced. Thus, scrapped
automobile converters are becoming an important poten-
-tial source of recycled platinum and palladium.
In general, two types of catalyst are
used in automotive emission-control equipment:
1. alumina pellets containing from 330 to
500 ppm platinum and 110-200 ppm
palladium, and
2. silicate monolithic honeycombs contain-
ing about 800 to 1,500 ppm platinum and
100 to 350 ppm palladium.
In both cases, small quantities of rhodium and other
precious metals have recently been added. The amount
of catalyst in one converter is from 0.4 to 208 kg.
Several processes for platinum and
palladium recovery from converter beads are already
known, among which the recovery methods involving
precious metals dissolution are best suited for large-
scale treatment of spent autornotive catalyst. In
U.S. Patent No. 3,985,854, for example, precious
metal values are recovered from catalysts by reacting
the catalyst with a mixture of hydrochloric acid and
an oxidizing agent, to leach out the precious metals
in the form of their soluble chloride complexes. The
leach solution must be capable of oxidizing the con-
tained precious metals present in the metallic state
and solubilizing the complexes thus formed. The
metals are then recovered from the solution by one of
several methods of precipitation.
Platinum and palladium dissolution requires
a high concentration of chloride ions in acid solution
and the presence of an oxidizing agent which is
usually nitric acid. The reactions involved are the
following:
3Pt + 4HN03 + 18HCl = 3H2PtCl6 + 4N0 + 8H20 (l)
3Pd + 2HN03 12HCl = 3H2PdC14 + 2N0 + 4H20 (2)
Such conditions are assured by leaching at 70-95C
with aqua regia solution wherein the ratio of acids
is HCl/HN03 = 3/l. Leaching with aqua regia at
temperatures above 70C is extremely difficult
because of the highly aggressive nature of both the
solution and the gaseous products of its decomposi-
tion. This solution instability is marked by a high
89
partial pressure ox gaseous hydrogen chloride andsu~stantial nitric acid decomposition with evolution
of nitrogen oxides. Such loss of oxidizing agent
(HN03) and source of chloride ions (HCl) that are
necessary to keep platinum and palladium ions in
solution slows the rate of metal dissolution result-
ing in an increased leaching time. This, in turn,
promotes undesirable dissolution of the alumina
substrate of the catalyst, with additional acids
consumption. Apart from the fact that HCl regene-
ration is technically difficult from environmental
and corrosion view points, the presence of usually
large amounts of aluminum in solution precludes the
possibility of recycling the solution for additional
leaching since this indicates that the amount of
free HCl present in solution has been severely dimi-
nished by alumina dissolution.
Summary of the Invention
It is therefore an object of this invention
to overcome the above drawbacks and to provide a
process for recovering precious metals from materials
containing same, which enables one to considerably
reduce the gaseous products evolved during leaching
as well as the extent of alumina dissolution, and
also to substantially decrease the acid consumption.
In accordance with the present invention,
there is thus provided in a process for recovering
precious metals from materials containing same by
leaching the precious metal with an acidic aqueous
medium containing chloride ions and a strong oxidiz-
ing agent to form a soluble chloride complex of the
~2~9~39
precious metal and separating the metal from the
complex, the improvement wherein the chloride ions
art derived at least in part from a non-acidic
chloride-containing leaching agent which is non-
volatile and soluble in the acidic aqueous mediumwithout crystallizing under the leaching conditions.
It has been unexpectedly found that by
replacing at least part of the hydrochloric acid
conventionally used for leaching precious metals by
a non-acidic chloride-containing compound which is
non-volatile and soluble in the leach solution without
crystallizing under the leaching conditions, the
partial pressure of hydrogen chloride and of other
gaseous products evolved during leaching can be
considerably reduced and the acid consumption
significantly decreased. Since chloride is the most
effective medium in which precious metals can be
brought into solution, such a non-acidic chloride-
containing leaching agent must have a crystallization
point which is high enough to permit a sufficiently
high concentration of dissolved chloride. Generally,
about 10 to 100% of the hydrochloric acid can be
replaced by this non-acidic chloride-containing leach-
ing agent.
Description of Preferred_Embodiments
Examples of suitable non-acidic chloride-
containing leaching agen-ts for the practice of the
invention include aluminum chloride (AlC13) and
magnesium chloride (MgCl~). It is also possible to
use calcium chloride (CaC12), sodium chloride (~aCl)
and potassium chloride (KCl) to a certain extent as
-- 5 --
I
~228~8~3
the applicability thereof is limited due to their
lower crystallization point. Aluminum ehloride is of
course preferred since it provides 3 moles of chloride
ions per mole of AlC13 used. Where the precious metal
bearing material to be treated is an alumina-based
catalyst, aluminum chloride is also preferably used
since an increased concentration of aluminum chloride
in the leach solution substantially reduces the rate
of dissolution of the alumina substrate of the
catalyst, aluminum chloride also has no leaching
effect on the alumina substrate. For example, it has
been observed that by inereasing the alumina chloride
eoncentration in the leach solution to about 200 g/l,
the acid consumption was deereased from about 9.0 to
6.7 moles of acids per kg of eatalyst pellets and
the extent of alumina dissolution redueed from 20%
to 12% of the initial quantity of alumina.
Generally, the total eoneentration of
ehloride ions present in the leaeh solution is at
least about 2 moles/liter, preferably from about 2 to
about 8 moles/liter. The preferred ehloride eoncen-
tration is about 4.5 moles/liter.
Where aluminum ehloride is used as the
non-acidic chloride-eontaining leaching agent, it
is preferably present in the leach solution in an
amount of about 0.2 to about 2.3 moles/liter. It
should be noted that 2 3 moles/liter AlCl3 is the
highest concentration permissible without AlCl3
undergoing crys-tallization and provides a sufficiently
high concentration of chloride ions so that the
6 --
~L2289~
hydrochloric acid may be completely dispensed with.
On the other hand, where a mixture of HCl, AlC13 and
oxidizing agent is used for leaching the precious
metals, the molar ratio of HCl:AlC13 is preferably
from ako~t 1:5 to about 3:1.
The leach solution is generally acidic
to solubilize readily the chlcride complex of the
precious metal formed. Generally, the concentration
of hydrogen ions is at least 0.5 mole/liter.
l'he leach solution must also contain a
strong oxidizing agent dissolved therein so as to
oxidize the precious metal and render it soluble in the
acid medium. Examples of suitable oxidizing agents
include nitric acid, chlorine, chlorates (such as
sodium or potassium chlorate), brornine, bromates (such
as sodiurn or potassium bromate~, iodine, iodates (such
as sodium or potassium iodates) and hydrogen peroxide.
Ni-tric acid is of course preferred since it also pro-
vides a source of hydrogen ions necessary to maintain
the acidity of the leach solution. The nitric acid
is preferably used in an amount of about 0.6 to about
2.2 moles/liter.
Generally, the leaching is carried out at a
temperature of about 70 to about 105C, preferably
about 90-105C.
The separation of the precious metal from
the chloride complex formed during leaching can be
effected in known manner. Examples of suitable sepa-
ration methods include electrolysis, precipitation
by cementation, sorption with simultaneous reduction,
~2~8g89
reduction under an overpressure of hydrogen and
reduction with hydrazine.
The process of the invention is partieu-
larly applicable to the recovery of platinum and
palladium from alumina-based catalysts containing
same. As mentioned previously, by using aluminum
ehloride, most of the alumina substrate i5 left intact.
Thus, after extraetion of the platinum and palladium,
the catalyst can be further treated with a sulfurie
aeid solution to leaeh out the alumina in the form of
aluminum sulfate hydrate (alum). Up to 80% of the
initial alumina content can therefore be recovered
as a valuable alum by-produet.
A preferred process for reeovering platinum
and palladium from an alumina-based eatalyst eomprises
the steps of:
a) treating the eatalyst at a temperature
of about 90 to about 105C with a leach solution
eontaining aluminum ehloride, hydroehlorie aeid
and nitric acid, in which the molar ra-tio of
HCl:AlC13 is from about 1:5 to about 3:1, to leaeh
out the platinum and palladium in the form of their
soluble chloride eomplexes, thereby providing a leaeh
liquor eontaining these complexes dissolved therein
b) eleetrolytically depositing a major
portion of the platinum and palladium from the leach
liquor obtained in step (a) serving as electrolyte,
while bleeding off a portion of spent electrolyte,
c) washing the leached catalyst from step
(a3 to extraet leaeh liquor entrapped therein and
~22~398~
containing dissolved chloride complexes, thereby
providing a wash liquor containing same,
d) neutralizing and purifying the portion
of spent electrolyte obtained in step (b) and combin-
ing the neutralized and purified electrolyte withthe wash liquor obtained in step (c); and
e) extracting the remaining portion of the
platinum and palladium from the combined electrolyte
and wash liquor obtained in step (d) by sorption of
the chloride complexes contained therein, with
simultaneous reduction to metallic state.
Step (a) can be effected for example by
feeding the leach solution to the inlet o-f a
leaching reactor containing a static bed of the
catalyst in crushed, bead or pellet form, and
allowing the leach solution to intimately contact
the catalyst. The leach liquor containing dissolved
chloride complexes can then be withdrawn from the
outlet of the reactor. The total time of leaching
may vary from about 50 to about l90 minutes.
The electrolysis which takes place in
step ~b) and involves a cathodic deposition of the
pla-tinum and palladium from the leach liquor serving
as electrolyte occurs in two steps:
1) palladium deposition and platinum
reduction from Pt4 to pt2 :
PdCl42 + 2e = Pd + 4Cl , E = 0.59V (3)
PtCl62 + 2e = PtCl42 + 2Cl , E = 0.64V (4)
2) platinum deposition:
30 PtCl42 + 2e = Pt + 4Cl , E = 0.75V (5)
~28~39
Good compact deposition from the leach
liquor onto graphite or titanium cathodes takes
place when the cathodic current density is less than
60 A/m2. For example, after 24 hours of electroly-
sis at 30 A/m2, the platinum concentration decreasesfrom 326 mg/l to 121 mg/l, and palladium from 104
mg/l to 29 mg/l. Total deposition of platinum and
sirnultaneous reduction in palladium concentration to
2.7 mg/l was observed after an additional 42 hours.
About one third of the spent electrolyte
must be bled off due to accumulation of other
metals in solution. The lead chloride concentra-
tion can be reduced by crystallization, while metals
such as iron, manganese, magnesium or aluminum are
partially precipitated as hydroxides during neu-
tralization in step (d). In general, only lead
dissolved from used catalysts accumulates in solu-
tion in the forrn of stable chloride complexes, and
when recirculated solution becomes saturated in
lead, crystallization oE lead chloride (PbC12) is
necessary. Lead chloride may be crystallized from
solutions containing 6 to 30 g/l Pb by simple cool-
ing. Chloride complexes of platinum and palladium
are stable in these solutions, and do not crystal-
lize as chlorides if their concentrations are low -
below a few grams per liter.
Two steps of washing are required for the
total removal of platinum and palladium from the inte-
rior of the leached pellets. During the first wash,
carried out by a solution of concentrated chlorides
-- 10 --
(AlC13 or NaCl), about 90% of platinum and palladium
adsorbed on the inside surface of the pellets is
recovered. A second wash with water recovers the
remaining platinum and palladium, the final wash
water containing less than 3 mg/l platinum and
palladium at a pH above 3.
The wash liquor is combined with the portion
of spent electrolyte which has been neutralized and
purified and the combined solutions are then treated
in the extraction step (e). In step (e), the platinum
and palladium are extracted from solution by sorption
of their chloride complexes, with simultaneous reduc-
tion to the metallic state, on AMBORAME (trade mark)
resin, produced by Rohm and Ices Ltd.
AMBORANE contains the strongly reducing
borane radical. The reduction reactions taking place
are as follows:
4R3N- BH3 + 6PtCl6 + 12H20 = 4R3N~I (Cl )
+ 6Pt + ~H3BO3 + 32Cl -I 20H+ (6)
2R3~ BH3 + 6PdC14 + 6~I20 = 2R3~H (Cl )
+ 6Pd + 2H3B03 + 22Cl + 10H (7)
Complete sorption of platinum from solution
is more difficult than for palladium. The former
requires an excess of AMBORANE relative to the
platinum and palladium content of the solution, and
a lengthy solution/resin contact time.
It is worth mentioning that the selectivity
of reduction described by reactions (6) and (7) is
less when other metal-bearing anionic complexes
(e.g. PbC14 , CuC13 , FeC14 ) are present. Traces
-- 11 --
~2~89~39
of these metals in the final product of sorption and
reduction can then be expected.
Metallic platinum and palladium are produced
by burning of loaned resin. The composition of final
product recovered from catalyst pellet leach solution
is about 72% platinum and 2~% palladium.
The yield of platinum sorption by AMBORANE
from concentrate leach solutions attains 95% and
palladium 89% respectively. Nevertheless, about
17-28 mg/l platinum and 4-12 mg/l palladium remains
in the spent solution after this process. Essentially
complete platinum and palladium precipitation from
solution after sorption is obtained by cementation
with metallic aluminum powder or granules. The
solution after final platinum and palladium precipi-
tation contains less than 3 mg/l Pt or Pd.
l`he process of the invention enables one
to recover more than 97% of the platinum and
palladium from catalyst pellets. Although the
invention is primarily directed to the recovery
of platinum and palladium from catalytic converters,
it is of course applicable to other precious metals
which form soluble chloride complexes, e.g. iridium,
rhodium and goldO Thus, gold may be recovered
according to the process of the invention from
electronic scrap material such as reject components,
circuit board trim, router dust, punchings, trimmings,
etc.
The following non-limiting examples
illustrate the invention.
~L22E~8~3
Examples 1-8
The continuous chloride leaching of a
catalyst pellets static bed in a two-liter
leaching reactor was effected at a temperature
of 90-104C using a leaching solution containing
varying proportions of HCl, AlC13 and H~03. The
composition of the pellets was ~27-441 ppm Pt,
203-211 ppm Pd, 3.1% Pb, 0.58% Mn, 0.55% Fe. The
leaching conditions and results obtained are reported
in Table 1 hereinafter.
Example 1 is given by way of comparison
since it illustrates the conventional leaching method
using only HCl and ~03.
As it is apparent from Table 1, by replacing
at least part of the hydrochloric acid by aluminum
chloride so that at least 10% of the chloride ions are
derived from AlC13, the quan-tity of gaseous products
evolved during leaching is reduced by at least 50%.
Example 8 illustrates the case where no HCl is present
in the leach solution, the chloride ions thus being
derived completely from AlC13.
It should also be noted in connection with
ExampLes 2 and 4 to 8 that the leach solution fed
to the inlet of the reactor contains some Pt and
Pd as use was made of a recycled leach solution.
Examples 1 and 3 were performed using fresh leach
solution.
Examples 9-15
Examples 2 to 8 were repeated, except that
MgC12 was used instead of AlC13. The leaching condi-
- 13 -
~2Z13~8~
tions and results obtained are reported in Table 2
hereinafter.
As it can be seen, the results obtained
with MgC12 are essentially the same as with AlC13.
It should be noted that the concentration of 3.42
moles/liter MgC12 in Example 15 is the highest concen-
tration permissible without MgC12 crystallization.
Examples 16-21
Examples 2 to 7 were repeated, with the
exception that CaC12 was used instead of AlC13 The
leaching conditions and results obtained are reported
in Table 3 hereinafter.
As it can be seen, the results obtained
with CaC12 are essentially the same as with AlC13.
lS It should be noted that the concentration ox 2.19
moles/liter CaC12 in Example 21 is the highest con-
centration permissible without CaC12 crystallization
Examples 22-27
Examples 2 to 7 were repeated, except that
NaCl was used instead of AlC13. The leaching condi
tions and results obtained are reported in Table 4
hereinafter.
As it can be seen, the results obtained
with NaCl are essentially the same as with AlC13.
It should be noted that the concentration of 4.11
moles/liter NaCl in Example 27 is the highest concen-
tration permissible without NaCl crystallization.
14 -
~221 3~3~39
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