Note : Les descriptions sont présentées dans la langue officielle dans laquelle elles ont été soumises.
82
RECOV~RY OF T.J~'.~D V~I.U~
R~C~GROUND OF T~ INVENTION
In standard methods of obtaining metallic lead from concentrates,
the standard procedure has been to treat the lead sulfide concentrates in
a blast furnace. However, the pyrometallurgical procedure possesses many
disadvantages and drawbacks. Primary among these disadvantages is that
the process will result in some major pollution problems such as the gen-
eration of sulfur oxide gas along with substantial fuming. The fuming
carries with it possible carcinogenic compounds which will contain lead,
cadmium, etc. Therefore, it is necessary to provide improved and safer
methods for obtaining metals such as lead in metallic or elemental form
by methods which will not contribute to pollution of the air or will be
safer to operate. The aforementioned lead smelting techniques will consist
of roast sintering the lead sulfide concentrate whereby a major portion
of the sulfur will be removed followed by melting in a blast furnace to
.
~` obtain the metallic lead.
In an effort to alleviate the pollution problem, it is necessary
., - .
to develop new processes for obtaining lead which will be competitive as
; .
an alternative to the eonventional smelting practices. Prior work in the
hydrometallurgical field resulted in developing a non-aqueous processing
route whereby lead sulfide concentrates are chlorinated at temperatures
above 300 C. to produce lead chloride and volatilized sulfur. However,
chlorination at these elevated temperatures will promote the formation of
volatile chlorides of contaminating elements such as iron, magnesiumr
aluminum, silicon, zinc as well as elemental sulfur which may be present
in the lead sulfide concentrate. Other hydrometallurgical processes which
have been developed include the use of ferric sulfate as a leach agent.
In this method, the lead sulfide is sulfated to form lead sulfate. This
step is then followed by earbonation of the lead sulfate to form lead
earbonate and thereafter the lead earbonate is sub~ected to dissolution
ln hydrofluosilieie acid for eleetrolysis to metallic lead. Yet another
hydrometallurgical method which is developed for the recovery oE lead
ls based on the use of a ferrlc chloride medium. This method lnvolves
2 -
em~
8Z
a leaching step whereby the lead sulfide is convcrted to lea~ chlorlde and
thereafter subjected to steps of solubilizing, crystallization, and
electro].ysis.
As will hereinafter be shown in greater detail, it has ncw been
discovered that metallic lead in a relatively pure state may be produced
in a simple and economical manner by a series of steps in which lead sources
are halogenated, water washed, brine leached, and the resulting solution
treated with a reducing agent and an oxidizing agent to remove any impurities
which may still be present, thereby permitting the recovery of the desired
lead in a purer state.
SPECIFICATION
This invention re].ates to a hydrometallurgical process for the ~-
- recovery of lead. More specifically the invention is concerned with an
- improvement in a process for obtaining metallic lead from lead sulfide
sources or concentrates whereby a greater proportion of impurities which
are present in the concentrate are removed, thereby leading to the
production of a lead source which is in a purer form than has heretofore
been obtainable.
It is therefore an object of this invention to provide an improved
process for the production of lead values.
A further object of this invention is to provide a hydrometal-
lurgical process for the production of relatively pure lead values from
lead sulfide concentrates.
In one aspect an embodiment of this invention resides in a process
for the recovery of lead values from a sulfur-containing lead bearing source
which comprises the steps of: (a) halogenating said lead bearing source at
an elevated temperature, tb) washing the halogenated mixture, (c) leaching
the washed mixture with brine, (d) filtering the resulting solution to separate
elemental sulfur and residue from soluble lead halide, and (e) recovering the
desired lead values, the improvement which comprises treating the soluble
lead halide with a reduclng agent, filtering to remove insoluble residue,
thereafter further treating the soluble lead halide with an oxidizing agent
and iron, and filtering the solution to remove insoluble residue before
,~ .
' cm/ _ 3 _
110~3~i82
recoverlng sald lead values.
A specific embodiment of this invention i.s found in a process for
the recovery of lead values from a sulfur-containing lead bearing source
which comprises chlorinating said lead bearing source at a temperature in
the range of from about 90 to about 120C., water washing the chlorinated
mixture, brine leaching the wash mixture at a temperature of from about 80
to about 120 C., filtering the resulting solution to separAte elemental
sulfur and residue from soluble lead chloride, treating said soluble lead
- chloride with a reducing agent comprising lead dust at a temperature in the
range of from about 80 to about 120C., filtering to remove insoluble residuer
further treating the soluble lead chloride with an oxidizing agent comprising
air and iron at a temperature in the range of from about 80 to about 120 C "
filtering to remove insoluble residue, and reco~ering lead values by electro-
winning the soluble lead chloride.
Other objects and embodiments will be found in the following further
detailed description of the present invention.
, As hereinbefore set forth the present invention is concerned with
an improvement in a hydrometallurgical process for the production of metallic
lead. The feed stock which is ùtilized for the present process will comprise
a lead sulfide source in the form of either flotation concentrates or raw
feed ores which are naturally rich in lead sulfide or complex sulfide ores
containing lead, zinc, copper sulfides, etc., although it is contemplAted
that a portion of the lead may be present in the form of lead carbonate or
lead oxide. The feed stock may be subjected to a drying step in order to
remove any water which may be present in the lead bearing source in order
that the material will remain in a fluid state during the processing a~d
does not cake, and a~so that evolution of water will not occur-during the
subsequent halogenation step to an extent which is great enough to form
significant quantities of a hydrogen halide such as hydrogen chloride~ -
hydrogen bromide, hydrogen fluoride, etc., or other detrimental reaction
products which could affect either the chemical or physical parameters
of the process. After drying the feed stock at elevated temperatures
whlch may range from about 100 to about 150 C. for a period of time which
cm/ ~ ~ _ 4 _
C
68Z
ls sufficient to reduce the water content of the feed to a value of 2%
or less, the dried feed is then subjected to halogenation. In contra-
distinction to prior art methods hereinbefore discussed, the present
invention utilizes a halogenation temperature of the lead sulfide at
relatively low values ranging from about 90 to about 120 C. The prior --
art methods such as the treatment of lead sulfide with a large excess
of ferric chloride will give elemental sulfur and thus leave an excess
of ferric chloride which is more corrosive in nature thus necessitating
the use of more expensive equipment and, in addition, is not as selective
in the chlorination of lead only, more impurity metals going into solution
,.. .
which will come over in the filtration step along with the soluble lead
chloride. The halogenation of the lead sulfide is effected at this tem-
perature of from 90 to about 120C. by treating said lead sulfide with a
- halogenating compound such as chlorine, bromine, fluorine, etc. The halo-
~ genation of the lead sulfide with the aforementioned halogen gas will
`:i
result in the formation of a lead halide such as lead chloride, lead bromide,
or lead fluoride with the attendant formation of elemental sulfur.
The resulting halogenated mixture is then subjected to a wash whereby
impurities such as the soluble metal halides will be removed prior to
subjecting the mixture to a brine leaching operation. The wash solution may
comprise water or aqueous solutions of calcium chloride, sodium chloride,
ammonium chloride, etc. The washing of the halogenated mixture will remove
such soluble metal chlorides as ferric chloride, copper chloride, zinc
chloride, cadmium chloride, etc., whereby the lead which is eventually
recovered will be in a purer form than that which has heretofore been
obtained. The washing of the halogenated mixture may be effected over
a relatively wide range of temperatures such as from about 5 to about
95 C., the amount of water or salt solution which is utilized for the
washing step varying according to the amount of halogenated mixture.
The wash solution is then separated from the solid halogenated mixture
and charged to a treatment step whereby the wash water or salt solution
may be treated for discharge or may also, if so desired, be treated for
the recovery of the metallic impurities which have been removed and
cm/ ~ ~ ~ 5 ~
.
i~9~82
separated from thc halogenated lead mixture. ~he sollds are then leached
by the addition of a brinc solution at an elevated temper~ture usually in
the range of from about R0 to about 120C., said brine solu~ion usu<~lly
comprising an aqueous sodium chloride solution cont~ining from about 20
to about 35% by weight of sodium chloride. During the brine leaching
step, the pH of the solution is maintained in the range of from about
4 to about 6 by the addition of acidic or caustic solution such as the
hydroxides of Group ]A of the Periodic Table including sodium hydroxide,
potassium hydroxide, lithium hydroxide, Group IIA of calcium, magnesium
oxide or halogen acids such as hydrochloric acid, hydrobromic acid/ etc.
By contolling the pH of the brine leaching solution in the aforesaid
range, other metallic impurities which are present in the solution such
as copper, silver, zinc, cadmium, antimony and possibly iron along with
: some unreacted sulfides will precipitate from the solution either by
hydrolysis or by reaction to form insoluble sulfides under the conditions
of the solution. The leaching of the mixture is effected for a period of
, time which may range from about 0.25 to about 2 hours or more in duration,
the residence time being that which is sufficient to dissolve the lead
halide.
Upon completion of the leaching step, the solution is then
subjected to a separation or filtration step in which the solid residue
which comprises elemental sulfur as well as any solid gangue impurities
are removed by separation from the sol~ble lead halide. The temperature
~ during the filtration or separation step is also effected at elevated
temperatures ranging from about 80 to about 120 C. whereby the lead
halide is maintained in a soluble form. Thereafter the liquid portion of
the separation is subjected to the action of a reducing agent which may
include among others elemental lead in the form of lead dust, sodium
sulfide, hydrogen sulfide, sulfur dioxide, etc. By subjecting the
solution to a purification, it is possible ta remo~e traces of metal
impurities such as silver, copper, mercury or any other metals which are
more noble in character than lead. If so desired, the step o~ solution
purification by adding reducing-reagent may be combined with the step of
~? cm/ ~ - 6 -
~10~68Z
brine leaching of the washed resldue.
The resulting solution which has been treated with a reducing
reagent of the typc hereinbefore set forth, although it is also contemplated
within the scope of this invention that any other compound which acts as a
reducing agent may also be used, is then subjected to another filtration
step whereby the solid residue which is formed by the reaction with the
; reducing agent is removed. Again the liquid portion is passed to a second
solution purification step wherein the solution is subjected to the action
of an oxidizing reagent. Examples of oxidizing agents which may be employed
in this second purification step would include compounds such as chlorine
gas, hydrogen peroxide, lead peroxide, sodium peroxide, potassium peroxide,
potassium permanganate, sodium permanganate, air, oxygen, etc. In addition
to the oxidizing agent, ferric iron is also added to the solution. In this
second solution purification zone any traces of antimony, bismuth, arsènic,
~ etc., will be removed with insoluble iron compounds while maintaining the
; p~ of the solution at a value greater than about 3. The resulting solution
is then subjected to a second separation or filtration step whereby the
residue which is insoluble in nature due to the oxidation of the metal
impurities is removed and the resulting solution which contains the still
soluble lead halide is passed to a lead recovery step. It is to be noted
that the temperature of the solution during the first purification step,
filtration, second purification step and second filtration is maintained
at an elevated range of from about 80 to about 120 C. in order to maintain
~ the lead halide in soluble form.
The filtrate which is recovered may then be treated in either
of two ways in order to recover lead values which are substantially pure
and free from any metal impurities. One method of recovering the desired
lead values is to pass the soluble lead halide to a crystallization zone
wherein the soluble lead halide is crystallized due to a temperature drop,
i.e., less than about 60 C., the solubility of the lead halide decreasing
as the temperature decreases.
The thus crystallized lead halide is then recovered and, in the
preferred embodiment of the invention, is dried to remove any trace of
- 7 -
~i~9~8Z
water which may still be present, the dryll'~ may be effected, lf so
desired, by placing the lcad halidc in an oven and sub~ecting the lead
halide to a temperature of about 100 C. in an atmosphere of air for a
period ranging from about 0.1 to about 4 hours or more, the duration of
the drying period being that which is sufficient to remove all traces
of water. Following the drying of the lead halide, it is then placed
..
in an appropriate apparatus and subjected to a temperature sufficient
to melt said halide until it assumes a molten form. This temperature
` may range from about 380 C. which is sufficient to melt lead bromide up
to about 875 C. which is sufficient to melt lead fluoride. The lead
halide in molten form is then admixed with a salt of a metal selected
:
from the group consisting of alkali metals and alkaline earth metals.
Examples of these salts of metals of Groups IA and IIA of the Periodic
~; Table will include lithium chloride, sodium chloride, potassium chloride,
rubidium chloride, cesium chloride, beryllium chloride, magnesium chloride,
calcium chloride, strontium chloride, barium chloride, lithium bromide,
sodium bromide, potassium bromide, rubidium bromide, cesium bromide,
beryllium bromide, magnesium bromide, calcium bromide, strontium bromide,
barium bromide, lithium fluoride, sodium fluoride, potassium fluoride,
rubidium fluoride, cesium fluoride, beryllium fluoride, magnesium fluoride,
calcium fluoride, strontium fluoride, barium fluoride, etc., in a fused
salt bath. In the preferred embodiment, the salt of a metal of Groups IA
or IIA of the Periodic Table will be comparable in the halide content to
the lead halide which is to undergo electrolysis, that is, if the lead
halide is lead chloride, the solid salt will comprise a chloride such as
sodium chloride, potassium chloride, lithium chloride, calcium chloride,
etc. In general, the salt of the metal of Groups IA or IIA of the
Periodic Table will be present in the fused salt mixture in an amount in
the range of from about 20% to about 40~ by weight of the mixture. It is
also contemplated within the scope of this invention that the lead halide
will undergo electrolysis in the presence of a mixture of at least two
salts of the metals of Groups IA and IIA of the Periodic Table, examples
of these mixtures comprising a sodium chloride-lithium chloride mixture,
t ~ ~ - 8 -
~i~9~l~32
A potasslum chlo~ide-lithium chloride mixture, a magnosium chloride-
calcium chloride mixture, a lithium bromide-potassium bromide mixture,
etc, In the fused salt bath the mixture of salts will be subjected to
electrolysis utilizing a sufficient voltage to effect said electrolysis
whereby metallic lead is deposited as a liquid which can be removed from
the fused salt. The lead may be removed continuously or batchwise.
As an alternative method for the recovery of the desired lead values,
the soluble lead chloride solution may be subjected to an aqueous electro-
winning process whereby the lead is extracted from the lead chloride solution
by eiectrolysis whereby the metailic lead is deposited out at the cathode
of the cell.
The present invention will be further illustrated with reference
to the accompanying drawing which illustrates a simplified f~ow diagram of
the present process. ~arious valves, coolers, condensers, pumps, controllers,
etc., ha~e been eliminated as not being essential to the complete under-
standing of the present invention. The illustration of these, as well as
other essential appurtenances will become obvious as the drawing is described.
Referring to the drawing, a charge stock of lead-containing con-
centrates, after being dried at an elevated temperature ranging from about
100 to 150 C. whereby the water content of the ore is reduced to a value
of 2% or less, is charged through line 1 to a halogenation zone 2. A halo-
genating agent such as chlorine gas, fluorine gas, bromine gas, etc. is
charged through line 3 to halogenation zone 2 for a period of time sufficient
to convert the lead sulfide to lead halide. The halogenation of the lead
sulfide to lead halide is effected at a temperature in the range of from
about 90 to about 120 C. In halogenation zone 2 the treatment of the
lead sulfide with the halogenating agent is accomplished in such a manner
as by stirring, mixing, shaking, fluidization, etc,, whereby all of the
lead sulfide is contacted with the halogenating agent. The resulting
mixture of elemental sulfur and lead halide is then passed through line
4 to wash zone 5. The mixture is contacted in wash zone 5 with an influx
of water or an aqueous salt solution of the type hereinbefore set forth
through line ~ whereby impurities such as soluble metal halides including
X cm/ ~ - 9 -
~9682
. . .
such compounds as iron chloride, copper chloride, zinc chlorlde, cadmium
chloride, etc., along with the elemental sulfur and lead halide are passcd
through line 7 into filtration zone 8. I~he soluble portion of the solution
is separated from the elemental sulfur and solid lead halide and removed
through line 9 for recycle to wash zone 5, a portion of the recycle being
-~ bled through line 10 and removed. The solids comprising elemental sulfur
and lead halide are then removed from filtration æone 8 through line 11 and
passed to brine leaching zone 12. In leaching zone 12 the solid product is
treated with an aqueous brine solution containing from about 20% to about
~ 10 30% by weight of sodium chloride, the addition of the brine solution being
- accomplished by passing said brine solution into leaching zone 12 through
line 13. The leaching step of this process is effected at elevated tem-
peratures ranging from about 80 to about 120 C. In addition, the p~ of
the brine leaching solution is maintained in a range of from about 4 to
about 6 during the leaching step by the addition of a caustic solution
such as sodium hydroxide, potassium hydroxide, etc., or a hydrohalic acid
such as hydrochloric acid, if required, through line 14. Upon completion
of the leaching step, the mixture is passed through line 15 to a separation
or filtration zone 16 whereby the soluble lead halide is separated from
elemental sulfur as well as any solid gangue impurities. The separation of
the soluble lead halide solution and the solid sulfur may be effected by
filtration or by flotation and settling whereby, after allowing the solid
residue containing elemental sulfur and/or impurities to settle, the liquid
is removed by conventional means such as decantation, etc. The solid sulfur
and residue are removed from filtration zone 16 through line 17 to a
recovery zone, not shown in the drawings, wherein the residue which contains
gangue, unreacted sulfides of the impurity metals such as zinc sulfide,
copper sulfide and iron sulfide as well as elemental sulfur is subjected
to a recovery treatment. The elemental sulfur may be separated from the
impurities and recovered by any method known in the art, one example of
such a recovery method which may be employed comprising froth flotation
method in which the sulfur is preferentially floated. In addition, a
scrubbing step to more fully liberate sulfur from the remainder of the
-- 10 --
~i cm/~ ~
6~32
residue may also be effected ~n the prescnce of a flotation promote~ such
as organic compounds readily available, for cx~mple, kerosene, The txeated
~ material is then tr,~ erre~ to a flotation ceil, a frothing agent is
-~ add~d, aeration is initiated, and the sulfur-laden froth is removed from
the cell. As an alternative method for the separation and recovery of
` elemental sulfur from impurities, the residue may also be treated with
aqueous ammonium sulfide in which the ammonium polysulfide which is formed
permits the recovery of elemental sulfur in a crystalline form. In like
manner, the impurities comprising various metals which are present in the
lead sulfide concentrate may also be recovered by conventional means such ~-
as cyanidation of the residue in a leaching operation to recover silver
or other precious metals.
The solution which contains soluble lead halide such as lead
chloride is removed from filtration zone 16 through line 18 and passed
to a first solution purification zone 19. In this solution purification
zone, the solution is subjected to the action of a reducing agent of the ~`
type hereinbefore set forth which is passed to solution purification zone
19 through line 20. It is to be noted that the temperature during the
steps of brine leaching, filtration and both solution purifications as
well as separation of the insoluble residues is maintained in a range of
from about 80 to about 120C. in order to maintain the lead halide in
soluble form. After subjecting the solution to the action of the reducing
agent, the resulting mixture is passed through line 21 to a separation
or filtration zone 22 whereby the residue is separated from the soluble
lead halide solution through line 23. The solution is then passed from
filtration zone 22 through line 24 to a second solution purification zone
25. In this second solution purification zone the soluble lead halide in
solution is subjected to the action of an oxidizing agent and elemental
iron which are passed into zone 25 through line 26. After thorough admixing,
the solution, which contains the soluble residue of impurity metals which
have been oxidized, is withdrawn from solution purification zone 25 through
llne 27 to filtration zone 2a whereby the solid residue is removed through
line 29 and the soluble solutlon passes through a line 30 to lead recovery
cm/ ~
1~9~8Z
`, .
zone 31. In lead recovery 31 the soluble lead halide so]ution may be
subjected to an aqueous electrowinning process after stripping the brine
leaching solution therefrom, said brine leaching solution which contains
sodium chloride as well as any caustic material or acid being removed
from lead recovery zone 31 through Iine 32 and recycled to leaching zone
12, a portion of the solution being bled off through line 33. Alternatively,
the soluble lead halide solution may be passed through line 34 to a
crystallization zone whereby the lead halide crystals will form due to a
temperature drop and thereafter the crystals separated from the barren
leach solution, said barren leach solution being recycled back to leaching
zone 12 for further use therein. After separation of the lead halide
crystals from the leach solution, the crystals may be passed to a drying
zone such as an oven wherein all traces of water are removed by heatins
at an elevated temperature in excess of 100 C. for a predetermined period
of time. The aforementioned dried lead halide crystals may then be removed
from the drying zone and passed to a fused salt bath wherein said crystals
are subjected to electrolysis in the presence of a salt of the type herein-
before set forth. By effecting the electrolysis at an elevated temperature
which is sufficient to maintain molten conditions, it is possible to remove
and recover metallic lead from the fused salt electrolysis zone while the
halogen molecules may be recycled, if so desired, back to the halogenation
; zone. It is to be noted that by utilizing such a flow system it is possible,
after leaching the stoichiometric amount of halogen which is necessary to
react with the lead sulfide, to reuse the halogen in a recycle or closed
system thereby obviating the necessity of adding an additional amount of
a halogen in any large quantities. This latter step will contribute to the
lower cost of the overall process used in obtaining metallic lead from lead
halide feed stocks.
While the above discussion has been descriptive of a continuous
method of operating the process of the present invention, it is also con~
templated that the recovery of metallic lead from a lead sulfide souxce
may also be effected in a batch type of operation. When this type of
cm/ ~ - 12 -
,
11~96~3Z
operation i9 used, a quantity of the charge stock ls placed ln a drying
apparatus such a.s an oven and subjected to a drying step at a temperatuxe
within the range hereinbefore set forth. Upon completion of the drying
step, the charge stock is then placed in an appropriate apparatus and is
thereafter subjected to the action of a halogenating agent. Inasmuch as
the halogenation of the lead sulfide is exothermic in nature, the heat of
reaction which is evolved will be controlled within the desired operating
range hereinbefore set forth, although it is contemplated that heating or
cooling means may be provided to stabilize the temperature of the reaction.
Upon completion of the conversion of the lead sulfide to the desired halide,
the ha].ogenated product is then washed to dissolve any soluble metal haiide
compounds other than lead which may be present as impurities in the charge
stock. The washed solid product is separated from the water by conventional
means such as filtration, decantation, etc., and thereafter placed in an
appropriate apparatus where it is subjected to the action of a brine
leaching solution, the lead halide being solubilized. After agitating the
solution for a predetermined period of time sufficient to dissolve the lead
halide while maintaining the pH of the solution in a range of from about 4
to about 6 by the addition of a controlled amount of caustic solution or
hydrohalic acid, if necessary, the soluble lead halide is separated from
elemental sulfur and residue and is thereafter recovered. The filtrate is
then placed in an apparatus and subjected to the action of a reducing agent
while maintaining the temperature in the range of from about 80 to about
~ 120 C. in order to maintain the lead halide in soluble form. Again the
solid residue which results from the action of the reducing agent is
separated from the liquid solution by conventional means following which
the liquid solution is placed in another apparatus and subjected to the
action of an oxidizing agent and iron while maintaining the temperature in
the aforementioned range. After removal of the solid residue which results
from the oxidizing reaction, the solution lS then either cooled to crystallize
out the lead halide compound fo~lowed by a molten salt bath electrolysis or
subjected to an aqueous electrowinning process for the recovery of metallic
lead.
cm~ 13 -
i82
The ollowing examples arc given for purposes of illustrating
the process of this invention. However, these examples are given merely
for purposes of illustration and are not intended to limit the generally
broad scope of the present invention in strict accordance therewith.
EXAMPLE 1
In this example 10 kilograms of a lead sulfide concentrate which
is relatively high in antimony, zinc, copper, iron and silver impurities
were reacted with 6.0 pounds of chlorine gas over a period of 5.4 hours
while maintaining the temperature of the reactor between 93 and 101 C.
Following this, the material was aged for an additional period of 1 hour .
at a temperature of about 85 C. An analysis of the lead concentrate before
and after chlorination is set forth below:
Analysis, Weight %
Before Post Conversion
ElementChlorination Chlorination %
Lead 67.8 54.4 98 -`
Iron 5.45 4.41 --
Copper 0.72 0.78 --
Zinc 4.00 3.1 --
Silver 43.32* ----- --
Antimony 0.30 ----- ~~
Bismuth ~0.10 ----- --
Sulfur 17.8 14.6 --
Chlorine ----- 21.0 --
* Ounces per ton
Following this, 30 grams of the chlorinated product were water
washed by treatment with 150 cc of water for a period of 30 minutes at
a temperature of 25 C. while maintaining the pH in the range of from
2.05 to 6Ø The filtrate of the water solution was analyzed with the
following results:
- 14 -
~ cm/ k~
682
_ 2 3 4 5
Lead
ppm 2325 3350 1125 3175 2075
Extraction %2.~34 4.1 1.4 3.9 7.54
Iron
ppm 2050 1950 645 1900 1800
Extraction %31.0 29.5 6.7 2~.7 27.2
Zinc
ppm 875 875 400 925 800
Extraction %18O8 18.8 ~.6 20.0 17.2
Copper
ppm 550 550 0 545 570
Extraction ~ 47 47 0 46.6 45.3
Following this, the concentrate from the water wash was subjected
to a brine leaching using a 25% sodium chloride solution as the brine leach.
100 cc of the filtrate was analyzed after maintaining the pH in a range of
from 4.7 to 5.9. The results are set forth in the following table:
1 2 3 4 5
_
Iron
ppm 110 140 15 160 110
Extraction % 0.8 1.0 0.1 1.2 0.8
Zinc
ppm 50 31.3 75 37.5 62.5
Extraction % .5 .3 .80 0.40 0.7
Copper
ppm 0 0 0 0 0
Extraction % 0 0 0 0 0
The filtrate or solution from the brine leach was then subjected to
a purification step by treatment with lead dust. Analysis of the solution
showed that it contained 10 parts per million of silver. To remove the silver
500 cc of the brine leach solution was treated for a period of 1 hour at a
; temperature of 100 C. with 200 milligrams of lead dust. The silver pre-
; cipitated out during the test and it was found that after separation of the
precipitate, the liquid portion which had contained 10 parts per million
before treatment contained not more than 1 part per million of silver and
less than 10 parts per million of antimony after treatment with lead dust.
The liquid which contains lead chloride in solution along with
30 other possible contaminants such as antimony, bismuth and arsenic may then
be treated with iron and an oxidi7ing agent such as air by passage of air
cm/ ~ - 15 -
~g~82
through the solutlon while maintalning the temperature at about 100 C.
After a period of about 1 hour, the solution may be filte~ed to ~emove
the precipitated impurities and the relatively pure lead chloride may be
precipitated by passing said solution into a crystallization apparatus at
a temperature below about 60C. The precipitated lead chloride may then be
admixed with sodium chloride followed by electrolysis, said electrolysis
of the Eused salts being effected at a temperature of about 550 C. using
a voltage of 2.4 volts. The desired metallic lead may be recovered by
tapping the electrolysis apparatus.
EXAMPLE II
In this example a lead sulfide concentrate which does not contain
as high a percentage of contaminants such as antimony~ zinc, iron, copper
and silver was chlorinated by drying 10 kilograms of the concentrate in an
oven at 110 C. for a period of 4 hours. The thus dried concentrate was
then chlorinated by treatment at a temperature of about 95C. for a period
of 4.5 hours using 2400 grams of chlorine gas. The thus treated concentrate
was aged for a period of l.S hours at 85C. Analysis of .he chlorinated
product showed a weight percent of:
Pb 58.6%
Cu 0.42%
Fe 2.4~
Zn 1.1%
In addition, it was found that 95.0% of the lead was converted to lead
chloride.
Thereafter four 50 gram portions of the lead chloride concentrate
were water washed and the filtrate made up to 200 cc. The filtrate was
analyzed with the following results:
1 2 3 4
Wash Water Vol. cc 150 150 150 150
Wash Temp. C. 25 100 100 25
Wash Water, NaCl % 0 0 5 5
Flltrate Vol. cc 200 200 200 200
Filtrate pH 2.2 3.9 4.1 2.6
Pb conc. in Filtrate, ppm S000 4000 3500 3000
Pb extraction in Filtrate, ~ 3.4 2.7 2.4 2.0
Fe conc. in Filtrate, ppm 1800 750 340 1700
Fe extraction ln Filtrate, ~ 30 12.5 5.6 28.3
cm/~ - 16 -
fi8Z
1 2 3 4
Zn conc. in Filtrate, ppm 210 - 250 2S5 215
Zn extraction in Eiltrate, % 7.6 9.1 9.3 7.8
Cu conc. in Filtrate, ppm 375 50 75 450
Cu extraction in Filtrate, %35.7 4.8 7.1 43
The concentrate which was recovered from the water wash was then
subjected to a brine leach using a 25% solution of sodium chloride, said
leaching being effected at a temperature of 100C. The filtrate from
the brine leach after separation of elemental sulfur and gangue was analyzed
with the following results:
1 2 3 4
Brine Leach Filtrate, cc 300 300 300 300
Brine Leach Filtrate, pH 4.5 4.5 4.8 4.8
Fe conc. in Filtrate, ppm 14 10 12 12
Fe extraction in Filtrate, % 0.8 0.6 0.7 0.7
Zn conc. in Filtrate, ppm 4.6 6.0 2.8 3.8
Zn extraction in Filtrate, % 0.6 0.7 0.3 0.5
Cu conc. in Filtrate, ppm 0.0 0.0 0.0 0.0
Cu extraction in Filtrate, % 0.0 0.0 0.0 0.0
Following this the solution may then be subjected to a reducing
reaction by treatment with 200 milligrams of lead dust while maintaining
the temperature of the solution at about 100 C. Eor a period of 1 hour.
Thereafter the residue which is formed by separation of silver and other
impurities may then be separated from said residue and subjected to an
oxidizing reaction wherein the solution is treated with an oxidizing agent
such as sodium peroxide plus ferric iron while maintaining the temperature
of the solution at about 80 to 120 C. for an additional period of about
1 hour. After separation of the solid impurities, the li~uid which contains
purified lead chloride in solution is then treated in a manner similar to
that set forth in Example I whereby relatively pure metallic lead may be
recovered by an electrolysis process.
cm/ ~ - 17 -
