Note: Descriptions are shown in the official language in which they were submitted.
1 15G048
This invention relates to the extraction of metal
compounds from metal bearing materials and more particularly
to t~e extraction and recovery of lead values in a calcium
plumbate and/or oxide product from minerals or lead bearing
materials. Silver which may be present in association with
the lead may be recovered as native silver, silver chloride,
sulfide or sulphate, or a silver complex with other metals,
or in some other form from which it can be recovered by
conventional techniaues.
The recovery of a high grade lead product suitable for
treatment for metal recovery from lead bearing minerals
has usually been accomplished by flotation concentration
of coarse grained lead sulphide deposits into a concentrate
containing greater than 50~ lead and pyrometallurgical
reduction of this concentrate in a blast ~urnace. The reserves
of these coarse grained lead sulphide deposits are rapidly
being depleted. The major new reserves of lead are being
found in fine grained massive sulphide deposits containing
~'
1 15~0~
sulphides of zlnc, lead, copper, silver, and iron. ~ecoveries
into high grade lead concenlrates are typically low from these
deposits, necessitating significan~ reduction in grade to
maintain economic recoveries. It will be necessary for some of
these deposits to resort to the production of bulk zinc, lead,
copper concentrates to insure high recoveries. Several new
processes are available for treating these low grade and bulk type
concentrates includlng ferric chloride leach processes, copper
chloride leach processes, sulphuric acid-oxygen pressure leach
processes and the sulphation roast process. The lead and silver in
the latter two processes report in a low grade lead sulphate-hematite
leach residue. In the ferric and cupric chloride leach processes,
leach filtrates are produced which contain lead and silver as
chlorides in a concentrated brine solution. The present process
has application for lead and silver recovery from the leach
residues and brine solutions generated in all of these processes.
Substantial reserves of lead and silver also exist in leach
residues from electrolytic zinc plants. These residues typically
assay 15-40~ lead as lead sulphate and for the most part are
considered as unsuitable as feed for a conventional lead smelter
except in small amounts. Ano~her source of low grade materials is
51ag ~rom lead smelters. Lead is prasently recovered from these
sla~s by energy intensive ~umin~ processes. The present process can
be employed directly to recov~r lsad and silver E~om zinc plant
residue5 and a~ter either sulphuric acid leaching or sulphation
roasting to ~ecover lead and silver ~rom slags.
It is known that lead ~ulphate and associated silv~r may be
solubilized by means o~oncentrated brines as proposed in Canadian
Pa~ent 1~,918, ~883); Canad~an Patent 228,518,(1919); and U.S. Bureau
~0 - 2 -
1 ~ 5~0~8
of .~.ines sulletin 1~7,(1918). Whilst these methods solve
th~ problem of separating the le~d and silver from the residues
there has been som~ economic difficulty in the subsequent
recovery of the lead and silver from the solution in a usable
form.
West Ger~an Patent 2,500,453, (1976) descr.ibes a method of leaching
lead sulphate containing material in sodium chloride solution
and after residue separation, precipitating the Pb from solution
with milk of lime~ The lead precipitates contain greater than
10~ chloride and 11~ sulphate and are not acceptable to conventional
lead smelters except in small amount and at depressed prices
due to the deleterious effects of chlorides on acid plant catalysts,
baghouses, and refractories.
Canadian Patent 228,518,(1gl93;United States Patent 4,063,933,
(197~; and processes currently being developed by the U.S. Bureau of
Mines, Minemet Recherche,(France), and Cominco Limi.ted (Canada),
advocate lead recovery hy the precipitation of lead chloride crystals
from solution by evaporation and/ox cooling. The subsequent recovery
of lead metal to be accomplished by electrolysis. Capital and
operating costs are projected to be much higher for these processes
than for conventional smelting.
Australian Patent 28,957,~1971)describes a method of precipitating
lead chloride from brine solution by c041 Lng followed by reacting
said lead chloride with water and calcium sulphate to produce a lead
~ulphate precipitate and calcium chloride solution. Although low
chloride levels in the lead sulphate were obtainable with rigorous
wa.~hing, the pxQduct ls a~ain ~uitable -to lead smelters in lim:Lted
quantit.ies and mu~t be t.reated on a sin-ter machine to remove
~15~
sulphur before the blast. Capital and operating costs are projected
to be high for the process since the brine solution must be heated
for high lead solubLlity and then cooled for lead chloride
precipitation.
Canadian Patent 228,518,(1919)describes a method of lead
recovery from concentrated brine solution by direct precipitation
as sulphide or sulphate. These precipitates are difficult to
wash and contain significant amounts of entrapped chlorides.
Again conventional lead smelters will accept only small quantities.
Canadian Patent 1~918 describes a method of precipitating
lead and silver from brine solution by cementation on metallic
zinc. Recently other researchers have rediscovered the cementation
techique and advocate either zinc or iron as cementation media.
High grade metallic lead and silver cementates are produced in
these processes which are acceptable to lead smelters at premium
prices at high tonnage. Considerable economic penalties are
incurred to produce good quality cementates, however, since the
cementation reagents are expensive and the zinc or iron in the
lean brine re~ulting from cementation must be recovered in a form
acceptable for sale or reuse. This can be accomplished only at
considerable cost.
It is an objective of this invention to provide a process
~or the extraction and r covery of lead and silver into a product
which will be acceptable t4 conYentional lead smelters in large
~onnages and at a premium price.
Furt~er objectiv~s are ~or the proce~s te consume minimum
energy and the reagent~ us~d to be recovered and either reu~ed
with high ef~tciency or credited in the sale o~ t:he laad product.
1 1~60~8
In accordance with a broad aspect of thi.s invention there
is provided a process comprising the steps of (1) preparing a
solution of lead chloride by dissolving lead sulphate contained
in an ore or process residue in an acidic concentrated chloride
brine; (2) separating the solution so formed from insoluble
gangue or other residue; (3) forming a prec.ipitate of lead
oxychloride by adding lime to said solution and separating said
lead oxychloride precipitate from the residual lean brine
solution; (4) reacting the said oxychloride precipitate with
oxygen and lime in a reactor at an elevated temperature to
produce a calcine containing most of the lead as calcium plumbates
and/or lead oxides; (5) washing said calcine in water and/or
dilute chloride brine to dissolve soluble chlorides; (6) separating
the resulting residue obtained from the resulting chloride brine;
and, (7) washing said residue containing calcium plumbates and/or
lead oxides with fresh water to remove residual chlorides.
In another broad aspect there is provided a process for
gaining lead and silver values eomprising the steps of (1)
preparing a solution of lead chloride and silver compounds by
dissolving lead sulphate and silver compounds contained ln an
ore or process residue in an aeidie eoneentrated ehloride brine;
(2) separating the solution so ~ormed ~rom insolubl~ ~angue or
o~her re~idue; t3) ~ormin~ a preeipita~e of lead oxychlo~ide and
silver eomp~unds hy adding lime ~o said ~olution and separating
said precipitate from the residual lean brine solution;
1 1560a,8
(4~ reacting the said oxychloride precipitate with oxygen
and lime in a reactor at an elevated temperature to produce a
calcine containing most of the lead as calcium plumbates and/or
lead oxides and most of the silver as silver or silver
compounds; (5) washing said calcine in water and/or dilute
chloride brine to dissolve soluble chlorides; (6) separating
the resulting residue from the resulting chloride brine; and,
~7) washing said residue containing calcium plumbates and/or
lead oxides, as well as silver and silver compounds, with
~resh water to remove residual chlorides.
In accordance with another aspect of this invention we
provide an improvement in a process comprising the step of
(1) preparing a solution of lead chloride by dissolving lead
sulphate contained in ore or process residues in concentrated
chloride brine, thereby also dissolving any silver associated
with the lead; (2) separating the solution so formed from the
insoluble gangue and other residues; (3) forming a precipitate
of lead oxychloride (and any silver which may be present) by
adding lime to the solution and separating the lead oxychloride
and silver precipitate from the residual lean brine solution;
(4) recycling the lean brine, normally after concentration
thereof such as by evaporation or by addition of further chloride
and also normally after re-acidification by the addition of
~urther acid, ~or reuse in the ~urth~r extraction o~ lead
sulphata a~ under s~eps (1) and ~2). ~h~ improvem~nt compr:isas
~5) xeac~in~ the said praaipita~e ~on~aining lead oxychlorida
-- 6 --
1 ~ S~04~
with oxygen such as by air and with excess lime present in the
precipitate, and if desired adding fresh lime, in a reactor at
a temperature above 325C for longer than one half hour to
produce a calcine containing most of the lead as calcium plumbates
and lead oxides, and containing any silver present as native
silver, silver chloride, sulfide or sulphate, and complexes of
silver with other materials; (6) repulping said calcine in water
and/or dilute chloride brine to remove soluble chlorides;(7)
separating the residue obtained in step (6) from the resulting
chloride brine; (8) recycling the brine resulting from step ~7),
with the optional treatment rnentioned above, for further extraction
of lead sulphate under the previous steps; (9) washing the said
residue from step (7) with fresh water to remove residual
chlorides; and (10) recycling the chloride brine obtained in
step (9) to step (6) and/or recycling the said chloride brine,
again with the optional treatment mentioned above for reuse
in the further extraction of lead sulphate under the previous
stép.
In one aspect the present invention provides, in a process
comprising the step of forming a precipitate of lead oxy~
chloride by adding lime to a chloride brine solution cont:aining
lead chloride, and separating said lead oxychloride precipitate
from the residual lean brine solution; the improvement which
comprlse~ reaqting the said oxychloride preaipitate Wi~h oxygen
and lime in a reactor at a temperature above 325C. ~or longer
than Q.S hour~ to produce a calcine containing most. of the lead
as calai~n plumbates and/or lead oxides; repulpln~ said aalcine
1 1 5~)'18
in water and/or dilute chloride brine to dissolve sol~ble chlor-
ides; separating the residue obtained from the resulting chloride
brine; and wash.ing said residue containing calcium plumbates and/
or lead oxides with fresh water to remove residual chlorides.
In another aspect the invention provides a process for gaining
lead and silver values comprising the steps of preparing a solution
of lead chloride and silver compounds by dissolving lead sulphate
and silver compounds contained in an ore or process residue in an
acidic concentrated chloride brine; separat:ing the solution so
formed from insoluble gangue or other resiclue; and forming a
precipitate of lead oxychloride and silver compounds hy adding
lime to said solution and separatiny said precipitate from the
residual lean brine solution, characterised by the further steps
of reacting the oxychloride precipitate with oxygen and lime in
a reactor at an elevated temperature to produce a calcine con-
taining most of the lead as calcium plumbates and/or lead oxides
and most of the silver as silver or silver compounds; washing
the calcine in water and/or dilute chloride brine to dissolve
soluble chlorides; and separating the resulting residue containing
calcium plumbates and/or lead oxides, as well as silver and
silver compounds, from the resulting chloride brine.
~ 7a -
~ 15~0~8
In the drawings which accompany this invention:
Figure 1 is a schematic flow sheet showing certain aspects
of the present invention;
Figure 2 is a graph showing the relationship between
calcium chloride addition and sulphate in solution;
Figure 3 is a schematic flow sheet showing a lead-silver
recovery plant employing certain aspects of this invention.
The advantages of producing a calcium plumbate product
are as follows:
.- 7b -
1 ~ 560~8
i) calci~n plumbate is not water or cold brine soluble and
will not react with chloride brines under neutral or
basic conditions to reform lead oxychlorides.
ii) entrained chlorides in the plumbate calcine can be easily
removed and reduced to very low levels by washing with water or
unsaturated brine solution.
iii) plumbate repulp solutions Eilter rapidly, leaving a dry
residue.
iv) plumbate products can be briquetted and fed directly into a
lead smelter blast furnace without prior sintering, increasing
smelter throughput for smelters in which the capacity is
limited by the sinter machine.
v~ as reported by Denev, D.G. et al in Dokl. Bolg. Akad. Nauk,
Vol. 26, 11, 1973, page 1485 calcium orthoplumbate is an
oxidant for lead sulphide at high temperature resulting in the
products PbO, CaO and SO2 and hence would make a good dilutant
for galena concentrate on a sinter machine.
vi) CaO is a product ofthe rPduction of calcium plumbate and is also
required as a slagging agent in lead blast furnaces, usually
at high tonnages. Accordingly, since the use of some
calcium plumbate as feed to a lead smelter would reduce the
requirement for lime, some credit should be given ~or the lime
in the plumbate product.
vii) the production of a calcium plumbate product allows for the
use o~ lime ~or the precipitation of the lead ~rom the brine
1-3ach olutlon and al~o a~ a reaatant in the high temperature
aonversion o~ oxychloxide to plumbate~ Llme i9 a relatively
ln~xpensive, ea~y to handle, envi~or~entally acceptable,
and readily available commodlty.
3~
~ :.3 ' _ ~ _
1 1 560~
~;iii) the use of lime results in the formation of calcium
chloride after the conversion of the metal chlorides
(lead, zinc, copper, iron) to oxides. This calcium
chloride is recycled in the brine to the lead sulphate
leach and results in the precipitation of most of the
sulphate as calcium sulphate into the leach residue.
Accordingly the plumbate product is low in sulphate.
Also, the low soluble sulphate in the leach enhances
the solubility of lead and silver allowing for leach
operation at lower temperatures, resulting in a lower
energy consumption and less maintenance due to decreased
corrosion. The effect of calcium chloride on sulphate
solubility in sodium chloride brine solution is shown
in Figure 2. The solubility of lead as lead sulphate in
269 gpl NaCl brine increases from about 13 gpl at 35C
to about 18 gpl lgrams per liter) at 35C when CaC12 is
added to yield a brine containing 34 qpi C~C12. Lead
solubility is directlv proportional to the brine saturation
and the sulphate concentration in the brine.
Since calcium chloride is a more expensive commodity than
sodium chloride a~d since there appears to be a lower limit to
the soluble sulphate in the brine leach solution attainable with
calcium chloride and since sodium chloride i5 easier to remove
~rQm leach re~idue by washing, it seems to be preferable but not
n~c~ary to use a concentrated sodium chloride brine as the ba~e
soluti~n u~ing the lime ad~i~ions ~n ~teps (3) and ~5) as the source
Q~ calcium chloride ~or sulphate removal. Small amoun~s o~ ~res~
NaC1 and CaC12 will be required to make up ~or losses in ~e leach
residue ~nd product.
~ 9 _
.. .. ~ . .. .
0 '1 ~
Lead and silver extractions into brine can be accelerated
by increasing the acidity by addition of an acid such as hydro~
chloric or sulfuric acid which will ensure at least mildly acidic
conditions. The optimum pH in ~he brine leach for high lead and
silver extraction, efficient residue washing, and low lime
consumption appears to be about 1.5.
Extractions of lead from lead sulphate material into brine
are very high and may approach 99~ with the proper choice of
reten~ion time, temperature, brine composition, and residue washing
t0chniques as long as the solubility limit of the lead is not
approached. ~ead extractions fall from 99% at 75~ of lead
chloride saturation to 96% at 86~ of saturation to 91% at 94
of saturation for brine leaching }n 269 gpl NaCl - 33 gpl
CaC12 - pH 1.5 brine at 35C and 1.5 hours leaching time.
The saturation limit of lead as lead chloride in this brine
is 18.3 gpl.
Silver extraction by brine is very depe~dent on the nature ~
and prior history of the lead sulphate containing material. Some
materials, usually those which have been very recently produced
in a roaster or leach process exhibit silver extractions greater
than 8Q%. Other materials, usually stockpiled, exhibit lower
silver extractions of about 50%. Silver recoveries can be
inareased from these materials by flotation recover~ of a silver
c~nc~ntr~e and using ~he~present proaess on th~ flotation
~all~ngs whiah con~ain mQst of the lead and all the remaining
silver. ~he silver ~lotatio~ conaentra~e and ~he plumbat~
prod~ct ~an then b~ combined ~or sale ~o aonventional l~a~ smeltexs.
Flotati~n proces~ei 8uch a~ desarib~ by M~rl~ama,E. and Yamamato,Y.
in AIME World Symposlum o~ Mining and M~allurgy o~ ~aad and Zinc,
,; 30 10
1 1S60~8
Vol. II, 1970, page 215 have been shown to yield silver concentrates
wlth hlgh silver assays and recoveries from lead sulphate
containing materials.
Another option in the present proces-s is the production of
separate silver and lead products. Silver can be removed from
the brine leach solution by cementation on a suitable metallic
medium such as zinc, iron, or lead. With proper stoichiometric
conditions, retention time, and pH nearly all the silver can be
recovered in a high grade metallic product containing some little
lead and copper as contaminants. Lead along with the solubilized
cementation agent would then be recovered in ~e plumbate product
as in steps (3) - (9).
When lead is precipitated from brine solution by the addition
of a base as in step (3) of the process, the lead compbunds formed
will depend on the pH or the mole ratio of base to lead chloride
and the total chloride concentration. Table l shows the effect of
these variables on the nature of the lead precipitate when lead
is precipitated from a brine solution containing 15 gpl lead as
lead chloride at 45C and a retention time of 1.5 hours. Shorter
retention times can be employed but chloride levels in the
precipitate will increase unless the temperature is increased above
45C. Silver is coprecipitated with lead. Most of the brine
soluble impurities which are present in the lead sulphate containing
starting material such as zinc, copper, iron, bismuth, and arsenic
also copr~cipitate with l~ad. T~e ba~t process eaonomics are
ob~ained with lime as the precipita~ion agent at an addition rate
he~ween l.Q and 5.5 mole ratio of lime to lead chloride. The excess
lime also acts as a 10cculating agent for oxychloxide precipita~e,
l 1560~
Table 1
Effect of the Mole Ratio of Base to Lead Chl.oride and the Total
Chloride Concentration on the Lead Precipitate
NaCl CaC12
Concen- Concen- Mole
Basetration tration Ratio Lead Precip.itate
(gpl)
CaO 269 15 0.75 PbOHCl
Na2C3 269 15 1.0 PbOHCl
NaOH 269 15 2.0 PbOHCl
CaO 269 15 1.5 3PbO-PbC12-nH2O ~
minor 2PbO-PbCl2~nH2O
CaO 269 33 1.5 3PbO~PbCl2~nH2O +
minor 2PbO~PbC12~nH2O
2 3 269 33 1.5 3PbO~PbC12 nH2O +
minor 2PbO-PbC12-nH2O
NaO~ 269 33 3.0 3PbO-PbC12-nH2O +
minor 2PbO-PbC12-nH2O
CaO 269 33 10.0 3PbO-PbC12-nH2O
NaOH 269 33 20.0 6PbO-PbCl2-nH2O
- 12
~"
0 ~ ~
resulting in iml)~oved solid~liquid seEJaration.
~ lthoLIgh it is desirable to produce a precipita-te containing
as little chloride as possible, very low chlorlde-oxychlorides
canrlot be precipitated from concentrated brine unless uneconomic
quantities of sodium hydroxide are used. Since they are very
soluble, all excess sodium hvdroxide and/or sodium carbonate
must be neutraliæed with hydrochloric acid before the lean
brine resulting from precipitate separation can be recycled to
the brine leach step (1). The use of sodium hydroxide precipitating
agent also results in excessive reagent calcium chloride makeup
requirements.
Ermilov, V.V. and Aitenov, S.A. in Trudy Institut Metallurgi
Obogashcheniia, Vol. 30, 1969, page 47 proposed a method for
producing a lead (iv) oxide precipitate from concentrated brine
by adding equal molar quantities of calcium oxide and calcium
hypochlorite to lead chloride. Although the reaction is irreversible and
regenerates calcium chloride into solution, and the product can easily
be washed to less than 0.5% chloride, the economics are unfavourable
due to the value of hypochlorite in comparison to lead metal and
lime.
If the lime addition in the precipitation step (3) was less than
2.5 mole ratio to lead, then after separation of the lead oxychloride
precipitate from the lean brine solution, the precipitate is
repulped with water or any reasonably unsaturated brine solution
produced in the proces~. Lime is added to the pulp to bring the
mole ratio ~f the total lime additlon in the process to bel;ween
.S and 5.S. A~tar solid/li~uid separation the pulp is subjected
to thermal tr~atmant. Alternatively the lead oxychloride precipitate
- 13 -
l 15~0~
may be blended with lime to increase the total mole ratio to
between 2.5 and 5.5, without repulpins and ~he blend subjected
to thermal treatment.
If the lime addition in step (3) was greater than 2.5 mole
ratio to lead, then repulping and/or further lime addition is
not necessary before thermal treatment.
The lead oxychloride-lime blend is heated in a reactor in
the presence of oxygen or air. The reactor can be a rotary kiln,
furnace, roaster, autoclave or any devi~e commonly used for thermal
treatment. The retention time in the reactor depends upon the
desired degree of conversion of oxychloride to plumbate and oxide,
the temperature in the reator, the lime to lead mole ratio, and the
oxygen partial pressure. The effects of these variables on the
nature of the calcine product are shown in Table 2. At reaction
temperatuxes above 500C sintering of the product appears and
volatilization of lead chloride begins. Preferred conditions
appear to be a totai process lime addition to lead mole ratio
of about 3, a reaction temperature of about 400C, a retention time
of about l hour, and an excess of oxygen for lead oxidation.
Pressure above atmosphericis not required for the reaction to be
complete within 2 hours, Sufficient air or oxygen can be supplied
for the reaction by convection, free or forced, or by pressurizing
the reactor.
'
~ 1 5 ~ 3
o o o o o o
ra 0
Y + ~ ~ Y Y
G N t`~
C~ U C~
ul Q R Q 5~ 0 0 Q R
1:4 ~ a~ 3 R ~ Pl O
~5 0000 ++ O 1~C~
.4 Q Q.a O O O Q Q
O ~ G Q~ 0 0P~
P~ V + ~ + + + ~ + ~ +
O 0 3 0 0 0 0 0 0 P~
u) ~ C~ Q Q Q Q R Q R Q Q
a~ o 0 ~ 0t~ 0~d ra ~4
UC~ VC~
a~
X
E~ h o
+
O ~ 00 0 0 ~ ~d 0 0
0
~
.~
C)
aJ
~ ~ - ~
a~ ,, O ~ ~ OO O O o o
Q O ~ ~ O o o ,1 ~ ~ ~ ~-
E~ S o
~ a)
g ~ .
rd
~ h u~
O 0 0 0 0 0 o o o o o o Ir~ R~
0 0 ooooooon~
O Cl.
~1 E~ ,1
0 E~
t)
~1
0
O
.r/ n~
a) 4
O
O ~P
~-1 0 O
~I h lo 0
Ul ~ OC:~ O~DO U'~ U
Ul ~ .. . ~ . .. . .
t) O
O ~
Pl
~ ~.5 -
11560~8
It issurprisfng that calcium orthoplumbate (Ca2PbO4) is
formed at high yields with such low temperatures and short
retention times. Calcium orthoplumbate in a pure form (>90%)
is a valuable commodity and is used in the manufacture of
primers for steel and galvanized steel, of pigments and of
binders for paints. It also has use in the plastics and resin
industries. The common commercial production method is the
reaction of PbO with lime and air or oxygen at temperatures above
700C. The reaction kinetics are reportedly slow below this
temperature and the reaction will not go below 500C. It has
baen reported by Denev, D.G et al in Dokl.Bolg.Akad.Nauk, Vol.26
11,1973, page 1485, however, that additions of small ~uantitles ;
of NaCl to the reaction mixture speed the kinetics. The present
invention differs from the customary practice in that the reagent
for calcination is lead oxychloride and not lead oxide. Also, ~ ;
the oxychloride i5 contaminated with significant quantities of
NaCl and CaC12. Accordingly, the kinetics and energetics of
plumbate ~ormation have been altered significantly from commercial
experience.
: 20 Figure 3 is a schematic plant layout for a particulsr embodiment
~- of the invention relating to example 1.
. :
The following examples illustrate the practice of our invention
:~ but should not be constxued as limiting.
S~mpl~s ~ ho~ sulp~uric acid leach ~es~due~ ~b~ained ~rom ~he
,
'~ s~lphatio~ roa~tln~ ~nd leachin~ o~ bulk zinc-lead-coppe~~5ilver
,
5ulphide ~o~cantrates assayi~ 30-32~ æn, 3.5-10~ Pb, ~.7~ Cu, 4.4 -
oz/~ ~t~o~ ounce per ~hort ton) ~ilv~x, and 14-23~ Lron were
processad accordingly to the
- 16 -
,
~ 15~0'1~
invention. A sample (example 10) of a hot acid leach residue
from a dead roast zinc plant was also processed. The residues
assayed as in Table 3.
The residues were leached in brines of composition
given in Table 4. Residue was added at a ratio of 15 gm of
contained lead per litre of hrine. The leaches were conducted
at 35-40C for 1.5 hours. Leach residues were allowed to
settle and the thickened residues filtered and washed with
~resh brine. Extractions of lead and silver are given in Table
5. All of the zinc, iron and copper as sulphates in the hot
sulphuric acid leach residues leached along with the lead
and silver. Copper and bismuth assays in the brine were 40 and
45 mgpl respectively.
One pregnant brine solution (example #6) was treated with zinc
dust at an addition rate of 0.5 gpl producing a cementate
containing 99% of the silver, 95% of the copper and 80~ of the
bismuth in the pregnant brine.
The remaining pregnant brines and the solution resulting ~rom
the zinc dust cementation test were then treated with lime at
1.5 mole ratio to lead in the brine solution. The temperature and
retention time were 45C and 1.5 hours respectively.
Precipitates were allowed to settle and the thickened
precipitat~s filtered.
~ he precipi~a~es in all tests wer~ ~hen hlend~d with lime,
~ xequir~d, to bring the total mole ratio of l.ime added to the
proce~s to lead in the precipitate to 1~75~5.5. The blends were
17
1 ~S60~
Table 3
Assays (wt%) of Hot Acid Leach Residues (Dry Basis)
Example ~
1 8 9 10
._ _
ZnFe2O4 1.5 1.5 5.0
(Zn,Fe)S 1.0 0.9 1.0
Fe23 52.8 63.2 10.0
PbSO~ 31.1 10.9 46.7
CaSO4 H2 6.8 6.2 6.0
SiO2 1.8 1.8 25.0
As 0.2 0.2 0.5
S 0.6 0.5 0.3
ZnSO4 1.1 4.0 0.5
CUSO4 0.1 0.3 0.1
Fe2(SO4)3 1.6 5.7 0.7
Ag (ppm) 639 310 250
Gangue 0.4 0.4 3.0
- ~ 18 ~
Table 4
Brine Compositions
Example ~ MgC12 NaCl CaC12 pH
(gpl) (gpl) (gpl)
1 - 300 - 1.5
2 - 250 - 1~5
3 - 269 15 1.5
4 ~ 269 33 1.5
- 269 33 ~.5
6 - 269 33 1.5
7 280 - 33 1.5
8 50 220 33 1.5
9 - 269 33 1.5
- 269 33 1.5
- 1
3~
. . . -- ,
~ 1 5~04~
Table 5
Extractions of Lead and Silver
Example# P~ extraction (%) Ag extraction (%)
1 93 65
2 75 50
3 92 62
4 99 77
96 47
6 98 75
7 95 70
8 95 70
:~ 9 97 71
~: 10 96 69
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1 1 5 ~
Table 6
Lime ~ddition and Plumbate Product ~Dry Basis) Assays ~wt.%)
Total
Example # Lime Ca Cl Fe Zn Ag (oz/ST)
to lead
1 3.0 60 22 0.6 1.4 1.5 34
2 1.75 67 12 4.8 1.8 1.9 38
3 ~.0 51 29 0.3 1.2 1.2 27
4 3.0 ~0 22 0.4 1.2 1.1 41
5.5 42 36 0.2 1.0 1.0 17
6 3.0 57 21 0.4 1.3 3.3 <1
7 3.0 59 22 0.4 1.3 1.3 37
8 3.0 60 22 0.5 1.2 1.3 38
9 3.0 54 20 0.4 3.0 3.0 50
3.0 61 22 0.5 0.7 0.7 40
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j~ '
",
0 ~ 8
treatecl in an oven with ;~ slow purge o~ fresh air for 1.0 hours
at 400C. The calcines were repulped to 50% pulp density for 15
minutes in rresh w~ter and filtered and displacement washed with
a volum~ of water equal to the calcine repulp water. The assays
(dry basis) of the resulting plumbate products are given in Table 6.
Example 11
A sample of the hot sulfuric acid leach residue used in
example 5 was repulped in water at 10~ pulp density, American Cyanamid
flot~tion reayents Aero 404 (trade mark~ promoter and Aerofroth 77A
(trade mark) (~rother) were added at 600 g. and 60 q. per metric ton of
residue respectively. After 5 minu~es conditioning time, flotation
was initiated and a stable froth was maintained for about 7.5 minutes.
The concentrate obtained assayed 18~ Pb, 24% Fe, 1.8% Zn, and
635 oz/ST silver. Silver recovery from the head material was 70%.
The tailings from the flotation were treated in the process
exactly similar to example 4. Silver recovery from the tailings
was 35%, result;ng in an overall silver recovery of 81~.
As will be apparent to those skilled in this art, the process
of this invention may be applied to the recovery of metals from
a variety of metallurgical products such as ores and concentrates,
~melter dusts, metal drosses, middling concentrates from ~lotation
~rocessing, slags and process residues, and other like sources oE
lead and silver,
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