Note : Les descriptions sont présentées dans la langue officielle dans laquelle elles ont été soumises.
i6~
BACKGROUND OF THE INVENTION
This invention relates to a process for the recovery of
metal values from complex sulfide concentrates and, more particu-
larly, relates to a process for hydrometallurgically treating
lead sulfide concentrates for the recovery of lead and other non-
ferrous and precious metal values.
Complex sulfide concentrates which may contain lead,
zinc, iron, arsenic, antimony, bismuth, copper, silver, gold and
the like values have historically been treated in pyrometallurgi-
cal processes. Hydrometallurgical processes normally have not
been able to cope with such complex compositions either technical-
ly or economically. However, rapidly increasing metal prices,
higher hygiene standards established for pyrometallurgical
processing, and advances in technology have made hydrometallurgi-
cal processing of complex sulfide-containing concentrates more
attractive.
A number of routes-for the hydrometallurgical processing
of complex lead sulfide-containing concentrates have been
considered, but most have been proved to be either unsuitable for
obtaining economical yields of metal values to be recovered or
were burdened with prohibitively high costs. Such routes include
- sulfuric acid pressure leach systems, followed by amine or ammonia
extract~on, and chloride-based leach systems.
Known prior art on processes using chloride-based
systems for the hydrometallurgical processing of lead sulfide-
containing concentrates usually disclose a leach wherein an
aqueous lixiviant is used which may contain one or moxe compounds
of the group consisting of ferric-, ferrous-, sodium-, magnesium-
and calcium-chloride and which may be acidified with either
hydrochloric or sulfuric acid. The leach usually is performed
hot at atmospheric pressure, lead chloride formed may then be
crystallized and subjected to electrolysis for recovery of lead,
and the lixiviant recovered and/or regenerated and recycled to the
leach.
Typical of these processes are those patented by
Ni~ls C. Christensen during the period of 1920 to 1930. For
example, according to United States Patent No. 1,435,891, which
issued on November 14, 1922, lead-zinc sulfide ore is leached
wikh hot ferric chloride, lead is preferentially dissolved, silver
is precipitated and the solution is electrolyzed or cooled and the
lead chloride crystallized, melted and electrolyzed. The ferric
chloride is regenerated by absorbing chlorine in the residual
ferrous chloride solution. The leach residue is treated with
sulfuric acid. According to United States Patent No. 1,441,063,
which issued January 2, 1923, lead, silver and copper sulfides are
leached with a hot chloride lixiviant which comprises sodium-,
calcium-, magnesium- or ferrous-chloride as well as ferric-chloride
and some hydrochloric acid; silver and copper are cemented from
the leach solution and lead is precipitated by electrolysis from
aqueous solution or from fused lead chloride. The lixiviant does
not act upon pyrite, chalcopyrite and some complex arsenical silver
compounds, but does act upon zinc blende to a limited extent.
More recently, a similar process has been disclosed in
United States Patent No. 3,929,597, which issued o~ December 30,
1975. According to this process, lead and silver are produced
from sulfides containing lead, silver, zinc and iron by leaching
with a ferric salt solution, at 25 - 100C., separating leach
solution from leach residue, leaching the residue with a sodium
chloride brine, at 50 - 100C., cooling the resulting solution to
crystallize and separate lead salts, cementing silver from the
_ ~ _
remaining solution and prod~cing lead by mol-ten lead salt
electrolysis. The ferric salt solution is regenerated by
contacting the leach solution which contains ferrous salt with
chlorine evolved in the electrolysis. A portion of the leach
solution is bled off. The residue from the brine leach is treated
in a sodium sulfide leach resulting in a sulfide bleed stream and
solids, which are treated in a second ferric leach followed by a
brine leach for further dissolution of values. A final residue
from these second leaches is removed from the process.
Prior art processes, including the processes of the
foregoing references, have several limitations. They do not
disclose techniques for the separate and economical recovery of
metal values contained in complex sulfides, for treating process
effluents for the recovery of values in such a manner that
pollution is obviated, or for possible integration in a metallur-
gical plant wherein complex lead containing sulfide concentrates
are treated separately from other concentrates. Moreover, the
prior art does not disclose techniques and conditions for more
selective separation of values in leashing or for careful control
~20 of the water mass-balance in the process to enable economic
operation.
~TATEMENT OF INVENTION
We have now developed a process which substantially
overc-omes the disadvantages of known prior art processes.
In a preferred embodiment of our invention there is
provided a process for the treatmen-t of complex lead sulfide-
containing concentrates additionally containing at least one
metal of the group consisting of iron, copper, zinc, silver,
arsenic, antimony, bismuth and gold which comprises the steps of
selectively leaching a concentrate with iron-containing lixiviant
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for converting lead sulfide in said concentrate to lead chloride
to produce a leach residue and a leach solution; subjecting said
leach residue to a two-stage, countercurrent, hot brine leach to
dissolve lead chloride in a brine-leach solution and to form a
brine-leach residue; subjecting the brine-leach solution to
crystallization by evaporative cooling to lower the temperature
of said solution to the range of from 20 to 30C. for the separate
recovery of crystalliæed lead chloride, residual brine and crystal-
lization condensate, and returning said residual brine to said
lQ brine leach; subjecting crystallized lead chloride in admixture
with about 8~ by we.ight of sodium chloride to electrolysis in a :
fused bath at a temperature in the range of from 410 to 500~C. for
production of lead and evolution of chlorine; sequentially
washing said brine-leach residue in two or more stages with
crystallization condensate to remove lead chloride from said
residue; xeturning condensate from the washing stages to said
crystallization and recovering said washed residue; dividing
: leach:solution into two or more portions; absorbing chlorine in
a first portion of said leach solution for the generation of ferric
chloride-containing solution; and treating a second portion of
~-
said leach solution for the recovery of values.
The iron-containing lixiviant comprises erric chloride
: . in aqueous solution in a concentration in the range of from 100 to
200 g/l ferric ion, and the leaching of concentrate is conducted
at a temperature in the range of from 20 to 60C. and for a period
of time of from 2 to 6 hours, whereby substantially all lead is
converted to lead chloride and major portions of any zinc, iron,
copper, arsenic, antimony, silver and gold remain in the leach
residue.
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.
Alternatively, th~ process of the invention contemplates
the use of aqueous solutions of ferrous chloride and hydrochloric
acid in concentrations in the range of 25 to 160 g/l ferrous
chloride and of 60 to 120 g/l hydrochloric acid for leaching of
concentrate, conducted at a temperature above at least about 70C.
with evolution of hydrogen sulfide and for a period of time in the
range of 0.5 to 2 hours, whereby substantially all lead is
converted to lead chloride, and gold and major portions of zinc,
iron7 copper, arsenic and antimony remain in the leach residue.
Sulfide ores which may be treated in an integrated
metallurgical plant mav contain such metals as lead, zinc, copper,
iroh, cobalt, nickel, arsenic, antimony, bismuth, indium, tin,,
tellurium and silver which occur in simple or complex sulfides,
and gold. The ores usually are subjected to a preliminary
beneficiation to produce concentrates which can be subsequently
processed for the economic recovery of values by pyro- and/or
hydro- metallurgical techniques. However, total separation of
values is never achieved in the preliminary concentrating treat-
ment, with the result that the subsequent metallurgical processes
~20 produce further concentrates, intermediates and residues which
must be selectivel-y treated to realize full recovery of values.
Concentrates that can be treated according to the process of the
present invention comprise sulfides containing lead, zinc, copper,
iron, arsenic, antimony, bismuth and silver, as well as gold and
small amounts of other metals.
We have found that it is important that t~e leach
according to the process of the invention is carried out selective-
ly for lead. While it is desired, therefore, that substantially
all lead be converted into lead chloride, the leaching of amounts
of zinc, copper, iron, arsenic, bismuth and precious metals is
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determined by the form in which these metals are present in the
concentrate. We have found that by using certain aqueous
lixiviants and selected leaching conditions, lead Sulf ide can
be almost completely converted into lead chloride and separated
from all other metal values, other metals can be preferentially
included in either the leach solution or the leach residue, while
some remaining metals will tend to divide between solution and
residue.
In selecting lixiviant and leaching conditions, certain
other points must be kept in mind. The recovery of metals from
leach solutions in the presence of a large amount of iron, i.e.
iron contained in spent lixiviant, is difficult to accomplish.
Secondly, the conversion of sulfides other than lead sulfide by
chloridizing leach reactions will create an imbalance between the
masses of chlorine consumed and recycled in the process, since
only chlorine from the subsequent electrolysis of lead chloride
becomes available for re-use in the lixiviant. Non-selective
leaching of sulfides thus leads to an imbalance in chlorine and
requires that additional chlorine be supplied to the process.
Thirdly, amounts of iron dissolved into leach solutions must
eventually be removed in a manner that is both economical and
non-polluting~
It is, therefore, an object of the present invention
to provide an improved process for the recovery of lead from
complex metal sulfide concentrates.
It is another object of the present invention to
provide a process for treating complex metal sulfide concentrates
- for the separate recovery of lead and other metal values in forms
which are suitable for further processing of these values.
It is a further object of the present invention to
provide a process for treating lead sulfide-containing
concentrates for the recovery of metallic lead and of concentrates
of other metal values in a form suitable for further processing
and recovery of such values.
It is a still further object of the present invention
to provide a process comprising selective leaching of metal values
from complex metal sulfide concentrates.
It is yet another object of the present invention to
provide a process for the recovery of lead from lead sulfide-
containing concentrates by the careful balancing of the amount
of liquid in the process.
BRIEF DESCRIPTION OF TE~E DRAWINGS
These and other objects and the manner in which they
can be attained will become apparent from the following detailed
description of the process of the invention as illustrated in the
drawings in which:
Figure 1 is a flowsheet of a first embodiment
of the process of our invention; and
~0 Figure 2 is a flowsheet of a second embodiment
of the process of our invention.
DESCRIPTION OF THE PREFERRED EMBODIMENT
With reference to Figure 1, concentrate is fed to
leach 1 wherein the concentrate is reacted with iron-containing
aqueous lixiviant capable of converting lead sulfide to lead
chloride. The concentrates may be fed to leach 1 either as
received from a concentrator or reground to obtain the required
fineness prior to feeding. Particle sizes of 100 mesh or smaller
are satisfactor~.
The aqueous lixiviant comprises ferric chloride which
eEfectively converts lead sulfide into insoluble lead chloride
and elemental sulfur. We have found that the leach is highly
selective when carried out at a temperature in the range of 20 to
60C for 2 to 6 hours using an amount of ferric chloride in the
lixiviant which is sufficient to convert the lead sulfide into
lead chloride and a carefully controlled excess of ferric chloride
which will be consumed in reactions with other compounds in the
feed. The preferred temperature is within the range of from 30 to
50C.
The lixiviant may contain from 100 to 200 g/l ferric
ion, as well as a small amount of hydrochloric acid, such as
for example 1 to 10 g/l, to ensure that no readily hydrolyzable
metals precipitate. After completion of leach 1, the reaction
mixture is subjected to a liquid-solids separation yielding a
leach residue and a leach solution.
~11 liquid-solids separations in the process are carried
out using conventional methods and apparatus well known in the art.
The leach residue, which consists of substantially all
the lead as lead chloride, substantially all the pyrite, gold,
; copper, arsenic and antimony, about 80% of the zinc and pyrrho-
tite, and a portion of the silver and the bismuth, which were
present in the original concentrate, as well as elemental sulfur
and gangue materials, is washed with water and is subjected to
brine leach 2. The lead chloride is extracted in a hot, concen-
trated aqueous brine and is subsequently separated from the other
metal values, sulfur and gangue materials in a brine-leach
solution ~aebrhgVa brine-leach residue.
The aqueous brine may be a concentrated calcium chloride
or sodium chloride solution. High brine concentrations yield
_ ~ _
maximum extraction of lead chloride, but too high concentrations
may cause separatlon of calcium or sodium chloride. We have
found that a substantially saturated sodium chloride brine
containing from 250 to 320 g/l sodium chloride gives the bes-t
results, especially with respect to separation of brine from
brine-leach residue and purification of residual brine. If
desired, the sodium chloride brine may contain a small amount of
calcium chloride to react with any lead sulfate that may be
present.
As the brine leach is performed with a substantially
saturated brine, which extracts lead chloride almost to the
saturation point of lead chloride in the brine, the brine leach
and subsequent liquid-solids separation must be carried out to
meet four objectives: Substantially all lead chloride must be
extracted, the solution must not become saturated in lead
chloride, the brine-leach residue must be washed in such a manner
that no lead chloride precipitates during washing and the use of
e~cessively large quantities of wash liquid must be avoided.
High concentrations of lead chloride in the brine-
leach residue and in the normally associated liquid would result
in the precipitation of lead ahloride upon dilution with wash
liquid. This precipitated lead chloride cannot be removed with
moderate amounts of water wash liquid. The use of excessive
quantities of wash liquid, including wash liquids for solids
obtained in other steps of the brine-leach crystallization
circuit, to be described, would upset the essential water mass-
~- balance and would necessitate costly evaporation of excess water.
We have ound khat by conducting the brine leach in
two stages in countercurrent fashion, the four above-named
objectives can be met. Using a two-stage countercurrent brine
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' '
leach, a solution which is substantially saturated with lead
chloride can be obtained from the first stage, while solution
associated with the solids obtained from the second stage,
consequently, has a low lead chloride content and remains well
below the saturation point of lead chloride. Washing of the
second stage solids can, therefore, be carried out with a small
amount of wash liquid without causing precipitation of lead
chloride.
Leach residue from leach 1 is leached in the first
stage 2_ of the countercurrent brine leach 2 with liquid from
the second stage 2b and residual brine which is recirculated
from crystallization 7, to be described. After liquid-solids
separation of the slurry from first stage 2a, the liquid, which
constitutes the brine-leach solution, is passed directly to
crystallization 7 and the solids are subjected to the second
stage 2b of the countercurrent leach with residual brine. After
li~uid-solids separation of the slurry from second stage 2b, the
liquid is returned to the first stage 2a and the solids, which
constitute the brine-leach residue, are subjected to sequential
washing in two or more stages with hot condensate from crystal-
lization 7. The wash liquid is returned to crystallization 7
and the washed brine-leach residue, substantially free of lead
and chloride, may be passed to roast 3.
The use of hot condensate for washing of the brine-
leach residue facilitates maintaining the water mass-balance by
eliminating the addition of water to the leach-crystallization
circuit. The washing of brine-leach residue for the removal
of lead chloride is essential if the residue is to be sold or
further treated, because contained lead chloride creates problems
in subsequent treatment of the residue and so inhibits its sale-
ability. It also represents loss of lead from the present
process.
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The two-stage, countercurrent brine leach is preferably
carried out at temperatures in the range of 80 to 100C. and is
usually completed in a time in the range of 10 to 30 minutes.
The leach is kept acidified, i.e. at an "apparent" pH of 0.5 or
less, to prevent hydrolysis of bismuth and antimony. The true pH
cannot be directly determined because of interference by high
salt concentrations. The "apparent" pH was read from a meter
standardized against dilute hydrochloric acid solutions of known
concentrations and pH.
The brine-leach residue, after washing, may be sold but
is preferably treated for recovery of values. 5uch treatment may
be accomplished by a number of methods and we prefer to convert
the lead chloride-free, brine-leach residue to calcine by
subjecting the residue to a roast 3. If desired, elemental
sulfur may be removed from the brine leach residue prior to
conversion to calcine.
In roast 3, any sulfides in the brine-leach residue are
converted substantially to oxides and the sulfide sulfur is
~ . .
burned to sulfur dioxide which may be converted into sulfuric
acid.
~ The roast may be advantageously carried out at a
temperature in the range o~ 900 to 1200C. in a suspension
roaster using conventional techniques.
The calcine is fed to sulfur dioxide leach 4 where the
caIcine is decomposed and dissolved in sulfuric acid with the aid
of sul~ur dioxide at elevated temperature and pressure. The
oxides and ferrites contained in the calcine are dissolved as
sulfates. Leach 4 is carried out in an autoclave and the
reaction mixture is maintained at a temperature in the range of
70 to 100C. and under a partial pressure of sulfur dioxide in
the range of 1 to 4 kg/cm2 for a period of about 2 hours.
Final acid concentration preferably is in the range of 10 to
20 g/~-
The reaction mixture is discharged from the autoclave
and treated in hydrogen sulfide precipitation 5 in which the
dissolved metal sulfates of copper, arsenic, antimony, bismuth
and silver form insoluble sulfides while zinc sulfate and ferrous
sulfate remain in solution. The precipitation takes place in one
or more enclosed, agitated vessels. Hydrogen sulfide add~tion is
controlled by monitoring the redox potential of the reaction
mixture. The temperature is maintained in the range of 20 to
100C. and the pressure is maintained at about atmospheric
pressure. After completion of the precipitation, the reaction
mixture is subjected to liquid-solids separation. The liquid
- fraction, which contains mainly zinc sulfate and ferrous sulfate,
may be further treated for the precipitation of iron, for
example as iron oxide or jarosite by known methods, and the
recovery of a zinc sulfate solution, which may be treated to
recover zinc sulfate, or to recover zinc for example by electro-
- 20 lysis. The solids fraction, a small portion of which may be
returned to precipitation 5 to improve crystal growth, is
repulped and subjected to flotation 6.
Flotation 6 is carried out in a known manner for the
; separation of sulfides and gold from silica and gangue materials.
The flotation concentrate and tailings are each subjected to
liquid-solids separation and the liquid fractions are used to
re-pulp the solids fraction obtained from precipitation 5. The
flotation concentrate solids fraction comprises a sulfide
concentrate containing copper, antimony, arsenic, bismuth, silver
and gold. This concentrate may be treated further, together with
'' .
.
a similar concentrate which is obtained from precipitation 12,
to be described, for the separate recovery of its values. The
flotation tailings solids fraction which contains mainly silica
and yangue minerals as well as some metal values may be
discarded or, alternati~ely, fed to a secondary brine leach 17,
to be described.
In crystallization 7, the brine-leach solution
obtained from brine leach 2 is cooled to a temperature in the
range of 20 to 30C. whereby substantially pure lead chloride
crystallizes. The crystallization of lead chloride preferably
is carried out in one or more crystallizers using the evapora-
tive cooling method under reduced pressure whereby lead
chloride crystals and residual brine are removed from the
crystallizer and whereby a condensate is obtained from the
vaporsO This condensate is important in the maintaining of a
water balance in the process. A portion of the condensate is
used in the washing of the brine-leach residue, a second,
.... .
minor portion is used in the washing of the crystallized lead
chloride and a third, minor portion is used in the washing of
brine purification residue.
The crystallized lead chloride is separated from the
residual brine, washed with condensate and subsequen~,ly dried
before being fed to molten salt electrolysis 9. The residual
brine is returned 'to first stage 2a of brine leach 2. To
ensure that a pure lead can be produced, the lead chloride must
be of high purity and it has been found necessary to control
the impurity content of the brine, To exercise this control, a
small portion of the circulating brine is subjected to
purification 8 in which the brine is neutralized by addition of
an alkaline material such as sodium h~droxide or lime to a pH
- 13 -
of from about 7~8 to 10, preferably about 8.5, whereby hydroxides,
hydroxy-chlorides or oxychlorides of such metals as zinc, iron,
copper, bismuth, arsenic, antimony, lead and silver are precipita-
ted. It is necessary before or during neutralization to sparge
an oxidizing gas such as air into the brine to oxidize iron from
the ferrous to the ferric state to provide a filterable precipi-
tate. If so desired, the oxidation may be carried out with
chlorine prior to neutralization. The precipitate is separated,
washed with a minor amount of condensate from crystallization 7
and fed to leach 1. The amount o brine to be treated in puri-
cation 8 is usually about 1 to 5% o the total amount of brine.
The dried, pure lead chloride is fed to electrolytic
cells for molten salt electrolysis 9. The cells contain a fused
mixture consisting preferably of about 92% lead chloride and
about 8% sodium chloride which form a eutectic mixture with a
rnelting point of about 410C. The lead chloride may be fed to
the cells directly or may be melted prior to feeding to the
cells. In the cells lead chloride decomposes into lead and
.
; chlorine. Molten lead is removed from the cells and solidified,
while chlorine is taken from the top of the cells and is fed to
ferric chloride generation 10. The electrolysis is operated at
a temperature in the range of 420 to 500C. It is understood
tha~ other compositions of the fused salt may he used such as
eutectic compositions of lead chloride and one or more salts
chosen from the group of alkali and alkaline-earth metal
chlorides. The operating temperature of the electrolysis depends
o~ the melting temperature of the eutectic composition used. The
current efficiency of the electrolysis is 98% or better. The
purity of the lead recovered from the cells is 99.9~ or better
and chlorine recovery is virtually 100%.
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The leach solution obtained from leach 1 is divided
into two or more portions. In this embodiment the solution is
divided into two portions. The first and major portion is
contacted with chlorine from electrolysis 9 in generation 10,
whereby the ferrous chloride in the solution is oxidized to
ferric chloride. The generation proceeds rapidly at temperatures
in the range of 25C. to the boiling point of the solution and
may be carried out in at least one absorption tower. The
generated ferric chloride-containlng solution is returned as the
iron-containing lixiviant to leach 1.
The second and minor portion of the leach solution,
comprising about 10 to 20~ of the total volume, is treated for
the recovery of values and for the elimination of unwanted
materials from the process in forms which do not create environ-
mental problems. This second and minor portion is first treated
in reduction 11, wherein any ferric chloride in the solution is
reduced to ferrous chloride. The reductan~ may be one of a number
of suitable compounds but the use of lead sulfide-containing
concentrate is preferred. Leach solution and concentrate,
containing an amount of lead sul~ide at least sufficient to
reduce any ferric iron to ferrous iron, are mixed and maintained
~t a temperature in the range o~ ~0 to 80C. for a period in the
range of 15 minutes to one hour. After completion of the
reduction, the reaction mixture is separated into a solids and a
liquid fraction. The former is fed to leach 1 and the latter to
hydrogen sulfide precipitation 12.
In precipitation 12, the solution is treated with
; hydrogen sulfide to precipitate sulfides of such metals as silver,
copper, bismuth, arsenic and antimony. The precipitation is
carried out at about atmospheric pressure in a closed vessel and
i ,
- 15
at a temperature in the range of 25 to 90C., while the addition
of hydrogen sulfide is regulated by monitoring the redox
potential and maintaining the pH at a value of about 0.5 by the
addition of lime, if necessary. The silver, copper, bismuth,
arsenic and antimony contained in the solution are substantially
completely precipitated and the precipitated sulfides are
separated from the liquid. A portion of the sulfides may be
recycled to the precipitation 12 to promote crystal growth and
the remaining portion is recovered and may be combined with the
flotation concentrate solids from flotation 6 and treated for the
separate recovery of values.
The liquid obtained after separation from precipitated
sulfides is treated with additional hydrogen sulfide and with
addition of a neutralizing agent in zinc sulfide precipitation 13.
The zinc in the solution is precipitated as substantially pure
zinc sulfide at about ambient pressure and at a temperature in
the range of 25 to 90C. while controlling the pH of the reaction
mixtùre at a value of about 1.5 by the addition of lime in the
form of a slurry. The zinc sulfide is separated from the liquid
and may be recovered as such and further treated or sold, or may
be fed to roast 3 for subsequent recovery of zinc in the zinc
sulfate-containing solution.
The solution obtained from precipitation 13 now
contains mainly ferrous chloride as well as calcium and magnesium -
chlorides. A portion of this ferrous chloride solution is fed
to oxidation 14 for precipitation and removal of excess iron and
such undesirable metals as accumulate in the process such as
magnesium. In oxidation 14, the solution is treated with ox~gen
in a pic~le liquor furnace at temperatures in the range of 500 to
750C. whereby metal chlorides are precipitated and converted to
:' .
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oxides and hydrogen chloride is e~7olved. The residual solids
which are mainly oxides of iron and magnesium are discarded and
e~olved hydrochloric acid may be absorbed in lixiviant~ The
remaining portion of the ferrous chloride solution is treated
with chlorine in generation 15.
~ eneration 15 is similar to generation 10 in that
ferrous chloride is reacted with chlorine from electrolysis 9 to
form ferric chloride at a temperature in the range of 25C. to
the boiling polnt of the solution. The generated ferric chloride-
containing solution may be combined with generated iron-containing
lixiviant from generation 10 as indicated by the broken line, but
preferably is fed to calcium removal 16. If desired, solution
obtained from precipitation 13 may be fed directly to an oxidation
14, whereby generation 15 is eliminated and wherein ferrous
chloride is oxidized to ferric chloride with simultaneous
precipitation of ferric oxide. This oxidation is described
hereinbelow in detail as step 14 with reference to Figure 2.
Ferric oxide is removed and all or a portion o~ the ferric
chloride solution is fed to calcium removal 16 as indicated by
~he broken line in Fiyure 1.
In calcium removal 16, calcium in the solution is
removed, for example, by addition of a stoichiometric amount of
sulfuric acid, or iron sulfate. After removal of precipitated
calcium sulfate, the solution is fed to leach 1 as iron-
containing lixiviant. In a preferred embodiment, we treat the
solution with lead sulfate which reacts with calcium chloride
; in the solution to form calcium sulfate and lead chloride. The
lead sulfate may be added to removal 16 as such or in the form
of a lead sulfate-containing concentrate or zinc plant leach
residue. Zinc plant leach residue is obtained from hydrometal-
- 17 -
;
lurgical treatment of primary leach residues obtained from roast-
leach or hydrometallurgical processes for the recove~y of zinc.
Such residues contain mainly lead sulfate, silica and gypsum, as
well as silver in elemental or combined form. In reacting lead
sulfate-containing concentrate or zinc plant leach residue with
the calcium chloride and ferric chloride-containing solution, the
lead sulfate in the concentrate or the residue is converted to
insoluble lead chloride according to PbSO4 + CaC12 > Pb~12~ +
CaSO4J and the silver is converted to a soluble silver chloride
complex. Other values also dissolve. The reaction goes to
substantial completion in a tim~ in the range o one to four
hours at a temperature in the range of 50C. to the boiling
point of the solution, preferably in the range of 50 to 70~C., at
atmospheric pressure. The iron in the solution must be present in
the ferric state to ensure that silver sulfide is converted to a
soluble silver chloride complex.
The mixture from calcium removal 16 is sub~ected to
liquid-solids separation and the liquid containing ferric
chloride is returned as lixiviant to leach 1. The solids may be
~ur~her treated for the recovery of lead and other values by, for
example, subjecting the solids to a secondary hot brine leach 17,
which is similar to brine leach 2, to dissolve lead and other
values, and to leave a residue which, after separation from
solution, may be discarded. As discussed above, the solids
contained in the tailings from flotation 6 may also be added to
secondary brine leach 17 for ultimate recovery of any residual
values contained in those solids. The solution containing lead
chloride and other values is fed to a crystallization, not shown,
for recoverv of lead chloride, or may be fed to crystallization 7.
3Q
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The embodiment of the process of the invention
illustrated in Figure 2 is similar to the embodiment illustrated
in Figure 1, the main differences residing in the use of a
different lixiviant in leach 1 and in the regeneration of the
lixiviant. In the embodiment of Figure 2, concentrate as
received from the concentrator, or concentrate reground to the
desired particle sizes of 100 mesh or smaller, is fed to leach 1
wherein it is contacted with aqueous iron-containing lixiviant
capable of converting lead sulfide to lead chloride. The
aqueous lixiviant comprises ferrous chloride and hydrochloric
acid~ The sulfides in the concentrate, upon reacting with the
lixiviant, are converted into chlorides and hydrogen sulfide.
We have found that the leach can be carried out with a selectivity
that is similar to that obtained by leaching with ferric chloride-
containing lixiviant as described above with reference to Figure
1. Lead sulfide in the concentrate is almost quantitatively
converted into lead chloride and hydrogen sulfide. A portion o~
the sulfides o~ silver, zinc and bismuth, and pyrrhotite react
similarly, forming chlorides and hvdrogen sulfide, while pyrite,
gold and sulfides of copper, arsenic and antimony remain mostly
- unreacted.
The lixiviant may contain ferrous chloride in an amount
in the range of 25 to 160 g/l ferrous ion and hydrochloric acid.
With low concentrations of hydrochloric acid in the lixiviant,
the amount of liquid to be treated becomes too large to be
practical, while with high concentrations, the lixiviant cannot
be regenerated to desired high concentrations~ The preferred
amount of hydrochloric acid in the lixiviant is in the range of
60 to 120 g/l. The leach is carried out at an elevated tempera-
ture above at least about 70C., as desired, under atmospheric or
-- 19 --
superatmospheric pressure. The leach preferably is carried out
in the range of from 90C. to the boiling point of the reaction
mi~ture under autogenous pressure, in a closed vessel and using
an amount of lixiviant sufficient to give a low free acid content
in the leach solution without adversely affecting the selectivity
of the leach. ~he leach preferably is carried out countercurrent-
ly in two stages with an amount of lixiviant sufficient to give
10 to 20 g/l free acid in the leach solution. Evolved hydrogen
sulfide is discharged to lixiviant regeneration 18, to be
discussed.
The leaching time is in the range of 0.5 to 2 hours.
Reaction mixture is fed to a liquid-solids separation for
separation into leach solution and leach residue. As ferrous
chloride and hydrogen chloride-containing lixiviant has a higher
activity towards certain iron compounds such as pyrrhotite in
the concentrate than ferric chloride-containing lixiviant, more
iron is leached into the leach solution~ while more sulfur is
removed as hydrogen sulfide. Consequently, the amount of leach
residue is less than that obtained according to the embodiment
~0 of the process illustrated in Figure 1. The larger amount of
ferrous chloride in the leach solution does not create any
problems with respect to the recovery of metal directly removed
from solution in a closed circuit process.
- The leach residue is treated using the same methods,
conditions and steps, i.e. steps 2 through 9, as discussed above
with reference to Figure 1.
The leach solution is divided into two or more portions.
In this embodiment the solution is divided into three portions.
A first portion is fed to generation 10 wherein solution is
reacted with chlorlne from electrolysis g to generate ferric
- 20 -
'.~
chloride from ferrous chloride. This generation 10 is identical
to generation 10 illustrated in Figure 1.
A second portion is treated for the recovery of silver,
arsenic, antimony and bismuth in hydrogen sulfide precipitation
12 and subsequently for the recovery of zinc in zinc sulfide
precipitation 13. Precipitations 12 and 13 are identical to
steps 12 and 13 described with reference to Figure 1. As no
ferric chloride is present in the leach solution, no reduction
step is required. The amount of the second portion of leach
solution depends on the amounts of silver, arsenic, antimony,
bismuth and zinc which are dissolved in leach 1 but is usually in
the order of about 10 to 20% of the total volume. The liquid
resulting from precipitation 13 is fed to oxidation 14.
The third and remaining portion of the leach solution
is fed directly to oxidation 14, together with the liquid
resulting from precipitation 13. It is essential that the amount
of iron in solutions fed to oxidation 14 is three times the
amount of iron which is dlssolved int,o the leach solutlon
obtained from leaching concentrate in leach 1. This requirement
determines the amounts of the three portions of ~he leach
solution ana ensures the mass-balance of iron in the process.
The pertinent reaction is represented by the following equation.
3FeC12 + 4 2 - > 2FeC13 + 12 Fe203
In oxidation 14, leach solution is reacted with oxygen or an
oxygen-bearing gas at elevated temperature and pressure in an
autoclave, whereby ferrous chloride is oxidized to ferric
chloride with simultaneous precipitation of ferric oxide. The
reaction may be carried out continuously at a temperature in the
range of 80 to 165C. under a partial pressure of oxygen in the
range of 100 to 200 psi and a retention time in the range of
- 21 -
lO to 120 minutes. In order to obtain non-hydrated ferric oxide
which can be easily separated from solution, the preferred
temperature range is 135 to 165C. An easily separable ferric
oxide can be obtained with a retention time in the range of 10 to
30 minutes. After completion of the reaction, the reaction
mixture is discharged from the autoclave and sub~ected to a
liquid-solids separation.
The solids fraction is removed from the process and the
liquid fraction is fed to calcium removal 16 which is identical
to removal 16 described with reference to Figure l. The solids
recovered from this step may be further treated in secondary
brine leach 17 as has been described. The liquid from removal 16
is fed to lixiviant regeneration 18. If desired, a portion of
the liquid fraction obtained from oxidation 14 may be directly
fed to regeneration 18, as indicated by the broken line.
In lixiviant regeneration 18, ferric chloride in
solutions obtained from generation 10, oxidation 14 and calcium
; removal 16 is reacted with hydrogen sulfide evolved in leach 1
according to the fo]lowing equation:
2~ 2FeCl3 ~ H2S ~ 2FeC12 + 2HCl + S
The sulfur in hydxogen sulfide is oxidized to elemental sulfur
and ferrous chloride and hydrochloric acid are formed. The
reaction is conducted at elevated temperatures in the range o~
40 to 160C. in one or more closed vessels, such as autoclaves
; or tubular reactors. The reaction mixture is maintained under
autogenous pressure when temperatures above the boiling point of
the reaction mixture are used. The reaction proceeds rapidly at
temperatures in the range of 40C. to the boiling point of the
solution and retention times of about 30 minutes are satisfactory.
- 22 -
~L~3 ~
After completion of the reaction, the elemental sulfur
is separated from the regenerated iron-containing lixiviant. The
separation may be carried out separately from the regeneration 18
in a liquid-solids separation when sulfur is formed below its
melting point. When sulfur is formed above its melting point,
liquid sulfur may be drained directly from the pressure vessel.
The recovered sulfur may be processed into a suitable
form or may be processed to produce sulfuric acidO Aqueous,
iron-containing lixiviant comprising ferrous chloride and hydro-
chloric acid is returned to leach 1.
The following examples illustrate the embodiments of
the process of the present invention.
Example 1
To demonstrate the selective leaching of complex lead
sulfide-containing concentrate, 500 g of concentrate having
particle sizes of 95% minus 325 mesh and assaying 39.20% lead,
~.35% zinc, 14.50% iron (mostly pyrite), 6.95% copper, 0.27%
bismuth ànd 0.21% silver was leached with 2~ o lixiviant
containing 112 g/R iron as ferric-chLoride at various temperatures
for different 1eaching times. Samples of leach solution were
taken at different time intervals and assayed. The distribution
of metals in the leach solution represented as percentages of the
amounts of metals in the original concentrate was calculated from
the assay results. The final residues were analyzed for lead and
the conversion of lead to lead chloride was calculated.
The data obtained are presented in Table I. The data
presented for iron have been corrected for the amount of iron in
the li~iviant. These data show that, by carrying out the leach
with ferric chloride at temperatures in the range of 30 to 50C.
using reaction times of up to 4 hours, substantially all lead is
- 23 -
converted to lead chloride, substantially all copper and iron
remain in the leach residue, between 10 and 20% of the zinc and
about 35% of the silver and about 80~ of the bismuth are
dissolved. Thus, lead can be selectively separated from zinc~
iron and copper and a major portion of the silver.
Table I
_ Distribution in leach solution
calculated as % of metals
Test TempO Time Lead contained in concentrate
No. C Min. Conversion
% Zn FeCu Bi Ag
__ ~5_ , .. , . __-~_............ ,.. . . .. , .,, ~ .~ ... , ~ . ~ . ~
1 90 15 _ 33 1710 71 57
_ 60 3344 80 84
_ liO . 79 5067 92 95
2 70 15 _ 28 <55 65 40
_ 37 <58 69 40
~ 120 99.8 48 <S15 74 40
3 S3 15 _ 10 ~52.0 55 35
_ 20 ~52.9 79 35
120 99.8 25 <53.5 79 35
__ . ~
4 33 15 _ 2.8 C51.4 47 35
_ 5.0 <51.7 59 35
120 _ 7.5 <52.0 71 35
240 99.a 10.7 ~52.3 84 35
Example ?
The leach of the previous example was repeated for a
lead sulfide concentrate in which iron was present as pyrite,
pyrxhotite, chalcopyrite and marmatite and which contained
41.5% lead, 6.3% zinc, 11.7% iron, 4.8% copper, 0.19% bismuth,
0.80% arsenic, 0.58% antimony and 21.1% total sulfur.
Concentrate was leached with a ferric chloride~containing
lixiviant at different temperatures and retention times and
the conversion of lead into lead chloride and the extraction
of other metals into the leach solution determined. The test
- 24 -
results are presented in Table II. The figures for iron in the
Table have been corrected for the amount of iron in the lixiviant.
The results show that leaching below 70C. can be carried out
with substantially complete conversion of lead to lead chloride
and with a selectivity which extracts minor portions of zinc and
iron and very small portions of copper, arsenic and antimony.
Table II
rest Tëmp Time ~ Lead ~ Extraction in Leach Solutlon
No C Min. Conversion _
O Zn Fe Cu BiAs Sb
_ _ .. .
1 30 240 99.0 5 13 ~ 652.5 2.5
2 70 120 99.8 50 19 16 94 8 9
3 gO 120 99.9 ~4 41 60 94 9 10
. ~ _ . .
Example 3
To demonstrate selective leaching with a ferrous
chloride and hydrochloric acid-containing lixiviant, 136 g. of
a lead, zinc and iron sulfide-containing concentrate (95~ minus
325 mesh) assaying 58.9% lead, 5.S~ zinc and 9.7~ iron (mostly
pyrrhotite) was leached with one litre of lixiviant containing
; 20 134 g/~ iron as ferrous chloride and varying amounts of hydro-
chloric acid at various temperatures and leaching times. After
separation of leach solution from leach residue, the leach
solution was assayed for free hydrochloric acid content and
the leach residue was leached with an excess sodium chloride
brine at 90C. for 15 minutes. After separation from brine,
the brine-leach residue was assayed for lead, zinc and iron.
The distribution of the lead, zinc and iron in the brine-leach
residue was calculated as percentages of the amount of these
metals present in the original concentrate. The data obtained
are presented in Table III.
The data show that lead can be substantially completely
converted to lead chloride and separated from the leach residue
by a brine leach, and, that by carrying out the leach at tempera-
tures from 90C.to the boiling point of the solution with
residence times between 0.5 and 2 hours and a residual free-acid
content in the leach solution of 20 to 40 g/~, about 90% of the
zinc can be separated from about 80% of the iron present in the
original concentrate~ The high extractions of lead and iron and
low extraction of zinc are desirable.
Table III
_ ~ ~ Free HCL g/R in Distributlon in brine-leach
TNst Temp. Time lixiviant leach residue calculated as % of
o. C. ~rs. solution metals contained in concentrate
Original ~ Pb Zn Fe
_ ~ .. ... . _. .
1 103 268.5 19 0.1 94 15
2 103 0.586.5 41 0.3 91 20
3 95 286.5 39 0.6 91 19
4 95 2124.5 59<0.1 42 10
5 ~70 2a6 . 5 59O ~ S .. . . . _ _ .
~
The leach of the previous example was repeated for a
leaa sulfide concentrate (95% minus 325 mesh) containing 49.3
lead, 7.6% zinc, 10.3% iron (mainly pyrite), 2% copper, 0.2%
bismuth, 1.7% arsenic, a very small amount of antimony and
gold and 54 ounces per ton of silver. 196 g. concentrate was
leached at 103C. for 0.5 hour with a lixiviant containing
134 g/~ iron as ferrous chloride and 85.5 g/~ hydrochloric
acid. The leach residue was leached with brine and the brine-
leach residue assayed. The distribution of the metals in the
brine leach residue as percentages of the amounts present in
the concentrate was calculated and the figures are given in
Table IV.
- 26 -
Table IV
Metal Distribution in Brine-Leach Residue Calculated
as Percentages of Metals contained in Concentrate
Pb Zn Fe Cu Bi As Sb Ag Au
_ _ _ _
0.1 89 80 99 2 96 98 45 100
. ~
The figures presented in Table IV show that lead can
be selectively separated from iron, z,inc, copper, bismuth,
arsenic, antimony, silver and gold by leaching in a ferrous
chloride and hydrochloric acid lixiviant followed by a brine
leach. The figures further.show that substantially all copper,
arsenic, antimony and gold are separated in the brine-leach
residue together with major portions of the zinc and iron,
while substantially all bismuth, minor portions of zinc and iron
and about half of the silver are extracted in the leach solution.
It follows from the data presented in Examples 1, 2,
3 and 4 that, by leaching complex lead sulfide-containing
concentrates with an iron-containing lixiviant capable of
converting lead sul~ide into lead chloriae under carefully
controlled conditions adapted to the composition of each
concentrate, lead can be substantially completely converted
into lead chloride and other metal values can be selectively
extracted into the leach solution or left in the leach residue.
Example 5
This example illustrates the treatment of the leach
residue in a two-stage countercurrent brine leach followed by
crystallization of pure lead chloride, the purification of
- 27 -
. .
brine and the washing of brine-leach residue to remove lead
chloride and chloride. 100 kg ferric chloride-leach residue
containing 50 kg lead as lead chloride was subjected to a
first brine leach at 35C. for 15 minutes with brine
containing 300 g/~ sodium chloride. The brine comprised
467 R spent brine from the crystallizer containin~ 7 kg lead
and 400 ~ brine containing 16 kg lead as chloride from the
second-stage thickener. First-leach mixture was charged to
the first-stage thickener, which yielded a 817 ~ overflow
containing 61 kg lead which was fed to the crystallizer, and
an underflow comprising 50 kg solids containing 8 kg lead and
50 ~ liquid containing 4 kg lead. The underflow was fed to a
second brine leach at 95C. with 400 ~ spent brine from the
crystallizer containing 6 kg lead and 300 g/~ sodium chloride.
After 15 minutes the second leach mixture was charged to the
second thickener from which was obtained 400 ~ overflow, which
was returned to the first brine leach, and an underflow
comprising 40 kg solids containing no lead and 50~ 1iquid -
containing 2 kg lead. After further separation, the solids
were washed sequentially with three portions of 40 ~ of hot
; crystalli~er-condensate each. All liquids were combined
giving 170R solution containing 2 kg lead, which was fed to
the crystalli~er. The washed brine-leach residue comprised
40 kg solids and 40 ~ liquid containing no lead and less than
0.1~ chloride.
In the crystallizer, the solution was cooled to 23C.
by evaporative cooling yielding 50 ~g lead as pure lead chloride
(99.8~), 867~ residual brine containing 13 kg lead which was
returned to the first- and second-stage brine leaches, and
120 ~ hot condensate, which was used to wash lead chloride and
brine from the final brine-leach residue.
-- 28 --
The presented data clearly show that proper washing
yields a substantially lead-free and chloride-free residue
and that no water is necessary to perform the washing over and
above the amount obtained in the evaporative crystallization
of lead chloride.
Example 6
This example illustrates that brine solutions can be
effectively purified and that a sodium chloride brine can be
more effectively purified than a calcium chloride brine. One
litre spent brine from the crystallizer was neutralized at
about 70C. with lime to a pH of about 8 while air was bubbled
through the solution. The resulting purified brine was
separated from precipitated solids. Test results are shown in
Table V~
- Table V
_ . . . .
Brlne Assay in mg/~
_ . ~ ......... _ _
Brine Sample pH TeOmcp Lime Pb Zn Fe Cu Bi Sb As Ag
CaCl~ 8rine _ _ _ _ ~ _ _ _ _ _
Unpurified 0.5 _ _ 30000 780 3400 210 260 40 36 215
Purified 8.0 69 4.9 2000 132 18 2 30 0.~ 0.5 180
NaCl Brine
. .
Unpurified 0.5 _ _ 20000 900 4800 240 270 49 40 240
Purified 8.5 68 17.0 225 2 4 Cl 6 <0.1 <0.1 2
Example 7
In this example it is shown that lead, which meets
the ASTM specification for corroding lead, can be produced
electrolytically from a molten salt eutectic mixture containing
92% by weight of lead chloride and 8% by weight of sodium
chloride.
- 29 -
~3
Lead chloride obtained from the test of Example 5
having a purity of 99.8~ was used to make the eutectic mixture.
The cell was a ceramic-lined vessel with graphite electrodes
spaced at 40 mm to which a current of 50A and a voltage of
3.5V were applied giving a current density of 128A/dm2. The
cell was operated at a temperature of 480C. and 200 g lead was
produced per hour with a current efficiency of 99~. The lead
was spectrographically analyzed and found to contain less than
30 parts per million of total impurities, i.e., Al, Sb, As, Bi,
Cu, Fe, Ag, Sn, Zn, Si, Ni and Ca.
Example 8
To demonstrate that the iron balance can be
maintained in the process, wherein a lead sulfide containing
concentrate is leached with a ferrous chloride and hydrochloric
acid-containing lixiviant, excess iron is rejected by oxidation
o ferrous chloride to ferric oxide with simultaneous produc-
tion of ferric chloride, and lixiviant is regenerated by
reaction of ferric chloride with hydr.ogen sulfide, 100 kg of a
lead concentrate containing 50% lead and 10% iron is treated
with 300~ lixiviant containing 40.2 kg iron as ferrous
chloride (134 g/~ iron) and 26.4 kg hydrochloric acid (88 g/~)
at 103C. for 1 hour. All of the lead and 50% of the iron are
converted to chlorides. The leach residue, containing 50 kg
lead and 5 kg iron, is fed to the brine leach, while 300~
leach solution, containing 45.2 kg iron, is split in three
portions. 200.4~ containing 30.2 kg iron is treated with
chlorine obtained from electrolysis of lead chloride, yielding
200.4~ ferric chloride solution containing 30.2 kg iron.
10~ leach solution containing 1.5 kg iron is treated for
removal of values leaving the same amount of solution and
~ 30 -
~p~
iron for treating in the oxidation. 89.6~ leach solution
containing 13.5 kg iron is fed directly to the oxidation.
The total amount of iron fed to the oxidation is 15 kg, i.e.,
three times the amount of iron dissolved in the leach solution.
In the oxidation, the solution is oxidized at 160C.
for 20 minutes under a partial pressure of oxygen of 7 atmos-
phere. The reaction equation is as follows:
3 FeC12 + 3/4 2--- 3 2FeC13 + 1/2 Fe2o3
As seen from this equation, one third of the iron is
precipitated as ferric oxide which requires that the solution
fed to the oxidation contains an amount of iron which is three
times as large. Thus, in order to maintain the iron mass-
balance in the process, the solution treated in the oxidation
must contain three times the amount of iron dissolved from the
concentrate in th~ leach.
The iron oxide is removed from the oxidation reaction
mixture leaving 99.6~ solution containing 10 kg iron as ferric
chloride. This solution is fed to the lixiviant regeneration,
. . .
together with the 200.4~ solution containing 30.2 kg iron
obtained from the chlorine treatment, wherein the solution is
~~ treated with hyarogen sulfidej obtàined from the leach o~
concentrate, at a temperature of 80C. for 30 minutes yielding
elemental sulfur and 300~ regenerated lixiviant containing
40.2 kg iron as ferrous chloride.
Example 9
This example illustrates the treatment of zinc plant
leach-residue, calcium removal and secondary brine leach.
155 g of a zinc plant leach residue containing lead, zinc and
calcium sulfate as well as silver was leached in one litre of a
solution containing 100 g/~ ferric ion as ferric chloride and
11 g/~ calcium ion at 100C. for 4 hours. After liquid-solids
separation, one litre leach solutlon was obtained. The leach
residue was leached in one litre brine containing 250 g/~
sodium chloride and 50 g/~ calcium chloride at 95C. for
15 minutes. Liquid-solids separation of the brine-leach
reaction mixture yielded one litre brine-leach solution and
80 g brine-leach residue. The compositions of solids and
liquids are given in Table VI.
Table VI
1 0 __ . . _ . .. . .
Material Unit C ~mposi :ion _
Pb Ca Zn Ag S(SO4)
~ .. ~ .... ... .. ~
Zinc-Plant Leach Residue % 36.8 5.75 2.6 4.4 10.5
(oz/t)
g 57.0 8.9 ~.0 0.023 16.3
Leach Solution g/l 4.9 1.5 3.7 0.016 2.2
Brine-Leach Solutiong/l 52.-0 18.0 trace 0~005 trace
Brine-Leach Residue % 0.3 23.0 0.35 (oz/t3 17.6
g 0.25 18.4 0.3 0.002 14.1
It is evident from the figures presented in Table VI
that lead, silver and zinc values contained in zinc plant leach
residue can be effectively recoverea, the lead in the brine-
leach solution and the silver and zinc mainly in the leach
solution. The calcium present in the ferric chloride-containing
solution is effectively removed.
